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30 March 2012
KATANGA MINING LIMITED
An Independent Technical Report on the Material Assets of Katanga Mining Limited, Katanga Province, Democratic Republic of Congo
REPO
RT
Qualified Person Company Name Qualification
Willem van der Schyff Golder Associates Africa Pri Sci Nat, BSc Hons (Geo), GDE (Mining Eng), GSSA
Report Number. 11613674-11195-1
Distribution: x1 Katanga Mining Limited x1 GAA Library x1 GAA Project File
Submitted to: Katanga Mining Limited Suite 300 204 Black Street Whitehorse Y1A 2M9 Yukon Territory Canada
KML - INDEPENDENT TECHNICAL REPORT (NI 43-101)
30 March 2012 Report No. 11613674-11195-1 i
To: Ontario Securities Commission, as Principal Regulator British Columbia Securities Commission Alberta Securities Commission Saskatchewan Financial Services Commission – Securities Division The Manitoba Securities Commission Autorité des Marchés Financiers Nova Scotia Securities Commission Prince Edward Island, Office of the Attorney General Securities Division New Brunswick, Securities Administration Branch Securities Commission of Newfoundland and Labrador Yukon, Registrar of Securities Northwest Territories, Registrar of Securities Nunavut, Registrar of Securities The Toronto Stock Exchange Katanga Mining Limited Dear Sirs/ Mesdames: Re: Katanga Mining Limited - Technical Report I, Willem van der Schyff, PriSciNat, (Registration Number 400176/05), am a qualified person (as such term is defined under National Instrument 43-101 – Standards of Disclosure for Mineral Projects ( NI 43-101 )) responsible for preparing or supervising the preparation of the entire technical report entitled An Independent Technical Report on the Material Assets of Katanga Mining Limited, Katanga Province, Democratic Republic of Congo dated March 30, 2012 (the Technical Report ) prepared for Katanga Mining Limited (the Corporation ). Pursuant to section 8.3 of NI 43-101, I hereby consent to the Corporation's public filing of the Technical Report and extracts from or summaries of the Technical Report, including extracts from or summaries of the Technical Report contained in the Corporation's Annual Information Form dated March 30, 2012 (the Annual Information Form). I confirm that I have read the Annual Information Form and that it fairly and accurately represents the information contained in the Technical Report that supports the disclosure. Dated this 30th day of March, 2012 /s/ Willem van der Schyff ___________________ Willem van der Schyff
Golder Associates Africa (Pty) Ltd.
Tel: Fax: www.golder.com
Golder Associates: Operations in Africa, Asia, Australasia, Europe, North America and South America
Reg. No. 2002/007104/07 Directors: FR Sutherland, AM van Niekerk, SAP Brown, L Greyling
Golder, Golder Associates and the GA globe design are trademarks of Golder Associates Corporation.
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Table of Contents
1.0 SUMMARY ............................................................................................................................................................... 16
1.1 Introduction ................................................................................................................................................. 16
1.2 Property Description and Location .............................................................................................................. 16
1.3 Ownership .................................................................................................................................................. 16
1.4 Geology and Mineralization ........................................................................................................................ 16
1.4.1 Geology ................................................................................................................................................. 16
1.4.2 Mineralisation ........................................................................................................................................ 17
1.5 Status of Material Assets ............................................................................................................................ 18
1.6 Mineral Resources and Ore Reserves ........................................................................................................ 18
1.6.1 Mineral Resources ................................................................................................................................ 18
1.6.1.1 Comparison of the 2011 and 2010 Mineral Resources ...................................................................... 19
1.6.2 Ore Reserve Estimate ........................................................................................................................... 20
1.6.3 Reconciliation with December 2010 Ore Reserve Estimate .................................................................. 21
1.7 Development and Operations ..................................................................................................................... 22
1.8 Interpretations and Conclusions ................................................................................................................. 22
1.9 Recommendations ...................................................................................................................................... 23
1.10 Economic Analysis ...................................................................................................................................... 23
2.0 INTRODUCTION ...................................................................................................................................................... 23
2.1 Qualified Person ......................................................................................................................................... 23
2.2 Site Visits .................................................................................................................................................... 24
2.3 Basis of Technical Report ........................................................................................................................... 24
2.4 Terms of Reference .................................................................................................................................... 24
3.0 RELIANCE ON OTHER EXPERTS ......................................................................................................................... 24
4.0 PROPERTY DESCRIPTION AND LOCATION ........................................................................................................ 26
4.1 KCC Rights: Summary of Permits ............................................................................................................... 26
4.2 KCC Rights: Joint Venture Agreements ...................................................................................................... 26
4.3 Replacement Reserves .............................................................................................................................. 28
4.4 Property Boundaries ................................................................................................................................... 28
4.5 Royalties, Duties and Other Fees ............................................................................................................... 30
4.5.1 Royalties payable to the State .............................................................................................................. 30
4.5.2 Surface Rights Fees payable to the State ............................................................................................. 30
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4.5.3 Royalties payable to Gécamines ........................................................................................................... 30
4.5.4 Pas de Porte payable to Gécamines ..................................................................................................... 30
4.5.5 Customs Duties and Taxes Payable ..................................................................................................... 30
5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY ..................... 30
6.0 HISTORY ................................................................................................................................................................. 31
7.0 GEOLOGICAL SETTING ......................................................................................................................................... 31
7.1 General Geology ......................................................................................................................................... 31
7.2 General Stratigraphy ................................................................................................................................... 31
7.3 Local Geology and Geological Models ....................................................................................................... 31
7.3.1 KTO Mine .............................................................................................................................................. 31
7.3.2 T-17 Open Pit ........................................................................................................................................ 33
7.3.3 KOV Open Pit ....................................................................................................................................... 33
7.3.4 Mashamba East Open Pit ..................................................................................................................... 34
7.3.5 Kananga Mine ....................................................................................................................................... 35
7.3.6 Tilwezembe Open Pit ............................................................................................................................ 35
