The Hassai Mine Envisaged Business Plan (CIL Gold Plant and VMS Concentrator) Red Sea State, Sudan NI 43-101 Technical Report Prepared for La Mancha Resources Inc. and Ariab Mining Company Prepared by: Bill Plyley, MAusIMM – La Mancha Resources Inc. Jean-Jacques Kachrillo – La Mancha Resources Inc. Graeme Baker MAusIMM – AMEC Minproc Limited Dean David FAusIMM – AMEC Minproc Limited Adam Coulson ACSM, CIMM – AMEC Earth & Environmental Ian Thomas MAusIMM – Sedgman Limited Remi Bosc MEFG – Arethuse Geology Sdn Bhd Clayton Reeves, MSAIMM – CSA Global (UK) Simon McCracken MAIG – CSA Global (UK) Effective Date of Report: Effective Date of Mineral Resources: Effective Date of Mineral Reserves: 22 October 2010 31 August 2010 31 December 2009 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report IMPORTANT NOTICE This report was prepared as a National Instrument 43-101 Technical Report for La Mancha Resources Inc. (La Mancha) by AMEC Minproc Limited (AMEC), Arethuse Geology (Arethuse), Sedgman Limited (Sedgman) and CSA Global Pty Ltd (CSA Global). The quality of information, conclusions and estimates contained herein is consistent with the level of effort involved in the consultants’ services, based on: i) information available at the time of preparation, ii) data supplied by outside sources, and iii) the assumptions, conditions and qualifications set forth in this report. This report is intended for use by La Mancha subject to the terms and conditions of its contracts with AMEC and the other consultants. This contract permits La Mancha to file this report as a Technical Report with Canadian Securities Regulatory Authorities pursuant to National Instrument 43-101 Standards of Disclosure for Mineral Projects. Except for the purposes legislated under provincial securities law, any other uses of this report by any third party is at that party’s sole risk. FINAL – Rev 0 – 22 Oct 2010 Page i The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table of Contents Disclaimer....................................................................................................................... i 1. 1.1 1.2 1.3 1.11 1.12 1.13 1.14 SUMMARY ................................................................................................................. 1 BACKGROUND ................................................................................................................................1 BUSINESS PLAN .............................................................................................................................1 RESOURCES AND RESERVES ......................................................................................................1 1.3.1 Gold Resources and Reserves ..........................................................................................1 1.3.2 VMS Resources.................................................................................................................3 USE OF INFERRED RESOURCES IN BUSINESS PLAN ................................................................4 MINING ............................................................................................................................................5 1.5.1 Kamoeb Open Pit ..............................................................................................................5 1.5.2 Heap Leach Tailings and Stockpile Reclaim ......................................................................7 1.5.3 Hadal Awatib Open Pit.......................................................................................................7 1.5.4 Hassai South Underground Mine Design ...........................................................................8 1.5.5 Production Schedules ........................................................................................................9 1.5.6 Mine Capital Cost ............................................................................................................14 1.5.7 Mine Operating Cost ........................................................................................................ 15 PROCESSING ................................................................................................................................ 15 1.6.1 CIL Plant .......................................................................................................................... 15 1.6.2 VMS Concentrator ........................................................................................................... 17 INFRASTRUCTURE ....................................................................................................................... 20 1.7.1 Power .............................................................................................................................. 20 1.7.2 Water ............................................................................................................................... 21 1.7.3 Accommodation ............................................................................................................... 21 1.7.4 Access and Port ..............................................................................................................21 ENVIRONMENTAL ......................................................................................................................... 21 CAPITAL COSTS ........................................................................................................................... 21 1.9.1 General............................................................................................................................ 21 1.9.2 CIL Plant Development .................................................................................................... 22 1.9.3 VMS Concentrator Development ..................................................................................... 22 OPERATING COSTS ..................................................................................................................... 22 1.10.1 General............................................................................................................................ 22 1.10.2 CIL Plant .......................................................................................................................... 23 1.10.3 VMS Concentrator ........................................................................................................... 23 PROJECT SCHEDULE .................................................................................................................. 23 FINANCIAL MODELLING ............................................................................................................... 24 CONCLUSIONS ............................................................................................................................. 27 RECOMMENDATIONS .................................................................................................................. 28 2. 2.1 2.2 2.3 2.4 2.5 INTRODUCTION ...................................................................................................... 30 BACKGROUND .............................................................................................................................. 30 SCOPES OF WORK....................................................................................................................... 30 PRINCIPAL SOURCES OF INFORMATION .................................................................................. 31 PARTICIPANTS AND PERSONAL SITE INSPECTIONS ............................................................... 31 INDEPENDENCE ........................................................................................................................... 33 1.4 1.5 1.6 1.7 1.8 1.9 1.10 FINAL – Rev 0 – 22 Oct 2010 Page ii The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 3. RELIANCE ON OTHER EXPERTS ........................................................................... 34 4. 4.1 4.2 4.3 4.4 4.5 4.6 4.7 PROPERTY DESCRIPTION AND LOCATION ........................................................... 35 LOCATION ..................................................................................................................................... 35 MINING CLAIM DESCRIPTION – GOLD ....................................................................................... 37 MINING CLAIMS – BASE METALS ................................................................................................ 42 OWNERSHIP OF MINERAL RIGHTS ............................................................................................ 42 MINERAL ROYALTIES................................................................................................................... 42 ENVIRONMENTAL OBLIGATIONS ................................................................................................ 42 RELATIONSHIP BETWEEN AMC AND THE SUDANESE GOVERNMENT .................................. 42 5. 5.5 5.6 5.7 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY ..................................................................................................... 43 ACCESS ......................................................................................................................................... 43 PORT FACILITIES.......................................................................................................................... 43 CLIMATE ........................................................................................................................................ 44 INFRASTRUCTURE ....................................................................................................................... 44 5.4.1 Buildings and Mine Camp ................................................................................................ 44 5.4.2 Other Offices ................................................................................................................... 44 5.4.3 Logistics .......................................................................................................................... 44 LAND USAGE ................................................................................................................................ 45 PHYSIOGRAPHY AND VEGETATION........................................................................................... 45 SURFACE AND GROUNDWATER ................................................................................................ 45 6. 6.1 6.2 HISTORY ................................................................................................................. 46 PRE-COMINOR .............................................................................................................................. 46 COMINOR ...................................................................................................................................... 46 7. 7.1 7.2 GEOLOGICAL SETTING .......................................................................................... 48 REGIONAL GEOLOGY .................................................................................................................. 48 LOCAL GEOLOGY ......................................................................................................................... 48 8. 8.1 8.2 DEPOSIT TYPES ..................................................................................................... 49 INTRODUCTION ............................................................................................................................ 49 GOLD DEPOSITS .......................................................................................................................... 49 8.2.1 Oxide and Quartz-Kaolinite-Barite (“SBR”) Gold Deposits ............................................... 49 8.2.2 Gold-bearing Quartz Veins .............................................................................................. 51 8.2.3 Gold-rich Barite Lenses Without Proximal Gossan Development .................................... 52 VOLCANOGENIC CU-ZN-AU-AG MASSIVE SULPHIDE DEPOSITS ............................................ 52 5.1 5.2 5.3 5.4 8.3 9. 9.1 9.2 MINERALISATION .................................................................................................. 53 BASE METAL MASSIVE SULPHIDE DEPOSITS ........................................................................... 53 GOLD DEPOSITS .......................................................................................................................... 53 9.2.1 Supergene (SBR) Deposits Overlying VMS Mineralisation .............................................. 53 9.2.2 Quartz Veins .................................................................................................................... 53 10. 10.1 EXPLORATION ........................................................................................................ 54 EXPLORATION METHODS ...........................................................................................................54 FINAL – Rev 0 – 22 Oct 2010 AMEC Page iii The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 10.2 10.3 SURVEYING .................................................................................................................................. 55 MAIN RESULTS ............................................................................................................................. 55 10.3.1 VMS – Prior to 2007 ........................................................................................................ 55 10.3.2 2007 VTEM Geophysical Survey ..................................................................................... 56 11. 11.1 11.2 11.3 11.4 DRILLING ................................................................................................................ 59 INTRODUCTION ............................................................................................................................ 59 DRILLING: 1993-2006 .................................................................................................................... 59 RC AND CORE DRILLING: 2008/09 .............................................................................................. 59 11.3.1 Hassai South Drilling ....................................................................................................... 60 11.3.2 Hadal Awatib East Drilling ............................................................................................... 63 11.3.3 Kamoeb Drilling ...............................................................................................................64 HEAP LEACH RESIDUE DRILLING ............................................................................................... 65 12. 12.1 12.2 12.3 12.4 SAMPLING METHOD AND APPROACH................................................................... 66 DRILLING, SAMPLING AND SAMPLE PREPARATION ................................................................ 66 RC AND CORE RECOVERY – HASSAI SOUTH AND HADAL AWATIB EAST ............................. 66 RC AND CORE RECOVERY – KAMOEB ...................................................................................... 66 AUGER RECOVERY – TAILINGS .................................................................................................. 66 13. 13.1 SAMPLE PREPARATION, ANALYSES AND SECURITY .......................................... 67 INTRODUCTION ............................................................................................................................ 67 13.1.1 Gold ................................................................................................................................. 67 13.1.2 Base Metals ..................................................................................................................... 67 13.1.3 Heap Leach Tailings Gold ............................................................................................... 67 SAMPLING, SAMPLE PREPARATION AND STORAGE ............................................................... 67 13.2.1 Gold Exploration: 1992 to 2007 ....................................................................................... 67 13.2.2 Gold Exploration: 2008/09 ............................................................................................... 68 13.2.3 Base Metal Sulphide Exploration ..................................................................................... 69 13.2.4 Heap Auger Drill Samples ............................................................................................... 70 DRY BULK DENSITY ..................................................................................................................... 71 13.3.1 Core Samples ..................................................................................................................71 13.3.2 Auger Samples ................................................................................................................72 13.2 13.3 14. 14.1 14.2 14.3 14.4 14.5 14.6 14.7 DATA VERIFICATION ............................................................................................. 73 DATA COLLECTION ...................................................................................................................... 73 ASSAY DATA QUALITY ................................................................................................................. 73 14.2.1 Blanks..............................................................................................................................75 14.2.2 Standards ........................................................................................................................ 75 14.2.3 Duplicates ........................................................................................................................ 79 14.2.4 Conclusions ..................................................................................................................... 79 KAMOEB TWIN HOLES ................................................................................................................. 80 GEOLOGICAL DATA...................................................................................................................... 80 SURVEY DATA .............................................................................................................................. 80 DENSITY DATA ............................................................................................................................. 80 DATABASE VERIFICATION........................................................................................................... 80 14.7.1 Database Consistency – Internal Review ........................................................................ 80 FINAL – Rev 0 – 22 Oct 2010 AMEC Page iv The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 14.8 14.7.2 External Independent Data Validation ............................................................................. 81 14.7.3 Independent Sampling and Analysis (Gold Only)............................................................. 83 AUGER PROGRAM DATA VERIFICATION ................................................................................... 83 14.8.1 Drill Logs and Sampling Information ................................................................................ 83 14.8.2 Assay Data Quality .......................................................................................................... 84 14.8.3 Auger Samples – Conclusions ......................................................................................... 84 15. ADJACENT PROPERTIES ....................................................................................... 85 16. 16.1 MINERAL PROCESSING AND METALLURGICAL TESTING ................................... 86 HEAP LEACH TESTWORK ............................................................................................................ 86 16.1.1 Heap Leach Testwork ...................................................................................................... 86 CIL TESTWORK ............................................................................................................................. 87 16.2.1 Quartz Ore (Kamoeb South Deposit) ............................................................................... 87 16.2.2 Heap Leach Residue ....................................................................................................... 91 16.2.3 Metallurgical Gold Recovery for the CIL Economic Model (Includes Operating Cost Adjustments for Acidic SBR Material) .............................................................................. 97 VMS CONCENTRATOR TESTWORK............................................................................................ 99 16.3.1 Introduction ...................................................................................................................... 99 16.3.2 Sample Selection............................................................................................................. 99 16.3.3 Flotation Testwork ......................................................................................................... 101 16.3.4 Tailings Cyanide Leaching Testwork ............................................................................. 107 16.3.5 Metallurgical Projection and Metallurgical Parameters for Design ................................. 107 CONCENTRATOR FLOW SHEET DEVELOPMENT.................................................................... 109 16.4.1 Process Design Criteria and Mass Balance ................................................................... 111 16.4.2 Comminution Circuit ...................................................................................................... 111 16.4.3 Flotation Circuit.............................................................................................................. 111 16.4.4 Regrind .......................................................................................................................... 113 16.4.5 Concentrate Handling .................................................................................................... 114 16.2 16.3 16.4 17. 17.1 17.2 17.3 MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES ............................. 115 GENERAL .................................................................................................................................... 115 VMS RESOURCES: HADAL AWATIB EAST AND HASSAI SOUTH ............................................ 115 17.2.1 Geological Model ........................................................................................................... 115 17.2.2 Cut-off and Domain Modelling ....................................................................................... 117 17.2.3 Down-dip Drill Holes within the Supergene Domain....................................................... 118 17.2.4 Overall Population Distribution and Top-cuts ................................................................. 118 17.2.5 Dry Bulk Density ............................................................................................................ 119 17.2.6 Correlations Between Elements..................................................................................... 120 17.2.7 Variography and Interpolation Parameters .................................................................... 120 17.2.8 Block Model ...................................................................................................................121 17.2.9 Confidence Classification and Mineral Resource Reporting under NI 43-101 ................ 126 KAMOEB RESOURCES ............................................................................................................... 127 17.3.1 Geological Model ........................................................................................................... 127 17.3.2 Cut-Off and Domain Modelling ...................................................................................... 129 17.3.3 Population Distribution and Top-cuts ............................................................................. 129 17.3.4 Dry Bulk Density ............................................................................................................ 130 FINAL – Rev 0 – 22 Oct 2010 AMEC Page v The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 17.4 17.5 18. 18.1 18.2 18.3 18.4 18.5 17.3.5 Variography and Interpolation Parameters .................................................................... 130 17.3.6 Block Model ...................................................................................................................131 17.3.7 Confidence Classification and Mineral Resource Reporting Under NI 43-101 ............... 134 TAILINGS RESOURCES .............................................................................................................. 136 17.4.1 Resources Estimated by Arethuse Using Conventional Resource Modelling Techniques – Heap 63A to 113 ..................................................................................... 136 17.4.2 Additional External Areas: Heaps 1-63, 67-71 .............................................................. 143 17.4.3 Material Stacked in 2008 – 2009 (Heaps 114 to 136) .................................................... 143 17.4.4 Additional Material to be Stacked, 2010-2012 ............................................................... 146 MINERAL RESOURCE STATEMENT .......................................................................................... 147 17.5.1 Overall AMC Resources – 31 December 2009 .............................................................. 147 17.5.2 Additional Heap Leach Tailings Resources ................................................................... 149 17.5.3 Mineral Reserve Statement ........................................................................................... 150 OTHER RELEVANT DATA AND INFORMATION ................................................... 151 MINING STUDIES – GENERAL STATEMENT REGARDING USE OF INFERRED RESOURCES ............................................................................................................................... 151 CSA MINING STUDIES – KAMOEB ............................................................................................. 151 18.2.1 Mining Study Background .............................................................................................. 151 18.2.2 Study Approach ............................................................................................................. 151 18.2.3 Mining Methods ............................................................................................................. 152 18.2.4 Pit Optimisation .............................................................................................................152 18.2.5 Mine Design ..................................................................................................................161 18.2.6 Waste Handling ............................................................................................................. 163 18.2.7 Mining Inventories ......................................................................................................... 163 18.2.8 Ore Production Schedules ............................................................................................. 164 18.2.9 Operating Costs............................................................................................................. 167 18.2.10 Mine Capital Costs ........................................................................................................ 169 CSA MINING STUDY – ACIDIC SBR ORE STOCKPILES AND HEAP LEACH TAILINGS .......... 170 18.3.1 Introduction .................................................................................................................... 170 18.3.2 Heap Leach Tailings and Acidic SBR Stockpile Reclamation Schedule ........................ 173 18.3.3 Heap Leach Tailings and Acidic SBR Stockpile Reclamation Operating Costs.............. 175 18.3.4 Heap Leach Tailings and Acidic SBR Stockpile Reclamation Mine Capital Costs ......... 177 AMEC MINING STUDIES – VMS DEPOSITS .............................................................................. 177 18.4.1 Mining Study Background .............................................................................................. 177 18.4.2 Study Approach ............................................................................................................. 177 18.4.3 Mining Methods ............................................................................................................. 177 18.4.4 Pit Optimisation .............................................................................................................179 18.4.5 Mine Design ..................................................................................................................185 18.4.6 Waste Handling ............................................................................................................. 190 18.4.7 Mining Inventories ......................................................................................................... 192 18.4.8 Ore Production Schedules ............................................................................................. 194 18.4.9 Mine Operating Costs .................................................................................................... 199 18.4.10 Mine Capital Costs ........................................................................................................ 202 GEOTECHNICAL INPUT .............................................................................................................. 204 18.5.1 Kamoeb South – AMC ................................................................................................... 204 18.5.2 AMEC Geotechnical Input – Introduction ....................................................................... 204 FINAL – Rev 0 – 22 Oct 2010 AMEC Page vi The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.6 18.7 18.8 18.9 18.10 18.11 18.12 18.13 18.14 18.15 18.16 19. 18.5.3 Geotechnical/Geological Domains ................................................................................. 204 18.5.4 Major Joint Set Orientation ............................................................................................ 205 18.5.5 Material Testing .............................................................................................................207 18.5.6 Rock Mass Classification ............................................................................................... 207 18.5.7 Design Criteria for Hadal Awatib and Hassai South Open Pits ...................................... 210 18.5.8 Design Criteria for Hassai South Underground Open Stoping ....................................... 212 HYDROGEOLOGY AND HYDROLOGY INPUT ........................................................................... 216 SEISMICITY ................................................................................................................................. 217 PROCESS PLANT DESCRIPTIONS ............................................................................................ 217 18.8.1 CIL Plant ........................................................................................................................ 217 18.8.2 VMS Concentrator Process Plant Description ............................................................... 221 PROJECT INFRASTRUCTURE AND SERVICES ........................................................................ 228 18.9.1 Water Supply .................................................................................................................228 18.9.2 Power Supply ................................................................................................................ 229 18.9.3 Accommodation ............................................................................................................. 230 18.9.4 Airstrip ........................................................................................................................... 230 18.9.5 VMS Concentrator Tailings Storage Facility .................................................................. 230 18.9.6 Other Infrastructure ....................................................................................................... 237 MARKETS .................................................................................................................................... 237 ENVIRONMENTAL AND SOCIAL CONSIDERATIONS ............................................................... 238 TAXES AND ROYALTIES ............................................................................................................ 239 CAPITAL COST ESTIMATE .........................................................................................................239 18.13.1 General.......................................................................................................................... 239 18.13.2 Mining ............................................................................................................................ 240 18.13.3 CIL Plant ........................................................................................................................ 240 18.13.4 VMS Concentrator ......................................................................................................... 241 OPERATING COST ESTIMATE ................................................................................................... 248 18.14.1 CIL Plant ........................................................................................................................ 248 18.14.2 VMS Concentrator ......................................................................................................... 252 18.14.3 Mining Costs ..................................................................................................................256 18.14.4 General and Administration ........................................................................................... 256 PROJECT ECONOMICS .............................................................................................................. 257 18.15.1 Overview ....................................................................................................................... 257 18.15.2 Individual Phase Description ......................................................................................... 263 PROJECT IMPLEMENTATION .................................................................................................... 272 18.16.1 Project Schedule ........................................................................................................... 272 18.16.2 Project Implementation Summary .................................................................................. 272 18.16.3 Project Implementation Scope ....................................................................................... 273 18.16.4 HSEC ............................................................................................................................ 273 18.16.5 Long Lead Items ............................................................................................................ 273 18.16.6 Logistics ........................................................................................................................ 274 18.16.7 Training ......................................................................................................................... 274 INTERPRETATION AND CONCLUSIONS .............................................................. 275 FINAL – Rev 0 – 22 Oct 2010 AMEC Page vii The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 20. 20.6 20.7 ADDITIONAL REQUIREMENTS FOR TECHNICAL REPORTS ON DEVELOPMENT PROPERTIES AND PRODUCTION PROPERTIES ....................... 277 BACKGROUND ............................................................................................................................ 277 MINING ........................................................................................................................................ 279 20.2.1 Overview ....................................................................................................................... 279 20.2.2 Geotechnical Evaluation ................................................................................................ 281 20.2.3 Grade Control and Mining.............................................................................................. 281 BENEFICIATION PLANT .............................................................................................................. 284 20.3.1 Overview ....................................................................................................................... 284 20.3.2 Process Flow Sheets ..................................................................................................... 285 20.3.3 Reagents ....................................................................................................................... 288 20.3.4 Gold Reconciliation at the Plant ..................................................................................... 288 20.3.5 Mine/Plant Gold Reconciliation ...................................................................................... 288 TAILINGS AND WASTE MANAGEMENT ..................................................................................... 289 INFRASTRUCTURE ..................................................................................................................... 289 20.5.1 Buildings and Mine Camp .............................................................................................. 289 20.5.2 Other Offices .................................................................................................................289 20.5.3 Logistics ........................................................................................................................ 289 20.5.4 Water Supply .................................................................................................................290 SOCIAL PROGRAM ..................................................................................................................... 291 FINANCIAL ANALYSIS ................................................................................................................ 291 21. 21.1 21.2 RECOMMENDATIONS ........................................................................................... 293 CIL PHASE ................................................................................................................................... 293 VMS PHASE................................................................................................................................. 293 22. 22.1 22.2 REFERENCES........................................................................................................ 294 GEOLOGY AND RESOURCES.................................................................................................... 294 GEOTECHNICAL ......................................................................................................................... 294 23. DATE AND SIGNATURE PAGE ............................................................................. 295 24. ILLUSTRATIONS ................................................................................................... 305 25. APPENDIX 1 RECENT HASSAI SOUTH, HADAL AWATIB AND KAMOEB DRILL INTERSECTIONS ....................................................................................... 306 20.1 20.2 20.3 20.4 20.5 FINAL – Rev 0 – 22 Oct 2010 AMEC Page viii The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report List of Tables Table 1.1 Table 1.2 Table 1.3 Table 1.4 Table 1.5 Table 1.6 Table 1.7 Table 1.8 Table 1.9 Table 1.10 Table 1.11 Table 1.12 Table 1.13 Table 1.14 Table 4.1 Table 4.2 Table 5.1 Table 14.1 Table 14.2 Table 14.3 Table 14.4 Table 14.5 Table 16.1 Table 16.2 Table 16.3 Table 16.4 Table 16.5 Table 16.6 Table 16.7 Table 16.8 Table 16.9 Table 16.10 Table 16.11 Table 16.12 Table 16.13 Table 16.14 Table 16.15 Table 16.16 Table 16.17 Table 16.18 Table 16.19 Table 16.20 Table 16.21 Table 16.22 Table 16.23 Gold Mineral Resources Summary (31 December 2009) ................................................................. 2 Mineral Reserves Summary (31 December 2009) ............................................................................ 2 Additional Gold Mineral Resources in Heap Leach Tailings Under Irrigation.................................... 3 VMS Mineral Resources, 31 December 2009 ................................................................................... 3 Pit Design Parameters ........................................................................................................................ 5 Hadal Awatib Open Pit Design Parameters ....................................................................................... 7 Kamoeb – Yearly Mining Schedule ..................................................................................................10 Mining Schedule for Heap Leach and Stockpile Reclaim ................................................................11 Mining Inventory for CIL Operation...................................................................................................12 Mine Capital Cost Estimate ..............................................................................................................14 Mine Operating Cost Estimate..........................................................................................................15 Average Concentrate Production .....................................................................................................20 Capital Cost Estimate, 5 Mt/a VMS Concentrator Phase ................................................................22 Financial Highlights for Proposed VMS Project, by Phase ..............................................................26 Coordinates of AMC’s Reserved Areas ...........................................................................................38 Coordinates of Mining Leases ..........................................................................................................40 Port Sudan Overview ........................................................................................................................43 Characteristics of Standard Reference Materials.............................................................................76 CRM Assay Results – Oxide Gold ...................................................................................................77 CRM Assay Results – VMS Gold .....................................................................................................78 CRM Assay Results – VMS Copper ................................................................................................78 CRM Assay Results – VMS Zinc......................................................................................................79 Quartz Ore Comminution Parameters..............................................................................................88 Quartz Composite Head Assay ........................................................................................................89 Gravity Separation Results Summary ..............................................................................................90 Quartz Ore Grind vs Leach Recovery Results .................................................................................90 Air vs Oxygen Sparging Leach Summary – Quartz Ore ..................................................................91 Quartz Ore Leach Cyanide Sensitivity .............................................................................................91 Heap Leach Bulk Composite Assay .................................................................................................92 Heap Leach Bulk Composite Gravity Separation Results Summary ..............................................93 Heap Leach Grind vs Leach Recovery Results ...............................................................................94 Lead Nitrate Addition Results ...........................................................................................................95 Pre-Aeration Testing Results............................................................................................................96 Heap Leach Variability Testing Results............................................................................................96 Breakdown of Heap Leach Mining Inventory Resources.................................................................97 Reagent Consumption and Costs for Heap Leach and Acidic SBR Material..................................99 Head Sample Chemical Analysis Results ......................................................................................100 Head Sample Mineral Distribution (% Mass) .................................................................................100 Rougher Flotation Kinetic Results ..................................................................................................102 Cleaner Flotation Results for Composite 1.....................................................................................104 Cleaner Flotation Results for Composite 3.....................................................................................105 Locked Cycle Testwork Results .....................................................................................................106 Tailings Cyanide Leach Testwork ..................................................................................................107 Testwork and Design Ore Grades..................................................................................................108 Design Concentrate and Tailing Grades ........................................................................................108 FINAL – Rev 0 – 22 Oct 2010 AMEC Page ix The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 16.24 Table 16.25 Table 16.26 Table 16.27 Table 17.1 Table 17.2 Table 17.3 Table 17.4 Table 17.5 Table 17.6 Table 17.7 Table 17.8 Table 17.9 Table 17.10 Table 17.11 Table 17.12 Table 17.13 Table 17.14 Table 17.15 Table 17.16 Table 17.17 Table 17.18 Table 17.19 Table 17.20 Table 17.21 Table 17.22 Table 17.23 Table 17.24 Table 18.1 Table 18.2 Table 18.3 Table 18.4 Table 18.5 Table 18.6 Table 18.7 Table 18.8 Table 18.9 Table 18.10 Table 18.11 Table 18.12 Table 18.13 Table 18.14 Table 18.15 Table 18.16 Design Calculated Stage Recoveries.............................................................................................108 Slurry Flows (m3/h) and Differences for the Different Ore Types...................................................111 Flotation Design Basis ....................................................................................................................112 Design Basis Concentrate Handling...............................................................................................114 Hadal Awatib East (HAE) and Hassai South (HASS) – Sample Statistics ....................................118 Hassai South – Density on Cores within Sulphide Mineralisation .................................................119 Hassai South – Primary Ore – Correlation Matrix ..........................................................................120 Block Model Definition ....................................................................................................................121 Kriging Search Ellipsoids ................................................................................................................122 Hadal Awatib East (HAE) and Hassai South (HASS) Resources Estimates as of 31 December 2009 .........................................................................................................................127 Proposed Gold Top-cuts for Different Domains (Au FA and Au Cy) .............................................130 Kamoeb Block Model Definition .....................................................................................................131 Kamoeb Interpolation Parameters..................................................................................................132 Kamoeb Group – NI 43-101 Gold Mineral Resources – 1 January 2010 .....................................135 Tailings Resource Basic Statistics – Au Fire Assay (g/t) ...............................................................138 Tailings – Block Model Details........................................................................................................139 Tailings – Grade Interpolation Parameters.....................................................................................140 Hassai Tailings Resources (Drilled, as of 31 December 2009) .....................................................143 Hassai Tailings from Active Heap Material, heaps 114-136, CSA September 2010 (Cyanidable Au) ..............................................................................................................................146 Material Currently Under Irrigation (Heaps 137-141, Cyanidable Au) ...........................................146 Oxide Mineral Resources, 31 December 2009 ..............................................................................147 Quartz Ore Mineral Resources, 31 December 2009 .....................................................................148 Tailings Mineral Resources, 31 December 2009 ...........................................................................148 Stockpile Mineral Resources, 31 December 2009.........................................................................148 Hassai Mine Combined Gold Mineral Resources, 31 December 2009.........................................149 VMS Mineralisation Mineral Resources, 31 December 2009 ........................................................149 Hassai Tailings from Active Heap Material, Heaps 114-136, CSA September 2010 (cyanidable Au) ...............................................................................................................................150 Hassai Mine Mineral Reserves, 31 December 2009 .....................................................................150 Kamoeb South Whittle Input Parameters .......................................................................................154 Kamoeb North Whittle Input Parameters .......................................................................................155 Kamoeb South and Kamoeb North Resource Model Extents .......................................................156 Mining Costs Applied in the Whittle Optimisations .........................................................................157 Processing Costs Applied in the Whittle Optimisations .................................................................157 Pit Design Parameters ....................................................................................................................161 Kamoeb South and Kamoeb North Waste Dump Quantities ........................................................163 Mining Inventory – Kamoeb South Open Pit ..................................................................................163 Mining Inventory – Kamoeb North Open Pit...................................................................................164 Kamoeb – Yearly Mining Schedule ................................................................................................165 Operating Costs Schedule – Kamoeb Open Pits...........................................................................168 Replacement Capital Cost Summary – Kamoeb Open Pits ..........................................................169 Heap Leach Tailings Inventory .......................................................................................................171 Acidic SBR Stockpile Inventory ......................................................................................................172 Hassai Acidic SBR Stockpile and Heap Leach Tailings Reclamation Schedule...........................174 Operating Costs Schedule – Heap Leach Tailings and Acidic SBR Stockpile Reclamation ........176 FINAL – Rev 0 – 22 Oct 2010 AMEC Page x The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.17 Table 18.18 Table 18.19 Table 18.20 Table 18.21 Table 18.22 Table 18.23 Table 18.24 Table 18.25 Table 18.26 Table 18.27 Table 18.28 Table 18.29 Table 18.30 Table 18.31 Table 18.32 Table 18.33 Table 18.34 Table 18.35 Table 18.36 Table 18.37 Table 18.38 Table 18.39 Table 18.40 Table 18.41 Table 18.42 Table 18.43 Table 18.44 Table 18.45 Table 18.46 Table 18.47 Table 18.48 Table 18.49 Table 18.50 Table 18.51 Table 18.52 Table 18.53 Table 18.54 Table 18.55 Table 18.56 Table 18.57 Table 18.58 Table 18.59 Table 18.60 Table 18.61 Table 18.62 Table 18.63 Whittle Input Parameters ................................................................................................................180 Hassai South – Underground Resource Model Update ................................................................181 Hassai South – Resource Model Consolidated Grades ................................................................181 Hassai South – Underground Resource Model Update ................................................................185 Hassai South – Underground Resource Model Consolidated Grades..........................................186 Hassai South – Underground Resource Model (Cut-off 1.5% Cueqm)..........................................186 Hadal Awatib Waste Dump Quantities ...........................................................................................191 Mining Inventory – Hassai South Underground .............................................................................193 Undiluted Mining Inventory Sensitivity – Hassai South Underground ...........................................193 Mining Inventory – Hadal Awatib 5 Mt/a Open Pit .........................................................................194 Hassai South – Underground Mining Schedule .............................................................................195 Hadal Awatib – 5 Mt/a Mining Schedule ........................................................................................197 Unit Operating Costs – Hassai South Underground ......................................................................200 Unit Operating Costs – Hadal Awatib Open Pit..............................................................................201 Capital Cost Summary – Hassai South Underground ...................................................................202 Capital Cost Summary – Hadal Awatib Open Pit Options .............................................................203 Summary of Probable Major Joint Set Orientations .......................................................................205 Summary of Previous Laboratory Testing ......................................................................................207 Hadal Awatib – Summary of Rock Mass Properties by Domain and Stope Zone ........................207 Hassai South – Summary of rock Mass Properties by Domain and Stope Zone .........................208 Summary of Bench Face Rock Mass Classification ......................................................................209 Summary of Assumed in situ Stress Regime.................................................................................209 Hadal Awatib – Summary of Existing and Proposed Slope Design Crieria...................................211 Hassai South – Summary of Existing and Proposed Slope Design Criteria .................................212 Summary of Simplified Design Rock Mass Properties ..................................................................213 Summary of Stope Surface Stability Criteria ..................................................................................214 Groundwater Chemical Analysis, Hadal Awatib and Hassai South...............................................217 Summary of Major Design Criteria .................................................................................................218 Crushing Equipment .......................................................................................................................222 Cleaner Flotation Equipment ..........................................................................................................224 Concentrate Handling Equipment ..................................................................................................224 Concentrate Production ..................................................................................................................225 Design Raw Water Consumption (no return from TSF).................................................................227 Water Pipeline Summary ................................................................................................................229 Plant Power Requirements (MW)...................................................................................................229 Environmental Assessment Scoring Criteria ..................................................................................231 Disposal System Ranking...............................................................................................................234 Mine Capital Cost Estimate Summary – Kamoeb, Hassai South and Hadal Awatib ....................240 Estimated Capital Costs, 3.0 Mt/a CIL Plant ..................................................................................241 Concentrator Capital Cost Estimate Summary by Area.................................................................246 Concentrator Phase Sustaining Capital Estimate Allowance ........................................................248 Total Annual Operating Cost Estimate, 3 Mt/a CIL Plant ...............................................................249 Annual Processing Plant Labour Costs, 3 Mt/a CIL Plant .............................................................250 LOM Plant Consumable Costs, 3 Mt/a CIL Plant...........................................................................251 Annual Maintenance Consumable Costs, 3 Mt/a CIL Plant...........................................................251 Estimated Annual Power Costs, 3 Mt/a CIL Plant..........................................................................252 Average Process Operating Costs, 5 Mt/a VMS Concentrator .....................................................253 FINAL – Rev 0 – 22 Oct 2010 AMEC Page xi The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.64 Table 18.65 Table 18.66 Table 18.67 Table 18.68 Table 18.69 Table 18.70 Table 18.71 Table 18.72 Table 18.73 Table 18.74 Table 18.75 Table 18.76 Table 18.77 Table 18.78 Table 18.79 Table 20.1 Table 20.2 Table 20.3 Table 20.4 Table 20.5 Summary of Operations Labour Structure, 5 Mt/a VMS Concentrator ..........................................254 Major Reagent Costs, 5 Mt/a VMS Concentrator...........................................................................254 Power Costs, 5 Mt/a VMS Concentrator ........................................................................................255 Consumables Cost Summary, 5 Mt/a VMS Concentrator .............................................................255 Maintenance Materials, 5 Mt/a VMS Concentrator ........................................................................256 Summary of LOM Mine Operating Cost Estimates ........................................................................256 AMC 2009 G&A Costs in Euros as Basis for Scoping Study.........................................................257 Financial Highlights for Proposed VMS Project, by Phase ............................................................260 Sensitivity Analysis Matrix– Gold Price ..........................................................................................261 Sensitivity Analysis Matrix – Operating Costs ................................................................................262 Production and Cashflow Projections – Heap Leach Project ........................................................264 Production and Cashflow Projections – CIL Phase .......................................................................266 Production and Cashflow Projections – VMS Phase .....................................................................268 Sensitivity of VMS Phase Economics to Metal Price Changes .....................................................269 Production and Cashflow Outcomes – CIL+VMS Phases ............................................................271 Current Equipment Lead Times .....................................................................................................274 AMC Production History .................................................................................................................279 AMC Mobile Equipment Fleet as at 31/12/2009 ............................................................................281 Geology Mine Reconciliation for Depleted Deposits ......................................................................283 Reagents Used at the Hassai Heap Leach Plant...........................................................................288 Cashflow Analysis of Current Operation ........................................................................................291 FINAL – Rev 0 – 22 Oct 2010 AMEC Page xii The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report List of Figures Figure 1.1 Figure 1.2 Figure 1.3 Figure 1.4 Figure 1.5 Figure 1.6 Figure 1.7 Figure 1.8 Figure 1.9 Figure 1.10 Figure 1.11 Figure 4.1 Figure 4.2 Figure 4.3 Figure 6.1 Figure 8.1 Figure 8.2 Figure 10.1 Figure 10.2 Figure 10.3 Figure 11.1 Figure 11.2 Figure 11.3 Figure 11.4 Figure 11.5 Figure 14.1 Figure 16.1 Figure 16.2 Figure 16.3 Figure 17.1 Figure 17.2 Figure 17.3 Figure 17.4 Figure 17.5 Figure 17.6 Figure 17.7 Figure 17.8 Figure 17.9 Figure 17.10 Kamoeb South – Pit Design – Plan View ........................................................................................... 6 Kamoeb North – Pit Design – Plan View............................................................................................ 6 Hadal Awatib – Pit Design 5 Mt/a ....................................................................................................... 8 Hassai South – Development Long Section ...................................................................................... 9 Kamoeb – Yearly Mining Profile .......................................................................................................10 Hassai Acidic SBR Stockpile and Heap Leach Tailings Yearly Reclamation Schedule Graph ................................................................................................................................................12 Hassai South – Underground Stoping Schedule (Ore) ....................................................................13 Hadal Awatib – 5 Mt/a Ore Mining Profile ........................................................................................13 Hadal Awatib – 5 Mt/a Mining Profile................................................................................................14 Hassai Mine Envisaged Business Plan – Summary Project Schedule ...........................................24 Metal Production Profile for Phased VMS Project ...........................................................................25 Location of the Hassai Project ..........................................................................................................36 Prospecting Licences........................................................................................................................37 Sudan Gold and Iron Concessions Map ..........................................................................................39 Hassai Project – Location of Mines and Prospects..........................................................................47 Diagrammatic Cross-section Showing Relationship of Ariab Deposits ...........................................50 Kamoeb Geology Map......................................................................................................................51 VTEM Geophysical Survey Basis ....................................................................................................56 VTEM Response Types ...................................................................................................................57 VTEM Geophysical Anomaly at Hadal Awatib .................................................................................58 Hassai South Drill Hole Location Plan (AMC, 2009 – 200x100 m Grid)..........................................61 Cross-section Through Hasai South Showing Relationship Between Intersected and True Thickness ..........................................................................................................................................62 Hadal Awatib East Drill Hole Location Plan (AMC, 2009 – 100x50 m Grid)....................................63 Hadal Awatib – Relationship Between Intersected and True Width ................................................64 Kamoeb Drill Hole Location and Topographic Plan - 250 x 250m grid (UTM coordinates 36N, Adindan Datum) .......................................................................................................................65 Kamoeb – 2003 AMC/OMAC Check Assay (Grove, 2003) .............................................................74 Heap Leach Grind Size NPV Trend .................................................................................................95 Locked Cycle Testwork Flow Sheet ...............................................................................................101 5 Mt/a Block Flow Diagram.............................................................................................................110 Hassai South – Main Cu-Au Ore Bodies – Long Section South to North (100 m Grid) ................116 Hadal Awatib East : Surface Trace of Main Mineralised Bodies ...................................................117 Hadal Awatib East – Validation Chart for Sulphide Ore – Grade per Vertical Profile ....................123 Hadal Awatib East Block Model : Oblique Views – Au, Cu, Zn, CuEq, AuEq, Resources Category..........................................................................................................................................124 Hassai South Block Model : Oblique Views – Au, Cu, Zn, CuEq, AuEq .......................................125 Kamoeb Area – Distribution of Mineralised Veins - 2009 model ...................................................128 Kamoeb South Validation Chart: Block Model vs Drill Holes Data by Cross-section....................134 Comparison of Topography November 2007 (Top), October 2008 (middle) and December 2009 (Bottom) .................................................................................................................................137 Tailings Model: Comparison Between Block Model and Drill Hole Data (2007/8 model) .............141 Grade Control v Stacked Grade .....................................................................................................145 FINAL – Rev 0 – 22 Oct 2010 AMEC Page xiii The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 17.11 Figure 18.1 Figure 18.2 Figure 18.3 Figure 18.4 Figure 18.5 Figure 18.6 Figure 18.7 Figure 18.8 Figure 18.9 Figure 18.10 Figure 18.11 Figure 18.12 Figure 18.13 Figure 18.14 Figure 18.15 Figure 18.16 Figure 18.17 Figure 18.18 Figure 18.19 Figure 18.20 Figure 18.21 Figure 18.22 Figure 18.23 Figure 18.24 Figure 18.25 Figure 18.26 Figure 18.27 Figure 18.28 Figure 18.29 Figure 20.1 Figure 20.2 Figure 20.3 Figure 20.4 Figure 20.5 Figure 20.6 Figure 20.7 Figure 20.8 Figure 20.9 Comparison of Remnant Resource Grade for Auger Drilling, Grade Control and Mill Stacking Data..................................................................................................................................146 Kamoeb South – 525 kt/a Optimisation Shell.................................................................................159 Kamoeb North – 525 kt/a Optimisation Shell .................................................................................160 Kamoeb South – Pit Design – Plan view........................................................................................162 Kamoeb North – Pit Design – Plan view ........................................................................................162 Kamoeb – Ore Production Profile...................................................................................................166 Kamoeb – Yearly Mining Profile .....................................................................................................166 Hassai Acidic SBR Stockpile and Heap Leach Tailings Yearly Reclamation Schedule ...............175 Hassai South – 5 Mt/a Optimisation Shell ......................................................................................184 Hadal Awatib – 5 Mt/a Optimisation Shell ......................................................................................184 Hassai South – Stoping Concept ...................................................................................................187 Hassai South – Decline Portal Location .........................................................................................188 Hassai South – Development Long Section ..................................................................................189 Hadal Awatib – Pit Design 5 Mt/a ...................................................................................................190 Hassai South – Oxide Waste Dump Location ................................................................................191 Hadal Awatib – Conceptual Waste Dump Location .......................................................................192 Hassai South – Underground Stoping Schedule ...........................................................................196 Hadal Awatib – 5 Mt/a Ore Profile ..................................................................................................198 Hadal Awatib – 5 Mt/a Mining Profile..............................................................................................198 Bench Face Mapping......................................................................................................................206 Hassai South Underground Deposit Development ........................................................................213 Hassai South Open Stope Stability Chart Design Guidelines........................................................214 Tailings Storage Facility Site Location Option ................................................................................231 Schematic of the Upstream Construction Method .........................................................................235 Metal Production Profile..................................................................................................................258 Metal Production Profile – Heap Leach Operations 2010-2013 ....................................................263 Metal Production Profile – CIL Phase, 2013-2018 .........................................................................265 Metal Production Profile – VMS Phase, 2015-2025 ......................................................................267 Metal Production Profile – Combined CIL and VMS Concentrator Phases ..................................270 Hassai Mine Envisaged Business Plan - Summary Project Schedule ..........................................272 Annual Mining Tonnages (Ore and Waste)....................................................................................277 Average Head Grade, Mine and Plant ...........................................................................................278 Annual Gold Production (kg) ..........................................................................................................278 Diagrammatic Representation of AMC’s Mining Flow Sheet .........................................................280 Illustration of In Situ Grade Control.................................................................................................282 AMC Gold Plant: Crushing/Milling Section Flow Sheet .................................................................286 AMC Gold Plant: Leaching Section Flow Sheet.............................................................................287 Water Sources 2009 .......................................................................................................................290 Sensitivity Analysis – Current Hassai Operation ............................................................................292 FINAL – Rev 0 – 22 Oct 2010 AMEC Page xiv The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 1. SUMMARY 1.1 BACKGROUND La Mancha Resources Inc. (La Mancha) owns a 40% interest in the Ariab Mining Company (AMC) through its purchase of 100% of Cominor in October 2006. AMC conducts an open-pit, heap leach gold operation in the Red Sea State of northeastern Sudan. Production commenced in 1991, with a total of over 2.2 million ounces (Moz) produced to date from multiple deposits. However, gold production has declined in recent years in line with falling head grades and poorer recoveries. 1.2 BUSINESS PLAN La Mancha has reviewed the remaining gold resources and other assets – including copper-bearing volcanogenic massive sulphides (VMS) lying beneath some of the open pits – and has developed a preliminary two-phase business plan, referred to as the VMS Project, to revitalise operations, based on: • A new 3 million tonnes per annum (Mt/a) CIL gold plant to treat: • − Heap leach residues with an average grade of 1.62 g/t Au, at a rate of up to 2 Mt/a − Remaining in situ oxidised gold ore, primarily from the Kamoeb deposit, and other stockpiled material, at a throughput of 1 Mt/a and an average grade of 3.06 g/t. A new 5 Mt/a copper concentrator to process supergene and fresh VMS mineralisation, initially from the Hadal Awatib and Hassai South deposits. 1.3 RESOURCES AND RESERVES 1.3.1 Gold Resources and Reserves 1.3.1.1 Gold Resources, 31 December 2009 Officially reported Measured, Indicated and Inferred gold Mineral Resources at 31 December 2009 are as shown in Table 1.1. These resources include: • Silica-Barite rock (SBR1) and Quartz mineralisation in nine existing areas, namely: − Hadal Awatib (Link, North, Pipe and Junction) − Hassai (North) − Kamoeb (North, South, East and West) − Shulai − Onur − Abderahman − Megzoub − Tagoteb − Umashar • Gold mineralised stockpiles • Heap leach tailings which were drilled during 2007, 2008 and 2009. 1 Weathering product of VMS mineralisation with enriched gold values. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 1 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 1.1 Gold Mineral Resources Summary (31 December 2009) Category Type Measured Heap leach Indicated Heap leach Tonnes Gold Ounces (kt) (g/t) 3 832 1.88 231 075 (oz) No cut-off 2 846 1.97 180 200 No cut-off tailings 886 5.24 149 250 1 577 6.26 317 700 3 823 3.82 469 500 (in situ) Oxide ore, quartz (Au g/t) Type Fire Assay (Intertek) Fire Assay (Intertek) Bulk deposit stockpiles Oxide ore, SBR Assay Bulk deposit tailings Gold ore Cut-off Grade No cut-off Cyanide soluble gold Bulk deposit (Hassai) 1 g/t (HadalAwatib East) Cyanide soluble gold 1.5 g/t (Others) (Hassai) 0.8 g/t Cyanide soluble gold (in situ) (Hassai) Fire Assay (Intertek) Total M+I Inferred Heap leach 12 964 3.23 1 347 725 1 178 2.11 78 252 706 5.79 131 500 Tailings Oxide Ore, SBR Fire Assay (Intertek) Bulk deposit (in situ) Oxide Ore, Quartz No cut-off 2 582 2.68 223 000 1 g/t (HadalAwatib East) Cyanide soluble gold 1.5 g/t (Others) (Hassai) 0.8 g/t Cyanide soluble gold (in situ) (Hassai) Fire Assay (Intertek) Total Inf. 4 467 3.02 432 752 Notes: - Mineral Resources estimated and classified according to CIMM categories by Remi Bosc, QP. - Assay methods and cut-off grades as shown in table. 1.3.1.2 Gold Reserves, 31 December 2009 Mineral Reserves as of the end of 2009 are reported to be 2.56 Mt at 4.99 g/t Au (Table 1.2), and are a sub-set of Mineral Resources. Table 1.2 Mineral Reserves Summary (31 December 2009) As at 31 December 2009 Tonnes g/t Au Ounces Cu Cu (%) (t) Traditional Ore Proven Reserves Probable Reserves Subtotal 1 893 000 4.67 284 000 1 893 000 4.67 284 000 664 000 5.92 126 400 664 000 5.92 126 400 Acidic Ore Proven Reserves Probable Reserves Subtotal FINAL – Rev 0 – 22 Oct 2010 AMEC Page 2 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 1.2 Mineral Reserves Summary (31 December 2009) As at 31 December 2009 Tonnes g/t Au Ounces Cu Cu (%) (t) Total Proven Reserves Probable Reserves - - - - 2 557 000 4.99 410 400 - Notes: - Mineral Reserves estimated and classified according to CIMM code under supervision of Bill Plyley, QP. - Cut-off grade variable according to material type and location, but typically 1.9 g/t for in situ mineralisation. - Gold metal price assumption $750/oz. 1.3.1.3 Additional Gold Resources, Post-2009 Additional Inferred resources have recently been estimated by CSA Global (CSA) to support the CIL expansion phase, comprising material currently undergoing heap leaching, as shown in Table 1.3. Table 1.3 Additional Gold Mineral Resources in Heap Leach Tailings Under Irrigation Category Tonnes Gold – CN Soluble Ounces (kt) (g/t) (oz) Indicated Inferred 514 0.91 14 600 1 329 1.42 58 800 Notes: - Based on cyanide soluble gold assays, using a material balance accounting approach. - Mineral Resources estimated and classified according to CIMM categories by S. McCracken, QP. - No cut-off grade applied, ie bulk deposit, all in situ material reported. Further material which will be stacked and processed by heap leaching prior to commencement of CIL operations is included as existing stockpiles or as in situ resources within Table 1.1. 1.3.2 VMS Resources Generally wide-spaced drilling at Hadal Awatib and Hassai South has allowed estimation of Indicated and Inferred VMS resources as shown in Table 1.4. Table 1.4 VMS Mineral Resources, 31 December 2009 Category Indicated Area/Type Tonnes Gold Copper Gold (kt) (g/t) (%) (oz) Copper (t) HA East, Cu>2% 508 0.78 2.80 12 000 14 200 HA East, Cu<2% 2 390 0.96 0.95 74 000 22 600 0.93 1.27 86 700 36 800 HS South, Supergene HS South, Primary Total Indicated 2 898 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 3 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 1.4 VMS Mineral Resources, 31 December 2009 Category Area/Type (kt) (g/t) (%) (oz) Inferred HA East, Cu >2% 2 930 0.75 2.50 71 000 HA East, Cu<2% 25 400 1.23 0.81 1 001 000 1 530 2.29 2.75 112 000 42 000 HS South, Primary 18 620 1.49 1.37 894 000 255 000 Total Inferred 48 480 1.33 1.19 2 078 000 576 000 HS South, Supergene Tonnes Gold Copper Gold Copper (t) 73 000 206 000 Notes: - HA = Hadal Awatib, HS = Hassai South. - Mineral Resources estimated and classified according to CIMM categories by R Bosc, QP. - Grades based on fire assay for gold, and triple acid digest/AAS finish for base metals; at Intertek, Jakarta. - Cut-off grade 0.8% copper equivalent (Cueq), where Cueq = Cu(%) + 0.63xAu(g/t). - The above relationship uses metal prices of $750/oz gold and $2.00/lb copper, and takes account of metallurgical recoveries. Drilling and mining has located VMS-style base metal mineralisation at 10 other deposits, indicating significant potential to expand the resource base, but insufficient drilling has been undertaken to define further resources at this stage. At this stage, insufficient metallurgical testwork and engineering studies have been undertaken to determine any VMS Mineral Reserves. 1.4 USE OF INFERRED RESOURCES IN BUSINESS PLAN The VMS Project as defined in the Business Plan is based partly on Inferred Mineral Resources which are defined under NI 43-101 as “that part of a Mineral Resource for which quantity and grade or quality can be estimated on the basis of geological evidence and limited sampling and reasonably assumed, but not verified, geological and grade continuity. The estimate is based on limited information and sampling gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes. Due to the uncertainty that may be attached to Inferred Mineral Resources, it cannot be assumed that all or any part of an Inferred Mineral Resource will be upgraded to an Indicated or Measured Mineral Resource as a result of continued exploration. Confidence in the estimate is insufficient to allow the meaningful application of technical and economic parameters or to enable an evaluation of economic viability worthy of public disclosure. Inferred Mineral Resources must be excluded from estimates forming the basis of feasibility or other economic studies.” FINAL – Rev 0 – 22 Oct 2010 AMEC Page 4 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Under National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101), a company subject to NI 43-101 is not permitted to report results of an economic analysis that includes Inferred Mineral Resources, once the project has advanced past a preliminary feasibility stage. However, La Mancha applied to the British Columbia, Alberta, and Ontario Securities Commissions for exemption from this limitation for the VMS Project, and such exemption has been granted subject to: • • Submission of a Technical Report that: − Summarises scientific and technical information concerning exploration, development and production activities on the Ariab Gold property − Addresses the impact of the development of the VMS Project on the existing Ariab Gold Project. Complies with relevant provisions of NI 43-101 regarding the speculative nature of a Preliminary Assessment based on Inferred Mineral Resources which prohibits their categorisation as Mineral Reserves. A statement that Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability is to be included in the Technical Report. 1.5 MINING As part of the VMS Project, mining studies for each of the two phases have been undertaken into: • Open pit mining of oxidised gold ore at Kamoeb and elsewhere, essentially continuing current mining practices • Retrieval of stockpiled material • Reclaim of heap leached material for reprocessing • Open pit mining of VMS mineralisation beneath the existing pit at Hadal Awatib, assuming 120 t excavators and 90 t trucks, in line with current practices • Underground mining by contractor at Hassai South by sub-level open stoping (SLOS), with paste backfill to overcome problematic hanging wall conditions. 1.5.1 Kamoeb Open Pit Pit optimisations were undertaken on the Kamoeb South and Kamoeb North deposits, and mine designs completed. Table 1.5 contains design parameters for the Kamoeb open pits, while pit designs for Kamoeb South and Kamoeb North are shown in Figure 1.1 and Figure 1.2. Table 1.5 Pit Design Parameters Batter Angle Bench Height Berm Width Ramp Grade Ramp Width (deg) (m) (m) (1 : x) (m) 63 10 5 10 15, 22 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 5 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 1.1 Kamoeb South – Pit Design – Plan View Figure 1.2 Kamoeb North – Pit Design – Plan View FINAL – Rev 0 – 22 Oct 2010 AMEC Page 6 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 1.5.2 Heap Leach Tailings and Stockpile Reclaim Heap leach tailings will be reclaimed at a rate of up to 2.0 Mt/a, while acidic SBR stockpiles will be reclaimed as required, ensuring that the fresh ore plant throughput of 1 Mt/a will be maintained for as long as possible. The existing heap leach tailings will be reclaimed by bulldozer and front end loader (FEL) into a mobile feeder system. This in turn transfers the reclaimed material to an overland conveyor, which has been assumed to be 2500 m in length. The overland conveyor feeds a storage bin at the milling area. Acidic SBR stockpile material will be reclaimed by bulldozer and FEL into trucks and transported to the ROM bin at the crusher plant. 