The Hassai Mine Envisaged Business Plan Red Sea State, Sudan

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
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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
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The Hassai Mine Envisaged Business Plan
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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
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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
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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
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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
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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
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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
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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
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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
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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
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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
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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
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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
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The Hassai Mine Envisaged Business Plan
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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
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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
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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.
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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
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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%
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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.
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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
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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.
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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
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•
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.
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•
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.
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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.
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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.
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−
•
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
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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.
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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.
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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.
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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.
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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
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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 ([email protected]/t)
-
4.6 @ 2.1
-
Indicated Resources ([email protected]/t)
-
6.8 @ 1.7
29.4 @ 1.1, 1.2
Mineral Reserves
Probable Reserves ([email protected]/t)
Inferred Resources ([email protected]/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.
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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.
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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
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•
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.
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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.
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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.
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•
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
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•
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.
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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
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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.
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Figure 4.1
Location of the Hassai Project
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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
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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'
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Figure 4.3
Sudan Gold and Iron Concessions Map
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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"
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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"
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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.
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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
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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.
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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.
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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.
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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.
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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.
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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.
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Figure 8.1
Diagrammatic Cross-section Showing Relationship of Ariab Deposits
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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)
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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.
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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.
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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).
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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
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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.
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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
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The Hassai Mine Envisaged Business Plan
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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.
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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).
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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.
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Figure 11.2
Cross-section Through Hasai South Showing Relationship Between Intersected and True Thickness
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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.
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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
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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.
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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.
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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
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•
•
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.
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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.
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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
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•
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
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•
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.
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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.
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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.
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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.
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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
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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
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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
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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.
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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.
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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.
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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
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−
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.
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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.
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15.
ADJACENT PROPERTIES
There are no adjacent properties relevant to this report.
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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.
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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.
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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.
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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.
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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
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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.
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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
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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
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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.
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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.
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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:
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•
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.
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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.
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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.
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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.
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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
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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.
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•
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.
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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.
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•
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
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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
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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.
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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
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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.
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Figure 16.3
5 Mt/a Block Flow Diagram
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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.
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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
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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.
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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
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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.
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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.
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The Hassai Mine Envisaged Business Plan
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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
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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.
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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.
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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.
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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
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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
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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 (%)
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Lower Supergene Limit
The Hassai Mine Envisaged Business Plan
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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.
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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.
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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.
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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.
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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.
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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.
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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
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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
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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.
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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.
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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.
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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
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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
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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
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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
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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
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−
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.
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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.
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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.
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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.
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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.
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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.
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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).
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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.
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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.
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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
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•
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.
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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.
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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$)
Whittle Schedule Parameters
Ore production rate
Annual discount rate
Mtpa
0.526
0.520
0.524
0.523
0.523
%
12.4
12.4
12.4
12.4
12.4
QRZ line
3,027,786
3,027,786
3,027,786
3,027,786
Other
6,199,752
6,199,752
6,199,752
6,199,752
Annual Quarts ore production
Pit slopes
degrees
38
38
38
38
38
Mining dilution
%
20.0
20.0
20.0
20.0
20.0
Mining recovery
%
95.0
95.0
95.0
95.0
95.0
525,000
Allocated general fixed expenses per tonne ore processed
QRZ line
5.77
5.77
5.77
5.77
Metallurgical recoveries %
Units
All rock
All rock
All rock
All rock
All rock
types
types
types
types
types
Gold recovery - Oxidised & Fresh (Au>2.25g/t)
%
80.00
80.00
80.00
92.40
92.40
Gold recovery - Oxidised & Fresh (Au<2.25g/t)
%
0.00
0.00
0.00
92.40
92.40
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Table 18.2
Kamoeb North Whittle Input Parameters
Ariab Mining Company ‐ Kamoeb Project Pit Optimisation Input Data – Kamoeb North Deposit Note: All costs & prices in US$ 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
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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.
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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.
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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.
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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
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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.
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−
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.
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Figure 18.3
Kamoeb South – Pit Design – Plan view
Figure 18.4
Kamoeb North – Pit Design – Plan view
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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
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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
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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
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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)
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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.
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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
$/t
20.07
21.87
19.96
19.94
20.54
20.09
19.99
18.30
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18.2.10
Mine Capital Costs
Due to the fact that the Hassai gold mine has been in operation since 1992, mine infrastructure is
already in place and a large fleet of major and minor mining equipment is owned by the company and
available.
18.2.10.1
Mining Infrastructure
All mining infrastructure is already in place and thus no capital expenditure is required to construct mine
production infrastructure.
Facilities already established for the open pit mine include:
•
Offices at the pit and in the central mine complex
•
Workshops at the pit and in the central mine complex
•
Service bays together with standing areas for mobile equipment at the pit and in the central mine
complex
•
Fuel Storage and refuelling areas for mobile equipment at the pit and in the central mine complex
•
Explosive storage facilities
•
Electrical supply infrastructure.
18.2.10.2
Mining Equipment
As the mine currently owns a large fleet of mining equipment, there is no need for any initial mining
equipment capital expenditure. As production increases to accommodate the CIL plant capacity and
the mining equipment ages, replacement capital will be required to augment the mining fleet.