7.3.7 Deposit Types ....................................................................................................................................... 36
8.0 MINERALISATION .................................................................................................................................................. 36
8.1 Mineral Resource Estimation Methodology ................................................................................................ 37
9.0 EXPLORATION AND DATA .................................................................................................................................... 37
9.1 Exploratory Data Analysis ........................................................................................................................... 38
9.1.1 KTO Mine .............................................................................................................................................. 38
9.1.2 T-17 Open Pit ........................................................................................................................................ 39
9.1.3 KOV Open Pit ....................................................................................................................................... 39
9.1.4 Mashamba East Open Pit ..................................................................................................................... 41
9.1.5 Kananga Mine ....................................................................................................................................... 41
9.1.6 Tilwezembe Open Pit ............................................................................................................................ 42
9.2 Variography Analysis .................................................................................................................................. 42
9.3 Estimation Parameters ............................................................................................................................... 45
9.3.1 Estimation Plan ..................................................................................................................................... 45
9.3.1.1 Density Assignment and Determinations ........................................................................................... 46
9.3.1.2 KTO Mine .......................................................................................................................................... 46
9.3.1.3 T-17 Open Pit .................................................................................................................................... 47
9.3.1.4 KOV Open Pit .................................................................................................................................... 48
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9.3.1.5 Mashamba East Open Pit .................................................................................................................. 48
9.3.1.6 Kananga Mine ................................................................................................................................... 49
9.3.1.7 Tilwezembe Open Pit ......................................................................................................................... 49
9.4 Summary .................................................................................................................................................... 50
10.0 DRILLING ................................................................................................................................................................ 50
10.1 KTO Mine ................................................................................................................................................... 50
10.2 T-17 Open Pit ............................................................................................................................................. 51
10.3 KOV Open Pit ............................................................................................................................................. 52
10.4 Mashamba East Open Pit ........................................................................................................................... 54
10.5 Kananga Mine............................................................................................................................................. 55
10.6 Tilwezembe Open Pit .................................................................................................................................. 55
11.0 SAMPLING METHOD AND APPROACH ................................................................................................................ 55
11.1 Historical Sampling ..................................................................................................................................... 55
11.2 Current Sampling ........................................................................................................................................ 56
11.2.1 Data Quality and Quantity ..................................................................................................................... 56
11.2.2 Sampling Preparation and Analyses ..................................................................................................... 57
11.2.2.1 Historical Sample Preparation and Analyses ..................................................................................... 57
11.2.2.2 Recent Sample Preparation and Analyses ........................................................................................ 57
11.3 Current Sample Preparation and Analyses ................................................................................................. 58
12.0 DATA VERIFICATION ............................................................................................................................................. 58
12.1 Quality Assurance and Quality Control ....................................................................................................... 58
12.2 QAQC Procedure ........................................................................................................................................ 58
13.0 MINERAL PROCESSING AND METALLURGICAL TESTING ............................................................................... 59
14.0 MINERAL RESOURCE ESTIMATES ...................................................................................................................... 59
14.1 Mineral Resource Statement ...................................................................................................................... 60
14.2 Comparison of the 2011 and 2010 Mineral Resources ............................................................................... 61
15.0 ORE RESERVE ESTIMATES .................................................................................................................................. 63
15.1 Overview of Mining Operations ................................................................................................................... 63
15.2 Development and Operations ..................................................................................................................... 66
15.3 Underground ............................................................................................................................................... 66
15.4 Kamoto Underground Life of Mine Plan - Previous Technical Study .......................................................... 67
15.4.1 Background ........................................................................................................................................... 67
15.4.2 Strategic Approach................................................................................................................................ 68
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15.4.3 Design Constraints Applicable to KTO .................................................................................................. 68
15.4.4 RAP General Layout – Previous Technical Studies .............................................................................. 69
15.4.5 Long Hole Retreat Stoping General Layout – Previous Technical Studies ........................................... 70
15.4.6 Cut-and-Fill Mining Methods and Variations – Previous Technical Studies .......................................... 70
15.4.7 Sub-Level Caving General Layout – Previous Technical Studies ......................................................... 71
15.5 Kamoto Underground Mining Strategic Update .......................................................................................... 71
15.5.1 Room and Pillar .................................................................................................................................... 71
15.5.2 Cut and Fill Transverse ......................................................................................................................... 72
15.5.3 Cut and Fill Longitudinal ........................................................................................................................ 74
15.5.4 Future Reserve Updates ....................................................................................................................... 74
15.5.5 Production Scheduling .......................................................................................................................... 75
15.5.6 KTO LOM Plan reconciliation ................................................................................................................ 76
15.6 T-17 Underground ...................................................................................................................................... 77