1.5.3 Hadal Awatib Open Pit Limited geotechnical data is available directly for the VMS deposits, but a site visit and review of structures and rock conditions in and around the existing open pits by a geotechnical specialist has provided a basis for the design assumptions adopted for open pit and underground mining of the VMS deposits. Pit optimisation was undertaken, and preliminary mine design completed. Table 1.6 contains design parameters for the Hadal Awatib open pit. The design includes a ramp, 22 m wide with a grade of 1:10. Table 1.6 Hadal Awatib Open Pit Design Parameters Sector Max. Interramp Angle Bench Face Angle Bench Height (m) o Weathered zone 46 North wall 49o South wall 47 o 10 65o 10 o 10 75 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 7 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report The pit design is shown in Figure 1.3. Figure 1.3 Hadal Awatib – Pit Design 5 Mt/a 1.5.4 Hassai South Underground Mine Design For the Hassai South underground mine it has been proposed that open stopes in the primary ore zone should have unsupported maximum dimensions of 35 m high (based on a 30 m sub-level spacing) by 30 m along strike and 17 m transverse width (stopes supported with Garford bulge cables in the back could be mined up to 40 m transverse width). It is recommended that these stopes are mined in a primary-secondary sequence, and bottom up, using rapid cycle times and paste backfill. For stopes in the supergene zone immediately below the pits, it is proposed that unsupported maximum stope dimensions should be 25 m high (based on a 20 m sub-level spacing) by 20 m along strike and 17 m transverse width. Additionally, 10 m rib pillars have been proposed every 40 m to provide global stability to the south pit wall (in which the access ramp is contained) during mining. A temporary crown pillar of 20 m to 30 m below the supergene zone has been recommended. Access is by decline with a portal in the lower part of the pit on the hanging wall side. Dual decline and level development requirements are designed to allow multiple stoping fronts on each level. Link drives have been included to simplify traffic flow between the west and the east sides of the mine. An additional fresh air intake has been included at the top of the East incline, which would be used as the second means of egress via an installed ladder way. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 8 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 1.4 is a long section outlining the underground development concept. Figure 1.4 Hassai South – Development Long Section 1.5.5 Production Schedules Production schedules have been determined for both the CIL and Concentrator plants, mining fleet requirements determined and capital and operating costs estimated. It has been assumed that mining at Kamoeb and reclaim of existing heaps is undertaken using existing equipment, whereas new equipment is assumed for open pit mining at Hadal Awatib. Underground mining will be undertaken by a contractor. 1.5.5.1 CIL Plant Feed The 3 Mt/a CIL Plant is fed by a combination of Kamoeb South and Kamoeb North open pit, stockpiled acidic SBR ore and ore from heap leach tailings. Table 1.7 summarises the mining schedule for Kamoeb. Figure 1.5 shows the total material movement schedule for Kamoeb. Table 1.8 displays the heap leach and stockpile reclaim schedule. Figure 1.6 shows the stockpiled ore and heap leach tailings reclaim movements. Note that for the financial model, La Mancha has added an additional 372 088 t at 4.29 g/t Au spread over Years 4, 5 and 6 to bring the scheduled tonnages up to 3.0 Mt/a. This material has been identified overlying VMS mineralisation in the Hadal Awatib pit design. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 9 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 1.7 Kamoeb – Yearly Mining Schedule Year Pit Total Ore Grade Total Waste Strip Total Rock Total Input to Input to Mined Ratio Mined Ounces Mill Mill (t) (t) Output from (t) (g/t) Mill (oz) 1 KamS 370 193 4.08 2 967 068 8.01 3 337 261 48 609 2 KamS 531 437 3.65 3 261 278 6.14 3 792 715 62 423 3 KamS 537 740 3.52 3 299 665 6.14 3 837 405 60 897 4 KamS 533 077 3.38 3 671 049 6.89 4 204 126 57 898 5 KamS/KamN 1 001 688 2.64 6 918 657 6.91 7 920 346 84 973 6 KamS/KamN 757 244 2.59 4 969 725 6.56 5 726 970 63 006 7 KamN Total 420 691 2.23 1 549 574 3.68 1 970 265 30 158 4 152 071 3.06 26 637 016 6.42 30 789 087 407 965 Note: Year 1 is 2010. Figure 1.5 Kamoeb – Yearly Mining Profile FINAL – Rev 0 – 22 Oct 2010 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 1.8 Mining Schedule for Heap Leach and Stockpile Reclaim Acidic SBR Heap Leach Tailings Metal. Tonnes Grade (t) (g/t) Tonnes Grade (t) (g/t) Acidic SBR washable 120 986 5.68 70 4 Acidic SBR non-wash/HL Tailings 466 923 6.00 92 2 000 000 1.62 5 Acidic SBR non-wash/HL Tailings 71 529 6.00 92 1 857 962 Year 1 Area Recovery (%) Total Metal. Recovery (%) Tonnes Grade (t) (g/t) Metal. Gold Recovery Production (%) (oz) 120 986 5.68 70 15 466 70 2 466 923 2.45 80 155 658 1.62 70 1 929 491 1.78 73 80 317 1.62 70 2 126 744 1.62 70 77 406 2 3 6 HL Tailings 2 126 744 7 HL Tailings 2 392 053 1.62 70 2 392 053 1.62 70 87 062 8 HL Tailings 3 000 000 1.62 70 3 000 000 1.62 70 109 189 9 HL Tailings 871 489 1.62 70 871 489 1.62 70 31 719 12 248 248 1.62 65 12 907 686 1.78 73 556 817 Total 659 438 5.94 88% FINAL – Rev 0 – 22 Oct 2010 AMEC Page 11 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 1.6 Hassai Acidic SBR Stockpile and Heap Leach Tailings Yearly Reclamation Schedule Graph The material included in the schedule includes some Inferred Resources and does not constitute a Mineral Reserve. It is referred to as a Mining Inventory and is made up as summarised in Table 1.9. The Mining Inventory includes ore that is included in current in situ reserves that will be mined and leached prior to 2013, with residual gold content estimated based on previous leaching recoveries. Table 1.9 Mining Inventory for CIL Operation Ore Type Traditional ore* Heap leach residue Acidic ore Total Mining Inventory Tonnage Gold Gold (kt) (g/t) (oz) 3 085 2.90 287 386 12 248 1.62 636 830 538 6.00 103 926 15 871 2.02 1 028 142 Traditional Ore* = SBR and Quartz mineralisation of the type previously heap leached, including 372 kt grading 4.29 g/t Au identified above VMS mineralisation in the proposed Hadal Awatib VMS pit.. 1.5.5.2 VMS Plant Feed The 5 Mt/a VMS concentrator is fed by a combination of underground and open pit ore as shown in Figure 1.7 and Figure 1.8 respectively. Figure 1.9 shows the total material movement schedule for Hadal Awatib. Again, Inferred Resources are included for this preliminary economic evaluation, and insufficient studies have been completed to confirm mining and processing parameters. Consequently, the schedule is based on a Mining Inventory which totals 29.36 Mt from the two sources. FINAL – Rev 0 – 22 Oct 2010 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 1.7 Hassai South – Underground Stoping Schedule (Ore) 2,000 1,800 1,600 1,400 Tonnes (kt) 1,200 1,000 800 600 400 200 0 Y1 Y2 Y3 Y4 Y5 Y6 Y7 Y8 Y9 Y10 Y11 Y12 Y13 Y14 Y11 Y12 Y13 Y14 Period Primary Supergene Figure 1.8 Hadal Awatib – 5 Mt/a Ore Mining Profile 4,000 3,500 3,000 Tonnes (kt) 2,500 2,000 1,500 1,000 500 0 Y1 Y2 Y3 Y4 Y5 Y6 Y7 Y8 Y9 Y10 Period Primary Supergene FINAL – Rev 0 – 22 Oct 2010 AMEC Page 13 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 1.9 Hadal Awatib – 5 Mt/a Mining Profile 30,000 25,000 Tonnes (kt) 20,000 15,000 10,000 5,000 0 Y1 Y2 Y3 Y4 Y5 Y6 Y7 Y8 Y9 Y10 Y11 Y12 Y13 Y14 Period Waste 1.5.6 Oxide Primary Supergene Mine Capital Cost Total mine capital cost of $151 M (Table 1.10) includes the purchase of a new mining fleet for Hadal Awatib, together with pre-production costs associated with the development of the Hassai South underground mine. Replacement capital costs have been included for mining at Kamoeb for a portion of the aging mining fleet. No capital costs were included for the heap leach tailings and stockpile reclaim operations which will use existing equipment. Table 1.10 Mine Capital Cost Estimate Area Items Kamoeb Open Pit Equipment Cost Estimate ($M) Hadal Awatib Open Pit Sub-total 8.8 Equipment 79.6 Infrastructure Sub-total Hasai South Underground Infrastructure 31.2 23.8 Sub-total Note: 5.0 84.6 Development Material movement Total Mine Capital Cost 8.8 2.6 57.6 151.0 Underground infrastructure includes preliminary works, surface works, portal development, ventilation and the paste plant. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 14 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 1.5.7 Mine Operating Cost It is assumed that open pit mining at Kamoeb and Hadal Awatib, and heap leach and stockpile reclaiming all will be undertaken as per current operations, ie utilising a local workforce. However, underground mining will be undertaken by a qualified mining contractor. Costs vary by location and change over time. Details are provided in Section 18, and average costs are summarised in Table 1.11. Open pit costs have been derived from historical AMC data, plus input from AMEC Minproc’s internal database, and include transport to the plant. Underground mining costs have been derived from AMEC Minproc’s internal database, and include costs associated with continuing underground development, backfill, mine services and management of operations. Table 1.11 Mine Operating Cost Estimate Ore Source Average Cost Estimate ($/t ore) Heap leach reclaim 1.14 Kamoeb open pit 19.87 Hadal Awatib open pit 14.14 Hassai South underground 26.17 1.6 PROCESSING 1.6.1 CIL Plant 1.6.1.1 Metallurgical Testwork A limited metallurgical testwork program was undertaken by Amdel Mineral Laboratories, Perth, on samples of Kamoeb (Quartz) ore and heap leach residue. This work included sample preparation, head sample chemical analysis, mineralogy, gravity separation, grind sensitivity, oxygen/air sparging and cyanidation leach testing. The testwork findings can be summarised as follows: • The Quartz ore is abrasive and exhibits strong compressive strength, with soft SAG milling characteristics and a medium ball milling work index • Heap leach material exhibits a soft to medium ball mill work index • Head assaying indicated average gold grades of 4.29 g/t and 2.15 g/t for the Quartz ore and heap leach residue samples, respectively • Significant levels of mercury were detected in both samples which will require monitoring in future testing, and may require addition of a mercury scrubber to the processing flow heet • The Quartz ore gave reasonable gravity response at 24% recovery, while heap leach material gave a poor response, achieving only 4% recovery to concentrate • Quartz ore appears to be insensitive to grind size below 80% passing 150 microns (P80 150 µm), whereas the heap leach material shows improving recovery with grind; NPV analysis shows that with the heap leach material in isolation, the best NPV was achieved at the finest grind tested of P80 53 µm. Grinding finer than P80 53 µm may present further increases in NPV FINAL – Rev 0 – 22 Oct 2010 AMEC Page 15 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report • Both ore types were found to be insensitive to the use of oxygen sparing over air sparging within the leach, with similar recoveries being obtained • Similarly, cyanide concentration did not appear to influence recovery over the range tested between 250 ppm to 1000 ppm NaCN • Neither pre-aeration nor lead nitrate addition improved the leach performance of the heap leach material. Anticipated recovery in an agitated leach circuit is expected to be 93.0% for the Quartz ores and 65.8% for the heap leach material, with preliminary NaCN consumptions of 1.45 kg/t and 1.38 kg/t respectively. 1.6.1.2 Anticipated CIL Production Plant Recoveries and Reagent Consumptions La Mancha has reviewed the testwork results and developed anticipated recoveries for different components of the proposed CIL Mining Inventory. This approach distinguishes between: • Quartz ore: 93% recovery based on preliminary testwork • SBR material: 95% recovery, based on production experience and tails grades from CIL teswork • Acidic SBR material:: 92% recovery, based on experience • Heap leach residue: 65% of fire assay gold, or 88% of cyanide soluble gold. Overall average recoveries are estimated to be 70% of head grade as shown in Section 16.2.4. The annual gold production profile used in the economic analysis was developed by La Mancha, by assigning recoveries for particular mill feed types to the annual mining schedule provided by CSA. CIL plant Lime and cyanide consumptions for heap leach residue are estimated to be 0.9 kg/t and 1.38 kg/t, based on testwork. These consumptions are expected to rise to 10.0 kg/t and 2.5 kg/t for Acidic SBR material, based on previous heap leach experience. 1.6.1.3 CIL Process Plant The 3.0 Mt/a plant feed grade is based on 2.0 Mt/a of heap leach tailings and 1.0 Mt/a of fresh ore, which gives a calculated average grade of 2.01 g/t Au. The proposed circuit includes: • Crushing: a new primary crushing circuit consisting of a coarse ore bin, feeder, jaw crusher, conveyor and crushed ore bin. The current circuit will be made redundant. • Reclamation of the existing spent heap leach residue by bulldozer or FEL into a mobile feeder system. The residue will be transferred to a 2500 m overland conveyor, which will feed a storage bin located adjacent to the milling area. • A SABC grinding circuit consisting of an open circuit SAG mill operating with a scats crusher, followed by a single overflow ball mill operating in closed circuit with a set of hydrocyclones. The SAG mill will process fresh ore at 1.0 Mt/a while the ball mill will receive an additional 2.0 Mt/a of heap leach residue feed. The target design cyclone overflow sizing will be P80 75 μm. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 16 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report • A gravity recovery circuit consisting of a centrifugal concentrator and an intensive leach reactor. • A leaching and adsorption circuit with a combined residence time of 24 hours. The leach circuit will consist of two leach tanks followed by six carbon adsorption tanks operating in series with an average tank sizing of 1881 m³. Design leach feed density will be 44.3% weight for weight (w/w) solids. • A split_AARL elution circuit utilising a combined acid wash and elution column. Elution batch size will be 11 t of carbon. Gold recovered from the elution process will be electrowon in two stainless steel electrowinning sludging cells operating in parallel. • Cyanide detoxification (detox) based on the SO2/Air process for the destruction of excess cyanide in the tailings slurry. Total residence time of the circuit is 90 minutes. • Tailings thickening to recover water from the tailings stream, to minimise tailings pumping demands and to minimise site water losses due to evaporation. The design thickener diameter is 25.0 m. • A square, unlined, paddock style tailings storage facility (TSF) utilising a perimeter spigoting system, with a central decant tower for excess water recovery and recycle back to the processing plant. A rise rate of 2.0 m/a is assumed in the preliminary design requiring the TSF dimensions to be 1035x1035 m. Plant operating costs are based upon La Mancha supplied processing schedule, feed grades, circuit recoveries and gold production quantities and are estimated at $12.80/t, or $250.71/oz Au. 1.6.2 VMS Concentrator 1.6.2.1 Metallurgical Testwork A limited metallurgical testwork program was undertaken by SGS Canada Inc. (SGS), using two ore composites. This work included sample preparation, head sample chemical analysis, mineralogical analysis, flotation testing, cyanidation leach testing and product characterisation testwork. Testwork to date has not included any comminution, thickening, or filtration work. Equipment sizing in these areas is, therefore, based on assumed parameters and AMEC’s experience from other projects. Testwork was performed on two composites, one each from the supergene and primary zone of the Hassai South deposit. No samples were collected from Hadal Awatib, and the grades of the composites were appreciably higher than the expected average grade of these ore types. Head assays and QEMSCAN mineralogical characterisation was undertaken, showing pyrite (55%) and chalcopyrite (23%) are the predominant sulphide minerals in both composites. The majority (>70%) of the chalcopyrite was liberated at P80 100 µm, but fine grinding would be required to liberate 18-25% of chalcopyrite that occurred as composites with pyrite. A total of 13 rougher kinetics and batch cleaner flotation tests were performed on Composite 1 and Composite 3. In addition, a locked cycle test was performed on each of the composites. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 17 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report The main conclusions from the testwork were: Composite 1 – Enriched Zone • • Batch Testing − Batch testing with three cleaning stages recovered 76% of the copper to a 32.5% Cu concentrate − In the same concentrate, 65% of the gold was recovered − Pyrite scavenger flotation was able to recover 92% of the gold in the rougher tail but with 72% of the rougher tail mass going to concentrate. − Gravity concentration was able to produce a high grade gold concentrate (713 g/t) but at only 2.3% recovery of gold in feed − Rougher flotation was able to recover more than 97% of the copper and more than 82% of the gold. Locked Cycle Testing − Locked cycle testing recovered 87% of the copper to a concentrate grading 30% Cu − Gold recovery to the same concentrate was 73% − Pyrite scavenger flotation was able to recover 90% of the gold in the rougher tail, but with the majority (70%) of the rougher tail mass going to concentrate − Rougher flotation was able to recover 93% of the copper and 84% of the gold. Composite 3 – Primary Zone • • Batch Testing − Batch testing with three cleaning stages recovered 80% of the copper to a 30% Cu concentrate (and 85% of the copper to a 19% Cu concentrate) − In the same concentrates, respectively, only 30 and 37% of the gold was recovered − Pyrite scavenger flotation was able to recover 96% of the gold in the rougher tail but with 87% of the rougher tail mass going to concentrate. − Gravity concentration was able to produce a moderate grade gold concentrate (35 g/t) but at only 1.5% recovery of gold in feed − Rougher flotation was able to recover more than 93% of the copper and about 60% of the gold. Locked Cycle Testing − Locked cycle testing recovered 85% of the copper to a concentrate grading 25% Cu. − Gold recovery to the same concentrate was only 38% − Pyrite scavenger flotation was able to recover 95% of the gold in the rougher tail but with the majority (88%) of the rougher tail mass going to concentrate − Rougher flotation was able to recover 88% of the copper but only 45% of the Au. The primary grind in each case was to a P80 of 69 µm. The secondary grind took the concentrate down to 29 µm. The flotation reagents are simple with a xanthate collector, a common frother and lime as a pH modifier to adjust the rougher and cleaners to alkaline flotation conditions. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 18 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Enriched zone ore is easier to float than primary ore. Primary ore has especially low gold recovery with much of the value remaining in the non-floated pyrite. Cyanide leach testwork on concentrator tailings indicated low recoveries and high cyanide consumptions, and this option has not been pursued further at this stage. Interpretation of the testwork results suggests the concentrate grade and recovery parameters as follows: • Hassai South Supergene: 7% mass recovery to concentrate grading 32% Cu and 22 g/t Au, with recoveries of 81% and 67% for Cu and Au, respectively • Hassai South Primary: 4.9% mass recovery to concentrate grading 5.5% Cu and 11 g/t Au, with recoveries of 90% and 56% for Cu and Au, respectively • Hassai South Supergene: 3.4% mass recovery to concentrate grading 25.1% Cu and 10 g/t Au, with recoveries of 85% and 29% for Cu and Au, respectively. 1.6.2.2 Process Plant The VMS process flow sheet is preliminary due to the limited amount of testwork completed to-date. It comprises: • Comminution Circuit − Ore is delivered by mine haul truck to a ROM ore pad. − Crushing: ore is loaded into the ROM bin by FEL, is withdrawn from the ROM bin by vibrating grizzly feeder and passes to the primary jaw crusher. The undersize from the vibrating grizzly and the primary crusher discharges onto the mill feed conveyor. A weightometer and a tramp magnet are mounted over the head pulley. − Grinding and classification: the grinding circuit consists of an open circuit 5.2 MW SAG mill, followed by a ball mill in closed circuit with two clusters of 400 mm hydrocyclones. The SAG mill trommel oversize falls into a bunker for removal by loader or bobcat. Cyclone overflow flows by gravity to a static trash screen prior to reporting to the rougher flotation circuit, while the cyclone underflow stream is returned to the ball mill. The ball mill is 7.3 m diameter inside shell, with an EGL of 10.2 m. It is powered by twin 4.5 MW motors, for a total power of 9.0 MW. The ball mill discharge flows through a trommel. Undersize from the trommel cascades into the common mill discharge hopper. • Rougher Flotation and Regrind − Rougher flotation is nominally carried out at P80 69 µm. The rougher circuit is in open circuit, with rougher tailings reporting to the tailings thickener. The rougher stage of flotation consists of two trains of 6 x 100 m3 forced air tank cells. The installed residence time for the rougher flotation cells is 40 minutes. Flotation is undertaken at elevated pH of 10.5. Aerofloat 238 is added as collector and methyl iso-butyl carbinol (MIBC) as frother. Rougher concentrate gravitates to a concentrate hopper, and is pumped to the regrind circuit. Flotation tailings are passed to the tails thickener feed tank. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 19 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report − • The regrind circuit consists of two ISAmill M1000s, with 500 kW motor units, operating in open circuit. The regrind circuit is designed to produce a regrind target of P80 30 µm. The feed to the regrind mill is deslimed with a 150 mm hydrocyclone cluster. The regrind rougher concentrate is transferred to the cleaner flotation circuit. Cleaner Flotation • − Reground rougher concentrate is treated in three stages of closed-circuit cleaning. The concentrate is combined with the second cleaner tailings and cleaner scavenger concentrate in the first cleaner cell feed box to form first cleaner feed. First cleaner flotation is carried out in six tank cells with a nominal residence time of 20 minutes total. − The first cleaner concentrate is pumped to the second cleaner feed where final concentrate is produced. The first cleaner tailings gravitate to the cleaner scavenger bank, from which the non-floating component is transferred to final tails. − The second cleaner stage and the cleaner scavenger stage each have a nominal residence time of 15 minutes. The second cleaner consists of four tank flotation cells, while the cleaner scavenger bank consists of three tank flotation cells. − Final cleaner concentrate is stored in an agitated tank to promote de-aeration, prior to pumping to the concentrate thickener. Concentrate Thickening, Filtration and Handling − Final cleaner concentrate passes over a trash screen to the concentrate thickener. Thickener overflow is pumped to the process water tank. Thickener underflow (65% w/w solids) is pumped to the concentrate storage tanks. Two concentrate storage tanks have been provided with a live capacity of 1000 m3 each, allowing a total storage capacity of 48 h. − Filter feed pumps feed two pressure filters. Dry cake (10% moisture) is dumped from the bottom of each filter to a conveyor belt that discharges into a storage bunker. The filtrate gravitates to an air/water separator in which the filtrate is de-aerated prior to being pumped back to the concentrate thickener. − Concentrate is packed in 2 t bulk bags, and trucked to Port Sudan. − Concentrate storage capacity at site and port are approximately 15 and 30 days respectively. − Average concentrate production rates vary according to feed, as shown in Table 1.12. Table 1.12 Average Concentrate Production Ore Type t/d at 10% Moisture Hassai South Supergene 1065.0 Hassai South Primary 745.6 Hadal Awatib 513.0 1.7 INFRASTRUCTURE 1.7.1 Power Power for the existing operation (5.5 MW) is provided by diesel generator, but there is insufficient capacity to support the plant expansions. Site power requirements are estimated to be 10.3 MW and FINAL – Rev 0 – 22 Oct 2010 AMEC Page 20 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 19.5 MW for the CIL plant and concentrator respectively, ie a total of 29.82 MW. It is proposed to construct a 77 km long, high tension (HT) power line to link with the national grid supply, as part of the CIL plant development. 1.7.2 Water The project currently is supplied from groundwater bores, but demand will increase very significantly once the new plants are installed. It is proposed to construct a water pipeline to bring water from the River Nile; this would be constructed as part of the CIL project, but with sufficient capacity to supply the 5 Mt/a concentrator once it comes on line. It is estimated that six pump stations will be required along the line. Preliminary discussions have commenced regarding acquisition of water permits, without raising any issues. 1.7.3 Accommodation The existing accommodation camp would require a significant upgrade to bring it up to acceptable standard, and allowance has been made in the capital costs for a new camp to be constructed. 1.7.4 Access and Port Existing roads linking the site to Port Sudan are considered adequate to support the expanded operations, including transport of concentrates and reagents. Port Sudan has adequate facilities for exporting concentrates and importing reagents and consumables for the expanded operation. AMC has a 4500 m2 fenced yard at Port Sudan, and it is anticipated that a 1800 m2 concentrate storage shed will be erected there to house VMS concentrate bags prior to container loading and shipping to customers. 1.8 ENVIRONMENTAL AMEC Earth and Environmental conducted a high-level environmental review to identify any serious issues and/or opportunities related to current operations and the proposed expansions. A number of areas were identified that require attention in order to improve monitoring, reporting and response systems. However, no major issues were identified that are likely to significantly impact on development of the expansion projects. 1.9 CAPITAL COSTS 1.9.1 General Capital costs for the CIL plant phase have been estimated by CSA (mine) and Sedgman (plant and infrastructure). Capital costs for the concentrator phase, including development of VMS mining areas, have been estimated by AMEC. Costs are expressed in United States dollars as of first or second quarter 2010: accuracy is ±35-40%. 2 Note: An additional 7.2 MW is required to operate the water pipeline, but this will be supplied by local diesel generator sets. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 21 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 1.9.2 CIL Plant Development The estimated capital cost for the 3.0 Mt/a CIL plant and related infrastructure is $184.4 M. Additional capital expenditure of $8.83 M is required for upgrading a portion of the current, aging mining fleet. The following infrastructure items contribute significantly to these capital costs: • Overhead powerline $25.4 M • Nile water pipeline $39.6 M • Overland conveyor $10.0 M • New 200 man camp $4.0 M 1.9.3 VMS Concentrator Development The capital cost estimate for the VMS concentrator phase includes mining and certain necessary infrastructure items, but excludes joint infrastructure, such as the power line and water pipeline, that have been costed as part of the CIL plant phase. Total capital costs for the concentrator phase are estimated to be $319.43 M, as summarised in Table 1.13. Area Open pit mine Underground mine Process plant Infrastructure Area infrastructure Regional infrastructure Miscellaneous Indirect costs Accuracy Provision Total Initial Capital Cost Table 1.13 Capital Cost Estimate, 5 Mt/a VMS Concentrator Phase Capital Cost ($M) 71.66 44.74 78.19 38.83 12.90 0.50 11.42 38.39 22.79 319.42 1.10 OPERATING COSTS 1.10.1 General As for capital costs, operating costs are expressed in United States dollars, of second quarter 2010. CIL-phase operating costs were estimated by CSA (mine) and Sedgman (plant and infrastructure), while AMEC estimated costs for the VMS Concentrator phase. G&A costs were provided by AMC based on current operating mine experience. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 22 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 1.10.2 CIL Plant Under this scenario, Kamoeb life of mine (LOM) mining operating costs are estimated at $83.3 M, equivalent to $2.87/t moved or $20.07/t of ore treated. An additional $14.8 M is expended on reclaiming stockpiled ore and heap leach tailings over the life of the plant, at an average cost of $1.15/t. Plant operating costs were based upon the following: • The proposed processing plant design; • The resulting reagent consumptions from metallurgical testing. • The AMC specified processing schedule, feed grades, overall recovery and gold production quantities. For the LOM, the plant operating costs were estimated at $203.2 M, equivalent to $12.80 /t of ore treated, or $250.71/oz Au recovered. G&A and Other costs have been provided by AMC based on current Hassai site data and are estimated to be $9.2 M/a for the CIL operation. 1.10.3 VMS Concentrator Underground mining costs for Hassai South have been estimated at $26.17/t ore, including ongoing development costs. Open-pit mining costs at Hadal Awatib are estimated to be $0.47/t of material or $14.14/t ore. Process operating costs for the VMS concentrator, including transport to port, are estimated to be $9.38/t of ore treated or $46.9 M/a, which equates to 39 ¢/lb copper shipped in concentrates. Off-site charges are not included in those operating costs. G&A specific to the VMS operation is estimated by AMC at $9.24 M/a. 1.11 PROJECT SCHEDULE A high level schedule for the project as proposed by La Mancha is outlined in Figure 1.10, and shows the construction of the 3 Mt/a CIL plant completed by 2013, followed by development of the 5 Mt/a VMS concentrator by 2015. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 23 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 1.10 Hassai Mine Envisaged Business Plan – Summary Project Schedule 2010 2011 2012 2013 2014 2015 2016 2017 2018 2019 2020 Hassai CIP Project Current Heap Leach Operation CIP PFS/DFS CIP EPCM CIP - Commission and Operate Ariab VMS Project Scoping Study Drilling and Testwork Program Prefeasibility Study Definitive Feasibility Study EPCM Plant - Commission and Operate It is recommended that construction of the CIL and VMS process plants and associated facilities be executed on an engineering, procurement and construction management (EPCM) basis. Long lead items such as ball mills, crusher, flotation cells and filters would be identified as a matter of priority during the feasibility study (FS) phase of the project, to allow early purchase of these key items. Use of second-hand equipment may provide some schedule and cost reductions. 1.12 FINANCIAL MODELLING La Mancha has prepared three post-tax financial models for preliminary economic assessment, covering: • Base Case: existing heap leach operation treating remaining oxide gold reserves through to the end of 2013. • CIL Project, starting in 2013 and treating then-extant oxide gold reserves, stockpiled acidic mineralisation and heap leach tailings resources. • VMS project, starting in 2015 treating the identified VMS resources. Modelling is based on a phased expansion and production profile. Current Heap Leach operations continue to end of 2012. Phase One CIL operates alone from 2013 to end of 2014. Phase two VMS operates in parallel to Phase One from 2015. The gold and gold-equivalent copper production profile is shown in Figure 1.11. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 24 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 1.11 Metal Production Profile for Phased VMS Project 500,000 450,000 VMS Copper as Gold eqv. 400,000 VMS Concentrate Heap Leach Residue Gold Production, oz 350,000 Acidic non‐washable ore 300,000 Acidic washable ore SBR ore 250,000 Quartz ore 200,000 150,000 100,000 50,000 ‐ 2010 2011 2012 2013 2014 2015 2016 2017 2018 2019 2020 2021 2022 2023 2024 2025 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 25 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 1.14 Financial Highlights for Proposed VMS Project, by Phase Heap Leaching Phase 1: CIL Phase 2: VMS USD 950/oz USD 950/oz USD 950/oz -- -- USD 2.19/lb 7% 7% 5% -- -- 3.5% Phase 1 & 2 Main Assumptions Gold price Copper price Royalties (%) Gold Copper Corporate tax rate 15% 10% 10% 2.6 @ 4.88 0.6 @ 6.0 - Additional Mineral Resources - 3.8 @ 1.9 - Measured Resources (Mt@g/t) - 4.6 @ 2.1 - Indicated Resources (Mt@g/t) - 6.8 @ 1.7 29.4 @ 1.1, 1.2 Mineral Reserves Probable Reserves (Mt@g/t) Inferred Resources (Mt@g/t Au, Cu%) Total Mining inventory Tonnes, Mt Grades Gold, g/t Copper, % 2.6 15.8 29.4 45.2 4.88 2.01 1.11 1.43 -- -- 1.22 1.22 Production: Commissioning 2010 - 2013 2013 2015 -- Yearly mill run rate, Mtpa 0.65 3 5 -- Gold recovered, ‘000 oz 299 811 378 1 189 Copper recovered, ‘000 t -- -- 323 323 73% 79% 36% -- -- -- 90% -- Gold (oz) 74 780 155 880 59 355 -- Copper (t) -- -- 51 516 4 6 10 6+ $185.6 M $319.4 M $505.0 M $4.9 M $35.9 M $40.8 M Metallurgical recovery Gold Copper Yearly production* Mine life, years Financials: Initial capital cost Total sustaining capital Average cash costs $ 482/oz Au $ 1.24/lb Cu*** - Internal rate of return 30% 11% 17% NPV @ 0% discount $195.8 M $230.9 M $447.1 M NPV @ 5% discount $149.8 M $122.7 M $238.7 M 1.9 3.9 varies Payback** , years Notes: * Costs for years when project is running at design rates. ** Calculated from commencement of production. *** Including gold credits. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 26 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Key assumptions and financial highlights from the CIL and VMS models are shown in Table 1.14. The assumptions include gold and copper prices of $950/oz and $2.19/lb, respectively, over the life of the project. Total gold and copper production for the two phases is estimated at 1.19 Moz and 0.323 Mt, respectively. Average cash costs are $482/oz gold in the CIL circuit and $1.24/lb copper produced from the concentrator (including off-site costs and gold credits). The CIL project shows an NPV of $149.8 M and an IRR of 30%, while the VMS project shows an NPV of $122.7 M and an IRR of 11%. The VMS Project as defined in the Business Plan is based partly on Inferred Mineral Resources which are defined under NI 43-101 as “that part of a Mineral Resource for which quantity and grade or quality can be estimated on the basis of geological evidence and limited sampling and reasonably assumed, but not verified, geological and grade continuity. The estimate is based on limited information and sampling gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes. Due to the uncertainty that may be attached to Inferred Mineral Resources, it cannot be assumed that all or any part of an Inferred Mineral Resource will be upgraded to an Indicated or Measured Mineral Resource as a result of continued exploration. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability. Consequently, the predicted financial outcomes must be treated with a high degree of caution. Sensitivity analysis confirms that both phases of the project are very sensitive to metal prices: an increase of 10% in gold price adds approximately $81 M to NPV, while a similar increase in copper price adds $91.5 M. The CIL and VMS plants each operate at full throughput for only 5 years. The financial outcomes are, therefore, significantly affected by extensions to life of the operation. The potential to increase resources, particularly of VMS material, is considered by AMC to be high, with VMS mineralisation known to lie at the base of four additional existing gold pits, and with a number of other untested electrical conductors identified during exploration. 1.13 CONCLUSIONS Scoping studies have been completed into the possible development of a 3 Mt/a CIL circuit primarily to re-process heap leach tailings, and a 5 Mt/a flotation circuit to treat VMS mineralisation. Resource modelling and mining studies have been undertaken to investigate extraction methods and develop mining schedules to supply feed to these plants. A preliminary geotechnical investigation has been undertaken to support the proposed mining methods and mine designs for the VMS deposits which have not previously been mined. A high-level environmental review indicates that acceptable environmental outcomes should be achievable, assuming standard engineering design and operating practices are employed. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 27 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report The success of the proposed expansion project relies on two key pieces of infrastructure, namely a 77 km long power line connecting to the National Grid and a 165 km pipeline bringing water from the Nile River. Discussions have commenced with the relevant authorities regarding access to power and water. Capital and operating costs have been developed for both process plants and related infrastructure, with the assumption that the power line and water pipeline are funded as part of the CIL project. A project schedule has been developed showing production starting in 2013 for the CIL plant and 2015 for the VMS concentrator. This schedule allows for completion of feasibility studies followed by plant design, the delayed start to the VMS concentrator reflecting the less advanced status of this phase in terms of resource definition, geotechnical studies, mine design, process testwork and mine development. Financial modelling indicates that both the CIL and VMS phases of the expansion project are economically viable, with NPVs of $149.8 M and $122.7 M, respectively. Base case metal prices were $950/oz for gold and $2.19/lb for copper. However, it must be noted that the bulk of the resources contributing to the VMS mine schedule are classified as Inferred, as is a portion of the CIL plant feed. Consequently, there is a high degree of uncertainty in these resources, and their use in economic modelling is not generally allowed under NI43-101. An exemption has been provided by the Canadian securities regulators, allowing use of Inferred resources for a preliminary economic assessment in this instance. The financial outcomes are particularly sensitive to metal prices: a 10% increase in either gold or copper price improves overall NPV by approximately $80-90 M. Plant throughput is at full capacity for only 5 years in both cases, and significant economic upside exists if additional reserves can be located. VMS mineralisation is known to exist at the base of six oxide gold pits, of which only two have been drilled sufficiently to allow resources to be modelled for use in this study. Of these two, the resources at Hassai South have been modelled using large blocks with partial mineralisation estimated within these blocks. In order to undertake underground mining studies, the mineralisation has been regularised and the associated reduction in grade has a significant impact on project economics. It is believed that improvements in the resource estimation/modelling of the Hassai South underground and Hadal Awatib open pit deposits would assist in more accurate spatial definition of the mineralisation and mining-related dilution, and in turn may have the effect of increasing the schedule grades. It should also be noted, however, that there will likely be a drop in the overall mining inventory tonnes, as contained metal would not be affected. Additional resources are expected to be identified at the other known VMS locations, as well as from testing the numerous other geophysical (electrical) conductors identified in the district, potentially allowing full production to be maintained for several more years. 1.14 RECOMMENDATIONS Based on the positive outcomes of the scoping study, additional work is indicated to allow completion of feasibility studies to confirm development of the expansion phases. For the CIL phase, such work would necessarily include: • Additional mining studies at FS level for in situ and tailings/stockpile reclaim mining • Definitive metallurgical testwork on fully representative samples FINAL – Rev 0 – 22 Oct 2010 AMEC Page 28 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report • Plant and infrastructure design and cost estimation to feasibility level • TSF design • Confirmation and design of external power supplies • Environmental studies. To this end, La Mancha has announced award of feasibility study work for the CIL phase, for completion by the end of first quarter 2011, with a view to making an investment decision in the first half of 2011. A budget of A$1.69 M has been approved for this work. In addition, Sudanese for Construction and Oil Services has been contracted to design and cost the water the water pipeline from the Nile River at an estimated cost of US$ 250,000. The VMS component of the project is much less advanced, particularly in terms of resource status. Consequently, a 100 000 m, $18 M exploration program has been approved to: • Convert Inferred VMS resources to Indicated and Measured categories • Test for additional VMS resources beneath the Hadayamet open pit. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 29 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 2. INTRODUCTION 2.1 BACKGROUND La Mancha Resources Inc. (La Mancha) owns a 40% interest in the Ariab Mining Company (AMC) through its purchase of 100% of Cominor in October 2006. AMC currently conducts an open-pit, heap leach gold operation in the Red Sea State of northeastern Sudan. Production commenced in 1991, with a total of over 2.3 million ounces (Moz) of gold produced to date from multiple deposits. However, gold production has declined in recent years, in line with falling head grades and poorer recoveries. The current operation comprises mining, crushing and stacking at a rate of 0.7 Mt/a, with heap leaching using cyanide to recover just over 62 000 oz of gold (2009 figures). Supporting infrastructure includes diesel-powered electric generators, a camp for approximately 600 persons, and water supplied from a nearby borefield which accesses a near-surface aquifer, augmented by a dam to capture any surface run-off. The site is linked by road to Port Sudan on the Red Sea, a distance of some 200 km. La Mancha has reviewed the remaining gold resources and other assets – including copper-bearing volcanogenic massive sulphides (VMS) lying beneath some of the open pits – and has developed a preliminary business plan to revitalise operations, based on: • A new 3 Mt/a CIL gold plant to treat: • − Remaining in situ oxidised gold ore, primarily from Kamoeb deposit, and stockpiles of acidic ore, at a maximum throughput of 1 Mt/a and annual average grade of 3.38 g/t reducing over time to 2.23 g/t Au, depending on the source − Heap leach residues with an average grade of 1.62 g/t Au, at a target rate of 2 Mt/a, increasing once other resources have been depleted. A new 5 Mt/a copper concentrator to process supergene and fresh VMS resources, initially from the Hadal Awatib and Hassai South deposits, with indications of potential in several other areas. 2.2 SCOPES OF WORK Sedgman Limited (Sedgman) was commissioned by La Mancha to complete a scoping study for the development of a CIL plant and infrastructure3. CSA Global (UK) (CSA) was brought in to confirm heap leach residue resources present at the point when CIL processing is scheduled to commence, and prepare mining plans for the Kamoeb deposits in order to feed the CIL plant. At the same time, La Mancha commissioned AMEC Minproc Limited (AMEC) to undertake scoping level assessment of the potential for a VMS concentrator to be developed, including: • A geotechnical assessment of ground conditions, including a review of conditions within existing pits • Mining studies for open pit and/or underground extraction of VMS mineralisation • Preliminary assessment of environmental conditions as they relate to the VMS concentrator and tailings storage facility (TSF). 3 A preliminary scoping study into the potential economics of a CIL operation was partly completed by Sedgman in 2008. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 30 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report In situ mineral resources comprise oxide gold and VMS resources that have been determined to NI 43-101 standard and previously reported by Remi Bosc of Arethuse Geology Sdn Bhd (Arethuse). Arethuse also completed resource estimation for that portion of the heap leach tailings tested by auger drilling in 2007/08 and 2009. Additional heap leach tailings resources have been determined by CSA for use in the scoping study, and the same company has determined a new Mining Inventory for open pit mining at Kamoeb under the CIL scenario. The scopes called for contributions to the La Mancha NI43-101 Technical Report discussing the potential development of new plants to treat oxide gold ores (CIL plant) and VMS ores (VMS concentrator). The Technical Report describes the resources, available geotechnical information, proposed open pit and underground mining, metallurgical testwork, process and plant design, operating and capital cost estimates, and preliminary economic assessment for the project. La Mancha’s business plan calls for initial CIL processing in 2013, while the VMS concentrator has been scheduled to commence production in 2015. However, La Mancha requested that power and water infrastructure sufficient for a combined project be included with the CIL plant, and preliminary work in these areas has been completed by Sedgman. In addition to work by AMEC, Arethuse, CSA and Sedgman, information for the Technical Report has been provided by AMC personnel with regards to project history (including exploration history) licensing/permitting, current operations (including site operating costs), current Mineral Reserves and the financial analysis. This Technical Report has been completed in accordance with Form 43-101F Techncial Report of the Canadian Securities Administrators National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101). The Report is based on the outcomes of resource, mining and engineering studies completed by AMEC, Arethuse, CSA and Sedgman as noted in this Technical Report. 2.3 PRINCIPAL SOURCES OF INFORMATION In contributing to this report, AMEC, Arethuse, CSA and Sedgman have relied on information provided by AMC regarding current operations, plus various data, reports, maps and technical papers listed in the References section at the conclusion of this report (Section 21) and on experience gained from similar deposits. 2.4 PARTICIPANTS AND PERSONAL SITE INSPECTIONS Details of Qualified Persons and responsibilities are as follows: • Bill Plyley, MAusIMM and Chief Operating Officer for La Mancha, was responsible for compilation of the Technical Report and provided specific information regarding Mineral Reserves, adjacent properties, heap leach processing and current operations, existing infrastructure, markets and sales conditions, G&A costs, project implementation and economic analysis, including provision of CIL processed grades and recoveries. Mr Plyley has visited the property on numerous occasions over the past 4 years. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 31 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report • Graeme Baker, MausIMM and Principal Mining Engineer for AMEC was responsible for those parts of Sections 1, 18, 19, 21 and 22 relating to open pit and underground mining of the Hadal Awatib and Hassai South VMS deposits, with the exclusion of geotechnical aspects which were provided by others. • Dean David, FAusIMM and Process Consultant to AMEC, was responsible for those parts of Sections 1,16, 18, 19, 21 and 22 relating to processing of the VMS mineralisation, including metallurgy, plant design, and capital and operating costs. Mr David visited the property in March 2010. • Adam Coulson, ACSM, CIMM and Senior Rock Mechanics Engineer for AMEC Earth & Environmental, was responsible for those parts of Sections 1, 18, 19, 21 and 22 relating to geotechnical conditions governing open pit and underground mining of the VMS deposits. Mr Coulson visited the property in March 2010. • Ian Thomas, MausIMM and Process Consultant for Sedgman, was responsible for those parts of Sections 1, 16, 18, 19, 21 and 22 relating to processing of gold mineralisation in the proposed CIL plant, including metallurgy, plant and infrastructure design, preliminary capital and operating costs with the exception of the processed grades and recoveries. Mr Thomas visited the property in December 2007. • Remi Bosc, Member European Federation of Geologists and Principal Consultant Arethuse Geology (Malaysia) was responsible for those parts of Sections 1, 14, 17, 19, 21 and 22 relating to data verification and estimation of resources other than those for undrilled heap leach tailings. Mr Bosc has visited the property on numerous occasions, most recently in August 2010. • Simon Mc Cracken, MAIG, Principal Geologist for CSA Global (UK), was responsible for those parts of Sections 1, 17, 19, 21 and 22 relating to estimation of gold resources in heap leach tailings not previously drilled. Mr McCracken visited the property 25-31 August 2010. • Clayton Reeves, MSAIMM and Principal Mine Engineer for CSA Global (UK), was responsible for those parts of Sections 1, 18, 19, 21 and 22 relating to mining of gold resources for the CIL phase of the project. Mr Reeves has spent in excess of seven weeks on site, most recently in September 2010. • Jean-Jacques Kachrillo was responsible for Sections 7 to 13 of the Technical Report, relating to geology, mineralisation and exploration, including sampling and analysis. Other Experts who assisted in providing background information, environmental review, engineering design, cost estimation and cash flow evaluation were : • Dr Abu Fatima, Ph.D. Geology, General Manager, AMC, was responsible for Sections 4, 5, 6 and 18.5 relating to the property description, location, access, climate, history and the hydrological/hydrogeological conditions on site. Dr Abu Fatima is site-based. • Phillip Rogers, B.Sc. (Hons), Ph.D., MIEEM, Environmental Manager for AMEC Earth and Environmental UK Ltd., who visited site to review environmental conditions particularly pertaining to selection of the VMS tailings disposal area. • Phil Payne, Consultant Estimator for AMEC, who compiled the capital cost estimate for the VMS plant. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 32 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report • Pier Chiti Associate Director Mining with AMEC Earth & Environmental provided input for the VMS tailings disposal facility location and design. 2.5 INDEPENDENCE AMEC, Arethuse, Sedgman and CSA are not associates or affiliates of La Mancha, or of any company associated with La Mancha. Fees for this work are not dependent in whole or in part on any prior or future engagement or understanding resulting from the conclusions of this report. These fees are in accordance with standard industry fees for work of this nature. All sections of the Technical Report have a Qualified Person (QP) taking responsibility for preparation or supervising the preparation. Independent QPs have signed off on exploration data quality, Mineral Resources and Mining Inventory, process testwork, plant design, engineering and costings. However, since Ariab is a producing property, AMC QPs have taken responsibility for the overall report and for providing information regarding project background, land ownership and licences, geology and exploration activities, Mineral Reserves, financial analysis and information regarding current operations. La Mancha is a public corporation. Its stock is traded on the Toronto Stock Exchange under the symbol LAM, and its registered office is 2001, Rue University, Bureau 400, Montreal, Quebec, Canada. Note that all costs and prices are quoted in United States Dollars, unless otherwise specified. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 33 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 3. RELIANCE ON OTHER EXPERTS AMEC, Sedgman, CSA and Arethuse (The Consultants) have not reviewed any legal issues regarding the land tenure, surface rights and permits, nor independently verified the legal status or ownership of the Property, and have relied upon opinion supplied by La Mancha in this regard. The Consultants have not attempted to verify the potential to acquire water from the River Nile to supply the expanded project. Similarly, while a preliminary review of environmental conditions has been completed by AMEC, reliance is placed on assurances by La Mancha regarding compliance with all government regulations for current operations. QPs employed by La Mancha take responsibility for a number of areas within this report, notably: • History of the project • Current operations • Geology and mineralisation • Exploration, sampling and analysis • Testwork for heap leach operation • Mineral Reserves • Taxes and royalties • Project economic modelling and evaluation. The results and opinions expressed in this report by the Consultants are conditional upon the aforementioned supplied data and information being current, accurate, and complete as of the date of this report, and the understanding that no information has been withheld that would affect the conclusions made herein. The Consultants do not assume responsibility for La Mancha’s actions in distributing this report other than its filing with security regulators. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 34 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 4. PROPERTY DESCRIPTION AND LOCATION 4.1 LOCATION Sudan is located along the northeast coast of Africa on the Red Sea, and is bordered to the east by Ethiopia and Eritrea, to the north by Egypt, to the northwest by Libya, to the west by Chad and the Central African Republic, and to the south by Uganda, Kenya and the Democratic Republic of Congo. The AMC mining operations are located in a remote area within the Red Sea State of Sudan, approximately 450 km northeast of Khartoum and 200 km west of Port Sudan (Figure 4.1). The area is referred to variously as the Hassaï project or region, or the Ariab mining district. In this report, map coordinates are displayed as latitude and longitude, or in UTM (Universal Transverse Mercator) coordinates from UTM zone 36 North, Adindan datum, Clarke 1880 ellipsoid. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 35 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 4.1 Location of the Hassai Project FINAL – Rev 0 – 22 Oct 2010 AMEC Page 36 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 4.2 MINING CLAIM DESCRIPTION – GOLD AMC is the owner of the exploration and mining concessions in the Ariab district. This ownership is governed by an original Ariab Concession Agreement which, in 1991, transferred all rights of the previous Ariab Mining Development Joint Venture to the newly established Ariab Mining Company. The whole Ariab Mining District is covered by the Reserved Areas as shown in Figure 4.2 and Table 4.1, representing a total surface of more than 20 000 km². A Reserved Area is granted by the Minister of Mines & Energy and confers to the holder an almost exclusive right to carry out exploration and general prospecting for any metal or natural resource in the ground. The right to carry-out detailed prospecting and exploration drilling for specific metals is granted through an Exclusive Prospecting License (EPL), valid for two years. An EPL is a pre-requisite for obtaining a Mining Lease. Figure 4.2 Prospecting Licences BLOCK 11 BLOCK 18 FINAL – Rev 0 – 22 Oct 2010 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 4.1 Coordinates of AMC’s Reserved Areas Name Corner Longitude Latitude Name Corner Longitude Latitude Ariab A1 35° 00 18° 55 Togni T1 34° 55 18° 15 A2 35° 35 18° 45 T2 34° 55 18° 00 A3 35° 35 18° 30 T3 36° 00 18° 00 A4 35° 00 18° 30 T4 36° 00 18° 15 Wadi Amur Musmar W1 36° 00 19° 25 S1 36° 00 19° 25 W2 35° 27 19° 25 S2 36° 17 19° 25 W3 35° 27 19° 06 S3 36° 17 18° 53 W4 34° 55 19° 06 W5 34° 55 18° 30 W6 35° 00 W7 35° 00 W8 35° 35 Shulai S4 36° 00 18° 53 D1 36° 00 17° 50 18° 30 D2 36° 20 17° 50 18° 55 D3 36° 20 17° 25 18° 45 D4 36° 00 17° 25 Derudeb W9 35° 35 18° 30 B1 34° 30 18° 30 W10 36° 00 18° 30 Bahora B2 34° 55 18° 30 M1 34° 55 18° 30 B3 34° 55 18° 00 M2 35° 55 18° 30 B4 34° 30 18° 00 M3 35° 55 18° 15 M4 34° 55 18° 15 All of the AMC deposits in the Ariab mining district are presently covered by Mining Leases (Table 4.2), each valid for gold and associated metals for a duration of 21 years. Additional new concession (18) Derudaib Coordinates: Latitude Longitude A 18° 00' 36° 00' B 18° 00' Sudan – Eritrea Border (37° 30’) C 17° 00' Sudan – Eritrea Border D 17° 00' 36° 45' E 17° 25' 36° 45' F 17° 25' 36° 00' FINAL – Rev 0 – 22 Oct 2010 AMEC Page 38 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 4.3 Sudan Gold and Iron Concessions Map FINAL – Rev 0 – 22 Oct 2010 AMEC Page 39 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 4.2 Coordinates of Mining Leases Name Grant Date No. Longitude Dim Dim 4 20/11/2004 10/2004 Hadal Auatib A Hadal Auatib B Hadal Auatib C Hadayamet A Hadayamet B 01/08/1991 01/08/1991 15/12/1993 03/05/2000 03/05/2000 454 455 25/1993 2/2000 3/2000 Latitude Name Grant Date No. 35°13'40.2” 18°39'52.3'' Oderuk 01/04/1997 1/97 35°13'40.2” 18°39'19.8'' 35°14'14.2” 18°39'52.3'' 35°20'21” 18°39'07'' 35°14'14.2” 18°39'19.8'' 35°20'21” 18°38'34'' 35°26'26” 18°47'13'' 35° 28' 34’’ 18° 33’ 58’’ 35°26'26” 18°46'08'' 35° 28' 34’’ 18° 34’ 15’’ 35°27'00” 18°47'13'' 35° 28' 52’’ 18° 34' 15’’ 35°27'00” 18°46'08'' 35°27'00” 18°47'13'' 35°27'00” Hamim South 03/05/2007 4/2008 Longitude Latitude 35°19'13” 18°39'07'' 35°19'13” 18°38'34'' 35° 28' 52’’ 18° 33’ 58’’ 35° 28' 45’’ 18° 34’ 27’’ 18°46'08'' 35° 28' 45” 18° 34’ 44’’ 35°27'34” 18°47'13'' 35° 29' 03’’ 18° 34' 44’’ 35°27'34” 18°46'08'' 35° 29' 03’’ 18° 34' 27’’ Hamim North Medadip 03/05/2007 03/05/2007 3/2008 35°27'34” 18°46'20'' 35° 24' 14’’ 18° 46’ 54’’ 35°27'34” 18°45'56'' 2/2008 35° 24' 14’’ 18° 47’ 07’’ 35°28'42” 18°46'20'' 35° 24' 57’’ 18° 47' 07’’ 35°28'42” 18°45'56'' 35°35'59” 18°41'16'' 35°36'32,8” 18°41'15.5'' 35° 17' 49.6" 18°36'30.6" 35°36'30,8” 18°40'10.5' 35°18'13.3" 18°36'46.9" 35°18'21.7" 18°36'36.3" 35°35'57” 18°40'11'' 35°36'32,8” 18°41'15.5'' 35°37'07” 18°41'15'' UmAshar Youneim 25/07/2007 03/05/2007 01/2008 05/2008 35° 24' 57’’ 18° 46' 54’’ 35° 17' 58.1" 18)36'19.4" 35° 26' 35" 18° 35' 0" 35° 26' 35" 18° 34' 30" 35°37'05” 18°40'10'' 35° 27' 5" 18° 35' 0" 35°36'30,8” 18°40'10.5' 35° 27' 5" 18° 34' 30" FINAL – Rev 0 – 22 Oct 2010 AMEC Page 40 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 4.2 Coordinates of Mining Leases Name Grant Date No. Longitude Latitude Name Grant Date No. Longitude Latitude Hassai 01/08/1991 176 35°23'09” 18°42'02'' Adasedakh 01/06/1996 14/1996 35° 18'47" 18° 40'00" 35°23'09” 18°41'30'' 35° 19'55" 18° 40'00" 34°24'17” 18°42'02'' 35° 18'47" 18° 38'55" Kamoeb A Kamoeb B 16/03/2004 03/16/2004 1/2004 21/2004 34°24'17” 18°41'30'' 35°22'12” 18°39'26'' 35° 19'55" 18° 38'55" 35° 18'33" 18° 38'34" 35°22'12” 18°38'54'' 35° 19'41" 18° 38'34" 35°23'20” 18°39'26'' 35° 18'33" 18° 38'01" 35°23'20” 18°38'54'' 35° 19'41" 18° 38'01" Baderuk Ganaet 01/06/1996 01/08/1991 15/1996 35°22'12” 18°38'54'' 35°15'29" 18°44'07" 35°22'12” 18°38'21'' 426 35°16'03" 18°44'07" 35°23'20” 18°38'54'' 35°15'29" 18°43'02" 35°23'20” 18°38'21'' 35°16'03" 18°43'02" FINAL – Rev 0 – 22 Oct 2010 AMEC Page 41 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 4.3 MINING CLAIMS – BASE METALS AMC obtained confirmation, by letter from the Geological Research Authority of the Sudan dated 6 December 2007, that AMC is still entitled to the deposits previously discovered by the Ariab Mining Development Joint Venture covered by the previous and present EPLs; and that AMC is entitled to carry out exploration programs for base metals in other areas within the bounds of the original Concession Agreement. 4.4 OWNERSHIP OF MINERAL RIGHTS AMC is the owner of the exploration and mining concessions in the Ariab Mining District. COMINOR is a shareholder of AMC with 40% of the shares. The Government of Sudan owns 56% of the shares and a private French company the remaining 4%. AMC was incorporated as a Sudanese company in September 1990. It is a private company limited by shares. 4.5 MINERAL ROYALTIES The gold operations are subject to the following royalties: • A Net Smelter Return (NSR) of 7% on revenue is payable to the Sudanese Ministry for Geology (GRAS) • A 2.25% Gross Profit tax is payable to La Mancha (via COMINOR) as an incentive fee. These royalties do not cover the possibility of base metal production and any royalties payable would, therefore, be subject to future negotiations in the event of base metal production. 4.6 ENVIRONMENTAL OBLIGATIONS In the Ariab Mining District, several pits have been mined for oxide gold and these resources are now exhausted. These pits have not been backfilled. However, in order to hinder wandering cattle and nomads entering abandoned pits to access water, AMC has constructed safety bunds. In addition, the Company has partially backfilled some of the pits with oxide waste to cover exposed sulphidic rock and prevent the formation of acidic water. A provision has been made by AMC to provide for such limited reclamation costs, pending final decision about the re-opening of the pits for base metal mining. 4.7 RELATIONSHIP BETWEEN AMC AND THE SUDANESE GOVERNMENT AMC maintains a strong working relationship with the Sudanese government. AMC’s Chairman is H.E. Dr. Abdel Bagi El Gailani, who is also the Sudanese Minister of Minerals. The Hassai project is presently the only significant mining venture operating in Sudan. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 42 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 5. ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY 5.1 ACCESS The Ariab district is accessible from Khartoum by the road north to Atbara (paved road in good condition), and then the Atbara-Port Sudan road (Figure 4.1). A gravel track provides access from the Atbara-Port Sudan road to the Hassai mine site (the western access road). From Port Sudan, Hassai can also be accessed via the Atbara-Port Sudan road to Haiya Junction and then a gravel track to the mine site (the eastern access road). The company maintains the local roads in the vicinity of the plant and mine sites. 5.2 PORT FACILITIES Port Sudan is the major regional port and is managed by Sea Ports Corporation (SPC). SPC was established in 1974 as an independent Sudanese maritime body responsible for construction, development and maintenance of ports, harbours and lighthouses. Port Sudan is divided into North Quays, South Quays and Green Harbour. Green Harbour is undergoing development as part of a long term project that consists of establishing extra quays to serve different port handling operations. There are also two other ports located south of Port Sudan; Port Digna (60 km south) and Al Khair Petroleum terminal (3 km south-east). AMC has a 4500 m2 fenced yard at Port Sudan. It is anticipated that a concentrate storage shed measuring 1800 m2 will be erected in this space to house VMS concentrate bags prior to container loading and shipping to customers. Table 5.1 Port Sudan Overview Port Area Data Northern Quay Southern Quay Green Harbour Berths 11 4 2 Length 1663 m 733 m 548 m Depth 8.7 m to 10.7 m 10.7 m to 12.8 m Area Not determined 4000 m 2 Not specified 5.0 Mt/a 3.0 Mt/a Not specified Bulk lime, molasses, edible oils Petroleum, containers, bulk grain Dry bulk cargo, seeds, containers Capacity Current Uses Port superstructure/equipment 14.2 m 2 gantry cranes 4RTG 35 quay cranes Mobile cranes, forklifts, r/or, tractors and trailers available Tugboats of 1600-2000 hp 4 pilot boats of 3600 hp 10 service boats of 180 hp 1 patrol boat of 2000 hp Storage and warehouses 27 warehouses GH storage area 436 000 m2 Harbour accommodates ships up to 50 000 t Operations Pilotage of vessels is compulsory 3 shifts operate on 24 hour basis FINAL – Rev 0 – 22 Oct 2010 AMEC Page 43 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 5.3 CLIMATE The property is in the desert of Northern Sudan where precipitation is infrequent. The climate is arid with a very hot season from June to September during which the maximum temperatures range from 45°C to 55°C and rainstorms may occur. The coolest period covers the months of January and February with daytime temperatures of 30°C and cool nights ranging from 10°C to 15°C. The dominant winds depend on the season: mainly from the west or northwest during the hot period, and from the north or northeast during the rest of the year. Intermittent rainfall can occur during the period from end of July up to October and the number of rain event averages 3-4. The average rainfall is around 30 mm and the highest recording was 80 mm. No evaporation rates are available at this time. 5.4 INFRASTRUCTURE 5.4.1 Buildings and Mine Camp The Hassai mine camp is approximately 3 km from the processing plant and accommodates approximately 600 personnel (expatriates and locals). It includes accommodation, dining halls, a bakery and local market, and recreational facilities. Mine buildings include offices, workshops, power house, etc. The AMC mine site is equipped with a clinic and physician on standby, available for workers and community alike. An emergency plan includes transportation by ambulance or air to the nearest hospital (25 km away). An on-site communication tower allows cellular phone communication through three mobile phone access providers, and internet access. A total of 17 diesel generators (totalling 5470 kVa) supply electricity to the plant and facilities. 5.4.2 Other Offices The head office building in Khartoum houses approximately 40 personnel, including general management, financial control and local purchasing. AMC also has a small office in Port Sudan for approximately ten personnel who are responsible for coordinating sea freight shipments, including the purchasing and transportation of supplies for Hassai (food, equipment, etc.). 5.4.3 Logistics Transportation from the port in Port Sudan to the mine site is carried out by a combination of subcontractors and company-owned trucks. The distance is approximately 200 km, and about 2000 t of consumables are transported each year. Airfreight cargo service into Sudan is provided through Lufthansa, Emirates, Egypt Air and other scheduled flights. A Twin Otter airplane owned by AMC is used for limited transport of personnel. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 44 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 5.5 LAND USAGE The population is semi-nomadic, with people camping around the few wells for most of the year. After the rain, they move toward the grazing areas where they also grow some crops, mainly millet or doura. Camels, sheep, goats and donkeys constitute the herds of domesticated animals in the region. The people belong to the Beja ethnic group, which is subdivided into two main tribes: the Hadendawa, who are concentrated in the mine area, and the Atman. 5.6 PHYSIOGRAPHY AND VEGETATION The region is characterised by chains of hills separated by sandy valleys that collectively form the main basin joining Khor Ariab and Wadi Amur, the latter flowing towards the Nile. Vegetation consists predominantly of sparse thorny shrubs and dry grasses in the valleys. Grasses cover the valleys for several months after heavy rains, serving as grazing grounds for sheep, goats and camels. 5.7 SURFACE AND GROUNDWATER Due to the extremely arid desert conditions, water resources in the district are scarce. Water for current operations is sourced from various points including: • Fresh as well as saline water sourced from a series of wells located at distances up to 100 km from the Hassai plant. • Basins (hafirs) protected by earth dams have been dug to store run-off rainwater; these basins now have a total capacity of over 340 000 m3. • Recycled sewage water has also been used in the leach process since 1996. The basement geology in the Hassai region consists of granite and volcanic rocks, and no sizable underground aquifers are known. Significant rainfall in recent years has greatly increased the water reserves. AMC estimates that current water reserves accessed by existing wells are sufficient to sustain production for at least two more years without any additional precipitation. The River Nile, 165 km from site, constitutes the only major water source in this part of Sudan. It is intended to acquire water rights and construct a pipeline to supply the significantly increased needs of the proposed CIL and VMS concentrator plants. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 45 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 6. HISTORY 6.1 PRE-COMINOR From 1977 to 1981, French-Sudanese teams conducted several exploration programs for various metals (W, Pb, Zn, Cu, Ag, Au, Cr, etc.) over five large areas in Sudan within the framework of a cooperation agreement. One such exploration program focused on 17 gossans in the central part of the Red Sea Hills (Ariab-Arbaat area). These gossans are the weathering products of massive sulphide deposits found at depth. The gold in most of the Ariab mining district is associated with these gossans; the exceptions are Kamoeb and Ganaet gold deposits, although these were also explored during the French-Sudanese programs. 6.2 COMINOR In 1981, a joint venture agreement was signed between BRGM and the Sudanese government to cover the mining development of three specific areas: Eyob (50 km south of the Ariab district), Ariab and Hamissana (approximately 200 km north of the Ariab area). Detailed exploration work was performed over these three areas from 1981 to 1984. Exploration covered most of the gossans in the Ariab district, although initial work focused on the polymetallic potential of the underlying sulphides. In 1983, however, the discovery of noteworthy gold concentrations in silica-kaolinite-(barite) rocks associated with the gossan at Hassai shifted the interest towards gold. During the same period, the Ganaet and Kamoeb gold deposits were also explored in detail. Groundwater exploration also began in 1983. From 1984 to 1987, exploration efforts were concentrated on the Ariab district. Major trenching work was carried out on all known gossans in the region: Hadal Awatib SW, Hadal Awatib E, Talaiderut, Oderuk, Baderuk and Adassedakh. In February 1985, sufficient data from surface trenching, percussion drilling, core logging and pits were collected at Hassai to justify the installation of a pilot plant. The first gold was poured at Hassai in March 1987. Delays in the negotiations between the parties involved in the joint venture caused work to be suspended from April 1987 to February 1988. Gold production and exploration work was reactivated in April 1988 and the pilot operation program was satisfactorily completed in December 1989. A feasibility report for the Ariab Gold Project performed by BRGM was submitted in May 1990. The report examined 10 known gold deposits in the district within a circular area measuring 25 km in diameter. In eight of these deposits (Adassedakh, Baderuk, Hadal Awatib West and East, Hassai South and North, Oderuk, Talaiderut), gold is associated with silica-kaolinite-barite rock and ferruginous gossans, which are in turn the near-surface expressions of underlying massive-sulphide mineralisation. At Kamoeb, gold is present in quartz veins, and at Ganaet it is associated with barite lenses. Mine production began in 1991 and has yielded over 2.3 Moz of gold to date from a large number of deposits (Figure 6.1). The following oxide deposits are now considered to be exhausted: Adassedakh, Baderuk, Baderuk N, Dim Dim 4, Dim Dim 5, Hadal Awatib E, Hadal Awatib W, Hadal Awatib N, Oderuk and Talaiderut Oderuk W, while those being mined at the end of 2009 were Hassai North, Hadal Awatib Link and Kamoeb. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 46 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 6.1 Hassai Project – Location of Mines and Prospects At present, gold from the oxidised part of massive sulphides is nearly depleted, and the existing pits are floored by massive sulphide. As of June 2010, most of the mining reserves (excluding stockpiles) are in two deposits: Kamoeb South and Hadal Awatib Link. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 47 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 7. GEOLOGICAL SETTING 7.1 REGIONAL GEOLOGY The regional geology has been described in previous NI 43-101 Technical Reports filed by La Mancha (eg La Mancha Resources Inc., December 2009). In summary, the Hassai project deposits lie within granite-greenstone terrane of the Arabian-Nubian Shield of Neoproterozoic age. 7.2 LOCAL GEOLOGY The oxidised VMS and quartz vein gold deposits and the VMS mineralisation of the Ariab mining district are within the Neoproterozoic Ariab greenstone belt. The host rocks comprise bimodal volcanic, volcaniclastic and siliciclastic strata and late- to post-tectonic granites. Most of the VMS deposits are present within specific stratigraphic units, commonly altered felsic tuffs. A more detailed description of the local geology, structure and age of the mineralisation is provided in La Mancha Resources Inc., December 2009. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 48 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 8. DEPOSIT TYPES 8.1 INTRODUCTION Information regarding deposit types is quoted below in italics from La Mancha Resources Inc., December 2009. “The Ariab area is rich in gold and base metals deposits. The three types of gold deposit are summarized in Section 8.2, while base metal-rich VMS deposits are described in Section 8.3. 8.2 GOLD DEPOSITS 8.2.1 Oxide and Quartz-Kaolinite-Barite (“SBR”) Gold Deposits “Oxide and quartz-kaolinite-barite gold deposits are the main type of gold mineralisation in the region and are characterized by gold enrichment in gossans and “silica-barite rocks” (“SBR”), both of which are the weathering products of underlying polymetallic massive sulphide deposits. The massive sulphides are volcanogenic in nature and are part of the Ariab Proterozoic greenstone belt. More specifically, most deposits are found within Unit D, the upper member of the differentiated volcanic sequence in the Ariab series. The VMS mineralisation is described in more detail in the next section. Figure 8.1 is a diagrammatic representation of a typical oxide-sulfate gold deposit in the Ariab area, and the relationship with underlying copper-zinc-gold VMS mineralisation. Oxide-sulfate gold deposits represent the main source of ore at Hassai since mining commenced. Only small resources of this type remain and some of these are currently being mined. Other lesser oxidesulfate gold deposits require additional exploration. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 49 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 8.1 Diagrammatic Cross-section Showing Relationship of Ariab Deposits FINAL – Rev 0 – 22 Oct 2010 AMEC Page 50 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 8.2.2 Gold-bearing Quartz Veins The mineralisation at Kamoeb consists of gold-bearing quartz veins, veinlets, stockworks and vein selvages distributed as shown in Figure 8.2. Timing of the gold mineralisation remains enigmatic. Lead isotopic dating of the main quartz veins at Kamoeb suggest they are early and related to the emplacement of the adjacent basement granite with or without gold mineralisation. These veins were subsequently deformed resulting in either remobilisation of the gold or a separate gold mineralising event. The mineralisation displays affinities with mesothermal gold deposits, sharing key geological features such as gold occurrence in moderately deformed quartz veins hosted by metamorphic rocks in a greenstone belt. After mining, it appears that at least some of the gold mineralisation is hosted in deformed wall rocks around the quartz veins. Figure 8.2 Kamoeb Geology Map Aplitic dyke Simplified geological map of Kamoeb quartz veins, labels as follows: KS1, Kamoeb south vein 1; KS2, Kamoeb south vein 2; KE, Kamoeb east; KN, Kamoeb north; KW, Kamoeb west. (Fatima, 2006) FINAL – Rev 0 – 22 Oct 2010 AMEC Page 51 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 8.2.3 Gold-rich Barite Lenses Without Proximal Gossan Development The Ganaet deposit is one of several small mineralised bodies that contain gold in baritic lenses. The Rawai and Hamim prospects have similarities to Ganaet. The mineralisation shares some features with SBR deposits, but lacks any spatial association with underlying VMS mineralisation and is not marked by well-developed gossan zones. It may represent the distal facies associated with VMS mineralisation. Following mining in 2007, it appeared that the barite lenses did not extend at depth, and that gold mineralisation was found in highly deformed corridors, with mylonite fabrics and green chloritic alteration. Gold is also contained in late brittle faults parallel with the fabric and with an apparent, association with magnetite. This strongly suggests shear zone type mineralisation that may have reworked original barite-hosted primary mineralisation. 8.3 VOLCANOGENIC CU-ZN-AU-AG MASSIVE SULPHIDE DEPOSITS Volcanogenic massive sulphide deposits can be classified into five types based on host rock compositions (Barrie and Hannington, 1999). From the most primitive to the most evolved in a chemical sense, the five host rock compositions considered are: mafic, bimodal-mafic, maficsiliciclastic, bimodal-felsic, and bimodal-siliciclastic. The mafic-siliciclastic VMS type has sub-equal proportions of mafic volcanic or intrusive rocks and turbiditic siliciclastic rocks; felsic volcanic rocks are minor or absent. There may be significant amounts of carbonate within the siliciclastic rocks, but the siliciclastic component always predominates. They are principally of Middle Proterozoic age and younger, and they are commonly complexly deformed. The deposits of Japan and the Windy Craggy deposit of British Columbia, Canada are type examples on land. The rifted continental margin in the Guaymas basin of the Gulf of California, the sedimented oceanic rift of Middle Valley and the Escanaba trough in the NE Pacific ocean, and the Atlantis II deeps of the Red Sea provide three distinct tectonic settings as analogs for the land-based deposits. Maficsiliciclastic VMS deposits are less numerous than most of the other types, but their average tonnage (average of 11.0 MT) is second only to the bimodal-siliciclastic VMS type (Barrie and Hannington, 1999). The VMS deposits of the Ariab mining district are classified as bimodal-siliciclastic type, similar to many of the large deposits in the Iberian Pyrite Belt, and to the Bisha VMS deposit in western Eritrea. It is important to note comparisons with other VMS globally that have similar characteristics to the large VMS deposits of the Ariab district. These deposits are all large, commonly with copper-rich bases4, and with layered, relatively zinc-rich tops. In addition, the deposits may have relatively barren pyritic central mid-sections.” 4 Stratigraphic footwall. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 52 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 9. MINERALISATION Descriptions of the mineralisation are included in previous Technical Reports (eg. La Mancha Resources Inc., January 2008), and covers remnant gold vein mineralisation at Kamoeb, SBR mineralisation at Hassai South, Hassai North and Yonim, and VMS deposits carrying copper, zinc and gold. 9.1 BASE METAL MASSIVE SULPHIDE DEPOSITS Exploration activities have intersected massive sulphides at 13 localities. The most interesting deposits in terms of economic potential are: • Hassai South: a single lens some 1000 m long and 5-20 m wide, dipping 60o S and extending to at least 400 m. Geophysical surveys indicate possible extensions to over 700 m. The upper, oxide portion has been mined as gold-bearing SBR ore, with supergene ore grading over 5% copper at the base of the pit. Published resources are 20.5 Mt at 1.49% Cu and 1.56 g/t Au, all of which is Inferred (La Mancha Resources Inc., October 2009). • Hadal Awatib: the largest of the known deposits, although apparently broken into multiple lenses at Hadal Awatib East, West and North, all of which have supported mining of gold ore from the oxide zone, for total production in excess of 1 Moz. The total strike length exceeds 2500 m, with widths of up to 100 m. Depth extensions have been drill-tested to 500 m with geophysical signatures down to at least 800 m. Dip is sub-vertical. Published sulphide resources for Hadal Awatib East (La Mancha Resources Inc., December 2009) are: − Indicated: 2.9 Mt at 1.27% Cu, 0.93 g/t Au − Inferred: 28.3 Mt @ 0.99% Cu, 1.18 g/t Au. Other deposits such as Hadayamet, Taladeirut, Adassedekh, Oderuk and Onur have undergone minimal drill testing at this stage, but interpretation of geophysical data suggests strike lengths of a few hundred metres, thicknesses of 15-40 m and good depth extent in all cases. Drilling has indicated supergene and primary sulphide mineralisation with a range of Cu/Zn ratios and generally minor gold in primary sulphides. 9.2 GOLD DEPOSITS 9.2.1 Supergene (SBR) Deposits Overlying VMS Mineralisation Weathering of VMS deposits has produced supergene gold-bearing gossans and silica-barite mineralisation. These deposits have typical strike lengths in the order of 700-3000 m and are 3-50 m thick. The bulk of this mineralisation has been mined, although mining continues on parts of Hadal Awatib Link. 9.2.2 Quartz Veins Kamoeb is the prime example of this style of mineralisation. Gold is associated with quartz veins, veinlets and vein selvages in brittle-ductile deformation settings. Kamoeb comprises four zones (Kamoeb North, South East and West, Figure 8.2) ranging from 1-10 m thick and with a cumulative strike length of more than 4 km. Each vein system forms the core of a hilly landform rising 50-100 m above the surrounds. The vein systems extend down-dip for at least 150 m in Kamoeb South. Ore is massive grey, pink and white quartz, with veins sheared and anastomosing along strike. Gold is fine– grained, although clusters up to 150 µm are developed at times. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 53 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 10. EXPLORATION 10.1 EXPLORATION METHODS Modern exploration activities in the area began in 1977. A non-exhaustive list of the techniques that led to the discovery and definition of actively mined orebodies and other deposits includes: • Reconnaissance prospecting and mapping at various scales (1:50 000 to 1:500) • Stream-sediment geochemistry • Follow-up multi-element lithogeochemistry • Reconnaissance percussion drilling • Airborne geophysical survey (VTEM 2007) • Ground geophysics (mainly spontaneous polarisation with minor gravimetric and EM surveys) • Landsat imagery analysis (1997) • Trenching • Core drilling • Reverse circulation (RC) drilling. Surface mapping and sampling identified a large number of visually and topographically distinct gossans developed over massive sulphides, leading directly to the discovery of the majority of the gold deposits. Mapping and sampling also identified gold-mineralised quartz veins and SBR-type mineralisation. Much of the previous work has been directed towards location of gold mineralisation to maintain feed to the current heap leach operation, although identified deposits are now largely exhausted. More recently there has been a focus on base metal VMS mineralisation, commencing with a 11 415 line-km helicopter-borne VTEM survey flown by Geotech Airborne Limited in 2007 (La Mancha Resources Inc., October 2009). This survey located all previously known VMS bodies, and provided additional information regarding dip and depth extensions to some deposits. In addition, a large number of targets were identified for future evaluation. Drill testing of VMS mineralisation has been undertaken to outline resources in two areas, Hassai South and Hadal Awatib East, this work being completed largely in 2008 and 2009. Drilling included short holes drilled within existing pits to test supergene-enriched upper parts of the deposits, and deeper angled holes from surface to define the primary massive sulphide lenses. All geological exploration data has been compiled into a geographic information system (GIS). FINAL – Rev 0 – 22 Oct 2010 AMEC Page 54 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 10.2 SURVEYING Areas tested by trenches and drilling have been surveyed by AMC's surveyors who have experience in the area dating back more than 15 years. Various types of theodolites, total station and a Leica Differential GPS (GPS station SR530 RTK) have been used over time, allowing the production of good quality topographic maps at scales of 1:1000 and 1:500. Plans are drafted using Microstation software. All surveys are referenced to the UTM system (Datum Adindan). Several triangulation benchmarks are available on site with precise coordinates surveyed by the Sudan Survey Department. All data are digitised using the UTM zone 36 North - Datum Adindan (Clarke 1880) projection. The Digital Elevation Model (DEM) was transferred in 3D-DXF format and uploaded to Surpac; a few minor corrections were made (minor cross-over line issues), but no major issue was detected. All the drill holes were surveyed by professional AMC survey teams using a Leica Differential GPS. Drill collar location files were transferred in a spreadsheet to the geological department prior to database upload. The collars are globally consistent with the DEM. 10.3 MAIN RESULTS The work described in this sub-section was carried out by AMC or by contractors under AMC control. For example the geophysical survey was subcontracted to Geotech and drilling was subcontracted to Longyear and GED. AMC’s activities extend over many years, and gold exploration is not described in detail, since most of the mineralisation has been mined. The reader can refer to three technical reports with additional information: • Technical Report February 2008 • Technical Report Hassaï Resources 2009 • Technical report Hadal Awatib resources 2009 When a drill hole is described, only apparent width is reported. The true width is taken in account in reserves and resources calculation. 10.3.1 VMS – Prior to 2007 From the early 1980’s to the opening of the Hassaï mine in 1992, some drill holes designed to test the SBR gold resources also intersected VMS mineralisation. The main VMS intersections reported during this period are included in Appendix 1. These results were generated in the early phases of exploration in the Ariab area when there was no particular focus on gold or base metals. The intersections of massive sulphides were generally complete and give a good indication of the potential grade and width of massive sulphides lenses. Later, when exploration focused on gold, numerous drill holes intersected massive sulphides, but did not drill through them. Although these holes do not give a good indication of the width of massive sulphide lenses, they provide information on the grade, particularly in the upper part of the lenses where enrichment is likely (eg at Hassai where drill hole HASS 053 intersected 10 m at 5% Cu and 2 g/t Au). Because the VMS was not the main goal of the drill program, these results were not followed-up FINAL – Rev 0 – 22 Oct 2010 AMEC Page 55 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report by infill holes and very often the results were partial and concerned only a part of massive sulphide body. However, they were sufficient to demonstrate the VMS potential of the area. 10.3.2 2007 VTEM Geophysical Survey A helicopter-borne Time Domain electromagnetic survey was flown between February 16th and April 11th, 2007. Characteristics of this survey are listed below: • Contractor : Geotech Airborne Limited • Methods : • − Electromagnetics (Time domain EM) − Magnetics (Total magnetic intensity) − Radar altimeter + Differential GPS Total line: 11415 km (5 blocks) − Block 13 : Hassai = 4762 km − Block 2 : Hadayamet = 1385 km − Block 4 : Zahateb = 1479 km − Block 1 Extension : Youneim = 2326 km − Block 2 Extension : Mandilu = 1348 km The method is summarised in Figure 10.1. Figure 10.1 VTEM Geophysical Survey Basis The types of VTEM responses from electro-magnetic conductors are shown in Figure 10.2. The results of the survey are very encouraging. Not only were all the known massive sulphides identified by the survey, but, thanks to the low and slow flight of the helicopter (20 m/s and 10 m/ s at an average 85 m above ground), additional details such as the dip of some conductors was recorded with unexpected accuracy. Moreover, at places such as Hadal Awatib and Hadayamet the depth extension of massive sulphide was confirmed. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 56 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 10.2 VTEM Response Types VTEM results from Hadal Awatib illustrate the accuracy and efficiency of the method. Figure 10.3 shows the VTEM response of the 2.2 km long Hadal Awatib deposit. The strong VTEM response from central part of Hadal Awatib that has not been drill tested should be noted. The numerous anomalies have been listed and classified as follows: they constitute targets for current and future exploration activities. • • Block North − Oderuk extension Priority 1 − Hadal Awatib Priority 1 − Baderuk extension Priority 1 − Younim East Priority 1 − Ganaet East Priority 1 − Rahadab Priority 1 − Shidimann West Priority 1 − Medadip South Priority 1 − Hadayamet West Priority 1 − Tidityu Priority 1 − Anomalies A to I Priority 1 − Joseph Priority 1 − Ganaet South Priority 2 − Abukurunt Priority 2 − Mandilu East, West and North Priority 2 − Eikidi Priority 2 − Ientai Priority 2 − Tedmi Priority 3 Block South (Zahateb) − Zahateb FINAL – Rev 0 – 22 Oct 2010 Priority 1 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 10.3 VTEM Geophysical Anomaly at Hadal Awatib (workshops) H adal Awatib Pipe U nexplor ed ar ea s o m a lie ctive a n Co n d u g n me n t li a Drilled boreholes, collars and projected traces Red-orange dots: planned boreholes H adal Awatib W est FINAL – Rev 0 – 22 Oct 2010 H adal Awatib N or th Black crosses: top of conductive anomalies, digitised on geophysical sections H adal Awatib East blocks A+ B H adal Awatib East blocks C+D The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 11. DRILLING 11.1 INTRODUCTION Numerous drilling campaigns have been undertaken since the 1980s, involving a series of contractors, using a variety of percussion, RC and diamond core drill rigs. The dominant method for resource definition (particularly for remaining resources) has been diamond drilling. Boart Longyear was the principal contractor between 1993 and 2006, completing 2172 holes totalling nearly 170 km of core in the AMC Reserve Area. Since 2007, RC drilling has been used in all near-surface targets (oxide zone), while core drilling, with RC pre-collar, was used for deeper investigations (VMS deposits). Much of the early work focussed on shallow gold mineralisation in oxidised VMS (SBR deposits, now largely mined out), or quartz veins (Kamoeb). A number of these drill holes intersected massive sulphides beneath oxide gold mineralisation, although frequently they did not penetrate the full thickness of the lens as VMS mineralisation was not the target. For example, at Hadal Awatib East, of 155 diamond drill holes completed in the period 1996 to 2002, only 9 holes penetrated into primary sulphides. 11.2 DRILLING: 1993-2006 A Boart Longyear LY44 operated by Boart Longyear completed the bulk of the resource holes through to 2006. Drilling and sampling were supervised by professional geological staff from AMC. Core diameter was PQ or HQ. Core lengths varied from 0.5 to 1.5 m, averaging 1 m. No core remains from this work: whole cores of mineralised intervals were analysed and barren intervals discarded. Little or no information is available regarding down-hole survey data from the period 1996 to 2004, although very few holes from this period provide information for remaining (unmined) resources. Holes from the period 2005 to 2006 were surveyed at 50 m intervals down hole; little more information is available, but, again, these holes provide little information relevant to the VMS and remaining gold resources. 11.3 RC AND CORE DRILLING: 2008/09 Drilling in 2008 to 2009 was undertaken by General Exploration Drilling Ltd (GED), using three different drill rigs: • G & K 850 track-mounted multipurpose drill rig • KL400 track-mounted multipurpose drill rig • Drilltech DK40 RC drill rig. The RC drill rigs were used mostly for small SBR targets in the oxide zone, as well as for Kamoeb in-fill drilling. Core drilling (or RC with diamond tails) was used mostly to define the VMS lenses at depth (Hassai South and Hadal Awatib East), and to investigate the overlying supergene-enriched zone. RC pre-collars were drilled for the diamond holes either with the DK40 or the multipurpose rig in RC configuration. The pre-collar hole was then cased and diamond core drilling commenced in NQ diameter. Where drilling conditions permitted, core runs were in 3 m lengths. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 59 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Drill hole deviation measurements were taken on a proportion of the non-vertical core holes, at variable intervals of up to 50 m, using an Eastman single shot camera or EZ-Reflex single shot probe. In the latter case, readings were hand-recorded onto a paper log and digitised in the office. Readings were adjusted from magnetic to true north by addition of 1°. The diamond core was then re-oriented and a line drawn perpendicular to the up-hole intersection with the foliation as a guide for cutting. Geological logging was then conducted including the recording of other technical data such as core recovery, core diameter, density and RQD. All technical data collected was then entered into an Excel spreadsheet and subsequently uploaded into the database. A hard copy version of the geological logs is printed and filed at Hassai Mine. In addition to the RC pre-collars, RC drilling was also used to drill some holes into the VMS from the base of pits. Drill cuttings from each hole were collected from a cyclone in a plastic bag at 1 m intervals. The weight of the sample was recorded as an indication of the recovery, and a geological log of the rock type and any indications of alteration and mineralisation recorded. All RC sample rejects and remaining half cores are stored at Hassai. 11.3.1 Hassai South Drilling A total of 155 diamond drill holes (13 839 m) was drilled between 1996 and 2002 with the Boart Longyear team (Monthel, et al. 2007) on a nominal grid of 25 x 50 m. This drilling was mostly focused on the oxide zone (SBR mineralisation). Only nine drill holes before 2002 penetrated to the sulphide zone. In 2008/09, a total of 51 holes (10 982 m) was completed at Hassai South to provide coverage at approximately 100 m along strike and down-dip in the sulphide zone (supergene and primary domains, Figure 11.1). FINAL – Rev 0 – 22 Oct 2010 AMEC Page 60 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 11.1 Hassai South Drill Hole Location Plan (AMC, 2009 – 200x100 m Grid) HASSAI SOUTH VMS AND DRILL HOLES TRACES 1/2000 2068700. 2068700. 2068800. 2068800. N 1891.0 1991.0 2091.0 2191.0 2391.0 2291.0 2591.0 2491.0 2691.0 2791.0 HASS_D069 HASS_D161 HASS_D068 2991.0 HASS_D177 HASS_D178 HASS_D073 HASS_D074 HASS_D162 HASS_D072 HASS_D155 HASS_D154 HASS_D152 HASS_D153 HASS_D179 HASS_D075 HASS_D076 HASS_D163 HASS_D067 HASS_D037 HASS_D038 HASS_D207 HASS_D040 HASS_D041 HASS_D159 HASS_D039 HASS_D157 HASS_D065 HASS_D066 HASS_D160 HASS_D048 HASS_D064 HASS_D206 HASS_D049 HASS_D158 2068500. HASS_D053 HASS_D054 HASS_D063 HASS_D042 HASS_D043 HASS_D205 HASS_D151 HASS_D045 HASS_D044 HASS_D204 HASS_D168 HASS_D169 HASS_D050 HASS_D055 HASS_D056 HASS_D150 HASS_D051 HASS_D057 HASS_D149 HASS_D046 HASS_D047 HASS_D170 HASS_D203 HASS_D052 HASS_D202 HASS_D090 HASS_D201 HASS_D091 HASS_D092 HASS_D060 HASS_D061 HASS_D062 HASS_D116 HASS_D115 HASS_D089 HASS_D131 HASS_D093 HASS_D132 HASS_D110 HASS_D099 HASS_D094 HASS_D100 HASS_D167 HASS_D127 HASS_D128 HASS_D140 HASS_D164 HASS_D141 HASS_D136 HASS_D137 HASS_D200 HASS_D194 HASS_D197 HASS_D196 HASS_D199 HASS_D095 HASS_D096 HASS_D165 HASS_D166 HASS_D097 HASS_D195 HASS_D098 HASS_D198 HASS_D101 HASS_D102 HASS_D133 HASS_D134 HASS_D138 HASS_D011 HASS_D139 HASS_D118 HASS_D113 HASS_D119 HASS_D114 HASS_D117 HASS_D148 HASS_D103 HASS_D104 HASS_D107 HASS_D108 HASS_D135 HASS_D208 HASS_D190 HASS_D012 HASS_D211 HASS_D077 HASS_D078 HASS_D146 HASS_D121 HASS_D122 HASS_D079 HASS_D080 HASS_D144 HASS_D145 HASS_D129 HASS_D130 HASS_D106 HASS_D105 HASS_D059 HASS_D058 HASS_D209 B.L. HASS_D125 HASS_D123 HASS_D126 HASS_D124 HASS_D081 HASS_D082 HASS_D083 HASS_D084 HASS_D085 HASS_D173 HASS_D171 HASS_D174 HASS_D175 HASS_D176 HASS_D142 HASS_D143 HASS_D111 HASS_D112 HASS_D087 HASS_D088 HASS_D147 2068600. HASS_D070 HASS_D172 2068500. HASS_D071 HASS_D156 HASS_D109 HASS_D120 HASS_D086 HASS_D014 HASS_D210 HASS_D189 HASS_D213 HASS_D212 HASS_D188 HASS_D225 HASS_D180 HASS_D001 HASS_D187 HASS_D186 2068400. HASS_D002 HASS_D220 HASS_D219 HASS_D184 HASS_D226 HASS_D230 HASS_D193 HASS_D028 HASS_D004 HASS_D228 HASS_D229 HASS_D192 HASS_D227 HASS_D216 HASS_D215 HASS_D185 2068100. 2068100. 2068200. 2068200. 2068300. 2068300. HASS_D214 HASS_D191 HASS_D222 HASS_D182 HASS_D183 HASS_D221 HASS_D224 HASS_D181 HASS_D223 2068400. 2068600. 2841.0 B.L. HASS_D217 HASS_D218 751600 751800 752000 752200 0 752400 752600 752800 753000 200 M The results of the drill holes completed in 2008 and 2009 from the southern edge of the pits are included in Appendix 1 as are those completed from the floor of the pit. The relationship between true width and observed width is variable, but generally the true width represents at least more than the half of the observed width, except for the holes drilled from the floor of the pits, as illustrated in Figure 11.2. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 61 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 11.2 Cross-section Through Hasai South Showing Relationship Between Intersected and True Thickness FINAL – Rev 0 – 22 Oct 2010 AMEC Page 62 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 11.3.2 Hadal Awatib East Drilling A total of 236 diamond drill holes (22 337 m) were drilled prior to 2000, focussing mostly on delimitation of the oxide zone. Of these drill holes, only eight targeted the VMS mineralisation, while an additional 29 went through the oxide zone and into the sulphide zone. From 2004 to 2006 two successive diamond drill campaigns were completed to define remaining resources at the Link Zone (oxide zone), currently being mined. Only four of these holes went into the sulphide domain. In 2008-09, 63 holes (7864 m) were completed into the VMS deposit, mainly from the base of the pit. The drill pattern is irregular, approximating 40 x 20 m in shallower parts, widening to 50 x 200 m at depth (Figure 11.3). Figure 11.3 Hadal Awatib East Drill Hole Location Plan (AMC, 2009 – 100x50 m Grid) HADAL AWATIB EAST A/B & C/D/E - VMS General Map 1/2000 78250 N 2078250 78250 8110 8160 8180 8330 8225 8440 8260 2078200 78200 8305 HAE_D109 HAE_R268 HAE_R275 HAE_R276 HAE_D108 HAE_D107 HAE_D323 HAE_D005B HAE_D005 78100 2078100 HAE_R260 HAE_R261 HAE_R262 78050 2078050 HAE_R265 HAE_R266 HAE_R267 HAE_R279 HAE_R271 2078000 8940 HAE_D076 HAE_D077 HAE_D243 HAE_D172 HAE_D173 HAE_D174 HAE_D206 HAE_D215 HAE_D185 HAE_D207 HAE_D183 HAE_D208 HAE_D184 HAE_D180 HAE_D181 HAE_D209 HAE_D182 HAE_D233 HAE_D234 HAE_D235 8160 77950 HAE_D121 HAE_D250 8180 2077950 HAE_D122 HAE_D112 HAE_D318 8225 8260 HAE_D248 HAE_D119 HAE_D118 HAE_D211 HAE_D249 HAE_D252 HAE_D236 HAE_D240 HAE_D111 HAE_D115 HAE_D116 HAE_D113 HAE_D114 HAE Link HAE_D258 Block C+D HAE_D253 8440 77850 8330 2077850 8615 758600 758500 758400 758300 758200 758100 8505 8870 HAE_D257 HAE_D221 HAE_D220 HAE_D230 HAE_D231 HAE_D191 HAE_D190 HAE_D159 HAE_D160 8800 HAE_D224 9045 58200 58300 58400 58500 58600 58700 0 58800 58900 59000 77850 HAE_D225 HAE_D226 HAE_R255 HAE_R256 HAE_D192 9140 58100 77900 AA' HAE_D222 HAE_D223 8940 8700 77950 HAE_D218 HAE_D219 HAE_D228 HAE_D229 HAE_D254 758900 HAE_D251 758800 Block A+B 78000 HAE_D156 HAE_D153 758700 2077900 78050 HAE_R310 HAE_D138 HAE_D163HAE_D139 HAE_D311 HAE_D128 HAE_D126 HAE_D129 HAE_D127 HAE_D124 HAE_D136 HAE_D137 HAE_D313 HAE_D125 HAE_D130 HAE_D312 HAE_D131 HAE_D179 HAE_D149 HAE_D210 HAE_D314 HAE_D161 HAE_D162 HAE_D150 HAE_D315 HAE_D157 HAE_D158 HAE_D151 HAE_D152 HAE_D132 HAE_D133 HAE_D227 HAE_D134 HAE_D135 HAE_D216 HAE_D154 HAE_D217 BB' HAE_D155 8305 77900 9170 9140 HAE_D145 HAE_D146 HAE_D144 HAE_R304 BB' HAE_D193 HAE_D194 HAE_D195 HAE_D142 HAE_R306 HAE_D143 HAE_R305 HAE_D196 HAE_D197 HAE_D170 HAE_D171 HAE_D198 HAE_D147 HAE_R308 HAE_D164 HAE_D199 HAE_D168 HAE_D140 HAE_D200 HAE_D148 HAE_D165 HAE_R307 HAE_D169 HAE_D141 HAE_D259 HAE_D201 HAE_D202 HAE_D166 HAE_D203 HAE_D167 HAE_D178 HAE_R309 HAE_D204 HAE_D176 HAE_D205 HAE_D186 HAE_D175 HAE_D177 HAE_D237 HAE_D239 HAE_D238 HAE_D321 HAE_D319 HAE_D214 HAE_D187 HAE_D188 HAE_D189 HAE_D241 HAE_D242 HAE_D084 HAE_D082 HAE_D083 HAE_D085 HAE_D120 HAE_D282 9045 HAE_D213 HAE_D212 HAE_D247 HAE_D246 HAE_D081 HAE_D080HAE_D244 HAE_D245 HAE_D325 HAE_D283 78100 HAE_D078 HAE_D079 HAE_D320 8110 78150 HAE_D043 HAE_D073 HAE_D074 HAE_D086 HAE_R297 HAE_R281 8800 8700 8870 HAE_D075 HAE_R295 HAE_R280 HAE_R272 HAE_R273 HAE_D324 HAE_R296 HAE_D317 HAE_R274 78000 HAE_D044 HAE_D096 HAE_D097 HAE_R278 HAE_R264 8615 HAE_R293 HAE_D095 HAE_D042 HAE_R294 HAE_R277 HAE_R269 HAE_R263 HAE_D110 HAE_D316 HAE_R270 8505 HAE_R286 HAE_R287 59100 9170 HAE_D232 59200 759300 AA' 759200 2078150 759100 78150 759000 78200 59300 100 M The results obtained from holes drilled on the southern edge of the pits are included in Appendix 1. The relationship between true width and observed width is variable, but generally the true width represents more than the half of the observed width, except for the holes drilled from the floor of the pits, as illustrated in Figure 11.4. FINAL – Rev 0 – 22 Oct 2010 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 11.4 Hadal Awatib – Relationship Between Intersected and True Width 11.3.3 Kamoeb Drilling The Kamoeb area was first drilled in 2001. Up to 2004, 289 diamond holes were completed for 21 351 m. Kamoeb South was drilled every 50 m, whereas Kamoeb West and North were drilled at approximately every 100 m, on an irregular pattern. In 2008/09, 136 RC holes and 2 RC holes with diamond tails were completed on the Kamoeb group, totalling 7109 m. In total 466 drill holes totalling 36 676 m and 20 111 assays are recorded in the database and were used for the delimitation of the Kamoeb group gold resources (Figure 11.5) as follows: • Kamoeb South and East: 340 drill holes, 28 656 m and 17 276 assays • Kamoeb West: 39 drill holes, 2588 m and 432 assays • Kamoeb North: 87 drill-holes, 5432 m and 2403 assays FINAL – Rev 0 – 22 Oct 2010 AMEC Page 64 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 11.5 Kamoeb Drill Hole Location and Topographic Plan - 250 x 250m grid (UTM coordinates 36N, Adindan Datum) Kamoeb North Diamond drilling RC drilling RC with Diamond tail Kamoeb North Kamoeb East Kamoeb West Kamoeb South 11.4 HEAP LEACH RESIDUE DRILLING In 2007/2008, and again in 2009, AMC undertook auger drilling of a number of heap leach pads where leaching had been completed. These pads were designated A to D, and lie in the vicinity of the Hassai Mine and process plant. Drilling was conducted by Dump & Dune using a cased auger. The then-active dumps were not drilled. The drill pattern was typically 25x20 m. A total of 606 holes were completed, for 6509.5 m (4419 samples). Drill collars and the heaps themselves were surveyed by the mine survey team using Garmin DGPS and total station equipment. Samples were collected over 1.5 m intervals and brought to the core-yard for sample splitting. Sample weights were recorded for each interval during the 2007/2008 program, and recovery was determined. FINAL – Rev 0 – 22 Oct 2010 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 12. SAMPLING METHOD AND APPROACH 12.1 DRILLING, SAMPLING AND SAMPLE PREPARATION The drilling approach is detailed in Section 11, sampling method, sample preparation and analysis in Section 13, and quality control methods employed in Section 14. 12.2 RC AND CORE RECOVERY – HASSAI SOUTH AND HADAL AWATIB EAST Core recovery and RQD has been measured systematically since 1992. All such data from the 2008/09 program have been loaded into the database. Prior to that, measurements were made but were not loaded systematically into the database. However, historical reports confirmed the results of recent drilling, ie that recoveries were average to poor in the oxide zone, improving to very good (>90%) in supergene and primary sulphide mineralisation. For recent diamond drilling at Hassai South, 90% of the primary mineralisation intervals were above 98% recovery, while 90% of supergene mineralisation intervals exceeded 90%. In most cases where RC drilling intersected base metal sulphide mineralisation, mineralised intercepts have now been replaced by core drilling. However, since 2007, where RC drilling supplies resource information samples (as at Hadal Awatib East), samples have been weighed, demonstrating acceptable recoveries in sulphide zones. All supergene and primary mineralised intersections at Hassai South were from core drilling. 12.3 RC AND CORE RECOVERY – KAMOEB Core recovery for Kamoeb is very good, approximating 100% in the mineralised zone 99% of the time. RC weights were systematically recorded for each metre drilled. Accuracy of weight measurement is not very precise, but indicates a satisfactory recovery of RC samples. Median weight is 31 kg with a coefficient of variation of 18%. 95% percent of samples are between 20 and 40 kg. 12.4 AUGER RECOVERY – TAILINGS Drilling samples from 2007/2008 drilling were weighed. Based on the nominal drill hole diameter (50 mm), theoretical recovery was calculated for a range of densities. Median theoretical recovery for 2007/2008 drilling over this range was between 79% and 98.6%, confirming good to very good recovery. Drilling sample recoveries for 2009 work were measured indirectly, but suggest similar recovery to 2007/2008 drilling. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 66 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 13. SAMPLE PREPARATION, ANALYSES AND SECURITY 13.1 INTRODUCTION Different approaches have been adopted for exploration samples for gold and base metal mineralisation. 13.1.1 Gold Core samples from 1992 to 2006 for gold were prepared on site and analysed at AMC’s mine laboratory using a cyanidable gold method. Since 2007 exploration samples (RC and core) for gold have been prepared mostly at Hassai and assayed by Intertek, Jakarta, using fire-assay methods. The Kamoeb resource estimate is based on a mix of both methods. Most of the oxide VMS (SBR) resource estimates are based on cyanidable gold assays. 13.1.2 Base Metals Base metal sulphide core samples from recent drilling (2008/09) were sent to the Intertek laboratory in Jakarta (half or quarter cores) for preparation and analysis (fire assay for gold, and acid attack with AAS for base metals). Earlier core samples were prepared on site (full core), and analyses undertaken off-site, but few records exist from this time and there is uncertainty about which laboratory undertook the work. The VMS resources (gold and base metals) are based very largely on the Intertek assays. 13.1.3 Heap Leach Tailings Gold Auger samples from drilling of heap leach pads A to D were prepared on site. The 2007/2008 samples were originally analysed on site for cyanidable gold, but the samples were thereafter sent to Intertek, Jakarta for reassay by fire assay, and the latter results were used for resource estimation. The 2009 samples were analysed at Intertek, Jakarta. Tailings resources are therefore based on fire-assay data. 13.2 SAMPLING, SAMPLE PREPARATION AND STORAGE 13.2.1 Gold Exploration: 1992 to 2007 13.2.1.1 Core Samples Historical drill hole sampling is described as follow by Monthel, et al. 2007: • Material is sun-dried if required • Whole core is jaw crushed to -16 mm (6 to 12 kg). • Riffle split to approximately 3 kg sub-sample • Grinding to -2 mm using a Nyberg cone crusher • Riffle split and a several hundred gram split is milled to 125 µm (preceded by a quartz charge if the bowl is dirty) • pH test: samples pH<1.5 are roasted for 2 hours at 650°C FINAL – Rev 0 – 22 Oct 2010 AMEC Page 67 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report • • Split to four by 75 g sub-samples for: − Cyanide gold analysis at AMC mine laboratory − Three samples retained as possible check samples for pH, total gold, polymetallics, etc. Once results have been provided, the geology department retains samples with gold >1.5 g/t, as well as sulphide samples. All gold assays were made at the AMC mine laboratory using a cyanidable gold leaching method (see section below). 13.2.1.2 Gold Analysis by Cyanide Leach (AuCy) This is undertaken at the AMC mine laboratory. A 20 g sub-sample is mixed with 50 ml of NaCN (5 g/L) and NaOH (1%), shaken for 3 hours, and analysed by AAS with a detection limit of 0.1 g/t. Data is manually compiled, entered into a spreadsheet and forwarded to the Geological Department. This method is in still use for grade control and plant control samples, as well as for reconnaissance drilling samples. 13.2.2 Gold Exploration: 2008/09 13.2.2.1 Core Samples Samples are prepared as presented in sub-section 13.2.3.2 au-dessous. Quarter cores were sent to Intertek, Jakarta. 13.2.2.2 RC Samples In 2008, RC samples were weighted and sun-dried if necessary. Full samples were then riffle split to 1 kg using a 3-stage riffle splitter. The coarse rejects were stored, and 1 kg sub-sample submitted to Intertek for further grinding and analysis. In 2009, following the setting-up of an exploration sample preparation laboratory on site, RC samples were split to 5 kg, then sent to the sample processing laboratory for cone crushing and splitting to 1 kg. Rejects are stored in sea containers at AMC’s camp. The 1 kg sample crush was pulverised using Labtechnic LM2 mills to an estimated -125 µm, and split to 250 g. The 250 g sample was submitted to Intertek for further grinding and analysis. The RC samples were packed in wooden boxes for road transport to Khartoum. From there they were airfreighted to Jakarta where they were collected by Intertek, and processed according to the laboratory PT01 procedure. This involved drying (105°C), crushing and pulverising the entire sample using Labtechnic LM2 mills to 95% <75 µm. 13.2.2.3 Analysis, Fire-assay All pulps were analysed by Intertek, using a 30 g or 50 g sub-sample taken for fire assay with AA finish. A repeat assay using gravimetric finish was undertaken for samples >50 g/t Au. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 68 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 13.2.3 Base Metal Sulphide Exploration From 1992 to 2007, full cores were sampled and prepared in AMC’s sample preparation laboratory. Sulphide samples were assayed in AMC’s mine laboratory for gold by cyanidable leach, but these results were not used for the sulphide resource estimation. The bulk of drilling for VMS resource estimates was undertaken in 2008 and 2009. Samples were prepared in both AMC’s and Intertek’s sample preparation facility, and analysed by Intertek using fireassay for gold and acid attack with AAS for silver and base metals. 13.2.3.1 Diamond Drill Holes, 1996-2002 Historical drill hole sampling is described as being done in an exhaustive manner (Monthel, et al. 2007). Whole core was sampled at intervals ranging from 0.75 to 1.5 m, depending on lithological contacts. Physical preparation of the samples was carried out by teams of workmen supervised by an experienced foreman at the AMC laboratory. Whole core was first crushed to -16 mm, split down to 3 kg, ground to 2 mm, split down to 200 to 400 g, and ring-milled to 125 μm. A quarter of the sample was submitted for assay, while the pulp reject was kept for further analysis if required. Very little remains from these early historical samples, as full core was used in sample preparation, most of pulp rejects have not been kept and supposedly barren cores were disposed. The lack of reference samples from historic drill hole is considered as a minor issue for the estimation of sulphide ore, since few such holes were used for resource estimation (eg only 9 of 60 intersections at Hassai South are from pre-2008 drilling), and the results from older holes are in good agreement with adjacent more recent drilling. 13.2.3.2 Diamond and RC Drilling, 2008-09 A 1 kg sample has been taken from each RC pre-collar sample. One hole on every other traverse is submitted for analysis. If there is any mineralisation, additional holes will be submitted but this appears unlikely. Logging and sampling of drill core was carried out by field geologists and experienced foremen, under the supervision of a senior geologist, and involved: • True depth correction compared to the driller depths • Technical logging including RQD and core recovery based on corrected depth and whole cores • Sampling line drawn along the cores • Sawing of the cores • Geological logging and photographs of all cores boxes • Sampling according to the geologist sampling plan • Always the same half of the core sent for assay • Samples intervals ranging between 0.5 and 1.5 m depending on the geology. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 69 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Half core samples were packed in wooden boxes, which were transported by truck to Khartoum. They were then airfreighted to Jakarta where they were collected by Intertek, and processed according to the laboratory PT01 procedure. This involved drying (105°C), crushing and pulverising the entire sample using Labtechnic LM2 mills to 95% <75 µm. For some samples, primary crushing followed by a first split to 1 to 2 kg may have occurred. All cores are stored in a core-shed, or in a sea container. All core boxes are properly labelled and carefully stored. 13.2.3.3 Base Metal Assaying 1996-2006 Drilling Details regarding assaying samples from the small proportion of diamond drill holes completed in the sulphide zone between 1996 and 2006 are scanty. Gold was assayed by fire assay (external laboratory) or by 3-hour cyanide leach and AAS (Hassai Mine laboratory). Silver and base metals are believed to have been assayed at an external laboratory, most probably OMAC in Ireland, as reported by Monthel et al., 2007, although no records are available. It appears that quality control samples were not included with these samples. Note that silver and lead contents are not an issue since the grades are low and usually below a reasonable cut-off. Given the relatively low proportion of historical assays compared to more recent assays, these historical base metals assays were included in the latest resource calculations. As might be expected, there is a significant difference between cyanidable gold and fire assay results in the sulphide samples, thus all cyanidable gold assays were discarded for the latest base metal sulphide resource estimate. 2008-2009 Diamond Drill Holes All the samples were submitted to Intertek, Jakarta, which is an independent laboratory with ISO17025 accreditation. The laboratory was visited by the author in May 2009 and is considered to meet international standards. Silver and base metals were assayed using the Intertek procedure GA30: • Triple acid digestion (HCL/HNO3/HClO4) followed by accurate volumetric finish • AAS analysis. Gold was assayed using fire assay with AAS finish, on a 30 g pulp. 13.2.4 Heap Auger Drill Samples Sample preparation in 2007 was undertaken by AMC at Hassai Mine as follows: • Weighing of sample • Drying • Crushing to D95 <2 mm FINAL – Rev 0 – 22 Oct 2010 AMEC Page 70 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report • Splitting to approximately 250 g; remainder (approximately 750 g) stored on site as reference material • Milling to 125 µm • Splitting into four tubes, one for assay and the remainder as duplicate and reference samples. Intertek analysis of 2007 samples was by fire assay on 30 g charges with AA finish. Samples from 2009 drilling were split on site to 250 g, with the remainder stored on site for reference. The 250 g sub-sample was submitted to Intertek, Jakarta, where it was milled to -75 µm in an LM2 mill, with a 50 g sub-sample taken for fire assay with AA finish. A repeat assay using gravimetric finish was undertaken for samples >50 g/t Au. As with base exploration samples, the auger samples were transported by truck to Khartoum. They were then airfreighted to Jakarta where they were collected by Intertek. A proportion of the rejects from 2007 drilling stored on site has been damaged. Samples from 2009 remain intact. 13.3 DRY BULK DENSITY 13.3.1 Core Samples Dry bulk density in the oxide zone and supergene/primary mineralisation was measured systematically on drill core, using a standard hydrostatic method with wax on half core lengths of 5 to 8 cm, as follows: • Air drying of the sample. • Weighing of the dry sampler, using an electromechanical balance (weight Ws). • Coating of the sample with a thin paraffin film of known density (0.925), and weighing in air (weight Wt). • Hydrostatic weighing of the paraffin coated sample immersed in water (weight Wti). A Mettler PM2000 electromechanical balance, fitted with a hydrostatic weighing device is used for the process. • Density is calculated using the following formula: d = Ws/((Wt-Wti)-(Wt-Ws)/0.925). Systematic density measurements from core drilling of supergene and primary mineralisation from 2008 drilling were undertaken using a similar procedure. Overall results can be summarised as follow: • • Hadal Awatib East − Oxide zone (SBR): 1380 samples, with density ranging from 1.7 to 2.6 (min=1.2, max=3.4) with an average density of 2.18 − Sulphide zone (supergene and primary): 58 samples, ranging between 1.0 and 4.9, averaging 4.39 Hassai South (sulphide) − Supergene: 343 samples ranging from 1.5 to 5.0, average 4.19 − Primary: 168 samples ranging from 2.68 to 4.93, average 4.31 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 71 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report • Kamoeb: 46 samples, ranging from 1.76 to 2.93, averaging 2.56 (median = 2.61). Additional data from GEOSICA suggest an average density of 2.8. 13.3.2 Auger Samples A total of 40 measurements were made at various points on heaps A to D in 2007, with a further three measurements made in 2009. Three methods were tested, with the “Hand-compacted” density method results used for converting volume to tonnage. The “Hand-compacted” method involves filling a box of 10 405 cm3 volume with dump material, using a shovel. Fill is compacted by hand every 5 cm. When full the box is weighed, dried and reweighed to allow calculation of wet and dry bulk density. The in situ heap material consists of agglomerated pellets up to approximately 1 cm top-size, and while it is deposited uncompacted, it packs well, and undergoes some compaction in the heaps due to deposition of overlying layers. The mean dry bulk density varies from 1.5 to 1.9 depending on the compaction factor. This variability is due to the nature of the heap material as well as the somewhat subjective nature of the test. Reconciliation with production data (2007) suggests a wet density of 1.65 to 1.7, with a corresponding dry density of 1.5 to 1.6. For resource estimation purposes, a conservative value of 1.5 was selected. No density quality control has been undertaken. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 72 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 14. DATA VERIFICATION 14.1 DATA COLLECTION Recent drill hole data is compiled in a master spreadsheet. The drill hole data collected by the exploration team and transferred to the master Excel spreadsheet is verified by the exploration geologist dedicated to the management of the data. Historical drill hole data were collected on a similar manner, using several mining softwares (SERMINE, GDM) and Excel spreadsheets. The data were compiled collected within a simple database and reexported for resources estimation. During the various migrations between systems, some issues appeared for duplicates and for samples depth from and depth to. Database validation was completed to identify and resolve such issues. A separate Excel-based database was employed for the heap auger program. Logs were recorded under supervision of an experienced geologist. Relevant information regarding data verification for this dataset is contained in Section 14.7. The master Excel spreadsheet is endorsed by the AMC Exploration Manager before use in resource modelling. 14.2 ASSAY DATA QUALITY Pre-2008 drill hole samples were not submitted with a proper quality assurance/quality control (QAQC) programme. However, these holes contribute little to the database for estimation of remaining resources, including the base metal sulphide resources. An Inter-laboratory test for a limited drilling program in 2005/06 was undertaken in Hadal Awatib East oxide zone. This indicated a consistent bias of 7 to 10% between AMC Mine laboratory (cyanidable gold) and OMAC (fire assay gold), the mine laboratory returning lower values (Monthel, et al. 2007). Similar tests were undertaken for Kamoeb, first in 2003 (Grove, 2003) with 107 samples sent to OMAC, Ireland and in 2009 (AMC, Bennet, 2009) with 100 samples from 2003/04 pulps sent to Intertek Jakarta. Conclusions are similar, and the results show an acceptable correlation between the different analyses, with the fire-assay being 5-10% higher than the cyanidable gold. The fire assay analyses tend to return higher values for samples >10 g/t Au, however the number of samples in this range are limited. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 73 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 14.1 Kamoeb – 2003 AMC/OMAC Check Assay (Grove, 2003) KAMOEB Check Assay QC Plot 25 OMAC FA_AU 20 y = 1.0584x R² = 0.9315 15 10 5 0 0 5 10 15 AMC CY_AU 20 25 A thorough quality control scheme was applied during the 2007/09 drilling campaign for all deposits, involving blanks, duplicates and certified reference materials (CRM, or standards) for gold and base metals inserted in the whole sequence of samples, as follows: • Hadalal Awatib East: 234 blanks (3.6% of samples), 434 duplicates (6.7%) and 397 standards (6%) have been interspersed in 6473 samples • Hassai South: 118 blanks (3.2%), 212 duplicates (5.7%) and 252 standards (6.75%), amongst 3733 samples • Kamoeb: • − In 2008, 140 blanks (3.4%) 137 duplicates (3.3%) and 144 standards (3.5%), amongst 4106 samples − In 2009, 140 blanks (1.8%) and 404 standards (5.3%), with 7623 samples Tailings: − 2008, 93 blanks (3.3%) 89 duplicates (3.2%) and 173 standards (6.2%) with 2794 samples − 2009, 36 blanks (2.2%) 105 duplicates (6.5%) and 84 standards (5.2%) with 1625 samples. Assay data from recent drilling is emailed to the project geologist by Intertek, and compiled with sample information in an Excel spreadsheet. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 74 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 14.2.1 Blanks These were made from sieved sand dune material, and were presented in 50 g tubes at a rate of approximately 2% to 4%. Results display a slightly elevated gold background (average 0.06 g/t Au) remain below a generally accepted level of 5 times the detection limit; these levels may indicate a low level of site pollution, but will not materially affect resource estimates. Outliers, as identified by anomalous multi-element signatures, reflect sample swap errors. In 2007/08, some anomalous blanks values were identified in the tailings dataset. A small number of blank samples returned values greater than 0.1 g/t, indicating possible contamination. Three batches of results from Heaps B and C were eliminated from the resource database on this basis. 14.2.2 Standards Eleven commercial standards (Standards or CRMs) supplied by Rocklabs, Gannet and Geostats were used for assay quality control. These samples represent a wide range of gold and base metal (Cu, Zn) grades and sulphide-bearing matrix types (Table 14.1). CRM was submitted as 30 to 50 g sachets accompanying exploration (primarily cut core) samples, at a rate of approximately 4-7%. A small number of sample handling errors was detected, including misidentified Standard number and exchange with exploration sample number; this is readily recognised on the basis of the characteristic multi-element signature of each Standard. Results were assessed in terms of coefficient of variation (COV = Standard deviation / average). A good laboratory with regular precise assay will have a COV) in the range of 3-5% for gold, and below 3% for base metals. Note that the lower the certified value is, or the lower the number of measurement is, the higher is the coefficient of variation. The Intertek analyses show acceptable to good COV for gold and base metals (Table 14.2 to Table 14.5). High COV values for two Standards are related to small statistical data-sets. No significant bias was observed between Intertek results and the accepted CRM values. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 75 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 14.1 Characteristics of Standard Reference Materials Laboratory Name Matrix Gannet PGO Sulfide Compass Au Cu Zn Pb Ag (ppm) (%) (%) (%) (ppm) 2.6 1.6 3.3 15.7 10.8 VMS Resources Gannet Sulfide 0.11 0.56 3.6 Geostats GBM-308-13 Cu Sulfide Ore (16%S) 1.858 0.963 0.325 19.8 Geostats GBM-308-14 Cu Sulfide Ore (32%S) 3.719 1.902 0.651 40.2 Gannet SU011 Sulfide (2.54%S) 0.97 0.4 Gannet SU008 Sulfide (0.27%S) 0.41 0.04 Gannet SU012 Sulfide (21.2%S) 2 0.64 Geostats G3018 Pyrite concentrate 1.19 Geostats G9076 Sulphide gold ore 7.25 Geostats G3055 Sulphide gold ore 2.43 Rocklabs K2 (HiSilK2) Siliceous (1.0% S) 3.47 Kamoeb Resources Rocklabs HiSilK2 Siliceous (1.0% S) 3.474 Rocklabs OxF65 Oxide material 0.805 Rocklabs OxH55 Oxide material 1.282 Rocklabs OxH66 Oxide material 1.285 Rocklabs Oxi54 Oxide material 1.868 Rocklabs OxJ64 Oxide material 2.366 Rocklabs OxL63 Oxide material 5.865 Rocklabs OxN62 Oxide material 7.706 Rocklabs OxP61 Oxide material 14.92 Rocklabs HiSilK2 Oxide material 3.474 Tailings Resources Rocklabs OxH55 Oxide material Rocklabs Oxi54 Oxide material Rocklabs OxE56 Rocklabs Si42 Oxide material Siliceous Rocklabs OxJ64 Oxide material Rocklabs OxF65 Oxide material 1.282 1.868 0.611 1.761 2.366 0.805 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 76 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 14.2 CRM Assay Results – Oxide Gold Standard Count Expected Value Average Value Coefficient of Bias (Au g/t) (Au g/t) Variation (%) (%) Tailings, 2008 OxH55 50 1.28 1.32 3.2 3.1 OxI54 47 1.87 1.83 3.4 2.3 OxE56 34 0.61 0.65 3.2 5.8 Si4 29 1.76 1.77 3.6 0.3 OxJ64 13 2.37 2.32 1.4 1.8 Tailings, 2009 OxH55 50 1.28 1.29 2.3 0.9 OxF65 34 0.81 0.80 2.7 -1.0 Blank 36 - 0.03 Kamoeb, 2008 G301-8 1.19 1.09 4.8 -8.5 G305-5 2.43 2.36 4.0 -2.8 G907-6 7.25 7.09 2.2 -2.2 G306-4 21.57 21.05 2.5 -2.4 G999-4 3.02 2.98 3.2 -1.4 Kamoeb, 2009 OxF65 72 0.81 0.81 2.21 0.7 OxH55 42 1.28 1.31 1.86 2.2 OxH66 15 1.29 1.32 1.47 2.7 OxI54 10 1.87 1.87 2.44 0.6 OxJ64 99 2.37 2.38 2.59 1.4 HiSilK 95 3.47 3.42 3.19 -0.7 OxL63 32 5.87 5.93 2.78 1.2 OxN62 20 7.71 7.35 1.32 -4.6 OxP61 19 14.92 14.59 1.25 -2.1 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 77 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 14.3 CRM Assay Results – VMS Gold Standard Count Expected Value Average Value Coefficient of Bias (Au g/t) (Au g/t) Variation (%) (%) Hadal Awatib East Su011 2 0.97 0.80 18.7 -22 SU008 10 0.41 0.42 7.9 2 SU012 2 2.00 2.11 6.4 5 G301-8 65 1.19 1.14 4.0 -4 G907-6 66 7.25 7.15 2.3 -1 G305-5 64 2.43 2.42 3.2 0 Hassai South Su011 16 0.97 0.93 5.4 -4.1 SU008 12 0.41 0.43 12.4 4.9 SU012 13 2.00 2.03 2.9 1.5 G301-8 20 1.19 1.14 3.0 -4.2 G907-6 16 7.25 7.15 2.4 -1.4 G305-5 19 2.43 2.42 1.3 -0.4 HiSilK2 7 3.47 3.40 1.3 -2.0 Table 14.4 CRM Assay Results – VMS Copper Standard Count Expected Value Average Value Coefficient of Bias (Au g/t) (Au g/t) Variation (%) (%) Hadal Awatib East PGO 33 2.60 2.59 2.2 0 Compass 35 0.11 0.11 4.7 3 GBM308-13 50 1.86 1.86 2.1 0 GBM308-14 66 3.72 3.68 2.1 -1 SU011 2 0.40 0.41 3.4 2 SU008 10 0.04 0.05 11.7 11 SU012 2 0.64 0.67 1.1 4 Hassai South PGO 31 2.60 2.59 2.7 -0.4 Compass 21 0.11 0.12 11.5 9.1 GBM308-13 52 1.86 1.85 2.1 -0.5 GBM308-14 41 3.72 3.64 2.2 -2.2 SU011 16 0.40 0.43 4.2 7.5 SU008 12 0.04 0.05 57.8 25.0 SU012 12 0.64 0.66 4.2 3.1 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 78 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 14.5 CRM Assay Results – VMS Zinc Standard Count Expected Value Average Value Coefficient of Bias (Au g/t) (Au g/t) Variation (%) (%) Hadal Awatib East PGO 33 1.60 1.61 2.2 1 Compass 35 0.56 0.55 1.9 -1 GBM308-13 50 0.96 0.93 2.0 -3 GBM308-14 66 1.90 1.83 2.4 -4 PGO 31 1.60 1.68 3.1 5.0 Compass 20 0.56 0.56 3.6 0 GBM308-13 52 0.96 094 1.6 -2.1 GBM308-14 41 1.90 1.83 2.5 -3.7 Hassai South 14.2.3 Duplicates Pulps duplicates were submitted to Intertek, either blindly (Kamoeb 2008, Tailings), or on carefully selected mineralised intervals. All cross-plot show acceptable dispersion around the X=Y line. 14.2.4 Conclusions Systematic quality control is a recent feature of exploration work at Hassai. However, it applies to the bulk of samples supporting remaining resources. Earlier drilling campaigns were supported from time to time by small sets of samples (usually about 30 samples) sent to an external laboratory for a cross-validation, and a conservative bias was identified in earlier gold analyses from the mine laboratory. The recent quality control procedure is considered acceptable, but has a few weaknesses: • Duplicates are organised by the external laboratory • No quality control follow-up has been set-up, allowing AMC to react in case of a fatal flaw • Field/Coarse duplicates should be included • Standards and blanks are 50 g sachets and tubes intercalated in 2 to 5 kg rock samples and are easily spotted • No coarse blanks have been sent that would help monitor the sample preparation quality. Recent quality control data indicates that the latest exploration results – which provide the majority of the database for remaining base metal and gold resources – are of acceptable quality for resource estimation for NI 43-101 purposes at this level of confidence. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 79 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 14.3 KAMOEB TWIN HOLES In 2008, a number of twin holes were completed to check diamond drill and cyanidable leach assay against RC drill and fire assay on a key area in the southeastern corner of Kamoeb. A number of statistical comparisons were completed, from which it appeared that diamond core cyanidable gold was some 7% to 8% higher grade than RC with fire assay. This is in contrast to previous studies that indicated cyanidable gold undercalled fire assay gold by 5% to 10%. 14.4 GEOLOGICAL DATA Geological data was recorded by a team of field and senior geologists employed by AMC. The geological data was collected manually at the drilling stage, then transferred to an Excel spreadsheet by a dedicated junior geologist. No independent systematic check of the quality of geological data has been made. 14.5 SURVEY DATA Collar survey data has been collected by qualified AMC mine surveyors. The collars coordinates were then transferred to an Excel spreadsheet and passed to the project geologist. Collar plots have been reviewed by AMC geologists and are believed to accurately represent drill hole locations. Down-hole survey data was examined for consistency, but no other quality control has been undertaken. A digital elevation model (DEM) was developed for each deposit by the AMC survey team using Leica DGPS and Total station methods, at different period of time, as the deposit were mined. Minor issues were noted and corrected when necessary. Extra care was given to the tailings where the DEM is a critical element to the volume estimate. All collars are consistent with the DEM, except when the deposit has been partially mined out. 14.6 DENSITY DATA No density quality control has been undertaken. 14.7 DATABASE VERIFICATION 14.7.1 Database Consistency – Internal Review Data is loaded into Surpac Vision. Routine checks undertaken in Surpac showed only minor issues that were addressed, as follows: • Depth consistency between collar and logs • Sample overlap • Survey consistency with the drill hole depth • Out of bound assays. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 80 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report The following errors were noted: • A few maximum depth errors (discrepancy between geological and assay log and collar or survey maximum depth) that are easily corrected. • Overlapping sample errors existed for some duplicate samples. • On occasions, diorite and schist are shown as mineralised, which suggest that there were errors in the geological log, either during logging or during the data entry. In Hadal Awatib East, after the independent data validation, it appeared that AMC had relogged the geology, but that geological logs had not been updated in the database. • Historical core recovery is often missing. • Some drill holes have collar survey information but no assay data. This was found to be due to no visible signs of mineralisation, and thus no samples were submitted for analysis. • A number of holes without lithological logs. These holes are largely from the oxide zone within mined-out pits, and are of minor importance with respect to remaining resources. It was noted that a small number of assay laboratory reports contained transcription errors, that could originate either from the laboratory or when creating sample submission sheets. These errors are minor and easily corrected. Collars were checked against the topography and are consistent. The lithological codes contain numerous typographic errors that are easily corrected. Two drill holes (Hadal Awatib_D160 and HASS_D230) had values inconsistent with adjacent readings or adjacent drill holes and indicate probable measurement errors. The massive sulphide interval in Hadal Awatib_D160 has not been used for wireframing, but its assays have been manually added into the sample-set used for statistics and interpolation. The error value in the lower part of HASS_D230 has been disregarded. 14.7.2 External Independent Data Validation In Hassai South and Hadal Awatib East, data were independently validated by a senior geologist from Arethuse, in order to verify the database compared to the original data. A proportion of the original data were scanned by AMC and provided to Arethuse, along with the database. Approximately 25% of all drill hole intercepts into massive sulphides were checked against the field logs and the laboratory data sheets. Verification of mineralised zones compared to core-box photos was not included in the scope. Errors were classified as: • Minor: minor typos in depth, or in numerical values. • Major: important numeric error, missing lines, etc. with no determinant error on the mineralised intervals. • Critical: Major error on a mineralised interval. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 81 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Criteria of success are as follow: • Above 5% of errors, the entire database should be re-verified and corrected • Between 2 and 5% errors, Phase B is repeated • Below 2% error, the database is considered internally validated. Hassai South The validation review demonstrated an overall error rate of below 2%. No critical errors were identified. A moderate number of errors were identified, both minor and major, but these were: • Dominantly lithological errors • Largely related to pre-2005 drilling, and thus representing a small part of the data used for estimation of massive sulphide resources and unmined gold resources. In this regard, over 90% of the sulphide resources are based on 2008/09 drilling, while Indicated oxide resources are derived largely from 2005/06 work, which contains few errors. Hadal Awatib East The number of errors was too large to completely validate the database, comprising: • Lithological reporting errors (2008-09 drill holes): these are not critical, and present only approximations. The approximation is of smaller order of magnitude compared to the broad geological model used for the resource estimate. • The historical assay database (prior to 2002 drill holes) present numerous significant imprecisions in depth reporting. • Intertek assay database (2008-09 drill holes) is valid. Although formally not validated, the Hadal Awatib East drill hole data-set can be used for the 2009 definition of indicated and inferred resources as per NI 43-101, as: • A major error-free set of data is not demonstrated, but major fatal flaws for resource definition are unlikely • Imprecision in resource definition due to the drill holes database errors is likely to have happened, but will concern mostly the inferred resources or already mined-out material • The consistent set of data acquired in 2008-09 can be validated despite the database status. Indicated sulphide resources have been estimated using this part of the data-set. Considering that: • Indicated sulphide resources have been estimated using mostly 2008-09 data (715 samples out of 730 samples – 98%) • Indicated oxide resources are based on historical drill holes 2005-06 • Inferred oxide resources are based on historical drill holes prior to 2000 • Inferred sulphide resources are base on 2213 samples, of which: − 8.7% (192 samples) are from drilling pre-2000 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 82 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report − 0.9% (21 samples) are from 2005-06 drilling − 90.4% (2000 samples) are from 2008-09 drilling. It is the relevant QP’s opinion that no major or critical errors are likely to significantly affect the materiality of the resources as they have been classified: mostly inferred and partially indicated. Indicated resources have been estimated with 2008-09 data that has the higher degree of confidence. Kamoeb The database was initially not fully set-up, and it appeared to be necessary to rebuild it from existing data. The author verified every drill hole and corrected the database from the original data: filed logs, manual cross-sections, laboratory datasheet, etc. The relevant QP was involved in the making of the database, and could not review it independently. Still, a number of tests, similar to those one undertaken for Hassai South and Hadal Awatib East were carried out to resolve any potential flaw, with acceptable results. It is the QP’s opinion that, no major or critical errors are likely to significantly affect the materiality of the resources as they have been classified. 14.7.3 Independent Sampling and Analysis (Gold Only) In 2007, Geostat visited site and undertook independent sampling and analysis of 40 samples, including 25 grade control drilling duplicates (rifle splits), 4 stockpile samples, 4 heap leach residue samples, and 7 grab samples of gold mineralisation in open pits. Samples were analysed for cyanidable gold at the AMC laboratory, and at SGS Lakefield, Ontario using an equivalent method as well as by 30 g fire assay. One sample pair was rejected prior to assessment, since the two results bore no relationship to each other (Mine Lab = 0.4 g/t, SGS = 24.2 g/t). Statistical analysis of the remaining duplicate data showed no notable difference between the two laboratories. 14.8 AUGER PROGRAM DATA VERIFICATION 14.8.1 Drill Logs and Sampling Information Validation of the 2007 electronic database by the author initially identified numerous errors. Consequently, all data was checked against original hard copies and all problems were corrected. A final validation was then undertaken by the senior geologist in charge of exploration at Hassai. Checks included assay and collar survey data. The 2009 master spreadsheet displayed no obvious logical errors. A brief verification against original hard copies of field logs confirmed the database information. Seven errors in sample sequencing were readily identified and corrected at this time. The author considers the auger drill hole database to be in good order and suitable for use in resource estimation to meet NI 43-101 standards. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 83 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 14.8.2 Assay Data Quality In 2007, for each batch of 30 drill samples quality control samples were added, comprising two gold CRM standards, one blank and three duplicates. • In general, results for the standards were acceptable, with occasional anomalies related to incorrect standard numbering. • A small number of blank samples returned values greater than 0.1 g/t, indicating possible contamination. Three batches of results from heaps B and C were eliminated from the resource database on this basis. • 100 coarse split duplicates were analysed independently at Amdel, Perth, using fire assay with ICP finish method. The results show close agreement with the Intertek results across the full range of values. • 138 sample pulps were sent to SGS, Balikpapan for inter-laboratory confirmation using 30 g fire assays. Correlation of the two datasets is very good, again confirming the quality of data from Intertek. • 89 pulp duplicates were sent separately to Intertek to evaluate assay reproducibility. Very good agreement was noted with one exception, which is probably a sample identification error. Precision is approximately ±5% at the 90% confidence level, which is well up to industry standard. In 2009, the same QAQC sample protocol was adopted, but applied to batches of 50 samples. • Results for standards fell within expected limits with the exception of two results (out of 89) which were clearly due to swapping of reference samples. Results for the remainder display low COVs indicative of good laboratory performance. • Blank samples were slightly mineralised; being dune sands collected distant from the mine, it is thought that the results reflect low-level contamination in the site preparation area or at Intertek’s laboratory. However, the average value is 0.03 g/t, with a maximum <0.1 g/t, which is considered minor with reference to the resource grade. • Sample preparation duplicates show good agreement with originals. between the two datasets. 14.8.3 No bias was observed Auger Samples – Conclusions While a number of issues arose concerning the 2007 drilling database, these have been resolved by careful cross-checking and revalidation. Care was taken to avoid such issues in the 2009 program. QC samples introduced with the 2007 assay program identified some problems related to incorrect identification of CRMs. Interlaboratory check analyses have confirmed the reliability of the Intertek results. Minor contamination was identified in analyses of blank samples. Three batches of sample results accompanying highly anomalous blanks have been rejected from the resource database as a precaution. The final dataset is considered reliable for resource estimation purposes. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 84 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 15. ADJACENT PROPERTIES There are no adjacent properties relevant to this report. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 85 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 16. MINERAL PROCESSING AND METALLURGICAL TESTING 16.1 HEAP LEACH TESTWORK 16.1.1 Heap Leach Testwork During heap leach processing, AMC has conducted permanent column tests to check, confirm and adapt its plant parameters to the specificity of the different ores processed. This section does not show any results from the actual heap leach operation at Hassaï nor the testwork conducted to design this operation, but only the day-to-day tests undertaken to control and predict the performance of the plant. The parameters of the current heap leach operation are given in Section 20.3. The purpose of a column leach test is not so much to duplicate in a laboratory test the results that can be expected from a commercial heap leaching operation, but to collect kinetic and reagent consumptions information on the ore being evaluated so that scale-up equations can be validated. This will allow the projection of the commercial heap leach operation's performance under different operating scenarios. With regards to AMC experience, it is usually not necessary to run more than three column leach tests on each ore type of a particular deposit to validate a kinetic model. It is then to be adapted to the mine plan to ensure the best prediction of the parameters. 16.1.1.1 Predictive Tests on Core Samples Exploration group provides a series of samples representing the deposit. These core samples are divided into 10 m intervals regardless of the grade. The laboratory then prepares two composites by level and crushes the samples down to 12.5 mm (d100), as this is the size of the crusher final product. A fraction of the crushed sample is then used to determine head grade of Au, Ag and Cu, and pH to look for the presence of acidgenerating elements. The rest of the crushed sample is submitted to size analysis on a sieve set of 12.5 mm, 10 mm, 8 mm, 5 mm, 2 mm and 1 mm. Each split is then analysed for gold content to be able to run a recovery by particle size analysis after leaching. Thereafter, 30 kg of material is agglomerated, cured and placed in a 2 m high by 200 mm diameter Plexiglas column. The columns are irrigated at a specific flowrate of 16 L/m²/h using a 0.3 g/L cyanide solution at a pH of 10.5. A goal of 1.5 m3/t is set as the final solution to solid target. After leaching, the residue sample is dried and analysed for gold by particle size. A metallurgical and mass balance is carried out to calculate: • Gold recovered by both solution and solid balances • Cyanide consumption • Lime and cement consumptions • Compaction. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 86 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report From its experience, AMC has been able to establish that laboratory columns always perform faster than field heaps, for two major reasons: • The ore is placed in the columns in a more uniform way resulting in a better percolation through the heap. • The solution to ore solid ratio (tonnes of solution per tonne of ore) is generally higher in laboratory columns than in field heap. Therefore, a correction factor (0.9) is used to forecast operational recovery. The other data are used to predict the reagent consumptions, the compaction of the heap and the need for pH modifier as lime to be used in the industrial heap leach operation. By testing the different ore types of the deposit, the metallurgist is then able to forecast the performance of the heap leach operation, which is still to be confirmed by tests on material stacked on the heaps. 16.1.1.2 Tests on Stacked Ore Samples are collected at the automatic sampler prior to the stacking operation in order to prepare representative composites of the ore processed. When the composite is formed, column tests are performed in the same manner as presented before. The metallurgist can then predict with more accuracy the real performance of the leaching operation as he has now a representative composite of the ore processed. The corrected recovery of these column tests is then the target to be achieved by the operation. 16.2 CIL TESTWORK 16.2.1 Quartz Ore (Kamoeb South Deposit) A total of 750 kg of quartz ore samples from the Kamoeb deposit was received for testing purposes. This was combined to form a “Quartz Composite” for the testing regime. 16.2.1.1 Comminution Parameters Testing was undertaken to determine the following comminution parameters for the Quartz Composite: − Bond Ball Mill Work Index (BBWi) at a closing screen size of 106 µm − Bond Rod Mill Work index (BRWi) − Bond Abrasion Index (Ai) − Bond Crushing Work Index (BCWi) − Uniaxial Compressive Strength Tests (UCS) − SAG milling parameters (SAG Mill Comminution (SMC)) Tests. Results of the testing are summarised in Table 16.1. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 87 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 16.1 Quartz Ore Comminution Parameters Parameter Unit Result BBWi kWh/t 14.2 BRWi kWh/t 13.0 Ai g 0.4460 BCWi kWh/t 20.9 UCS MPa SG Dwi Mia 99.3 2.42 3.7 kWh/t 13.4 A 58.9 b 1.12 Axb 66.0 Analysis of the comminution results characterise the Quartz ore as being: • Of medium hardness with respect to ball and rod milling applications • Moderately to highly abrasive • Strong with respect to crushing applications • Soft for the purposes of SAG milling. 16.2.1.2 Ore Characterisation Head assay analysis was conducted on the Quartz Composite sample with a summary of the results being presented in Table 16.2. The head assay results shows that the Quartz ore contains gold grades which may be suited to CIP/CIL circuit leaching. The preg-robbing potential of the ore looks to be low and the potential for copper interference appears to be minimal. The mercury levels suggest that it may be necessary to provide capture equipment within the gold room facilities, and it will be necessary in subsequent testing to analyse for mercury in leach products and tailings. 16.2.1.3 Mineralogy Optical mineralogical analysis was undertaken to determine the gold occurrences within the Quartz ore. A sample was sized over a 100 µm sieve, with both the retained and passing size fractions subsequently being subjected to heavy liquid separation. Optical examination was then performed on each of the sinks fractions. A total of 10 gold occurrences were detected in the coarse sinks material, primarily associated with goethite, with two occurrences being aggregates of fine, low silver gold. In the fine sinks material, eight occurrences were detected, five of these being liberated gold particles with the remainder being associated with goethite. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 88 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report The results showed that the gold occurs in Quartz ore as fine-grained material which is either predominantly liberated or associated with goethite. Table 16.2 Element Quartz Composite Head Assay Assay Unit Ag 2.0 ppm ANC <0.5 kg H2SO4/t As 6 ppm Au 4.29 ppm Bi 1.0 ppm Ca 0.21 % CO2 <0.1 % Cu 70 ppm Cu (CN Sol) 12 ppm Hg 1.09 ppm Mg 0.75 % Ni 25 ppm Pb 120 ppm Pd <10 ppb Pt <10 ppb S 0.04 % 0.02 % Sb 1.0 ppm Si 40.2 % 2- S 16.2.1.4 Te 7.0 ppm Total C 0.05 % Total Organic C 0.05 % Zn 45 % Gravity Separation The response of the Quartz ore to gravity recovery techniques was assessed by treating a 30 kg subsample of the ore through a laboratory Knelson concentrator and subjecting both the obtained concentrate and gravity tailings stream to cyanide leaching. The feed, concentrate and tailings streams produced were then analysed for both gold and silver grades. A summary of the results is given in Table 16.3. The results of gravity testing indicated that around one-quarter of the gold present in the Quartz ore would be amenable to gravity separation techniques for the recovery of gold, though not for the recovery of silver. While the gold recovery value appears to be attractive, the mass recovery achieved in the testing is still around 10 times higher than what would be achieved in an operating plant and therefore would have a high probability of over-stating the true recovery. Based on previous experience, the result is at the low end of what would be considered viable for further investigation. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 89 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Results of the tailings leaching suggest a good recovery will be achieved without the use of a gravity circuit, and thus it is recommended that a gravity circuit is not included in future Hassaï flowsheets. Table 16.3 Gravity Separation Results Summary Composite Au Quartz Concentrate Au Concentrate 0.25 414 24.1 97.4 Tailings Tailings 3.33 75.9 93.0 Total Recovery Ag Concentrate Ag Total Recovery 94.1 Concentrate 0.25 20 3.7 86.9 Tailings Tailings 1.49 96.3 32.8 Total Recovery 16.2.1.5 Total Recovery 34.8 Grind Sensitivity Samples of the Quartz Composite were processed in a cyanide leach to determine the influence of grind size on precious metal recovery. A total of three grind sizes were tested, these being P80 212 µm, 150 µm and 106 µm, with each ground sample being leached in 1000 ppm NaCN with >15 ppm oxygen in solution. Table 16.4 summarises the results obtained from testing. Table 16.4 Quartz Ore Grind vs Leach Recovery Results Gold Grind P80 (µm) Residue Head Grade Grade (g/t) Silver Recovery (%) Residue Head Grade Grade (g/t) Recovery (%) 212 0.30 4.52 93.4 1.00 1.69 41.0 150 0.23 4.61 95.0 1.00 1.73 42.3 106 0.22 4.40 95.1 1.00 1.75 43.9 The results of the testing indicated that recovery was independent of grind size at a grind of 150 µm and below. All further testing with the Quartz ore was subsequently undertaken at a 150 µm grind. 16.2.1.6 Oxygen vs Air Sparging in Leach Two samples of the Quartz Composite were leached over 48 hours in a 1000 ppm cyanide solution to assess the effect of air sparging compared to oxygen sparing. The results in Table 16.5 demonstrate FINAL – Rev 0 – 22 Oct 2010 AMEC Page 90 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report that air sparging produced essentially the same recovery result when compared to oxygen. It was recommended that all further tests be undertaken with air sparging. Table 16.5 Air vs Oxygen Sparging Leach Summary – Quartz Ore Gold Gas Sparged Residue Head Grade Grade (g/t) Silver Recovery (%) Residue Head Grade Grade (g/t) Recovery (%) Oxygen 0.23 4.61 95.0 1.00 1.73 42.3 Air 0.25 4.95 94.9 1.00 1.86 46.1 16.2.1.7 Leach Cyanide Sensitivity The influence of cyanide concentration in the leach was investigated by targeting 250 ppm, 500 ppm and 750 ppm cyanide concentrations over a 48 hour leach. The results shown in Table 16.6 indicate that good gold recoveries can be achieved with low cyanide levels being maintained in the leach conditions. Table 16.6 Quartz Ore Leach Cyanide Sensitivity Gold NaCN Concentration (ppm) Head Grade Residue Grade (g/t) Recovery (%) Silver Head Grade Residue Grade (g/t) Recovery (%) 250 0.32 4.64 93.1 1.00 1.79 44.2 500 0.33 4.67 92.9 1.00 1.78 43.9 750 0.27 4.30 93.7 1.00 1.73 42.1 16.2.1.8 Quartz Ore Reagent Consumption A review of the leaching test results was undertaken to estimate the reagent consumptions for the quartz ore to be utilised in the processing plant engineering design. Due to the rapid leach dissolution achieved in testing, 24 hour reagent consumptions were chosen as the basis for plant design. For the Quartz ore, the following reagent consumptions were estimated: • NaCN, 1.45 kg/t. • Quicklime, 0.32 kg/t. The above results were from an average of six leaching tests. 16.2.2 Heap Leach Residue A total of 118 individual heap leach residue samples were provided for testing, of which 1/6 were combined to form a Heap Leach Bulk Composite to be used for metallurgical testing. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 91 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 16.2.2.1 Comminution Parameters The particle sizing of the heap leach material is anticipated to be F100 8.0 mm and, as such, the most likely processing route would involve ball milling prior to leaching. As a result, only the BBWi parameter was determined during the laboratory testing regime. The value for BBWi was determined to be 12.0 kWh/t, indicating that it would be classed as a moderate to soft ore for ball milling purposes. 16.2.2.2 Ore Characterisation Head assay analysis was conducted on the Heap Leach Bulk Composite sample with a summary of the results being presented in Table 16.7. Table 16.7 Element Heap Leach Bulk Composite Assay Assay Unit Ag 13.0 ppm ANC 7.1 kg H2SO4/t As 220 ppm Au 2.15 ppm Bi 14.0 ppm Ca 0.59 % CO2 0.1 % Cu 365 ppm Cu (CN Sol) 40 ppm Hg 13.8 ppm Mg 0.16 % Ni 0.009 ppm Pb 888 ppm Pd 15 ppb Pt 10 ppb S 2.88 % S 2- 1.21 % Sb 25.8 ppm Si 33.0 % Te 27.2 ppm Total C 0.04 % Total Organic C 0.02 % Zn 150 % The head assay results show that significant levels of both gold and silver remain in the heap leach tailings. There also appears to be higher levels of copper than in the Quartz ore, which may need further consideration. Of note, the levels of both arsenic and mercury are significantly higher than for the Quartz Ore and will need to be studied further in future testing. As with the Quartz ore, these FINAL – Rev 0 – 22 Oct 2010 AMEC Page 92 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report mercury levels suggest that it may be necessary to provide capture equipment within the gold room facilities. 16.2.2.3 Mineralogy Optical mineralogical analysis was undertaken to determine the gold occurrences within the Heap Leach Bulk Composite. A sample was sized over a 100 µm sieve with both the retained and passing size fractions subsequently being subjected to heavy liquid separation. Optical examination was then performed on each of the sinks fractions. The analysis found no optically detectable occurrences of gold in either the coarse or fine sinks fractions. This suggests that the gold is present as very fine particles, and that gravity recovery techniques would be ineffective on heap leach material. 16.2.2.4 Gravity Separation The response of the Heap Leach Bulk Composite to gravity recovery techniques was assessed by treating a 30 kg sub-sample of the ore through a laboratory Knelson concentrator and subjecting both the obtained concentrate and gravity tailings stream to cyanide leaching. The feed, concentrate and tailings streams produced were then analysed for both gold and silver grades. A summary of the results is given in Table 16.8. Table 16.8 Heap Leach Bulk Composite Gravity Separation Results Summary Composite Au Heap Leach Concentrate Tailings Mass Pull % 0.38 Head Grade g/t 18.2 Distribution % 4.1 94.5 Leach Recovery % Head Grade g/t 1.72 Distribution % 95.9 Leach Recovery % 37.5 % 39.8 Mass Pull % 0.38 Head Grade g/t 14.7 Distribution % 1.0 Total Recovery Ag Concentrate Tailings Total Recovery Leach Recovery % 64.7 Head Grade g/t 8.30 Distribution % 99.0 Leach Recovery % 51.8 % 51.9 The results from Table 16.8 show that the gravity recovery from the Heap Leach Bulk Composite was poor, confirming the observations made in the mineralogical examination. As with the quartz ore testing, the mass pull achieved around 10 times higher than what would be achieved in an operating FINAL – Rev 0 – 22 Oct 2010 AMEC Page 93 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report plant and therefore would be overstating the true recovery further. It is recommended that no further consideration be given for gravity circuit recovery from the heap leach material. 16.2.2.5 Grind Sensitivity Samples of the Heap Leach Bulk Composite were processed in a cyanide leach to determine the influence of grind size upon precious metal recovery. A total of four grind sizes were tested, these being P80 150 µm, 106 µm, 75 µm and 53 µm, with each ground sample being leached in 1000 ppm NaCN with >15 ppm oxygen in solution. Table 16.9 summarises the results obtained from testing. Table 16.9 Heap Leach Grind vs Leach Recovery Results Gold Grind P80 (µm) Residue Head Grade Grade (g/t) Silver Recovery (%) Residue Head Grade Grade (g/t) Recovery (%) 150 0.95 1.90 50.0 8.00 12.24 34.6 106 0.81 1.96 58.6 8.00 12.70 36.9 75 0.77 2.03 62.0 7.00 11.70 39.9 53 0.61 2.01 69.7 7.00 12.10 42.0 The results indicated a definite improvement in both gold and silver extraction with finer grind. Economic analysis was performed to determine the optimal grind size, based upon the following assumptions: • BBWi of 12.0 kWh/t • Gold price of $650/oz • Power cost of $100/MWh • Capital cost of $1.2 M/MW installed grinding power • A discount rate of 12.0%. The results of the net present value (NPV) determination are given graphically in Figure 16.1 and demonstrate that the overall economics favoured the finer grind; it is possible that grinds finer than 53 µm may have merit. It is recommended that this analysis be reviewed during any subsequent feasibility study. 16.2.2.6 Influence of Lead Nitrate Comparative leach tests were undertaken to study the influence of lead nitrate upon leach recovery. This was trialled at two grind sizes (P80 75 µm and P80 53 µm) and with both air and oxygen addition. The results, presented in Table 16.10, indicated that the addition of lead nitrate did not appear to significantly change the leach recovery obtained. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 94 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 16.1 Heap Leach Grind Size NPV Trend 60,000 NPV ('000 USD) 50,000 40,000 30,000 20,000 10,000 0 50 70 90 110 130 150 170 Grindsize (μ m) Table 16.10 Lead Nitrate Addition Results Gold Silver Lead Nitrate Addition (mL) Residue Grade Air 15 0.49 1.80 72.8 7.00 11.9 41.2 Air 75 0.46 1.71 73.1 8.00 12.9 37.8 75 Air 150 0.68 1.97 65.5 7.00 11.7 40.3 53 Oxygen 15 0.73 2.04 64.3 7.50 12.5 40.0 53 Oxygen 75 0.72 2.08 65.3 8.25 13.4 38.4 53 Oxygen 150 0.66 1.96 66.3 7.50 12.7 40.8 Grind P80 (µm) Gas Sparged 75 75 16.2.2.7 Head Grade (g/t) Recovery (%) Residue Grade Head Grade (g/t) Recovery (%) Influence of Pre-Aeration Pre-aeration was investigated on two tests to determine if this improved metal extraction in a cyanide leach. Pre-aeration times of 2 hours and 4 hours were trialled, followed by a 24 hour leach. The results in Table 16.11 indicate that pre-aeration had no appreciable influence upon the leach performance on heap leach material. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 95 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 16.11 Pre-Aeration Testing Results Gold Head Grade Pre-aeration Time (hours) Residue Grade 2 0.74 2.04 63.9 4 0.71 2.02 65.0 (g/t) 16.2.2.8 Recovery (%) Silver Residue Grade Head Grade Recovery (%) 7.50 12.4 39.3 8.00 13.1 38.9 (g/t) Heap Leach Variability Leaching Tests Twelve heap leach variability tests were conducted on selected individual samples of the heap leach material. Each test was conducted at a grind of P80 53 µm, oxygen sparged, at a cyanide level of 1000 ppm over 48 hours. The results are presented in Table 16.12. Table 16.12 Heap Leach Variability Testing Results Gold Heap Leach Sample Number Residue Grade Head Grade (g/t) Recovery (%) Silver Head Grade Residue Grade (g/t) Recovery (%) 6139 0.40 1.07 62.6 5.50 13.15 58.2 6351 0.32 0.95 66.3 2.50 9.08 72.5 6487 0.61 1.62 62.4 20.5 26.3 21.9 6556 0.78 3.03 74.3 6.25 8.41 25.7 6742 1.17 5.00 76.6 4.50 6.31 28.7 6782 0.55 2.38 76.9 3.00 4.40 31.8 6856 0.50 6.81 92.7 5.00 6.39 92.2 6993 0.54 2.35 77.2 3.00 4.70 36.0 7492 0.65 1.23 46.9 4.00 5.16 22.4 7632 0.76 2.18 65.1 11.5 19.8 41.8 7696 1.55 2.83 45.2 4.50 6.71 32.9 7706 0.99 1.74 43.0 5.50 8.05 31.7 Results of the variability testing gave a range of gold recoveries between 43.0% and 92.7%, averaging 65.8%, being similar to that of the Heap Leach Bulk Composite. Examination of the leach profiles demonstrate rapid dissolution of both gold and silver, with almost all being achieved in the first two hours, suggesting that shorter leach residence times could be utilised in the final flowsheet design. 16.2.2.9 Heap Leach Residue Reagent Consumption A review of the leaching test results was undertaken to estimate the reagent consumptions for the Heap Leach Residue to be utilised in the processing plant engineering design. As with the Quartz Ore, leaching dissolution was seen to be rapid and as such, the 24 hour reagent consumptions were chosen as the basis for plant design. For the Heap Leach Residue, the following reagent consumptions were estimated: FINAL – Rev 0 – 22 Oct 2010 AMEC Page 96 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report • NaCN, 1.38 kg/t. • Quicklime, 0.90 kg/t. The above results were from an arithmetic average of 18 leaching tests comprising both the composite and variability samples. 16.2.3 Metallurgical Gold Recovery for the CIL Economic Model (Includes Operating Cost Adjustments for Acidic SBR Material) Metallurgical gold recoveries for mining inventory material are defined for the economic model. In most cases these are based on metallurgical testwork, but some mining inventory categories did not have metallurgical testwork conducted to allow more accurate estimation of gold recoveries in a CIL Processing facility. These gold recoveries were then estimated based on previous Heap Leach results and CIL testwork. These methods used and results are discussed as follows: Kamoeb Quartz Material Metallurgical testwork on Kamoeb Quartz ore stockpiles gave a gold recovery of 92.9%. For the purposes of the economic model, 93% gold recovery was used. Further metallurgical testwork is recommended to confirm this recovery for the expanded Kamoeb mining inventory. Heap Leach Residue (Not Sampled) The mining inventory for Heap Leach Residue fell into to three categories: • Heap Residue sampled and tested supported by fire assays • Heap Residue already placed but not sampled or tested. Cyanide soluble assay data and metal balances are available, but cyanide soluble grade underestimates total gold content • Heap Residue not-yet-placed, ie. material identified for heap leaching before 2013, with predicted grades based on cyanide soluble assays and metal balance. Relative quantities of each material type are indicated in Table 16.13 Table 16.13 Breakdown of Heap Leach Mining Inventory Resources Material Resource Resource Assay Tonnage Grade Au Recovery Metal Category Estimate Classification Method (t) (g/t) (%) (kg) 1 Pre-2008 M+I Fire Assay 6 703 000 1.9 65 8 338 1 Pre-2008 Inferred Fire Assay 1 178 000 2.1 65 1 582 2 2008/09 M+I CN Soluble 488 000 0.9 88 382 2 2008/09 Inferred CN Soluble 1 329 000 1.4 88 1 662 3 2010 Inferred CN Soluble 2 550 000 0.9 88 1 949 12 248 000 1.6 70 13 914 Total For category #1 above, gold recoveries were based on metallurgical testwork, which indicated average recovery of 66%. For the purposes of economic modelling 65% gold recovery has been used. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 97 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report For categories 2 and 3 above, the gold recovery is based on the metallurgical testwork for Category #1, but adjusted to compensate for gold not accounted for in the cyanide soluble assay method. For the testwork on samples for category #1 inventory, it was found that the cyanide soluble assay method accounted for only 74% of the gold in the samples used for the metallurgical testwork composite sample. Therefore the gold recovered from the metallurgical testwork represented a higher component of the cyanide soluble head grade than the fire assay head grade used for the economic modelling. To adjust for this difference, the gold recovery for fire assay (category #1) material was adjusted by 1/0.74 to reflect the expected recovery from the cyanide soluble gold component of the material. As a result the CIL gold recoveries for the Categories differ: • Category #1: Gold recovery is 65% of Fire assay gold • Category #2: Gold recovery is 88% of cyanide soluble assay gold • Category #3: Gold recovery is 88% of Cyanide soluble assay gold For the economic model the weighted average grade and recovery were used to best show how the material will be reclaimed. Further studies are required to confirm the method used for the recoveries. Metallurgical testwork will indicate gold recovery for CIL from feed sources post-2009 to establish an ultimate CIL tail grade and the gold recovery differential, (contribution) of the CIL Plant over the heap leach process. Acidic SBR Material There have been no column leach tests or CIL leach tests on Acidic SBR. However, it represents only a small proportion of the CIL phase mining inventory. Gold recovery was estimated based on results from CIL testwork on the Heap Residue. Heap Residue was originally from SBR material mined in the early years of the project life. The primary difference between SBR and Acidic SBR is the presence of acidic sulfates in the later. It has been shown in heap leach tests that similar recoveries to SBR can be achieved with Acidic SBR if the ore is washed first to remove acidic compounds. The non-washable Acidic SBR that forms a part of the CIL Plant feed inventory in the economic model in this report, is highly sulfated to the point that the material is difficult to wash. However, this highly sulfated (oxidised) state also means that gold locked the original sulfides may be more readily available to cyanidation. CIL processing alleviates the washing issue. It is planned to feed the Acidic SBR as a small component of the plant feed to minimise any detrimental effects of the fine sulfates on plant control. 16.2.3.1 Gold Recovery from Acidic SBR material Based on CIL tail grades and original mined grades that averaged 12.5 gAu/t for SBR material, (adjusted to reflect fire assay gold for the years up to 2007) CIL Gold recovery for the original material is expected to be 94.4%, given a CIL test tail grade of 0.70 g/t Au. The gold recovery for Acidic SBR was reduced to 92% for the economic model to compensate for potential sulfide gold. The CIL Plant is expected to recover 95% of the gold from other SBR material. Both of these assumptions will require confirmation by future testwork. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 98 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 16.2.3.2 Operating Cost for the Acidic SBR Acidic SBR consumes substantially more lime than Oxide SBR. It also consumes more Cyanide due to an elevated soluble copper content. Previous heap leaching of washable Acidic SBR had lime consumption of 10 kg/t of ore and Cyanide consumption of 2.5 kg per tonne of ore. It is expected that non-washable SBR is as acidic so the same reagent consumptions are used. Table 16.14 shows the typical lime and cyanide consumptions for heap leach residue (from testwork), and the associated assumptions for Acidic SBR material. Table 16.14 Reagent Consumption and Costs for Heap Leach and Acidic SBR Material Heap Leach Residue Reagent Consumption Price Acidic SBR Cost Consumption Price Cost (kg/t) ($/kg) ($/t) (kg/t) ($/kg) ($/t) Lime 0.9 0.27 0.24 10.00 0.27 2.65 NaCN 1.38 2.11 2.91 2.50 2.11 5.28 Total 3.15 7.93 The difference between the two material type reagent consumption costs is $4.78/t of plant feed material. For ease of modelling, an additional $5/t was added to the base operating cost for Acidic SBR plant feed. 16.3 VMS CONCENTRATOR TESTWORK 16.3.1 Introduction A limited metallurgical testwork program was developed and undertaken by SGS Canada Inc. (SGS), using two ore composites. This work included sample preparation, head sample chemical analysis, mineralogical analysis, flotation testing, cyanidation leach testing and product characterisation testwork. Testwork to date has not included any comminution, thickening, or filtration work. Equipment sizing in these areas is, therefore, based on assumed parameters and AMEC’s experience from other projects. 16.3.2 Sample Selection Three composites were prepared from the samples received: • Composite 1 – HASS-D207 From Hassai South supergene zone • Composite 2 – HASS-D214A From Hassai South Primary zone • Composite 3 – HASS-D221 From Hassai South Primary zone Testwork was performed on Composites 1 and 3 only. A sub-sample from each of the three composites was submitted for head assay analyses, with results presented in Table 16.15. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 99 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 16.15 Head Sample Chemical Analysis Results Element Cu Unit % Au g/t 7.44 1.94 Pb Zn % % 0.080 1.13 0.032 0.56 S % 40.0 43.0 = % 38.1 39.8 S Composite 1 7.62 Composite 3 1.91 Additionally, a sub-sample at a grind size of 80% passing 100 µm of Composite 1 and Composite 3 was submitted for high-definition QEMSCAN mineralogical characterisation, with results as shown in Table 16.16. Table 16.16 Head Sample Mineral Distribution (% Mass) Mineral Pyrite Pyrrhotite Chalcopyrite 2º Cu Sulphides Other Cu Minerals Sphalerite Other Zn Minerals Galena Molybdenite Arsenopyrite Other Sulphides Silicates Mg-Chlorite Oxides Carbonates Sulphates Apatite Others Total Composite 1 Composite 3 55.4 0.1 22.8 0.6 0.0 1.8 0.0 0.1 0.0 0.0 0.0 16.1 0.3 0.7 2.2 0.0 100.0 80.0 0.0 6.2 0.1 0.0 1.2 0.0 0.0 0.0 0.0 0.0 7.0 1.6 0.2 2.8 0.7 0.1 0.0 100.0 Pyrite (55%) and chalcopyrite (23%) are the predominant sulphide minerals in Composite 1, accounting for 78% of the mineral mass. Similarly for Composite 3, pyrite and chalcopyrite account for 80% and 6% of the mineral mass, respectively. Silicates are the most abundant non-sulphide gangue mineral, accounting for 7% of the mineral mass. In Composite 1, 77.5% chalcopyrite is free and liberated, while 18.4% is associated with pyrite. A further 3.2% is either associated with silicates or in complex association. These are unlikely to be recoverable by sulphide flotation. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 100 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report In Composite 3, 88% of pyrite is liberated, and approximately 10% associated with chalcopyrite. In addition, 72% chalcopyrite is liberated, while 25% is associated with pyrite. The majority of pyrite (97%) is liberated, and 2% is associated with chalcopyrite. Chalcopyrite liberation is likely to improve with a finer primary grind in both composites as 18-25% of chalcopyrite is associated with pyrite. Due to the high pyrite content of both composites, a primary grind targeting full chalcopyrite liberation may be necessary to improve copper recovery and flotation kinetics. 16.3.3 Flotation Testwork A total of 13 rougher kinetics and batch cleaner flotation tests were performed on Composite 1 and Composite 3. In addition, a locked cycle test was performed on each of the two composites as per the flow sheet shown in Figure 16.2. Figure 16.2 Locked Cycle Testwork Flow Sheet FINAL – Rev 0 – 22 Oct 2010 AMEC Page 101 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 16.3.3.1 Flotation Kinetic Testwork Two rougher kinetics flotation tests were performed on Composite 1 and Composite 3. A summary of test conditions and results are shown in Table 16.17. Table 16.17 Rougher Flotation Kinetic Results Composite Conditions Comp 1 Concentrate Froth Time Copper Gold (min) Recovery Recovery (%) (%) Test F1 Ro Con 1 2 27.5 55.2 99 µm Ro Con 1-2 4 46.1 70.6 pH=9.5-10.0 Ro Con 1-3 7 61.8 78.3 37.5 g/t Aero 238 Ro Con 1-4 10 64.2 80.4 Ro Con 1-5 15 71.6 82.9 Comp 1 Ro Tail - 28.4 17.1 Test F3 Ro Con 1 4 37.2 55.6 69 µm Ro Con 1-2 8 62.0 67.3 pH=10.5 Ro Con 1-3 14 86.8 75.1 50 g/t Aero 238 Ro Con 1-4 20 94.6 78.8 Ro Con 1-5 26 97.3 83.5 Ro Tail - 2.7 16.5 Test F2 Ro Con 1 1 8.7 9.0 98 µm Ro Con 1-2 3 16.7 14.5 pH=9.0-10.0 Ro Con 1-3 6 22.2 17.1 87.5 g/t Aero 238 Ro Con 1-4 11 49.9 26.4 Ro Con 1-5 15 68.1 42.8 Comp 3 Ro Tail - 31.9 57.2 Test F4 Ro Con 1 4 44.0 27.8 68 µm Ro Con 1-2 8 76.4 36.6 pH=10.5 Ro Con 1-3 14 86.7 42.5 50 g/t Aero 238 Ro Con 1-4 20 90.1 46.9 Ro Con 1-5 26 90.9 49.3 Ro Tail - 9.1 50.7 Comp 3 The main conclusions from the testwork are: Composite 1 • A cumulative rougher concentrate (Test F3) was produced, grading 15.8% Cu at 97% Cu recovery and 12.1 g/t Au at 84% Au recovery. • Rougher concentrate produced in Test F3 had a significantly improved copper grade and recovery when compared to the concentrate produced in Test F1, probably due to improved liberation, as the primary grind size was reduced from 99 µm to 69 µm. The improved liberation resulted in greater copper selectivity and pyrite depression. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 102 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report • Mass pull in both tests was high, at approximately 45%. Pyrite recovery was significant at 40-45%. • The gold recovery in the cumulative rougher concentrate was similar in both tests at approximately 83%, indicating that variation in primary grinding between a P80 of 99 µm to 69 µm had a minimal impact on gold recovery. Composite 3 • Test F4 produced the best results overall, with a cumulative rougher concentrate grading 11.5% Cu at 91% Cu recovery and 6.43 g/t Au at 49.3% recovery. • Copper and gold recoveries observed in Test F4 were significantly higher than those produced in Test F2, probably due to improved liberation from the primary grind size reduction from P80 of 98 µm to 68 µm. • Mass pull decreased significantly from 26% in Test F2 to 15% in Test F4 with corresponding pyrite recovery of 27% and 11%, respectively. • Metal recoveries, specifically gold, into cumulative rougher concentrate were significantly lower for Composite 3 in comparison to Composite 1. Approximately half of the gold reports to the rougher tailings and is likely to be associated with pyrite. 16.3.3.2 Batch Cleaner Testwork To assess the copper and gold grade-recovery relationships, batch cleaner tests were performed for Composite 1 and Composite 3. Various flotation conditions including pH, collector dosage, froth time, regrinding and flotation feed (gravity tailings) were investigated. Composite 1 Results • Test F5 investigated modified rougher conditions compared to Test F3. Reduced froth time, lower collector dosage and a higher pH were tested with the objective of reducing pyrite recovery. Although the rougher concentrate in Test F5 recovered only 28% of the pyrite, copper recovery was also reduced to 87%. • In addition, excessive stage-recovery losses of 31% and 16% were noted for copper and gold respectively, from the first cleaner to the rougher circuit. This is an indication of insufficient collector addition and froth time. • In Test F6, flotation conditions were altered to target greater metal recoveries. Rougher conditions in Test F6 followed those of Test F3. Collector dosage was increased to 40 g/t in the cleanercleaner scavenger circuit and the regrind size was decreased to P80 51 µm. The metal recoveries were significantly increased in the rougher and cleaner circuit as a result of these changes. Table 16.18 summarises the conditions and results for Composite 1. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 103 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 16.18 Cleaner Flotation Results for Composite 1 Conditions Concentrate Mass Assay Recovery Recovery Cu Au Cu Au (%) (%) (g/t) (%) (%) Rougher con 28.0 22.9 18.7 86.7 76.5 Ro. 68 µm, pH=10.0-10.5 1 cleaner con st 12.9 31.9 32.3 55.4 60.5 40 g/t Aero 238 2nd cleaner con 12.1 33.1 33.4 54.1 58.9 Test F5 Froth time 20 min rd 3 cleaner con 11.7 33.4 33.9 52.6 57.6 Cleaner scav. con 6.1 28.7 9.7 23.5 8.6 20 g/t Aero 238, Froth time 23 min Rougher tail 72.0 1.4 2.2 13.3 23.5 Test F6 Rougher con 37.9 19.4 15.5 97.2 81.8 st 23.7 29.1 22.0 91.4 72.9 Clean. 30 µm, pH= 11.3-11.8 Ro 65 µm, pH=10 1 cleaner con 50 g/t Aero 238 2 cleaner con 22.8 30.2 22.5 91.0 71.5 Froth time 26 min 3rd cleaner con 22.0 31.0 22.6 90.1 69.3 Clean. 51 µm, pH= 11.0-11.5 nd Cleaner scav. con 3.3 9.9 6.6 4.3 3.0 Rougher tail 62.1 0.3 2.1 2.8 18.2 40 g/t Aero 238, Froth time 24 min Rougher con 46.3 16.1 13.7 98.7 86.8 Ro 55 µm, pH=10.0-10.5 1st cleaner con 25.9 26.4 20.9 90.6 73.7 75 g/t Aero 238 2 cleaner con 24.5 27.7 21.6 89.8 72.0 Test F11 Froth time 26 min nd rd 3 cleaner con 16.8 31.1 26.3 69.4 60.4 Cleaner scav. con 5.6 8.0 7.1 5.9 5.4 40 g/t Aero 238, Froth time 20.5 min Rougher tail 53.7 0.2 1.8 1.3 13.2 Test F12 Rougher con 44.0 16.9 14.6 98.0 85.0 Ro 65 µm, pH=10.0-10.5 1st cleaner con 25.6 27.5 22.5 92.9 76.3 75 g/t Aero 238 2nd cleaner con 23.2 29.8 23.7 91.4 73.2 Froth time 21 min 3 cleaner con 17.7 32.5 27.6 75.7 64.7 Clean. 55 µm, pH= 11.3-11.8 Clean. 30 µm, pH= 11.3-11.8 45 g/t Aero 238, Froth time 21 min rd Cleaner scav. con - - - - - Rougher tail 56.0 0.3 2.0 2.0 15.0 To further investigate the sensitivities of rougher performance to collector dosage, froth time, and pH, Tests F11 and F12 were conducted at increased collector dosage and with varying froth time. The following observations are made: • Comparing Tests F11 and F6, the rougher concentrate recorded increased gold recovery at the expense of lower concentrate grade. The additional gold recovered is likely to be embedded within larger pyrite grains. Without liberation, these gold particles are likely to be rejected with the pyrite in the cleaner circuit. • In comparing Tests F6, F11 and F12, it was found that increased collector dosage caused excessive pyrite deportment to the rougher concentrate, increasing reliance on the cleaner circuit for concentrate upgrading. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 104 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report • In addition, at a primary grind size of 65 µm, some of the gold recovered in the rougher concentrate is likely to be in the form of fine inclusions embedded within pyrite grains which would be lost in the middling streams in the cleaner circuit, or would be rejected into the cleaner scavenger tailings. Therefore, it is important to optimise the rougher collector dosage, froth time and pulp pH to ensure that the gold-embedded pyrite is directed to the correct stream flow. An alternative option is finer primary grinding which, combined with optimised flotation conditions, could provide better selectivity in the rougher circuit. Composite 3 Results Tests F7 and F8 investigate the effect of pH. The pulp pH for Test F7 was 10-10.5 and 11.0-11.5 in the rougher and cleaner circuits, respectively. Test F8 was conducted at rougher circuit pH of 10.5 and cleaner circuit pH of 11.3-11.8. The froth time in the cleaner circuit of Test F8 (second and third cleaner) was also shortened. Table 16.19 summarises the conditions and results for Composite 3. Table 16.19 Cleaner Flotation Results for Composite 3 Conditions Test F7 Ro. 70 µm, pH=10.0-10.5 50 g/t Aero 238, Froth time 26 min Clean. 29 µm, pH= 11-11.5 40 g/t Aero 238, Froth time 21.5 Test F8 Ro 69 µm, pH=10.5 50 g/t Aero 238, Froth time 26 min Clean. 29 µm, pH= 11.3-11.8 40 g/t Aero 238, Froth time 19.5 min Test F13 Ro 64 µm, pH=10.5 57.5 g/t Aero 238, Froth time 26 min Clean. 64 µm, pH= 11.3-11.8 40 g/t Aero 238, Froth time 18 min Concentrate Mass Assay Recovery Recovery Cu Au Cu Au (%) (%) (g/t) (%) (%) Rougher con 30.0 5.9 4.3 95.4 59.6 st 12.4 13.6 7.7 91.7 44.5 1 cleaner con nd 2 cleaner con 10.1 16.4 8.7 90.5 40.9 3rd cleaner con 8.1 20.0 9.8 88.4 37.0 Cleaner scav. con 2.9 0.8 2.1 1.2 2.8 Rougher tail 70.0 0.1 1.2 4.6 40.4 Rougher con 16.4 10.8 6.0 93.5 50.5 st 1 cleaner con 8.9 19.2 9.5 91.1 43.5 cleaner con 7.7 22.1 10.4 90.4 41.1 3 cleaner con 6.7 25.0 11.2 89.0 38.4 nd 2 rd Cleaner scav. con 1.1 1.8 2.8 1.1 1.6 Rougher tail 83.6 0.1 1.2 6.5 49.5 Rougher con 27.4 6.5 4.2 93.0 58.8 1st cleaner con 16.0 10.5 5.6 88.5 46.2 2nd cleaner con 10.5 15.8 7.5 86.4 40.2 rd 3 cleaner con 8.6 18.9 8.4 85.4 37.1 Cleaner scav con 4.1 1.6 2.8 3.4 5.9 Rougher tail 72.6 0.2 1.1 7.0 41.2 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 105 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report The main conclusions for this testwork are: • In comparison to Test F7, the rougher concentrate of Test F8 had similar copper recovery (93%), but a lower gold recovery of 50%. • Mass pull was significantly reduced to 16% and pyrite recovery decreased to 13%. It is likely gold is in the form of fine inclusions embedded within larger pyrite grains, thus the lower gold recovery is associated with the lower pyrite recovery. Consequently, gold recovery was increased from 37% in Test F7 to 48% in Test F8 in the pyrite scavenger concentrate. • The third cleaner concentrate produced in Test F8 graded 25.0% Cu at 89% Cu recovery and 11.2 g/t Au at 38% Au recovery. This was a higher grade at similar recoveries to Test F7. The combined effects of higher pulp pH in the rougher and cleaner circuits as well as the shorter froth time in the cleaner circuit are the likely cause. Flotation conditions in Test F13 were similar to Test F8, but without regrinding. The third cleaner concentrate, grading 18.9% Cu at 85% Cu recovery and 8.4 g/t Au at 37% Au recovery, had lower copper and gold grades as well as lower copper recovery to the final cleaner concentrate compared to Test F8. The lower metal grades found were a result of greater gangue and pyrite recovery, likely related with the extent of liberation. 16.3.3.3 Locked Cycle Testwork A singled locked cycle test was conducted for each composite. Conditions of the test where similar to Tests F6 and F8 for Composite 1 and 3, respectively. The results are summarised in Table 16.20. Table 16.20 Locked Cycle Testwork Results Sample Composite 1 Stage Assay Recovery on Stage Recovery Cu Au % Stage Cu (% Stage) (%) (g/t) (%) (%) Rougher con 45.4 15 13.6 93.2 84.4 Rougher tail 54.6 0.9 2.1 6.8 15.6 3rd cleaner con 86.7 30.2 25.1 93.2 94.2 rd 3 cleaner tail 13.3 14.7 11.0 6.8 5.8 Clean scav. con 19.5 16.2 7.9 73.0 34.0 Clean scav. tail 80.5 1.8 3.4 27.0 66.0 Overall recovery 21.3 - - 87.3 73.1 Rougher con 14.6 11.5 6.10 88.1 45.4 Rougher tail 85.4 0.3 1.26 11.9 54.6 rd 91.1 25.1 11.6 99.1 95.5 3 cleaner con Composite 3 Mass rd 3 cleaner tail 8.9 2.3 5.6 0.9 4.5 Clean scav. con 15.2 2.3 3.6 35.9 28.4 Clean scav. tail 84.8 0.74 1.7 64.1 71.6 Overall recovery 6.5 - - 84.9 38.3 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 106 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Under the selected conditions for Composite 1, a final cleaner concentrate grading 30.2% Cu and 25.1 g/t Au was achieved at 87% Cu and 73% Au recovery. For Composite 3, final cleaner concentrate grading 25.1% Cu and 11.6 g/t Au was achieved at 85% Cu and 38% Au recovery. Mass pulls in the rougher circuit for both composites were not consistent throughout the locked cycle test, with greater copper losses to the pyrite scavenger circuit when compared to batch test results overall. Copper recovery in the final cleaner concentrate was lower than expected as a result. 16.3.4 Tailings Cyanide Leaching Testwork The primary purpose of these tests was to approximately determine the potential for gold recovery. Parameters evaluated included leach retention time, feed particle size, and cyanide concentration. Testwork conditions and results are presented in Table 16.21. Table 16.21 Tailings Cyanide Leach Testwork Composite 1 Test Number Composite 3 CN-1 CN-3 CN-5 CN-7 CN-2 CN-4 CN-6 CN-8 Test feed F-6 LCT-1 LCT-1 LCT-1 F-7 LCT-1 LCT-1 LCT-1 P80, mm 65 69 12 69 ~70 67 11 67 NaCN, g/L 2 5 2 2 2 5 2 2 Slurry, % w/w 25 20 20 20 25 20 20 20 NaCN 5.68 14.83 14.69 8.69 3.31 19.20 8.94 6.09 CaO 1.22 2.19 1.55 3.61 0.71 1.15 1.92 1.92 5h 35 - 51 36 29 - 50 32 8h - 40 - - - 39 - - Cons. kg/t Au Recovery, % Cu Recovery, % Residue grade Head grade 24 h 39 44 61 41 34 41 57 38 48 h 42.3 46.9 66.7 42.3 37.1 45.4 61.4 38.8 72 h - - - 44.6 - - - 36.5 96 h - - - 45.4 - - - 38.1 48 h 24.7 26.0 57.3 - 44.1 30.8 57.4 - 96 h - - - 30.9 - - - 41.4 Au, g/t 1.58 1.47 0.98 1.54 0.79 0.77 0.51 0.88 Cu, % 0.28 0.47 0.26 0.44 0.06 0.20 0.12 0.18 Au, g/t 2.73 2.77 2.93 2.82 1.25 1.41 1.32 1.42 0.37 0.64 0.61 0.64 0.11 0.29 0.28 0.31 Cu, % As a general comment, the recoveries were low and cyanide consumption high. consumption appears to be due to the high rates of addition. 16.3.5 Metallurgical Projection and Metallurgical Parameters for Design 16.3.5.1 Flotation High cyanide The differences in grades between the composites used in the testwork and those used as head grades for the study (see Table 16.22), necessitated estimation of the recoveries for use in the design. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 107 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 16.22 Testwork and Design Ore Grades Hassai South Supergene Hassai South Primary (HSS) Grades Hadal Awatib (HSP) (HA) Composite 1 Design Composite 3 Design No Sample Design Copper grade, % 7.62 2.75 1.91 1.37 - 0.99 Gold grade, g/t 7.44 2.29 1.94 1.49 - 1.18 Zinc grade, % 1.13 0.68 0.56 0.68 - 0.68 The design recoveries for HSP were estimated assuming similar concentrate and tailing grades as those reported in the testwork (Test F8 and the Locked Cycle Test). Recoveries and grades for HSS were assumed, in the expectation that this type of ore will behave in similar way to the composite used for the testwork (Composite 1), but assuming lower concentrate grades to account for the lower ore design grade. For HA, recoveries and grades were assumed considering that this type of ore will behave in a similar way to HSP, but assuming lower concentrate grades to account for the lower ore design grade (except for final concentrate that was assumed to be the same as for HSP). Design concentrate and tailing grades assumed for the different flotation stages are presented in Table 16.23. The calculated stage recoveries are presented in Table 16.24. Table 16.23 Design Concentrate and Tailing Grades Hassai South Super. Stream Hassai South Primary Hadal Awatib Cu Au Cu Au Cu Au (%) (g/t) (%) (g/t) (%) (g/t) Rougher concentrate 15.0 Calc. 10.8 Calc. 8.0 Calc. Cleaner 1 concentrate 25.0 18.0 19.0 9.0 15.0 6.0 Cleaner 1 tail grade 2.0 Calc. 1.0 Calc. 0.8 Calc Cleaner scav. concentrate 12.0 5.0 2.35 3.6 1.5 2.0 Cleaner scav. tail 1.5 Calc. 0.75 Calc. 0.7 Calc. Cleaner 2 concentrate 32.0 22.0 25.1 11.0 25.1 10.0 Cleaner 2 tail 8.0 Calc. 2.26 Calc. 2.0 Calc. Table 16.24 Design Calculated Stage Recoveries Ore Type Hassai South Super. Hassai South Primary Hadal Awatib Stage Mass Cu Au Mass Cu Au Mass Cu Au Rougher 15.8 86.2 72.6 11.9 93.6 50.5 11.3 91.0 49.2 Cleaner 1 51.6 93.0 92.5 44.7 93.9 69.8 39.9 92.6 57.5 Cleaner scavenger 4.8 28.6 15.3 15.6 36.7 17.8 12.5 23.4 8.5 Cleaner 2 70.8 90.7 86.6 73.3 96.8 89.6 56.3 94.2 93.8 Global 7.0 81.0 67.0 4.9 90.0 36.0 3.4 85.0 29.0 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 108 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Other metallurgical design parameters derived from the testwork are as follows: • Residence time required for the flotation area is based on laboratory flotation times with a residence time scale-up factor. Based on kinetic testwork data available for rougher flotation, it has been assumed that the rougher recoveries can be achieved with a design froth time of 40 minutes. Residence times of 20, 15 and 15 minutes, respectively, were selected for cleaner 1, cleaner scavenger and cleaner 2 stages. • The circuit design includes two cleaner stages only. Froth time for cleaner 2 is scaled up from the total testwork time of cleaner 2 and cleaner 3 combined. • Collector dosage of 50, 20, 10 and 10 g/t, has been selected for the rougher, cleaner 1, cleaner scavenger and cleaner 2 stages, respectively. • Based on the batch and locked cycle testwork performed for both composites, a P80 of 69 µm for grinding and 30 µm regrinding, has been selected. 16.4 CONCENTRATOR FLOW SHEET DEVELOPMENT The concentrator facility includes all ore processing facilities from primary crushing to concentrate storage, and pumping and storage of process tailings. The plant design is based on a single processing train configuration, with a nameplate capacity of 5 Mt/a. The process flow sheet is shown in Figure 16.3. The key unit steps in the process flow sheet are as follows: • Target primary grind to P80 69 µm in the crushing/grinding circuit • Conventional rougher flotation • Regrind of rougher concentrate to a P80 30 µm • A three stage flotation cleaning circuit • Concentrate thickening, filtration and transport • Tailings thickening and disposal. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 109 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 16.3 5 Mt/a Block Flow Diagram FINAL – Rev 0 – 22 Oct 2010 AMEC Page 110 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 16.4.1 Process Design Criteria and Mass Balance Due to limited testwork undertaken to-date, the design criteria are in many instances derived from similar plants, benchmarks and assumptions. Areas in which assumptions have been made will require verification by testwork in future studies. Mass balances were developed for the three main ore types, Hassai South Supergene (HSS), Hassai South Primary (HSP) and Hadai Awatib (HA). Variations in flow rate as high as 108% for the same stream are observed when comparing the three mass balances (Table 16.25). The maximum differences result from the HSS ore type, but limited tonnage of this ore type exist and the HSS mass balance flows have not been used for equipment sizing. Table 16.25 3 Slurry Flows (m /h) and Differences for the Different Ore Types Stream HSS HSP HA Diff. Max HSP and HA and Min. Diff. (%) (%) 5 Rougher concentrate 226.9 170.4 161.7 40 Total feed to cleaner 1 408.2 322.4 316.5 29 2 Cleaner 1 tail 253.8 217.9 222.9 16 2 Cleaner 1 concentrate 154.4 104.5 93.6 65 12 85.2 55.1 65.3 55 16 Cleaner 2 concentrate 109.6 76.7 52.8 108 45 Cleaner scavenger tail 242.5 184.6 193.8 31 5 Cleaner scavenger con 11.4 33.3 29.1 193 14 Concentrate thickener feed 156.5 109.6 75.4 108 45 Concentrate thickener overflow 123.9 86.7 59.7 108 45 Concentrate thickener underflow 32.6 22.8 15.7 108 45 Feed to concentrate filter 37.2 26.0 17.9 108 45 Concentrate cake 16.3 11.4 7.8 108 45 1284.1 1282.8 1300.6 1 1 Cleaner 2 tail Total Tailings to Tails Thickener 16.4.2 Comminution Circuit In the absence of testwork data, a Bond ball mill work index (BWI) of 15 kWh/t was assumed for the grinding circuit design. Due to the high pyrite content of the ore, and subsequent high specific gravity, the power required for a single mill would be 11 MW, hence the circuit was split in two stages comprising a SAG and a ball mill in combination. 16.4.3 Flotation Circuit The flotation circuit flow sheet developed consists of three steps including rougher, cleaner 1, cleaner scavenger and cleaner 2 (recleaner) stages. Rougher concentrate is pumped to a regrind circuit prior to reporting to the cleaner circuit. Cleaner 1 concentrate feeds the cleaner 2 stage, and cleaner 2 concentrate reports to the concentrate thickener. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 111 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Cleaner 1 tail feeds the cleaner scavenger stage, and cleaner scavenger concentrate is recycled to the cleaner 1 feed (optionally to the regrind stage). Cleaner scavenger tail is combined with the rougher tail and discarded. The cleaner stage is separated into cleaner, and cleaner scavenger banks to produce a separate cleaner circuit tail, rather than recycle cleaner tail to rougher or scavenger feed. This circuit has several advantages over a single cleaner bank type circuit, as it allows selection of a primary metallurgical goal for each stage. For example, maximising concentrate grade in the cleaner circuit and minimising copper and gold losses in the cleaner scavenger circuit. The flotation circuit design utilises Outotec, forced air, mechanically agitated, tank cells throughout. The design criteria residence time results in the selection of six cells in three pairs for the rougher and cleaner 1 stage, three cells for the cleaner scavenger stage and four cells in two pairs for the cleaner 2 stage. The selection of the cell volume and quantity is based on achieving the necessary residence time and the mass pull required by the design mass balance. The selected design criteria for the flotation circuit are provided in Table 16.26. Table 16.26 Flotation Design Basis Description Value Source Rougher Flotation Cell type Design residence time, min Effective volume of cell, % Number of trains Number of cells, units per train 3 Cell volume, m /cell 3 Air feed per cell, Am /min Rougher concentrate, Solids SG, t/m3 Solids in rougher concentrate, % Solids in rougher conc. after spray water, % Forced air 40 85% 2 6 100 9-20 4.7 32% 25% AMEC recommendation Assumed or estimated data AMEC recommendation AMEC recommendation AMEC recommendation Calculated Vendor data Assumed or estimated data Assumed or estimated data Assumed or estimated data Cleaner 1 Flotation Cell type Design residence time, min Effective volume of cell, % Number of trains Number of cells, units per train Cell volume, m3/cell Air feed per cell, m3/min 3 Rougher concentrate, Solids SG, t/m Solids in rougher concentrate, % Solids in rougher conc. after spray water, % Forced air 20 85% 1 6 20 3-7 4.7 30% 25% AMEC recommendation Assumed or estimated data AMEC recommendation AMEC recommendation AMEC recommendation Calculated Vendor data Assumed or estimated data Assumed or estimated data Assumed or estimated data FINAL – Rev 0 – 22 Oct 2010 AMEC Page 112 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 16.26 Description Cleaner Scavenger Flotation Cell type Design residence time, min Effective volume of cell, % Number of trains Number of cells, units per train Cell volume, m3/cell Air feed per cell, Am3/min 3 Rougher concentrate, Solids SG, t/m Solids in rougher concentrate, % Solids in rougher conc. after spray water, % Cleaner 2 Flotation Cell type Design residence time, min Effective volume of cell, % Number of trains Number of cells, units per train 3 Cell volume, m /cell 3 Air feed per cell, Am /min Rougher concentrate, Solids SG, t/m3 Solids in rougher concentrate, % Solids in rougher conc. after spray water, % Flotation Design Basis Value Source Forced air 15 85% 1 3 20 3-7 4.9 20% 20% AMEC recommendation Assumed or estimated data AMEC recommendation AMEC recommendation AMEC recommendation Calculated Vendor data Assumed or estimated data Assumed or estimated data Assumed or estimated data Forced air 15 85% 1 4 10 2-5 4.6 30% 25% AMEC recommendation Assumed or estimated data AMEC recommendation AMEC recommendation AMEC recommendation Calculated Vendor data Assumed or estimated data Assumed or estimated data Assumed or estimated data To date, no residence time optimisation testwork has been undertaken. AMEC believes that further optimisation of flotation residence time has the potential to reduce total equipment requirements. 16.4.4 Regrind For this study, a circuit based on ISAmill units was assumed, due to proven higher power efficiencies for fine grinding. The regrind was considered in open circuit, with a desliming cyclone removing fines from the regrind mill feed (assumed 50% of the rougher concentrate). Regrind mill size and installed power were established utilising information from the vendor database for high pyrite concentrates at the required feed and product particle size. The vendor data showed a maximum power value of 12.6 kWt/h (average 8.8 kWt/h). Mill sizing considered that all rougher concentrate will feed the mill (bypassing the desliming cyclone). Final selection of regrind equipment should be based on a trade-off study, considering capital and operating cost estimates for both options. It will be necessary to establish operating parameters such as regrind power efficiencies, optimum media types and media consumption for the different types of mills by conducting regrind testwork. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 113 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 16.4.5 Concentrate Handling Design criteria for concentrate thickening and filtration have been based on design parameters for a similar copper concentrate. Two Larox pressure filters were selected based on the design mass balance and the assumed design criteria. If alternative filter technology other than the selected Larox type filter is considered viable (eg horizontal plate and frame), additional appropriate testing will be required to ensure correct filter sizing. The selected design criteria are provided in Table 16.27. Table 16.27 Design Basis Concentrate Handling Description Value Source High rate AMEC recommendation Settling rate, t/m .h 0.1 Assumed or estimated data Thickener diameter, m 22 Calculated data 65% Assumed or estimated data Thickening Thickener type 2 Underflow density, % w/w Filtration Filter type Availability,% Operating hours, h/d 2 Auto. Pressure Filters AMEC recommendation 80.0% Assumed or estimated data 19.2 Assumed or estimated data Specific filtration rate, kg/m .h 360 Assumed or estimated data Cake moisture, %w/w 10% Assumed or estimated data Number of filters, units 2 AMEC recommendation 2t Maxibags Assumed or estimated data 4.6 Assumed or estimated data Concentrate delivery method Dry concentrate SG FINAL – Rev 0 – 22 Oct 2010 AMEC Page 114 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 17. MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES 17.1 GENERAL The mineral resources identified at Hassai Mine as of the end of 2009 comprise: • VMS resources – with the exception of very minor adjustments to wireframes for Hassai South, the VMS resources remain the same as those reported in La Mancha,October 2009 (Hassai South) and La Mancha, December 2009 (Hadal Awatib East). No subsequent drilling or resource modelling has been undertaken for these deposits. • Gold resources currently being mined or proposed for mining including: • − Current stockpiles − Kamoeb Quartz ore, modelled and reported in La Mancha 2008, updated after infill drilling in 2009, and subsequently depleted by on-going mining activities − Remnant in situ SBR and other oxide ores, also estimated and reported in La Mancha 2008 Tailings from past heap leaching operations. Drill tested in 2007/08 and 2009, but resources not formally included in any prior Technical Report. The resources contemplated for the Hassai Mine Envisaged Business Plan comprise: • Those included under the current heap leach operating plan through to the end of 2012, comprising: − Existing heap leach tailings up to end of 2008, drilled and estimated by Arethuse and included in the end-2009 resource statement. − Additional heap leach tailings placed in 2009, as determined by CSA − Planned heap leach tailings to be generated between 2010 and the end of 2012, as determined by CSA from the AMC mining schedule and expected recoveries. • In situ and stockpiled gold resources remaining at the end of 2012, determined from the above resources and the current AMC mining schedule. • VMS resources determined by Arethuse as included in the end-2009 resource statement. 17.2 VMS RESOURCES: HADAL AWATIB EAST AND HASSAI SOUTH 17.2.1 Geological Model Hadal Awatib East and Hassai South are the largest known VMS deposits in the Hassaï district. The sulphides in the VMS deposits were originally overlain by a thick, well-preserved oxidation profile. The sulphides have been leached in the top 60-120 m, leaving an oxide cap and a quartz-kaolinitebarite white powdery residue below (“SBR” rock) that has high gold content and very low base metal content. The oxide and SBR has been largely mined out, although limited resources remain. The host rock to the two VMS deposits is sheared green mafic schist, and there is little or no stratigraphic information available to assist with the interpretation of the VMS bodies other than the sulphide lenses themselves. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 115 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Hassai South deposit (Figure 17.1) comprises two main lenses striking east-west over approximately 1000 m and dipping 60o south, with little tectonic disruption. Towards the edges of the deposit the lens breaks up into two or three separate lenses with intervening beds of sheared volcanics. These lenses thin and pinch out to the east. Two zones exist, namely a copper-enriched supergene zone and a primary zone. The supergene zone is present at the bottom of the Hassaï South pit and extends down dip for 40-60 m. The primary massive sulphide continues down dip from the base of the supergene zone for >250 m. The current drilling campaign has intersected massive sulphide to a depth of 400 m below the pit floor, and it remains open at depth. The Hadal Awatib East deposit (Figure 17.2) has been subjected to several deformation events. The mineralisation is broadly continuous and consists of two mineralised lenses, with limited bifurcations, but the lenses are folded and faulted. The orebody trends 105/285o for 1200 m length and has a vertical extent of 150-350 m below the oxidation limit. The separation between supergene and primary mineralisation was difficult to identify during core logging, and there is no appreciable supergene enriched (Cu or Au) domain. Leaching of Zn has been observed in the top 50 m. The distinction between supergene and primary ore types, believed to be a soft limit, has been set at 50 m below the base of oxidation, but has been used for dry bulk density assignment only. While chalcopyrite stringer mineralisation has been recorded, it has not been logged systematically. No coherent zones are recognised and this style of mineralisation is not included in the mineral resources at this stage. Figure 17.1 Hassai South – Main Cu-Au Ore Bodies – Long Section South to North (100 m Grid) Actual pit Oxidation Limit Supergene Primary Actual pit Oxidation Limit Supergene Limit FINAL – Rev 0 – 22 Oct 2010 SBR Supergene Primary Supergene Limit The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 17.2 Hadal Awatib East : Surface Trace of Main Mineralised Bodies AB 4 Link CD 8 6 7 5 3 2 1 6 17.2.2 Orebody Number Main vms axis 9: all purple small mineralised bodies with limited extension Cut-off and Domain Modelling The 3D modelling, geological and block models, geostatistical analysis and grade modelling have been conducted using Surpac Vision V6.1. Geological sections have been interpreted on vertical crosssections, and wireframes were snapped to the drill holes. For the sulphide zones, a copper equivalent (CuEq) cut-off grade of 0.8% was used to guide interpretations, based on the formula CuEq % = Cu % + 0.63 x Au ppm. This relationship takes account of metal prices ($750/oz gold and $2.00/lb copper) and recoveries. The following rules were applied during interpretation: • At least 1 m mineralised interval • Maximum 2 m barren interval, with final average grade above cut-off grade • Flexibility applied to give consistent mineralised body outlines. For the oxide zone at Hadal Awatib East a 1 g/t Au cut-off was considered, which is consistent with current and historical mining. The Hadal Awatib East oxide zone has been modelled and is now being mined. At Hassai South some residual acidic high grade SBR ore remains at the bottom of the pit but the oxide zone has not been modelled. Pb and Ag are not considered economic at this stage, even as by-products. Zn is slightly higher grade and has been estimated separately. Modelling of Zn zones was possible using a 0.75% cut-off at Hadal Awatib East and 1% at Hassai South. For Hassai South, a minor revision of the wireframe was undertaken in April 2010, leading to slightly different resources from those published in 2009. This revised model is considered to be a slight improvement and has been used to develop the Mining Inventory for the scoping study. FINAL – Rev 0 – 22 Oct 2010 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 17.2.3 Down-dip Drill Holes within the Supergene Domain At Hassai South, AMC has drilled a number of holes from the bottom of the pit, many of which did not fully intersect the supergene sulphide lenses. An over-estimation is possible in some cases. The shape of the variographic ellipsoid (see variography, Section 17.2.8) shows that the variability is at a maximum in both the across-strike and down-dip directions, and that both directions should be drilled. 17.2.4 Overall Population Distribution and Top-cuts Core and RC samples have a typical and generally consistent length of 1 m and samples were composited to this length for statistical and variographic analysis, and for grade interpolation. Basic statistics for the two deposits are provided in Table 17.1. Table 17.1 Hadal Awatib East (HAE) and Hassai South (HASS) – Sample Statistics HAE Oxide zone HAE Sulphide Zone Au Cy Au FA ppm Cu% Zn (0.75%Zn envelope) Number of composites 2169 2880 3354 1609 Min Max Mean Median Variance 96.66 9.69 5.1 135.12 0 32.6 1.09 0.95 1.18 20.9 1.19 0.64 2.74 0.01 7.05 1.56 1.31 1.19 Standard Deviation 11.62 1.09 1.65 1.09 Coefficient of variation 1.2 0.99 1.39 0.7 Upper cut-off 35 g/t 7.5 g/t 6.3 to 8% 4% HASS Supergene HASS Primary AuFa g/t Cu% Zn% (Cu-Au envelope) Zn (1% Zn envelope) AuFa g/t Cu% Zn% (Cu-Au envelope) Zn (1% Zn envelope) Number of composites 398 458 458 27 765 825 825 133 Min Max Mean Variance 0.02 23.6 2.2 9.66 0.02 15.02 2.98 6.6 0 4.02 0.29 0.31 0.57 2.94 1.48 0.36 0.03 17.7 1.49 1.69 0.02 6.13 1.43 1.46 0 4.59 0.4 0.43 0.03 4.63 1.7 0.66 Standard Deviation 3.11 2.57 0.56 0.6 1.3 1.21 0.65 0.82 Coefficient of variation 1.41 0.86 1.91 0.41 0.87 0.85 1.63 0.48 Upper cut-off 13g/t No No No No No No No FINAL – Rev 0 – 22 Oct 2010 AMEC Page 118 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Top-cuts have been looked at carefully for the individual ore bodies. In most cases, a top-cut is not required, however at Hadal Awatib East, the oxide zone shows higher grade dispersion of gold, thus a 35 g/t Au top-cut has been applied, this being typical for AMC’s SBR deposits. Some high Cu and Au values need to be cut; these values most likely represent stringers in the feeder zone, which have not yet been delimited. At Hassai South, some higher Au grades in the supergene zones have been cut, to limit the number of outlier values. The comparison of statistics between raw samples, composites and composites samples with top-cuts, show very acceptable results. 17.2.5 Dry Bulk Density In Hassai South, density distribution (Table 17.2) reflects the presence of a few barren intervals (silicate) and the presence of both massive and disseminated sulphides. Despite being systematic, the density measurement spatial distribution is not yet complete. Density measurement in the supergene zone was mostly done on down-dip drill holes, and on the western end only in the primary zone. Table 17.2 Hassai South – Density on Cores within Sulphide Mineralisation Supergene Primary Supergene Primary Composites Composites Density Density Number 343 168 10.0 Percentile 2.83 3.00 Min 1.48 2.68 20.0 Percentile 3.84 3.52 Max 5.03 4.93 30.0 Percentile 4.13 3.97 Mean 4.19 4.31 40.0 Percentile 4.28 4.48 Median 4.45 4.77 50.0 Percentile 4.45 4.77 Trimean 4.40 4.55 60.0 Percentile 4.58 4.82 Biweight 4.42 4.48 70.0 Percentile 4.67 4.84 Std dev 0.71 0.73 80.0 Percentile 4.73 4.86 Cov 0.17 0.17 90.0 Percentile 4.79 4.88 95.0 Percentile 4.83 4.90 97.5 Percentile 4.85 4.90 Density does not always correlate with the Cu + Zn or Au values. As a result a uniform density was preferred for each mineralised domain. The mean density of 4.19 in the supergene domain and of 4.31 in the primary domain are considered conservative because they are significantly below the median, corresponding approximately to the 35th percentile. These densities are comparable to similar projects. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 119 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report In Hadal Awatib, where sulphide density measurements are scarce, the following uniform densities have been assigned for the domains: • Oxide (SBR): 2.1 • Supergene sulphide: 4.2, compared with 4.19 in Hassai South • Primary sulphide: 4.4, compared with 4.31 in Hassai South • Indicated resources in the sulphide domains (mostly supergene) have been assigned a 4.2 density in order to be conservative. Detailed analysis of the data set suggests that the silicate host rock has a density of 2.7-2.8 which is similar to other deposits in the area. 17.2.6 Correlations Between Elements A correlation matrix has been reviewed for each domains for both Hassai South and Hadal Awatib East. Results are shown for Hassai south primary ore in Table 17.3, and similar relationships were observed for the other domains. Correlation between the different elements is typically average or low, as is the correlation between density and any element. This suggests that the definition of sub-domains would enhance the overall quality of future models. Table 17.3 Hassai South – Primary Ore – Correlation Matrix Density Density Au FA Cu Zn Cu Eq Au FA 0.3648 1 Cu 0.0456 0.1478 1 Zn -0.1778 -0.0045 -0.1073 1 Cu Eq 0.1948 0.5194 0.9396 -0.0779 1 Cu+Zn -0.0114 0.1192 0.9681 0.2225 0.9 17.2.7 Cu+ Zn 1 1 Variography and Interpolation Parameters Variography and ordinary kriging (OK) interpolation was conducted on the following primary domains: • • Hassai South − Cu-Au supergene with a CuEq 0.8% cut-off − Cu-Au primary with a CuEq 0.8% cut-off Hadal Awatib East − Cu and Au sulphide zone with a CuEq 0.8% cut-off: Orebody 6 − Zn sulphide zone with a 0.75 % Zn cut-off: Orebody 6 (Zn envelope) − Au Oxide with a 1g/t Au cut-off : Orebodies 3, 4, 5. The remaining domains were interpolated using inverse distance square (IDS) with parameters based on variographic analysis of adjacent domains. All Indicated resources have undergone their own variographic analysis and OK interpolation. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 120 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report The samples were composited to 1 m lengths, and upper cut-offs were applied as discussed previously. The 3D variographic analysis, based on variographic maps and standards tests, led to the definition of a 3D ellipsoid parallel to the mineralised body. All variograms were modelled using 1 or 2 nested spherical variogram models. In Hassai South: • The variogram ellipsoid for the primary domain is parallel to the ore body as expected, with the major axis east-west, a semi-major axis down dip, and a minor axis across dip with a very short range. Ranges on the major axis are very long. • The variogram ellipsoid in the supergene domain shows a semi-major axis range close to the minor axis range; it suggests an heterogeneity developed in both directions: horizontally across the deposit strike, and vertically parallel to the oxidation direction. • Most of the minor axes were not possible to model, in which case the down-hole variogram was used with a correcting factor. • The Zn variogram cannot be modelled. • It was not possible to model the density variogram in the primary domain, hence an average density was assigned to the model for each domain. 17.2.8 Block Model 17.2.8.1 Block Model Definition Block model dimensions are provided in Table 17.4. Table 17.4 Block Model Definition Hadal Awatib East Hassai South Y X Z Y X Z Minimum Coordinates 2068000 751500 -300 2077600 759200 0 Maximum Coordinates 2068900 753300 580 2079200 759840 640 5 100 20 20 20 10 2.5 2.5 1.25 User Block Size Min. Block Size Rotation No sub blocks, ore percentage per block Bearing: 0 o Dip: 0 o Plunge: 0 o Bearing: o 75 Dip: 0 o Plunge: 0 o In Hassai South, kriging neighbourhood analysis (KNA) test work was carried out on the copper data, in order to minimise the conditional bias and to maximise kriging efficiency (KE) during the estimation. Block model parameters such as block size, number of samples, search ranges and discretisation levels were optimised, resulting in selection of blocks measuring 5x100x20 m (YXZ). In Hadal Awatib East, due to the high heterogeneity between different areas, the block size was based on the drill holes spacing in the supergene area drilled for Indicated resources: 20x20x10 m block size compared with a drill spacing of 40x20 m. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 121 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 17.2.8.2 Grade Interpolation Kriging and search ellipsoids determined from the variographic analysis are presented in Table 17.5. Table 17.5 Kriging Search Ellipsoids Domain Type of interpolation Top-cut KAMS1_ 2 KAMS1 _3 KAMS1 _4 KAMS1 _5 OK OK ID2 ID2 ID2 30 g/t Kriging ellipsoid Nugget (c0) 3x3x3 discretisation points 1st structure KAME2 _2 KAME2 _3 OK OK OK 30 g/t 30 g/t 30 g/t 9 5.5 12 13 9 12.6 8.5 23.5 17 12 23 35 12 20 30 7.2 120 7 150 25 110 13 160 9 180 Major / semimajor KAMW3 _1 KAMN4 _1 KAMN4 _2 KAMN4 _3 ID2 ID2 ID2 ID2 1.4 1.4 2.1 2.1 2 Major / minor 9 11 9 12 8 Bearing Major 240 55 220 55 330 0 127.4 -28.8 137.4 -28.8 0 0 45 54 54 120 150 150 150 150 110 160 180 150 150 150 250 1.4 1.5 1.5 1.5 2.1 2.1 2.1 2 1.5 1.5 1.5 1.5 Major / minor 9 11 5 5 5 9 12 8 5 5 5 5 Bearing Major 240 220 240 220 127.4 330 127.4 137.4 305 255 240 235 Plunge Search ellipsoid KAME1 _1 2nd structure (c2, a2) (c1, a1) Dip Max. distance Major / semimajor 55 55 55 55 -28.8 0 -28.8 -28.8 55 65 45 50 Plunge 0 0 0 0 54 45 54 54 0 0 0 0 Min 20 20 3 2 3 20 20 20 3 3 3 3 Max 100 100 30 30 30 100 100 100 30 30 30 30 Dip Nb informing samples 1st pass KAMS1_ 1 of Interpolation was conducted within separate domains, based on their separate samples sets, ie. hard boundaries were employed between domains. OK interpolation was performed when possible, with IDS used where data were too scattered for variographic analysis. All but a few blocks were interpolated. Non-interpolated blocks were assigned a value equal to the mean of the surrounding blocks. 17.2.8.3 Model Validation The block model grades have been compared against the composite samples, per vertical profile, or per horizontal level for both the oxide and the primary domains and were found to be closely comparable. Figure 17.3 shows an example of a validation plot for Hadal Awatib East. There is a slight underestimation for Cu and CuEq in Hadal Awatib East, likely linked to the capping. Block model volume reports have been checked against the wireframes volumes. Differences are negligible. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 122 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 17.3 Hadal Awatib East – Validation Chart for Sulphide Ore – Grade per Vertical Profile FINAL – Rev 0 – 22 Oct 2010 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 17.2.8.4 Model Presentation Sections through the models are displayed in Figure 17.4 and Figure 17.5. Figure 17.4 Hadal Awatib East Block Model : Oblique Views – Au, Cu, Zn, CuEq, AuEq, Resources Category Au ppm AuEq ppm Cu % CuEq % Zn % Resources Categories Indicated Inferred FINAL – Rev 0 – 22 Oct 2010 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 17.5 Hassai South Block Model : Oblique Views – Au, Cu, Zn, CuEq, AuEq Au (g/t) Existing pit Lower Supergene Limit Existing pit Lower Supergene Limit Cu (%) Existing pit Lower Supergene Limit Cu Eq (%) Existing pit Zn (%) FINAL – Rev 0 – 22 Oct 2010 Lower Supergene Limit The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 17.2.9 Confidence Classification and Mineral Resource Reporting under NI 43-101 Confidence classification is mostly based on drill-hole spacing, but includes a review of geological continuity, grade continuity, QAQC and database quality. The overall geological continuity of the VMS sulphide domains is reasonably assumed but still often inferred, especially for the continuity at depth and in between widely spaced drill holes. It is supported by geophysics, surface geological mapping, and outcrop on the pit floor. Geological continuity and grade continuity has been well demonstrated for Ore Body 6 (the western part of Hadal Awatib East) and for the oxide domain. Drilling of the supergene zone at Hassai South is potentially biased and some over-estimation of the grades may have happened due to the large numbers of down-dip drill holes compared to the acrossstrike drilling. On the other hand, the acidic SBR ore (which contains high Au grades) has not been evaluated. The boundary between acidic SBR and supergene sulphide ore is irregular and may not have been accurately delineated. The contact between supergene and primary sulphide ore has been conservatively chosen at RL 390 m, as information is insufficient for more precise definition. The wireframe was slightly updated in April 2010 for the mining inventory and the previously published figure has been updated here. Samples supporting the HAE Link oxide gold deposit (Indicated resource) mainly date back to 2005-06, were well-documented and are of acceptable standards, although QAQC data is limited by current standards. The database for recent drilling is of acceptable standing, but the database for historical drill holes is questionable. The weaknesses in the databases are not believed to have impaired the materiality of the resources at this level of confidence. There is a shortage of density measurements over the sulphide domains. The assignment of a uniform bulk density is considered appropriate and conservative. Therefore, it is the author’s opinion that the Resource estimates as shown in Table 17.6 can be classified as Indicated and Inferred Mineral Resources according to NI 43-101 standards. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 126 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 17.6 Hadal Awatib East (HAE) and Hassai South (HASS) Resources Estimates as of 31 December 2009 Category Area/Type Tonnes Gold Copper Gold Copper (kt) (g/t) (%) (oz) (t) - Oxides Indicated Inferred HAE Link 205.5 9.53 - 63 000 HAE - other 127.7 8.85 - 36 300 - Total Indicated 333.2 9.27 - 99 300 - HAE Link 3.5 12.54 - 1 400 - HAE - other 65.5 8.32 - 17 500 - Total Inferred 69.0 8.53 - 18 900 - Sulphides Indicated HA East, Cu>2% 508 0.78 2.80 12 000 14 200 HA East, Cu<2% 2 390 0.96 0.95 74 000 22 600 0.93 1.27 86 700 36 800 HS South, Supergene - HS South, Primary - Total Indicated Inferred 2 898 HA East, Cu>2% 2 930 0.75 2.50 71 000 73 000 HA East, Cu<2% 25 400 1.23 0.81 1 001 000 206 000 1 530 2.29 2.75 112 000 42 000 HS South, Supergene HS South, Primary 18 620 1.49 1.37 894 000 255 000 Total Inferred 48 480 1.33 1.19 2 078 000 576 000 Note: - Au was assayed by Fire Assay (30 g pulp) in sulphide zone, and by Cyanide Leach in oxide zone; Cu and Zn were assayed by triple acid digestion with AAS finish. - Sulphide cut-off CuEq 0.8%, and oxide cut-off Au 1g/t. Cut-offs based on metal prices of $750/oz gold and $2.00/lb copper, and include relevant recoveries and costs to produce metals. - HAE LINK is currently being mined. The resources of HAE Link have been updated following the Hadal Awatib re-evaluation. - Sulphide blocks of Hadal Awatib East have been filtered at 2% cut-off to allow separate reporting of high grade Cu mineralisation. Significant high grade Cu zones have been statistically highlighted and are expected to correspond to the stringer/feeder zones. Thse will require additional drilling because of their specific interest. Zn grades are too low to be reported as a resource at this stage of the study, but have been modelled for metallurgical purposes. 17.3 KAMOEB RESOURCES All work has been carried out using Surpac Vision V6.1. 17.3.1 Geological Model The Kamoeb group comprises four distinct quartz veins sets (Figure 17.6). Gold is associated with the main quartz veins, but can also be found in the immediate wall rocks. The veins have been mapped and modelled in 3D, with input from recent mining. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 127 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 17.6 Kamoeb Area – Distribution of Mineralised Veins - 2009 model Kamoeb North Kamoeb North Kamoeb East Kamoeb West Kamoeb South Kamoeb East is located within a shear corridor with intense crenulation and faulting (Chevalier, 2009). Quartz veins have been partly dismembered, and remobilisation of gold within the wall rock is especially important at this site. Despite the high density of drilling, geological continuity is poorly understood. Kamoeb South is very continuous along two main quartz veins dipping 50-70o to the south. Kamoeb North and Kamoeb West, despite being sheared, appear to be two distinct sets of relatively continuous, but thin veins that pinch and swell. Coarse gold was commonly observed in polished sections from Kamoeb. AMC submitted ten samples for screen fire assay to Intertek, Jakarta in 2009. The laboratory concluded that although there was some coarse gold (>75 µm) the coarse gold effect was not important and would not affect the accuracy of the assay. FINAL – Rev 0 – 22 Oct 2010 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 17.3.2 Cut-Off and Domain Modelling Veins have been interpreted and wireframed at a cut-off grade of 0.8 g/t using cyanidable gold (Au Cy) for 2001/2004 drill-holes, and fire assay gold (Au FA) for 2008-2009 drill-holes. The cut-off grade of 0.8 g/t Au is based on current AMC operating practices. Dilution rules are as follows: • At least 1 m mineralised interval • Maximum 2 m barren interval, with the final average grade remaining above cut-off grade • Flexibility has been applied to give consistent mineralised body outlines. Interpretation and wireframing was done on cross-sections, and all solids were snapped to drill holes. Wireframes were manually adjusted where geometry could have been an issue. Barren intervals within veins have been modelled as voids. Voids and solids have been intersected. All wireframes have been clipped to the topography (1 January 2010). Five volumes (sets of closed wireframe) have been individualised for Kamoeb South, three for Kamoeb East, one for Kamoeb West, and three for Kamoeb North. Due to the poor geological understanding, modelling in Kamoeb East remained interpretative. 17.3.3 Population Distribution and Top-cuts Capping (top-cut) was considered where the coefficient of variation (standard deviation/mean) exceeded 1.2, at the 97.5% percentile, and where outliers appeared on a standard log-normal population distribution. Where possible, no top-cut was applied, in order to avoid being overly conservative, while excluding high-grade nuggets that would lead to over-estimation of grades. For practical reason a uniform top-cut has been preferred. Basic statistics are summarised in Table 17.7, and have been determined on 1 m composites, with no cut-off or top-cut applied. Most of the samples were also 1 m long, and the statistical difference between samples and composites is not significant. For variographic analysis and interpolation purposes, the samples within the orebody wireframes were composited to 1 m. Fire assay and cyanidable gold data have slightly different distributions, but it was considered that mixing of the two populations was acceptable for the overall estimates. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 129 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 17.7 Proposed Gold Top-cuts for Different Domains (Au FA and Au Cy) KAMS1_3 KAMS1_4 KAMS1_5 KAME2_1 KAME2_2 KAME2_3 KAMW3_1 KAMN4_1 KAMN4_2 KAMN4_3 value Kamoeb North West KAMS1_2 Maximum Kamoeb Kamoeb East KAMS1_1 Kamoeb South 53.00 29.80 8.20 5.31 18.20 76.00 123.00 54.60 12.00 17.26 2.17 4.79 Mean 4.34 3.10 2.81 2.12 5.42 8.59 4.88 4.28 2.59 2.67 Median 2.76 2.02 1.58 1.44 3.60 6.50 2.16 2.14 2.10 1.87 1.20 1.11 0.85 0.66 0.99 1.03 2.13 1.46 0.74 0.90 17.90 13.16 8.20 5.31 18.20 31.39 34.85 20.90 7.30 8.87 30 - - - - 30 30 30 - - 1.36 3.12 3.20 Coefficient of 0.58 0.39 variation th 97.5 Percentile Proposed top-cut 4.79 - - Note that KamS1 and KamE2 veins have now been mined, but their data were used for modelling and separate basic statistics have been conducted as well. Results of these statistics have been carefully examined as part of the capping analysis. 17.3.4 Dry Bulk Density Based on the data set presented in Section 17.2.5, uniform densities of 2.55 for ore and 2.8 for waste have been used for the resource estimates and are considered to be conservative. 17.3.5 Variography and Interpolation Parameters Variography was conducted on 1 m composite with a top-cut applied. It was conducted on the main vein sets where data distribution was sufficiently dense for variographic analysis, ie KamS1_1, KamS1_2, KamE2_1, KamE2_2, KamE2_3. Variogram modelling commenced with an omni-directional variogram and a variogram map within the main mineralised plan. All variograms were modelled using two nested spherical variogram models. All models are based on normal variograms. Ellipsoid parameters are given along with the Surpac convention in Table 17.9. Variogram structures are poorly defined. Part of the reason for this may be the difference between AuCy and AuFA populations. All nested variograms have a strong nugget effect (20-30%), a high variance and a strong first variogram structure. The shapes of the kriging ellipsoids are in accordance with vein geometry, except in Kamoeb East where strong shearing has affected the veins. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 130 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Detailed kriging parameters, directly derived from the variographic analysis, are presented in Table 17.9. 17.3.6 Block Model 17.3.6.1 Block Model Definition. KNA was undertaken in 2008 as part of the Kamoeb South and Kamoeb East study, and block sizes of 10x10x10 m or 10x10x5 m were proposed. For practical reason 10x10x5 m has been preferred with small sub-block to allow accurate volume calculation (Table 17.8). This is consistent with drilling at 50x25 m to 25x25 m for Kamoeb South and East. Less consideration was given to Kamoeb North and East given their Inferred Resource status. Table 17.8 Kamoeb Block Model Definition Y X Z Minimum Coordinates 2061820 750640 300 Maximum Coordinates 2063420 752880 780 10 10 5 Min. Block Size 2.5 2.5 1.25 Rotation -30 0 0 User Block Size 17.3.6.2 Grade Estimation Interpolation parameters for OK and IDS are specified in Table 17.9 for each domain. Two passes of OK have been used to maximise the number of informing samples. IDS has been used for the few blocks not informed by OK. Blocks interpolated with IDS were all classified as Inferred. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 131 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 17.9 Kamoeb Interpolation Parameters Domain Type of interpolation Top-cut Kriging ellipsoid KAMS1_1 KAMS1_2 KAMS1_3 KAMS1_4 KAMS1_5 KAME1_1 KAME2_2 KAME2_3 KAMW3_1 KAMN4_1 KAMN4_2 KAMN4_3 OK OK IDS IDS IDS OK OK OK ID2 ID2 ID2 ID2 30 g/t 30 g/t 30 g/t 30 g/t Nugget (c0) 9 5.5 12 13 9 1st structure 12.6 8.5 23.5 17 12 3x3x3 (c1, a1) 23 35 12 20 30 discretisation 2nd structure 7.2 7 25 13 9 points (c2, a2) 120 150 110 160 180 1.4 1.4 2.1 2.1 2 9 11 9 12 8 240 220 330 127.4 137.4 Dip ( ) 55 55 0 -28.8 -28.8 Plunge (o) 0 0 45 54 54 Major / semimajor Major / minor Bearing Major o () o FINAL – Rev 0 – 22 Oct 2010 AMEC Page 132 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 17.9 Kamoeb Interpolation Parameters Domain Max. distance Major / semimajor Search ellipsoid Major / minor Bearing Major (o) Dip (o) samples 1st pass No. informing samples 2nd pass KAMS1_2 KAMS1_3 KAMS1_4 KAMS1_5 KAME1_1 KAME2_2 KAME2_3 KAMW3_1 KAMN4_1 KAMN4_2 KAMN4_3 120 150 150 150 150 110 160 180 150 150 150 250 1.4 1.5 1.5 1.5 2.1 2.1 2.1 2 1.5 1.5 1.5 1.5 9 11 5 5 5 9 12 8 5 5 5 5 240 220 240 220 127.4 330 127.4 137.4 305 255 240 235 55 55 55 55 -28.8 0 -28.8 -28.8 55 65 45 50 Plunge ( ) 0 0 0 0 54 45 54 54 0 0 0 0 Min 20 20 3 2 3 20 20 20 3 3 3 3 Max 100 100 30 30 30 100 100 100 30 30 30 30 Min 5 5 5 5 5 Max 20 20 20 20 20 ID2 o No. informing KAMS1_1 Method ID2 ID2 ID2 ID2 No. informing Min 5 5 5 5 5 samples Max 30 30 30 30 30 Max distance 250 250 250 250 250 3rd pass FINAL – Rev 0 – 22 Oct 2010 AMEC Page 133 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 17.3.6.3 Model Validation Block model volume reports have been checked against the wireframes volumes. Differences are negligible. The block model grades have been compared against the composite samples, per vertical profile, or per horizontal level for both the oxide and the primary domains, and are very comparable. An example is shown in Figure 17.7 for cross-sections through the two main vein sets KMS_1 and KMS1_2. Figure 17.7 Kamoeb South Validation Chart: Block Model vs Drill Holes Data by Cross-section 17.3.7 Confidence Classification and Mineral Resource Reporting Under NI 43-101 Geological continuity and grade continuity have been well demonstrated in Kamoeb South. At Kamoeb East, understanding of the geology is incomplete and geological continuity is sometimes only inferred. Grade continuity in Kamoeb East is acceptable, due to the high density of samples. For Kamoeb West and North, geological and grade continuity is still inferred, but acceptable for resource estimation at that level. The database has been consolidated for this exercise, and internal controls show that it is of acceptable standards. Sampling and assay are a patchwork of practices that appear to be acceptable and compatible. The use of cyanidable gold results in an undercall of the true gold grade, and therefore the overall result will tend to be conservative compared to a total gold method. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 134 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report The bulk density is uniform and is considered slightly conservative. Resources classification is based on continuity of the veins, their density of drilling and the geological continuity: • • Kamoeb South and East − Where estimated by OK in two passes, the main veins are considered Indicated − Low quality interpolation (3rd pass) for the two main veins are considered Inferred. Blocks lying distant from the last drill holes fence have been manually downgraded to Inferred resources − Low continuity veins are considered Inferred − Some areas where geological continuity is inferred for Kamoeb East have been downgraded to Inferred. Kamoeb West and Kamoeb North have been classified as Inferred resources. Note that Indicated resources are supported by a high density of informing samples, and that most of them have been estimated through the 1st pass of kriging, implying at least 20 informing samples. Despite the high density of drilling in some areas, no resources have been classified as Measured, since: • The mixture of assay and drilling practices makes it difficult to ensure the necessary precision • Assays remain relatively imprecise with high nugget effect • Geology in Kamoeb East and at the intersection of Kamoeb East and South remains poorly understood. Therefore, it is the author’s opinion that the resource estimate as shown in Table 17.10 can be classified as Indicated and Inferred Mineral Resources according to NI 43-101 standards. Table 17.10 Kamoeb Group – NI 43-101 Gold Mineral Resources – 1 January 2010 Ore Type Category Indicated Location Tonnage (kt) Kamoeb South Kamoeb East Total Ind. Quartz Inferred Grade (g/t Au) Metal (oz Au) 3 309 3.63 386 600 514 5.01 82 900 3 823 3.82 469 500 Kamoeb South 207 3.58 23 800 Kamoeb East 96 5.91 18 200 Kamoeb West 234 2.42 18 200 Kamoeb North 2 045 2.47 162 800 2 582 2.69 223 000 Total Inf. Notes: - Au was analysed by Fire Assay (30 g or 50 g pulp) or cyanidable gold (3 hr leach). Cut-off is 0.8 g/t, which is in line with current operations, and assumes a metal price of $750/oz. - Resources are inclusive of reserves. - Mining resumed in December 2009. Mine depletion since then has not been subtracted. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 135 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 17.4 TAILINGS RESOURCES Heap leach residue (tailings) have been drilled in 2007/2008 and 2009 for all heaps stacked up to end of 2007, and resources were estimated using resource modelling techniques by Arethuse. More recently (September 2010), stacked heaps which were still active at the time of drilling have been estimated by CSA using a metallurgical balance approach. Forecast tailings from heap leach operations over the period 2010 to 2012 are implied from the AMC mining schedule, using historical recoveries to forecast tailings grades, although these are not included in the current resource statement. All 3D models, geological and block models have been created using Surpac Vision V6.0 and above. 17.4.1 Resources Estimated by Arethuse Using Conventional Resource Modelling Techniques – Heap 63A to 113 17.4.1.1 Topography The different heap leach pads were frequently reshaped by earth-moving equipment, and several stages of surveying were necessary to follow up on the material movement. Survey was sufficiently detailed to individualised the different heaps that were drilled. Corrections were made in order to model minor features and improve the precision in volume modelling, such as pavement for the conveyor belt (30 cm), berms, etc. The tailings are divided into four blocks, A, B, C and D, of one or two layers, with a third layer added to Block D. Heap leach pads have been stacked in 6 m slices using a moving stacker. They are stacked in a northsouth direction, with the exception of the upper layer of Block C which was deposited in an east-west direction. Every heap is about 50 m wide. A 150 m length of deposit represents about 63 000 m3 (or 110 000 t). Blocks B and C were reshaped by bulldozer in 2008 to prepare room for additional heaps in 2009, and some material has been retrenched from Block D. Topography evolved again in 2009, by raising new heaps on the existing ones. Changes to the heap leach pads are shown in Figure 17.8. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 136 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 17.8 Comparison of Topography November 2007 (Top), October 2008 (middle) and December 2009 (Bottom) HLPD HLPB HLPC HLPA Additional Heaps HLPD HLPC HLPB HLPA D: Retranchement of 14,500m3 C: Remodelling of top layer: - 42,200 m3 Additional stacks A: Remodelling of bottom layer: - 26,900 m3 Additional stacks Additional stacks FINAL – Rev 0 – 22 Oct 2010 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 17.4.1.2 Domain Modelling Heaps have been modelled and interpolated as whole blocks. Each block has been modelled separately as A, B, C and D, Upper and Lower, thus creating eight distinct models that have been estimated independently. • Blocks B and C Upper and Lower were fully drilled in 2007. Volumes and tonnage have been remodelled and adjusted in late 2008 • Block A Lower was partially drilled only as it was covered by active heaps in 2008 that were still active during the second drilling campaign in 2009 • Block D Upper and Lower was drilled out in the 2007 and 2009 campaigns. Since then, a few additional stacks have been raised above the Upper level, but are still active at the time of the resources estimates • The lower surface of the heaps is still imprecise; some of the lower volumes could not be evaluated with enough precision and were classified at this stage as Inferred. 17.4.1.3 Overall Population Distribution and Top-cuts Basic statistics for the tailings data set are shown in Table 17.11. Table 17.11 Tailings Resource Basic Statistics – Au Fire Assay (g/t) File Block A Block B Block C Block D Block D1 Block (2007/2008) (2009) D2(2009) Number of samples 593 315 344 1154 1276 1362 Minimum value (g/t) 0.34 1.39 0.83 0.01 0.43 0.01 Maximum value (g/t) 11.9 9.54 10.1 29.17 29.17 17.5 Mean 2.00 3.54 2.16 1.93 1.95 1.41 Median 1.68 3.26 1.97 1.74 1.75 1.22 Variance 1.90 2.04 0.92 1.32 1.40 0.73 Standard Deviation 1.38 1.43 0.96 1.15 1.18 0.85 Coefficient of Variation 0.69 0.4 0.44 0.60 0.61 0.60 2.5 Percentile 0.59 1.55 1.02 0.85 0.76 0.51 5.0 Percentile 0.82 1.71 1.12 0.98 0.88 0.60 10.0 Percentile 0.94 1.90 1.31 1.10 1.03 0.70 25.0 Percentile 1.26 2.56 1.53 1.36 1.33 0.91 50.0 Percentile 1.68 3.26 1.97 1.74 1.75 1.22 75.0 Percentile 2.23 4.33 2.55 2.25 2.32 1.69 90.0 Percentile 3.25 5.37 3.17 2.90 3.02 2.31 95.0 Percentile 5.00 6.15 3.97 3.27 3.52 2.78 97.5 Percentile 6.23 7.27 4.34 3.95 4.19 3.22 100.0 Percentile 11.90 9.54 10.1 29.17 29.17 17.5 Top-cut 5.00 6.15 4.00 3.30 NA NA Cut-off 0.82 1.70 1.10 0.95 NA NA FINAL – Rev 0 – 22 Oct 2010 AMEC Page 138 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Au assays show a log-normal distribution, with a few outlier higher grades. In earlier work, top-cut and lower cut-off corresponded to the 5th and 95th percentiles. Both were applied for variographic analysis for Blocks A, B and C in 2007/2008, whereas only the top-cut has been used for kriging interpolation. In 2009, a slightly different approach was preferred. influence of high values was restricted to one block. 17.4.1.4 Interpolation was run using GEMS and the Dry Bulk Density The mean dry bulk density varies from 1.5 to 1.9 depending on the compaction factor. This variability is due to the nature of the heap material as well as the somewhat subjective nature of the density test. Reconciliation with production data (2007) suggests a wet density of 1.65 to 1.7, with a corresponding dry density of 1.5 to 1.6. For resource estimation purposes, a conservative uniform value of 1.5 was selected. 17.4.1.5 Variography and Interpolation Parameters Variography has been conducted separately for each Block. Samples are uniformly of 1.5 m. All grade values have been trimmed-off for variographic purpose with a top-cut and a lower cut, as deduced from the statistical analysis. The standard variographic tests have been done including variographic maps using different plans and down-hole variography. Kriging parameters, a direct consequence of the variographis analysis, are presented in Table 17.13. 17.4.1.6 Block Model Block Model Definition The difference between block model reports and solids averages 0.05%, and the block-model volume estimates are therefore considered to be accurate. KNA test work was carried out after the first drilling campaign, using blocks in all Blocks, to maximise KE during grade estimation, by optimising the block model parameters such as block size, number of samples, search ranges and discretisation levels. Orientation of the block-model (strike 105o) is an average orientation between Blocks A-B-C (strike 95o) and Block D (strike 117o). Sub-blocks have been used for a correct volume estimate. Table 17.12 Tailings – Block Model Details Type Y X Z Minimum Coordinates 2 069 600 753 400 480 Maximum Coordinates 2 071 200 756 600 672 User Block Size 25 25 3 Min. Block Size 6.25 6.25 0.75 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 139 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Estimation All of the delimited domains were estimated independently in 2008, and updated in 2009. OK was used for each the pads using all 1.5 m samples within that pad. When a sample was located at the intersection of two pads, its centroid determined to which pad it belonged. The continuity of the grade through the limits of the individual stacks did not suggest that the physical stacks limits are statistical limits. What is more, the search ellipse average range of 50-65 m means that the neighbouring stacks having limited influence on a block. Kriging parameters have been determined from the variography. Search parameters are concordant with the kriging ellipsoid. A three pass search strategy was adopted to maximise the interpolated blocks without compromising interpolation quality. Table 17.13 Tailings – Grade Interpolation Parameters Block A 1 Block B1 Block C1 Block D1 and C2 and D2 Block D2 Block D1 0 Kriging Ellipsoid o Ellipsoid bearing ( ) 97 7 5 25 0 0 0 0 0 0 0 Ellipsoid plunge ( ) 0 0 0 0 0 0 Major/semi-major 1.35 1 1 1.6 1 1 o Ellipsoid dip ( ) o Major/minor 3 2.5 1.5 2.7 1 1 Variogram model 2 Spherical 2 Spherical 2 Spherical 1 Spherical 3 spherical 3 spherical C0 (nugget): 0.15 (28%) 0.21 (18%) 0.15 (38%) 0.11 (36%) 0.15 (15%) 0.07 (7%) 0.28 Sill1: 0.2 0.34 0.12 0.19 0.26 Range1: 25 m 21 m 22 m 65 m 4m 4m Sill2: 0.18 0.6 0.12 0.33 0.31 Range2: 65 m 50 m 68 m 30 m 38 m Sill3: Range3: Top-cut 5 ppm 6.15 ppm 4 ppm 0.26 0.34 100 m 80 m 80 100 3.3 ppm First Pass Major search radius 65 m 50 m 65 m 65 m Minor search radius 20 m 20 m 20 m 20 m 80 100 Informing samples 25-50 25-50 25-50 25-50 25-50 25-50 200 Second Pass Major search radius 130 m 100 m 130 m 130 m 160 Minor search radius 40 m 40 m 40 m 40 m 160 200 Informing samples 20-50 20-50 20-50 20-50 25-50 25-50 Major search radius 250 m 250 m Third Pass 250 m 250 m Minor search radius 75 m 75 m 75 m 75 m Informing samples 10-50 10-50 10-50 10-50 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 140 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Model Validation The final grade estimate was validated statistically and graphically against the input drill hole samples and revealed a good correlation (Figure 17.9). Figure 17.9 Tailings Model: Comparison Between Block Model and Drill Hole Data (2007/8 model) Comparison of block-model result with drill-holes samples - per individual stacks 160 4.5 Block-model Au FA Average Drill-hole Au FA Average Number of samples per stack 4 140 3.5 120 Au ppm 100 2.5 80 2 60 Number of samples 3 1.5 40 1 20 0.5 103 102 101 99 100 93.3 93.2 93 93.1 92 91 90 89 88 87 86 85 84 83 82 81.2 80 81.1 79 78 77 76 75 73 71 70 69 66 65 64 0 63 0 Individual stacks Preliminary Reconciliation with Production Data – 2007/2008 Model • Tonnage A preliminary reconciliation with plant data was carried out in 2008 in order to verify the tailings resource model, as well as to support further potential tailings resources. Average re-calculated wet density is estimated at 1.69, but is quite variable between 1.4 and 3.0. Tonnages from block models are dry tonnes, whereas production data tonnes are wet tonnes, before leaching. Using an estimated 7% moisture, the wet bulk density for the block models is 1.6. Both densities are quite comparable, although the tonnage estimated by the plant is slightly higher (approximately 5%). Therefore tonnages can be considered as comparable, and cross-validated. • Grades Individual grades are difficult to compare with precision, since: − Production data are calculated from mass balance, in cyanidable gold − Leaching does not apply to individual stacks, but typically to a set of three stacks. Therefore only overall average grades have been compared, with results as follows: − Total ratio of Au_Fa 3DModel (dry weight) / Au_Cy calculated (wet weight) = 1.33 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 141 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report − After doubtful ratios have been discarded, the average ratio of Au_Fa 3DModel (dry weight) / Au_Cy calculated (wet weight) = 1.37. Production grades are calculated using mass. Therefore AuFa/AuCy ratio should be corrected by the density difference ratio (1.5/1.69 = 13%). The corrected ratio from mass difference, Au FA / Au Cy corr .= 1.18. This difference between Au Cy and Au FA is slightly higher than that typically existing between Fire Assay and Cyanidable Au at Hassai, but is considered acceptable. 17.4.1.7 Confidence Classification and Reporting in Compliance with NI 43-101 No issues were noted with the topographic survey. There is a significant volume uncertainty at the bottom of each pad, and the lower portion of every pad has been classified Inferred. Assays are considered reliable. Sample size may be a bit small and may induce some imprecision. Confidence classification of resources for the model has, therefore, been based largely on kriging quality, as follows: • Measured Resource: first pass interpolation • Indicated Resources: second pass Interpolation and nearest sample < 50 m • Inferred Resource or geological potential: third pass Interpolation. Materials at the base of the pads have been classified as inferred: grade is properly estimated but volume uncertainty is high. Block C and part of Block A have been reshaped by bulldozer. Volume has been conserved, and average grades are conserved, although the detailed grade distribution has been altered and traceability is damaged. These areas have therefore been re-classified as Indicated and Inferred, instead of Measured, Indicated and Inferred. Low confidence (Inferred) resource for a portion of Block A1, corresponds to an area that has not been drilled. Block D has been completed and re-estimated in 2009. It is a large, coherent body, modelled with a significant quantity of assay data, and a conservative density. Its classification is based on grade interpolation quality, measured here using KE. KE, derived from the Block and Kriging variances, is a measure of the variability of the estimate and the choice of parameters of estimation. Block estimates with a KE < 0.33 are classified as Inferred, KE between 0.33 and 0.66 as Indicated, and KE > 0.66 as Measured. The estimated block grades show good correlation with adjacent composite grades. No cut-off was applied, as the residue appeared to be a relatively homogeneous bulk body, where selective mining is unlikely. Final resource estimates have been classified as Mineral Resources in line with NI 43-101 standards, as provided in Table 17.14. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 142 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 17.14 Hassai Tailings Resources (Drilled, as of 31 December 2009) Category Measured Tonnage Gold Grade (kt) (g/t) Contained Gold (oz) 3 832 1.88 231 075 Indicated 2 846 1.97 180 223 Total M+I 6 677 1.92 411 298 Inferred 1 178 2.11 78 252 Total Inferred 1 178 2.11 78 252 Notes: - Grades based on fire assay gold (Intertek Jakarta). - No cut-off applied, tailings being considered as a bulk deposit. 17.4.2 Additional External Areas: Heaps 1-63, 67-71 Historical heaps (numbers 1 to 63 and 67 to 71 (partially)) were removed and deposited in three external sites as follows: • Hassai Hadaymet Road 27 500 m2 • Hassai 55 800 m2 • Banat 44 500 m2 Total for the three areas is about 127 500 m2, with an average height of 1-3 m of unconsolidated material, and their volume and tonnage are difficult to assess by traditional drilling methods. Tonnage potential is thought to be limited and has not been classified as a resource. 17.4.3 Material Stacked in 2008 – 2009 (Heaps 114 to 136) CSA visited site between the dates 25-31 August 2010 to review resource estimates of material in heaps 114-136 which were not accessible for drilling in 2008. All information within this section summarises the submitted CSA report “Hassai Heap Leach Remnant Resources” (Report #R230.2010, dated 16 September 2010). The review utilised weightometer tonnages and conveyor sampling grades recorded during stacking, and a comparison made with mining grade control data. The weightometers used are situated on both the Quartz and SBR conveyors ahead of agglomeration. Grade samples are collected by a cross-feed automatic sampler on the Quartz feed line, while grab samples are taken on the SBR feed conveyor. Gold analysis is undertaken at the mine laboratory, utilising cyanide extraction rather than fire assaying techniques. CSA noted that the design and operation of the automatic sampler is likely to lead to unrepresentative samples with potential to be biased low if segregation occurs on the belt. CSA further commented that the location of the site laboratory close to the plant, and the poor dust management was likely to lead to contamination of samples. CSA recommended that the the laboratory be moved to a clean dust free air conditioned environment; that it be equipped with appropriate dust control equipoment, and that regular round robin checks with other laboratories be completed. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 143 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Hassai leach heaps are typically 200-300 m long by 50 m wide, with up to three layers of material placed, successively on top of geotextile liners. Following depletion, the top of the pad is lowered by approximately a third and the material is pushed out to lengthen the base on which the next heap is to be stacked. This is done to ensure that there is sufficient room for moving the conveyor and stacking equipment around, and for vehicular access. A new layer of geotextile is laid on top of the modified pad, and then the next heap of material is put on top of that and irrigated. At the time of writing, a third layer has been added to about a third of the stacks. Modifying the heap shapes after stacking without complimentary resurveying impacts resource estimation accuracy, as the exact location of material is unknown. Some of the moved material ends up in front of the next stack, and other portions are used to form bunds. Records of the amount of material moved by bulldozing and grading are not kept. For these reasons, the resource classification used in this case is indicated. In addition to this, occasionally material from exhausted heaps is recycled and used to create a permeable barrier above the geotextile fabric. No records of tonnage moved are kept of material that is reused in this manner, but is thought to be less than 2% of the total stack volume. Resources contained in the heaps are estimated on the basis of conveyor weights and sampled grades. Comparison with grade control data, points to known deficiencies in both the grade control and mill data. These deficiencies include: • Grade control samples are collected by open-hole blast-hole drilling, which typically produces poor quality samples. • Grade control samples are taken at the rate of about 1 per 20 tonnes for 1x5 m drilling and 1 per 30 tonnes for 1.5x5 m or 2.5x2.5 m spaced drilling, whereas conveyor sampling is at a rate of about 1 per 10 tonnes of stacked material. Samples are combined into a daily composite which is subsequently split and assayed. • QAQC sampling of the conveyor samples and cyanide leachate includes regular checks against Certified Reference Material (CRM) and duplicate sampling, whereas grade control QAQC consists of the inclusion of some blanks in the laboratory, but doesn’t include field duplicates and insertion of CRMs in the field. • Weights in the mill are measured directly at multiple points, whereas weights for grade control purposes are derived from dig plan volumes and application of sparsely taken SG measurements. • The mill recycles some material from old heaps to form permeable protective layers above the geotextile. This material is not always accounted for in the trucking database, and in particular if it was derived from a depleted heap rather than one of the tailings storage areas. • The leachate travels through up to three heaps before being stripped of gold. Careful gold assays and flow volume measurements as well as actual metal produced, allow recoveries to be estimated from each of the heaps. However the system is not perfect and some heaps were indicated to contain very low grades that are less than what were predicted to remain using experimental recoveries. • Both grade control and stacked grades are based on cyanide soluble gold assays rather than fire assays. The two methods provide significantly different results. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 144 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report There are significant differences between grade control and stacked grades for individual heaps, but ten-point moving averages show that, overall, the mill weights are about 12% higher than the grade control weights. CSA believes that whereas the individual errors can be high, the cumulative error is within that expected for an Indicated Resource classification. In particular, this is so because the full resource can be sighted on surface. Figure 17.10 shows the grade control versus stacked contained metal. The cumulative difference error in contained gold of the grade control sampling with respect to the Mill stacking sampling is less than 10%. Figure 17.10 Grade Control v Stacked Grade Fifty-four of the exhausted heaps were sampled using auger drill holes during 2007 and 2009 by Arethuse Consulting. Measured, Indicated and Inferred Resources have been estimated for these heaps by Arethuse and are presented in Section 17.4.1 and the end-2009 resource tabulation. Fire assays with appropriate QAQC controls were used to inform these resource estimates, whereas heaps that have not been drilled are assayed for cyanide soluble gold only. Figure 17.11 shows comparative assays for remnant resources in heaps where auger drilling, grade control and Mill stacking data have been used. The auger drilling grades have been estimated using fire sssay whereas grade control and mill stacking samples have been assayed for cyanide soluble gold. There is a strong similarity between the results. The auger drilling results are better controlled, but all results exhibit similar grade ranges. The remnant resources in the remaining exhausted heaps, prior to reshaping, have been estimated using the mill balance and stacking data, and are shown in Table 17.15. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 145 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 17.11 Comparison of Remnant Resource Grade for Auger Drilling, Grade Control and Mill Stacking Data Table 17.15 Hassai Tailings from Active Heap Material, heaps 114-136, CSA September 2010 (Cyanidable Au) Category Tonnage Gold Grade Contained Gold (kt) (g/t) (oz) 14 600 Measured - Indicated 514 0.91 Total M+I 514 0.91 14 600 Inferred 1 329 1.42 58 800 Total Inferred 1 329 1.42 58 800 17.4.4 Additional Material to be Stacked, 2010-2012 17.4.4.1 Material that is Currently Being Leached (Heaps 137-141) It is not possible to provide an Indicated resource estimate for heaps that have not completed the leaching process. For the purposes of this exercise, heaps that are currently being leached and that have an unknown final leach recovery may be assigned to Inferred Resources as, although their location is known, and the input grade and tonnage has been satisfactorily measured, the final recovery and remaining grade can only be broadly estimated. The grades presented are in situ at the completion of stacking (Table 17.16), but will decline during leaching. Table 17.16 Material Currently Under Irrigation (Heaps 137-141, Cyanidable Au) Category Tonnage Gold Grade Contained Gold (kt) (g/t) (oz) Inferred 586 1.69 30 700 Total Inferred 586 1.69 30 700 Notes: - Resources estimated by S. McCracken, QP. - Au values represent cyanidable gold. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 146 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 17.4.4.2 Material that is Being Stacked but that has not yet been Exposed to Cyanide This material is currently included within the existing Mineral Resources and Mineral Reserves. Until stacking of new heaps is completed it is not possible to estimate accurately either pre-irrigation tonnes and grade or the amount of remnant resources they might contain. In order that these planned heaps can be used as resources in the CIL scoping study, CSA recommends that two separate but sequential recoveries are applied to the in-pit resources. The first estimates recovery by the heap leach process and the second estimates the later recovery by the CIPL process. 17.4.4.3 Heaps from Planned/Scheduled Production that will be Completed Prior to Commissioning of the CIL Plant As above, this material is included within existing Mineral Resources and Mineral Reserves. In order to convert this to tailings material available for reprocessing as part of the CIL Preliminary Ássessment, sequential recoveries are applied to in-pit resources. 17.5 MINERAL RESOURCE STATEMENT 17.5.1 Overall AMC Resources – 31 December 2009 Mineral Resources for Hassai Mine as of 1 January 2010 have been prepared by Remi Bosc of Arethuse, a Member of the European Federation of Geologists and a qualified person under NI43-101. The Mineral Resources for gold mineralisation have been tabulated by “Ore Type” in Table 17.17 to Table 17.20, and combined for reporting by category in Table 17.21. Table 17.17 Oxide Mineral Resources, 31 December 2009 Ore Type Category Location Indicated Hadal Awatib East Hassai North Other SBR Oxide Total Ind. Inferred Tonnage (kt) 330 231 Grade Metal (g/t Au) 9.28 4.47 (oz Au) 98 400 33 200 1 016 1 577 5.69 6.26 69 8.53 Hadal Awatib East Hassai North Other SBR Total Inf. 186 100 317 700 18 900 9 3.17 900 628 5.52 111 700 706 5.79 131 500 Notes: - Grades based on cyanide soluble gold (Hassai mine laboratory). - Cut-off grade: 1g/t HadalAwatib East; 1.5 g/t Hassai North and Other SBR deposits. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 147 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 17.18 Quartz Ore Mineral Resources, 31 December 2009 Ore Type Category Location Tonnage Grade (kt) (g/t Au) (oz Au) 3 309 3.63 386 600 Kamoeb South Indicated Kamoeb East 514 5.01 82 900 3 823 3.82 469 500 207 3.58 23 800 Kamoeb East 96 5.91 18 200 Kamoeb West 234 2.42 18 200 Total Ind. Quartz Inferred Metal Kamoeb South Kamoeb North Total Inf. 2 045 2.47 162 800 2 582 2.69 223 000 Notes: - Grades based on cyanide soluble gold (Hassai mine laboratory) and fire assay gold (Intertek Jakarta). - Cut-off 0.8 g/t. Table 17.19 Tailings Mineral Resources, 31 December 2009 Ore Type Category Location Measured Indicated Tailings Tonnage Grade Metal (kt) (g/t Au) (oz Au) 3 832 1.88 231 075 2 846 1.97 180 223 6 677 1.92 411 298 Inferred 1 178 2.11 78 252 Total Inf. 1 178 2.11 78 252 Total M+I Hassai Heap leach Notes: - - Grades based on fire assay gold (Intertek Jakarta). No cut-off applied, tailings being considered as a bulk deposit. Table 17.20 Stockpile Mineral Resources, 31 December 2009 Ore Type Category Ores Indicated Quartz Ores Acidic Total Indicated Tonnage Grade Metal (kt) (g/t Au) (oz Au) 3 1.77 150 219 3.22 22 700 664 5.92 126 400 886 5.24 149 250 Note: - Grades based on Cyanide soluble gold (Hassai mine laboraotory). FINAL – Rev 0 – 22 Oct 2010 AMEC Page 148 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 17.21 Hassai Mine Combined Gold Mineral Resources, 31 December 2009 Category Tonnage Grade Metal Measured (kt) 3 832 (g/t Au) 1.88 (oz Au) 231 075 Indicated 9 132 3.81 1 116 650 Total M+I 12 964 3.23 1 347 725 Inferred Total Inf. 4 466 3.03 432 752 4 466 3.03 432 752 Notes: - Mineral Resources estimated and classified according to CIMM categories by Remi Bosc, QP. - Assay methods and cut-off grades as shown in table. Mineral Resources for VMS mineralisation were similarly compiled by Remi Bosc of Arethuse and are reported by location and category according to NI43-101 in Table 17.22. Table 17.22 VMS Mineralisation Mineral Resources, 31 December 2009 Category Area/Type Indicated Gold Copper Gold Copper (kt) (g/t) (%) (oz) (t) HA East, Cu>2% 508 0.78 2.80 12 000 14 200 HA East, Cu<2% 2 390 0.96 0.95 74 000 22 600 2 898 0.93 1.27 86 700 36 800 HA East, Cu>2% 2 930 0.75 2.50 71 000 73 000 HA East, Cu<2% 25 400 1.23 0.81 1 001 000 206 000 1 530 2.29 2.75 112 000 42 000 HS South, Supergene HS South, Primary Total Indicated Inferred Tonnes HS South, Supergene - HS South, Primary 18 620 1.49 1.37 894 000 255 000 Total Inferred 48 480 1.33 1.19 2 078 000 576 000 Notes: - HA = Hadal Awatib, HS = Hassai South - Mineral Resources estimated and classified according to CIMM categories by R Bosc, QP - Grades based on fire assay for gold, and triple acid digest/AAS finish for base metals; at Intertek, Jakarta - Cut-off grade 0.8% copper equivalent (Cueq), where Cueq = Cu(%) + 0.63xAu(g/t) - The above relationship uses metal prices of $750/oz gold and $2.00/lb copper, and takes account of metallurgical recoveries. 17.5.2 Additional Heap Leach Tailings Resources Additional heap leach tailings resources have been estimated under the supervision of Simon McCracken of CSA, a Qualified Person under NI43-101, and are reported in Table 17.23. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 149 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 17.23 Hassai Tailings from Active Heap Material, Heaps 114-136, CSA September 2010 (cyanidable Au) Category Tonnage Gold Grade Contained Gold (kt) (g/t) (oz) Measured - Indicated 514 0.91 Total M+I 514 0.91 14 600 14 600 Inferred 1 329 1.42 58 800 Total Inferred 1 329 1.42 58 800 Notes: - Resources estimated by S.McCracken, QP - Grades based on material balance accounting, with grades determined as cyanidable gold. 17.5.3 Mineral Reserve Statement Mineral Reserves as of the end of 2009 have been determined under the supervision of Bill Plyley, of La Mancha, a Qualified Person under NI43-101, and are reported in Table 17.24. These reserves are determined in relation to the continued mining and processing of heap leach ore: no reserves have been determined for the proposed CIL or VMS phases, since the necessary resource, mining, geotechnical, processing and engineering studies have not been completed in sufficient detail to allow sufficient confidence in material movements, costs and recoveries to be determined. Table 17.24 Hassai Mine Mineral Reserves, 31 December 2009 Category Probable Tonnage Gold Grade (t) (g/t) Gold (oz) 2 557 000 4.99 410 400 Notes: - Mineral Reserves prepared under supervision of Bill Plyley, QP - Cut-off grade takes account of metal price ($750/oz Au), recoveries and operating costs, and varies according to material type. Typically it is 1.0 g/t Au. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 150 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18. OTHER RELEVANT DATA AND INFORMATION 18.1 MINING STUDIES – GENERAL STATEMENT REGARDING USE OF INFERRED RESOURCES As discussed in Section 1.4, the NI43-101 regulations do not permit the use of Inferred Resources in public reporting of mining studies and economic analysis for projects, such as Ariab, that have advanced to at least a preliminary feasibility level. However, La Mancha has applied for and been granted an exemption under Section 9.1 of NI 43-101 in order to prepare a preliminary assessment of the economic potential of the Inferred Mineral Resources to form the foundation of future developments of the Ariab gold project. Inferred Mineral Resources are considered too speculative geologically to have economic considerations applied to them that would enable them to be categorised as Mineral Reserves and there is no certainty that the preliminary assessment will be realised. Therefore, the terms “Mining Inventory” and “Potentially Mineable Material” have been used in this report to identify and distinguish Mineral Resources including Inferred Mineral Resources falling within conceptualised mine plans from NI43-101 compliant Mining Reserves. 18.2 CSA MINING STUDIES – KAMOEB 18.2.1 Mining Study Background CSA carried out the mining portion of a scoping study evaluation of AMC’s Kamoeb quartz vein gold deposits, various SBR acidic ore stockpiles and approximately 12 Mt of heap leach tailings dumps, the aim of which was to develop a mining strategy for the Hassai CIL gold plant scoping study. The scope of work included examination of production at Kamoeb deposit and SBR acidic ore stockpiles and tailings dumps for scheduling and financial analysis. The Kamoeb deposit was partially mined between 2003 and 2007. Mining at the deposit resumed in early 2010. The Kamoeb deposit has been divided into two main areas namely Kamoeb South and Kamoeb North. Kamoeb South is sub-divided into Kamoeb East and Kamoeb South while Kamoeb North is sub-divided into Kamoeb North and Kamoeb West. At this stage Kamoeb North contains only inferred resources while Kamoeb South contains predominantly Indicated resources. The Scoping study included the inferred resources at Kamoeb North to delineate Potentially Mineable Material and the Indicated resources at Kamoeb South to delineate probable reserves so that mining operations may progress at Kamoeb South. 18.2.2 Study Approach The following work was undertaken: • Mining method selection • Preliminary estimation of operating costs • Pit optimisations to determine practical pit limits whilst maximizing the project value, using Whittle Four-X software • Sensitivity analysis to determine what factors may influence the project • Pit design on selected Whittle shells • Reporting of inventories within the pit design FINAL – Rev 0 – 22 Oct 2010 AMEC Page 151 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report • Generation of mining schedules using the pit inventory information • Generation of scoping level operating and capital cost estimates. 18.2.3 Mining Methods 18.2.3.1 Kamoeb South Kamoeb South is composed of relatively continuous veins, dipping 50o to 70o to the south, while Kamoeb East comprises a set of veins where gold has been remobilised by late shearing events. Despite a high density of drilling, the geological continuity is poor at Kamoeb East. The ore zone has been outlined for approximately 150 m below the current workings, and remains open at depth. Based on the shallow depth, shape and orientation of the orebody, the fact that it has already been exposed extensively at the Kamoeb East and Kamoeb South areas, and that it is situated in oxidised material, a surface mining method was determined to be the safest and most cost-effective method of mining Kamoeb South deposit. Open Pit Analysis Based on the fact that the Hassai gold mine has been in operation for many years as an open pit operation, has a large pool of skilled and semi-skilled labour currently working on the operation and an existing open pit fleet of 60 t trucks and 120 t excavators, a conventional open pit mining method was considered. 18.2.3.2 Kamoeb North Kamoeb West, despite being sheared, appears to be a set of two relatively continuous, but thin veins. The ore zone has been outlined for approximately 100 m below surface, and remains open at depth. Kamoeb North is classified as inferred at this stage and will require further exploration drilling to increase the classification of the deposit to allow for reserve estimation. Based on the shallow depth, shape and orientation of the ore body, the fact that it outcrops along a ridge of hills, and that it is situated in oxidised material, a surface mining method was determined to be the safest and most cost-effective method of mining the Kamoeb North deposit. Open Pit Analysis Again, conventional open pit mining was considered, taking account of the existing fleet and local operator experience. 18.2.4 Pit Optimisation Pit optimisations were carried out for both Kamoeb South and Kamoeb North. Analyses were carried out on an ore production rate of 525 kt/a. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 152 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.2.4.1 Approach Whittle 4X pit optimisation software (Whittle) was used to generate optimal pits for the deposits, based on analysis of the resource model. Whittle allows the generation of a series of nested optimal pits, where each successive outline is for a slightly higher product price than the previous one. This is done for a range of prices, from the lowest for which ore can be profitably mined to the highest expected in the future. These pits are then interrogated at the base case costs and prices to establish their respective values. Selection of the optimal pit is normally based on maximising the project Net Present Value (NPV), but maximum cash flow can also be used as a selection criteria. As no capital has been included in the Whittle analysis, NPVs are only “Operating NPV” and therefore should be used only for relative ranking purposes. The Operating NPV is often overstated. Whittle incorporates time-discounting of money and assumes two extreme mining sequences (best and worst cases) for optimal pit selection. The best-case mining sequence mines the nested pits, starting with the smallest pit outline and mining subsequent pits until the largest pit is mined out. The worstcase mining sequence mines to the final pit outline bench by bench. The best case scenario returns a higher NPV due to the increased cash flow during the earlier years as a result of mining internal pits with lower strip ratios and/or higher grades. In consultation with AMC, a balance of maximum DCF and ore tonnes was used to pick the optimal pit shells to be used as the basis for the ultimate pit design for both deposits. 18.2.4.2 Optimisation Input Parameters Pit optimisation was carried out using all classified mineralisation (Measured, Indicated and Inferred) contained within the resource models. A list of financial and physical parameters were supplied to CSA Global by AMC and used as inputs for the Whittle optimisations. Table 18.1 and Table 18.2 outline the various input parameters used in the Whittle optimisations for the Kamoeb deposits. The input parameters for the Kamoeb South deposit assumed that mining starts in 2010. Due to the fact that the CIL plant will be commissioned in 2013, the processing method utilised for gold extraction from Kamoeb South ore is heap leaching for the first 3 years and CIL for the remainder of the life of the deposit. Mining at Kamoeb North deposit will commence in 2014 and ore processing method is thus assumed to be CIL for the life of the deposit. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 153 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.1 Kamoeb South Whittle Input Parameters Ariab Mining Company ‐ Kamoeb Project Pit Optimisation Input Data – Kamoeb South Deposit Note: All costs & prices in US$ Metal price Gold Year 1 Year 2 Year 3 Year 4 Year 5 US$/oz 1014 950 950 900 850 Royalties & other 7.9 7.9 7.9 7.9 7.9 % Payable 92.1 92.1 92.1 92.1 92.1 US$/oz net 934 875 875 829 783 US$/gram net 30.03 28.13 28.13 26.65 25.17 0.526 0.520 0.524 0.525 0.525 Throughput rate (Mtpa) Operating Costs - Ore Cost (HL) Units Quantity Quantity Quantity Quantity Quantity Processing US$/t ore 19.04 19.04 19.04 20.01 20.01 Ore transportation cost US$/t ore 1.61 1.61 1.61 1.61 1.61 Total annual general fixed expenses – Ore > 2.25 g/t (US$) 16,057,213 9,227,539 Allocation of general fixed expenses : Operational fixed costs US$/t ore 18.40 19.12 19.39 5.77 5.77 Year 1 Year 2 Year 3 Year 4 Mining cost ore increment US$/t ore 1.47 1.47 1.47 1.47 1.47 QRZ line 60.2% 62.5% 63.4% 32.80% Total process cost US$/t ore 40.52 41.24 41.51 28.85 28.85 SBR line 39.8% 37.5% 36.6% 67.20% Operating Costs - Ore Cost (CIL) Units Quantity Quantity Quantity Quantity Quantity Processing US$/t ore 20.01 20.01 20.01 20.01 20.01 QRZ line 9,659,644 10,037,755 10,178,897 3,026,633 Ore transportation cost US$/t ore 1.61 1.61 1.61 1.61 1.61 SBR line 6,397,569 6,019,458 5,878,316 6,200,906 Operational fixed costs US$/t ore 5.77 5.77 5.77 5.77 5.77 Mining cost ore increment US$/t ore 1.47 1.47 1.47 1.47 1.47 Annual Quarts ore production Total process cost US$/t ore 28.85 28.85 28.85 28.85 28.85 525,000 Selling costs - Product Cost Units Quantity Quantity Quantity Quantity Quantity Allocated general fixed expenses per tonne ore processed Cost of refining US$/gram au 0.053 0.053 0.053 0.053 0.053 Total selling cost US$/gram au 0.053 0.053 0.053 0.053 0.053 Allocated annual general fixed expenses (US$) QRZ line 18.40 19.12 19.39 5.77 Total annual general fixed expenses – Ore < 2.25 g/t (US$) Stockpiling Units Quantity Quantity Quantity Quantity Quantity Ore rehandling cost US$/t rehandled 0.28 0.28 0.28 0.28 0.28 9,227,539 Allocation of general fixed expenses : Mining cost - average, including CAPEX Units Amount Amount Amount Amount Amount Year 1 Year 2 Year 3 Year 4 Mining cost - waste and portion of ore US$/ t mined 1.99 1.99 1.99 1.99 1.99 QRZ line 32.8% 32.8% 32.8% 32.8% Total - Mining cost US$/ t mined 1.99 1.99 1.99 1.99 1.99 Other 67.2% 67.2% 67.2% 67.2% Increment per 10m additional depth US$/ t mined Allocated annual general fixed expenses (US$) Whittle Schedule Parameters Ore production rate Annual discount rate Mtpa 0.526 0.520 0.524 0.523 0.523 % 12.4 12.4 12.4 12.4 12.4 QRZ line 3,027,786 3,027,786 3,027,786 3,027,786 Other 6,199,752 6,199,752 6,199,752 6,199,752 Annual Quarts ore production Pit slopes degrees 38 38 38 38 38 Mining dilution % 20.0 20.0 20.0 20.0 20.0 Mining recovery % 95.0 95.0 95.0 95.0 95.0 525,000 Allocated general fixed expenses per tonne ore processed QRZ line 5.77 5.77 5.77 5.77 Metallurgical recoveries % Units All rock All rock All rock All rock All rock types types types types types Gold recovery - Oxidised & Fresh (Au>2.25g/t) % 80.00 80.00 80.00 92.40 92.40 Gold recovery - Oxidised & Fresh (Au<2.25g/t) % 0.00 0.00 0.00 92.40 92.40 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 154 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.2 Kamoeb North Whittle Input Parameters Ariab Mining Company ‐ Kamoeb Project Pit Optimisation Input Data – Kamoeb North Deposit Note: All costs & prices in US$ Metal price Gold Year 2 Year 3 Year 4 Year 5 US$/oz 1014 950 950 900 850 Royalties & other 7.9 7.9 7.9 7.9 7.9 % Payable 92.1 92.1 92.1 92.1 92.1 US$/oz net 934 875 875 829 783 US$/gram net 30.03 28.13 28.13 26.65 25.17 0.526 0.520 0.524 0.523 0.523 Total annual general fixed expenses (US$) Units Quantity Quantity Quantity Quantity Quantity 9,227,539 Throughput rate (Mtpa) Operating Costs - Ore Cost Year 1 Processing US$/t ore 20.01 20.01 20.01 20.01 20.01 Ore transportation cost US$/t ore 1.61 1.61 1.61 1.61 1.61 Operational fixed costs US$/t ore 5.77 5.77 5.77 5.77 5.77 Allocation of general fixed expenses : Year 1 Year 2 Year 3 Year 4 Mining cost ore increment US$/t ore 1.47 1.47 1.47 1.47 1.47 QRZ line 32.8% 32.8% 32.8% 32.8% Total process cost US$/t ore 28.85 28.85 28.85 28.85 28.85 SBR line 67.2% 67.2% 67.2% 67.2% Selling costs - Product Cost Units Quantity Quantity Quantity Quantity Quantity Cost of refining US$/gram au 0.053 0.053 0.053 0.053 0.053 QRZ line 3,027,786 3,027,786 3,027,786 3,027,786 Total selling cost US$/gram au 0.053 0.053 0.053 0.053 0.053 SBR line 6,199,752 6,199,752 6,199,752 6,199,752 Mining cost - average, including CAPEX Units Amount Amount Amount Amount Amount Annual Quarts ore production Mining cost - Waste and portion of ore US$/ t mined 1.99 1.99 1.99 1.99 1.99 525,000 Total - Mining cost US$/ t mined 1.99 1.99 1.99 1.99 1.99 Increment per 10m additional depth Allocated annual general fixed expenses (US$) US$/ t mined Allocated general fixed expenses per tonne ore processed QRZ line 5.77 5.77 5.77 5.77 Whittle Schedule Parameters Mtpa 0.526 0.520 0.524 0.523 0.523 Annual discount rate Ore production rate % 12.4 12.4 12.4 12.4 12.4 Pit slopes degrees 38 38 38 38 38 Mining dilution % 20.0 20.0 20.0 20.0 20.0 Mining recovery % 95.0 95.0 95.0 95.0 95.0 All rock All rock All rock All rock All rock Metallurgical recoveries % Units types types types types types Gold recovery - Oxidised & Fresh (Au>2.25g/t) % 80.00 80.00 80.00 80.00 80.00 Gold recovery - Oxidised & Fresh (Au<2.25g/t) % 92.40 92.40 92.40 92.40 92.40 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 155 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.2.4.3 Resource and Mining Models Kamoeb South and Kamoeb North The resource models used as the basis for pit optimisations are Surpac models supplied by Arethuse. The Kamoeb South and Kamoeb North resource models were originally a single model that was split into two models for ease of use. The resource models are located on a rotated grid. The resource model extents are recorded in Table 18.3. The supplied resource models included waste for pit optimisation. Table 18.3 Kamoeb South and Kamoeb North Resource Model Extents Block Model Parameters X Y Z Minimum Coordinates 2 061 820 750 640 300 Maximum Coordinates 2 063 420 752 880 780 Parent Block Size (m) 10 10 5 Minimum block size (m) 2.5 2.5 1.25 Number of cells 30 36 40 Rotation -30 0 0 Prior to optimisation, the resource models required a number of operations performed within Datamine to make them suitable for application within Whittle. These models needed to be prepared to allow for the differing mining costs, cost adjustment factors and sensitivities that were required to be analysed. The following processes were applied: • Any absent or negative geological or physical values were resolved • Unnecessary geological flags or attributes were removed (to expedite the optimisation process) • Differing rock type codes were created to distinguish ore from waste within Whittle • Bulk tonnes were created for each block • Metal content for each block were created as applicable • A mining cost adjustment factor (MCAF) attribute as created for use within the Whittle optimisation • A processing cost adjustment factor (PCAF) attribute was created for use within the Whittle optimisation. This process created the block models 4xks00.i.dm and 4xkn00.i.dm that could be used for the Whittle optimisations. The models were then imported into Whittle. Cross-checks were performed and these confirmed that the engineering model quantities (Whittle input quantities) matched those of the original resource model. 18.2.4.4 Topography A single surface topography covering Kamoeb South and Kamoeb North was supplied by AMC. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 156 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report All material above the topographical files was coded as air. 18.2.4.5 Pit Slopes At the time that the Whittle optimisation was completed no detailed geotechnical data was available for Kamoeb deposit. An overall slope angle of 38o was assumed to be suitable for the pit optimisation in order to allow for ramps and a minimum pit base width of 20 m to allow movement of mining equipment on the pit floor. 18.2.4.6 Mining and Processing Costs This study assumes that AMC undertakes owner-mining operations as at present. The costs used for the Whittle optimisations are fully allocated costs which cover all direct fixed costs associated with the ore and waste extraction and ore processing. These costs are expressed as a cost per tonne. The remaining fixed costs of investment and overhead and indirect costs (mine management costs, head office costs etc) are expressed as an annual fixed cost which is divided between the SBR and quartz mining operations. Table 18.4 summarises the mining costs and Table 18.5 the processing costs applied to the Whittle optimisations. Table 18.4 Mining Costs Applied in the Whittle Optimisations Cost Type Waste Ore 1.99 3.46 Fully Allocated Costs ($) No bench-by-bench MCAF was applied to the reference mining cost in the Whittle optimisations as all material is in the shallow oxidised zone. Table 18.5 Processing Costs Applied in the Whittle Optimisations Cost Type Fully Allocated Costs ($) 18.2.4.7 Processing Ore Transportation 19.04 1.61 Mining Dilution Ore recovery can be affected by: • Ore zone geometry and regularity • Effectiveness of grade control delineation of ore on the working bench floor • Mine planning and scheduling • Proper control of blasting • Proper control of loading operations • Proper haul truck dispatch and dump control. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 157 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Based on experience at the current mining operations at Kamoeb, a 20% mining dilution factor was applied to the project within Whittle to account for dilution that may be expected to occur during the course of mining. 18.2.4.8 Mining Recovery Mining dilution can be impacted by: • Mining width • Loading equipment selection and dimensions • Ore zone geometry and regularity • Relative friability between the ore and the host rocks • Treatment of thin waste bands within the ore zones • Good geological and grade control • Good control of drilling and blasting design and practice • Finding the right trade-off between ore recovery and dilution • Employee training and awareness. Based on experience at the current mining operations at Kamoeb, a 95% mining recovery factor was applied to the project within Whittle to account for the amount of mineralised material that might be lost during mining operations. 18.2.4.9 Metal Prices Base Case gold prices used for the optimisation study varied between $1014/oz in Year 1 gradually decreasing to $850/oz in Year 5. 18.2.4.10 Cut-off Grades Cut-off grades are determined in the optimisation on an individual block basis. Each of the deposits has separate recovery and process costs attributed. The block value is calculated from the metal price, recoveries, grades and process costs. 18.2.4.11 Discount Rate A discount rate of 12.4% was applied to calculate the discounted cash flow for the optimisation. 18.2.4.12 Optimisation Results Optimisation was carried out to determine the approximate mine life for the Project. Indicated material was included in the “Base Case” optimisations at Kamoeb South and Inferred material was included in the “Base Case” optimisations at Kamoeb North. The base metal prices and production constraints were also applied. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 158 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report After consultation with AMC, a balance of maximum DCF and ore tonnes were used to pick the optimal pit shells to be used as the basis for the ultimate pit design for both deposits. The selected optimisation shells were as follows: • Kamoeb South, 525 kt/a 2.58 Mt @ 3.66 g/t Au, SR of 5.43:1 Operating NPV of $63.34 M Total operating cost of $47.50/t ore. Kamoeb North, 525 kt/a 1.31 Mt @ 2.56 g/t Au, SR of 5.47:1 Operating NPV of $26.30 M Total operating cost of $40.32/t ore. ` • Figure 18.8 and Figure 18.9 show plan views of the selected Whittle shell for each case. Figure 18.1 Kamoeb South – 525 kt/a Optimisation Shell FINAL – Rev 0 – 22 Oct 2010 AMEC Page 159 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 18.2 Kamoeb North – 525 kt/a Optimisation Shell 18.2.4.13 Sensitivity Analysis Sensitivities on mill feed tonnes were performed on both deposits, showing the following: • • Kamoeb South − Due to the relatively high mining strip ratio, varying the mining cost has a relatively low impact on the size of the shell. Unit processing cost variations of +20% and -10% have a moderate impact on the size of the shell. − Variations in metal price and metallurgical recovery had a significant impact on the size of the shell. Kamoeb North − Due to the lower grades at Kamoeb North and the relatively high mining strip ratio, varying the mining cost has a moderate impact on the size of the shell. Reducing the unit processing cost resulted in a relatively low impact on the size of the shell, whereas increasing the unit cost had a significant impact on the size of the shell. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 160 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report − Due to the lower grades at Kamoeb North, variations in metal price and metallurgical recovery had even more significant impact on the size of the shell. 18.2.5 Mine Design 18.2.5.1 Kamoeb South and Kamoeb North The open pit mines at Kamoeb South and Kamoeb North have been designed to produce 525 kt/a of ore. The criteria taken into account include: • Statutory and safety measures • Policy decisions • Equipment dimensions. Table 18.6 shows the geotechnical parameters, as agreed by AMC and CSA, used to design the pits for each mining area, based on the pit shells identified by the Whittle optimisations: Table 18.6 Pit Design Parameters Batter Angle Bench Height Berm Width Ramp Grade Ramp Width (deg) (m) (m) (1 : x) (m) 63 10 5 10 15, 22 Ramp widths of 22 m were based on the width of the selected CAT 775 haulage truck, with an allowance for a bund wall on the open side of the ramp and enough breadth between the trucks for them to pass safely. It was agreed that, due to the relatively small size of the pit, few trucks would be required to haul rock from the deeper portions of the pit. The ramp was thus narrowed to 15 m from the second last bench as there would be no need for trucks to pass each other on the last two benches. The ramp gradient is based on what the selected haulage truck can manage while maintaining maximum production. The pits were designed to have a minimum pit base width of 20 m at all times to ensure sufficient space to manoeuvre mining equipment in the pit for load and haul operations. Figure 18.3 shows the pit design for Kamoeb South and Figure 18.4 the design for Kamoeb North. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 161 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 18.3 Kamoeb South – Pit Design – Plan view Figure 18.4 Kamoeb North – Pit Design – Plan view FINAL – Rev 0 – 22 Oct 2010 AMEC Page 162 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.2.6 Waste Handling 18.2.6.1 Kamoeb Table 18.23 details the waste dump requirements. Table 18.7 Kamoeb South and Kamoeb North Waste Dump Quantities Description Waste Production In situ Bulk Density Volume Requirement (t) bcm lcm Kamoeb South (Scenario 1) 10 922 596 2.8 3 900 927 5 071 205 Kamoeb North (Scenario 3A) 16 824 665 2.8 6 008 808 7 811 452 Kamoeb North (Scenario 3) 9 812 351 2.8 3 504 411 4 555 734 18.2.7 Mining Inventories 18.2.7.1 Kamoeb South All material above the marginal cut-off grade of 0.8 g/t has been coded as “ore”. Table 18.8 details the Kamoeb South open pit mining inventory per bench. As the inventory includes indicated material only, a probable reserve can be reported once process and engineering parameters and costs are confirmed. Table 18.8 Mining Inventory – Kamoeb South Open Pit Bench Total tones In situ ore In Ore tonnes Waste Bench Head Tonnes situ (Including 5% loss tones Strip Ratio grade Au Au and 20% dilution) grade (t) (t) (t/t) (g/t) (t) 580-590 387 387 570-580 20 645 20 645 560-570 271 353 3 501 3.45 3 991 267 362 66.98 2.88 550-560 953 185 73 056 4.95 83 284 869 900 10.44 4.12 540-550 2 091 690 248 173 4.91 282 917 1 808 773 6.39 4.09 530-540 3 029 891 370 728 4.42 422 629 2 607 262 6.17 3.68 520-530 3 051 297 381 782 4.24 435 231 2 616 066 6.01 3.53 510-520 2 816 240 337 028 4.21 384 211 2 432 028 6.33 3.51 500-510 2 550 464 272 297 4.02 310 419 2 240 046 7.22 3.35 490-500 1 931 768 218 261 3.71 248 817 1 682 951 6.76 3.09 480-490 1 348 679 176 863 3.54 201 623 1 147 056 5.69 2.95 470-480 829 699 147 829 3.38 168 525 661 174 3.92 2.82 460-470 440 735 95 728 3.43 109 130 331 604 3.04 2.86 450-460 191 562 56 939 3.40 64 910 126 652 1.95 2.83 440-450 32 028 16 903 3.68 19 269 12 758 0.66 3.07 19 559 624 2 399 087 4.12 2 734 959 16 824 665 6.15 3.43 Total FINAL – Rev 0 – 22 Oct 2010 AMEC Page 163 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.2.7.2 Kamoeb North All material above the marginal cut-off grade of 0.8g/t has been coded as “ore”. Table 18.9 details the Kamoeb North open pit mining inventory per bench. As the inventory includes inferred material only, a reserve cannot be reported. Table 18.9 Mining Inventory – Kamoeb North Open Pit Bench Total tones In situ ore In Ore Tonnes Waste Bench Head Tonnes situ (Including 5% loss tones Strip Ratio grade Au Au and 20% dilution) grade (t) (t) (t/t) (g/t) (t) 580-590 4 249 817 1.16 932 3 317 3.56 0.97 570-580 43 653 9 013 1.26 10 275 33 378 3.25 1.05 560-570 164 555 32 005 1.96 36 486 128 069 3.51 1.63 550-560 686 586 85 228 2.43 97 160 589 426 6.07 2.02 540-550 1 658 262 130 460 2.77 148 725 1 509 538 10.15 2.31 530-540 2 049 060 152 312 3.00 173 636 1 875 424 10.80 2.50 520-530 2 121 266 188 067 3.02 214 396 1 906 870 8.89 2.52 510-520 1 939 541 199 238 2.99 227 131 1 712 410 7.54 2.49 500-510 1 480 066 192 281 2.91 219 201 1 260 865 5.75 2.42 490-500 783 518 164 911 2.70 187 999 595 519 3.17 2.25 480-490 262 581 74 230 2.37 84 622 177 959 2.10 1.98 470-480 36 127 14 517 2.07 16 549 19 577 1.18 1.73 11 229 463 1 243 080 2.80 1 417 111 9 812 351 6.92 2.33 Total 18.2.8 Ore Production Schedules 18.2.8.1 Kamoeb South and Kamoeb North Datamine was used to report quantities and grades, and custom-built Excel spreadsheets were used for the scheduling of Potentially Mineable Material. In general, the following steps were undertaken in the scheduling process: • Definition of ore and waste within the pit limits using Datamine • Production of bench inventories using Datamine • Transfer of bench inventories to spreadsheet • Produce preliminary schedule. A “ramp up” profile of 70% heap leach plant capacity in Year 1 up to 100% heap leach plant capacity in Year 2 and 100% CIL plant capacity in Year 5 is used in the schedule. A Life of Mine (LOM) mining schedule was created incorporating each of the pits within Kamoeb South and Kamoeb North. The aim of the schedule was to find a balance between mining high grade material as early as possible to get maximum returns, whilst minimising the use of stockpiles in order to keep rehandling costs to a minimum. The optimum sequence for mining the deposits thus appears at this FINAL – Rev 0 – 22 Oct 2010 AMEC Page 164 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report stage to be to mine Kamoeb South first to take advantage of the higher gold grades in the Kamoeb East area, followed by Kamoeb North. The LOM mining schedule was created by sequencing the tonnes and associated grades at Kamoeb South and Kamoeb North to ensure the highest returns as early as possible. The ore extraction was scheduled by bench ensuring that initially the heap leach mill feed capacity was satisfied incorporating a one year mill feed ramp up and finally the CIL plant capacity was satisfied. The waste from each of the pits was then scheduled by bench allowing the strip ratios to be determined for each period. The associated head grade profile was also scheduled for the LOM. The overburden at Kamoeb South has already been excavated during the previous mining phase. Any overburden for the rest of the pits at Kamoeb North will be excavated during mining operations. Summary schedule data is outlined in Table 18.10, and Figure 18.5 and Figure 18.6 are graphical representations of the mining profile. Table 18.10 Kamoeb – Yearly Mining Schedule Year Pit Total Ore Grade Input Total Waste Strip Total Rock Total Ounces Input to Mill to Mill Mined Ratio Mined Output from Mill (t) (g/t) (t) (t) (oz) 1 KamS 370 193 4.08 2 967 068 8.01 3 337 261 48 609 2 KamS 531 437 3.65 3 261 278 6.14 3 792 715 62 423 3 KamS 537 740 3.52 3 299 665 6.14 3 837 405 60 897 4 KamS 533 077 3.38 3 671 049 6.89 4 204 126 57 898 5 KamS/KamN 1 001 688 2.64 6 918 657 6.91 7 920 346 84 973 6 KamS/KamN 757 244 2.59 4 969 725 6.56 5 726 970 63 006 7 KamN 420 691 2.23 1 549 574 3.68 1 970 265 30 158 4 152 071 3.06 26 637 016 6.42 30 789 087 407 965 Total FINAL – Rev 0 – 22 Oct 2010 AMEC Page 165 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 18.5 Kamoeb – Ore Production Profile Millions Kamoeb Ore Production Schedule 1.20 4.50 4.00 1.00 Tonnage (t) 0.80 3.00 2.50 0.60 2.00 0.40 1.50 1.00 0.20 0.50 0.00 ‐ 1 2 3 4 5 Year Ore (t) Head Grade (g/t) Figure 18.6 Kamoeb – Yearly Mining Profile FINAL – Rev 0 – 22 Oct 2010 6 7 Head Grade (g/t) 3.50 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.2.9 Operating Costs Open pit mining costs have been derived from the historical AMC site operating cost data. Operating costs are the combination of the mining technical services costs and the direct mining costs. The direct mining operating costs includes the drilling, blasting, excavation and haulage costs to the ROM pads or the waste dumps in close proximity to the pit collar and the ownership costs for the major items of mining equipment. Technical Services includes ore transport from pit ROM pad to the plant, workshop costs, quarry GSE which includes explosives costs and grade control costs. Table 18.11 outlines the operating cost schedule and unit operating cost data for the Kamoeb mining operation. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 167 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.11 Operating Costs Schedule – Kamoeb Open Pits Kamoeb Mine Production and Heap Leach Tailings and Acidic SBR Stockpile Reclamation Operating cost Schedule Description Units Total Year-1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Kamoeb South and Kamoeb North Production Waste Tonnes Mined kt 26 637 2 967 3 261 3 300 3 671 6 919 4 970 1 550 Ore Tonnes Mined kt 4 152 370 531 538 533 1 002 757 421 Ore Au Grade g/t 3.06 4.08 3.65 3.52 3.38 2.64 2.59 2.23 Total Rock Tonnes Mined kt 30 788 3 337 3 793 3 837 4 204 7 920 5 727 1 970 Strip Ratio t/t 6.42 8.01 6.14 6.14 6.89 6.91 6.56 3.68 Ore Haulage $ (‘000) 6 685 596 856 866 858 1 613 1 219 677 Grade Control (Laboratory) $ (‘000) 1 846 264 264 264 264 264 264 264 Quarry GSE $ (‘000) 23 473 2 093 3 004 3 040 3 014 5 663 4 281 2 378 Quarry Workshops $ (‘000) 24 749 2 309 3 196 3 231 3 205 5 783 4 438 2 587 Sub-total Technical Services $ (‘000) 56 753 5 262 7 320 7 400 7 341 13 322 10 202 5 906 Ore Mining Costs $ (‘000) 5 681 507 727 736 729 1371 1036 576 1 216 Kamoeb South and Kamoeb North Production Operating Costs Technical Services Waste Mining Costs $ (‘000) 20 901 2 328 2 559 2 589 2 880 5 429 3 899 Sub-total Direct Mining Costs $ (‘000) 26 582 2 835 3 286 3 325 3 610 6 799 4 936 1 791 Total Mining Costs $ (‘000) 83 335 8 096 10 606 10 725 10 951 20 121 15 138 7 698 Unit Cost per Tonne Rock Mined $/t 2.71 2.43 2.80 2.80 2.60 2.54 2.64 3.91 Unit Cost per Tonne of Ore Mined $/t 20.07 21.87 19.96 19.94 20.54 20.09 19.99 18.30 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 168 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.2.10 Mine Capital Costs Due to the fact that the Hassai gold mine has been in operation since 1992, mine infrastructure is already in place and a large fleet of major and minor mining equipment is owned by the company and available. 18.2.10.1 Mining Infrastructure All mining infrastructure is already in place and thus no capital expenditure is required to construct mine production infrastructure. Facilities already established for the open pit mine include: • Offices at the pit and in the central mine complex • Workshops at the pit and in the central mine complex • Service bays together with standing areas for mobile equipment at the pit and in the central mine complex • Fuel Storage and refuelling areas for mobile equipment at the pit and in the central mine complex • Explosive storage facilities • Electrical supply infrastructure. 18.2.10.2 Mining Equipment As the mine currently owns a large fleet of mining equipment, there is no need for any initial mining equipment capital expenditure. As production increases to accommodate the CIL plant capacity and the mining equipment ages, replacement capital will be required to augment the mining fleet. It should be noted that no significant additional costs have been allowed for landing new equipment in Sudan (ie. import taxes, duties, etc.). This needs to be investigated in the next level of study. Table 18.12 outlines the replacement capital cost data. Table 18.12 Replacement Capital Cost Summary – Kamoeb Open Pits Description Assumptions Total Cost ($’000) Mining Equipment Replacement Capital for Major and minor mining 8 829 equipment Total Replacement Capital Cost 8 829 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 169 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.3 CSA MINING STUDY – ACIDIC SBR ORE STOCKPILES AND HEAP LEACH TAILINGS 18.3.1 Introduction Heap leach tailings will be reclaimed at a rate of 2.0 Mt/a while acidic SBR stockpiles will be reclaimed as required, ensuring that the fresh ore plant capacity of 1.0 Mt/a will be maintained. The existing heap leach tailings will be reclaimed by bulldozer and FEL into a mobile feeder system. This in turn transfers the reclaimed material to an overland conveyor, which has been assumed to be 2500 m in length. The overland conveyor feeds a storage bin at the milling area. Acidic SBR stockpile material will be reclaimed by bulldozer and FEL into trucks and transported to the ROM bin at the crusher plant. Table 18.15 shows the heap leach tailings inventory and Table 18.14 shows the acidic SBR stockpile inventory. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 170 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.13 Heap Leach Tailings Inventory Heap Leach Tailings Mining Inventory Gold in Tailings Heaps Source for CIL Feed Number Pre 2008 Resources 63A-113 Classification M&I Ore Estimation (t) Method 6 677 000 Drilling, Assay Au Grade Gold Gold Rec Recovery (g/t) (kg) (kg) (%) Comment Fire Assay 1.91 12 793 8 315 65% Leaching completed Fire Assay 2.11 2 434 1 582 65% Leaching completed CN soluble 0.91 469 412 88% Leaching completed CN soluble 1.58 1 339 1 176 88% Leaching completed CN soluble 1.15 553 486 88% Block model Pre 2008 Resources 63A-113 Inf 1 178 000 Drilling, Block model 2008 - 2009 Heap 114-119 M&I 514 000 Metallurgic 120-129 Inf 847 000 Metallurgic Residue 2008 - 2009 Heap al balance Residue 2008 - 2009 Heap al balance 130-136 Inf 482 000 Residue Metallurgic al balance 2nd cycle leaching essentially complete - mass balance as per 31st July 2010 2010-2013 estimated Inf 2 550 000 production CIL Feed (Heap Metallurgic CN soluble 0.9 2 219 1 949 88% al balance 12 248 000 Expected metallurgical balance 1.62 19 808 13 921 70% Residue) at 1/1/2014 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 171 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.14 Acidic SBR Stockpile Inventory Acidic SBR Stockpile Mining Inventory As at Dec. 31, 2009 Acidic SBR Ore Au Grade Gold Gold Rec Recovery (t) (g/t) (kg) (kg) (%) Washable Probable Reserves 120 986 5.68 687 481 0.70 Probable Reserves 538 452 6.00 3 231 2 972 0.92 Total Reserves 659 438 5.94 3 918 3 453 Non-Washable FINAL – Rev 0 – 22 Oct 2010 AMEC Page 172 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.3.2 Heap Leach Tailings and Acidic SBR Stockpile Reclamation Schedule The LOM reclamation schedule for the acidic SBR stockpile and the heap leach tailings was created by scheduling the tonnes required to supplement the Kamoeb ore to satisfy the 1 Mt/a fresh mined ore plant requirement and to satisfy the 2 Mt/a heap leach tailings plant requirement. At this stage the grades for both the acidic SBR stockpiles and the heap leach tailings have been kept constant for the LOM. Table 18.15 below presents a summary of the total annual acidic SBR ore and heap leach tailings reclaimed. It also shows the annual mill feed grade and the final recovered gold ounces, based on input provided by AMC. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 173 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.15 Hassai Acidic SBR Stockpile and Heap Leach Tailings Reclamation Schedule Acidic SBR Year Area Tonnes Grade (t) (g/t) Heap Leach Tailings Metal. Tonnes Grade (t) (g/t) Recovery 1 (%) Total Metal. Tonnes Grade (t) (g/t) Recovery Acidic SBR washable 120 986 5.68 70% Acidic SBR non-wash/HL 466 923 6.00 92% 2 000 000 1.62 71 529 6.00 92% 1 857 962 2 126 744 (%) Metal. Gold Recovery Production (%) (oz) 120 986 5.68 70% 15 466 70% 2 466 923 2.45 80% 155 658 1.62 70% 1 929 491 1.78 73% 80 317 1.62 70% 2 126 744 1.62 70% 77 406 2 3 4 Tailings 5 Acidic SBR non-wash/HL Tailings 6 HL Tailings 7 HL Tailings 2 392 053 1.62 70% 2 392 053 1.62 70% 87 062 8 HL Tailings 3 000 000 1.62 70% 3 000 000 1.62 70% 109 189 9 HL Tailings 871 489 1.62 70% 871 489 1.62 70% 31 719 12 248 248 1.62 65% 12 907 686 1.78 73% 556 817 Total 659 438 5.94 88% FINAL – Rev 0 – 22 Oct 2010 AMEC Page 174 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 18.7 Hassai Acidic SBR Stockpile and Heap Leach Tailings Yearly Reclamation Schedule 18.3.3 Heap Leach Tailings and Acidic SBR Stockpile Reclamation Operating Costs Heap leach tailings reclamation mine operating costs consist of the cost of equipment and labour to load heap leach tailings ore onto a mobile feeder system. This in turn transfers the reclaimed material to an overland conveyor for transportation to the plant. The overland conveyor is costed under the operating plant. Acidic SBR stockpile reclamation operating costs consist of load and haul costs of acidic SBR stockpile material to the ROM pads. Table 18.16 outlines the operating cost schedule and unit operating cost data for the Heap Leach Tailings and Acidic SBR Stockpile Reclamation operation. FINAL – Rev 0 – 22 Oct 2010 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.16 Operating Costs Schedule – Heap Leach Tailings and Acidic SBR Stockpile Reclamation Kamoeb Mine Production and Heap Leach Tailings and Acidic SBR Stockpile Reclamation Operating cost Schedule Description Units Total Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 H.L. Tailings Acidic SBR Stockpile Reclamation Heap Leach Tailings Tonnes Reclaimed kt 12 248 2 000 1 858 2 127 2 392 3 000 871 Heap Leach Tailings Au Grade g/t 1.62 Acidic SBR Stockpile Tonnes Reclaimed kt 659 121 1.62 1.62 1.62 1.62 1.62 1.62 467 72 Acidic SBR Stockpile Au Grade g/t 5.94 5.68 6.00 6.00 Total Tonnes Reclaimed kt 12 908 121 2467 1929 2127 2392 3000 871 Heap Leach Tailings Reclamation Costs US$ (‘000) 13 963 2 280 2 118 2 424 2 727 3 420 993 Acidic SBR Stockpile Reclamation Costs US$ ('000) 857 157 607 93 Total Reclamation Costs US$ ('000) 14 820 157 2887 2211 Unit Cost per Tonne of Ore Reclaimed US$/t 1.15 1.30 1.17 2424 2727 3420 993 1.15 1.14 1.14 1.14 1.14 H.L. Tailings and Acidic SBR Stockpile Costs FINAL – Rev 0 – 22 Oct 2010 AMEC Page 176 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.3.4 Heap Leach Tailings and Acidic SBR Stockpile Reclamation Mine Capital Costs Due to the fact that the Hassai gold mine has been in operation since 1992, mine infrastructure is already in place and a large fleet of major and minor mining equipment is owned by the company and available. No capital costs have thus been allocated to heap leach tailings and Acidic SBR stockpile reclamation. 18.4 AMEC MINING STUDIES – VMS DEPOSITS 18.4.1 Mining Study Background AMEC carried out a scoping study evaluation of AMC’s Hassai South and Hadal Awatib VMS deposits, the aims of which were to develop a mining strategy for the VMS concentrator scoping study and to assist in planning future exploration of these deposits. The scope of work included examination of production at both 2 Mt/a and 5 Mt/a, with 5 Mt/a being selected by AMC as the final production rate for scheduling and financial analysis. 18.4.2 Study Approach The following work was undertaken: • Mining method selection • Preliminary estimation of operating costs • Pit optimisations to determine practical pit limits whilst maximizing the project value, using Whittle Four-X software • Sensitivity analysis to determine what factors may influence the project • Pit design on selected Whittle shells • Reporting of inventories within the pit design • Evaluation of underground mining options for material remaining below ultimate pits • Preliminary underground mine designs for selected underground resources • Reporting of inventories within the underground stope designs • Generation of mining schedules using the pit and underground stope inventory information • Generation of scoping level operating and capital cost estimates. 18.4.3 Mining Methods 18.4.3.1 Hassai South The Hassai South VMS deposit is a fairly regular ore-body, with an approximate strike length of 1.2 km, dipping at approximately 65-70º and exposed in the bottom of an abandoned oxide pit. The ore zone has been outlined for approximately 300 m below the existing pit, and remains open at depth. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 177 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report The orebody dips under an existing oxide waste dump, thus any cutback on the pit will require rehandling of the oxide waste. This alone suggests that the greatest potential for VMS mining is likely to be from underground. There are, however, certain underground “upsides” should a cut-back be viable on the existing pit: • The existing oxide pit has had a few wall failures (one of some significance), that will require remedial work to ensure longer term pit stability for underground access • There is a significant amount of supergene material in the bottom of the pit, which is of high value, and is not likely to be ideal material for use as a crown pillar • While cutting back the pit, provision can be made for the location of the portal (larger berm above portal, flatter slope, etc.). Open Pit Analysis Based on historical information and studies provided by AMC, standard open pit mining was considered, taking account of the existing fleet size of 90 t trucks and 120 t excavators, and local operator experience. Once the final scoping open pit dimensions and strip ratios are known, the suitability of this fleet for the resulting mining schedule can be reassessed. Underground Analysis The orebody dip, strike length and thickness lend themselves to sub-level open stoping (SLOS). Preliminary site geotechnical investigations (Section 18.2) indicates that the hanging wall is of a fair rock mass condition, and a maximum unsupported strike length of 30 m could be achievable for a 30 m sub-level spacing. In order to maintain global stability and maximise extraction stopes, backfill will be required, and paste fill has been assumed. 18.4.3.2 Hadal Awatib The Hadal Awatib VMS deposit is interpreted as a series of folded lenses of varying width, with a fairly large regular pod of mineralisation at the southwest end. The orebody is exposed in the bottom of an existing oxide pit, however the geology appears very complex and AMC advises that more drilling is required to improve understanding of the deposit. Open Pit Analysis Again, standard open pit mining was considered, taking account of the existing fleet and local operator experience. Underground Analysis Due to the relatively complex nature of the deposit, and the current level of understanding of the orebody, it was decided that underground evaluations would be considered only for material remaining below any VMS open pit, and that any underground evaluation would be carried out at a high level. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 178 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.4.4 Pit Optimisation Pit optimisations were carried out for both VMS deposits. Analyses were carried out on VMS ore production rates of 2 Mt/a and 5 Mt/a. 18.4.4.1 Approach Whittle 4X pit optimisation software (Whittle) was used to generate optimal pits for the deposits, based on analysis of the resource model. Whittle allows the generation of a series of nested optimal pits, where each successive outline is for a slightly higher product price than the previous one. This is done for a range of prices, from the lowest for which ore can be profitably mined to the highest expected in the future. These pits are then interrogated at the base case costs and prices to establish their respective values. Selection of the optimal pit is normally based on maximising the project NPV, but maximum cash flow can also be used as a selection criteria. As no capital has been included in the Whittle analysis, NPVs are only “Operating NPV” and therefore should be used only for relative ranking purposes. The Operating NPV is often overstated. Whittle incorporates time-discounting of money and assumes two extreme mining sequences (best and worst cases) for optimal pit selection. The best-case mining sequence mines the nested pits, starting with the smallest pit outline and mining subsequent pits until the largest pit is mined out. The worstcase mining sequence mines to the final pit outline bench by bench. The best case scenario returns a higher NPV due to the increased cash flow during the earlier years as a result of mining internal pits with lower strip ratios and/or higher grades. In consultation with AMC, the maximum cash flow shell was selected as the basis for the ultimate pit design for both deposits. 18.4.4.2 Optimisation Input Parameters Pit optimisation was carried out using all classified mineralisation (Measured, Indicated and Inferred) contained within the resource models. Load/haul and blasting costs were supplied by AMC, based on the current site costs. Processing costs, anticipated metallurgical recoveries and selling costs were provided by AMEC. Base Case metal prices used for the study were $2.19/lb for copper and $900.00/oz for gold. A summary of the Whittle input parameters used in pit optimisation is provided in Table 18.17. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 179 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.17 Whittle Input Parameters ITEM VALUE SOURCE DISCOUNT RATE Discount Rate % 10 La Mancha % of NSR % of NSR % of NSR 7.90 La Mancha Unit US$/lb US$/oz 2.19 900.00 La Mancha La Mancha 2.8 Remi Bosc 1.37 2,204.62 32,150 0.03215 La Mancha ROYALTY Sudanese Ministry for Geology (GRAS) La Mancha (via Cominor) Total Royalty METAL PRICES METAL PRICES Copper Gold GENERAL Default Density Conversion Factors FOREX Pound units per metric tonne Ounce units per metric tonne Ounce units per gram Unit Euro:US times times times PROCESS Oxide CIP Recovery Cu Au Zn Concentrate Grade Cu Au Zn % % % 92.4 % g/t % Hassai South Supergene Primary Hadal Awatib Supergene Primary 81.0 67.0 7.0 90.0 36.0 14.0 85.0 29.0 10.0 85.0 29.0 10.0 AMEC Minproc AMEC Minproc / CIP Scoping Study AMEC Minproc 32.0 25.1 25.1 25.1 AMEC Minproc REVENUE CALCULATIONS COPPER SULFIDE Payable Metal Payable Metal Minimum Deduction Minimum Deduction %Cu %Au %Cu Au g/t 96.5 100 1.10 1.0 AMEC Minproc AMEC Minproc AMEC Minproc AMEC Minproc US$/lb Cu 0.20 AMEC Minproc US$/lb Cu US$/lb Cu % % 0.00 0.00 0 0 AMEC Minproc AMEC Minproc AMEC Minproc AMEC Minproc US$\oz payable Au 4.00 AMEC Minproc %Zn US$/1%Zn 2.0 2.00 AMEC Minproc AMEC Minproc %Au 100 La Mancha Metal Price Payable US$/g Au 26.6491 Selling Costs Base Charge US$/g Au 0.05 Unit US$/wmt con US$/wmt con US$/wmt con US$/wmt con US$/wmt con US$/wmt con 34.00 5.90 0.00 10.00 50.00 104.00 % 9 AMEC Minproc US$/t conc US$/dmt conc US$/dmt conc US$/dmt conc US$/dmt conc US$/dmt conc US$/dmt conc US$/dmt conc 6.40 37.36 6.48 0.00 10.99 54.95 114.29 230.47 AMEC Minproc Treatment Charge Base Charge (TC/RC) Price Partisipation Upper Price Lower Price Upscale Down Scale Refining Charge Penalties Limit Rate CIP Gold Payable Metal Kamoeb 2010 Optimisation Parameters CONCENTRATE COSTS COPPER SULFIDE Transport Cost ‐ mine to port Port Charges Packing Cost Insurance Interest Shipping Costs Moisture Content Con Marketing Transport Cost ‐ truck to port Port Charges Packing Cost Insurance Interest Shipping Costs Total Cost Hassai CIP Scoping Study AMEC Minproc AMEC Minproc AMEC Minproc AMEC Minproc AMEC Minproc PROCESS OPERATING COSTS Supergene/Primary Processing Site G&A Ore Transport Total Operating Cost US$/t ore US$/t ore US$/t ore US$/t Ore Oxide ‐ CIP Processing Site G&A Ore Transport Total Operating Cost Hassai South 2Mtpa 5Mtpa Hadal Awatib 2Mtpa 5Mtpa 8.36 5.54 0.62 14.52 8.30 5.54 2.47 16.31 5.75 2.22 0.62 8.58 5.69 2.22 2.47 10.37 Hassai South 2Mtpa 5Mtpa Hadal Awatib 2Mtpa 5Mtpa AMEC Minproc AMC Site Costs AMC Site Costs US$/t ore US$/t ore US$/t ore US$/t Ore 13.49 13.49 13.49 13.49 CIP Scoping Study 0.62 14.11 0.62 14.11 2.47 15.96 2.47 15.96 AMC Site Costs % % 100 0 US$/t.km km US$/t Hassai Sth 0.21 3 0.62 Hadal Awatib 0.21 12 Pit located adjacent to plant 2.47 US$/t Hassai Sth 1.34 ‐ 8.6 Hadal Awatib 2.8 ‐ 8.1 Varies by bench Hassai Sth Hadal Awatib deg 48 +/‐ 2deg 43 +/‐ 2deg AMEC Minproc deg 39 +/‐ 2deg 41 +/‐ 2deg AMEC Minproc Mtpa 2.0 & 5.0 MINING OPERATING COSTS Mining Recovery Mining Dilution Ore Overhaul Cost Unit Cost Distance to ROM Overhaul Cost Operating Cost Estimate Mining Cost PIT SLOPES North Wall All mRL Sensitivity South Wall All mRL Plant Throughput AMEC Minproc Diluted Resource Models AMC Site Costs AMC Site Costs / AMEC Minproc AMEC Minproc FINAL – Rev 0 – 22 Oct 2010 AMEC Page 180 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.4.4.3 Resource and Mining Models Hassai South The first resource model supplied by AMC was a “partials” model based on a very large block size (5x100x20 m - YXZ). To reduce mining dilution, a smaller block size model was supplied, that would better match a likely minimum mining unit (5x20x10 m), so that a consolidated block grade could be applied. The following formula was used to define Cueq in the resource models: All Ore Types: Cueqr = Cu (%) + 0.63 x Au (g/t) Table 18.18 reports the updated partials resource model at a cueqr cut-off grade of 0.8%. Table 18.18 Hassai South – Underground Resource Model Update Category Volume bcm Tonnes r Cu Au Zn Cueq (%) (g/t) (%) (%) 389 559 1 632 250 2.67 2.15 0.09 4.02 Primary 4 380 677 18 880 716 1.36 1.49 0.19 2.30 Total VMS 4 770 236 20 512 966 1.47 1.54 0.18 2.44 Supergene The partial percentage grades for the block needed to be converted into consolidated grades, as it was not considered appropriate to assume the ore could be selectively mined within a block. Metal tonnes by element were calculated by material type (ie. oxide, supergene, primary), and then added together and divided by the total tonnes of the block to give the consolidated block grades. For optimisation, a block needs to be of a single material type; as such the material type with the largest block proportion percentage was used to define the material type for the block. Table 18.19 details the updated consolidated grade resource model using a Cueqr cut-off grade of 0.8%. Table 18.19 Hassai South – Resource Model Consolidated Grades Category Volume Tonnes (bcm) Cu Au Zn cueqr (%) (g/t) (%) (%) 678 000 2 402 758 1.70 1.40 0.06 2.58 Primary 5 548 000 21 641 999 1.11 1.22 0.16 1.88 Total VMS 6 226 000 24 044 757 1.17 1.24 0.15 1.95 Supergene Hadal Awatib The resource model used as the basis for pit optimisation is a Surpac model supplied by AMC. The resource model was located on a rotated grid, and had sub-blocks down to a very small size (2.5x2.5x1.25 m). To convert this model into a mining model, blocks were consolidated into what was considered a reasonable size minimum mining unit based on the current mining fleet. In order to FINAL – Rev 0 – 22 Oct 2010 AMEC Page 181 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report maintain model integrity, a mining block size of 5x5x2.5 m was chosen as an increment of the smallest sub-block size. The resource model was expanded to include waste for pit optimisation. 18.4.4.4 Net Smelter Return Net smelter return (NSR) is defined as the payment received by the mine after the smelter, refiner or buyer has deducted all their charges, and all the transport, storage, insurance, assaying and marketing costs associated with delivering the product to the customer have been subtracted. The NSR was calculated on a block by block basis for each model. The final coded “mining model” was exported into Whittle and resource reports were generated to confirm the integrity of the model coding. The value of any oxide CIL plant feed was not included in the NSR calculation, as this was treated as a VMS plant NSR only. However, the value of the CIL plant feed was calculated in Whittle, and used to potentially assist in paying for incremental stripping. Some concerns were expressed by the resource geologist about the confidence in the Hassai South oxide material, so the impact of this material on the results was assessed as part of the sensitivity study. 18.4.4.5 Topography Surface topography for Hassai South was supplied by AMC. For Hadal Awatib several topographical files were supplied by AMC. These were merged with the supplied ultimate pit design to form the final topographical file used for the Hadal Awatib evaluation. All material above the topographical files was coded as air. 18.4.4.6 Pit Slopes Pit slopes were based on site observations of the current oxide mining operations (refer to Section 18.2). 18.4.4.7 Mining Costs Load and haul costs were supplied by AMC, based on existing oxide mining unit costs for excavators and 40-60 t trucks. As the mining costs were for oxide operations and only a single rate was available (rates by bench not available), AMEC adjusted the values based on its internal cost database to develop bench rates, and make allowance for the increase in pit depth. Mining G&A cost was assumed to be included in the unit processing cost. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 182 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.4.4.8 Mining Dilution Both resource models were considered to be diluted models; as a result no further dilution was added for pit optimisation. A mining recovery of 100% was applied. 18.4.4.9 Metal Prices Base Case metal prices used for the study were $2.19/lb for copper and $900.00/oz for gold. 18.4.4.10 Cut-off Grades Cut-off grades are determined in the optimisation on an individual block basis. Each of the deposits has separate recovery and process costs attributed. The block value is calculated from the metal price, recoveries, grades and process costs. 18.4.4.11 Discount Rate A discount rate of 10% was applied to calculate the discounted cash flow for the optimisation. 18.4.4.12 Optimisation Results Optimisation was carried out to determine the approximate mine life for the Project. Measured, Indicated and Inferred material was included in the “Base Case” optimisations, and the base metal prices and production constraints applied. The marginal cut-off grade (NSR $/t) by option and deposit was as follows: • Hassai South, 5 Mt/a $8.58/t • Hadal Awatib, 5 Mt/a $10.37/t After consultation with AMC, the maximum cash flow shell was selected as the basis for the ultimate pit designs. The selected optimisation shells were as follows: • Hassai South, 5 Mt/a 5.7 Mt @ $53.01/t NSR, SR of 7.6:1 Operating NPV of $112.0 M Total operating cost of US$34.44/t ore • Hadal Awatib, 5 Mt/a 16.1 Mt @ $38.07/t NSR, SR of 4.2:1 Operating NPV of $165.3 M Total operating cost of US$28.05/t ore Figure 18.8 and Figure 18.9 detail a plan view of the selected Whittle shell for each case. Whittle shells are coloured blue with the starting topography coloured green. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 183 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 18.8 Hassai South – 5 Mt/a Optimisation Shell Figure 18.9 Hadal Awatib – 5 Mt/a Optimisation Shell FINAL – Rev 0 – 22 Oct 2010 AMEC Page 184 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.4.4.13 Sensitivity Analysis Sensitivities were performed on both deposits for the 2 Mt/a case, showing the following: • • Hassai South − As the grade of the oxide material was relatively low, exclusion of this material from the Whittle evaluation results in a marginal impact on operating NPV and only a minor impact on the resulting shell − Due to the relatively high mining strip ratio, varying the mining cost has a very significant impact on the size of the shell and the pit financials. Unit processing cost variations of +20% and -10% have a moderate impact on the results − Varying the slopes by ±2 degrees had a marginal impact on the overall results. Hadal Awatib − Exclusion of the oxide material in the resource model resulted in a very significant impact on operating NPV (-47%), but had very little impact to the overall size of the shell − Overall, the relatively low average NSR results in the deposit being sensitive to any cost increases; an increase of 20% in mining costs had a significant impact on the results, as did a 20% increase in operating costs, both in terms of shell size and the pit financials − Varying the slopes by ±2 degrees had a marginal impact on the size of the shell, but flattening of the slopes had a significant impact on the pit financials. 18.4.5 Mine Design 18.4.5.1 Hassai South Open Pit A design was developed for the 5 Mt/a option, but no sensible cutback option was possible; extremely small cutbacks were indicated on the northern side of the pit, with “bull noses” on the southern side (near the waste dump peaks). Underground A smaller block size (5x10x5 m) model was provided as being more appropriate for underground mining analysis. Table 18.20Table 18.20 details the updated partials resource model using a Cueqr cut-off grade of 0.8%. Table 18.20 Hassai South – Underground Resource Model Update Category Volume Tonnes (bcm) Supergene Cu Au Zn cueqr (%) (g/t) (%) (%) 390 245 1 635 127 2.67 2.14 0.09 4.02 Primary 4 379 643 18 876 265 1.36 1.49 0.19 2.30 Total VMS 4 769 888 20 511 392 1.47 1.54 0.18 2.44 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 185 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.21 details the updated consolidated grade resource model at the same cut-off grade. Table 18.21 Hassai South – Underground Resource Model Consolidated Grades Category Volume Tonnes (bcm) Supergene r Cu Au Zn cueq (%) (g/t) (%) (%) 599 250 2 203 788 1.91 1.56 0.07 2.89 Primary 5 256 500 20 943 393 1.17 1.28 0.17 1.98 Total VMS 5 855 750 23 147 181 1.24 1.31 0.16 2.07 Orebody parameters suggested SLOS with paste fill as the most likely method of mining. All oxide mineralisation was considered as waste for the purposes of the underground evaluations. The following formulas were developed to define the Cueq based on the updated revenue factors: • Supergene Ore: Cueqm = Cu (%) + 0.501 x Au (g/t) • Primary Ore: Cueqm = Cu (%) + 0.228 x Au (g/t) Typical average unit mining costs for SLOS with paste fill were applied to develop a “preliminary” marginal cut-off grade of 1.5% cueqm for stope definition. Table 18.22 details the updated consolidated resource model using a Cueqm cut-off grade of 1.5%. It should be noted that these results assume that no blocks below the cut-off fall in the stopes, and that no additional ore loss and dilution factors have been applied for underground mining. Table 18.22 Hassai South – Underground Resource Model (Cut-off 1.5% Cueqm) Category Volume Tonnes (bcm) Supergene m Cu Au Zn cueq (%) (g/t) (%) (%) 461 500 1 765 056 2.19 1.76 0.09 3.07 Primary 2 238 000 9 347 917 1.60 1.61 0.28 1.97 Total VMS 2 699 500 11 112 973 1.69 1.64 0.25 2.14 A level interval of 30 m was used, except for the sulphide zone, where 20 m was applied. As the concept involved early removal of supergene stopes, a crown pillar has been planned for 350-380 mRL. A temporary crown pillar has also been allowed for at 230-260 mRL to allow top-down stoping to be carried out above this location and bottom-up stoping below. Once the two stoping zones are completed, this pillar will be extracted on retreat. Once all stoping is complete, the ultimate crown will be removed. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 186 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 18.10 outlines the conceptual stoping levels for Hassai South. Figure 18.10 Hassai South – Stoping Concept As the current resource is considered to be relatively shallow (approximately 300 m below the current pit base), decline access was applied. The existing pit ramp is on the hanging wall side of the deposit, and it was decided to locate the portal on the same side, to avoid crossing the bottom of the pit and sterilising part of the supergene mineralisation in maintaining underground access. Figure 18.11 indicates the conceptual decline portal location for Hassai South. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 187 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 18.11 Hassai South – Decline Portal Location The following basic decline development design parameters were applied: Decline/Incline Development Width Height Gradient Curve Radius 5.0 m 5.5 m 1 in 7 straight sections 1 in 8 curves 22 m Other Development Level Access Level Development Ventilation Drives Ventilation Rises 5.0 mW x 5.0 mH 5.0 mW x 5.0 mH 4.5 mW x 4.5 mH 3.5 mD An allowance of 15% was applied to decline development centreline designs to allow for stockpiles and miscellaneous stripping. This allowance was increased to 23% for level development to allow for additional stripping for stope slots and the like. Due to the ore zone strike length, dual declines were designed so that multiple stoping fronts could be set up on each level, thus increasing the potential production rate from underground. Link drives have been included to simplify traffic flow between the west and the east sides of the mine. This will also allow for the potential application of road trains if required, without the need for turning loops. However, an additional link drive may be required at the bottom of the mine for this to be workable. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 188 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 18.12 outlines a long section of the underground development concept. Figure 18.12 Hassai South – Development Long Section Although no detailed ventilation work has been carried out, proposed production rates suggest that an additional fresh air intake would likely be required. Consequently one has been included at the top of the East Incline, which would be used as the second means of egress via an installed ladder way. 18.4.5.2 Hadal Awatib An open pit design was provided for the 5 Mt/a VMS processing option. The following pit design parameters were applied: North Wall Bench Height Berm Width Batter Angle Ramp Width Ramp Grade 10 m 4m 65 o 22 m 1 in 10 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 189 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report South Wall Bench Height Berm Width Batter Angle Ramp Width Ramp Grade 10 m 6.5 m 75° 22 m 1 in 10 Figure 18.13 shows the pit design for the 5 Mt/a option. Figure 18.13 Hadal Awatib – Pit Design 5 Mt/a 18.4.6 Waste Handling 18.4.6.1 Hassai South Development waste will initially be placed on the top of the existing oxide waste dumps. However, once stoping commences, waste will either be dumped into mined-out stopes, or dumped into the bottom of the pit (once the supergene stopes have been removed and backfilled). Figure 18.14 outlines the location of the existing Hassai South oxide waste dump locations. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 190 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 18.14 Hassai South – Oxide Waste Dump Location 18.4.6.2 Hadal Awatib Table 18.23 details the waste dump requirements. Table 18.23 Hadal Awatib Waste Dump Quantities Option Req. Dump (Mbcm) Option 5 Mt/a 37.6 Waste Types by Option Total Unclassified Mineralised (Mbcm) (Mbcm) (Mbcm) 28.9 28.5 0.45 A conceptual waste dump design was completed (Figure 18.15). Detailed topography of the proposed waste dump location will be required to generate a more accurate design at the next level of evaluation. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 191 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 18.15 Hadal Awatib – Conceptual Waste Dump Location 18.4.7 Mining Inventories 18.4.7.1 Hassai South All material inside the ore drives and stope outlines was coded as ore. Supergene ore stopes have been assumed to be stoped in 20 m strike lengths, with 10 m rib pillars left between every second stope (ie. 10 m pillar per 50 m strike length), thus a 20% ore loss factor was applied (material left in support pillars). Primary ore stopes have been assumed to be stoped in 30 m strike lengths, with no rib pillars required. Typical ore loss factors for SLOS are in the range of 5-10% (material left in stopes and on walls); a 5% ore loss factor was applied for the scoping study. Dilution factors associated with SLOS are typically in the range of 10-15% to allow for overbreak and stope wall failures. Potentially some dilution had already been included in the grade consolidation, therefore stopes had a 10% dilution factor applied. Stopes are typically located inside a wider “ore zone”, and, as such, it was considered that including dilution as purely waste would not be reasonable. The average grade of the resource blocks outside of the stope outlines was assessed, and a dilution grade 0.49% Cu and 0.60 g/t Au was applied. Table 18.24 details the “base case” Hassai South underground mining inventory. As these inventories contain inferred material, a reserve cannot be reported. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 192 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.24 Mining Inventory – Hassai South Underground Material Tonnes Cu Au Zn (Mt) (%) (g/t) (%) 1.2 1.22 1.20 0.16 Development Ore Stoping Supergene 1.7 1.59 1.27 0.06 Primary 9.6 1.36 1.40 0.24 12.5 1.37 1.36 0.21 Total VMS Feed As a sensitivity, dilution factors associated with SLOS were set to 0%, in the event that sufficient dilution had already been included in the block consolidation of grades. Ore loss factors were however still applied, as loss of material in support pillars and stope recovery still needed to be considered. Table 18.25 details the resulting “un-diluted” underground mining inventory. Table 18.25 Undiluted Mining Inventory Sensitivity – Hassai South Underground Material Tonnes Cu Au Zn (Mt) (%) (g/t) (%) 1.22 1.20 0.16 Development Ore 1.2 Stoping Supergene 1.4 1.76 1.38 0.07 Primary 8.8 1.44 1.47 0.26 11.5 1.46 1.43 0.23 Total VMS Feed Further work is required on the resource model before more definitive dilution factors can be determined, and refinement of the stope shapes can be carried out. As the original resource blocks were based on a partial percentage ore approach, there is essentially no spatial aspect to the location of this ore, thus accurate definition of stope shapes is impossible, leading to the need for consolidation of grades within each block. This then raises the issue of dilution, which is difficult to measure in the current ore model as the ore has no spatial aspect. Moving forward, a reasonable minimum mining unit needs to be incorporated in the resource estimation process, from the construction of ore interpretations through to block size selection for resource estimation. Possibly most importantly, the grade field applied to the blocks needs to assume the entire block will be mined. 18.4.7.2 Hadal Awatib All material above the marginal NSR cut-off grade of $10.37 has been coded as ore. Table 18.26 details the Hadal Awatib open pit mining inventory. As the inventory includes inferred material, a reserve cannot be reported. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 193 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.26 Mining Inventory – Hadal Awatib 5 Mt/a Open Pit Total Tonnes Cu Au Zn (Mt) (%) (g/t) (%) SR SG Waste 78.0 Oxide 0.4 0.09 4.30 0.02 2.44 2.74 Supergene 8.9 0.96 0.96 0.63 4.01 Primary 8.0 1.26 0.88 0.72 4.19 Mineralised Waste 1.5 0.10 0.16 0.65 Total VMS Feed 16.9 1.10 0.92 0.67 18.4.8 Ore Production Schedules 18.4.8.1 Hassai South 3.43 4.7 4.10 Surpac was used to report quantities and grades, and custom-built Excel spreadsheets were used for the scheduling of the Hassai South Project. The schedule, which is summarised in Table 18.27, was intended to produce ore as quickly as feasible from the underground mine. In general, the following steps were undertaken in the scheduling process: • Definition of ore within the stope and ore drive outlines using Surpac • Production of stope and development inventories using Surpac • Transfer of inventories to spreadsheet • Transfer of development quantities to spreadsheet • Produce preliminary schedule. Figure 18.16 displays graphical representations of the underground ore mining profile. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 194 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.27 Hassai South – Underground Mining Schedule Mining Schedule ‐ Hassai South Underground CAPITAL Unit Total m‐1 m‐2 m‐3 Y1 Y2 Y3 1,207 896 644 774 1,066 508 136 112 327 127 101 307 327 197 124 54 148 Y4 Y5 Y6 Y7 Y8 Y9 Y10 Y11 Y12 Y13 Y14 Development Dimensions Width Height Diameter Description Horizontal Development Main Decline East Incline East Decline Link Drives Link#1 Link#2 Level Access Vent Drives 5 5 5 5 5 5 4.5 5.5 5.5 5.5 5.5 5.5 5 4.5 Vertical Development Vent Rises FAR RAR 3.5 3.5 5mW x 5.5mH 5mW x 5.5mH 5mW x 5.5mH 5mW x 5.5mH 5mW x 5.5mH 5mW x 5mH 4.5mW x 4.5mH (m) (m) (m) (m) (m) (m) (m) 3,169 644 1,697 192 213 1,142 590 3.5mD 3.5mD (m) (m) 124 399 Waste Tonnes Trucking OPERATING 923 192 (t) (t.km) 583 978,457 156 106,741 222 351,866 205 519,850 (m) (kt) (%) (g/t) (%) 14,914 1,182 1.22 1.21 0.16 927 69 1.25 0.99 0.10 6,840 525 1.18 0.99 0.16 7,147 589 1.25 1.43 0.18 127,596 843,167 740,829 6 218 50 1,744 40 1,389 41 1,427 37 1,283 42 1,472 35 1,208 33 1,153 23 782 21 712 14 478 9 295 11,797 9,269 114,641 90,075 113,908 89,499 109,226 85,821 121,608 95,549 118,059 92,761 103,924 81,655 113,501 89,180 75,681 59,463 67,601 53,115 46,738 36,723 29,327 23,042 130 1.31 1.02 0.02 1,261 1.39 1.11 0.17 1,253 1.33 1.32 0.11 1,201 1.40 1.30 0.21 1,338 1.39 1.65 0.22 1,299 1.38 1.44 0.20 1,143 1.45 1.48 0.21 1,249 1.38 1.58 0.28 832 1.38 1.33 0.28 744 1.42 1.18 0.30 514 1.45 1.26 0.33 323 1.49 1.34 0.29 Development Dimensions Width Height Diameter Description Horizontal Development Ore Drives 5 5 5mW x 5mH Tonnes Cu Au Zn Trucking (t.km) 1,711,592 351 12,159 Tonnes Trucking # (m) (t) (t.km) Vertical Development Slots Included in Stope quantities Stoping Production Drilling Blasting Production Tonnes Cu Au Zn Trucking Backfill Top Down Stopes Bottom Up Stopes Stopes Not Filled Total (m) (m) 1,026,012 806,152 (kt) (%) (g/t) (%) 11,286 1.39 1.38 0.22 (t.km) 37,955,985 (m3,000) (m3,000) (m3,000) (m3,000) 1,665 611 596 2,872 329,410 3,269,956 4,482,944 4,059,358 4,912,182 4,089,368 3,920,294 4,640,413 2,875,629 2,366,166 1,811,158 1,199,107 15 343 185 138 214 97 310 92 134 101 326 210 85 32 328 227 60 2 289 15 343 323 253 53 307 117 41 57 215 9 3 179 191 135 135 84 84 1,249 1.38 1.58 0.28 832 1.38 1.33 0.28 744 1.42 1.18 0.30 514 1.45 1.26 0.33 323 1.49 1.34 0.29 7 7 VMS ‐ ROM Feed Supergene Tonnes Cu Au Zn (kt) (%) (g/t) (%) 1,802 1.58 1.27 0.06 36 1.53 1.21 0.07 243 1.41 1.11 0.04 594 1.44 1.13 0.04 266 1.31 1.09 0.05 220 1.62 1.29 0.09 69 2.01 1.14 0.14 229 1.98 1.66 0.08 147 2.02 1.89 0.07 Primary Tonnes Cu Au Zn (kt) (%) (g/t) (%) 10,666 1.34 1.38 0.24 34 0.96 0.75 0.15 411 1.08 0.92 0.18 1,255 1.29 1.28 0.24 987 1.33 1.39 0.12 982 1.35 1.30 0.24 1,269 1.35 1.67 0.22 1,070 1.25 1.40 0.22 996 1.36 1.42 0.23 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 195 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 18.16 Hassai South – Underground Stoping Schedule 2,000 1,800 1,600 1,400 Tonnes (kt) 1,200 1,000 800 600 400 200 0 Y1 Y2 Y3 Y4 Y5 Y6 Y7 Y8 Y9 Y10 Y11 Y12 Y13 Y14 Period Primary 18.4.8.2 Supergene Hadal Awatib Surpac was used to report quantities and grades, and custom-built Excel spreadsheets were used for the scheduling of the Hadal Awatib Project. In general, the following steps were undertaken in the scheduling process: • Definition of ore and waste within the pit limits using Surpac • Production of stage inventories using Surpac • Transfer of stage inventories to spreadsheet • Produce preliminary schedule. Although no stage designs were carried out, the pits have been divided in half (East-West) in an effort to try and bring some ore forward in the schedule. The Eastern end had a lower initial pre-strip, so this was used as Stage 1, with Stage 2 being the Western end. Hadal Awatib open pit mining schedules were developed to fill the shortfall from the underground schedule in the required VMS plant feed for each of the feed rate options. Summary schedule data is outlined in Table 18.28, and Figure 18.17 and Figure 18.18 are graphical representations of the mining profile for the 5 Mt/a schedule. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 196 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.28 Hadal Awatib – 5 Mt/a Mining Schedule Mining Schedule ‐ Hadal Awatib 5Mtpa Open Pit Unit Total m‐1 m‐2 m‐3 Y1 Y2 Y3 Y4 Y5 Y6 Y7 Y8 Y9 Y10 Y11 Y12 Y13 Y14 OPEN PIT MINING Waste Oxide Supergene Primary Total Ore Total Mining kbcm kbcm kbcm kbcm kbcm kbcm 28,926 152 2,216 1,911 4,127 33,206 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 6,376 0 0 0 0 6,376 9,133 63 80 0 80 9,276 6,005 33 498 159 656 6,694 3,449 56 903 23 926 4,431 2,349 0 730 194 924 3,273 962 0 5 870 875 1,837 652 0 0 666 666 1,318 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 Total Ore Total Mining kt kt SR 17,266 96,802 4.6 0 0 0.0 0 0 0.0 0 0 0.0 0 17,258 0.0 456 25,316 54.5 2,716 19,236 6.1 3,884 13,542 2.5 3,756 10,384 1.8 3,634 6,362 0.8 2,819 4,703 0.7 0 0 0.0 0 0 0.0 0 0 0.0 0 0 0.0 0 0 0.0 0 0 0.0 0 0 0.0 kt kt 32,008 64,794 0 0 0 0 0 0 12,646 4,612 12,729 12,587 6,528 12,708 105 13,437 0 10,384 0 6,362 0 4,703 0 0 0 0 0 0 0 0 0 0 0 0 0 0 Total Mining - By Stage stage1 stage2 CIP ‐ ROM Feed Oxide Tonnes Cu Au Zn (kt) (%) (g/t) (%) 372 0.09 4.29 0.02 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 152 0.08 3.47 0.03 81 0.02 3.44 0.01 139 0.13 5.70 0.03 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 Supergene Tonnes Cu Au Zn (kt) (%) (g/t) (%) 8,891 0.96 0.96 0.63 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 304 1.07 0.79 0.50 1,970 1.37 0.76 0.86 3,649 0.85 1.06 0.49 2,947 0.81 1.00 0.67 20 0.65 1.10 0.54 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 Primary Tonnes Cu Au Zn (kt) (%) (g/t) (%) 8,002 1.26 0.88 0.72 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 665 1.45 0.80 0.90 96 1.50 0.89 0.41 808 1.34 0.87 0.70 3,614 1.32 0.88 0.66 2,819 1.09 0.91 0.77 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 Supergene Tonnes Cu Au Zn (kt) (%) (g/t) (%) 1,802 1.58 1.27 0.06 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 36 1.53 1.21 0.07 243 1.41 1.11 0.04 594 1.44 1.13 0.04 266 1.31 1.09 0.05 220 1.62 1.29 0.09 69 2.01 1.14 0.14 229 1.98 1.66 0.08 147 2.02 1.89 0.07 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 Primary Tonnes Cu Au Zn (kt) (%) (g/t) (%) 10,666 1.34 1.38 0.24 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 34 0.96 0.75 0.15 411 1.08 0.92 0.18 1,255 1.29 1.28 0.24 987 1.33 1.39 0.12 982 1.35 1.30 0.24 1,269 1.35 1.67 0.22 1,070 1.25 1.40 0.22 996 1.36 1.42 0.23 1,249 1.38 1.58 0.28 832 1.38 1.33 0.28 744 1.42 1.18 0.30 514 1.45 1.26 0.33 323 1.49 1.34 0.29 0 0.00 0.00 0.00 Oxide Tonnes Au (kt) (g/t) 372 4.29 0 0.00 0 0.00 0 0.00 69 2.64 116 4.07 187 5.04 0 0.00 0 0.00 0 0.00 0 0.00 0 0.00 0 0.00 0 0.00 0 0.00 0 0.00 0 0.00 0 0.00 Supergene Tonnes Cu Au Zn (kt) (%) (g/t) (%) 10,694 1.07 1.02 0.54 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 2,501 1.35 0.86 0.59 3,678 0.97 1.03 0.48 3,284 0.87 1.03 0.61 697 0.96 1.04 0.57 368 1.56 1.40 0.28 166 1.97 1.84 0.10 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 Primary Tonnes Cu Au Zn (kt) (%) (g/t) (%) 18,668 1.30 1.17 0.44 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 1,799 1.28 1.08 0.38 1,322 1.34 1.22 0.28 1,716 1.35 1.15 0.41 4,303 1.33 1.09 0.54 4,632 1.17 1.05 0.60 1,235 1.33 1.35 0.30 1,249 1.38 1.58 0.28 832 1.38 1.33 0.28 744 1.42 1.18 0.30 514 1.45 1.26 0.33 323 1.49 1.34 0.29 0 0.00 0.00 0.00 TOTAL VMS Tonnes Cu Au Zn (kt) (%) (g/t) (%) 29,362 1.22 1.11 0.48 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 0 0.00 0.00 0.00 4,300 1.32 0.95 0.50 5,000 1.07 1.08 0.43 5,000 1.04 1.07 0.54 5,000 1.28 1.08 0.54 5,000 1.20 1.08 0.58 1,401 1.40 1.41 0.28 1,249 1.38 1.58 0.28 832 1.38 1.33 0.28 744 1.42 1.18 0.30 514 1.45 1.26 0.33 323 1.49 1.34 0.29 0 0.00 0.00 0.00 VMS ‐ ROM Feed UNDERGROUND MINING VMS ‐ ROM Feed PROCESS SCHEDULE CIP ‐ ROM Feed VMS ‐ ROM Feed FINAL – Rev 0 – 22 Oct 2010 AMEC Page 197 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 18.17 Hadal Awatib – 5 Mt/a Ore Profile 4,000 3,500 3,000 Tonnes (kt) 2,500 2,000 1,500 1,000 500 0 Y1 Y2 Y3 Y4 Y5 Y6 Y7 Y8 Y9 Y10 Y11 Y12 Y13 Y14 Y10 Y11 Y12 Y13 Y14 Period Primary Supergene Figure 18.18 Hadal Awatib – 5 Mt/a Mining Profile 30,000 25,000 Tonnes (kt) 20,000 15,000 10,000 5,000 0 Y1 Y2 Y3 Y4 Y5 Y6 Y7 Y8 Y9 Period Waste Oxide Primary Supergene FINAL – Rev 0 – 22 Oct 2010 AMEC Page 198 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.4.9 Mine Operating Costs Due to the lack of experience locally in Sudan, the assumption has been made that all underground mining will be carried out by a suitably qualified expatriate mining contractor. Underground mining costs have been derived from an AMEC Minproc internal cost database. For open pit mining, it has been assumed that an entirely new mining fleet will need to be purchased for the VMS open pit mining, and that this will be operated by the existing labour force. The excavator fleet assumed is similar to the largest diggers currently on site (120 t); however the trucking fleet has been increased to 90 t trucks, which are a better fit for this size of excavator. Open pit mining costs have been derived from the historical AMC site operating cost data along with the AMEC Minproc internal cost database. 18.4.9.1 Hassai South A breakdown of the allowances is as follows: • Horizontal Development: operating horizontal development cost includes all ore drive development costs. • Vertical Development: vertical development cost includes all stoping slot development costs. Slots have been assumed to be developed using a raise bore between 30 m levels. • Drill and Blast: includes all stope production and drilling costs. Development drill and blast costs have been included in the horizontal development costs. • Material Movement: material movement cost includes all costs associated with the bogging and trucking of both development and stope tonnes. • Backfill: costs include all costs associated with the backfilling of stopes with paste fill. The unit operating cost applied includes an allowance for the extension of paste fill lines. In an effort to reduce costs, paste fill cost was not included for the crown pillar retreats. The assumption was made that rib pillar locations could be found that wouldn’t impact significantly on the total stope tonnes and grades, thus no additional ore loss was applied for the crown pillar removals. Although this was considered reasonable for a scoping study evaluation, moving forward an assessment will need to be carried out to determine the method for crown pillar removal, and what additional measures may be required (ie rib pillar sizes, fill every second stope, cost saving versus ore loss trade-off, etc.). • Mine Services: mine services is an allowance for all underground service related items such as the installation, operation, maintenance, relocation and removal of the following: − Secondary ventilation − Compressed air lines and compressor costs − Water and pump lines and pump costs − Power feed lines and supply costs − Communication lines − Firing lines FINAL – Rev 0 – 22 Oct 2010 AMEC Page 199 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report • • Supervision and Control: an allowance for all management of the underground operations such as: − Mine management − Contract management − Supervision − Mining engineering − Survey control − Geological control − Safety/environmental Owner’s costs for the management and control of the operations have also been allowed. Summary unit cost data is outlined in Table 18.29. Table 18.29 Unit Operating Costs – Hassai South Underground VMS Ore Mined (Mt) 12.5 Unit Operating Costs Horizontal development ($/t ore) 2.98 Vertical development ($/t ore) 0.98 Drill and blast ($/t ore) 4.23 Material movement ($/t ore) 6.02 Backfill ($/t ore) 1.96 Mine services ($/t ore) 3.00 Supervision and control ($/t ore) 7.00 Total – Underground ($/t ore) 26.17 18.4.9.2 Hadal Awatib Open Pit – Operating Costs • Drill and Blast: site unit costs were used for drill and blast cost; drilling cost is assumed to include grade control drilling. The following unit costs were applied: − Drilling $0.36/t − Blasting $0.21/t Applying these unit costs equates to approximately $1.66/bcm mined. • Load and Haul: load and haul costs have been built up from the AMEC Minproc internal cost database, from typical hourly operating costs. Unit costs applied excluded fuel and labour, as these were tabulated separately. Cycle times were calculated using approximate haulage destinations for both ore and waste • Ancillary: ancillary equipment costs have been built up from the AMEC Minproc internal cost database, applying typical hourly operating costs. Unit costs applied excluded fuel and labour, as these were tabulated separately • Ore Overhaul to Plant: site unit costs were used for ore overhaul cost ($0.21/t.km). The average haul distance applied was 12 km, and it was assumed this rate included labour and fuel FINAL – Rev 0 – 22 Oct 2010 AMEC Page 200 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report • Fuel: typical fuel burn rates for the mining fleet were used to calculate the fuel usage, and the fuel cost supplied by site was applied ($0.41/L) • Hydrology: an allowance for pit dewatering, ground support and activities associated with mining significantly larger and deeper pits than is currently underway • Labour: a typical labour force was built up based on the size of the operations and numbers of mining fleet. Local labour rates provided by AMC were applied to calculate the total labour cost • General: the majority of the general mining costs were included in the plant G&A cost, eg. laboratory services, accommodation and flights, environmental, contractors and consultants, freight and logistics, emergency response and ERT, and safety. However, the following has been allowed for under this line item: − Survey and GC consumables − Safety and training consumables − Mining software and computing upkeep − Open pit office costs − Mine dispatch support. Adjustments – Hadal Awatib Operating Costs It has been assumed that 100% of the fleet operating costs are attributed to Hadal Awatib, as the equipment is assumed to be fully utilised in that pit. Table 18.30 outlines the unit operating cost data for the 5 Mt/a schedule. Table 18.30 Unit Operating Costs – Hadal Awatib Open Pit CIL ore mined (Mt) 0.4 VMS ore mined (Mt) 16.9 Waste mined (Mt) 79.5 Total Mined (Mt) 96.8 Unit Operating Costs Drill and blast ($/t ore) 2.33 Load and haul ($/t ore) 3.06 Ancillary ($/t ore) 2.07 Ore overhaul to plant ($/t ore) 2.58 Fuel ($/t ore) 1.79 Hydrology ($/t ore) 0.39 Labour ($/t ore) 1.47 General ($/t ore) 0.44 Total – Open Pit ($/t ore) 14.14 ($/t) 2.47 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 201 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.4.10 Mine Capital Costs As the underground operating costs have been based around typical contract rates, equipment capital recovery is built into the contractor rates, and, as such, no allowance is made for mining fleet purchase/sustaining capital. Costs associated with underground infrastructure and the capital development (such as decline development) have been included as capital items. For open pit mining, it has been assumed that an entirely new mining fleet will be purchased for the VMS open pit mining, and that this will be operated by the existing labour force. It should be noted that no significant additional costs have been allowed for landing new equipment in Sudan (ie. import taxes, duties, etc.). This needs to be investigated in the next level of study. 18.4.10.1 Hassai South The capital cost summary for Hassai South underground is outlined in Table 18.31. Table 18.31 Capital Cost Summary – Hassai South Underground VMS Ore Mined (Mt) 12.5 Capital Costs Infrastructure ($M) 31.2 Horizontal Development ($M) 21.7 Vertical Development ($M) 2.1 Material Movement ($M) 2.6 Total – Underground ($M) 57.6 Infrastructure ($/t ore) 2.50 Horizontal Development ($/t ore) 1.74 Vertical Development ($/t ore) 0.17 Unit Capital Costs Material Movement ($/t ore) 0.21 Total – Underground ($/t ore) 4.62 Costs associated with infrastructure and the capital development (such as decline development) have been included as capital items. The following items have been allowed for in the capital cost estimate. • Infrastructure: infrastructure has an allowance for the following capital items: − Preliminary works: geotechnical, contractor mobilisation, contractor demobilisation, rehabilitation and raise bore mobilisation − Surface works: repair to site roads, stabilise/repair pit wall failures, raw water supply, potable water supply, waste dump works − Buildings: offices, workshop, surface magazine, wash-down bay, air compressor, and communications − Office equipment: furniture, computer hardware, computer software, survey equipment, and ventilation equipment − Safety equipment: emergency response, cap lamps, self rescuers, and ER vehicle − Light vehicles FINAL – Rev 0 – 22 Oct 2010 AMEC Page 202 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report − Portal: decline portal works − Ventilation: main ventilation fans and equipment, regulators, vent doors (drive through), pump stations, electrical supply and communications − UG infrastructure: refuge chambers, ladder ways, refuelling bay, service bay, magazine, crib room, and toilets − Paste plant: including borehole and initial pipe-work • Horizontal Development: capital includes all decline, incline and link drive development costs, as well as all access drives, and ventilation drives • Vertical Development: capital includes all ventilation and ladder way rises • Material Movement: material movement costs include all costs associated with the bogging and trucking of capital development tonnes. 18.4.10.2 Hadal Awatib It has been assumed that an entirely new mining fleet will need to be purchased for the VMS open pit mining, comprising 120 t excavators and 90 t trucks. Replacement capital has been included as required when the replacement hours have been surpassed. Table 18.32 outlines the capital cost data for the 5 Mt/a schedule. Table 18.32 Capital Cost Summary – Hadal Awatib Open Pit Options CIL ore mined (Mt) 0.4 VMS ore mined (Mt) 16.9 Waste mined (Mt) 79.5 Total Mined (Mt) 96.8 ($M) 5.0 Capital Costs Infrastructure Hydrology ($M) 0.0 Equipment ($M) 79.6 Mine Services Capital ($M) 0.0 Mine administration and technical operating ($M) 0.0 Total – Open Pit ($M) 84.6 Infrastructure ($/t ore) 0.30 Hydrology ($/t ore) 0.00 Equipment ($/t ore) 4.71 Mine services capital ($/t ore) 0.00 Mine administration and technical operating ($/t ore) 0.00 Total – Open Pit ($/t ore) 5.00 Unit Capital Costs FINAL – Rev 0 – 22 Oct 2010 AMEC Page 203 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.5 GEOTECHNICAL INPUT 18.5.1 Kamoeb South – AMC At least 16 samples were previously taken to assess the density of the wall rock, and an average of 2.8 g/cm3 was calculated. RQD information was routinely measured during core logging. A 3-D surface was modelled to correspond to RQD >95%. In addition to economic parameters, slope design was also defined according to rock type and RQD surface. The parameters proposed for Kamoeb South are as follows: • Slope angle between 40° and 60°, depending on the wall • Final bench height: 10 m • Bench height: 2.5 m, rarely 5 m • Bench angle: 55° to 80° • Ramp width: 23 m; 15 m for the last 2 benches • Ramp slope angle: 8%; 10% for the last 2 benches. The pits are designed to accommodate 40 to 60 t trucks. 18.5.2 AMEC Geotechnical Input – Introduction A site visit was made by Adam Coulson of AMEC to inspect conditions in the existing open pits, particularly Hadal Awatib East and Hassai South. Limited bench scale mapping was undertaken to confirm rock mass classification and major joint set orientation, while eight diamond drill hole cores were reviewed. Data was also collected and reviewed from previous geotechnical studies, while discussions were held with the mine manager and senior mine geologist. From this information, scoping study level rock mass classification for the Q-system (Barton et. al., 1974), the CSIR Rock Mass Rating (RMR) classification system (Bieniawski, 1976 &1989), Geological Strength Index (GSI) (Hoek et. al., 1995) and Laubscher's Mining Rock Mass Rating (MRMR) (Laubscher, 1990) have been made for the two key domains at Hadal Awatib East and Hassai South. 18.5.3 Geotechnical/Geological Domains 18.5.3.1 Domain 1: Green Chlorite Schist (SCHI) Metamorphosed andesite, foliated parallel to the orebody strike and dip. stronger at Hassai South. This unit forms the wall rock at both deposits. 18.5.3.2 Foliation appears to be Domain 2: Volcanogenic Massive Sulphides (VMS) The massive sulphides forming the orebodies are generally fine grained, pyritic and homogenous through the centre of the intersections but can be interbedded or disseminated at the wall rock contacts. Significant microfracturing exists in the fine grained VMS, and may be a result of post extraction oxidisation. At Hadal Awatib, 1 cm-spaced fracturing perpendicular to the core axis may indicate a low tensile strength of the material, or high stress. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 204 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.5.3.3 Domain 3: Diorite and Basic Dykes (DYK) These are identified at both deposits, but are more prevalent at Hadal Awatib. These dykes are generally 1-2 m thick, but may exist in swarms. These dykes cut the mineralisation and do not appear to cause a structural problem for pit wall stability, but could be a source of stope dilution. 18.5.4 Major Joint Set Orientation Based on preliminary baseline bench mapping, the major joint sets that govern stability of the pit walls and of the underground deposit have been identified in Figure 18.19 with joint orientation summarised in Table 18.33. It should be noted that these joint orientations pertain to the wall rock; during the site visit it was not possible to observe the jointing within the VMS, which can only be obtained for oriented core or transverse exposure. Consequently, jointing in the VMS has been assumed. At both locations the dominant joint set is related to the foliation joint set which appeared to be more strongly defined at Hassai South, followed by a vertical to sub-vertical joint set and a horizontal to subhorizontal joint set generally dipping at 10o to the north. A sporadic but relatively persistent sub-vertical joint set was identified, and may have been the potential initiator of the wedge/plane failures that have occurred in the south walls of some existing pits (eg. Hassai South southern pit wall). Table 18.33 Summary of Probable Major Joint Set Orientations Typical Pit Domain Joint Set Spacing (m) o Dip o ( ) ( ) Hadal Awatib Green Chlorite Schist 1 (Sub Vertical – Foliation) 0.5 to 1 073 65 North and South Pit Walls 2 (Horizontal) 5 to 10 213 10 HW, Central and FW Rock 3 (Vertical – Joints and Dykes) 2 to 4 223 79 4 (Sub Vertical – Sparse/Random) > 10 213 38 Hassai South Green Chlorite Schist 1 (Vertical – Foliation) 0.1 to 0.75 083 63 North and South Pit Walls 2 (Horizontal) 2 to 3 273 10 HW and FW Rock 3 (Vertical – Joints and Dykes) 2 to 4 314 73 4 (Sub Vertical – Sparse/Random) > 10 293 48 Hassai South VMS 1 (Sub Vertical – Orebody Dip) 3 090 60 2 (Horizontal) 5 270 10 3 (Vertical) 5 180 90 --- --- (Typical Assumed Orientations) Random 1 Strike1 Strike is based on the right-hand rule. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 205 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 18.19 Bench Face Mapping (a) and (b) poles and contours for Hassai South , (c) and (d) Hadal Awatib a. b. HAS HAS Foliation Horizontal Joints Sub Vertical Joints Dykes and Vertical Joints c. d. HAE HAE Foliation Horizontal Joints Sub Vertical Joints Dykes and Vertical Joints The assumed joint set orientations in the VMS forms the basis for determination of open stoping dimensioning using the Mathews-Potvin Open Stope Stability graph method. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 206 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.5.5 Material Testing At present, limited material testing (11 uniaxial compressive tests, 8 Brazilian tensile, 5 direct shear tests and 12 basic friction angle tests) has been performed at Ariab, and most tests have concentrated on the weaker gossan and chlorite schists, with no testing performed on fresh VMS. The summary of these test results as they are stated in the INTECSA-INARSA report (2002), including assumed values for the VMS, are summarised in Table 18.34. Table 18.34 Summary of Previous Laboratory Testing Rock Type Density UCS Tensile Young's (t/m3) (MPa) Strength Modulus Direct Shear C (MPa) (GPa) (MPa) Reference phi SCHI 2.65 - 2.69 110 9.9 72.5 0.02 28 Intecsa-Inarsa (2002) Clayey Schist 2.1 - 2.4 16 13.7 19.6 --- --- Intecsa-Inarsa (2002) Gossan 2.3 <15 --- --- --- --- Intecsa-Inarsa (2002) VMS 4.5 - 5 150 --- --- --- --- Assumed AMEC 2010 18.5.6 Rock Mass Classification Rock mass classification for the open pits was performed from bench mapping with additional review and basic logging of selected drill hole cores from both orebodies. Core in the VMS and contacts was either half split or quarter split, making exact verification of RQDs within this unit difficult, however, information on natural joint conditions could be determined. It is important to note that these are only preliminary estimations and for further studies additional geotechnical data should be obtained through bench mapping, geotechnical logging of oriented core, hydrogeological investigations and additional rock strength testing. The summarised average joint properties based on Barton's Q-system and determined rock mass classification based on RMR (Bieniawski, 1976,1989), and GSI (Hoek et. al., 1995) from the core logging are summarised for the two VMS deposits on Table 18.35 and Table 18.36. Table 18.35 Hadal Awatib – Summary of Rock Mass Properties by Domain and Stope Zone Borehole Rock Type FF AVG RQD Jn Jr Ja Q' RMR'76 (Joint/m) AVG Assumed AVG AVG AVG (%) RMR' GSI AVG 89 AVG Calc AVG Hadal Awatib HW SCHI 1.25 72 12 1.6 1.4 7.2 61 69 64 Combined HW VMS 7.17 72 12 1.5 0.8 12.4 63 63 58 Per Zone MID SCHI 1.12 91 12 1.6 1.1 12.6 66 72 67 FW VMS 3.95 80 12 1.6 0.8 13.8 67 70 65 FW SCHI 1.32 85 12 1.5 1.1 9.9 64 75 70 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 207 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.36 Hassai South – Summary of rock Mass Properties by Domain and Stope Zone Borehole Rock FF AVG RQD Jn Jr Ja Q' RMR'76 RMR' GSI Type (Joint/m) AVG Assumed AVG AVG AVG AVG 89 AVG Calc AVG (%) Hasai HW SCHI 1.47 96 12 1.4 1.2 10.5 65 74 69 South Ore VMS 2.48 90 12 1.6 0.8 14.4 68 71 66 Combined FW SCHI 2.83 92 12 1.4 1.4 7.7 62 67 62 Per Zone DVMS 2.76 92 12 1.1 1.5 5.5 59 64 59 1 Note the average RQDs (based on AMC logging), and joint properties are based on the averages for each domain. The joint number (Jn) of 12 (3 joint sets plus random) has been assumed for all domains based on the major joint sets identified from mapping, except for the Dyke in which a Jn of 9 (3 joint sets) has been assumed. The values for Q' are determined on the average joint characteristic per interval, and the value of RMR'76 is calculated using Bieniawski's equation (RMR = 9LogeQ' + 44). The values of RMR'89 are calculated independently based on the average joint and rock mass characteristic per interval, and used through the GSI (GSI = RMR'76 = RMR'89 – 5), to compare agreement of values. Verification of the RQDs on intact core was performed and it was determined that the values determined by AMC are reasonable. As can be seen for both deposits, the overall rock mass rating for VMS is slightly greater than the wallrock SCHI. The RQD values tend to be lower in the VMS, however, this is offset by better joint conditions than in the SCHI. The ore zone can be classified as a Fair to Good rock mass based on the Q-system and the SCHI can be classified as a Fair rock mass. Overall the VMS ore zone at Hassai South is slight better quality than at Hadal Awatib, while the converse is true of the SCHI. The relevance of the core discing identified at Hadal Awatib needs to be verified with material testing, as does the appearance of microfracturing. At this stage there is no significant difference in the rock mass quality between the supergene and the primary zones. These general values will be used for the scoping-level underground mining design. The rock mass classification for the open pits is based on the bench mapping and using the average RQDs determined from the core logging. The summary of the bench joint mapping for each deposit is summarised in Table 18.37. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 208 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.37 Summary of Bench Face Rock Mass Classification Wall Hadal Awatib Link South Hassai South RQD Rock Assumed Type (%) RMR'8 Jn Jr Ja Q' RMR'76 Assumed AVG AVG AVG AVG Calc 12 1.6 1.2 10.3 65 GSI MRMR1 AVG AVG 71 66 45 9 AVG HW SCHI 90 DIO DYK 85 9 1.5 1.3 11.3 66 64 59 40 2 10 12 1.0 3.0 0.3 33 --- 33 20 HW SCHI 90 12 1.7 1.3 9.6 65 72 67 45 Fault 10 15 1.0 10.0 0.1 20 34 29 13 Gossan South Wall 1 2 MRMR Adjustment Factors: Weathering = 1 and 0.9 (Gossan); Joint Orientation = 0.85; Blasting = 0.8; Stress = 1.0 A single Gossan exposure was reviewed on the West wall of Hadal Awatib by the ramp and values have been estimated. The rock mass ratings agree relatively well with those obtain from the core logging for the SCHI and have been used as the basis for empirical slope evaluation. 18.5.6.1 In Situ Stress Regime At present, no in situ stress data exists for the region, however, the majority of the major faults trend NW-SE and would infer a maximum principal stress direction oriented WNW-ESE, to initiate shear deformation and folding. However, for this study a worst case local stress orientation for the Hassai South deposit has been assumed to be oriented in a north-south direction, based on similar observed maximum principal stress orientation perpendicular to the orebody strike and foliation such as at Brunswick Mine (Canada) and Mount Isa Mines (Australia). Generally, for tabular orebodies similar to the Hassai VMS deposit the eventual induced stress direction becomes oriented perpendicular to the orebody strike after progressive mining. Additionally, the present pit will have disturbed the in situ stress orientation such that a north-south direction is more appropriate. The stress magnitudes are also unknown, and are assumed based on experience, for assessment of stope dimensioning (Table 18.38). Table 18.38 Summary of Assumed in situ Stress Regime Principal Stress 1 Mag (MPa/m) K MPa @ (σ1/σ3) 300 m Depth 000 1.5 12.1 Plunge Trend 0 σ1 (Horiz) 0.0405 σ2 (horiz) 0.0324 0 090 1.2 9.7 σ3 (Vert)1 0.027 90 000 1 8.1 Average Overburden Density = 2.75 t/m FINAL – Rev 0 – 22 Oct 2010 AMEC Page 209 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.5.7 Design Criteria for Hadal Awatib and Hassai South Open Pits For this level of study detailed slope stability analysis is not warranted based on the present level of information. Preliminary assessment of potential inter-ramp angles (IRAs) has been performed based on the empirical pit slope design chart developed from open pit experience in the mining industry (after Laubscher, 1990). This acts as a basis for assessment and comparison to existing pit slopes and performance, but is only a guide and not a definitive design tool. Based on estimated maximum pit depths at Hadal Awatib and Hassai South of approximately 170 m and 150 m respectively, the basic IRAs, including a design factor of safety of approximately 1.35, are as follows: • Gossan 44o • Hadal Awatib 47o • Hassai South 49o. These IRAs are broadly in-line with previous studies for Hadal Awatib (INTECSA-INARSA (2002) and Hadayamet (ANETA, 1999)) in which limit equilibrium analysis was performed (albeit based on limited testing and mapping data), determining that for a 90 m deep pit the maximum IRA in SCHI should not exceed 50o to 55o, and for gossan 43o to 46o. The pit slope design criteria has been further refined based on the preliminary joint orientations and review of the existing pits in the following sections. The design criteria have been based on the current mine fleet and operating practice which includes 10 m high benches and bench widths varying from 4 to 8 m, dependent on rock mass conditions. A review of conditions in the existing Hadal Awatib and Hassai South pits indicates that current design criteria have worked relatively well, with the following observations: • Hadal Awatib A minor bench wedge failure was noted in the South wall, but is localised to two benches and, based on the clean catch benches below, was probably noted and removed during the final wall mining. A review of the existing bench face or batter angles, bench widths and IRAs was made. Preliminary kinematic stability assessment suggests that the face angle in the North wall should be matched to the foliation dip to prevent sliding failure, while in the South wall toppling failure potential exists and the bench face angle can be increased but a larger berm width should be developed. Generally, the actual bench face angle in the North wall has been matched to the dominant foliation (Table 18.39 and Table 18.40), which is standard practice, and the South wall has been steepened, with an increase in the berm width. Based on the performance of the North wall at Hassai South and the slightly lower wall height of the North wall (130 m versus 170 m), there is the potential to increase the IRA of the North wall by reducing the berm width. One issue with the reduction in berm width to 4 m is that wall control becomes important for the final pit wall, as this narrow width can be significantly reduced with back break, making the function of the catch berm ineffective. Controlled blasting using a pre-shear is recommended for the North wall with the reduced berm width. Other potential options would be to increase the bench height to 15 m, such that the berm width could be increased to 6 m. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 210 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.39 Hadal Awatib – Summary of Existing and Proposed Slope Design Crieria N. Wall Hadal Awatib (Hadal Awatib) Units S. Wall Actual N. Wall N. Wall Actual S. Wall S. Wall Avg Gossan SCHI Avg Gossan SCHI SCHI Bench Height m 10 10 10 10 10 10 Berm (Bench) Width m 5.2 5 4 6.78 7 6.5 Bench Berm Spacing Batter Angle (BFA) Wall Height Inter Ramp Wall Angle (IRA) • SCHI # 1 1 1 1 1 1 deg 66.4 65 65 77.6 75 75 m 70 30 130 100 30 170 deg 45 46.0 49.1 48 45.9 47.4 Hassai South Review of the Hassai South pit indicates that the current pit design criteria has worked relatively well for the north wall of the Hassai pit. However, the south pit wall has suffered two failures, the smaller being a two to three bench wedge/plane failure in the west end of the pit and the larger failure over five to six benches in the east end. The former failure is based on the formation of a wedge developed on sub-vertical jointing, and is relatively minor in comparison to the later failure which could have occurred through a number of factors: failure on a similar structure with toppling failure, failure on a fault or shear, or over-steepening of the South pit wall beyond a stable slope angle. A review of the actual bench face angle, berm widths and IRAs for a typical section is summarised, with the design recommendations for pit deepening in Table 18.40. As can be seen, the South wall was developed relatively steeply, with limited berm widths - some of these were noted to be under 3 m. Plane failure along the foliation set will dominate in the North wall and requires flattening the bench face angle to the foliation. Toppling failure on the same set in the South wall, in which the face angle could have been steepened, however, the berm width should also have been increased. This was not the case here for the South wall. This failure was reported to occurred close to a year after completion, but was not the result of a precipitation event. A tension gash at the crest of the pit was identified and monitored up until the pit was completed, but was not monitored following this and the exact failure date is unknown. Based on the identification of the sub-vertical dipping structure, which may have been an initiator, the proposed IRA for the South wall has been reduced to below this, such that similar failure would only result in single bench failures. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 211 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.40 Hassai South – Summary of Existing and Proposed Slope Design Criteria N. Wall Hassai South Units Actual Avg S. Wall N. Wall SCHI Actual Avg SCHI S. Wall SCHI SCHI Bench Height m 10 10 10 10 Berm (Bench) Width m 4.2 4 3.2 7 Bench Berm Spacing Batter (Bench Face) Angle Wall Height Inter Ramp Wall Angle 18.5.8 # 1 1 1 1 degrees 63.9 64 67.4 75.00 m 90 140 100 150 degrees 47.5 48.40 54.2 45.93 Design Criteria for Hassai South Underground Open Stoping The general thickness (5 to 40 m), dip of the orebodies (60o to 70o), and lateral and vertical extent, suggests that sub-level open stoping (SLOC) with backfill is the most suited mining method. Stope dimension recommendations are based on a combination of the empirical Modified Stability Graph method (Potvin, 1988; Nickson 1992, Hadjigeorgiou et. al., 1995) and estimation of stress through numerical stress modelling or analytical methods. Preliminary stope dimensions are based on the former with an estimate of the induced stress conditions based on experience using the Kirsch stress approximation (Hoek and Brown, 1980). Underground mining of the Hadal Awatib deposit has not been considered at this time, primarily due to the complexity of the multiple lenses. Hassai South contains a supergene zone of around 20 to 30 m thickness below the bottom of the pit with the bulk of the orebody contained within the primary ore zone (Figure 18.20). Underground mining consideration has been given to preferential extraction of the supergene zone directly below the bottom of the pit in the early stages of mining. In order to achieve, this backfilling of smaller stopes with rib pillars will be essential in order to maintain stability of the overhanging South pit wall. Additionally, paste backfill is recommended to reduce the potential for surface water infiltration through storm events, which could flood the bottom of the pit. A nominal temporary crown pillar below this zone has been considered to allow separation of the upper supergene zone - which can be mined in a top-down sequence - from the lower zone which could commence in a bottom-up primary-secondary stope sequence. This pillar will add additional stability to the zone and provide an additional barrier against potential water infiltration. The intention would be to mine this lower grade pillar at the completion of mining at a lower extraction ratio. The advantages of a bottom-up sequence versus top-down, are increased stability and lower binder (cement) costs for stopes. 18.5.8.1 Open Stope Dimensioning Based on preliminary core logging, simplified rock mass properties have been assumed (Table 18.41). As joint orientation data is not known for the ore zone and is limited for the country rock, two rock mass qualities based on the number of potential joint sets have been considered to obtain an upper and lower boundary for the ore and country rock. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 212 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 18.20 Hassai South Underground Deposit Development Super Gene Zone Crown Pillar Primary Zone Table 18.41 Summary of Simplified Design Rock Mass Properties Stope Zone and Rock Unit UCS RQD Jn Jr Ja Q' RMR'76 (MPa) Ore Zone (VMS) Upper Boundary 150 88 9 1.5 0.75 19.6 71 Ore Zone (VMS) Lower Boundary 150 88 12 1.5 0.75 14.7 68 HW/FW SCHI Upper Boundary 110 100 9 1.3 1.25 11.6 66 HW/FW SCHI Upper Boundary 110 100 12 1.3 1.25 8.7 64 This range of Q' values is used with the Modified Open Stope Stability factors to determine a stability number, N'. For this analysis the A=factor - which is related to induced stress and the intact strength (UCS) of the rock - has been assumed as 1.0 for hanging walls, indicating a relaxed stope surface under low stress. A stress factor of 2 (based on the Kirsch equations, Hoek an Brown, 1980) has been applied to determine induced stresses for a mean average depth of 300 m below surface. The current deepest stope considered in this study is 420 m below surface. Evaluation of the grade cut-off shells applied to the zone indicates that the hanging walls of the stope can vary in dip between 60o and 70o. This change in the dip affects the stability of the hanging wall and has also been evaluated in determination of potential stope dimensions. The determined stability numbers and calculated recommended hydraulic radii (HR, =Area/Perimeter) based on unsupported and supported stope surfaces are summarised in Table 18.42, and have been plotted on the empirical stability graph Figure 18.21. It should be noted that the design line for the unsupported case is in the unsupported transition zone and thus assumes only temporary stability of one to two months, which can be achieved using paste backfill. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 213 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.42 Summary of Stope Surface Stability Criteria Lower Bound Case Rock UCS Type (MPa) HW Criteria (60 ) SCHI 110.0 8.7 1.0 0.3 5.0 13.0 8.2 10.3 HW Criteria (70 o) SCHI 110.0 8.7 1.0 0.3 5.9 15.3 8.7 10.6 Back Criteria – Ore VMS 150.0 14.7 0.55 0.2 2 3.2 5.2 8.4 End Wall Criteria - Ore VMS 150.0 14.7 0.55 0.3 8 19.4 9.4 --- o Q' A B C N' HR HR Unsupported Supported Upper Bound Case HW Criteria (60 o) SCHI 110.0 11.6 1.0 0.3 5.0 17.3 9.0 10.7 HW Criteria (70 ) SCHI 110.0 11.6 1.0 0.3 5.9 20.5 9.6 11.0 Back Criteria – Ore VMS 150.0 19.6 0.55 0.2 2.0 4.3 5.7 8.8 End Wall Criteria - Ore VMS 150.0 19.6 0.55 0.3 8.0 25.8 10.4 --- o Figure 18.21 Hassai South Open Stope Stability Chart Design Guidelines Modified Stability Graph (after Potvin, 1988 modified, Nickson 1992, Hadjigeorgiou et. al, 1995) 1000 Stable Zone Unsupported Design Line Stability Number, N' 100 Unsupported Supergene HW Transition Zone Supported Transition Zone 10 Stable With Support Upper Bound Lower Bound Caved Zone 1 HW - 60 deg HW - 70 deg Supported Design Line Back Ends 0.1 0 5 10 15 Hydrauliuc Radius (m) 20 25 (ref. Hoek et. al., 1995) FINAL – Rev 0 – 22 Oct 2010 AMEC Page 214 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Based on these hydraulic radii, recommended stope dimensions have been determined as follows: Hanging Wall Dimensions The potential hanging wall dimensions for an assumed sub-level interval of 30 m (floor to floor) considering 5 m high drifts are: Unsupported Case − 70o Hanging wall: Unsupported Strike Length = 33 to 41 m − 60o Hanging wall: Unsupported Strike Length = 28 to 34 m Supported Case (Garford Cables) − 70o Hanging wall: Supported Strike Length = 51 to 56 m − 60o Hanging wall: Supported Strike Length = 43 to 48 m Recommendations: from a cost perspective, an unsupported hanging wall is preferable, hence for this study a recommended maximum strike length of 30 m has been considered for primary (and secondary) stopes. Back Dimensions Based on the hanging wall evaluation, assuming stopes will be 30 m along strike: Unsupported Case Stope Strike 30 m = Unsupported Ore Thickness = 16 to 18 m Supported Case (Garford Cables) Stope Strike 30 m = Supported Ore Thickness = 38 to 42 m Recommendations: Panel stopes if ore thickness > 17 m (ie mine the hanging wall stope at a maximum ore width of 17 m, then backfill, followed by mining and backfilling of the footwall stope) or install cable bolt support (Garford Cables) and mine full width up to 40 m. End Walls Based on a sub-level spacing of 30 m: Unsupported Case Stope Sub-level Spacing 30 m, Unsupported Ore Thickness = 50 to 68 m Recommendations: Stope end walls should not be a problem and should be stable based on the current assumptions. For the supergene zone the sub-level spacing has been reduced to 20 m and smaller stopes will be mined to increase stability below the bottom of the open pit on the South pit wall. Additionally, as these stopes will be mined using a top-down sequence with in-pit drilling, mining the stopes below this zone will require mining under backfill, and thus a smaller dimension is preferable to reduce the potential for failure of the paste fill. As no back will be maintained, the critical surface is the hanging wall, and FINAL – Rev 0 – 22 Oct 2010 AMEC Page 215 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report stopes have been reduced to 20 m along strike. Assuming a hanging wall dip of 60o the resultant HR = 5.7. Additionally, in order to maintain global stability of the South wall for the current study, 10 m wide (along strike) rib pillars have been proposed every 40 m along strike. Below the bottom of the supergene zone, a crown pillar of 20 to 30 m high is recommended. For the present study these pillars have not been evaluated for stability, and would require numerical stress modelling for confirmation. 18.5.8.2 Other Mining Considerations At the present time the portal location is assumed to be located in the far west end of the pit continuing on from the current pit ramp. At this time no consideration has been given to the general stability of the ramp or portal in relation to mining of the supergene mineralisation, or to the rock mass conditions. Consideration as to the stability of the main access in relation mining should be undertaken as part of any further studies. A preliminary mining block sequence has been assumed for the study, but detailed stope sequencing has not been considered and requires more detailed rock mechanics data for assessment in combination with numerical stress modelling. This should also include assessment of the underground infrastructure in relation to mining. One key consideration that has not been addressed in this study is the effective mining below the present wall failure in the open pit. Potential options that need to be considered are the stabilisation of the upper wall of the open pit around the failure, through a partial cut-back of the affected zone, and cleaning and removal remotely of failed material with stabilisation of the pit wall using ground support, or a reduced recovery of the immediate ore below this failure. 18.6 HYDROGEOLOGY AND HYDROLOGY INPUT Since mining to-date has taken place above the water table, no hydrogeological studies of the mine area have been undertaken. Minor water is present in the base of some pits, and the appearance of VMS ore marks the water table level. During open pit mining at Hadal Awatib East and Hassai South, subterranean water appeared in the supergene zone, between 10 and 20 m above the massive sulphides. Little data is available for flowrates; values of 13 m3/h and 3-5 m3/h were recorded for Hadal Awatib and Hassai South, respectively; transmissivity was weak. Chemical analyses of the water are shown in Table 18.43. Groundwater is strongly acid, and high in sulphates and iron, with often high Cu and Zn present, as expected. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 216 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.43 Groundwater Chemical Analysis, Hadal Awatib and Hassai South Parameter Hadal Awatib pH Fe (mg/L) Hassai South 1.9 1.22 - 2.89 64 000 10 000 – 35 000 Cu (mg/L) 21 410 - 1610 Zn (mg/L) 1515 251 - 1220 Mg (mg/L) 2800 SO4 (mg/L) 110000 Open pit wall design assumes dewatered/depressurised wall conditions, while underground design assumes minimal pumping of water from the workings. These assumptions need to be confirmed as part of any feasibility study for the VMS concentrator project. The climate is very dry with rare precipitation. However, wadis are present, marking intermittent surface drainage channels, and need to be taken into account in mine, plant and waste management design. 18.7 SEISMICITY A preliminary assessment of the seismic risk at the Hassai Project site was undertaken. Although Sudan is generally characterised by low seismic activity, several large earthquakes which have resulted in loss of life and damage to property have been recorded. A search of the USGS database covering the period 1973 to date identified 157 events of magnitude >M4 within 600 km of the site. However, the vast majority of these were associated with the main East African Rift following the Red Sea, with epicentres typically >300 km from site. The site can therefore be described as being relatively aseismic. It is unlikely that events associated with the Red Sea fault system will be felt at site. The peak ground acceleration (pga) at site associated with an M4.3 event, 300 km away, will probably be less than 0.01 g (based on typical attenuation laws). Nonetheless, apparently random events do occur, such as a single M4.3 event some 180 km from site in 1973. It is difficult to predict where and when such events will occur. For that reason, a pga of 0.05 g has been adopted for conservative reasons for preliminary feasibility design purposes. This is slightly more than that inferred by the USGS Hazard map for Africa and Europe. 18.8 PROCESS PLANT DESCRIPTIONS 18.8.1 CIL Plant The process plant is designed to process 3.0 Mt/a, comprising of 2.0 Mt/a of heap leach residue and 1.0 Mt/a of fresh ore. The flow sheet is based on conventional comminution and CIL processes. The plant design is based on ore feed grades of 5.0 g/t Au and 4.0 g/t Ag for the fresh ore, and 1.5 g/t Au and 1.2 g/t Ag for the reclaimed heap leach material. Leach extractions of 90.2% Au and 70.0% Ag from all ore blends were FINAL – Rev 0 – 22 Oct 2010 AMEC Page 217 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report assumed in order to size the elution circuit and goldroom. A gravity recovery circuit has been included, although initial test results suggest that ultimately this may not be required. The design criteria used are summarised in Table 18.44. Table 18.44 Summary of Major Design Criteria Description Value Nominal annual dry ore throughput Unit 3 000 000 t/a Kamoeb Ore Grade Gold Nominal 5.00 g/t Silver Nominal 4.00 g/t Heap Leach Residue Grade Gold Nominal 1.50 g/t Silver Nominal 1.20 g/t Plant Feed Grade Gold Nominal 2.67 g/t Silver Nominal 2.13 g/t Overall Recovery on New Feed Gold Design 91.0 % Silver Design 62.4 % Gold Nominal 233 803 oz/a Silver Nominal 128 397 oz/a Throughput Nominal 1 000 000 t/a Availability Design 70.0 % Minimum Throughput Design 163 t/h Nominal 2 000 000 t/a Metal Production Crushing (Kamoeb) Heap Leach Reclaim Throughput Grinding Circuit type SABC Availability Design 91.3 % SAG Feed Rate (fresh) Ball Mill Feed Rate Design 125 t/h Design 375.0 t/h Note: − The overall recovery is based on a leach recovery of 90%, in addition to gold recovered through gravity concentration. A simplified flow sheet is provided in Figure 16.3, and a brief description of the plant is as follows. 18.8.1.1 Heap Leach Tailings Reclaim The existing heap leach residue will be reclaimed by bulldozer or front end loader into a mobile feeder system. This system in turn transfers the reclaimed material to an overland conveyor, which has been assumed to be 2500 m in length. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 218 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report The overland conveyor feeds a storage bin located adjacent to the milling area. The bin then discharges the heap leach tailings feed at a measured rate via a variable speed controller into the feed of the overflow ball mill. Heap leach residue will be reclaimed at a rate of 2.0 Mt/a. 18.8.1.2 Crushing A new crushing circuit is proposed, consisting of a new coarse ore bin, feeder, jaw crusher, conveyor and crushed ore bin. The current circuit will be made redundant and the scoping study has not allowed for any of this equipment to be re-used. This new circuit is capable of treating both Kamoeb South ore and acidic SBR ore types. The crushing circuit has been sized to process 1.0 Mt/a of ore. 18.8.1.3 Grinding The proposed grinding circuit consists of a single SAG mill operating in open circuit, followed by a single overflow ball mill, operating in closed circuit with a set of hydrocyclones. The SAG mill operates with a scats crusher, though this requirement will be dependent upon a data review during the feasibility study stage. The target cyclone overflow sizing is P80 75 μm. The SAG mill processes only fresh ore at a rate of 1.0 Mt/a: the proposed mill is sized at 1.5 MW. The overflow ball mill processes both the discharging slurry from the SAG mill plus reclaimed heap leach tailings at a rate of 2.0 Mt/a. The ball mill will be 5.0 MW in size. Comminution requirements are based on previous testwork. circuit will be required during the feasibility study stage. 18.8.1.4 Further optimisation of the proposed Gravity A bleed stream of cyclone underflow is gravity fed over a DSM screen to scalp material at approximately 5 mm. The screen oversize reports to the mill discharge hopper, while the screen undersize reports to a single centrifugal concentrator. Concentrator tails report to the mill discharge hopper, while the concentrate reports to the intensive cyanidation unit for further processing. The centrifugal concentrator will produce, on average, one tonne of concentrate per day. The intensive cyanidation unit will treat this concentrate batchwise with a leaching time of 22 hours. 18.8.1.5 Leaching and Adsorption Cyclone overflow from the grinding circuit is gravity fed through a trash screen into the first leach tank. The leach circuit consists of two leach tanks with a combined residence time of approximately six hours. Oxygen or air injection into the leach tanks may be required, depending on the results of the testwork program. Leach feed density has been assumed to be 44.3% w/w solids. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 219 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report The adsorption circuit consists of six tanks with a total residence time of 18 hours. Loaded carbon from the first tank is pumped to the loaded carbon screen for rinsing, prior to entering the elution column. Carbon is pumped counter-current to the slurry flow by recessed impellor pumps. Each adsorption tank has a single cylindrical intertank screen with an 800 μm aperture to retain carbon. Average tank sizing is 1881 m³. 18.8.1.6 Elution/Goldroom Loaded carbon from the first adsorption tank is pumped via the loaded carbon screen to the elution column. The column is designed for a combined acid wash and elution duty. Carbon is acid washed with 3% hydrochloric acid, prior to rinsing and elution using the split-AARL method. Barren carbon is transferred to the regeneration kiln mounted on top of the adsorption tanks, prior to returning to the adsorption circuit. Pregnant liquor from the elution circuit is electrowon onto stainless steel cathodes in two electrowinning cells. The cathodes are removed and manually stripped with high pressure water as required (nominally once per week). Sludge from the electrowinning cells is dried prior to smelting to produce doré. The elution batch sizes have been determined such that one elution cycle per day is required. The batch size is 11.0 t at assumed carbon loadings of 2000 g/t Au and 1250 g/t Ag. The elution time has been estimated at 9.6 hours. 18.8.1.7 Cyanide Detoxification A cyanide detoxification (detox) circuit has been included within the revised circuit to reduce the tailings cyanide solution content to low levels. This is based upon the SO2/Air process, utilising sodium metabisulphite (SMBS) solution, copper sulphate and hydrated lime as reagents. Total residence time is 90 minutes. As this process has yet to be tested with the Hassaï ore, the values used have been assumed based upon Sedgman experience, and testing is required to verify the circuit requirements during the feasibility study stage. Tailings from the adsorption circuit gravitate over the carbon safety screen, to capture any misreporting carbon, with the screen undersize gravitating into the first detox tank. Air is injected into the tank along with lime slurry, SMBS and copper sulphate. Slurry overflows the first tank into the second detox tank, where further lime slurry is added to maintain pH along with further air injection. The second detox tank overflows into the CIL tails sump. Individual tank live volume is 556 m³. 18.8.1.8 Tailings Tailings slurry from the cyanide detox circuit is pumped to a high rate thickener for water capture. Thickener overflow is gravity fed to the process water dam. Target thickener underflow density has been assumed to be 60% solids and is pumped to the TSF located adjacent to the processing plant. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 220 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Thickener specific values have been assumed in this study at 0.80 t/m².h and need to be verified by laboratory testing during any subsequent feasibility study. The current TSF design has assumed a square paddock style facility, which is unlined. Tailings are deposited into the facility by a perimeter spigoting system with beaching of the solids occurring adjacent to the dam wall and a decant pond forming in the centre. A rise rate of 2.0 m/a has been assumed in the current design, however this needs to be confirmed by testing. Decant water from the TSF is recovered by a pump located in a central decant tower and returned to the process water dam for reuse in the process plant. The tailings thickener diameter is 25.0 m. The tailings dam dimensions have been assumed to be 1035 x 1035 m. No specific site has been nominated (see Section 18.9.5.1). 18.8.1.9 Reagents Cyanide, caustic soda, SMBS, hydrated lime and copper sulphate are mixed manually as required in dedicated facilities. Bulk storage facilities for quicklime and hydrochloric acid are installed, along with an automated flocculant mixing facility for the tailings thickener. Caustic soda, SMBS, hydrochloric acid and copper sulphate are added to the circuit via dedicated dosing pumps. Cyanide and hydrated lime slurry are circulated through a ring main, with injection into the process being achieved by automated timed dosage. Quicklime is added as a dry powder to both the SAG and ball mill feed belts. Equipment sizing has been matched to the processing options and assumptions made. 18.8.2 VMS Concentrator Process Plant Description Equipment availability assumptions for both circuits are 91.3% for all plant areas, with the exception of the filtration and concentrate handling area which operates at an availability of 80%. 18.8.2.1 Ore Delivery and ROM Pad Ore is delivered by mine haul truck to a ROM ore pad. The ROM pad storage is sized to ensure anticipated mine stoppages do not restrict plant feed. 18.8.2.2 Crushing The crushing plant is sized to operate 24 hours/day, 7 days/week at an instantaneous throughput of 625 t/h. Table 18.45 shows the equipment sizing for the crushing area. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 221 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.45 Crushing Equipment Stage ROM bin 200 t Vibrating grizzly feeder Jaw crusher 1.6 m W x 7.3 m L Metso C140, 200 kW Ore is loaded into the ROM bin by FEL. The ROM bin is sized to hold approximately 20 minutes of feed at the average throughput rate. Ore is withdrawn from the ROM bin by vibrating grizzly feeder. Oversize discharges into the primary jaw crusher operating at 100 mm closed side setting (CSS). The undersize from the vibrating grizzly and the primary crusher discharges onto the same conveyor belt. The primary crusher conveyor discharges to the mill feed conveyor. A weightometer and tramp magnet are mounted over the head pulley. 18.8.2.3 Grinding and Classification The grinding circuit consists of an open circuit SAG mill, followed by a ball mill in closed circuit with two clusters of 400 mm hydrocyclones. The SAG mill is 6.1 m diameter, with an effective grinding length of 8.8 m and is powered by a single 5.2 MW motor. The mill is designed for overflow discharge via a trammel, with the undersize reporting to the mill discharge hopper. The SAG mill trommel oversize falls into a bunker for removal by loader or bobcat. The SAG mill motor is selected with a hyper-synchronous SER drive. Grinding media for the SAG mill is added as required onto the SAG mill feed conveyor. Ball loading is achieved via a hopper and feeder system located above the SAG mill. For the ball mill, the grinding media is added via a kibble. Duty and standby cyclone feed pumps draw from the ball mill discharge hopper and pump to two clusters of 400 mm diameter hydrocyclones (13 operating and two standby at each cluster). The target P80 of the cyclone overflow is 70 µm. Cyclone overflow is moved by gravity to a static trash screen prior to reporting to the rougher flotation circuit. The cyclone underflow stream is returned to the ball mill. The ball mill is 7.3 m diameter inside shell, with an EGL of 10.2 m. The mill is powered by twin 4.5 MW motors, for a total power of 9.0 MW. The ball mill discharge flows through a trommel. Undersize from the trommel cascades into the common mill discharge hopper. A four tonne crane is provided to assist in cyclone cluster maintenance activities. Other maintenance tasks in the mill area require the use of a mobile crane. A relining machine and mill platform access ramp are provided to allow for change-outs of liners and lifters. The SAG and ball mills each have a jacking cradle system and inching drive for maintenance purposes. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 222 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.8.2.4 Rougher Flotation and Regrind Rougher flotation is nominally carried out at the cyclone overflow pulp density of 35% solids and an optimum grind size P80 of 69 µm. The rougher circuit is in open circuit, with rougher tailings reporting to the tailings thickener. After flowing through the trash screen, the cyclone overflow from the grinding circuit gravitates to the rougher flotation cells. The rougher stage of flotation consists of two trains of 6 x 100 m3 forced air tank cells. Cells are installed as three pairs of twinned cells, ensuring sufficient stages to reduce pulp short circuiting, while minimising the static head requirement across the rougher circuit to ensure gravity flow to the rougher tails hopper. The installed residence time for the rougher flotation cells is 40 minutes, based on the scaled-up results from the laboratory tests. Flotation is undertaken at elevated pH of 10.5 to aid in the depression of pyrite. Aerofloat 238 is added as collector and methyl iso-butyl carbinol (MIBC) as frother. Rougher concentrate gravitates through launders to a concentrate hopper, and is pumped to the regrind circuit. Flotation tailings gravitate from the final cell to the rougher tailings hopper and is pumped to the tails thickener feed tank. The regrind circuit consists of two ISAmill M1000 units with 500 kW motors operating in parallel in open circuit. The regrind circuit is designed to produce a regrind target P80 of 30 µm. The feed to the regrind mill is deslimed in advance of the mill to remove fines with a 150 mm hydrocyclone cluster. The cluster contains 13 operating and two spare cyclones. The desliming cyclone underflow reports to the mill feed hopper, and the overflow gravitates to the regrind mill discharge hopper. The regrind media feeder transports grinding media from the media hopper to the mill feed hopper. The media combines with the mineral slurry and is transported via feed pump into the grinding chamber of the mill. Media is slowly consumed during the milling process and fresh media is added to retain the required charge. The media is added in a controlled manner to maintain constant power draw for each mill at the set-point specific for that mill. In this way, the size distribution of the product from the regrind circuit is also controlled. The regrind rougher concentrate from the mill gravitates to the regrind mill discharge hopper, and is pumped to the cleaner circuit. 18.8.2.5 Cleaner Flotation Table 18.46 shows the equipment sizing for the cleaner flotation area. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 223 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.46 Cleaner Flotation Equipment Stage Option B First cleaner (cleaner 1) 6 cells of 20 m3 Cleaner scavenger 3 cells of 20 m 3 Second cleaner (cleaner 2) 4 cells of 10 m 3 The discharge slurry from the regrind circuit is treated in three stages of closed-circuit cleaning. Regrind rougher concentrate is combined with the second cleaner tailings and cleaner scavenger concentrate in the first cleaner cell feed box to form first cleaner feed. First cleaner flotation is carried out in six tank cells with a nominal residence time of 20 minutes total. The first cleaner concentrate is pumped to the second cleaner feed where final concentrate is produced. The first cleaner tailings gravitate to the cleaner scavenger bank, from which the nonfloating component is transferred to final tails. The second cleaner stage has a nominal residence time of 15 minutes and the cleaner scavenger stage 15 minutes. The second cleaner consists of four tank flotation cells, while the cleaner scavenger bank consists of three tank flotation cells. Final cleaner concentrate is stored in an agitated tank in the cleaner flotation area to allow residence time for concentrate de-aeration, prior to pumping to the concentrate thickener. 18.8.2.6 Concentrate Thickening, Filtration and Handling Table 18.47 shows the equipment sizing for the concentrate handling area. Table 18.47 Concentrate Handling Equipment Major Equipment Concentrate trash screen Concentrate thickener Concentrate filters 1.2 m W x 3.6 m L 22 m diameter high rate 2 x pressure filters of 50 m 2 Final cleaner concentrate is pumped from the flotation area to the concentrate area. Pump discharge passes over the concentrate trash screen to protect the thickener and downstream filter. Underflow from the trash screen flows into the concentrate thickener. Thickener overflow is pumped to the process water tank. Thickener underflow is removed by pumps at 65% w/w solids and pumped to the concentrate storage tanks. Thickener underflow can also be recycled back to the thickener to ensure that underflow density can be maintained during times of low concentrate production. Two concentrate storage tanks are provided with a live capacity of 1000 m3 each, allowing a total storage capacity of 48 h. Filter feed pumps feed two pressure filters. Dry cake (10% moisture) is dumped from the bottom of each filter to a set conveyor belt that discharges into a storage bunker. The filtrate gravitates to an air/water separator in which the filtrate is de-aerated prior to being pumped back to the concentrate thickener. The filter building has a concrete floor and push walls to suit operation of FELs. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 224 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report A FEL is used to transfer concentrate from the intermediate stockpile to the concentrate packaging system which packs the concentrate in two tonne bulk bags. Bulk bag delivery is included due to bulk handling issues at Port Sudan. Bulk bags of concentrate are loaded by forklift onto trucks for transport to Port Sudan. The average concentrate productions are presented in Table 18.48 for each ore type. Table 18.48 Concentrate Production Ore Type t/d at 10% Moisture Hassai South Supergene 1065.0 Hassai South Primary 745.6 Hadal Awatib 513.0 The site and port storage capacity is included at 15 and 30 days respectively, for the average concentrate production rates tabulated above. 18.8.2.7 Reagents Nominal on-site reagent storage is included with a stock capacity of 30 days to ensure stocks are sufficient to allow continued operation during periods of road outages or supply delays. The use of a separate reagent shed and mixing/storage area minimises unnecessary personnel interactions with this potentially hazardous area. Hydrated Lime Hydrated lime (Ca(OH2), is delivered to site in bulka bags, and is loaded into the feed hopper using a hoist. Solids are transferred by screw feeder for continuous mixing in a mixing tank. Duty/standby pumps circulate milk of lime through the lime ring main. Lime dosing at each individual distribution point is controlled by on/off pinch valve using time-based control loops. Milk of lime is distributed to the following locations: • Mill discharge hopper • Regrind surge tank • Cleaner scavenger cell 1 • Cleaner 1 concentrate hopper. Flocculant Powdered flocculant, Magnafloc 1011 or equivalent, is delivered to site in bulka bags, and is loaded into the feed hopper using a hoist, for batch mixing in the vendor-supplied flocculant mixing package. The resulting 0.25% w/w solution is transferred to a storage tank with 24 hour residence time for aging. Flocculant is metered to each of the concentrate and tailings thickeners using duty/standby variable speed positive displacement pumps. Metering is achieved by calibrating pump speed to flocculant flow rate. Flocculant is diluted to approximately 0.025% using process water just prior to the addition point. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 225 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Collector Collector (Aerofloat 238 or equivalent), is delivered in liquid form in 210 L drums. A drum pump is used to transfer collector from the drum to the storage tank. Collector is pumped to the following dosage points using dedicated diaphragm metering pumps for each point: • Mill discharge hopper • Rougher cells 1 and 7 • Rougher cells 5 and 11 • Regrind surge tank • Cleaner 1 tailings hopper. Frother MIBC is delivered in liquid form in 210 L drums. A drum pump is used to transfer this frother from the drum to the storage tank, from where it is pumped to the dosage points using dedicated diaphragm metering pumps for each point. Delivery of frother is to the following locations: • Rougher cells 1 and 7 • Cleaner 1 cell 1 • Cleaner scavenger cell 1 • Cleaner 2 cell 1. 18.8.2.8 Air Systems Plant and Instrument Air Plant and instrument air are provided from the same compressor and dryer system. Four sequencecontrolled compressors supply a plant air receiver at 1000 kPa. The plant air is distributed directly from the receiver. Instrument air is treated through dryers and filters to remove moisture and particulates. A pressure reducing valve reduces plant and instrument air pressure to nominally 700 kPa. Instrument air is maintained during pressure fluctuations by using a pressure sustaining valve on the plant air distribution line. Blower Air Four blowers (including one standby) provide high volume, low pressure air for flotation. Automatic flow controllers at each cell are used to control air addition. 18.8.2.9 Water Raw Water System Raw water consumption projected for the combined CIL and concentrator operation is shown in Table 18.49. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 226 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.49 Demand Design Raw Water Consumption (no return from TSF) Value (m³/d) 3.0 Mt/a CIL plant 5.0 Mt/a VMS Flotation Plant 800 Man Camp 5 544 11 261 160 Dust Control Allowance 1 500 Community Allowance 1 000 Sub-total Contingency (9%) Total 19 465 1 675 21 140 Existing sources are totally inadequate to meet this demand, and raw water will be supplied from the Nile River by overland pipeline to a lined raw water pond. Raw water is used as make-up water for the process water distribution system, and is also distributed to the following points: • Gland water system • Concentrate trash screen • Lime mixing • Flocculant mixing. Process water is stored in the process water dam, where recovered water from the process and fresh raw water are collected and distributed to the plant. The process water pond has a capacity of 24 hours at nominal capacity. The process water pumps operate in a duty/standby arrangement. Process water distribution is by pressurised header, and is distributed to the following locations: • Crushing dust suppression system • SAG/ball mill and mill discharge hopper • Flotation concentrate water sprays • Dilution of flocculant at the concentrate and tailings thickeners • Fire water system. Fire Water System From the fire water dam, a system of pumps supply water to the fire water header and to the hydrants distributed through the plant. The fire water storage satisfies a 4 hour fire-only storage requirement. Gland Water Filtered raw water is used for gland water. Antiscalant addition may be required. There are two gland water services, the gland water being split into low pressure and high pressure systems. Both systems operate from the same feed tank using dedicated duty/standby pumps for each system. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 227 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report High pressure gland water is delivered to: • Mill cyclone feed pumps • Filter feed pumps • Tailings discharge pumps • Process water pumps. Low pressure gland water is delivered to all other pumps requiring gland water, and also supplies the filter requirements in the concentrate area (cloth wash water tank and filter pressing water tank). Potable Water Potable water is supplied from the CIL plant. Potable water for eye wash and safety shower distribution is via duty/standby pumps. The shower water system is a pressurised header. Due to high ambient temperatures, the shower water system includes a cooling tower to maintain acceptable water temperatures. 18.8.2.10 Concentrator Tailings Management The concentrator produces approximately 4.7 Mt/a of tailings material requiring disposal on site. A review of disposal options was completed, taking account of the fact that water conservation and recovery is important, both from an environmental and financial perspective. From this it was concluded that a slurry disposal method would be adopted for the scoping study, with discharge of the total tailings stream at the plant thickener underflow density (55% w/w), and delivery via a centrifugal pumping system to the TSF for sub-aerial deposition. Details regarding the option study, TSF design and operation are included in Section 18.9. 18.9 PROJECT INFRASTRUCTURE AND SERVICES 18.9.1 Water Supply Water supplies for the current Hassaï operation are reliant upon bores and surface storage dams, and are limited. The supply of raw water for the expanded project is expected to be sourced via a new pipeline from the Nile River, approximately 165 km away. This pipeline is sized to meet the requirements of both the VMS concentrator and the CIL plant. It is estimated that six pump-stations will be required to transport water to site. Although the CIL plant is expected to be developed and commissioned prior to the VMS plant, the full capacity requirement for all site operations is to be constructed during the CIL plant construction phase. The pipeline is constructed primarily from mPVC (extruded on site) with steel sections used in some areas as dictated by the ground conditions. Table 18.50 summarises the preliminary pipeline design. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 228 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.50 Water Pipeline Summary Description Unit Pumping Stages Pipe Size, OD System Op Pressure (max) Pipe Rating Total Static Head Value - 6 mm 500 m 140 bar PN 9 - 20 m 150 Hydraulic Power / Stage kW 322 Installed Power / Stage kW 600 m³/hr 961 km 165 t 6253 Nominal Flowrate Total New Pipe The pipeline has been sized to provide 21 140 m³/d of water (nominally 961 m³/h operating 22 h/d). This flowrate will be sufficient to supply the projected demands of a 3.0 Mt/a CIL plant plus the proposed future 5.0 Mt/a VMS flotation plant (Table 18.49). Line sizing has reached the limits of commercially available mPVC pipe, while it has been estimated that six pumping stages will be required. At the plant site, water will be stored a new, lined raw water dam, which overflows into a new, lined process water dam. The capital cost estimate has been developed by Sedgman, while AMEC has included capital costs for additional water reticulation infrastructure required for the VMS concentrator. 18.9.2 Power Supply The current operating site maintains 17 generators to supply the operating power requirements. It is anticipated that these units will be converted to meet emergency power requirements for the proposed Hassai CIL plant. This capability could function as emergency power supply for both the CIL and VMS concentrator in the future. Power for the CIL and VMS concentrator projects is to be sourced via an overland line from the existing national power grid supply. The project is approximately 77 km from the nearest potential grid connection. As the CIL plant is anticipated to precede the VMS, the required overland transmission lines and associated power infrastructure for both plants will be installed as part of CIL project. Power required on site for the CIL and Concentrator components is shown in Table 18.51. Table 18.51 Plant Power Requirements (MW) Total installed power (excluding standby equipment) Utilised power Emergency power CIL Concentrator 10.3 24.0 19.5 2 x 2 MW generator sets FINAL – Rev 0 – 22 Oct 2010 AMEC Page 229 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Preliminary conversations with NEC, the national electrical supply company, have confirmed that there is no shortage of power available from the Atbara - Port Sudan line. Line construction would be undertaken by NEC, financed by AMC, and construction is estimated to take a minimum of 1 year. The power cost is estimated to be $0.074/kWh, based on pricing supplied by AMC. 18.9.3 Accommodation The existing Hassai camp is located approximately 3 km from the current heap leach operation. This camp currently accommodates 600 personnel (expatriates and locals) and includes accommodation, mess hall, bakery, local market and recreational facilities. On-site communication allows mobile phone and internet access. A new accommodation village will be developed for the CIL and VMS plants, to accommodate workforce requirements. A 200 person camp has been allowed for in the capital cost estimate by Sedgman for the CIL operation, while accommodation for a further 500 persons is included by AMEC in the VMS concentrator estimate. 18.9.4 Airstrip A fully maintained airstrip is located at the site and is used for the heap leach operation. This airstrip is utilised currently for air freight to site and transport of personnel to and from Khartoum. AMC maintains a Twin Otter aircraft which is based on site. 18.9.5 VMS Concentrator Tailings Storage Facility 18.9.5.1 Site Selection During the site visit by AMEC Minproc and AMEC E&E personnel to site in early March 2010, a proposed site for the location of the TSF was identified. The site is shown in Figure 18.22, together with the proposed location of the VMS plant and the northern part of the existing leach pads. The site would appear to be suitable based on the availability of land, the absence of communities and the absence of obvious sensitive environments in the area. The topography of the area is very flat. A preliminary site assessment identified abundant potential embankment construction materials, including material from old waste rock dumps. However, geotechnical, hydrological, social and environmental issues will require further investigation as part of any feasibility studies. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 230 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 18.22 Tailings Storage Facility Site Location Option The results of the preliminary review indicate that the key aspects to be considered for the selection of the TSF site include the following: • Key environmental aspects: protection of the Wadi Amur from potential contamination • Key technical requirements: the approximate TSF storage volume is 38.5 Mm3, which provides storage for 50 Mt, or 10 years of production, assuming an average density of 1.3 t/m3. 18.9.5.2 Preliminary Environmental Assessment To consider the relative environmental effects of the site and to establish potential mitigation requirements, sites were assessed using information pertaining to socio-economic and political factors as summarised in Table 18.52. Table 18.52 Environmental Assessment Scoring Criteria Score Level of Objection Resources and Risk 1 None or Minor objection expected Minor resources required to counter objection 2 Objection expected Major resources required to counter objection and provide mitigation 3 Significant objection expected Risk of onerous mitigation measures, requires mitigation to be measures incorporated into technical design FINAL – Rev 0 – 22 Oct 2010 AMEC Page 231 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report The approach used required consideration of the following factors: • The quality, detail and extent of available data • AMEC's experience of similar tailings disposal impacts and effects • A judgement on the relative significance of the impacts in the context of the social and environmental setting. The assessment establishes a qualitative score for each site in respect of the environmental factors considered. It is recognised that individual factors have a different level of significance and the scores reflect the likelihood of objections to the development based on each. The resources necessary to counter each objection and the extent of the appropriate mitigation measures are reflected in the scores. Therefore, a score of 1 is given to perceived minor objections to reflect minimal environmental effect, whereas a score of 2 or 3 is used to highlight differences between a minimal effect and situations of unfavourable environmental responses. The criteria to measure the selected site (and any other potential site identified in future) include the following: Village Impact The site identified scored 1, the lowest possible score, as there are no villages in the vicinity of the mine site of the proposed TSF site. Catchment The surface water catchment for the site is believed to be minimal and, therefore, according to available mapping there should be no significant implications for this site. It is recognised, however, that the proposed site could be in the upstream side of the Wadi Amur and, therefore, mitigation measures to minimise the risk of contamination will have to be assessed and implemented. Ownership/Land Use It was observed during the site visit that the area of the proposed TSF does not have, and due to its characteristics is unlikely to ever have, a land use. No local communities were observed in the vicinity of the project site, and, on this basis, the site is considered appropriate. Visual Impact AMEC believes that the site is appropriate for this type of development. Potential Impact on Water Supply Two aspects to the potential impact on water supplies need to be considered: • Groundwater vulnerability • Cost of water piping from the Nile and the impact of that installation/operation. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 232 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Based on the above observations and subject to confirmation of the assumptions, it is considered that the site north of the proposed VMS plant site is appropriate for the location of the TSF. 18.9.5.3 Tailings Delivery Options Water conservation and recycling has been identified as a key parameter in the operation of the tailings facility, and, therefore, the three options with contrasting philosophies for water management were analysed in some detail as follows: • Filter cake disposal: the tailings stream is passed through additional thickeners and/or a pressure belt filter system to dewater the total product to the optimum moisture content. Filter cake is delivered to the deposition area for discharge by either articulated truck or a dedicated overland conveyor system. The site filter cake disposal scheme would include sequential raising of the TSF while controlling surface run-off and silt discharge. Pre-deposition works will include clearance of vegetation, topsoil stripping and stockpiling (if any), the formation of seepage collection drains, silt collection and drainage collection ponds. Stormwater will be diverted around the TSF by the open drain formed adjacent to the periphery, and may report to a small water retention dam. Upon reaching the final design height, the surface of the facility will be re-profiled to a peripheral drainage system and a series of silt traps installed. The limited volume of supernatant released from the tailings will report with normal surface run-off to the supernatant pond where it will be returned to the process plant. This pond will be significantly smaller than envisaged for a slurry disposal scheme. • Paste/thickened tailings disposal: the tailings stream is discharged via additional thickeners to reduce the pulp density to at least 70% solids by weight. Tailings are then pumped to the depository by positive displacement pumps where discharge is undertaken by either co-disposal with stripped overburden or rock waste from the mining operations, or discharged from a series of open end discharge points to form a natural cone-shaped depository. The depository will be surrounded by a low perimeter wall to retain any supernatant water released from the conical pile and to control storm water run-off. Resultant supernatant water will reclaimed for return to the process plant. Due to the availability of potential sites in close proximity to the proposed VMS plant site, and the requirement to recover water, a paste transport system is considered potentially beneficial for the project. • Slurry disposal: this method considers the discharge of the total tailings stream at the plant thickener underflow density (55% w/w), and delivery via a centrifugal pumping system to the TSF for sub-aerial deposition. Due to the high water content, the variability of the particle size distribution of the slurry, specific gravity and the transporting fluid viscosity, the solids undergo hydraulic separation upon sub-aerial discharge. The coarse fraction is deposited adjacent to the discharge point while the finer fraction is held in suspension and carried forward to the supernatant pond. To enhance the separation of interstitial water, tailings are discharged via spigots, cyclones or open pipes, which are continually relocated along the TSF periphery to form a layered free-draining beach. Catchment run-off and supernatant water released by the tailings will report to the supernatant pond and the floating decant equipped with submersible pumps. Generally rainfall and run-off water will be retained within the TSF without release to the environment. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 233 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report The site identified near the proposed VMS plant site would be suitable for the construction of a tailings facility using this transportation method. However, the water losses (entrained water, seepage and evaporation) will be high, and a trade-off study is needed to confirm the costs of the overall system. The alternatives were assessed in terms of: • • Environmental impacts, primarily: − Uncontrolled escape of effluent and solids by seepage and run-off − Wind erosion of the tailings solids from the surface Capital and operating costs. Table 18.53 shows the results of the disposal system ranking study. In general, conventional slurry disposal is favoured in terms of costs, whereas the filtered and paste/thickened systems have potential environmental benefits. For the purposes of the scoping study, it has been assumed that the tailings produced will be in slurry form with approximately 55% density (w/w). The tailings discharge system will include equipment and costs for the tailings thickener discharge to the TSF. Table 18.53 Disposal System Ranking Supernatant pond Filter Cake Paste /Thickened Slurry 1 2 3 Topography Location with respect to plant site 1 1 1 Seepage 1 2 3 Surface erosion/stability 3 2 1 Wind erosion 3 1 2 Ease of future expansion 1 2 3 Geotechnical 1 1 1 Operational supervision 2 2 1 Storm Water Management Capital expenditure 2 2 1 Operational expenditure 2 2 1 17 17 17 Total 1 - is most advantageous, 3 - is least advantageous 18.9.5.4 TSF Design The upstream method with day walling for dam raises is a method widely used in Africa and will be applied to the project to take advantage of its relatively low capital and ongoing costs. Figure 18.23 shows a schematic section of the construction. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 234 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 18.23 Schematic of the Upstream Construction Method Tailings Dam Area and Rate of Rise The rate of rise of a tailings dam constructed using the upstream method is paramount in the definition of the stability of the dam. As the upstream section of the embankment is founded on tailings previously deposited, the consolidation and drainage of the tailings should be sufficient to allow the construction of the following raises without compromising dam stability. General guidelines indicate that the maximum rate of rise for a tailings dam being constructed using the upstream method should not exceed 2-3 m at any time during the life of the facility. Given the particular circumstances of the site, it is considered appropriate to use a maximum rate of rise of 3 m per year, as the lack of rainfall and the high temperatures prevalent on site will promote desiccation and consolidation of the tailings mass. The layout of the tailings facility will be mandated by the required tailings disposal capacity for the ultimate life of mine. The preliminary process schedule calls for disposal of approximately 30 Mt/a, equivalent to a total deposition capacity requirement of approximately 23 Mm3. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 235 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Based on the above, the footprint of the TSF, up to the downstream toe of the starter embankment, will measure approximately 160 ha, with a perimeter of approximately 5000 m. The resultant TSF will reach a total height of some 17 m after 10 years of deposition. Tailings Storage Facility Components The design will need to cover the following elements as a minimum: • Pre-deposition earthworks, including site preparation and the construction of a starter embankment some 5 m high for the deposition of 2 years production • Storm-water diversion channels • Seepage collector drains and shallow seepage cut-off drains • Return water dam complex, made up of a lined return water dam and an unlined storm water dam • Floating barge and pump on the pool delivering supernatant water to the return water dams • Tailings delivery and return water pump station platform at the return water dam complex • Tailings delivery ring-feed • Topsoil stockpile (should this be required) • Access roads. No HDPE liner has been considered for the tailings facility. This assumption will need to be confirmed as part of any prefeasibility study. 18.9.5.5 Closure General A “best practice” closure plan will be developed during detailed design, based on guidelines similar to those prepared by the Ontario Ministry of Northern Development and Mines (1995) or the Minerals Industry Research Organisation of UK (MIRO, 1999). The plan will incorporate a long-term objective for closure and rehabilitation, which will permit the mine operator to leave the site in a condition that requires limited ongoing maintenance. The tailings surface will be drained, graded, sheeted with stripped overburden, and planted with natural vegetation as appropriate. An open-channel spillway will be excavated to the south of the facility to allow for the free discharge of uncontaminated surface waters. All tailings delivery and water reclaim pipes will be salvaged and their associated drains and access roads rehabilitated. That portion of the downstream slope of the TSF embankment not yet re-vegetated at closure will be treated accordingly. The embankment toe drains, sumps and monitoring boreholes will be kept open and monitored on a regular basis until such a time that any seepage is proved not to be detrimental to the environment. Prior to closure, the TSF will comprise an upstream embankment confining some 30 Mt of process waste, sufficiently drained so that a stable upper surface will have been established. Closure activities will commence during the final year of operation, to ensure that all the objectives can be achieved efficiently and cost effectively. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 236 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report In accordance with international best practice for TSF sites, data will be collated throughout the deposition period to ensure that an appropriate closure strategy is adopted. This information will include data on tailings geotechnical and geochemical properties, as well as on vegetation types, hydrology, meteorology, etc., which will in turn be incorporated into the closure planning documents. To achieve these objectives, the closure plan will define a number of key periods of activity during the closure of the TSF, which are envisaged as: • • Pre-abandonment period: during the last year of operating the process plant, preparatory works will include the modification of the tailings disposal system to ensure achievement of the final surface topography toward a future storm water run-off channel while also maintaining the final consolidation of the upper surface of the tailings. This will entail the following: − Control of deposition to create the tailings pond in its final location − Stockpiling barren (stripped waste) material to enable the construction of the tailings dry cover − Possible inclusion of nutrients and seed stock into the final tailings deposition layers as determined from vegetation trials Post-abandonment: during the post-abandonment period, input will be required to achieve the final surface topography commensurate with the agreed after-use and to ensure its long-term integrity. 18.9.6 Other Infrastructure The current heap leach operation is supported by offices, workshops, storage facilities, a laboratory, fuel, lubricant and explosives storage. In general these are considered to be adequate to support the CIL operation, but additional infrastucture will be required for the VMS concentrator and underground mining operations. No detailed engineering has been undertaken, but allowances have been made for these additional items in the capital cost estimate. 18.10 MARKETS AMC has not completed a review of the market for copper concentrates, nor initiated any discussions with potential customers. It is believed that: • Smelting capacity has continued to grow, exceeding the increase in global concentrate output. No significant growth in mine production is foreseen in the medium term. • The main growth in smelting capacity has been in China; concentrate imports rose 18% during 2009. Utilisation rates at smelters are at historical lows. • The position of smelters has eased due to higher sulphuric acid prices, better free metal credits and copper cathodes premiums. However, they still face an adverse market. • Recent annual benchmark treatment and refining charges (TC/RCs) reflect the tight market, as follows: − 2009 calendar year: TC $75/t and RC $0.75/lb − 2009 mid-year: TC $50/t and RC $0.50/lb − 2010 calendar: TC $46.5/t and RC $0.465/lb. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 237 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report • Price participation has disappeared since 2009. • China drives the spot market, supported by internal measures allowing smelters to compete aggressively in the spot market. • Lack of spot availability has created aggressive competition not only among merchants but also among smelters. Based on the above, it is believed that the positive outlook for copper concentrates will persist, with low TC/RCs expected in the coming years as the supply deficit persists and low smelting capacity rate utilisation remains. Consequently, there should be a demand for AMC’s copper concentrates, assuming high quality concentrates can be achieved. The underlying assumption for financial analysis is that TC/RCs will average $50/t concentrate and $0.50/lb copper over the life of the project, which is in line with industry forecasts. It is important to note that actual charges will depend on the quality of concentrate being supplied including factors such as concentrate grade and impurity levels. Although zinc can be recovered during the flotation process, no testwork has been completed to optimise its recovery and produce separate concentrates for sale. Consequently, zinc sales and zinc treatment costs have been excluded from the scoping study. 18.11 ENVIRONMENTAL AND SOCIAL CONSIDERATIONS A preliminary environmental review was undertaken by AMEC Earth and Environmental, including a site visit to gather relevant information. The review was based on the current legislative context in Sudan and focussed on identifying fatal flaws and opportunities. The review included the following specific tasks: • Review all available and appropriate information • Assess current compliance with local, regional, national and international regulations and standards • Examine potential remediation and reclamation issues • Identify potential issues that might arise with project expansion, including the need for an SEIA and other environmental requirements • Identify potential significant environmental impacts that might arise from the expansion project • Time requirements • Provide input into ongoing data collection activities to facilitate permitting and compliance processes. A number of areas were identified which will require attention and recommendations for remedial actions and mitigations are provided along with information setting out roles and responsibilities and time requirements in order to implement an appropriate system. Potential fatal flaws included: • No environmental management system (EMS), which consequently leads to lack in planning, mitigation and compliance • Existing SHE auditing and compliance is not effective in supporting the project to address environmental issues FINAL – Rev 0 – 22 Oct 2010 AMEC Page 238 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report • Establishing where the original baseline reports and data are, and ensuring documentation is controlled • No Environmental Emergency Plan is evident. The abovementioned actions need to be undertaken in order to meet the International Finance Corporation guidelines, but there is no evidence for an environmental fatal flaw that would affect either the CIL or VMS projects. Opportunities include: • Excellent relations with national and local representatives, provision of services and resources • Water – management, measurement, quality were all taken seriously and a management plan exists • Waste – signs of best practices were evident, but improvements are needed in overall management and planning, including for mine waste. 18.12 TAXES AND ROYALTIES AMC’s current gold production is subject to a Net Smelter’s return (“NSR”) of 7% to be paid to the state of Sudan. This level of royalty was acknowledged by the Sudanese government to be above industry average and was reviewed downwards in more recently awarded exploration permits. La Mancha’s most recent permit in Sudan, the Nuba Mountain project, has a 5% NSR on gold and a 3.5% NSR on base metals attached to it. Furthermore, AMC’s net earnings have been taxed at 15% rate by the Sudanese government since 2008. Prior to that, AMC was exempt from paying any tax on its net earnings, an incentive that had been specially granted by the state to encourage the start-up of the mining industry in the country. In the preparation of this report, La Mancha took the conservative approach of applying the current royalty to the first phase of the VMS project while reducing the tax level to 10% to reflect the importance of the investment to be realised. As La Mancha believes that the second phase of the project represents a material change in terms of product mix, they have considered it as a new project for royalties purposes. La Mancha has therefore retained the assumption of NSRs of 5% and 3.5% respectively for gold and base metals in addition to a reduced tax rate of 10%. These assumptions appear to be conservative in light of past agreement with the Sudanese government, but remain to be agreed upon. 18.13 CAPITAL COST ESTIMATE 18.13.1 General Mining capital costs for both the CIL and VMS phases of development are contained within the relevant parts of Section 18.1, and are summarised in Section 18.13. CIL Plant capital cost estimation was undertaken by Sedgman and VMS Concentrator costs estimated by AMEC. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 239 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.13.2 Mining As described in Sections 18.2 to 18.4, total mine capital cost includes the purchase of a new mining fleet for Hadal Awatib, together with pre-production costs associated with the development of the Hassai South underground mine (Table 18.54). Only replacement capital costs have been included for mining at Kamoeb, for a portion of the aging mining fleet. No capital costs are included for the heap leach tailings reclaim operations which will use existing equipment. Table 18.54 Mine Capital Cost Estimate Summary – Kamoeb, Hassai South and Hadal Awatib Area Items Cost Estimate Equipment 8.8** ($M) Kamoeb Open Pit Hadal Awatib Open Pit Hasai South Underground Sub-total 8.8 Equipment 79.6 Infrastructure 5.0 Sub-total 84.6 Infrastructure 31.2 Development 23.8 Material movement 2.6 Sub-total Total Mine Capital Cost 57.6 151.0 Note: - ** includes capital from 2010 for current operations; $3.22 m of this amount is attributed to the CIL Project Underground infrastructure includes preliminary works, surface works, portal development, ventilation, and the paste plant. 18.13.3 CIL Plant 18.13.3.1 Methodology The designs for the processing plant are based on a limited amount of metallurgical testwork, data from which was used to establish preliminary design criteria, process flow diagrams, a preliminary mass balance, and an equipment list. Budget prices were obtained for major equipment only. Minor equipment pricing was derived from the Sedgman database of previous projects with an escalation factor applied where appropriate. Estimates for other construction costs were based upon factors derived from Sedgman experience with similar projects. These costs included: • Structural • Concrete • Platework • Piping FINAL – Rev 0 – 22 Oct 2010 AMEC Page 240 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report • Electrical • Transport. An allowance for a 200 man accommodation village has been included, determined at a rate of $20 000 per man. 18.13.3.2 Summary Capital Cost The estimated capital cost for the 3.0 Mt/a CIL processing plant is $184.4 M, as summarised in Table 18.55. All costs are in United States dollars as of second quarter 2010 and are judged to have an accuracy of ±40%. Table 18.55 Estimated Capital Costs, 3.0 Mt/a CIL Plant Area Cost ($000s) Plant 74 299 Nile pipeline 39 641 Powerline 25 364 Camp 4 000 Contingency @10% 11 948 Subtotal – Direct 155 652 EPCM 15 503 Insurance 2 335 First Fill and three month consumables 4 081 Spares 1 276 Capital spares 3 000 Contingency @10% 2 484 Subtotal – Indirect 28 765 Grand Total 184 417 18.13.4 VMS Concentrator 18.13.4.1 Introduction This section summarises the capital cost estimates for the VMS concentrator and related facilities and infrastructure. All costs are estimated in United States dollars as at the first quarter 2010 (1Q10) and are judged to have an accuracy of ±30%. Owner’s costs and contingencies were provided by AMC. Where appropriate, the exchange rates used in the estimate are: • A$1.00 = US$0.87 • EUR 1.00 = US$1.37 • ZAR 1.00 = US$0.127 The total estimated value of the capital expenditure for the 5 Mt/a plant is $319.4 M. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 241 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.13.4.2 Estimate Structure and Basis The capital cost estimates have been structured into the following major categories: • Direct Costs: those expenditures that include supply of equipment and materials, freight to site and construction labour at site relevant to the particular option. • Indirect Costs: indirect costs are those expenditures covering temporary construction facilities plus engineering, procurement and construction management (EPCM) services, together with the supervision of commissioning of the works. • Accuracy Provisions: plant accuracy provisions have been assigned to each discipline item to cover expected growth in the estimate based on experience with other projects of this nature at a scoping level. The provision ranges from 10% for equipment and equipment installation to 25% for piping. • The accuracy provisions for Mining and the TSF have been incorporated into their Direct Costs. • The accuracy provisions do not consider allowances for such subjective risks such as currency exchange rate fluctuations, construction market forces, environmental considerations, community input considerations, unusual weather conditions, labour availability, difficult ground conditions, change to statutory regulations, charges and taxes, and scope changes. Unit rates for bulk materials, ie. earthworks, concrete, steelwork, platework, etc., have been derived from recent projects and studies undertaken by AMEC for mining facilities in Africa. Typical all-up rates used in the estimate are listed below: • Concrete (in place) $1040/m³ (average) • Steelwork (supplied and erected) $5480/t (average) • CS platework (supplied and erected) $5500/t (average) • Site erected carbon steel tanks $8500/t (average) • Shop rubber lining $320/m² • Site rubber lining $506/m² Equipment costs are based on budget quotes received from vendors for major items such as crushers, screens, mills, filter and thickeners. In-house database or allowances have been used for certain equipment items where recent costing is available. The following vendor pricing for major equipment is incorporated into the estimates: • Crushers Metso • Grinding mills Outotec • Flotation cells Outotec • Regrind mills Xstrata Technologies • Filters Larox • Thickeners Outotec FINAL – Rev 0 – 22 Oct 2010 AMEC Page 242 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report An allowance on the net price of the equipment has been made to include for packers, wedges, grouting, guarding and signage. 18.13.4.3 Plant, Area and Regional Roads Road and hardstand costs were based on minimal bulk earthworks (cut to fill) and construction consisting of a nominal 500 mm embankment from suitable local fill sheeted with 200 mm of base course. Typical rates applied are listed below: • Strip topsoil and dispose $10.7/Bm³ • Embankment construction, from local fill $13.1/Cm³ • Base course – deliver, grade and compact $68.3/Cm³ 18.13.4.4 Buildings The costs for buildings have been developed using historical data from recent projects and studies undertaken by AMEC for mining facilities in Africa. Allowances have made for fully fitting out with air conditioning, verandas, partitions, workstations, computers, electrical appliances, etc. Nominal sums have been included for workshop equipment tools and services, and laboratory equipment. 18.13.4.5 Plant Mobile Equipment The mobile equipment fleet cost basis is derived from indicative prices received from various suppliers and in-house data. An allowance of 15% for mobilisation to the mine site has been added to the net amount. 18.13.4.6 Bulk Fuel Storage and Distribution Allowance has been made for two carbon steel storage tanks along with associated retention bunds, bunkering and distribution pumps, fire protection, unloading pipework, etc. Distribution pipelines are allowed to the mine refuelling tanker point and the light vehicle refuelling point. 18.13.4.7 Electric Power Supply It is proposed that power is supplied by a 77 km 132 kV overhead line connected to the Sudanese main grid – which has been costed separately with the CIL plant. Allowance has been made for a HV switchyard and step-down transformer at the project site. 18.13.4.8 Communications An allowance based on in-house data includes receival mast to mine, PABX interface, backbone transmission, voice/data cabling, administration voice and data network, fixed voice services and twoway mine voice mobile radio. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 243 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.13.4.9 Permanent Accommodation Expansion Permanent workforce accommodation is based on single status. Costs for the accommodation are based on historical data from previous studies for site accommodation in Africa. No allowance has been made for vacant accommodation when personnel are on R&R. The village will be fully selfcontained with all amenities. A rate of $12 500 per person has been estimated. 18.13.4.10 Water Supply and Distribution As the water pipeline from the Nile is anticipated to be installed during the construction of the Hassai CIL project, this has been excluded from the concentrator estimate. 18.13.4.11 Material Quantities Preliminary global quantities for earthworks, concrete, steelwork, and platework have been determined from in-house data for similar installations and assessed from conceptual sketches prepared for this study. Rates as noted earlier have been applied to these quantities. 18.13.4.12 Construction Labour Labour rates for the estimate have been calculated in-house based on rates from recent projects and studies undertaken by AMEC in Africa. The rates include accommodation and travel costs, supervision, construction plant and cranes, temporary facilities, and contractor’s mark-ups, to give a direct labour cost per hour. Site construction hours have been calculated using Australian norms as the basis. A productivity factor of 0.4 has been applied to these norms to reflect the estimated hours considered applicable in this region of Sudan. For earthworks, a productivity factor of 0.8 has been applied. 18.13.4.13 Piping Estimate Piping costs for the process plant have been calculated by applying factors to the equipment supply cost. These factors vary depending on the work breakdown structure (WBS) area. They are based on AMEC’s experience of similar installations. The off-plot lines have been calculated by applying material plus installation rates to preliminary quantities. 18.13.4.14 Electrical and Instrumentation Electrical and instrumentation costs have been calculated by applying factors to the equipment supply cost. These factors vary depending on the WBS area. These factors are based on in-house data for similar installations. 18.13.4.15 Freight Ocean, inland and local freight costs have applied as a percentage of the equipment and material costs. A percentage of 4% has been applied for local freight and 10% for ocean/inland freight. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 244 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.13.4.16 Preliminaries Preliminaries are those items that must be included in the estimate, but are applicable across a number of areas. They consist of: • Mobilisation and demobilisation of contractors • Heavy lift cranes for the installation of crushers, mills, filters and thickeners (included in labour gang rate) • Cost of vendors’ representatives to be present when commissioning equipment • Assistance by contractors when carrying out the commissioning the plant • Construction camp • Temporary facilities for the EPCM contractor. For this level of estimate, these items are simply factored off the appropriate areas. 18.13.4.17 Tailings Storage Facility TSF costs include the supply and installation of the tailings delivery system from the concentrator, the decant system, return pipeline from floating decant barge to the process plant, embankment instrumentation, seepage collection and return pumps, and perimeter diversion channels. The estimate is a preliminary estimation based on preliminary bill of quantities and estimated distances for piping systems as well as a typical layout for a tailings facility. Rates and quantities have not been calculated and the cost is based on escalation from similar facilities. The availability of approved fill materials from the open pit works has yet to be confirmed. As a consequence, suitable earth fill from this source has not been assumed at this stage. 18.13.4.18 Capital Spares Included is an amount equal to 5% of the ex-works equipment cost to cover capital and start-up spares; an allowance has been added to this to cover freight to site. 18.13.4.19 Indirect Costs – Temporary Facilities and EPCM Indirect costs have been calculated by applying a percentage amount to the Direct Costs. A percentage of 1.0% has been applied for temporary facilities such as the establishment of construction management temporary facilities, temporary infrastructure, etc. A percentage of 6% has been applied for the construction camp supply and running costs. For the EPCM, various percentages (dependent on work area) were applied to undertake detail design, procurement, project management, construction management and commissioning if undertaken by an engineering company familiar with this task. Also included within this EPCM amount are such things as specialist sub-consultants, travel costs, R&R, hire vehicles, accommodation and messing, insurances, etc. The EPCM overall averaged approximately 11.0% for the 5.0 Mt/a plant. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 245 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.13.4.20 Qualifications and Clarifications • GST or any like tax has not been included • No import duty (if applicable) has been included for importation of equipment • Minimal site preparation has been included. No allowance has been made for rock excavation, dewatering or sub-strata improvement. It has been presumed that the site is generally clear and ground conditions are suitable for the construction of the proposed works • It is assumed that suitable borrow pits for building materials are in close proximity and, where there is a requirement for select fill, it can be produced with a minimum of screening and water conditioning • Bulk fuel storage and dispensing to service the mining fleet is included • CAR or goods in transit insurances have been included within the EPCM amount • First fill and consumables are included. 18.13.4.21 Capital Cost Summary The cost breakdowns by WBS area, including for mining, are given in Table 18.56. Table 18.56 Concentrator Capital Cost Estimate Summary by Area WBS Area No. WBS Area Title 5.0 Mt/a (US$) Direct Costs Underground Open Pit Infrastructure 27 664 000 Horizontal development 14 053 725 Vertical development 1 500 000 Material movement 1 532 097 Infrastructure Equipment Mining Total 4 900 000 66 762 000 116 411 822 3200 Comminution/beneficiation 33 594 649 3350 Flotation 24 486 848 3400 Concentrate handling 9 309 606 3850 Tails disposal 5 185 792 3900 Reagents 1 440 889 3950 Air distribution 1 880 105 3955 Water distribution 1 292 845 3960 Main pipe racks Total – Process Plant 1 002 982 78 193 716 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 246 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.56 Concentrator Capital Cost Estimate Summary by Area WBS Area No. WBS Area Title 5.0 Mt/a (US$) 4050 Site preparation and improvements 4100 Substation buildings 1 166 148 4150 Plant buildings 3 075 022 4200 Fire protection 258 239 4250 Water treatment 200 000 4300 Sewage disposal and treatment 224 292 4350 Mobile equipment 4400 Bulk fuel storage and distribution 4450 Tailings delivery line 4500 Tailings storage facility 4550 833 396 5 502 750 257 895 Incl. with TSF 20 955 607 Incl. with TSF 4600 Storage ponds 4650 Main HV switch yard 3 470 000 4700 Control systems 1 644 972 4750 Communications Total – Plant Infrastructure 657 403 588 348 38 834 071 5050 Permanent accommodation 6 000 000 5100 Water distribution lines 1 500 000 5150 Electrical power distribution 1 500 000 5250 Area roads 3 850 000 5300 Area communications Total – Area Infrastructure 50 000 12 900 000 6050 Regional roads 0 6100 6150 Electrical power feeder Water transmission line 0 0 6250 Port facilities 500 000 Total – Regional Infrastructure 500 000 7050 First Fill reagents and consumables 7100 Ocean freight 4 225 437 7150 Spares 2 143 663 7250 Mobilisation and demobilisation 2 323 903 7300 Commissioning Total – Miscellaneous Total Direct Cost 1 129 275 1 594 706 11 416 984 258 256 594 Indirect Cost Construction facilities 1 418 448 Construction camp 8 510 686 EPCM PCM for turnkey packages Total Indirect Cost Total Direct Cost 27 540 379 916 512 38 386 025 296 642 619 Accuracy provision 22 789 684 Total Initial Capital 319 432 303 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 247 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.13.4.22 Sustaining Capital Sustaining capital has been estimated for the proposed project life and includes the following: • Mining – includes mine infrastructure, progressive underground mine development and equipment • Process plant – includes sustaining capital for plant replacement and upgrades over time, and tailings facility wall lifts to increase capacity progressively. • General capital – includes an allowance for capital required to upgrade site buildings, mobile equipment replacement, computer replacement etc. A summary of the sustaining capital allowance is shown in Table 18.57. Table 18.57 Concentrator Phase Sustaining Capital Estimate Allowance Area Estimate Mining 25 745 147 ($) Tailings storage facility Process plant and process infrastructure General capital Total Sustaining Capital 18.14 OPERATING COST ESTIMATE 18.14.1 CIL Plant 18.14.1.1 Methodology 3 728 873 27 501 530 4 700 000 61 675 550 Sedgman has developed operating cost estimates for the CIL process plant as described herein. The operating costs have been divided into the following discrete cost centres: • Labour • Operating consumables • Maintenance materials • Power • Transport. The operating costs have been determined from a variety of sources including: • Supplier quotations • AMC advice • Sedgman database for similar operations • First principal estimates. Costs are presented in United States dollars and are based on prices obtained during the second quarter of 2010. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 248 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report The costs have been developed for a plant availability of 91.3% (8000 h/a) including scheduled and unscheduled maintenance. Table 18.58 summarises the total CIL plant operating costs for the life of mine (LOM). Table 18.58 Total Annual Operating Cost Estimate, 3 Mt/a CIL Plant Area Units Cost ($) Labour $M 29.60 Operating consumables $M 106.38 Maintenance consumables $M 6.32 Power $M 31.49 Transport $M 29.42 $M 203.21 Total – Cost/annum Labour $/t ore 1.87 Operating consumables $/t ore 6.70 Maintenance consumables $/t ore 0.40 Power $/t ore 1.98 Transport $/t ore 1.85 Total – Cost/tonne of ore treated $/t ore 12.80 Labour $/oz Au 36.52 Operating consumables $/oz Au 131.24 Maintenance consumables $/oz Au 7.80 Power $/oz Au 38.86 Transport $/oz Au 36.29 Total – $/oz of gold produced $/oz Au 250.71 18.14.1.2 Labour A summary of estimated labour costs is provided in Table 18.59. Labour costs are based on recent rates from Botswana, which are expected to be similar to those in Sudan. A total plant workforce of 192 has been estimated, excluding both mining and administration personnel. The workforce is comprised mainly of local labour, with expatriates initially used for senior management roles. 18.14.1.3 Operating Consumables Identified consumables and their anticipated consumption rates are given in Table 18.60. The analysis assumes the exhaustion of fresh ore supplies at the end of operating year 4 and with it the cessation of crushing circuit operations. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 249 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.59 Annual Processing Plant Labour Costs, 3 Mt/a CIL Plant Position Number Total Cost Total Cost ($/a) ($/t dry feed) Operating Labour Senior Foreman Day Foreman 4 1 229 516 47 876 0.08 0.02 Shift Foreman 4 191 504 0.06 Control Room Operators Crush/Reclaim 4 16 42 156 168 624 0.01 0.06 Grind 8 84 312 0.03 Leach/Adsorption Elution/Goldroom 8 8 84 312 84 312 0.03 0.02 Reagents 8 84 312 0.03 Water/Tails/Services Mobile Equipment 8 8 84 312 84 312 0.03 0.03 Trainees 16 140 384 0.05 Laboratory Metallurgical Clerk 16 1 168 624 10 539 0.06 0.00 Maintenance Mech/Elect/Inst Supervisors 4 229 516 0.08 Shift Tradesmen 24 790 584 0.26 Shift Electrical Shift TA's 8 24 263 528 252 936 0.09 0.08 Align & Cond. Monitoring 1 47 876 0.02 Mobil Equip Ops Materials Controller 2 4 21 078 191 504 0.01 0.06 Maint. Planner 2 95 752 0.03 Day Crew Tradesmen 2 65 882 0.02 Trades Assistants Maintenance Clerk 2 1 21 078 10 539 0.01 0.00 Mill Manager 1 382 899 0.13 Plant Metallurgist Junior Metallurgists 2 3 555 314 172 137 0.19 0.06 Mechanical Engineer 1 277 657 0.09 Mech/Elect/Inst Supt 1 277 657 0.09 Electrical Engineer Chief Chemist 1 1 277 657 57 379 0.09 0.02 0.25 120 724 0.04 192.25 5 595 714 1.87 Professional Consultants Total FINAL – Rev 0 – 22 Oct 2010 AMEC Page 250 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.60 LOM Plant Consumable Costs, 3 Mt/a CIL Plant Identified Consumables Usage Rate $’000 Crusher Consumables Primary Crusher liners 70 000 t/set 842 Grinding Consumables SAG mill grinding media 0.910 kg/t 5 791 4 sets/a 3 960 1.150 kg/t 21 866 2 sets/a 3 969 Quicklime 1.250 kg/t 2 562 Sodium cyanide 0.823 kg/t 48 795 Sodium hydroxide 0.117 kg/t 702 Hydrochloric acid 0.053 kg/t 514 SAG mill liners Primary ball mill grinding media Primary ball mill liners Reagents Activated carbon 0.020 kg/t 910 Flocculant 0.025 kg/t 1 587 SMBS 0.690 kg/t 6 571 Copper sulphate 0.081 kg/t 3 278 Hydrated lime 0.480 kg/t 2 590 Miscellaneous Laboratory consumables 87 Plant assay costs 681 Diesel costs – mobile plant 85 450 L/a 181 Diesel costs – process plant 703 560 L/a 1 489 Total 18.14.1.4 106 376 Maintenance Consumables Maintenance consumables were estimated at 5% of the total mechanical and electrical capital costs per annum, as per industry standard. Mobile plant expenditure was assumed at $30 000 per annum. Table 18.61 summarises the maintenance consumable costs. Table 18.61 Annual Maintenance Consumable Costs, 3 Mt/a CIL Plant Description Cost ($/a) Fixed Plant 1 165 000 Mobile Plant 30 000 Total 18.14.1.5 1 195 000 Power Costs Power for the Hassai project has been assumed to be predominantly derived from the electrical grid supply. The unit cost for the grid power has been specified by AMC at $0.074/kWh. Power supply to the Nile pipeline has been assumed to be derived from diesel generators; Sedgman has assumed a diesel consumption of 0.26 L/kWh, with the diesel cost of $0.40/L. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 251 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report The installed operating equipment power demand was determined from vendor data or first principal estimates. Utilisation and load factors were then applied to the total to obtain power usage. Table 18.62 summarises the annual power costs for the Hassai CIL plant and associated services. Table 18.62 Estimated Annual Power Costs, 3 Mt/a CIL Plant Description Estimated Power Estimated Cost Estimated LOM Cost (MWh/a) ($’000/a) ($’000) Crushing 1 550 115 459 Reclaim 1 225 91 538 Grinding 47 988 3 551 18 787 Leach/adsorption 10 784 798 4 222 Elution 114 8 45 Gold room 297 22 116 3 183 236 1 246 410 30 161 2 034 150 796 67 584 5 001 26 370 9 315 969 5 125 76 899 5 970 31 495 Tailings Reagents Services Subtotal – Process Plant Nile River pipeline Total Power Usage 18.14.1.6 Transport Costs Transport costs have been estimated based upon sea container rates of $7717.5/container, and road transport rates of $0.187/t. 18.14.1.7 Pre-production Costs Allowances for pre-production labour and associated costs have been excluded from this estimate. 18.14.1.8 Overall Operating Costs Overall process operating costs for the 3.0 Mt/a CIL Plant are estimated to be $203.2 M for the LOM, as summarised in Table 18.58. These costs translate to $12.80/t ore for processing costs, or $250.71/oz recovered. 18.14.2 VMS Concentrator The operating costs for the VMS concentrator has been developed by AMEC based on the process design criteria, mass balance and mine plan described earlier in this report. Costs are in United States dollars and reflect an estimate base date of June 2010. The accuracy of the operating cost estimate is ±30% and reflects the plant operating at design capacity. AMEC obtained information for reagents, consumables and maintenance costs based on its in-house database, while AMC provided the operating cost estimate for transport, labour rates and salaries, and unit power cost. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 252 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report The operating cost estimate has been based on the process design criteria, mass balance, mine plan and schedule developed by AMEC. 18.14.2.1 Exclusions No allowance has been made in this operating costs estimate for items such as: • Concentrate treatment and refining charges • Financing charges • Contingency • Escalation or exchange rate variations • Depreciation and accounting effects • General and Administration (including administrative management and support, logistics, environmental, safety, camp and security personnel). These costs have been compiled by AMC based on current Hassai operational costs. 18.14.2.2 Exchange Rates Where appropriate, an exchange rate of A$1.00 = US$0.75 has been applied. 18.14.2.3 Process Plant Operating Costs The average plant operating costs are outlined in Table 18.63. These costs are based on a design basis throughput and feed grade. Annual costs based on a yearly mine plan were prepared for financial analysis. 18.14.2.4 Labour The manning schedule was developed reflecting 12 hour continuous shift rosters and coverage requirements for leave entitlements. The process manning schedule is summarised in Table 18.64. Labour costs have been based on data supplied by AMC from the Hassai site, with expatriate rates based on AMEC’s database. Table 18.63 Average Process Operating Costs, 5 Mt/a VMS Concentrator Labour 5 568 965 Power 11 603 311 Reagents Consumables Maintenance materials Product transport Total US$/t ore 6 226 532 13 367 481 2 719 925 7 418 638 46 904 851 9.38 US$/t Cu 856 US$/lb Cu 0.39 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 253 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.64 Summary of Operations Labour Structure, 5 Mt/a VMS Concentrator Section Number Production 226 Metallurgy and Laboratory 34 Maintenance 27 Total Process Plant Operating Labour 18.14.2.5 287 Reagents Reagent consumption rates are based on testwork data and on information supplied by AMC. Reagent supply costs are based on AMEC’s database, and include transport to the Project site. Table 18.65 summarises average consumption quantities, and unit and average annual costs for major reagents and water. Table 18.65 Major Reagent Costs, 5 Mt/a VMS Concentrator Reagent Unit Cost Annual Consumption 3 (t/ML/m ) 3 Water Annual Cost (US$) $0.62/m 3 972 800 2 463 136 Collector $2140/t 270 577 800 Frother $3140/t 177 557 036 Flocculant $3540/t 198 702 336 Quicklime $194/t 9 908 1 926 224 Total Cost 18.14.2.6 6 226 532 Power The power consumption is estimated from the installed power values presented in the equipment lists. Suitable utility factors are applied to reflect the operating power draw of each equipment item, and the resultant operating power draw is converted to the annual power usage requirement by application of the relevant annual operating hours for the equipment item. The power unit cost used by AMEC is US$0.074/kWh, as supplied by AMC for national grid power. Power costs for each plant area and throughput case are summarised in Table 18.66. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 254 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.66 Power Costs, 5 Mt/a VMS Concentrator Crushing/ grinding (3200) 7 926 960 Flotation (3350) 1 938 927 Concentrate filtration (3400) 261 790 Tailings handling (3850) 645 857 Reagents (3900) 25 915 Services (3950 and 3955) 753 542 Administration and Other 50 320 Total Power Costs 18.14.2.7 11 603 311 Consumables Consumables include items such as crushing and grinding media and liners, packaging materials required for the despatch of the intermediate product, filter cloths, lubricants and laboratory reagents. The consumable costs were supplied by AMEC and based on costs from similar projects. The total consumables cost for each case are summarised in Table 18.67. Table 18.67 Consumables Cost Summary, 5 Mt/a VMS Concentrator Consumable Unit Cost Crusher liners $4.0/kg Mill liners, AG mill $2.4/kg 360 000 Mill liners, regrind mill $2.4/kg 360 000 Mill liners, SAG $2.4/kg 600 000 Laboratory/samples Grinding media, ball mill $20/sample $ 800 000 547 500 $1100/t 3 300 000 Grinding media, SAG $1100/t 4 400 000 Grinding media, regrind mill $5040/t 277 200 Allowance 250 000 Diesel for mobile equipment/miscellaneous Allowance 1 500 000 Mobile and hire equipment Allowance 100 000 Copper concentrate filter cloths Bags for product, 2 t Total Consumables Costs 18.14.2.8 $8/bag 872 781 13 367 481 Maintenance Materials The maintenance section of the operating cost estimate has been based on percentage factors applied to the direct capital cost for the various plant areas. The percentage factors applied and annual costs are summarised in Table 18.68. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 255 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.68 Maintenance Materials, 5 Mt/a VMS Concentrator Plant Area Units Basis (%) Comminution/beneficiation % 5.0 Flotation % 2.0 489 737 Concentrate handling % 2.0 186 192 Tails disposal % 2.0 103 716 Reagents % 2.0 28 818 Air distribution % 1.0 18 801 Water distribution % 1.0 12 928 Mobile equipment $ Allowance 200 000 Total Maintenance Costs ($/annum) 18.14.2.9 $ 1 679 732 2 719 925 Product Transport Product transport costs of $34/t have been based on cost data supplied by AMC for transport from site to Port Sudan. At full production, an average of 217 647 t/a of concentrate will be transported at an annual cost of $7.4 M. 18.14.3 Mining Costs As discussed earlier in Section 18, mine operating costs and schedules have been determined by CSA for Kamoeb and heap leach tailings and stockpile reclaim mining, and by AMEC for mining at Hadal Awatib and Hassai South. Costs are summarised in Table 18.69, with details in the relevant parts of Sections 18.2 to 18.4. Table 18.69 Summary of LOM Mine Operating Cost Estimates Area Cost $/t ore Heap leach tailings and stockpile reclaim 1.15 Kamoeb 20.07 Hadal Awatib open pit 14.14 Hassai South underground 26.17 18.14.4 General and Administration G&A and Other costs have been provided by AMC based on 2009 Hassai site data as shown in Table 18.70, revised for increased throughput. The total of 6 845 330 Euros translates to approximately $9.41 M. Going forward, these costs are estimated to be $8.500 M/a for the CIL operation due to improved efficiencies and reduced manpower requirements, increasing again to $9.241 M/a for the combined CIL+VMS operation. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 256 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.70 AMC 2009 G&A Costs in Euros as Basis for Scoping Study Hassai Safety Hassai Management Medical Hassaï Internal logistic on site Camp maintenance Camp food Camp cleaning Atbara house PZU office expenses Warehouse expenses Plane External beneficiaries Training Ariab development fund AMC Board Khartoum Bank charges Total 18.15 PROJECT ECONOMICS 18.15.1 Overview 51 626 1 161 124 458 553 309 428 574 537 658 109 116 105 6 221 175 826 444 253 372 312 291 429 57 834 247 126 163 825 1 736 318 20 704 6 845 330 La Mancha has prepared financial models for: • The Base Case, ie the existing heap leach project, reserves for which are due to be exhausted by the end of 2013. • The CIL Phase, starting in 2013 and treating then-existent oxide gold reserves, stockpiled acidic mineralisation and heap leach residue resources as identified in this report. • The VMS Phase, starting in 2015 using the VMS Mining Inventory defined in the AMEC mining study. The production profile (gold and gold equivalent copper) for the combined heap leach + CIL + concentrator envisaged business plan is illustrated in Figure 18.24. Copper has been related to gold taking account of metal prices and metallurgical recoveries. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 257 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 18.24 Metal Production Profile 500,000 450,000 VMS Copper as Gold eqv. 400,000 VMS Concentrate Heap Leach Residue Gold Production, oz 350,000 Acidic non‐washable ore 300,000 Acidic washable ore SBR ore 250,000 Quartz ore 200,000 150,000 100,000 50,000 ‐ 2010 2011 2012 2013 2014 2015 2016 2017 2018 2019 2020 2021 2022 2023 2024 2025 It is emphasised that economic evaluations of the CIL and VMS phases are based in part on mining of Inferred Mineral Resources which are too speculative geologically to have economic considerations applied to them that would enable them to be categorised as Mineral Reserves. It cannot be assumed that all or any part of an Inferred Mineral Resource will be upgraded to an Indicated or Measured Mineral Resource as a result of continued exploration. The NI43-101 regulations confirm that confidence in the estimate is considered insufficient to allow the meaningful application of technical and economic parameters or to enable an evaluation of economic viability worthy of public disclosure. On that basis, Inferred Mineral Resources must be excluded from estimates forming the basis of feasibility or other economic studies. However, the regulators have allowed an exemption for La Mancha to report the results of the preliminary assessment based on Inferred Resources, providing the above qualifications are clearly noted. The term Mining Inventory has been introduced to distinguish the Inferred Resources contained within preliminary mining designs from Mineral Reserves. Inputs for the financial models include: • Production schedules and mining costs prepared by AMC for the Base Case, CSA for the CIL Phase and AMEC for the VMS Phase, as shown in tables in relevant parts of Section 18.1. Note that for the financial model, La Mancha has added an additional 372 088 t at 4.29 g/t Au spread over 3 years to bring the scheduled tonnages up to 3.0 Mt/a. This material has been identified overlying VMS mineralisation in the Hadal Awatib pit design. • Process recoveries determined by AMC (current operation and CIL) and AMEC (VMS), as provided in Section 16. • Capital and site operating costs by AMC (current operation), Sedgman (CIL) and AMEC (VMS), as provided in Sections 18.1 and 18.13. Capital investment is assumed to be over a 2 year period for each expansion phase, at a ratio of 30:70. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 258 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report • Off-site operating costs as determined by AMC for current heap leach and proposed CIL operations. • Off-site operating costs for copper concentrates of $50/t for transport ex-Port Sudan, TCs of $50/t concentrate and RCs of $0.50/lb copper. • Payment terms: • − 96.5% of contained copper in concentrates − 100% payable gold. Gold and copper prices of $950/oz and $2.19/lb, respectively, over the life of the project. The models include provisions for royalties and income tax, and are on a 100% equity financing basis. A discount rate of 5% has been applied. Note that at this level of study, no provision has been made for working capital or for delay in accruing income from concentrate sales. Key assumptions and financial highlights from the CIL and VMS models are shown in Table 18.71. Total gold and copper production is estimated at 1.19 Moz and 0.323 Mt, respectively. Average cash costs are $482/oz gold in the CIL circuit and $1.24/lb copper from the concentrator (including gold credits). The basic findings for the three scenarios are as follows: • The current operation is scheduled to process 2.6 Mt between 2011 and 2013, producing 0.30 Moz, which provides a NPV of $36 M at a 5% discount rate. • The CIL Phase processes 15.3 Mt for 2013-2018, producing 0.81 Moz, and shows an NPV (5%) of $149.8 M and an IRR of 30%. Average cash costs are $482/oz gold in the CIL circuit. • The VMS Phase treats 29.4 Mt between 2015 and 2025, producing 0.32 Mt of contained copper and 0.38 Moz of contained gold in concentrates. The NPV is $122.7 M and the IRR 11%. • The combination of the CIL and VMS Phases operating between 2013 and 2025 have an NPV (5%) of $239 M and an IRR of 17%. Sensitivity analysis was undertaken for the expansion phases, confirming that the combined project (CIL and VMS, without production from 2010 – 2012) is very sensitive to revenue (metal price or grade): an increase of 10% in gold price adds approximately $80 M to NPV (NAV), while a similar increase in copper price adds $91 M. The combined project is also highly sensitive to operating costs with an increase of 10% leading to a $100 M reduction in NPV, whereas a similar increase in capital cost reduces the combined NPV by $47 M. NPV (NAV) is calculated using 5% discount rate. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 259 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.71 Financial Highlights for Proposed VMS Project, by Phase Heap Leaching Phase 1: CIL Phase 2: VMS USD 950/oz USD 950/oz USD 950/oz -- -- USD 2.19/lb 7% 7% 5% -- -- 3.5% Phase 1 and 2 Main Assumptions Gold price Copper price Royalties (%) Gold Copper Corporate tax rate 15% 10% 10% 2.6 @ 4.88 0.6 @ 6.0 - Measured Resources (Mt@g/t) - 3.8 @ 1.9 - Indicated Resources (Mt@g/t) - 4.6 @ 2.1 - Inferred Resources (Mt@g/t Au, Cu%) - 6.8 @ 1.7 29.4 @ 1.1, 1.2 2.6 15.8 29.4 4.88 2.01 1.11 1.43 -- -- 1.22 1.22 2010 - 2013 2013 2015 -- Mineral Reserves Probable Reserves (Mt@g/t) Additional Mineral Resources Total Mining Inventory Tonnes, Mt Grades Gold, g/t Copper, % 45.2 Production Commissioning Yearly mill run rate, Mt/a 0.65 3 5 -- Gold recovered, ‘000 oz 299 811 378 1 189 Copper recovered, ‘000 t Metallurgical recovery Gold Copper Yearly production* -- -- 323 323 73% 79% 36% -- -- -- 90% -- Gold (oz) 74 780 155 880 59 355 -- Copper (t) -- -- 51 516 4 6 10 6+ $185.6 M $319.4 M $505.0 M $4.9 M $35.9 M $40.8 M Mine life, years Financials Initial capital cost Total sustaining capital Average cash costs $ 482/oz Au $ 1.24/lb Cu*** - Internal rate of return 30% 11% 17% NPV @ 0% discount $195.8 M $230.9 M $447.1 M NPV @ 5% discount $149.8 M $122.7 M $238.7 M 1.9 3.9 varies Payback** , years * Excludes low production in final year ** Calculated from commencement of production *** Including gold credit FINAL – Rev 0 – 22 Oct 2010 AMEC Page 260 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.72 Sensitivity Analysis Matrix– Gold Price Metal Prices Sensitivity Analysis Copper Price NAV @ 5% discount rate 238,666 4,000 4,400 4,800 5,200 5,600 6,000 6,400 6,800 7,200 7,600 8,000 8,400 8,800 9,200 600 -207,502 -125,856 -47,603 29,175 105,531 181,740 257,948 334,095 410,223 486,352 562,480 638,608 714,736 790,865 650 -164,234 -83,892 -6,031 70,728 146,995 223,203 299,412 375,547 451,675 527,803 603,931 680,060 756,188 832,316 700 -121,356 -42,410 35,187 111,905 188,124 264,332 340,535 416,663 492,792 568,920 645,048 721,177 797,305 873,433 750 -79,652 -929 76,405 153,044 229,253 305,461 381,652 457,780 533,909 610,037 686,165 762,293 838,422 914,550 800 -38,677 39,980 117,050 193,601 269,809 346,018 422,196 498,325 574,453 650,581 726,710 802,838 878,966 955,094 850 900 2,192 43,061 80,713 121,314 157,589 198,128 234,050 274,500 310,259 350,708 386,467 426,917 462,634 503,072 538,762 579,200 614,891 655,328 691,019 731,456 767,147 807,585 843,275 883,713 919,404 959,841 995,532 1,035,970 Gold Price 950 1,000 1,050 1,100 1,150 1,200 83,866 124,668 165,470 206,272 247,074 287,747 161,907 202,446 242,985 283,523 324,062 364,601 238,666 279,163 319,632 360,090 400,539 440,989 314,949 355,399 395,849 436,298 476,748 517,198 391,158 431,608 472,057 512,507 552,956 593,406 467,366 507,816 548,256 588,694 629,131 669,569 543,509 583,947 624,384 664,822 705,259 745,697 619,637 660,075 700,513 740,950 781,388 821,825 695,766 736,203 776,641 817,078 857,516 897,953 771,894 812,332 852,769 893,207 933,644 974,082 848,022 888,460 928,897 969,335 1,009,772 1,050,210 924,151 964,588 1,005,026 1,045,463 1,085,901 1,126,338 1,000,279 1,040,716 1,081,154 1,121,591 1,162,029 1,202,467 1,076,407 1,116,845 1,157,282 1,197,720 1,238,157 1,278,595 1,250 328,347 405,139 481,439 557,647 633,856 710,006 786,134 862,263 938,391 1,014,519 1,090,648 1,166,776 1,242,904 1,319,032 1,300 368,913 445,635 521,888 598,097 674,305 750,444 826,572 902,700 978,829 1,054,957 1,131,085 1,207,213 1,283,342 1,359,470 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 261 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.73 Sensitivity Analysis Matrix – Operating Costs Operating and Capital Costs Sensitivity Analysis Capital Costs NAV @ 5% discount rate 238,666 -14% -12% -10% -8% -6% -4% -2% 0% 2% 4% 6% 8% 10% 12% 14% -14% 443,289 433,623 423,956 414,290 404,623 394,957 385,290 375,624 365,957 356,291 346,624 336,958 327,292 317,625 307,959 -12% 423,734 414,068 404,401 394,735 385,069 375,402 365,736 356,069 346,403 336,736 327,070 317,403 307,737 298,070 288,404 -10% 404,180 394,513 384,847 375,180 365,514 355,847 346,181 336,514 326,848 317,182 307,515 297,849 288,182 278,516 268,849 -8% 384,625 374,959 365,292 355,626 345,959 336,293 326,626 316,960 307,293 297,627 287,960 278,294 268,627 258,961 249,295 -6% 365,070 355,404 345,737 336,071 326,404 316,738 307,072 297,405 287,739 278,072 268,406 258,739 249,073 239,406 229,740 -4% 345,516 335,849 326,183 316,516 306,850 297,183 287,517 277,850 268,184 258,517 248,851 239,185 229,518 219,852 210,185 Inputs: Operating Costs -2% 0% 2% 325,948 306,331 286,634 316,281 296,665 276,967 306,615 286,998 267,301 296,948 277,332 257,634 287,282 267,666 247,968 277,616 257,999 238,301 267,949 248,333 228,635 258,283 238,666 218,968 248,616 229,000 209,302 238,950 219,333 199,636 229,283 209,667 189,969 219,617 200,000 180,303 209,950 190,334 170,636 200,284 180,667 160,970 190,617 171,001 151,303 Op cost over-run 4% 266,936 257,270 247,603 237,937 228,270 218,604 208,937 199,271 189,604 179,938 170,271 160,605 150,938 141,272 131,606 6% 247,238 237,572 227,905 218,239 208,572 198,906 189,240 179,573 169,907 160,240 150,574 140,907 131,241 121,574 111,908 0% 8% 227,541 217,874 208,208 198,541 188,875 179,208 169,542 159,875 150,209 140,542 130,876 121,210 111,543 101,877 92,210 Cap cost over-run 10% 207,790 198,124 188,457 178,791 169,124 159,458 149,792 140,125 130,459 120,792 111,126 101,459 91,793 82,126 72,460 12% 187,997 178,331 168,664 158,998 149,331 139,665 129,998 120,332 110,666 100,999 91,333 81,666 72,000 62,333 52,667 0% 14% 167,973 158,307 148,640 138,974 129,307 119,641 109,974 100,308 90,642 80,975 71,309 61,642 51,976 42,309 32,643 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 262 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report The CIL and VMS phases each operate at full throughput for only 5 years. The financial outcomes are, therefore, significantly affected by increases in Mining Inventory. The potential to increase resources, particularly of VMS material, is considered to be high, with VMS mineralisation known to lie at the base of four additional pits and with a number of untested electrical conductors identified during exploration. In addition, there is a belief that the underground inventory may be artificially diluted and that changes to the resource modelling approach might increase head grade to the plant. 18.15.2 Individual Phase Description 18.15.2.1 Current Operations: Heap Leaching Current operations are based on heap leaching of Mineral Reserves of 2.6 Mt of ore grading 5 g/t Au for 410 000 oz of contained gold. The mine-life is expected to be 4 years to the end of 2013. Plant feed consists of Quartz Ore, SBR Oxide Ore and Acidic Ore. The theoretical Production Profile is shown in Figure 18.25. Figure 18.25 Metal Production Profile – Heap Leach Operations 2010-2013 500,000 Gold Production, oz 450,000 400,000 Acidic non‐washable ore 350,000 Acidic washable ore 300,000 SBR ore 250,000 Quartz ore 200,000 150,000 100,000 50,000 ‐ 2010 2011 2012 2013 2014 2015 2016 2017 2018 2019 2020 2021 2022 2023 2024 2025 Processing capacity averages 700 kt of ore per annum with gold recovery averaging 73%. production averages 75 000 oz per annum. Gold Theoretical production and financial outcomes are illustrated in Table 18.74. It is important to note that economic outcomes do not exactly reflect projections for current years as the gold price used is $950/oz, whereas the current realised gold price is substantially above this amount. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 263 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Acidic ore stockpiled from Oxide ore pits comprises a significant portion of the plant feed. In order to treat this material, the addition of a washing plant was studied in 2006. This upgrade would be economically viable, but, the capital cost (then estimated at $35.5 M) for the washing plant and water pipeline from the Nile River is significant and not easy to support on this reserve. Table 18.74 Production and Cashflow Projections – Heap Leach Project 2,010 2,011 2,012 2,013 Tot / Avg Physical Data Tonnes of Ore Mined Tonnes of waste Tonnes milled Gold Grade (g/t) Recovery (%) Gold Production (oz) 474,695 643,272 537,740 300,000 1,955,706 5,694,566 4,669,373 3,299,665 2,000,000 15,663,604 751,376 631,835 650,598 582,000 2,615,808 4.22 4.96 3.75 6.92 4.88 70% 73% 74% 74% 73% 71,728 73,079 57,871 96,456 299,134 Profit and Loss Statement (in '000 US$) Revenues 68,142 69,425 54,978 91,633 284,177 Cost of Sales Mining and Milling Costs Mining Costs Haulage Costs Milling cost G&A and Other Costs Office / Administration Government Royalties Stock Variation 64,126 40,605 16,978 710 22,917 23,521 9,241 4,770 9,510 48,110 34,276 14,036 969 19,271 13,834 9,241 4,860 -267 47,085 30,117 9,440 833 19,843 16,969 9,241 3,848 3,879 51,882 26,534 5,658 465 20,411 25,348 9,241 6,414 9,693 211,203 131,532 46,113 2,977 82,442 79,671 Gross Margin Depreciation & Amortization of capital assets Gross Margin Cominor Mine Operating Income Income tax Net Earnings (Loss) 19,892 22,815 4,016 21,314 7,892 39,751 72,974 13,198 80 11,098 426 11,428 158 10,223 795 45,946 1459 -9,101 10,643 -3,377 30,323 28,487 -1,365 1,596 -507 4,548 4,273 -9,101 9,046 -3,377 25,775 22,342 13,606 19,878 11,929 45,690 91,103 Cash Flow Statement (In '000 USD) Cash flows from operating activities Cash flows from investing activities Non-Plant Capital Expenditures Washing Plant for Acidic SBR Water line Cash flows from financing activities 9,066 9,066 1,880 1,380 500 35,000 0 10,000 25,000 0 0 0 45,946 10,446 10,500 25,000 0 0 0 0 0 Free Cash Flow to Equity 4,540 17,998 -23,071 45,690 45,157 NPV @ 5% Discounting 35,684 While the current heap leach operation is viable, the requirement for additional water, the high operating costs for the wash plant and the existence of approxiumately 10 Mt of gold-bearing heap leach residue led to the consideration of a CIL plant installation. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 264 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.15.2.2 Phase One Expansion: CIL Plant Processing of ore via a CIL Plant provides a number of opportunities due to substantially increased efficiencies. CIL provides washing of acidic ore as above, but incorporated in the plant and water pipeline designs. In addition CIL provides the following improvements: • Reduced unit operating costs due to increased capacity from 0.7 Mt/a to 3.0 Mt/a • Increased gold recovery from 73% to 93% for fresh ore • Increased mining inventory due to lower costs and higher recovery, especially for the existing Quartz Resource • Ability to treat the Heap leach residue recovering 70% of the remaining gold. At a much higher capacity the CIL extends the mine-life by five years to 2018. Average gold production for the CIL Plant is 156 000 oz during the full-year operations, ie. double that of the heap leach operation. The production profile is shown in Figure 18.26. Note that the recoveries used to determine metal production have been determined by La Mancha, taking account of the difference between cyanidable and fire assays as appropriate (see Section 16.2.4). Figure 18.26 Metal Production Profile – CIL Phase, 2013-2018 500,000 Gold Production, oz 450,000 400,000 Heap Leach Residue 350,000 Acidic non‐washable ore 300,000 Acidic washable ore 250,000 SBR ore Quartz ore 200,000 150,000 100,000 50,000 ‐ 2010 2011 2012 2013 2014 2015 2016 2017 2018 2019 2020 2021 2022 2023 2024 2025 Quartz Ore contribution is significant from the end of 2012. Reduced costs lower the cut-off grade for the existing mine design, allow deeper development of existing pits, and development of new pits. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 265 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Acidic Ore is blended in the feed over the first two years. Gold production from this ore is increased due to higher gold recovery over heap leaching. Three plant capacities were examined: 2 Mt/a, 3 Mt/a and 4 Mt/a. The NPV increased with increasing capacity because the heap leach residue is easily reclaimed at higher rates (it does not require mining capital). The 3 Mt/a capacity case was chosen to provide a reasonable ‘stand-alone” mine-life and to allow water capacity for the subsequent VMS concentrator. Table 18.75 contains the physical production data and related projected cashflow analysis. Table 18.75 Production and Cashflow Projections – CIL Phase 2,011 2,012 2,013 2,014 2,015 2,016 2,017 2,018 Tot / Avg Physical Data Tonnes milled Gold Grade (g/t) Recovery (%) Gold Production (oz) 3,000,000 3,000,000 3,000,000 3,000,000 3,000,000 871,489 17,905,297 2.61 2.09 1.96 1.92 1.62 1.62 2.27 83% 82% 80% 78% 70% 70% 79% 209,801 165,133 150,687 144,230 109,566 31,828 811,245 Profit and Loss Statement (in '000 US$) 199,311 Revenues 156,876 143,152 137,018 Cost of Sales Mining and Milling Costs Mining Costs Haulage Costs Milling cost 92,683 54,183 10,342 3,106 40,735 83,852 61,912 19,484 3,671 38,758 74,607 56,087 14,088 3,598 38,400 64,717 46,626 4,847 3,379 38,400 G&A and Other Costs Office / Administration Government Royalties (X% of sale) Stock Variation 8,500 13,952 16,049 8,500 10,981 2,459 8,500 10,021 0 8,500 9,591 0 106,628 Gross Margin Depreciation & Amortization of capital assets Gross Margin Cominor Mine Operating Income Income tax Net Earnings (Loss) 104,087 30,237 57,606 17,265 41,820 12,149 0 0 3,420 993 38,400 11,155 8,500 7,286 0 3,000 2,117 0 12,972 770,683 390,729 272,776 89,216 20,680 246,178 45,498 53,948 18,507 73,025 68,545 72,302 46,482 36,010 2133 36,010 1460 36,010 1371 36,010 1446 36,010 10,461 930 259 190,512 7599 72,750 38,475 33,906 37,737 11,401 197,040 2,770 379,953 7,275 3,847 3,391 3,774 1,140 277 19,704 65,475 34,627 30,515 33,964 10,261 2,493 177,336 117,534 73,096 66,526 69,974 46,271 12,954 386,356 Cash Flow Statement (In '000 USD) Cash flows from operating activities Cash flows from investing activities Non CIP Capital Expenditures Mill - CIP 3mtpa Direct costs Mill - CIP 3mtpa Indirect costs Water line Sustaining CAPEX 54,725 130,912 0 3,220 34,203 79,808 8,630 20,136 11,892 27,748 0 0 0 0 0 375 0 0 750 0 0 750 0 0 750 0 0 750 0 0 0 190,512 3,220 114,011 28,766 39,641 4,875 375 750 750 750 750 0 0 0 0 0 0 0 0 Free Cash Flow to Equity -54,725 -130,912 117,534 72,721 65,776 69,224 45,521 12,204 195,843 NPV @ 5% Discounting 149,764 Cash flows from financing activities IRR 0 30% The CIL Phase has been evaluated separately from the Heap Leach operation. Only revenue and costs related to the CIL are used. CIL production begins in 2013. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 266 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Capital expenditure of $186 M consists of: • $3 M for additional mine fleet for the extended Quartz (Kamoeb) mine activity • $40 M for the water pipeline from the Nile River • $143 M for plant and infrastructure (including a 77 km powerline). Cash operating costs decrease from $706/oz gold with heap leaching to $482/oz for the CIL plant. Gold production increases from 299 koz with the heap leach only to 1 034 koz with CIL included. Gold production during CIL operation is 811 koz. Heap leach production to the end of 2012 is 203 koz. Infrastructure for the CIL phase was designed with consideration for the expected VMS phase. Water and power supply was sized to suit the CIL operation running in parallel with the VMS operation. 18.15.2.3 Phase Two Expansion: VMS Phase Resource identification of approximately 50 Mt of VMS mineralisation created a new expansion opportunity and a need to consider the CIL and VMS development in a new business plan. Figure 18.27 illustrates how the VMS project changes the AMC business plan. Copper production has been converted to gold equivalent, taking account of metal prices and recoveries, for illustration purposes in the chart, but is shown separately in Table 18.76. Figure 18.27 Metal Production Profile – VMS Phase, 2015-2025 500,000 450,000 VMS Copper as Gold eqv. 400,000 VMS Concentrate Heap Leach Residue Gold Production, oz 350,000 Acidic non‐washable ore 300,000 Acidic washable ore SBR ore 250,000 Quartz ore 200,000 150,000 100,000 50,000 ‐ 2010 2011 2012 2013 2014 2015 2016 2017 2018 2019 2020 2021 2022 2023 2024 2025 Three capacities were reviewed during the scoping study: 2, 3.5 and 5 Mt/a. The 5 Mt/a case gave the best NPV, although the 3.5 Mt/a case gave the best match between mine capacity and plant. The lower production “tail” after year 2020 represents the Hassai Underground mine production after the Hadal Awatib pit is exhausted. This case was used for the economic model to test the strength of the FINAL – Rev 0 – 22 Oct 2010 AMEC Page 267 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report proposal. However, AMC is confident that additional open pit resources will be defined at Hadal Awatib and other pits filling the capacity after 2020. The physical inputs and cashflow model for the VMS Phase are shown in Table 18.76. Table 18.76 Production and Cashflow Projections – VMS Phase 2,013 2,014 2,015 2,016 2,017 2,018 2,019 2,020 2,021 2,022 2,023 2,024 2,025 Tot / Avg 69 17,258 0 959 24,859 0 4,485 16,520 4,300 204 4,998 9,659 5,000 192 4,957 6,629 5,000 186 4,972 2,728 5,000 230 4,118 1,884 5,000 216 1,143 0 1,401 70 1,249 0 1,249 62 832 0 832 41 744 0 744 38 514 0 514 27 323 0 323 17 29,362 79,536 29,362 1,281 47,462 51 62,498 48 61,776 47 62,692 58 62,346 54 22,827 22,816 12,803 10,152 18 16 10 9 7,509 7 5,014 4 377,896 322 22,827 22,816 12,803 10,152 17 15 10 9 7,509 6 5,014 4 377,896 310 950 4,800 950 4,800 950 4,800 Physical Data Tonnes Tonnes Tonnes Tonnes of ore mined, '000 of waste mined, '000 of ore Milled, '000 of Concentrate Produced, ’000 t Metal in Concentrate Gold from concentrate, oz Copper from concentrate, '000 t Sales Data Metal Sales, t and oz Gold from concentrate, oz Copper from concentrate, '000 t Metal Prices, USD Gold, USD/oz Copper, USD/t Metal Revenue, '000 USD Gold from concentrate Copper from concentrate 0 0 0 0 47,462 49 62,498 46 61,776 45 62,692 56 62,346 52 950 4,800 950 4,800 950 4,800 950 4,800 950 4,800 950 4,800 950 4,800 0 0 0 45,089 59,373 58,687 59,557 59,229 0 236,813 222,895 215,821 266,880 250,698 950 4,800 950 4,800 950 4,800 950 4,800 21,686 21,675 12,163 9,645 7,133 4,763 359,001 81,940 71,825 47,742 43,972 31,125 20,015 1,489,725 P&L Statement (In '000 USD) Revenues 0 Cost of Sales Mining Costs Underground Mining Total - Underground Underground mining ($/t) 9,241 Open Pit Mining Total - Open Pit Open pit mining ($/t) Total Mining Costs 0 281,902 282,268 274,508 326,437 309,927 103,625 93,500 59,905 53,616 38,258 24,778 1,848,726 9,241 181,759 156,967 157,948 159,227 227,044 89,117 66,027 46,171 41,383 33,250 26,359 1,203,734 3,330 48 27,631 42 56,999 31 31,224 25 29,842 25 31,966 24 31,464 24 28,244 30,973 19,931 16,227 11,291 25 25 24 22 22 7,143 22 326,265 26 28,645 1.66 31,975 41,921 1.67 69,552 39,072 2.04 96,071 35,176 2.62 66,400 35,285 3.40 65,126 26,643 4.19 58,608 23,539 5.01 55,003 1,054 0 0 0 0 0.00 0.00 0.00 0.00 0.00 29,298 30,973 19,931 16,227 11,291 0 0.00 7,143 231,334 0 557,600 43,865 10.20 45,899 9.18 45,667 9.13 47,336 9.47 46,807 9.36 15,212 14,329 10.86 11.48 Milling costs Milling ($/t) 9,478 11.39 9,066 12.19 7,855 15.28 6,839 21.20 292,354 9.96 Concentrate Treatment & Shipping Cost 0 0 20,369 19,172 18,563 22,955 21,563 7,048 6,178 4,106 3,782 2,677 1,722 128,134 Refining Charge 0 0 5,631 5,300 5,131 6,345 5,961 1,948 1,708 1,135 1,045 740 476 35,420 Government Royalties Gold Copper 0 0 0 0 0 0 10,543 2,254 8,288 10,770 2,969 7,801 10,488 2,934 7,554 12,319 2,978 9,341 11,736 2,961 8,774 3,952 1,084 2,868 3,598 1,084 2,514 2,279 608 1,671 2,021 482 1,539 1,446 357 1,089 939 238 701 70,090 17,950 52,140 -31,975 -69,552 -3,961 186 3,730 2,422 76,733 22,417 0 0 0 0 0 0 9,241 9,241 9,241 9,241 9,241 9,241 9,241 9,241 9,241 9,241 9,241 9,241 9,241 120,136 Stock Variation G&A and Other Costs Gross Margin -9,241 82,883 14,509 27,473 13,735 12,233 5,007 -1,581 644,992 320 4,546 51,466 59,103 58,919 58,942 54,681 14,968 14,401 9,602 8,577 5,930 3,721 345,177 -9,561 -13,787 48,677 66,198 57,641 108,268 28,202 -459 13,072 -923 -5,302 299,814 0 0 4,868 6,620 5,764 10,827 2,820 -9,561 -13,787 43,809 59,578 51,877 97,442 25,382 Cash flows from operating activities -41,216 -78,793 Cash flows from investing activities Mine Development - Capital Costs Underground Mining Open Pit Mining Total Mining 133,809 185,623 Depreciation Mine Operating Income Income tax Net Earnings (Loss) -9,241 100,143 125,301 116,560 167,210 4,132 3,656 1,307 413 366 -459 11,765 3,719 3,290 0 0 0 32,985 -923 -5,302 266,830 Cash Flow Statement (In '000 USD) Process Plant Plant Infrastructure Area Infrastructure Regional Infrastructure & Miscellaneous Indirect Cost Sustaining Capex Accuracy Provision Cash flows from financing activities Free Cash Flow to Equity NPV @ 5% Discounting IRR 19,734 53,169 72,903 25,016 18,493 43,509 23,458 11,650 3,870 3,575 11,516 0 6,837 54,736 27,184 9,030 8,342 26,870 0 15,953 0 0 -175,025 -264,417 5,007 -1,581 612,007 13,955 91,315 118,867 114,526 158,806 156,796 3,855 7,552 6,914 10,301 4,011 4,011 4,011 3,426 2,255 1,385 381,108 11,757 1,363 13,120 0 2,435 2,435 500 1,262 1,762 0 2,903 2,903 0 4,925 4,925 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 600 0 600 57,607 84,550 142,157 835 1,420 5,790 4,011 5,376 0 0 0 0 0 77,359 115,012 106,975 151,892 146,495 36,926 26,166 13,321 11,867 4,011 4,011 4,011 3,426 0 0 0 0 32,915 22,155 9,310 8,442 78,194 38,834 12,900 11,917 38,386 35,930 22,790 2,255 785 0 0 0 2,752 -2,966 230,899 122,712 11% FINAL – Rev 0 – 22 Oct 2010 AMEC Page 268 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Project economics are positive due primarily to the relatively low operating cost ($1.24/lb of copper, including gold credits) in the first five years of VMS concentrator operation. Production for the VMS concentrator Phase is all related to copper concentrate production and sales. The treatment and sale terms are based on recent average industry performance. Currently, there is a shortage of concentrate, and therefore more favorable concentrate sale terms than shown in the model below. Given the high capital modeled with a mis-match of mine and mill capacity, AMC is encouraged that this phase provides an IRR of 11%. Sensitivities of NPV (5% discount rate) and IRR for this phase to metal prices are shown in Table 18.77. Table 18.77 Gold Price, $/oz Sensitivity of VMS Phase Economics to Metal Price Changes NPV 122,712 750 850 950 1150 1250 1350 4,000 -113,950 -86,305 -58,768 -3,787 23,254 50,170 4,800 68,948 95,830 122,712 176,156 202,832 229,509 Copper Price, $/t 5,600 6,400 248,229 426,998 274,906 453,647 301,582 480,296 354,936 533,593 381,612 560,242 408,289 586,890 7,200 605,590 632,238 658,887 712,184 738,833 765,482 8,000 784,181 810,830 837,478 890,776 917,424 944,073 4,000 -4% -1% 1% 4% 6% 7% 4,800 8% 10% 11% 14% 15% 17% Copper Price, $/t 5,600 6,400 17% 25% 19% 26% 20% 27% 22% 29% 23% 30% 24% 31% 7,200 32% 33% 34% 36% 37% 38% 8,000 39% 40% 41% 43% 43% 44% Gold Price, $/oz IRR 18.15.2.4 0 750 850 950 1150 1250 1350 Combined CIL and VMS Phase expansions The combined expansion phases are shown as incremental to the existing operations, withboth phases operating in parallel from 2015 to give the production profile shown in Figure 18.28. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 269 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 18.28 Metal Production Profile – Combined CIL and VMS Concentrator Phases 500,000 450,000 VMS Copper as Gold eqv. 400,000 VMS Concentrate Heap Leach Residue Gold Production, oz 350,000 Acidic non‐washable ore 300,000 Acidic washable ore SBR ore 250,000 Quartz ore 200,000 150,000 100,000 50,000 ‐ 2010 2011 2012 2013 2014 2015 2016 2017 2018 2019 2020 2021 2022 2023 2024 2025 Cash-flow and financial analysis for the combined expansion phases is shown in Table 18.78. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 270 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.78 Production and Cashflow Outcomes – CIL+VMS Phases 2,010 2,011 2,012 2,013 2,014 2,015 2,016 2,017 2,018 2,019 2,020 2,021 2,022 2,023 2,024 2,025 Tot / Avg 0 0 0 0 0 0 0 0 0 0 47,462 49 62,498 46 61,776 45 62,692 56 62,346 52 22,827 22,816 12,803 10,152 17 15 10 9 7,509 6 5,014 4 377,896 310 0 0 0 209,801 165,133 150,687 144,230 109,566 31,828 0 0 0 0 0 0 0 811,245 950 4,800 950 4,800 950 4,800 950 4,800 950 4,800 950 4,800 950 4,800 950 4,800 950 4,800 950 4,800 950 4,800 Revenues CIP VMS 0 0 0 0 0 199,311 156,876 425,054 419,287 378,596 356,674 309,927 103,625 93,500 59,905 53,616 38,258 24,778 2,619,408 0 199,311 156,876 143,152 137,018 104,087 30,237 0 0 0 0 0 0 0 770,683 0 0 281,902 282,268 274,508 326,437 309,927 103,625 93,500 59,905 53,616 38,258 24,778 1,848,726 Cost of Sales 0 0 0 93,425 84,593 247,866 213,184 207,054 173,492 227,044 89,117 66,027 46,171 41,383 33,250 26,359 1,548,965 Mining & Milling Costs CIP VMS 0 0 0 54,183 31,975 61,912 56,087 46,626 41,820 12,149 0 69,552 139,936 112,299 110,794 105,945 101,810 0 0 0 0 0 0 44,510 45,303 29,409 25,293 19,146 13,982 Government Royalties CIP VMS 0 0 0 13,952 0 10,981 0 10,021 10,543 9,591 10,770 7,286 10,488 2,117 12,319 0 11,736 0 3,952 0 3,598 0 2,279 0 2,021 0 1,446 0 939 53,948 70,090 Stock Variation CIP VMS 0 0 0 16,049 -31,975 2,459 -69,552 0 -3,961 0 186 0 3,730 0 2,422 0 76,733 0 22,417 0 0 0 0 0 0 0 0 0 0 18,507 0 VMS Concentrate Transport and Treatment Cost VMS Refining Charge 0 0 0 0 0 0 0 0 0 0 20,369 5,631 19,172 5,300 18,563 5,131 22,955 6,345 21,563 5,961 7,048 1,948 6,178 1,708 4,106 1,135 3,782 1,045 2,677 740 1,722 476 128,134 35,420 G&A and Other Costs 0 0 0 9,241 9,241 9,241 9,241 9,241 9,241 9,241 9,241 9,241 9,241 9,241 9,241 9,241 120,136 0 0 72,283 177,188 206,103 171,542 183,182 82,883 Physical Data Metal Sales, t and oz VMS Gold from concentrate, oz Copper from concentrate, '000 t CIP Gold, oz P&L Statement (In '000 USD) Metal Prices Gold, USD/oz Copper, USD/t Gross Margin Evaluation / Eploration expenses Depreciation CIP VMS Mine Operating Income Income tax Net Earnings (Loss) 3,000 15,000 0 0 950 4,800 950 4,800 0 105,886 0 0 950 4,800 0 950 4,800 0 950 4,800 0 0 0 0 0 36,010 320 36,010 4,546 36,010 51,466 -3,000 -15,000 0 69,556 31,727 89,711 110,989 0 36,010 59,103 0 0 6,956 3,173 8,971 11,099 -3,000 -15,000 0 62,600 28,554 80,740 99,890 -3,000 -15,000 0 83,004 950 4,800 0 36,010 58,919 0 0 14,509 27,473 13,735 12,233 0 0 0 0 10,461 58,942 0 54,681 0 0 14,968 14,401 0 9,602 0 8,577 76,612 113,779 28,202 -459 13,072 4,132 3,656 11,378 2,820 68,951 102,401 7,661 25,382 1,307 413 366 -459 11,765 0 3,719 3,290 272,776 849,954 5,007 -1,581 1,070,443 0 0 5,930 0 18,000 0 3,721 190,512 345,177 -923 -5,302 516,753 0 54,144 -923 -5,302 0 462,609 Cash Flow Statement (In '000 USD) Cash flows from operating activities Cash flows from investing activities CIP VMS Cash flows from financing activities Free Cash Flow to Equity NPV @ 5% Discounting IRR 2,017 164,256 195,190 167,611 174,226 156,796 3,000 54,725 130,912 133,809 185,623 0 54,725 130,912 0 0 3,000 0 0 133,809 185,623 0 0 0 0 0 36,926 26,166 13,321 11,867 5,007 -1,581 1,016,807 13,955 0 13,955 3,855 0 3,855 7,552 0 7,552 6,914 0 6,914 10,301 0 10,301 4,011 0 4,011 4,011 0 4,011 4,011 0 4,011 3,426 0 3,426 2,255 0 2,255 1,385 0 1,385 569,745 185,637 384,108 0 0 0 0 0 0 0 0 0 0 0 0 32,915 22,155 9,310 8,442 2,752 -2,966 447,061 -6,000 -69,725 -130,912 -50,805 -183,607 150,301 191,334 160,059 167,313 146,495 238,666 17% FINAL – Rev 0 – 22 Oct 2010 AMEC Page 271 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Production from the current operations for 2010 to 2012 is not included in the above analysis, although free cash from this period will go toward the required cash for the CIL Phase development. In addition, CIL cash-flow in the model reduces the required cash for the VMS concentrator phase. All modelling is based on identified resources (including Inferred Resources),and has excluded any upside potential. However, AMC is confident further potential exists at the project for feed for both the CIL and VMS concentrator phases. Other oxide pits that bottom in sulfides are thought to provide a means to extend cash-flow beyond 2019. Also optimisation of the mine schedules in future studies will improve the cash-flow profile for the final years of operation, improving NPV. 18.16 PROJECT IMPLEMENTATION At this stage of the project, the implementation document prepared by AMC is intended as a guide only, to allow an understanding of typical project implementation timelines and strategy. 18.16.1 Project Schedule A high level schedule for the project is outlined in Figure 18.29, and shows the construction of a 3 Mt/a CIL plant preceding the development of a 5 Mt/a VMS concentrator. Figure 18.29 Hassai Mine Envisaged business Plan - Summary Project Schedule 2010 2011 2012 2013 2014 2015 2016 2017 2018 2019 2020 Hassai CIP Project Current Heap Leach Operation CIP PFS/DFS CIP EPCM CIP - Commission and Operate Ariab VMS Project Scoping Study Drilling and Testwork Program Prefeasibility Study Definitive Feasibility Study EPCM Plant - Commission and Operate 18.16.2 Project Implementation Summary Construction of the process plant and associated facilities, including supporting infrastructure, is recommended to be executed on an engineering, procurement and construction management (EPCM) basis. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 272 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 18.16.3 Project Implementation Scope The Project scope of work will include the provision of facilities for mining, process plant, utilities and services, waste disposal and the associated infrastructure to support the construction work and ongoing operations. Responsibility for each element of the project will be assigned by AMC. It is envisaged that the division of responsibilities for the project implementation phase will be similar to that listed below. AMC’s scope will include: • Finance, governmental approvals, environmental approvals and licences • Land purchases, security, medical, taxes and duties • Mining planning and operations • Engagement of specialist consultants and contractors for mining, process waste disposal and other specialist scopes. The EPCM contractor’s scope will include: • Process plants • On site infrastructure including roads • Regional road upgrades • Assistance with management of off-site infrastructure works • Management of specialist consultants and contractors as required. 18.16.4 HSEC The management of health, safety and the protection of the environment and community (HSEC) are major considerations when executing projects. A project management plan will be developed based on a platform of risk assessment and subsequent risk elimination and mitigation. These factors are to be included at every stage from the study, design, construction through to commissioning and handover of successful projects. 18.16.5 Long Lead Items Long lead items will be identified and procured as a matter of priority. For the expected flowsheets, crushers, mills and flotation cells are expected to be key long lead items, and can significantly impact the critical path. Table 18.79 gives an indication of the lead times on major process plant equipment. The ball mills are clearly the long lead items that will impact on the project critical path, while other equipment items such as the crusher, flotation cells and filters will need to be procured very early in the engineering phase of the project. Suppliers will be approached during the definitive feasibility study FS phase of the project to establish the requirements for securing a purchase option and thereby initiating the purchase process in time to meet the project schedule. In addition, the options for second hand equipment will be sought during the FS. Use of second-hand equipment may provide some significant schedule and cost reductions. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 273 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 18.79 Current Equipment Lead Times Equipment Lead Time Crushers 20 weeks Ball mill 40-50 weeks Ball mill (sourced from China) Flotation cells 40 weeks approximately 26-35 weeks (depending on vendor) Thickener 30 weeks (depending on vendor) Pressure filter 9 working months approximately 18.16.6 Logistics The project procurement philosophy is based on sourcing the bulk of the equipment and materials from within Sudan (if possible) and other suppliers in Africa. However, most major equipment items are expected to be sourced internationally. Where overseas shipping is required, an experienced freight and logistics company will be utilised to manage the logistics. Once the goods are received and cleared through a port, management and control will shift to the designated transport company. Transportation of over-dimension loads will require an appointed third party specialist to conduct load surveys and movement supervision. 18.16.7 Training The EPCM contractor may assist AMC with the training of operations and maintenance personnel and assist with the development of training packages for the local workforce. This will ensure a smooth transition from construction to operations. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 274 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 19. INTERPRETATION AND CONCLUSIONS Scoping studies have been completed into the possible development of a 3 Mt/a CIL circuit to reprocess heap leach tailings and any remaining oxide gold resources, and a 5 Mt/a flotation circuit to treat VMS mineralisation. Resource modelling and mining studies have been undertaken to investigate extraction methods and develop mining schedules to supply feed to these plants. A preliminary geotechnical investigation has been undertaken to support the proposed mining methods and mine designs for the VMS deposits which have not previously been mined. A high-level environmental review indicates that acceptable environmental outcomes should be achievable, assuming standard engineering design and operating practices are employed. The success of the proposed expansion projects rely on two key pieces of infrastructure, namely a 77 km long power line connecting to the National Grid and a 165 km pipeline bringing water from the Nile River. Discussions have commenced with the relevant authorities regarding access to power and water. Capital and operating costs have been developed for the mines, both process plants and related infrastructure, with the assumption that the power line and water pipeline are funded as part of the CIL project. A project schedule has been developed showing production starting in 2013 for the CIL plant and 2015 for the VMS concentrator. This schedule allows for completion of feasibility studies followed by plant design, the delayed start to the VMS concentrator reflecting the less advanced status of this phase in terms of resource definition, geotechnical studies, mine design, process testwork and mine development. Financial modelling indicates that both the CIL and VMS phases of the expansion project are economically feasible, with NPVs of $149.8 M and $122.7 M, respectively. Base case metal prices were $950/oz for gold and $2.19/lb for copper. However, it must be noted that the bulk of the resources contributing to the VMS mine schedule are classified as Inferred. Consequently, there is a high degree of uncertainty in these resources, and their use in economic modelling is not recommended under NI 43-101. An exemption has been provided by the regulatory authorities, allowing use of Inferred resources in this instance, but the Inferred Resources are too speculative geologically to have economic considerations applied to them that would enable them to be categorised as Mineral Reserves. It cannot be assumed that all or any part of the Inferred Mineral Resource will be upgraded to an Indicated or Measured Mineral Resource as a result of continued exploration. The financial outcomes are particularly sensitive to metal prices: a 10% increase in either gold or copper price increases overall NPV by approximately $80-90 M. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 275 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Plant throughput is at full capacity for only five years in both cases, and significant economic upside exists if additional reserves can be located. VMS mineralisation is known to exist at the base of six oxide gold pits, of which only two have been drilled sufficiently to allow resources to be modelled for use in this study. Of these, the resources at Hassai South have been modelled using large blocks with partial mineralisation estimated within these blocks. In order to undertake underground mining studies, the mineralisation has been regularised and the associated reduction in grade has a significant impact on project economics. It is believed that improvements in the resource estimation/modelling of the Hassai South underground and Hadal Awatib open pit deposits would assist in more accurate spatial definition of the mineralisation and mining-related dilution, and in turn may have the effect of increasing the schedule grades. It should also be noted, however, that there will likely be a drop in the overall mining inventory tonnes, as contained metal would not be affected. Additional resources are expected to be identified at the other known VMS locations, as well as from testing the numerous other electrical conductors identified in the district, with the potential to allow full production to be maintained for several more years. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 276 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 20. ADDITIONAL REQUIREMENTS FOR TECHNICAL REPORTS ON DEVELOPMENT PROPERTIES AND PRODUCTION PROPERTIES 20.1 BACKGROUND The AMC mining operation started in 1991. Since then, more than 2.3 Moz of gold have been produced from 12 pits distributed over an area of 800 km². The AMC licences essentially cover the entire mining district, which consists of a Proterozoic greenstone belt hosting gold-enriched VMS deposits and orogenic gold in quartz structures. All deposits discovered up until now were discovered by surface work. Until 2004, only SBR-type deposits (ie. supergene enrichment of gold-enriched VMS deposits) had been mined and processed. A new grinding circuit was introduced in 2007 to treat gold more effectively from quartz vein mineralisation, which now forms the bulk of the feed. Figure 20.1 shows the annual tonnage of ore and waste mined, Figure 20.2 the average mine and plant determined head grade of cyanidable gold, and Figure 20.3 the annual gold production. Figure 20.1 Annual Mining Tonnages (Ore and Waste) 18,000,000 16,000,000 14,000,000 12,000,000 10,000,000 t ORE MINED 8,000,000 WASTE 6,000,000 4,000,000 2,000,000 0 1991 + 1992 1993 1994 1995 1996 1997 1998 1999 2000 2001 2002 2003 2004 2005 2006 2007 2008 2009 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 277 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 20.2 Average Head Grade, Mine and Plant 25 20 15 g/t plant mine 10 5 0 1991 + 1993 1992 1994 1995 1996 1997 1998 1999 2000 2001 2002 2003 2004 2005 2006 2005 2006 2007 2008 2009 Figure 20.3 Annual Gold Production (kg) 6,000.0 5,000.0 4,000.0 kg 3,000.0 2,000.0 1,000.0 0.0 1991 + 19921993 1994 1995 1996 1997 1998 1999 2000 2001 2002 2003 2004 2007 2008 2009 Summary statistics of production activity since start-up in 1992 are indicated in Table 20.1. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 278 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Mining in the Ariab Mining District has been by open pit. The different types of ores mined have comprised SBR, sulphated SBR frequently encountered at the bottom of the supergene enrichment zone and close to the underlying VMS, and quartz ore from the Kamoeb deposit. All material is processed in a crushing plant, the quartz material undergoing additional milling, before being mixed with cement in an agglomerator and stacked on heaps. Gold is recovered by conventional heap leaching, due to constraints in the availability of large quantities of water. Despite declining head grades, annual gold output remained at 4-6 t since 1996 through a steady increase in crushing and processing capacity, but decreased to approximately 2 t by 2009. Gold recovery has averaged 80%, but has decreased slightly, particularly since 2003. Table 20.1 AMC Production History Year Waste tonnes Ore Mined tonnes Ore Processed Cyan. Au tonnes (g/t) 1991/2 Gold Recovered Cyan. Au Au Recovery (g/t) (Kg) (%) 834 630 113 385 11.16 97 240 12.18 982.0 83 1993 2 594 550 135 607 15.50 127 056 15.53 1 862.7 94 1994 3 116 970 271 016 13.28 168 435 18.16 2 679.6 88 1995 4 965 530 277 967 15.05 203 349 19.92 3 750.4 93 1996 9 188 507 432 794 13.95 383 910 15.15 4 571.2 77 1997 11 170 830 574 994 10.32 432 403 12.43 4 556.9 85 1998 11 045 460 515 887 12.03 651 133 10.92 5 671.0 80 1999 13 305 145 583 401 10.62 702 776 10.30 5 565.7 77 2000 14 522 940 519 908 11.23 763 515 8.80 5 772.9 86 2001 15 127 500 1 046 160 8.14 754 034 8.49 5 415.9 85 2002 15 179 000 993 450 5.85 800 258 8.21 5 263.9 80 2003 14 606 541 1 034 215 7.21 920 000 6.99 5 172.9 80 2004 13 117 600 851 102 6.04 915 864 6.22 4 281.0 75 2005 9 396 000 1 530 772 5.78 1 038 793 5.78 4 738.6 79 2006 9 266 000 1 017 327 5.24 937 467 4.43 3 156.3 76 2007 5 746 410 541 747 4.58 888 621 4.06 2 703.0 75 2008 3 703 000 458 776 4.07 809 275 4.07 2 276.0 69 2009 2 999 405 177 367 3.11 725 303 3.78 1 921.8 70 Total 15 9886018 11 075 875 5.70 11 319 432 7.73 70 342.0 80 20.2 MINING 20.2.1 Overview A flow sheet of the mining operation is presented in Figure 20.4. The fleet of mining equipment needed to run the mine (see mobile equipment list in Table 20.2) requires an integrated 200 km road system that links all the pits with the central processing plant. Formerly each pit had its own workshop, but with concentration of mining activity on smaller and shorter life quarries, only mobile lubrication and light maintenance are carried-out in the pits, all heavy mechanical repair work being carried out in the central workshop at Hassai camp site. Actively mined open pits at the end of 2009 are Hassai North, Hadal Awatib Link and Kamoeb. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 279 The Hassai Mine Envisaged Business Pla NI 43-101 Preliminary Assessment Repo Figure 20.4 Diagrammatic Representation of AMC’s Mining Flow Sheet FINAL – Rev 0 – 22 Oct 2010 AMEC Page 2 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 20.2 AMC Mobile Equipment Fleet as at 31/12/2009 Equipment Number Make Mine 60-tonne dump truck 11 Caterpillar 775 D 11 40-tonne dump truck 5 Caterpillar 771 C, D 5 Plant 20-tonne truck 3 TATRA 3 20-tonne trucks 12 rented locally, different make 12 Excavator 2 Liebherr R984 B 2 Excavator 4 Liebherr R974 B 4 Excavator 1 Caterpilar 385C 1 Excavator 3 Caterpillar 330 L, BL 2 Wheel loader 2 Caterpillar 428 B, C 1 1 Wheel loader 4 Caterpillar 966 F, G 1 3 Wheel loader 1 Caterpillar 988 G 1 Wheel loader 3 Liebher 530 Drill 1 Atlas Copco F7 1 Drill 2 Sandvik DP 1500 2 Drill 3 Tamrock Tam 700 3 Bull dozer 3 Caterpillar D8 1 Bull dozer 4 Caterpillar D10 4 Motor Graders 4 Caterpillar 14H 3 Motor Graders 1 Caterpillar 16H 1 20.2.2 Geology 1 3 1 1 1 Geotechnical Evaluation Given the cost implications of the high strip ratio in most of the AMC mining operations, geotechnical evaluations of pit slopes are conducted periodically by external consultants for every major open pit. The mine staff subsequently follows up on their recommendations. Two stability incidents – due to unnoticed faults – have been recorded during operations. The last one was on the southern wall at the western end of the Hassai South open pit. The very dry climate facilitates rock cohesion, but does not prevent occasional breaches due to dilatation, in particular in winter when day and night temperatures show high contrasts. 20.2.3 Grade Control and Mining Delineating the ore to be mined within each pit is the responsibility of the grade control department. Drilling plans for sampling are designed for each bench (presently 2.5 m high) with a drill spacing along and across the mineralisation adapted to the size and shape of each deposit (for instance 5x1.5 m in Hassaï) as given by exploration and other pre-mining data. Grade control drilling is focused on the mineralisation, but extends into barren wallrocks. On each bench plan, a series of control points spaced 20-30 m apart are surveyed by the survey department, on which the sampling grid is laid out by the grade control geologists. Drilling and sampling is conducted by the blast drill crew. After assaying at the mine laboratory, the grade control geologists delineate the ore and waste limits using SERMINE software. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 281 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report The ore-waste boundary is first placed half way between mineralised and barren sampling holes, then refined and smoothed using all locally available geological data. Ore volumes are defined by either assigning a unit volume to each sample, or having the outlines digitised. Densities are derived from exploration data at every 10 to 25 m. Ore block grades are calculated by averaging sample data within the block, and ore tonnages by applying an average density to in situ ore volumes. The ore delineation plans (see principle on Figure 20.5), on which low grade, medium grade and high grade ore limits are assigned different colours, and acidic ores are identified, are handed over to the mine captains and to the surveyors, for physical marking out with coloured rocks on the pit floor before blast hole drilling and after blasting (ore is generally blasted and excavated before waste). Figure 20.5 Illustration of In Situ Grade Control (Figures are Average Gold Grades Measured on a 2.5 m Sub-bench) Such plans are also used to forecast mine production, as well as ore being stockpiled close to the mines before shipment to the plant. These tonnages are called “mined” or “geological”. Mine production, including new additions to the ore stockpiles as well as direct shipments to the plant, is calculated (on an undiluted basis) as follows: • From the grade control sampling map, the ore/waste boundary is drawn according to the grade of the sample holes (at a grid of 5x2.5 m by 2.5 m depth, depending on the deposit). • Each area is being classified according to grade: − HGO : high grade ore, average grade > 10 g/t − MGO : medium grade ore, average grade between 5 and 10 g/t − GLGO : good low grade, average grade between 4 and 5 g/t − BLGO : bad low grade, average between 1.5 and 4 g/t. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 282 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Tonnage is calculated by the number of loaded dumpers. Grade is the arithmetic average grade of each area. In addition, sulphate-bearing SBR material, which requires washing or even thermal treatment before being blended with other material, is identified. In general, there is little, if any, high grade ore left in the stockpile since this is immediately shipped, together with medium grade ore, to the plant. In the field, each heap is identified by a sign, indicating origin, ore class and average grade, this being readily traceable in the database. 20.2.3.1 Geology-Mine Reconciliation Differences between initial reserve estimates for each 10 m bench are compared to production data as calculated from grade control maps, each time a given 10 m bench has been mined out. Table 20.3 illustrates the results of the comparison between geological (reserve) and production estimates for depleted deposits. Table 20.3 Geology Mine Reconciliation for Depleted Deposits Exploration Estimate Depleted Deposit Au g/t Contained Au Pre-production Mining Mining v Exploration Estimate (%) Au g/t (kg) Baderuk 11.7 2 110 Contained Au Au (kg) 10.2 Contained Au (kg) 3 084 -13 46 Adassedakh 17.6 3 046 12.9 4 111 -27 35 Taladeirut 14.4 3 538 13.6 3 322 -6 -6 Oderuk 10.0 6 978 7.7 6 798 -23 -3 Baderuk North 5.1 132 4.2 120 -18 -9 Hadayamet 7.2 12 013 6.5 15 551 -10 29 Dim Dim 5 9.7 567 6.2 431 -36 -24 Hadal Awatib West, North 19.6 17 865 13.7 17 861 -30 0 Hadal Awatib East A+B 9.9 13 096 9.3 18 917 -6 44 Oderuk West 7.5 216 6.8 292 -9 35 Ganaet 6.8 342 6.3 674 -8 97 Weighted Total 13.0 59 771 10.1 70 195 -22 17 Results show that, historically, geological estimations of reserves based on exploration data from the SBR deposits tend to overestimate mined grade and underestimate the tonnage to an even greater degree, resulting in an overall conservative evaluation with respect to contained metal. Apart from estimation uncertainties in the original geological reserve estimate, such discrepancies are for a great part due to dilution. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 283 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 20.2.3.2 Dilution “Geological” ore tons are based on a strictly defined in situ volume multiplied by an estimated density. Mining processes, however, do not adhere strictly to the ore limits defined by the grade control department. Ore is usually excavated using an excavator with a 2.5 to 2.8 m wide bucket, a size that may not be adapted to the width of the mineralisation being mined. This can lead to excavating significantly more material than initially measured and planned, especially when the mineralisation is narrow, a more and more common situation in the pits AMC is now mining. When the mineralisation is thick, the effect is significantly lower than when the mineralisation is particularly thin. The percentage, ranging from 8-30%, varies from deposit to deposit, and in general is difficult to calculate. An additional factor of dilution is the softness of the ore/waste boundaries leading locally to sloughs. Dilution adds waste or sub-economic material to the ore mined and sent to the stockpile and subsequently processed by the plant. Except for errors in delineating the actual ore/waste boundaries, dilution always adds material, and this effect always goes in the same direction. Due to the different pits and ore types, dilution is not consistent in all operations, but averages approximately 15% on an annual basis. 20.3 BENEFICIATION PLANT 20.3.1 Overview Gold is extracted at the beneficiation plant through heap leaching. A circuit designed to treat SBR ore has been in use for the last 19 years and has been expanded several times by adding several crushers and hoppers. An additional “quartz” circuit was implemented in 2005 to handle the Kamoeb South ore through a ball mill, silo and screens adapted to ground (800 µm) material. This circuit has experienced some difficulty in handling the very fine mineralised clayey schist that is inevitably extracted along with the clean quartz vein material. The circuit is being fine-tuned (and was modified during 2007 with the addition of an SBR/Quartz ore mixer) and integrated into the SBR circuit. Both materials are then mixed with cement and lime (for acidic material) and fed into two agglomerators, where a cyanide solution is added as the moistening agent. Current crushing and milling activities run at about 2500 t/d. Individual heaps are 7 m high x 45x200 m. The annual throughput of ore being leached is 0.7 Mt (2009 figure). The heap is leached with sodium cyanide in three distinct cycles for a total duration of three months. The resulting pregnant solution is sent to an absorption-desorption (ADR) unit where the gold is captured on activated carbon. Metallurgical recoveries average approximately 72%. Gold smelting is performed on-site and the 25 kg doré bars (gold ±62%, silver ±30%; copper and iron impurities) are sent to a refiner. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 284 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 20.3.2 Process Flow Sheets Figure 20.6 and Figure 20.7 on the following pages describe the process flow sheet. SBR ores are mined and are sent to the SBR line where the material is sent to a series of crushers and screens to be crushed at <12.5 mm. Quartz material is sent to the quartz line through a series of two crushers and a screen to a silo (size of the feed <8 mm), a series of three Mogensen screens and a ball mill to be ground to <800 µm. The two types of ores are blended in an ore mixer, then cement and lime are added, and the product sent to one of two agglomerators where a cyanide solution is added. The agglomerated material is sent via land conveyors to a stacker and piled on the leach pads. The method used is the traditional heap leach process including one cycle with cyanide solution (0.35 g/L), one cycle with lower cyanide concentration (0.2 g/L) solution and one washing cycle with industrial water. All solutions are collected in ponds. The pregnant solution goes to the self-contained ADR unit, where the gold is stripped and won in an electrolytic cell. After smelting, the doré bars are temporarily kept in the gold room before shipment. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 285 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 20.6 AMC Gold Plant: Crushing/Milling Section Flow Sheet FINAL – Rev 0 – 22 Oct 2010 AMEC Page 286 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Figure 20.7 AMC Gold Plant: Leaching Section Flow Sheet FINAL – Rev 0 – 22 Oct 2010 AMEC Page 287 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 20.3.3 Reagents The description and quantities of the principal reagents used in the operation are listed in Table 20.4. Table 20.4 Reagents Used at the Hassai Heap Leach Plant Cement 10 kg/ t ore Lime 0-20 kg/t ore Cyanide 250 g/t ore Water 300 L/t ore Sulphated, acidic material of pH = 1.5 is found near the redox zone at all SBR deposits. It has been mainly stockpiled near the waste dumps in Hadal Awatib and in Hadayamet. The acid ore of Hadal Awatib is pre-treated by washing to dissolve the iron sulphates that impede the proper circulation of percolating fluids. The ore is generally spread on the waste dump in a 1.5 m thick layer and sprinkled with water in the amount of 0.8 to 1 m3/t. This static washing method is, however, unsuitable for the highly acidic ores of Hadayamet which are gradually fed in small quantities and with the addition of lime (to the order of 20 kg/t and more) to the batch of neutral ores (SBR and/or quartz). 20.3.4 Gold Reconciliation at the Plant Internal reconciliation at the plant is performed on a daily basis, with the processing of the data for the mass balance, the follow-up of the leaching recovery curve and the evaluation of the final recovery of the heap. The daily dry tonnage and computed grades for material going onto the heap are gathered as input data. Tonnage is measured by a belt conveyor balance, and corrected according to data from the weighbridge and for humidity, lime and cement added to the conveyor. Sampling is done in a bin chute leading to the agglomerator using an automatic sampler. At regular time intervals, a slit runs from one side of the bin to the other at a constant speed and collects 300-500 kg of material per day (d100 = 16 mm, d90 = 12.5 mm) in a minimum of two samples. Each sample passes through a laboratory cone crusher, and is then split and assayed. A gold balance is managed daily for each active heap, and provides an individualised recovery for each heap at the end of the day. Solutions percolating through any heap are assayed each day (pregnant and sterile solution) and the volume is measured. Doré gold and gold adsorbed on carbon are further reconciled with the gold in solution entering the ADR circuit. 20.3.5 Mine/Plant Gold Reconciliation Reconciliation of gold between the mine and the plant is not easy due to the blending of different sources of ore at the plant site. Ore transported to the plant is weighed on a weighbridge. The mine tonnage is then readjusted according to this weight. The ore is sampled in a systematic way during its passage to the plant, in order to calculate head grade. Mine production is then corrected when the exact quantity of gold is known, after its passage to the plant. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 288 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 20.4 TAILINGS AND WASTE MANAGEMENT The waste to ore ratio ranges from 7 to 39 depending on the pit. Most of the mined material is waste which amounts to more than 160 Mt in total. Waste dumps are created near every pit to store sterile material. Organic soil is not produced in the desert environment and is thus not included in the stockpiles. One of the waste dumps at Hadal Awatib acts as a blanket for washing sulphated ore. Tailings amount to approximately 11 Mt and are kept at the plant site. Earlier tailings were removed and stored in areas not far from the present active heaps. The historical metallurgical balance reveals that the average grade of the tailings may reach 1.57 g/t cyanidable Au. The company plans to retreat these tailings at the end of the mine life and is investigating suitable gold recovery processes. In all SBR pits, gold concentration drops in the massive sulphide near the redox zone, effectively defining the bottom of the ore zone. However, the few intersections into this sulphide zone show a potential for profitable base metal extraction given the current base metal prices. The reactivation of some of these pits might be worth considering following a proper assessment of this potential. 20.5 INFRASTRUCTURE 20.5.1 Buildings and Mine Camp The Hassai mine camp is approximately 3 km from the processing plant and accommodates about 600 personnel (expatriates and locals). It includes accommodation, dining halls, a bakery and local market, and recreational facilities. An on-site communication tower allows cellular phone communication through three mobile phone access providers and Internet access. Seventeen diesel generators (totalling 5470 kVa) supply electricity to the plant and facilities. 20.5.2 Other Offices The head office building in Khartoum houses approximately 40 personnel, servicing general management, financial control and local purchasing departments. AMC also has a small office in Port Sudan for seven personnel who are responsible for coordinating sea freight shipments, including the purchasing and transportation of supplies for Hassai (food, equipment, etc.). 20.5.3 Logistics Transportation from Port Sudan to the mine site is carried out by a combination of sub-contractors and company-owned trucks. The distance is approximately 200 km, and about 20 000 t of consumables are transported each year. Airfreight cargo service, is provided through Lufthansa, Emirates, Egypt Air and other through their scheduled flights. A Twin Otter airplane owned by AMC is used for limited personnel transportation and emergency purposes. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 289 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 20.5.4 Water Supply Water supply is a key issue given the desert climate at the mine area. Moreover, the basement geology in the Hassai region consists of granite and volcanic rocks and there is therefore no sizable underground aquifer. Total water consumption in 2009 was around 430 000 m³. The following steps have been implemented to deal with the water supply situation: • Fresh as well as saline water is sourced from a series of wells located at distances up to 100 km from the Hassai plant. • Basins (hafirs) protected by earth dams have been dug to store run-off rainwater; these basins now have a total capacity of over 340 000 m3. • Recycled sewage water is being used in the leach process since 1996. Significant rainfalls in recent years greatly increased the water reserves. AMC estimates that current water reserves (underground) are sufficient to sustain production for at least two more years without any further precipitation. Water supply sources are shown in Figure 20.8. Note: the water consumption per tonne of ore processed has increased over the years, reflecting the need for water to wash some acidic material, and also the increase of water consumption due to an increasing local population around the site. Figure 20.8 Water Sources 2009 Recycled water 2% Hafirs water 19% Recycled water 8% Fresh water wells 71% FINAL – Rev 0 – 22 Oct 2010 AMEC Page 290 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 20.6 SOCIAL PROGRAM The Red Sea Development Fund was formally established in 1998. The funds are mostly spent on the operating costs of the local Bir Ajam village, as well as targeted donations and capital projects in the Red Sea Hills region. They are presently administered by the Red Sea Hills region for sustainable development projects. Funding was $200 000 in 2004 and $250 000 in 2005 and the years after. AMC directly employs about 600 persons from the nearby region, and it is estimated that the company contributes to the welfare of all inhabitants in a area extending 120 km from the mine site. 20.7 FINANCIAL ANALYSIS Table 20.5 shows the analysis of the current AMC operations. Note the values are in US Dollars for 100% of AMC. The gold price assumption is $950/oz as used in other areas of this report. NPV and discounted cashflow are based on 5% discounting as used in other sections of this report. Table 20.5 Cashflow Analysis of Current Operation 2010 2011 2012 2013 Tot / Avg Physical Data Tonnes of Ore Mined Tonnes of waste Tonnes milled 474 695 643 272 537 740 300 000 1 955 706 5 694 566 4 669 373 3 299 665 2 000 000 15 663 604 2 615 808 751 376 631 835 650 598 582 000 Gold Grade (g/t) 4.22 4.96 3.75 6.92 4.88 Recovery (%) 70% 73% 74% 74% 73% 71 728 73 079 57 871 96 456 299 134 Gold Production (oz) Profit and Loss Statement (in '000 US$) Revenues 0 0 0 0 284 177 Cost of Sales 64 126 48 110 47 085 51 882 211 203 Mining and Milling Costs 40 605 34 276 30 117 26 534 131 532 16 978 14 036 9 440 5 658 46 113 Mining Costs Haulage Costs Milling cost G&A and Other Costs 710 969 833 465 2 977 22 917 19 271 19 843 20 411 82 442 23 521 13 834 16 969 25 348 79 671 Office / Administration 9 241 9 241 9 241 9 241 Government Royalties 4 770 4 860 3 848 6 414 19 892 Stock Variation 9 510 -267 3 879 9 693 22 815 Gross Margin Depreciation and Amortisation of capital assets Gross Margin Cominor Mine Operating Income Income tax Net Earnings (Loss) 4 016 21 314 7 892 39 751 72 974 13 198 11 098 11 428 10 223 45 946 80 426 158 795 1459 28 487 -9 101 10 643 -3 377 30 323 -1 365 1 596 -507 4 548 4 273 -9 101 9 046 -3 377 25 775 22 342 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 291 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Table 20.5 Cashflow Analysis of Current Operation 2010 2011 2012 2013 Tot / Avg Cash flows from operating activities 13 606 19 878 11 929 45 690 Cash flows from investing activities 9 066 1 880 35 000 0 45 946 Non-Plant Capital Expenditures 9 066 1 380 0 0 10 446 500 10 000 0 10 500 Cash Flow Statement (In '000 USD) Washing Plant for Acidic SBR Water line 25 000 Cash flows from financing activities Free Cash Flow to Equity NPV @ 5% Discounting 91 103 25 000 0 0 0 0 0 4 540 17 998 -23 071 45 690 45 157 35 684 While the current planned operations generate a substantial NPV @ 5% discounting, options to better utilise the water pipeline investment led to the proposed business plan including a CIL Plant that is a subject of this report. The current operation is robust in the current market environment with the NPV dropping to zero only below a gold price of $750/oz. Gold prices shown in Figure 20.9 are based on variances of 10% increments from -30% to +30%, and are not suggestive of gold price projections. In October 2010, the spot gold price exceeded $1350/oz. Figure 20.9 Sensitivity Analysis – Current Hassai Operation 100,000 80,000 Gold Price NPV @ 5%, USD, 000 Operating Cost (variance) ‐30% 60,000 ‐20% 40,000 +10 20,000 ‐10% +20 +30 0 665 ‐20,000 760 855 950 1045 Gold Price, USD/oz 1140 1235 Cash flow and NPV are less sensitive to operating cost variances than to gold price/gold production variances (Figure 20.9). Note that the NPV is zero above cost increases of 29%. Within the accuracy of the current operating cost estimates, the project is quite robust. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 292 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 21. RECOMMENDATIONS With very few exceptions, work to-date has been completed to a scoping level, and significant additional studies are required in all disciplines in order to confirm numerous assumptions and provide a sound basis for developing the CIL and/or VMS concentrator options to a definitive FS level. 21.1 CIL PHASE For the CIL phase, La Mancha has approved a budget of A$1.69 M in order to complete feasibility study work. The program is due for completion in the second quarter of 2011, and comprises: • Upgrading of Inferred resources of heap leach tailings to the Indicated category through development of an improved metallurgical balance. • Additional mining studies for both the open pit and stockpile and tailings reclaim activities, in order to refine the schedule, capital and operating costs. • Further metallurgical testwork on representative samples in order to define flowsheets and design criteria at a feasibility level. This will require collection of samples representative of the various domains, followed by comminution leaching and tailings testwork. • Tailings disposal: selection of the optimum disposal site is required, followed by TSF and tailings handling design. • Plant engineering to a FS level. • Infrastructure: confirmation of requirements, followed by FS-level design, specifically for the water pipeline, HT power line and accommodation village. • Environmental studies, and initiation of permitting as required, to support the final designs for mine, plant and infrastructure. • Project implementation: development of strategy and schedule for project development. • Capital and operating cost estimation for final FS designs. La Mancha intends to undertake the FS work in two stages, namely preliminary and final feasibility studies, and to this end has approved programs costed at A$0.907 M and A$0.781 M, respectively. Additional design of the water pipeline has been awarded to Sudanese for Construction and Oil Services at an estimated cost of US$ 250 000. 21.2 VMS PHASE The VMS phase is at a much earlier stage of development, with a poorer understanding of resource, mining, processing and cost parameters, which leads to a lower confidence in the economics of the concentrator phase. At this time, the highest priority is to expand and increase confidence in the resources, which will, inter alia allow for an improved understanding of mining options. Consequently, La Mancha has set aside a budget of $18 M for the VMS phase in order to undertake a drilling program of 100 000 m designed to: • Upgrade existing Inferred Resources of VMS material to Indicated and Measured status • Develop VMS resources beneath the exhausted gold open pit at Hadayamet, 30 km east of Hassai. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 293 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 22. REFERENCES 22.1 GEOLOGY AND RESOURCES Arethuse Geology, 2009. Hadal Awatib East VMS Cu-Au Resources Estimates. Independent report to AMC, October 2009. Barrie, C. T., and Hannington, M. D., 1999. Volcanic-associated Massive Sulphide Deposits: Processes and Examples in Modern and Ancient Settings: Introduction: In Reviews in Economic Geology Volume 8: Volcanic-Associated Massive Sulphide Deposits: Processes and Examples in Modern and Ancient Settings, Barrie, C. T., and Hannington, M. D., editors, p. 1-11. CSA Global (UK), 2010. Hassai Heap Leach Remnant Resources Report #R230.2010, dated 16 September 2010. Internal company report. Abu Fatima, M.A. (2006). Metallogenic genesis and geotechnic evaluation of the polymetallic massive sulphide and associated gold deposits at Arib-Arbaat Belt, Red Sea Hills, NE Sudan. PhD. Thesis, Geologie et Gestion des Ressources Minerales et Energetiques, Universite Henri Poincare, Nancy 1, France. La Mancha Resources Inc., January 2008. Hassai Mine, Sudan, NI 43-101 Technical Report. La Mancha Resources Inc., October 2009. Estimates, NI 43-101 Technical Report. Hassaï South Cu-Au VMS Deposit, Sudan, Resource La Mancha Resources Inc., December 2009. Hadal Awatib East Cu-Au VMS Deposit, Sudan, Resource Estimates, NI 43-101 Technical Report. Monthel J (2007). AMC Geology Exploration activity from 08/2005 to 03/2007. results and assessments. Final Report, Internal report AMC. 22.2 GEOTECHNICAL ANTEA 1999. Open Pits of Hadayamet and Hadal Auatib East (Sudan): Determination of the slopes of the pit walls. Internal report for Ariab Mining Company (AMC) December, 1999. Barton, N. R., Lien, R. and Lunde, J., 1974. Engineering Classification of Rock Masses for the Design of Tunnel Support, Rock Mech. 6(4), 189-239. Bieniawski, Z.T., 1989. Engineering rock mass classifications. New York: Wiley. Hoek, E., P.K. Kaiser and W.F. Bawden, 1995. Support of Underground Excavations in Hard Rock. Balkema. pp. 215. Hoek, E and Brown, E.T., 1980. Underground Excavations in Rock. Institute of Mining and Metallurgy, London. INTECSA-INARSA, 2002. Geological and Geotechnical Study and Final Slope Design at Hadayamet Open Pit. Internal report for Ariab Mining Company (AMC) May, 2002. Laubscher, D.H. 1990. A geomechanics classification system for the rating of rock mass in mining. J.S.Afr.Inst. Min. Metall. Vol. 90, no 10. pp 257-273. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 294 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 23. DATE AND SIGNATURE PAGE Project Name: Hassai Mine CIL Project Title: Hassai Mine CIL Project NI 43-101 Preliminary Assessment Report Location: Red Sea State, Sudan Effective Dates: Effective Date of Technical Report: 22 October 2010 Effective Date of Mineral Reserves: 31 December 2009 Effective Date of Mineral Resources: 31 August 2010 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 295 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report CERTIFICATE OF QUALIFIED PERSON – Adam Coulson I, Adam Coulson, do hereby certify that: • I am the Senior Rock Mechanics Engineer for AMEC, 160 Traders Boulevard East, Suite 110, Mississauga, Ontario, L4Z 3K7 Canada. • I graduated BEng, from Camborne School of Mines, UK in 1990, MSc. (Eng) from Queens University, Canada in 1996, and PhD from the University of Toronto, Canada in 2009. • I am a Professional Engineer and Member of the Canadian Institute of Mining and Metallurgy. • I have worked as an Engineer for a total of 20 years since my graduation from university. • I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43- 101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43 101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101. • I am responsible for the preparation of those parts of the Technical Report titled “The Hassai Mine Envisaged Business Plan (CIL Gold Plant and VMS Concentrator), Red Sea State, Sudan” Revision 0, dated 22 October, 2010 (the “Technical Report”) relating to geotechnical inputs to the VMS mining studies. • I visited the site for three days in March 2010. • I have read the relevant sections of the Technical Report and, as at the date of this certificate, to the best of my knowledge, information and belief, those sections of the Technical Report contain all scientific and technical information that is required to be disclosed to make that part of the Technical Report not misleading. • I am independent of the issuer applying all of the tests in section 1.4 of NI 43-101. • I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form. • I consent to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication by them for regulatory purposes, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report. Dated this 22nd Day of October, 2010 _______________________________ (Signed) Adam Coulson, P.Eng., PhD., CIMM Senior Rock Mechanics Engineer AMEC FINAL – Rev 0 – 22 Oct 2010 AMEC Page 296 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Certificate of Qualified Person - William F. Plyley I, William F. Plyley, do hereby certify that: • I am the Chief Operating Officer of La Mancha Resources Inc., with an office at Level 1, 12 St Georges Terrace, WA 6000, Australia. • I graduated from the University of Nevada, Reno (Mackay School of Mines) with a Metallurgical Engineering Degree, BSc. in 1982. • I am a member of The Australasian Institute of Mining and Metallurgy. • I have worked in the mining industry as a metallurgical engineer and in general management positions for over 35 years. • I have read the definition of “qualified person” set out in National Instrument 43-101 of the Canadian Securities Administrators (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I am a “qualified person” for the purposes of NI 43-101. • My last visit to the Hassai Mine site was on the 10th of October 2010 for a duration of 3 days. • I am responsible for supervising the compilation and overall preparation of the technical report entitled “The Hassai Mine Envisaged Business Plan (CIL Gold Plant and VMS Concentrator), Red Sea State, Sudan NI 43-101 Technical Report” (the “Technical Report”), dated 22 October 2010, relating to the Hassaï Mine in Sudan, in which La Mancha Resources Inc. owns an indirect 40% interest. In particular I am responsible for those parts of the Technical Report relating to adjacent properties, mineral processing for the heap leach operation, projected recoveries and the recovery schedule for the CIL circuit, market, taxes and royalties, G&A and other operating costs, project implementation, economic analysis and additional information for operating properties. • As Managing Director of La Mancha Resources Australia, I have been responsible for the Hassai mine since September of 2006. Consequently, I am not independent of the issuer applying thetest in section 1. of NI 43-101. • I have read the NI 43-101 and Form NI 43-101F1. The parts of the Technical Report to which I have contributed have been prepared in compliance with NI 43-101 and NI 43-101F1. • I have read the Technical Report, and, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading. Dated, this 22nd day of October 2010 (signed) William Plyley (MAusIMM) Chief Operating Officer La Mancha Resources Inc. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 297 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report CERTIFICATE OF QUALIFIED PERSON – Clayton Reeves I, Clayton Reeves, (Member of the South African Institute of Mining and Metallurgy), do hereby certify that: • • • • • • • • • • • • I am Principal Mine Engineer for CSA Global (UK), 2 Peel House, Barttelot road, Horsham RH12 1DE, United Kingdom. I graduated with a B.Sc. honours degree in Engineering (Mining) from the University of the Witwatersrand, South Africa in 1997. I am a Member of the South African Institute of Mining and Metallurgy. I have worked as a Mine Engineer for a total of 13 years since my graduation from university. I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43- 101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101. I am responsible for the preparation of those parts of the Technical Report titled “The Hassai Mine Envisaged Business Plan (CIL Gold Plant and VMS Concentrator), Red Sea State, Sudan” Revision 0, dated 22 October, 2010 (the “Technical Report”) relating to mine reserves, mine design, costing and scheduling of material to supply the continuing heap leach and proposed CIL plant operations, as contained in Section 18 and as summarised in Section 1 of the Technical Report. I have visited the Hassai Mine property several times between June 2009 and August 2010, for a cumulative total of over seven weeks on site. I have previously undertaken work for La Mancha Resources for the property that is the subject of this Technical Report. I have read the relevant sections of the Technical Report and, as at the date of this certificate, to the best of my knowledge, information and belief, those sections of the Technical Report contain all scientific and technical information that is required to be disclosed to make that part of the Technical Report not misleading. I am independent of the issuer applying all of the tests in section 1.4 of NI 43-101. I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form. I consent to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication by them for regulatory purposes, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report. Dated this 22nd Day of October, 2010 ________________________________ (Signed) Clayton Reeves, B.Sc., Member, South African Institute of Mining and Metallurgy Principal Mine Engineer CSA Global FINAL – Rev 0 – 22 Oct 2010 AMEC Page 298 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report CERTIFICATE OF QUALIFIED PERSON – Dean David I, Dean David, do hereby certify that: • I am a Process Consultant for AMEC, Level 14, 140 St Georges Terrace, Perth, Western Australia, 6000. • I graduated with a Bachelor of Applied Science (Metallurgy) from the South Australian Institute of technology (now the University of South Australia), Australia in 1982. • I am a Fellow of the Australasian Institute of Mining and Metallurgy, membership number 102351 • I have worked as a Process Engineer for a total of 28 years since my graduation from university. • I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant3 work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101. • I am responsible for the preparation of those parts of the Technical Report titled “The Hassai Mine Envisaged Business Plan (CIL Gold Plant and VMS Concentrator), Red Sea State, Sudan” Revision 0, dated 22 October, 2010 (the “Technical Report”) relating to the metallurgy, plant design and costing of the VMS concentrator. • I visited the site in March 2010 for 2 days. • I have read the relevant sections of the Technical Report and, as at the date of this certificate, to the best of my knowledge, information and belief, those sections of the Technical Report contain all scientific and technical information that is required to be disclosed to make that part of the Technical Report not misleading. • I am independent of the issuer applying all of the tests in section 1.4 of NI 43-101. • I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form. • I consent to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication by them for regulatory purposes, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report. Dated this 22nd Day of October, 2010 ________________________________ (Signed) Dean David, FAusIMM Process Consultant AMEC FINAL – Rev 0 – 22 Oct 2010 AMEC Page 299 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report CERTIFICATE OF QUALIFIED PERSON – Graeme Baker I, Graeme Baker, do hereby certify that: • I am Principal Mining Engineer for AMEC, Level 14, 140 St Georges Terrace, Perth, Western Australia, 6000. • I graduated with a Bachelor of Engineering Degree in Mining (Honours) from the University of Ballarat, Australia in (1998). • I am a Member of the Australasian Institute of Mining and Metallurgy. • I have worked as a Mining Engineer for a total of 12 years since my graduation from university. • I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43- 101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101. • I am responsible for the preparation of those parts of the Technical Report titled “The Hassai Mine Envisaged Business Plan (CIL Gold Plant and VMS Concentrator), Red Sea State, Sudan” Revision 0, dated 22 October, 2010 (the “Technical Report”) relating to open pit and underground mining of the Hadal Awatib and Hassai South VMS deposits, with the exclusion of geotechnical aspects which were provided by others. • I have not visited the mine site. • I have read the relevant sections of the Technical Report and, as at the date of this certificate, to the best of my knowledge, information and belief, those sections of the Technical Report contain all scientific and technical information that is required to be disclosed to make that part of the Technical Report not misleading. • I am independent of the issuer applying all of the tests in section 1.4 of NI 43-101. • I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form. • I consent to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication by them for regulatory purposes, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report. Dated this 22nd Day of October, 2010 ________________________________ (Signed) Graeme Baker BEng. (Mining), MAusIMM Principal Mining Engineer AMEC Minproc Limited FINAL – Rev 0 – 22 Oct 2010 AMEC Page 300 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report CERTIFICATE OF QUALIFIED PERSON – Ian Thomas I, Ian Thomas, do hereby certify that: • I am employed as Process Consultant by Sedgman Ltd, Level 4, 170 Burswood Road, Burswood, Western Australia, 6100. • I graduated with a Bachelor of Applied Science in Metallurgy from Bendigo College of Advanced Education (now Latrobe University), Australia in 1979. • I am a Member of the Australasian Institute of Mining and Metallurgy (#102227). • I have worked as a Metallurgist for a total of 31 years since my graduation from university. • I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43- 101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101. • I am responsible for the preparation of those parts of the Technical Report titled “The Hassai Mine Envisaged Business Plan (CIL Gold Plant and VMS Concentrator), Red Sea State, Sudan” Revision 0, dated 22 October, 2010 (the “Technical Report”) relating to processing of gold mineralisation in the proposed CIL plant, including metallurgy, plant and infrastructure design, preliminary capital and operating costs, with the exception of the processed grades and recoveries. • My last visit to the site and Khartoum was on the 6th of December 2007 for five days. • I have previously undertaken work for La Mancha Resources for the property that is the subject of this Technical Report. • I have read the relevant sections of the Technical Report and, as at the date of this certificate, to the best of my knowledge, information and belief, those sections of the Technical Report contain all scientific and technical information that is required to be disclosed to make that part of the Technical Report not misleading. • I am independent of the issuer applying all of the tests in Section 1.4 of NI 43-101. • I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form. • I consent to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication by them for regulatory purposes, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report. Dated this 22nd Day of October, 2010 ________________________________ (Signed) Ian Thomas, MAusIMM Process Consultant Sedgman Ltd FINAL – Rev 0 – 22 Oct 2010 AMEC Page 301 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report CERTIFICATE OF QUALIFIED PERSON - Jean-Jacques Kachrillo I, Jean-Jacques Kachrillo, do hereby certify that: • I am the Vice President of Exploration for La Mancha Resources Inc. I hold office at Tour Areva – 1, place de la Coupole, Paris La Defense and have been employed as such since 2007. • I graduated with a) Engineer Geologist of Ecole Nationale Superieure de Geologie et Prospection Miniere de Nancy (France) in 1974 b) 3rd cycle Thesis of Institut National Polytechnique de Nancy (“INPL”) in 1976. • I am a registered Geoscientist with Ordre des Geologues du Quebec. • I have worked as a geologist for a total of 33 years since my graduation from INPL. • I have read the National Instrument 43-101 of the Canadian Securities Administrators (“NI 43101”). I certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I am a “qualified person” for the purposes of NI 43-101 and that the Technical Report has been prepared in compliance with this Instrument. • My last visit to the Hassai Mine property was 22 August 2010 for a duration of 5 days. • I am responsible for the preparation of those parts of the Technical Report entitled “The Hassai Mine Envisaged Business Plan (CIL Gold Plant and VMS Concentrator) NI 43-101 Technical Report” (the “Technical Report”), dated 22 October 2010, relating to the Hassaï Mine in Sudan, in which La Mancha Resources Inc. owns an indirect 40% interest, concerning geology, mineralisation, exploration, drilling, sampling and sample preparation/assaying and sample security. • As an employee of La Mancha Resources Inc., I am not independent of the issuer as defined in section 1.4 of National Instrument 43-101. • I have read NI 43-101 and Form NI 43-101F1 and the Technical Report has been prepared in compliance with both. • As of the date hereof, and, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading. Dated, this 22nd day of October 2010 ________________________________ (Signed) Jean-Jacques Kachrillo (Geoscientist with Ordre des Geologues du Quebec) Vice President, Exploration La Mancha Resources Inc. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 302 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report CERTIFICATE OF QUALIFIED PERSON – Remi Bosc I, Remi Bosc, do hereby certify that: • I am an Independent Consulting Geologist with Arethuse Geology Sdn Bhd, and reside and maintain an office at 7236 Rotan Tunggal, 27600 Raub, Pahang, Malaysia. • I graduated from Ecole National Supérieure de Géologie de Nancy (France) as an ‘Ingénieur Géologue’ in 1994, and of Ecole des Mines de Paris (France), as ‘Mastere des grandes ecoles’ in Environmental Management in 2002. • I am registered as European Geologist with the European Federation of Geologists N°737. • I have worked as a geologist in mineral exploration and mining since my graduation in 1994. I have been assessing and reporting resources in Industrial Minerals for 3 years from 2003 to 2006. I have participated in gold, industrial minerals and base metals resources assessment in Malaysia, the Middle East and Sudan since 2006. • I have read the definition of “qualified person” set out in National Instrument 43-101 of the Canadian Securities Administrators (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfil the requirements to be a “qualified person”. • I am responsible for the preparation of those parts of the Technical Report titled “The Hassai Mine Envisaged Business Plan (CIL Gold Plant and VMS Concentrator), Red Sea State, Sudan” Revision 0, dated 22 October, 2010 (the “Technical Report”) relating to data verification and mineral resource estimation (other than for heap leach tailings resources generated post-2009), as contained in Sections 14 and 17, and as summarised in Section 1 of the Technical Report. • My last visit to the site was on the 24th of August 2010 for eight days. • I am independent of the issuer as defined in section 1.4 of National Instrument 43-101. • I have not received, nor do I expect to receive, any interest, directly or indirectly, in the project or in securities from La Mancha Resources Inc., its affiliates or subsidiaries. • I have read National Instrument 43-101 and Form 43-101F1 and the portions of the Technical Report for which I am responsible have been prepared in compliance with both. • I am not aware of any material fact or material change with respect to the subject matter of this Technical Report that is not contained in the said report and the omission of which would make the Technical Report misleading; • I consent to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication by them for regulatory purposes, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report. Dated this 22nd Day of October, 2010 ________________________________ (Signed) Remi Bosc, European Federation of Geologists N°737 Independent Consulting Geologist Arethuse Geology Sdn Bhd FINAL – Rev 0 – 22 Oct 2010 AMEC Page 303 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report CERTIFICATE OF QUALIFIED PERSON – Simon McCracken I, Simon McCracken, do hereby certify that: • I am Principal Geologist for CSA Global (UK), 2 Peel House, Barttelot Road, Horsham RH12 1DE, United Kingdom. • I graduated with a BApp.Sc. from the Royal Melbourne Institute of Technology (Australia) in 1988. • I am a Member of the Australian Institute of Geoscientists. • I have worked as a Geologist for a total of 20 years since my graduation from university. • I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43- 101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101. • I am responsible for the preparation of those parts of the Technical Report titled “The Hassai Mine Envisaged Business Plan (CIL Gold Plant and VMS Concentrator), Red Sea State, Sudan” Revision 0, dated 22 October, 2010 (the “Technical Report”) relating to Hassai heap leach tailings remnant resources, other than those reported in Section 17.4.1 by Arethuse Geology following drilling. • My last visit to the site was on the 25th of August 2010 for 7 days. • I have read the relevant sections of the Technical Report and, as at the date of this certificate,to the best of my knowledge, information and belief, those sections of the Technical Report contain all scientific and technical information that is required to be disclosed to make that part of the Technical Report not misleading. • I am independent of the issuer applying all of the tests in section 1.4 of NI 43-101. • I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form. • I consent to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication by them for regulatory purposes, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report. Dated this 22nd Day of October, 2010 ________________________________ (Signed) Simon McCracken BappSc MAIG Principal Consultant CSA Global (UK) LtdQualified Persons: FINAL – Rev 0 – 22 Oct 2010 AMEC Page 304 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 24. ILLUSTRATIONS Illustrations are included within the body of the report as appropriate. A list of illustrations is included in the table of contents. FINAL – Rev 0 – 22 Oct 2010 AMEC Page 305 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report 25. APPENDIX 1 RECENT HASSAI SOUTH, HADAL AWATIB AND KAMOEB DRILL INTERSECTIONS FINAL – Rev 0 – 22 Oct 2010 AMEC Page 306 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Hadal Awatib Long Holes from South Edge of Pit, 2007-2009 Hole ID From To Thickness Au Cu Zn Ag (m) (g/t) (%) (%) (g/t) HAE 113.00 138.00 25.00 0.54 1.14 0.79 2.84 D251 145.00 163.90 18.90 1.03 0.89 1.63 7.89 203.00 221.50 18.50 1.59 1.47 0.52 8.44 260.50 285.00 24.50 1.28 0.20 1.56 17.29 303.00 306.00 3.00 1.58 0.14 1.78 26.00 321.00 341.00 20.00 1.19 1.06 2.11 17.85 109.90 1.11 090 1.33 11.24 HAE 147.60 163.10 15.50 1.29 0.85 1.92 11.20 D252 171.50 196.00 24.50 1.24 0.83 1.08 5.20 203.00 229.30 26.30 1.00 0.85 1.14 3.42 242.30 303.00 60.70 1.81 0.43 0.21 12.29 323.40 387.75 64.35 1.63 0.45 0.22 12.21 191.35 1.52 0.58 0.59 10.05 HAE 179.06 199.00 19.94 0.36 1.41 0.35 0.00 D253 214.00 234.00 20.00 1.07 1.93 0.56 8.90 39.94 0.72 1.67 0.45 4.46 HAE D254 174.90 189.00 14.10 0.68 2.28 1.52 5.57 168.90 171.90 3.00 0.68 0.33 0.37 7.90 D256 135.00 153.00 18.00 1.14 0.24 1.55 12.28 HAE 180.00 194.00 14.00 0.29 2.66 0.66 1.71 D257 221.65 232.43 10.78 1.68 0.67 0.67 12.73 24.78 0.90 1.79 0.66 6.51 9.48 HAE D255 HAE HAE 178.65 202.00 23.35 1.05 0.59 2.07 D258 206.40 211.70 5.30 1.29 3.03 0.14 9.02 224.50 245.50 21.00 2.20 0.41 0.25 12.32 49.65 1.56 0.77 1.09 10.63 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 307 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Hadal Awatib – Holes from Base of Pit, 2007-2009 Hole ID From To Thickness Au Cu (m) (g/t) (%) 194 10.00 16.00 6.00 0.76 2.62 195 13.20 20.20 7.00 1.78 2.81 196 9.00 39.00 30.00 2.03 2.83 45.00 50.00 5.00 1.74 3.24 197 7.00 38.00 31.00 3.48 2.92 198 2.00 3.00 1.00 2.11 3.00 199 11.00 23.00 12.00 2.79 2.78 28.00 37.00 9.00 0.52 3.97 200 22.00 42.00 20.00 0.96 2.69 201 18.00 31.00 13.00 1.28 2.48 202 4.00 10.00 6.00 13.11 14.00 46.00 32.00 1.14 203 204 205 206 207 1.00 12.00 11.00 5.58 12.00 47.00 35.00 1.45 17.10 3.98 4.70 1.00 9.50 8.50 16.00 24.00 8.00 0.46 6.00 18.00 12.00 >29,6 21.00 50.00 29.00 4.80 3.43 21.70 26.66 4.96 2.27 3.45 30.50 44.60 14.10 3.50 4.71 6.50 17.00 10.50 2.18 17.00 50.00 33.00 5.76 2.17 7.16 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 308 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Hassai Long Holes 2007-2009 Hole ID From To Thickness Au Cu Zn Ag (m) (g/t) (%) (%) (g/t) HASS_D208 144.70 150.00 5.30 1.91 0.58 HASS_D209 184.45 191.00 6.55 1.20 0.88 HASS_D210 HASS_D211 225.00 255.00 30.00 1.46 1.34 225.00 233.00 8.00 1.36 2.96 251.00 255.00 4.00 2.00 3.20 185.00 203.00 18.00 1.51 1.21 216.00 221.00 5.00 1.25 2.81 247.50 261.20 13.70 1.14 0.60 HASS_D212 Barren HASS_D213 274.70 296.30 21.60 2.41 2.36 HASS_D214 289.00 311.20 22.20 2.01 1.07 HASS_D215 265.00 275.50 10.50 2.25 1.93 HASS_D216 358.00 392.50 34.50 1.77 1.70 HASS_D217 Stopped due to high deviation HASS_D218 626.00 662.00 36.00 0.60 0.12 HASS_D219 222.50 234.80 12.30 1.25 1.79 HASS_D220 341.00 358.00 17.00 1.90 1.94 375.00 380.00 5.00 0.93 1.79 267.80 288.00 20.20 1.78 2.02 325.00 367.00 42.00 3.11 1.63 372.00 393.00 21.00 1.51 0.75 HASS_D221 HASS_D222 HASS D223 HASS_D224 HASS_D225 230.90 246.00 15.10 2.61 2.22 248.00 263.00 15.00 1.55 1.49 304.65 333.30 28.65 1.68 1.70 339.00 342.70 3.70 1.94 1.21 260.00 282.60 22.60 2.03 1.91 289.00 292.00 3.00 2.10 2.50 HASS_D226 248.40 261.50 13.10 2.35 3.18 HASS_D227 212 214 2.00 0.89 1.42 HASS_D228 299.00 309.20 10.20 0.91 2.29 HASS_D229 177.00 181.50 4.50 1.41 0.47 HASS_D180 172.75 179.75 7.00 1.83 1.86 1.64 10.65 HASS_D180 193.55 198.5 4.95 0.92 0.68 0.25 5.00 HASS_D181 146.4 147.05 0.65 1.84 1.10 0.02 5.00 HASS_D181 154.96 155.35 0.39 1.47 1.62 0.78 9.00 HASS_D181 170.75 178.58 7.83 1.01 2.21 0.19 5.06 HASS_D182 146.74 149.28 2.54 0.95 2.66 0.04 5.00 41.00 HASS_D230 Barren HASS_D182 158.25 158.5 0.25 7.41 4.11 0.79 HASS_D182 167.35 178.78 11.43 0.93 1.36 0.05 5.19 HASS_D183 157.93 158.93 1.00 3.48 2.36 0.03 5.00 HASS_D183 180.9 187.03 6.13 1.67 2.50 0.37 6.26 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 309 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Hassai Long Holes 2007-2009 Hole ID From To Thickness Au Cu Zn Ag (m) (g/t) (%) (%) (g/t) HASS_D183 188.23 193.28 5.05 1.01 0.36 0.10 5.00 HASS_D184 156.5 168 11.50 1.32 1.25 0.44 5.00 HASS_D184 174 174.7 0.70 1.47 3.25 1.03 12.00 HASS_D184 176.82 181.55 4.73 0.91 1.65 0.19 5.00 HASS_D185 194.8 206.18 11.38 1.20 1.41 0.61 5.07 HASS_D185 214.68 222.7 8.02 0.99 1.31 0.26 5.25 HASS_D186 145.98 154.04 8.06 0.44 0.53 0.23 5.24 HASS_D186 154.04 168.2 14.16 1.36 2.45 0.45 6.57 HASS_D187 156.64 160.41 3.77 1.54 0.26 0.14 5.00 HASS_D187 163.66 179.73 16.07 1.08 1.63 0.33 5.78 HASS_D188 128.15 141.08 12.93 0.77 0.42 0.04 5.00 HASS_D188 141.08 158.58 17.50 1.41 2.67 0.36 6.12 HASS_D189 134.05 147.93 13.88 1.05 0.67 0.06 5.00 HASS_D189 148.63 150.46 1.83 1.59 0.68 0.01 0.82 HASS_D189 152 159.5 7.50 1.03 1.23 0.54 5.89 HASS_D190 109 121.3 12.30 0.71 0.18 0.03 5.00 HASS_D190 125.47 131.43 5.96 0.10 0.32 0.04 5.00 HASS_D190 131.43 135.06 3.63 0.62 0.31 1.18 6.34 HASS_D190 136.03 139.68 3.65 1.69 0.69 0.06 5.00 HASS_D191 240.27 264.39 24.12 1.79 1.48 0.61 5.89 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 310 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Hole ID Hadal Awatib – VMS Drilling Results, from Pit 2008-2009 Block A+B From To Thickness Au Cu Zn (m) (g/t) (%) (%) HAE R 260 13.00 15.00 2.00 0.98 2.42 0.44 10.00 HAE R 261 HAE R 262 13.00 6.00 24.00 14.00 11.00 8.00 0.71 0.91 1.93 3.68 0.28 0.20 1.18 8.44 HAE R263 inc 1.00 3.00 17.00 7.00 16.00 4.00 0.66 1.55 4.68 11.04 HAE R264 0.00 20.00 20.00 1.23 3.01 0.77 7.00 inc HAE R265 16.00 10.00 19.00 24.00 3.00 14.00 2.35 1.16 13.27 0.94 1.26 25.33 Ag (g/t) CS AA CS BB 8.69 20.75 46.00 51.00 5.00 3.10 4.60 HAE R266 1.00 19.00 4.00 23.00 3.00 4.00 1.56 0.83 0.45 1.34 29.67 4.25 HAE R267 12.00 21.00 9.00 0.21 1.66 HAE R268 1.00 12.00 11.00 9.10 99.27 2.00 6.00 4.00 21.10 258.00 CS CC inc HAE R269 6.00 10.00 4.00 1.73 4.99 HAE R270 1.00 29.00 20.00 36.00 19.00 7.00 1.34 0.87 0.80 2.20 HAE R271 0.00 19.00 19.00 1.18 0.77 22.00 46.00 37.00 58.00 15.00 12.00 1.04 0.75 0.69 3.11 HAE R272 47.00 50.00 HAE R273 HAE R274 17.00 3.00 33.00 11.00 16.00 8.00 1.00 0.67 0.34 9.25 1.23 0.63 8.79 5.57 1.25 4.80 8.75 0.91 1.36 0.94 HAE D 316 98.60 141.00 42.40 1.63 2.00 107.00 61.00 8.40 24.00 1.56 1.09 5.34 HAE D 317 98.60 37.00 64.60 67.00 2.40 1.17 4.29 250.00 281.00 278.00 286.00 28.00 5.00 0.83 2.33 1.34 0.51 5.00 16.00 11.00 19.24 51.64 6.00 18.00 11.00 27.00 5.00 9.00 37.64 1.72 83.80 11.67 36.00 61.00 25.00 0.92 61.00 82.00 77.00 85.00 16.00 3.00 1.61 2.20 91.00 93.00 2.00 4.22 2.39 34.00 48.00 44.00 71.00 10.00 23.00 0.63 0.83 4.63 1.86 0.78 14.10 5.91 80.00 114.00 34.00 1.70 0.59 0.60 11.35 6.00 17.00 11.00 1.04 3.12 1.72 10.64 20.40 CS DD HAE R275 inc HAE R277 HAE R 278 HAE R279 0.70 1.87 8.88 11.63 22.67 49.50 8.73 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 311 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Hadal Awatib – VMS Drilling Results, from Pit 2008-2009 Block A+B From To Thickness Au Cu Zn (m) (g/t) (%) (%) Hole ID HAE R280 inc HAE R281 HAE R282 Ag (g/t) 33.00 37.00 4.00 0.72 11.95 104.00 130.00 26.00 1.19 5.71 1.17 14.15 25.25 104 133 107 174 3.00 41.00 1.99 0.78 14.07 0.35 2.13 29.67 3.80 182 194 12.00 1.79 0.58 18 26 22 34.00 4.00 8.00 2.86 2.13 37 48 11.00 1.14 3.87 0.80 10.82 129 4 144 48 15.00 44.00 1.03 1.34 5.60 0.63 1.44 1.23 12.07 6.61 13.00 71 79 8.00 0.80 3.13 0.66 6.94 inc 222.4 222.4 266.5 226 44.10 3.60 1.54 0.81 1.62 5.33 1.57 0.74 15.02 15.50 inc 251 254 3.00 3.79 5.01 2.21 40.00 3 40 24 54 21.00 14.00 0.58 1.10 1.90 0.69 0.74 1.82 68 105 37.00 0.57 2.82 87 94 7.00 0.95 6.56 HAE R283 inc 4.64 5.24 0.83 13.29 CC EE HAE R286 5 21 16.00 7.47 21.81 inc 5 12 7.00 21.53 45.60 HAE R287 6 18 8 33 2.00 15.00 3.625 1.12 HAE D 318 HAE D 319 HAE D 320 10.5 9.07 1.25 36 46 10.00 1.30 0.46 0.73 0.00 24.00 24.00 44.00 24.00 20.00 1.07 1.20 0.97 1.43 54.00 73.00 19.00 1.13 0.65 77.00 9.00 107.00 34.00 30.00 25.00 0.57 1.26 3.54 8.90 34.00 46.00 12.00 1.09 1.71 53.00 192.80 66.30 196.10 13.30 3.30 1.04 0.58 3.22 12.74 2.59 26.36 216.20 248.80 32.60 1.03 15.55 2.00 16.40 14.40 1.15 4.35 142.00 157.60 149.00 163.00 7.00 5.40 0.86 1.16 9.58 2.47 1.63 169.00 182.00 13.00 1.24 0.98 2.43 21.29 15.11 CC FF HAE R293 barren HAE R294 8 31 23.00 1.55 HAE R295 62 88 77 129 15.00 41.00 1.61 1.16 0.37 0.65 1.98 1.34 62 106 44.00 0.83 0.55 1.74 8.8 121 162 41.00 1.18 1.09 1.51 15.48 HAE R296 19.43 13.2 13.6 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 312 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Hadal Awatib – VMS Drilling Results from Pit, Block C+D, 2008-2009 Hole ID HAE D 259 From To Thickness Au Cu Zn Ag (m) (g/t) (%) (%) (g/t) 33.85 42.13 8.28 1.13 5.39 0.42 13.45 45.10 62.15 17.05 1.26 1.13 0.32 12.24 HAE R 304 29.00 33.00 4.00 0.76 1.91 0.22 5.75 HAE R 305 0.00 8.00 0.66 4.96 0.79 9.75 1.01 10.64 13.00 27.00 14.00 0.92 4.07 33.00 42.00 9.00 1.33 0.63 8.56 HAE R 306 9.00 21.00 12.00 1.23 0.63 8.56 inc 16.00 19.00 3.00 2.35 13.27 1.26 25.33 23.00 46.00 23.00 0.80 2.87 2.51 7.70 16.10 HAE R307 33.00 43.00 10.00 1.03 2.13 1.33 57.00 70.00 13.00 1.24 0.53 2.28 28.00 35.00 7.00 1.46 4.23 37.00 65.00 28.00 1.54 0.97 1.23 0.00 6.00 6.00 0.86 1.77 0.24 11.00 25.00 14.00 0.80 0.96 3.80 25.00 47.00 22.00 0.70 1.41 1.17 HAE R 310 0.00 5.00 5.00 1.29 0.25 15.00 39.00 24.00 0.71 0.41 1.39 HAE R 311 10.00 50.00 40.00 0.68 3.19 1.54 24.00 25.25 1.25 0.67 HAE R308 HAE R 309 HAE R 312 HAE R 313 9.00 16.29 15.11 6.00 4.30 drill hole interrupted HAE R 314 0.00 1.00 1.00 3.61 HAE R 315 13.00 29.50 16.50 1.52 5.81 8.00 51.00 0.84 16.67 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 313 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Kamoeb Mineralised Intersections Hole ID From To Thickness AuCy Au FA (m) (g/t) (g/t) 3 11.33 0 kame2_1 KAM_D006 8.5 11.5 17.5 31.5 14 6.5 0 KAM_D007 12 17.5 5.5 8.14 0 26.5 39 12.5 3.11 0 KAM_D192 10.3 18 7.7 10.08 0 25 63.79 38.79 7.03 0 KAM_D197 3 19.5 16.5 6.12 0 25.5 39.5 14 4.72 0 KAM_D198 0 42.3 42.3 9.55 0 KAM_D199 26 29.8 3.8 8.83 0 33.5 44 10.5 6.9 0 25 36 11 4.78 0 58 66 8 23.21 0 KAM_D201 31 32.7 1.7 25.4 0 KAM_R339 0 4 4 0 6.55 KAM_R342 0 14 14 0 4.08 KAM_R343 3 22 19 0 10.94 KAM_R345 6 19 13 0 6.95 KAM_R346 31 60 29 0 10.84 KAM_R347 0 31 31 0 7 KAM_R348 55 82 27 0 6.41 KAM_D200 kame2_2 KAM_D020 61.87 66.25 4.38 0.09 0 KAM_D026 12.5 34.5 22 2.22 0 0 KAM_D027 25 36.5 11.5 4.12 40.5 46.5 6 1.1 0 50 53 3 1.07 0 24 39 15 7.08 0 43 49 6 22.12 0 52 58 6 1.33 0 61 71 10 3.49 0 KAM_D038 30.2 31.5 1.3 5.36 0 39 40 1 6.6 0 KAM_D064 50 53 3 10.47 0 KAM_D072 25 28 3 1.53 0 KAM_D073 55 63.5 8.5 3.39 0 70 83 13 4.02 0 KAM_D074 45 47.5 2.5 5.5 0 KAM_D075 46 49 3 1.07 0 67 69 2 1.35 0 KAM_D028 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 314 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Kamoeb Mineralised Intersections Hole ID KAM_D076 From To Thickness AuCy Au FA (m) (g/t) (g/t) 41 49 8 2.45 0 75 78.5 3.5 1.66 0 KAM_D201 0 21 21 2.93 0 KAM_D202 0 11 11 11.6 0 KAM_R351 0 25 25 0 4.01 KAM_R352 4 22 18 0 10.95 KAM_R354 0 7 7 0 2.04 8 18 10 0 6 KAM_R355 2.52 15 12.48 0 2.42 20 45 25 0 2.37 3 5 2 0 5.97 9 12 3 0 16.17 KAM_R356 KAM_R358 0 11 11 0 5.31 KAM_R359 19 24 5 0 6.59 34 35 1 0 0.76 KAM_R360 51 55 4 0 1.61 60 69 9 0 2.47 74 77 3 0 1.24 KAM_R364 32 49 17 0 1.41 KAM_R365 35 40 5 0 4.4 56 57 1 0 0.81 30 33 3 0 3.79 36 37 1 0 1.59 35 37 2 0 10.62 KAM_R366 KAM_R372 KAM_R373 KAM_R374 KAM_R404 41 51 10 0 3.71 59 61 2 0 2.21 41 44 3 0 3.57 63 64 1 0 2.94 52 54 2 0 63.25 37 40 3 0 0.78 42.29 46.13 3.84 0 0.14 59.28 60 0.72 0 0.07 64 69 5 0 3.98 KAM_D022 86.5 87.5 1 1.8 0 KAM_D023 125 127 2 2.75 0 KAM_D026 37 41 4 1.8 0 KAM_D027 67.5 69.5 2 2.8 0 KAM_D074 57.5 58.5 1 1.4 0 KAM_D075 21 29 8 2.32 0 77 80.5 3.5 2.77 0 KAMS_R028 KAMS_R029 kame2_3 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 315 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Kamoeb Mineralised Intersections Hole ID KAM_D076 From To Thickness AuCy Au FA (m) (g/t) (g/t) 29 38 9 8.61 0 88 90 2 6.85 0 95.5 104.5 9 6.3 0 14 26.5 12.5 2.92 0 KAM_D088 91 91.5 0.5 5.5 0 KAM_D191 30 35 5 2.58 0 KAM_D202 40 44 4 8.05 0 KAM_R343 29 31 2 0 4.96 KAM_R346 67 71 4 0 5.39 KAM_R348 26 31 5 0 7.53 KAM_R350 7 8 1 0 1.27 26 30 4 0 4.6 KAM_D077 KAM_R351 KAM_R353 KAM_R354 35 41 6 0 2.92 33 35 2 0 0.96 3.97 2 16 14 0 28 31 3 0 2.8 34 36 2 0 0.89 41 48 7 0 1.47 20 22 2 0 1.52 29 30 1 0 1.19 41 44 3 0 9.16 49 50 1 0 1.06 KAM_R356 20 23 3 0 8.38 KAM_R357 19 21 2 0 2.1 KAM_R358 36 40 4 0 1.79 44 46 2 0 3.17 KAM_R359 64 66 2 0 15.4 KAM_R365 78 79 1 0 1.11 KAM_R366 39 48 9 0 4.53 KAM_R370 85 87 2 0 1.77 KAM_R383 111 112 1 0 1.53 KAMS_R030 83 89 6 0 1.65 KAMS_R031 108 110 2 0 2.03 KAMS_R032 113 116 3 0 3.29 KAMS_R033 144 148 4 0 1.25 KAM_D172 24 25 1 3.9 0 KAM_D173 35 39 4 2.1 0 50 52.5 2.5 1.16 0 KAM_D174 35 35.8 0.8 0.8 0 KAM_D175 19 21 2 1.05 0 kamn4_1 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 316 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Kamoeb Mineralised Intersections Hole ID KAM_D176 From To Thickness AuCy Au FA (m) (g/t) (g/t) 13 23 10 4.88 0 46 47 1 1.46 0 23.5 30 6.5 4.68 0 46 48.5 2.5 2.7 0 KAM_D178 70.5 71.5 1 1.9 0 KAM_D182 44 51 7 1.19 0 KAM_D183 52.35 57 4.65 2.22 0 KAM_D187 12 15.5 3.5 4.16 0 KAM_D188 39 44 5 2.38 0 47.5 51 3.5 4.93 0 KAM_D189 53.7 54 0.3 1.7 0 56 62 6 1.92 0 68 71.5 3.5 1.27 0 90 92 2 1.9 0 KAM_D177 KAM_D190 KAM_D234 KAM_D235 KAM_D236 KAM_D237 97 99 2 3.3 0 102 106 4 2.15 0 12.5 16.2 3.7 1.89 0 26 37 11 1.41 0 0 5.35 5.35 0 0 24 28.3 4.3 3.84 0 34 35.5 1.5 1.6 0 44 49 5 0.97 0 5 11.5 6.5 2.22 0 14.5 16 1.5 4.9 0 26 27.5 1.5 1.4 0 37.5 43.5 6 2.16 0 45 49 4 2.17 0 5 7 2 1.03 0 11 13.4 2.4 2.89 0 44.3 47.5 3.2 1.17 0 KAM_D239 4 4.5 0.5 1.5 0 45 45.2 0.2 2.3 0 KAM_D240 40.6 42.7 2.1 0.37 0 KAM_D241 30 30.8 0.8 2 0 37 37.8 0.8 2.1 0 46.5 50.8 4.3 2.41 0 22.5 31 8.5 1.44 0 34.1 38.7 4.6 4.18 0 35.29 39.19 3.9 0 0 57.7 62.6 4.9 2.76 0 40.8 42.7 1.9 2.55 0 53.8 57 3.2 2.77 0 KAM_D243 KAM_D245 KAM_D246 KAM_D247 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 317 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Kamoeb Mineralised Intersections Hole ID KAM_D248 From To Thickness AuCy Au FA (m) (g/t) (g/t) 58 62.8 4.8 1.25 0 73.5 76.3 2.8 1.01 0 KAM_D249 21.7 26.65 4.95 4.02 0 KAM_D251 89.3 91.1 1.8 2.93 0 KAM_D255 8.6 11.55 2.95 4.53 0 27 28 1 3.1 0 31 34 3 2.07 0 KAM_D256 KAM_D257 36.8 39.2 2.4 2.17 0 6 13.5 7.5 2.89 0 12.2 16.4 4.2 2.45 0 25.49 25.83 0.34 0 0 28.68 30 1.32 0 0 KAM_D258 20.5 23 2.5 4.4 0 KAM_D259 18.3 20 1.7 6.52 0 KAM_D260 34.5 36 1.5 3.13 0 7.5 8.9 1.4 1.71 0 16.5 17.5 1 4.6 0 45 51.5 6.5 6.32 0 31.5 32.5 1 3 0 34.4 35 0.6 2.7 0 47 62 15 4.01 0 39.1 44 4.9 1.22 0 47 47.8 0.8 5.8 0 KAM_D265 24.5 24.6 0.1 2.7 0 58 61.5 3.5 1.51 0 KAM_D266 40.5 43.2 2.7 4.5 0 KAM_D262 KAM_D263 KAM_D268 22 24 2 1.13 0 KAM_D269 18.3 23.5 5.2 2.71 0 KAM_D270 28 28.6 0.6 1.7 0 44.2 47.5 3.3 6.75 0 6.5 7.5 1 1 0 0 KAM_D271 KAM_D272 8.8 11.8 3 6.84 17.1 30 12.9 2.23 0 40 42 2 1.95 0 KAM_D273 21.4 26 4.6 1.9 0 KAM_D274 56.1 57 0.9 2.1 0 68.2 69.45 1.25 2.43 0 18 18.7 0.7 2.8 0 0 KAM_D275 KAM_D276 19.5 38.5 19 3.22 43.7 44.6 0.9 1.3 0 2 7 5 6.08 0 17.8 21 3.2 1.48 0 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 318 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Kamoeb Mineralised Intersections Hole ID KAM_D277 From To Thickness AuCy Au FA (m) (g/t) (g/t) 25.5 31.5 6 2.35 0 1.5 4.5 3 2.43 0 7.6 13.2 5.6 5.56 0 32.2 33 0.8 2.4 0 KAM_D278 10.7 15.9 5.2 6.26 0 KAM_D279 28.15 35.2 7.05 3.55 0 KAM_D280 41.9 50.5 8.6 3.24 0 KAM_D287 24.1 26.8 2.7 0.63 0 KAM_D289 20.6 21.2 0.6 1 0 KAM_R418 34 36 2 0 1.76 KAM_R419 34 36 2 0 1.25 53 58 5 0 1.2 61 64 3 0 2.06 KAM_R420 29 37 8 0 1.74 KAM_R421 48 53 5 0 1.11 67 90 23 0 0.97 2 16 14 0 2.35 KAM_R422 KAM_R423 4 16 12 0 2.45 27 31 4 0 3.84 KAM_R424 11 15 4 0 4.1 KAM_R425 11 21 10 0 1.02 31 33 2 0 1.29 KAM_R426 12 13 1 0 1.03 1.36 0 0 kamn4_2 KAM_D179 15.2 18 2.8 0 kamn4_3 KAM_D284 31.2 31.65 0.45 3.4 0 KAM_D285 19.4 22.8 3.4 1.94 0 KAM_D286 16.1 19 2.9 4.21 0 0 kams1_1 KAM_D001 3 4.5 1.5 5.23 0 KAM_D002 17 19.5 2.5 4.92 0 KAM_D003 26 28 2 4.92 0 KAM_D004 36.5 38.5 2 1.2 0 KAM_D005 58.5 62 3.5 3.83 0 0 KAM_D009 107 110 3 8.02 KAM_D010 136.5 138.5 2 4.2 0 KAM_D019 4.5 21.5 17 4.33 0 KAM_D020 32 51.5 19.5 7.52 0 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 319 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Kamoeb Mineralised Intersections Hole ID KAM_D021 KAM_D022 From To Thickness AuCy Au FA (m) (g/t) (g/t) 46.5 67.5 21 3.08 0 76 77.8 1.8 6.87 0 45 69 24 2.39 0 73.5 79.5 6 3 0 KAM_D023 94.5 112.5 18 2.68 0 KAM_D024 138.5 140.5 2 2.35 0 142.5 150 7.5 3.1 0 KAM_D031 14 16 2 6.2 0 KAM_D032 44 47 3 2.55 0 0 KAM_D033 65.8 69.5 3.7 2.44 KAM_D034 107.5 108.5 1 2.1 0 KAM_D039 18.5 31.5 13 6.97 0 KAM_D040 24 41 17 3.71 0 KAM_D041 61.5 68.5 7 3.97 0 KAM_D042 85 90 5 4.8 0 KAM_D043 103.5 106.3 2.8 3.63 0 KAM_D044 119 122.5 3.5 3.33 0 KAM_D048 29 32.5 3.5 6.17 0 KAM_D049 51 57 6 2.52 0 KAM_D050 79.5 85 5.5 3.29 0 KAM_D051 11.13 37 25.87 0 0 KAM_D052 95 98.5 3.5 6.8 0 KAM_D054 19.5 24 4.5 8.6 0 25 38 13 6.62 0 KAM_D055 28.5 30 1.5 9 0 33 48.5 15.5 11.71 0 KAM_D056 55 66 11 3.19 0 KAM_D057 74.5 79 4.5 12.01 0 KAM_D058 100.5 104.5 4 3.42 0 KAM_D059 124 126.5 2.5 8.6 0 KAM_D060 79.12 81 1.88 0 0 KAM_D061 28.5 41.5 13 5.77 0 45 57 12 6.98 0 KAM_D062 47.5 51.5 4 6.88 0 56 69 13 6.59 0 KAM_D063 89.5 93.5 4 1.18 0 KAM_D064 24 44 20 6.89 0 KAM_D065 54 71 17 4.22 0 0 KAM_D066 76 78.5 2.5 9.9 82.2 87.5 5.3 1.6 0 KAM_D067 110.5 119.5 9 3.22 0 KAM_D068 134.5 148.5 14 3.49 0 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 320 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Kamoeb Mineralised Intersections Hole ID From To Thickness AuCy Au FA (m) (g/t) (g/t) KAM_D069 98.3 98.6 0.3 1.9 0 KAM_D070 60 60.5 0.5 3 0 KAM_D075 5.23 11.53 6.3 0 0 KAM_D076 8 21 13 2.85 0 KAM_D077 41.5 50 8.5 3.14 0 KAM_D078 42.5 61.5 19 7.36 0 KAM_D079 52.5 69 16.5 6.99 0 70.5 74.5 4 9.17 0 KAM_D083 45.5 49 3.5 2.26 0 KAM_D084 70.7 73.7 3 2.97 0 KAM_D085 91 92 1 2.4 0 KAM_D086 124.32 126 1.68 0 0 KAM_D088 66 86 20 5.58 0 KAM_D090 21.5 24.5 3 1.87 0 KAM_D091 51.2 54 2.8 3.28 0 KAM_D092 73 78 5 2.42 0 KAM_D093 96 106 10 2.86 0 KAM_D096 150 158 8 4.65 0 KAM_D101 6 10 4 3.25 0 KAM_D102 23.8 28 4.2 7.73 0 KAM_D103 51 56 5 7.46 0 KAM_D104 86.5 94.3 7.8 5.54 0 KAM_D105 13.5 17 3.5 2.81 0 KAM_D106 34.5 36.2 1.7 2.42 0 KAM_D107 49.3 52.3 3 0.9 0 KAM_D108 83 85.5 2.5 4.98 0 KAM_D113 13.5 16.5 3 5.47 0 KAM_D114 36.5 38.5 2 3.1 0 41.5 43.2 1.7 2.18 0 KAM_D115 58 60 2 1.6 0 KAM_D116 70.5 74 3.5 1.54 0 KAM_D122 12.5 17.5 5 4 0 KAM_D123 33.5 36.5 3 1.2 0 KAM_D124 16 20 4 3.2 0 KAM_D125 112 122 10 4 0 KAM_D138 59 65 6 2.5 0 KAM_D139 78 81 3 1.47 0 KAM_D140 41 42.5 1.5 4.43 0 KAM_D141 61.7 63.7 2 2.45 0 KAM_D142 89 93 4 2.33 0 KAM_D191 43 54.8 11.8 8.2 0 KAM_D193 64 82.5 18.5 4.68 0 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 321 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Kamoeb Mineralised Intersections Hole ID From To Thickness AuCy Au FA (m) (g/t) (g/t) KAM_D195 48 48.5 0.5 1.1 0 KAM_D402 148.55 153 4.45 0 1.86 156 166 10 0 5.98 KAM_D403 171 180 9 0 4.11 11.11 KAM_R360 5 12 7 0 20 33 13 0 5.07 KAM_R361 27 28 1 0 0.84 KAM_R362 10 26 16 0 3.46 KAM_R363 0 29 29 0 3.89 KAM_R364 0 5.83 5.83 0 7.7 KAM_R365 0 6 6 0 8.26 13 16 3 0 1.34 KAM_R366 0 17 17 0 3.15 KAM_R367 0 1 1 0 1.54 KAM_R369 105 111 6 0 4.98 KAM_R370 63 83 20 0 2.17 KAM_R371 38 49 11 0 2.69 KAM_R372 10 15 5 0 9.22 19 23 4 0 4.71 KAM_R373 18 35 17 0 5.56 KAM_R374 24 37 13 0 3.22 38 41 3 0 5.84 KAM_R375 42 43 1 0 1.91 KAM_R377 47 48 1 0 0.85 KAM_R383 83 108 25 0 1.89 KAM_R385 22 24 2 0 1.86 KAM_R386 17 19 2 0 2.84 24 27 3 0 2.24 KAM_R387 27 37 10 0 1.65 KAM_R388 16 18 2 0 3.99 KAM_R391 40 41 1 0 1.47 KAM_R392 179 186 7 0 4.93 KAM_R404 0 35 35 0 3.24 KAMS_R001 37 39 2 0 3.8 KAMS_R002 68 71 3 0 4.26 KAMS_R003 87 90 3 0 2.91 KAMS_R004 104 105 1 0 7.48 KAMS_R005 125 127 2 0 3.61 KAMS_R007 15 20 5 0 7.73 KAMS_R008 64 71 7 0 4.5 KAMS_R009 72 75 3 0 1.84 KAMS_R010 96 99 3 0 5.83 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 322 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Kamoeb Mineralised Intersections Hole ID From To Thickness AuCy Au FA (m) (g/t) (g/t) KAMS_R011 111 113 2 0 2.66 KAMS_R012 33 50 17 0 5.02 KAMS_R013 60 65 5 0 5.23 KAMS_R014 92 96 4 0 4.99 KAMS_R015 132 134 2 0 8.29 KAMS_R016 124 128 4 0 1.94 KAMS_R017 37 45 8 0 5.95 KAMS_R018 55 62 7 0 3.91 KAMS_R019 76 80 4 0 3.38 9.84 KAMS_R020 93 98 5 0 KAMS_R021 123 125 2 0 2.87 KAMS_R022 33 50 17 0 3.05 KAMS_R023 51 54 3 0 5.99 57 64 7 0 2.98 KAMS_R024 71 79 8 0 3.95 KAMS_R025 93 98 5 0 6.23 KAMS_R026 124 129 5 0 6.77 KAMS_R027 147 160 13 0 3.18 KAMS_R028 18 36 18 0 5.16 KAMS_R029 46 60 14 0 6.88 KAMS_R030 49 62 13 0 3.57 66.89 81 14.11 0 2.23 62 76 14 0 2.27 78 102 24 0 4.52 KAMS_R031 KAMS_R032 94 108 14 0 2.14 KAMS_R033 113 136 23 0 3.03 KAMS_R034 134 146 12 0 4.48 KAMS_R035 13 17 4 0 3.88 20 40 20 0 3.4 33 41 8 0 3.92 45 60 15 0 5.4 KAMS_R037 56 61 5 0 7.45 KAMS_R038 77 79 2 0 5.97 KAMS_R039 88 92.84 4.84 0 8.3 KAMS_R036 0 kams1_2 KAM_D008 14 35.5 21.5 2.36 0 KAM_D011 10 20.5 10.5 2.5 0 29.5 36 6.5 3.05 0 41 42 1 2.3 0 16.5 32.5 16 4.04 0 43 45 2 1.7 0 KAM_D012 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 323 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Kamoeb Mineralised Intersections Hole ID From To Thickness AuCy Au FA (m) (g/t) (g/t) 47 51 4 1.08 0 41.5 47 5.5 3.36 0 57 60 3 5.3 0 63 72 9 4.31 0 67 72.5 5.5 3.57 0 77 88 11 5.15 0 KAM_D015 97.5 102 4.5 1.6 0 104 108 4 2.68 0 KAM_D016 132 141.3 9.3 2.26 0 KAM_D013 KAM_D014 KAM_D023 75 81 6 3.08 0 KAM_D024 130.5 134.5 4 2 0 KAM_D035 35 36 1 6.8 0 KAM_D036 56.5 58.5 2 10.65 0 KAM_D037 80.4 81 0.6 3.5 0 KAM_D041 17 25 8 1.74 0 KAM_D042 47 52.5 5.5 1.48 0 KAM_D043 75.5 77 1.5 0.87 0 KAM_D044 101.56 102.99 1.43 0 0 KAM_D050 33.5 36.5 3 1.43 0 KAM_D052 50.5 54 3.5 3.67 0 KAM_D053 78 79 1 1.45 0 KAM_D056 21.5 27 5.5 2.15 0 KAM_D057 53 54 1 1.2 0 KAM_D058 84 85 1 1 0 KAM_D059 109.5 114.5 5 2.78 0 KAM_D065 19 28.5 9.5 3.07 0 KAM_D066 59.5 60.5 1 6.3 0 KAM_D067 98 99 1 0.6 0 KAM_D068 123 125 2 2.65 0 KAM_D085 11.7 13.2 1.5 2.1 0 KAM_D086 49.5 54 4.5 5.92 0 KAM_D087 76 80.5 4.5 3.63 0 KAM_D094 16 23 7 1.63 0 24.5 31 6.5 3.13 0 18.5 20.5 2 2.6 0 26.5 40.5 14 5.25 0 15 20.5 5.5 6.34 0 28.5 33.5 5 4.34 0 KAM_D095 KAM_D096 KAM_D097 KAM_D098 37.5 47 9.5 3.05 0 35.5 41.5 6 1.95 0 43.5 46 2.5 2.32 0 61.5 67.5 6 1.7 0 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 324 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Kamoeb Mineralised Intersections Hole ID KAM_D099 From To Thickness AuCy Au FA (m) (g/t) (g/t) 69.5 71 1.5 1.87 0 89.8 91.6 1.8 3.91 0 96.3 97.3 1 1.5 0 KAM_D100 5.5 6.5 1 1.1 0 15.5 22.5 7 4.91 0 KAM_D109 21.5 24.5 3 9.17 0 KAM_D110 59.5 64.2 4.7 3.13 0 KAM_D111 86.2 96 9.8 3.99 0 KAM_D112 113 129.5 16.5 1.73 0 KAM_D117 11.5 16.5 5 5.42 0 20.5 26.5 6 11.05 0 29.5 35.5 6 5.62 0 37.5 51.5 14 4.94 0 59 69 10 2.47 0 0 KAM_D118 KAM_D119 KAM_D120 77 92 15 3.09 KAM_D121 107.5 124 16.5 2.16 0 KAM_D126 20.5 26.5 6 4.55 0 KAM_D127 23.5 30.3 6.8 4.63 0 54 58.5 4.5 4.64 0 20 21 1 2.6 0 46.5 48 1.5 1.27 0 KAM_D128 KAM_D129 51 59.5 8.5 2.3 0 47 49.5 2.5 3.28 0 54 60 6 3.29 0 67 75.5 8.5 1.48 0 KAM_D130 68.3 69 0.7 2.4 0 71 76.5 5.5 3.23 0 KAM_D131 109.2 110.6 1.4 1.8 0 KAM_D132 41 42.26 1.26 1.83 0 KAM_D133 59 67 8 6.54 0 KAM_D134 98.8 105 6.2 2.45 0 KAM_D135 41.7 43 1.3 8.4 0 KAM_D136 63 66 3 3.93 0 KAM_D137 98.5 107.5 9 4.12 0 KAM_D143 36 39 3 1.17 0 52 55 3 0.97 0 54 57.5 3.5 2.14 0 62 63.5 1.5 3.3 0 KAM_D144 KAM_D145 72.5 74 1.5 1.87 0 56 58.5 2.5 1.92 0 72 81 9 2.56 0 91 97.5 6.5 1.34 0 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 325 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Kamoeb Mineralised Intersections Hole ID From To Thickness AuCy Au FA (m) (g/t) (g/t) 99.5 104.5 5 1.3 0 KAM_D146 76.5 82 5.5 3.02 0 86.5 89.5 3 1.77 0 KAM_D149 31.5 33 1.5 1.13 0 KAM_D150 11.5 19 7.5 6.44 0 KAM_D204 12 20 8 6.05 0 KAM_D205 17 25 8 3.54 0 KAM_D206 KAM_D207 KAM_D208 0 5 5 0.55 0 11 22 11 1.95 0 26 32 6 2.39 0 0 5.86 5.86 0 0 25.7 27.5 1.8 1.3 0 35 36 1 1 0 52.3 55 2.7 0.85 0 0 0 26 26 2.33 41.5 45.5 4 2.7 0 49 55 6 1.13 0 KAM_D209 34.5 39.5 5 1.32 0 KAM_D211 38 44 6 2.65 0 KAM_D212 70 72 2 2.2 0 KAM_D213 45.5 47.5 2 4.25 0 71.5 73.2 1.7 1.88 0 KAM_R384 5 11 6 0 2.54 KAM_R392 35 42 7 0 6.57 KAM_R393 29 33 4 0 3.09 34 39 5 0 5.29 59 65 6 0 4.58 KAM_R394 8 16 8 0 4.67 KAM_R395 0 1 1 0 0.88 12 38 26 0 2.65 0 3 3 0 1.92 18 22 4 0 1.59 24 25 1 0 0.76 33 49 16 0 1.01 KAM_R396 KAM_R398 KAM_R399 KAM_R400 7 11 4 0 4.07 16 27 11 0 1.96 40 54 14 0 2.85 0 8 8 0 1.51 1.99 11 24 13 0 39 43 4 0 4.33 8 18 10 0 4.39 37 39 2 0 1.64 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 326 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Kamoeb Mineralised Intersections Hole ID KAM_R401 From To Thickness AuCy Au FA (m) (g/t) (g/t) 47 50 3 0 1.25 13 20 7 0 1.85 KAM_R405 8 12 4 0 5.11 KAM_R406 43 47 4 0 6.7 KAM_R407 63 72 9 0 5.38 KAM_R408 56 70 14 0 3.01 KAM_R409 19 26 7 0 4.85 27 32 5 0 1.85 KAM_R410 11 23 12 0 2.57 KAM_R411 22 33 11 0 1.74 KAM_R412 43 50 7 0 1.98 KAM_R413 59 64 5 0 3.61 KAM_R414 63 66 3 0 1.31 KAM_R415 24 25 1 0 2.69 47 49 2 0 0.82 KAMS_R003 6.01 9 2.99 0 5.06 KAMS_R004 25 28 3 0 2.25 KAMS_R005 46 48 2 0 5.83 KAMS_R006 68 72 4 0 6.04 KAMS_R008 25 28 3 0 1.48 KAMS_R009 24 32 8 0 2.06 KAMS_R010 59 60 1 0 0.75 KAMS_R013 26 32 6 0 6.26 KAMS_R014 58 63 5 0 2.32 KAMS_R015 114 115 1 0 1.12 KAMS_R016 104 105 1 0 2.98 KAMS_R017 10 12 2 0 1.6 KAMS_R018 31 32 1 0 2.16 KAMS_R019 59 60 1 0 2.87 KAMS_R020 77 80 3 0 1.52 KAMS_R021 105 114 9 0 1.44 KAMS_R022 1 12 11 0 2.75 KAMS_R023 29 37 8 0 1.76 KAMS_R024 58 59 1 0 1.66 KAMS_R025 81 83 2 0 1.73 KAMS_R026 110 115 5 0 1.19 KAMS_R027 138 139 1 0 1.19 KAMS_R032 79 83 4 0 2.67 KAMS_R033 104 109 5 0 2.8 KAMS_R034 127 132 5 0 1.11 kams1_3 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 327 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Kamoeb Mineralised Intersections Hole ID KAM_D043 KAM_D049 From To Thickness AuCy Au FA (m) (g/t) (g/t) 111.78 112.71 0.93 0 0 117.5 118.5 1 8.2 0 64.5 66.5 2 1.1 0 69 70 1 1.7 0 4.28 KAM_R391 2 8 6 0 KAMS_R007 25 26 1 0 1 29 30 1 0 1.07 KAMS_R014 KAMS_R016 103 105 2 0 5.38 108 109 1 0 1 119 121 2 0 1.11 130 131 1 0 1.29 136 137 1 0 0.95 kams1_4 KAM_D023 66.5 67 0.5 1.3 0 KAM_D085 29.58 30.62 1.04 0 0 KAM_D117 40.5 42 1.5 0.97 0 KAM_D118 58 60 2 3 0 KAM_D215 30.6 34.5 3.9 1.28 0 KAM_R393 47 50 3 0 3.32 KAM_R408 74 75 1 0 3.9 KAMS_R003 27 28 1 0 1.04 KAMS_R004 38 39 1 0 3 KAMS_R031 45 47 2 0 2.5 KAMS_R032 72 73 1 0 0.87 KAM_D031 24 25 1 1.4 0 KAM_D080 18.5 21.5 3 2.87 0 KAM_D081 35.5 39 3.5 6.29 0 KAM_D089 64.5 66.5 2 9.55 0 KAM_D151 31 33 2 2.5 0 KAM_D152 52.5 54 1.5 5.23 0 KAM_D153 80.5 82.5 2 3.2 0 KAM_D155 27.5 28 0.5 2.8 0 KAM_D160 89.7 91.7 2 1.75 0 KAM_D161 110.5 112 1.5 4.8 0 KAM_D162 7 12.5 5.5 4.48 0 KAM_D163 25.5 31 5.5 3.5 0 KAM_D164 62 66 4 2.55 0 kams1_5 kamw3_1 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 328 The Hassai Mine Envisaged Business Plan NI 43-101 Preliminary Assessment Report Kamoeb Mineralised Intersections Hole ID From To Thickness AuCy Au FA (m) (g/t) (g/t) KAM_D165 81 83 2 1.9 0 KAM_D166 28.3 33.5 5.2 0.95 0 KAM_D168 35 35.6 0.6 3.5 0 KAM_D169 45 51.15 6.15 2.36 0 0 KAM_D170 90 92 2 2.15 KAM_D171 100 103 3 2.37 0 KAM_D217 30 32 2 1.74 0 KAM_D218 48 50.5 2.5 2.36 0 KAM_D219 14 15 1 3.7 0 KAM_D220 34 37 3 1.5 0 KAM_D221 52.5 53.5 1 1.7 0 KAM_D222 24.5 24.9 0.4 2.9 0 KAM_D225 39.3 42 2.7 1 0 KAM_D226 31 32.5 1.5 1.8 0 KAM_D227 40 43 3 2.37 0 KAM_D229 26.8 27.2 0.4 0.8 0 KAM_D230 29.5 30 0.5 1.1 0 KAM_D231 52 53 1 3.3 0 KAM_D232 32 33 1 5.3 0 KAM_D233 46.5 48 1.5 1.53 0 FINAL – Rev 0 – 22 Oct 2010 AMEC Page 329
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