It should be noted that no significant additional costs have been allowed for landing new equipment in
Sudan (ie. import taxes, duties, etc.). This needs to be investigated in the next level of study.
Table 18.12 outlines the replacement capital cost data.
Table 18.12
Replacement Capital Cost Summary – Kamoeb Open Pits
Description
Assumptions
Total Cost
($’000)
Mining Equipment
Replacement Capital for Major and minor mining
8 829
equipment
Total Replacement Capital Cost
8 829
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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.
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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
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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
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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.
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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%
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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.
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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
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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.
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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.
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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 $900.00/oz for gold.
A summary of the Whittle input parameters used in pit optimisation is provided in Table 18.17.
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Table 18.17
Whittle Input Parameters
ITEM
VALUE
SOURCE
DISCOUNT RATE
Discount Rate
%
10
La Mancha
% of NSR
% of NSR
% of NSR
7.90
La Mancha
Unit
US$/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
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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
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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.
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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.
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Figure 18.8
Hassai South – 5 Mt/a Optimisation Shell
Figure 18.9
Hadal Awatib – 5 Mt/a Optimisation Shell
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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
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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.
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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.
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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.
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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
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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.
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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.
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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.
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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.
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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.
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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
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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
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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
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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
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The Hassai Mine Envisaged Business Plan
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•
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
($/t)
2.47
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18.4.10
Mine Capital Costs
As the underground operating costs have been based around typical contract rates, equipment capital
recovery is built into the contractor rates, and, as such, no allowance is made for mining fleet
purchase/sustaining capital. Costs associated with underground infrastructure and the capital
development (such as decline development) have been included as capital items.
For open pit mining, it has been assumed that an entirely new mining fleet will be purchased for the
VMS open pit mining, and that this will be operated by the existing labour force.
It should be noted that no significant additional costs have been allowed for landing new equipment in
Sudan (ie. import taxes, duties, etc.). This needs to be investigated in the next level of study.
18.4.10.1
Hassai South
The capital cost summary for Hassai South underground is outlined in Table 18.31.
Table 18.31
Capital Cost Summary – Hassai South Underground
VMS Ore Mined
(Mt)
12.5
Capital Costs
Infrastructure
($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
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−
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
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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.
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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.
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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.
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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
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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.
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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
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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.
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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.
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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.
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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.
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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)
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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
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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.
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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
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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.
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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.
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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.
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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.
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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.
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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.
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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.
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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.
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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.
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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.
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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.
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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
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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.
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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
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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.
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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.
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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.
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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.
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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.
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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.
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•
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
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•
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.
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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
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•
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.
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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
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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.
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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.
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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.
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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
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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
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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.
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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
$/oz Au
250.71
18.14.1.2
Labour
A summary of estimated labour costs is provided in Table 18.59. Labour costs are based on recent
rates from Botswana, which are expected to be similar to those in Sudan. A total plant workforce of 192
has been estimated, excluding both mining and administration personnel. The workforce is comprised
mainly of local labour, with expatriates initially used for senior management roles.
18.14.1.3
Operating Consumables
Identified consumables and their anticipated consumption rates are given in Table 18.60. The analysis
assumes the exhaustion of fresh ore supplies at the end of operating year 4 and with it the cessation of
crushing circuit operations.
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Table 18.59
Annual Processing Plant Labour Costs, 3 Mt/a CIL Plant
Position
Number
Total Cost
Total Cost
($/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
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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 $0.40/L.
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The installed operating equipment power demand was determined from vendor data or first principal
estimates. Utilisation and load factors were then applied to the total to obtain power usage.
Table 18.62 summarises the annual power costs for the Hassai CIL plant and associated services.
Table 18.62
Estimated Annual Power Costs, 3 Mt/a CIL Plant
Description
Estimated Power
Estimated Cost
Estimated LOM Cost
(MWh/a)
($’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.
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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
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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.
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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.
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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.
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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.
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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.
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•
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.
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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 ([email protected]/t)
-
3.8 @ 1.9
-
Indicated Resources ([email protected]/t)
-
4.6 @ 2.1
-
Inferred Resources ([email protected]/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 ([email protected]/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
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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
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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
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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.
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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.
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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.
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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.
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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
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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%
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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.
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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.
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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%
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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.
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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.
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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.
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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.
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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.
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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
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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.
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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.
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Figure 20.4
Diagrammatic Representation of AMC’s Mining Flow Sheet
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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.
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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.
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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.
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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.
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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.
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Figure 20.6
AMC Gold Plant: Crushing/Milling Section Flow Sheet
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Figure 20.7
AMC Gold Plant: Leaching Section Flow Sheet
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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.
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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.
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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%
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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
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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.
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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.
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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.
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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
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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
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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
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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
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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
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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
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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
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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
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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
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25.
APPENDIX 1
RECENT HASSAI SOUTH, HADAL AWATIB AND
KAMOEB DRILL INTERSECTIONS
FINAL – Rev 0 – 22 Oct 2010
AMEC
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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
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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
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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
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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
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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
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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
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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
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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
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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
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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
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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
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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
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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
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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
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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