15.6.1 Background ........................................................................................................................................... 77
15.6.2 Mining Strategy ..................................................................................................................................... 77
15.6.3 Future Ore Reserve Updates ................................................................................................................ 78
15.6.4 Production Scheduling .......................................................................................................................... 78
15.6.5 T-17 underground LOM Plan reconciliation ........................................................................................... 81
15.7 Kamoto East underground .......................................................................................................................... 82
15.7.1 Background ........................................................................................................................................... 82
15.7.2 KTE LOM Plan Reconciliation ............................................................................................................... 85
15.8 Modifying Factors ....................................................................................................................................... 86
15.8.1 KTO Underground ................................................................................................................................. 86
15.8.2 T-17 Underground ................................................................................................................................. 87
15.8.3 KTE Underground ................................................................................................................................. 87
15.9 Ore Reserve Estimate ................................................................................................................................ 88
15.9.1 Ore Reserve Estimate – KCC Underground Operations ....................................................................... 88
15.10 Ore Reserve Estimate ................................................................................................................................ 88
15.10.1 Ore Reserve Estimate as at December 31, 2011 .................................................................................. 88
15.11 Reconciliation with December 2010 Ore Reserve Estimate ....................................................................... 89
16.0 MINING METHODS ................................................................................................................................................. 90
16.1 LOM Planning Process ............................................................................................................................... 90
16.1.1 Overview ............................................................................................................................................... 90
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16.1.2 Mining Model ......................................................................................................................................... 91
16.1.2.1 Dilution ............................................................................................................................................... 91
16.1.2.2 Mining Loss ....................................................................................................................................... 91
16.1.3 Pit Optimisation ..................................................................................................................................... 91
16.1.4 Pit Design .............................................................................................................................................. 92
16.1.5 Scheduling Units ................................................................................................................................... 92
16.1.6 Production Scheduling .......................................................................................................................... 92
16.2 T-17 Open Pit ............................................................................................................................................. 92
16.2.1 Modifying Factors .................................................................................................................................. 94
16.2.2 Pit Optimisation ..................................................................................................................................... 94
16.2.3 Pit Design .............................................................................................................................................. 94
16.2.4 Production Scheduling .......................................................................................................................... 96
16.2.5 T-17 Extension LOM Plan Reconciliation .............................................................................................. 98
16.3 KOV Open Pit ............................................................................................................................................. 98
16.3.1 Background ........................................................................................................................................... 98
16.3.2 Modifying Factors .................................................................................................................................. 98
16.3.3 Pit Optimisation ..................................................................................................................................... 99
16.3.4 Pit Design .............................................................................................................................................. 99
16.3.5 Production Scheduling ........................................................................................................................ 102
16.3.6 KOV LOMP Reconciliation .................................................................................................................. 104
16.4 Mashamba East ........................................................................................................................................ 105
16.4.1 Background ......................................................................................................................................... 105
16.4.2 Modifying factors ................................................................................................................................. 105
16.4.3 Pit Optimisation ................................................................................................................................... 106
16.4.4 Pit design ............................................................................................................................................ 106
16.4.5 Production scheduling ......................................................................................................................... 107
16.4.6 Mashamba East LOM reconciliation.................................................................................................... 110
16.5 KCC Open Pit Ore Reserve Statement ..................................................................................................... 110
16.6 Life of Mine Plan ....................................................................................................................................... 111
16.7 Risk ........................................................................................................................................................... 113
16.7.1 KTO ..................................................................................................................................................... 113
16.7.2 KTE ..................................................................................................................................................... 113
16.7.3 T-17 ..................................................................................................................................................... 113
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16.8 Surface Operations ................................................................................................................................... 113
16.9 General ..................................................................................................................................................... 114
17.0 RECOVERY METHODS: PLANT AND PROCESSING......................................................................................... 114
17.1 General Process Commentary.................................................................................................................. 114
17.2 Status Update ........................................................................................................................................... 114
17.3 Processing Facilities ................................................................................................................................. 114
17.3.1 KTC Original Facilities ......................................................................................................................... 114
17.3.2 KTC Current Operations ..................................................................................................................... 115
17.3.2.1 Oxide Ore Crushing and Milling ....................................................................................................... 115
17.3.2.2 Sulphide Ore Crushing and Milling .................................................................................................. 115
17.3.2.3 Oxide Flotation................................................................................................................................. 115
17.3.2.4 Sulphide Flotation ............................................................................................................................ 116
17.3.3 KTC Project Development ................................................................................................................... 116
17.3.3.1 Crushing, Materials Handling and Ore Storage ............................................................................... 116
17.3.3.2 Milling Circuits.................................................................................................................................. 116
17.3.3.3 Flotation Circuits .............................................................................................................................. 117
17.3.3.4 Concentrate Handling ...................................................................................................................... 117
17.3.4 Luilu Refinery Original Facilities .......................................................................................................... 119
17.3.5 Luilu Refinery: Current Operations ...................................................................................................... 120
17.3.5.1 Concentrate Reception .................................................................................................................... 120
17.3.5.2 Sulphide Concentrate ...................................................................................................................... 120
17.3.5.3 Oxide Concentrate ........................................................................................................................... 120
17.3.5.4 Leach Circuits .................................................................................................................................. 120
17.3.5.5 Copper Circuits ................................................................................................................................ 120
17.3.5.6 Cobalt Circuit ................................................................................................................................... 121
17.3.6 Luilu Copper Electro-Refinery ............................................................................................................. 121
17.3.6.1 Sulphide Concentrate ...................................................................................................................... 121
17.3.6.2 Oxide Concentrates ......................................................................................................................... 121
17.3.6.3 Sulphide Concentrate Roasting ....................................................................................................... 121
17.3.6.4 Primary Leaching ............................................................................................................................. 121
17.3.6.5 Thickening and Residue Handling ................................................................................................... 122
17.3.6.6 Copper Solvent Extraction ............................................................................................................... 122
17.3.6.7 Electro winning ................................................................................................................................ 122
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17.3.6.8 Iron and Aluminium Precipitation ..................................................................................................... 123
17.3.6.9 Copper Precipitation ........................................................................................................................ 123
17.3.6.10 Zinc Solvent Extraction .................................................................................................................... 123
17.3.6.11 Cobalt Recovery .............................................................................................................................. 123
18.0 PROJECT INFRASTRUCTURE ............................................................................................................................ 125
19.0 MARKET STUDIES AND CONTRACTS ............................................................................................................... 125
19.1 Markets ..................................................................................................................................................... 125
19.1.1 Copper ................................................................................................................................................ 125
19.1.2 Cobalt .................................................................................................................................................. 126
19.2 Contracts .................................................................................................................................................. 128
20.0 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL COMMUNITY IMPACT .......................................... 128
20.1 Terms of Reference .................................................................................................................................. 128
20.1.1 Regulatory Context ............................................................................................................................. 128
20.1.2 The Equator Principles ........................................................................................................................ 129
20.1.3 The International Finance Corporation Performance Standards ......................................................... 130
20.1.4 The World Bank Group Environmental Health and Safety (EHS) Guidelines ...................................... 130
20.1.5 KCC Environmental Policy .................................................................................................................. 130
20.1.6 Information Sources Reviewed ........................................................................................................... 131
20.1.6.1 Site Areas visited ............................................................................................................................. 131
20.1.7 Limitations of the Review .................................................................................................................... 132
20.1.8 Results of Review ............................................................................................................................... 132
20.1.8.1 Authorisations and Operating Licences ........................................................................................... 132
20.1.8.2 EIS and EMP ................................................................................................................................... 132
20.1.9 Key Environmental Risks Relating to KCC Operations ....................................................................... 133
20.1.9.1 Critical Risk ...................................................................................................................................... 133
20.1.10 Overview ............................................................................................................................................. 133
20.1.11 Impact of the Tailings Discharge ......................................................................................................... 134
20.1.12 KCC Remediation Measures ............................................................................................................... 134
20.2 Moderate Risk........................................................................................................................................... 134
20.2.1 Community Health and Safety – Site Access ...................................................................................... 134
20.2.2 General ............................................................................................................................................... 134
21.0 CAPITAL AND OPERATING COSTS ................................................................................................................... 134
21.1 Capital Cost Estimates ............................................................................................................................. 134
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21.2 Operating Cost Estimates ......................................................................................................................... 137
22.0 ECONOMIC ANALYSIS ........................................................................................................................................ 138
22.1 Principal Assumptions .............................................................................................................................. 138
22.2 Cash flow forecasts .................................................................................................................................. 139
22.3 Valuation of KCC ...................................................................................................................................... 142
22.4 Taxation, Royalties and Other Interests .................................................................................................... 142
22.5 Sensitivity Analysis ................................................................................................................................... 143
22.6 Risk Analysis ............................................................................................................................................ 144
22.7 Mining Risks ............................................................................................................................................. 144
22.8 Processing risks ........................................................................................................................................ 145
22.9 Capital risks .............................................................................................................................................. 145
22.9.1 Escalation of Costs ............................................................................................................................. 145
22.10 Operating Risks ........................................................................................................................................ 146
22.10.1 Poor Condition of Railways ................................................................................................................. 146
22.10.2 Underdeveloped In-Country Institutional Infrastructure and Capacity ................................................. 146
22.10.3 Senior Management and Technical Expertise ..................................................................................... 146
22.10.4 Artisanal Miners .................................................................................................................................. 146
22.11 Sovereign risk ........................................................................................................................................... 146
22.12 Economic and Market Risks ..................................................................................................................... 147
22.12.1 Commodity Prices: .............................................................................................................................. 147
22.12.2 Operating Costs: ................................................................................................................................. 147
22.12.3 Currency Risk: .................................................................................................................................... 147
22.13 Environmental and Social Risks ............................................................................................................... 147
22.13.1 Tailings Slurry Discharge to the Luilu River: ....................................................................................... 147
22.13.2 Community Health and Safety – Site Access: ..................................................................................... 147
22.13.3 Storm Water Management: ................................................................................................................. 147
22.13.4 Other Risks: ........................................................................................................................................ 147
23.0 ADJACENT PROPERTIES .................................................................................................................................... 148
24.0 OTHER RELEVANT DATA AND INFORMATION ................................................................................................ 148
25.0 TAILINGS AND WASTE ........................................................................................................................................ 148
25.1 Location Description ................................................................................................................................. 148
25.2 Description ................................................................................................................................................ 148
26.0 CLOSURE .............................................................................................................................................................. 148
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26.1 Approach and Limitations to Closure Cost Update ................................................................................... 148
26.1.1 Approach ............................................................................................................................................. 148
26.1.2 Limitations ........................................................................................................................................... 149
26.1.3 Available Information ........................................................................................................................... 150
26.1.4 Battery Limits ...................................................................................................................................... 150
26.1.4.1 Available Information ....................................................................................................................... 150
26.1.4.2 General Additions by GAA ............................................................................................................... 151
26.1.5 Assumptions and Qualifications .......................................................................................................... 152
26.1.5.1 General ............................................................................................................................................ 152
26.1.5.2 Site Specific ..................................................................................................................................... 153
26.1.5.3 Additional Allowances ...................................................................................................................... 155
26.1.6 Unit Rates ........................................................................................................................................... 155
26.1.6.1 General Surface Shaping ................................................................................................................ 155
26.1.6.2 Rehabilitation of Evaporation Ponds ................................................................................................ 155
26.1.6.3 Establishing of Vegetation (General) ............................................................................................... 156
26.1.6.4 Vegetation Establishment on Tailings .............................................................................................. 156
26.1.6.5 Vegetation Establishment on Waste Rock Dumps .......................................................................... 156
26.1.6.6 Creation of Enviro-Bund .................................................................................................................. 156
26.1.6.7 Groundwater Monitoring .................................................................................................................. 156
26.1.6.8 Surface Water Monitoring ................................................................................................................ 157
26.1.6.9 Rehabilitation Monitoring ................................................................................................................. 157
26.1.6.9.1 High Intensity Rehabilitation Monitoring ....................................................................................... 157
26.1.6.9.2 Low Intensity Rehabilitation Monitoring ........................................................................................ 157
26.1.6.10 Rehabilitation Care and Maintenance .............................................................................................. 157
26.1.6.10.1 High Intensity Care and Maintenance .............................................................................................. 157
26.1.6.10.2 Low Intensity Care and Maintenance ............................................................................................... 158
26.1.7 Closure Cost Assessment ................................................................................................................... 158
26.1.7.1 Infrastructural Areas ........................................................................................................................ 159
26.1.7.2 Mining Areas .................................................................................................................................... 160
26.1.7.3 General Surface Reclamation .......................................................................................................... 162
26.1.7.4 Water Management ......................................................................................................................... 162
26.1.7.5 Post-closure Aspects ....................................................................................................................... 163
26.1.7.6 Additional Allowances ...................................................................................................................... 163
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26.1.8 Summary of Update Closure Costs ..................................................................................................... 165
26.1.9 Matters requiring further attention ....................................................................................................... 165
26.1.10 Conclusion .......................................................................................................................................... 166
27.0 INTERPRETATIONS AND CONCLUSIONS ......................................................................................................... 166
28.0 RECOMMENDATIONS .......................................................................................................................................... 166
28.1 Dewatering................................................................................................................................................ 167
28.2 Tailings and Waste ................................................................................................................................... 167
29.0 REFERENCES ....................................................................................................................................................... 168
30.0 DATE AND SIGNATURE PAGES ......................................................................................................................... 169
TABLES Table 1: Mining Assets ...................................................................................................................................................... 18
Table 2: Mineral Processing Assets .................................................................................................................................. 18
Table 3: KCC: Consolidated Mineral Resources as at 31 December 20111,2,3,4 ................................................................ 18
Table 4: Comparison of the 2011 and 2010 Mineral Resources1,2,3,4,5 .............................................................................. 19
Table 5: KCC Ore Reserve Estimate1,2,3 ........................................................................................................................... 20
Table 6: Ore Reserve Reconciliation1,2,3 ........................................................................................................................... 21
Table 7: Plant and Processing Developments ................................................................................................................... 22
Table 8: NPV of KML’s holding in KCC ............................................................................................................................. 23
Table 9: Reliance on Other Experts .................................................................................................................................. 25
Table 10: Legal Tenure of KCC ......................................................................................................................................... 26
Table 11: Chronological developments for Material Assets............................................................................................... 31
Table 12: Independent Consultants .................................................................................................................................. 37
Table 13: Statistical Analysis of Kamoto Principal ............................................................................................................. 38
Table 14: Statistical Analysis of Etang South .................................................................................................................... 38
Table 15: Statistical Analysis of Etang North ..................................................................................................................... 38
Table 16: Statistical Analysis of T-17 Open Pit ................................................................................................................. 39
Table 17: Statistical Analysis of KOV Open Pit; Virgule .................................................................................................... 39
Table 18: Statistical Analysis of KOV Open Pit; Oliviera ................................................................................................... 40
Table 19: Statistical Analysis of KOV Open Pit; FNSR ..................................................................................................... 40
Table 20: Statistical Analysis of KOV Open Pit; Kamoto East ........................................................................................... 40
Table 21: Statistical Analysis of Mashamba East Open Pit ............................................................................................... 41
Table 22: Statistical Analysis of Kananga Mine ................................................................................................................. 41
Table 23: Statistical Analysis of Tilwezembe Open Pit ...................................................................................................... 42
Table 24: KTO Mine except Etang North: Omni-directional Variography Parameters ....................................................... 43
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Table 25: Etang North: Variography Parameters .............................................................................................................. 43
Table 26: T-17 Open Pit: Omni-directional Variography Parameters ................................................................................ 43
Table 27: KOV Open Pit: Omni-directional Variography Parameters for Virgule ............................................................... 44
Table 28: KOV Open Pit: Omni-directional Variography Parameters for Oliviera .............................................................. 44
Table 29: Mashamba East Open Pit: Omni-directional Variography Parameters .............................................................. 44
Table 30: Tilwezembe Open Pit: Variogram Parameters for Manganiferous Dolomites .................................................... 44
Table 31: Tilwezembe Open Pit: Variogram Parameters for Breccia ................................................................................ 45
Table 32: Tilwezembe Open Pit: Variogram Parameters for Tillites and Argillites ............................................................. 45
Table 33: Kananga Mine: Variogram Parameters for Upper Orebody ............................................................................... 45
Table 34: Kananga Mine: Variogram Parameters for Internal/Middle Zone....................................................................... 45
Table 35: GCM criteria for assigning density values ......................................................................................................... 46
Table 36: KTO Mine: Density Determinations on various lithologies ................................................................................. 47
Table 37: New Density Determinations for Etang North .................................................................................................... 47
Table 38: T-17 Open Pit: Density Determinations on Various Lithologies ......................................................................... 47
Table 39: New Density Determinations for T-17 ................................................................................................................ 48
Table 40: Mashamba East Open Pit: Density Determinations on Various Lithologies ...................................................... 48
Table 41: Kananga Mine: Density Determinations on Various Lithologies ........................................................................ 49
Table 42: Tilwezembe Open Pit: Density Determinations on Various Lithologies ............................................................. 49
Table 43: Kananga Mine: Density Determinations on Various Lithologies ........................................................................ 50
Table 44: Holes drilled in Etang North ............................................................................................................................... 51
Table 45: Summary of 2010 drill data for T-17 .................................................................................................................. 52
Table 46: Summary of 2010 drill data for Kamoto East ..................................................................................................... 53
Table 47: Summary of 2010 drill data for Virgule .............................................................................................................. 54
Table 48: Summary of 2010 drill data for Oliviera ............................................................................................................. 54
Table 49: Historical Core Recoveries per Area ................................................................................................................. 56
Table 50: KCC: Consolidated Mineral Resources as at 31 December 20111,2,3,4 .............................................................. 60
Table 51: Comparison of the 2011 and 2010 Mineral Resources as at December 31, 20111,2,3,4 ..................................... 61
Table 52: KCC Ore Reserve Estimate1,2,3 ......................................................................................................................... 66
Table 53: Summary of mining methods and previous studies ........................................................................................... 67
Table 54: KTO Life of Mine schedule ................................................................................................................................ 75
Table 55: KTO reconciliation ............................................................................................................................................. 76
Table 56: T-17 Underground LOM production ................................................................................................................... 80
Table 57: T-17 underground reconciliation ........................................................................................................................ 82
Table 58: KTE LOM production ......................................................................................................................................... 83
Table 59: KTE Reconciliation ............................................................................................................................................ 85
Table 60: KTO Mine modifying factors .............................................................................................................................. 86
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Table 61: KTO Mine: Modifying factors for the PPCF Method ........................................................................................... 87
Table 62: T-17 Underground modifying factors ................................................................................................................. 87
Table 63: KTE modifying factors ....................................................................................................................................... 88
Table 64: KTO Ore Reserves as at 31 December 20111,2,3 .............................................................................................. 88
Table 65: KCC Ore Reserve table1,2,3 ............................................................................................................................... 89
Table 66: Ore Reserve reconciliation1,2,3 ........................................................................................................................... 90
Table 67: Selected SMU dimensions per pit ..................................................................................................................... 91
Table 68: T-17 pit design criteria ....................................................................................................................................... 96
Table 69: T-17 extension LOM production profile ............................................................................................................. 96
Table 70: T-17 extension reconciliation ............................................................................................................................. 98
Table 71: KOV optimisation parameters............................................................................................................................ 99
Table 72: KOV pit design criteria ..................................................................................................................................... 101
Table 73: KOV LOM production profile ........................................................................................................................... 102
Table 74: KOV LOM Plan reconciliation .......................................................................................................................... 105
Table 75: Mashamba East pit optimisation parameters................................................................................................... 106
Table 76: Mashamba East Pit design criteria .................................................................................................................. 107
Table 77: Mashamba East LOM production profile ......................................................................................................... 107
Table 78: Mashamba East LOM reconciliation ................................................................................................................ 110
Table 79: Surface mining operations Ore Reserve table as at December 31, 20111,2,3 .................................................. 110
Table 80: Global refined copper market balance (Source: USGS 2006-2009, ICSG 2010-2012) ................................... 125
Table 81: Copper price forecast ...................................................................................................................................... 126
Table 82: Global refined cobalt market balance .............................................................................................................. 127
Table 83: Cobalt price forecast ....................................................................................................................................... 127
Table 84: Summary of relevant Environmental and Social DRC legislation relating to KCC ........................................... 128
Table 85: Capital Expenditure by operational area ......................................................................................................... 135
Table 86: Capital Expenditure by year ............................................................................................................................ 136
Table 87: Operating costs by area .................................................................................................................................. 137
Table 88: Major Operational Expenditure ........................................................................................................................ 138
Table 89: LOM Ore and Metals Volumes ........................................................................................................................ 138
Table 90: KCC cash flow for 2012 to 2021 ...................................................................................................................... 140
Table 91: KCC cash flow for 2022 to 2030 and LOM total .............................................................................................. 141
Table 92: Project Valuation ............................................................................................................................................. 142
Table 93: NPV of KML’s holding in KCC ......................................................................................................................... 142
Table 94: Royalty, tax and import duty assumptions ....................................................................................................... 142
Table 95: Sensitivity of the value of KML’s interest in KCC to changes in metal prices .................................................. 143
Table 96: Sensitivity of the value of KML’s interest in KCC to changes in operating costs ............................................. 143
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Table 97: Sensitivity of the value of KML’s interest in KCC to changes in capital expenditure........................................ 143
Table 98: Closure cost assessment ................................................................................................................................ 159
Table 99: Overall cost comparison .................................................................................................................................. 164
FIGURES Figure 1: Property Boundaries of KCC's Exploitation Permits ........................................................................................... 29
Figure 2: Plan view of KTO Mine Resource Model ............................................................................................................ 32
Figure 3: Oblique view of KTO Mine with Resource Model ............................................................................................... 32
Figure 4: Oblique view of T-17 Open Pit Resource Model ................................................................................................ 33
Figure 5: Oblique view of KOV Open Pit Resource Model ................................................................................................ 34
Figure 6: Oblique view of Mashamba East Resource Model and Planned Pit Layout ....................................................... 34
Figure 7: Oblique view of the Kananga Resource Model .................................................................................................. 35
Figure 8: Oblique view of Tilwezembe Pit with Resource Model and Pit Layout ............................................................... 35
Figure 9: KCC overall mining layout .................................................................................................................................. 63
Figure 10: KCC ROM Production profiles.......................................................................................................................... 65
Figure 11: KCC recovered Cu profiles 2011 ...................................................................................................................... 65
Figure 12: General configuration of mining areas at Kamoto underground mine .............................................................. 67
Figure 13: Room and Pillar mining layout (Zone 8) ........................................................................................................... 69
Figure 14: Cut and Fill longitudinal method ....................................................................................................................... 71
Figure 15: Room and Pillar with cemented backfill ............................................................................................................ 72
Figure 16: Schematic representation of Transversal Cut and Fill ...................................................................................... 73
Figure 17: Schematic representation of longitudinal cut and fill ........................................................................................ 74
Figure 18: KTO LOM Plan profile ...................................................................................................................................... 76
Figure 19: KTO development profile ................................................................................................................................. 76
Figure 20: T-17 current pit, showing open pit extensions and underground access .......................................................... 77
Figure 21: Conceptual design of T-17 underground design .............................................................................................. 78
Figure 22: T-17 underground appropriate design and schedule........................................................................................ 79
Figure 23: T-17 Underground LOM profile ........................................................................................................................ 80
Figure 24: T-17 Underground recovered Cu profile ........................................................................................................... 81
Figure 25: T-17 underground development profile ............................................................................................................ 81
Figure 26: KTE conceptual development layout ................................................................................................................ 82
Figure 27: Kamoto East Underground LOM profile ........................................................................................................... 84
Figure 28: KTE development requirements ....................................................................................................................... 84
Figure 29: KTE recovered Cu profile ................................................................................................................................. 85
Figure 30: Reserve contributions over time ....................................................................................................................... 89
Figure 31: T-17 Extension prerequisite construction work................................................................................................. 93
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Figure 32: T-17 extension with crusher protection and road and river diversions ............................................................. 94
Figure 33: T-17 open pit .................................................................................................................................................... 95
Figure 34: T-17 current pit survey showing section line A-A ............................................................................................. 95
Figure 35: Section line A-A of the T-17 pit extension relative to the 2011 T-17 pit survey ................................................ 96
Figure 36: T-17 extension tonnage and copper grade profile ............................................................................................ 97
Figure 37: T-17 ore type profile excluding T-17 extension ................................................................................................ 97
Figure 38: T-17 reserve profile excluding T-17 extension ................................................................................................. 97
Figure 39: KOV dilution curve for a 10x10x5 SMU ............................................................................................................ 99
Figure 40: KOV open pit .................................................................................................................................................. 100
Figure 41: KOV pit design relative to the end of 2011 pit survey ..................................................................................... 101
Figure 42: Section B-B through KOV indicating the ultimate pit versus the current pit .................................................... 101
Figure 43: KOV LOM profile ............................................................................................................................................ 102
Figure 44: KOV ore type profiles ..................................................................................................................................... 103
Figure 45: KOV Reserve profile ...................................................................................................................................... 103
Figure 46: KOV ultimate pit with mined out area and pit block designs ........................................................................... 104
Figure 47: Isometric view showing mined out and adjusted pit designs .......................................................................... 104
Figure 48: Mashamba East losses and dilutions ............................................................................................................. 106
Figure 49: Mashamba East pit design ............................................................................................................................. 107
Figure 50: Mashamba East LOM profile .......................................................................................................................... 108
Figure 51: Mashamba East ore type profiles ................................................................................................................... 109
Figure 52: Mashamba East Reserve profile .................................................................................................................... 109
Figure 53: Life of Mine plan ROM ore profile for the combined KCC operation............................................................... 111
Figure 54: LOM Plan waste profile for the combined KCC operation .............................................................................. 111
Figure 55: LOM recovered copper profile for the combined KCC operation .................................................................... 112
Figure 56: LOM Plan recovered cobalt profile for the combined KCC operation ............................................................. 112
Figure 57: KCC Filtration, Bagging and Storage Plant .................................................................................................... 116
Figure 58: The Kamoto Concentrator Phase 4 Block Flow Diagram ............................................................................... 118
Figure 59: Earthworks underway for the construction of the SX Plant ............................................................................ 122
Figure 60: Electrowinning plant under conversion ........................................................................................................... 123
Figure 61: The Luilu Phase 4 Block Flow Diagram ......................................................................................................... 124
Figure 62: The London Metal Exchange copper price from January 2007 to date (Source: LME) .................................. 126
Figure 63: The London Metal Exchange cobalt price from April 2010 to date (Source: LME) ......................................... 127
Figure 64: Site visit route................................................................................................................................................. 132
Figure 65: Sensitivity of the value of KML’s holding in KCC to changes in key variables ................................................ 144
Figure 66: Map of Infrastructure on perimeters of the convention AJVA - KCC .............................................................. 152
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APPENDICES APPENDIX A Document Limitations
APPENDIX B Abbreviations and Glossary of Terms
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1.0 SUMMARY 1.1 Introduction Golder Associates Africa (Pty) Ltd (GAA) was commissioned by Katanga Mining Limited (KML) to prepare this Independent Technical Report (ITR), which complies with the Canadian Securities Administrators' National Instrument 43-101 - Standards of Disclosure for Mineral Projects (NI 43-101), in respect of the Material Assets (as defined below) owned and operated by KML.
1.2 Property Description and Location This ITR covers the following operations, projects and infrastructure of KML and its subsidiaries located in the Kolwezi District in the Katanga Province of the Democratic Republic of Congo (DRC), which are collectively referred to herein as the Material Assets:
Mining Assets, namely:
Kamoto Underground Mine (KTO);
T-17 Open Pit and underground extension;
KOV Open Pit and underground extension (Kamoto East Underground Mine) (KTE);
Mashamba East Open Pit;
Tilwezembe Open Pit;
Kananga Mine; and
Extension of Kananga Mine.
Processing Assets , namely:
the Kamoto Concentrator (KTC); and
the Luilu Metallurgical Plant (Luilu Refinery).
Infrastructure necessary for the production of saleable metals.
1.3 Ownership Kamoto Copper Company SARL (KCC) owns the Material Assets, including the mining and exploitation rights related to the Mining Assets. KML holds a 75% stake in KCC. La Generale des Carrieres et des Mines (GCM) and La Société Immobilière du Congo (SIMCO), which are state-owned mining companies in the DRC, own the other 25% of KCC.
1.4 Geology and Mineralization 1.4.1 Geology The mineralized zones are at the western end of the Katangan Copperbelt, one of the great metallogenic provinces of the world, containing some of the world’s richest copper and cobalt deposits.
These deposits are hosted mainly by metasedimentary rocks of the late proterozoic Katangan system, a 7km thick succession of sediments with minor Volcanics, Volcanoclastics and intrusive rocks. Geochronological data indicates an age of deposition of the Katangan sediments of about 880 million years and deformation during the Katangan orogeny at less than 650 million years. This deformation resulted in the NS-SE trending Lufilian Arc, which extends from Namibia on the west coast of Africa through to Zambia, lying to the south of the DRC. Within the DRC, the zone of deformation extends for more than 300 km from Kolwezi in the north-west to Lubumbashi in the south-east.
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Stratigraphically, the rich copper and cobalt deposits found in Zambia and the DRC are localized in the Roan Supergroup (the Roan). The Roan occurs at the base of the Katanga succession, unconformably overlying the basement rock of Kibaran age (mid-proterozic). The Roan is separated from the overlying rocks of the Kundelungu and Nguba Supergroups by a conglomerate, the Grand Conglomerate. The Nguba (previously known as the Lower Kundelungu) is composed of sandstones and shales with a basal conglomerate, while the Kundelungu consists essentially of sediments and is separated from the Nguba by a conglomerate, the (French) ‘petit conglomerat’.
Within the Lufilian Arc are large-scale E-W to NW-SE trending folds with wavelengths extending for kilometres. The folds are faulted along the crests of the anticlines through which rocks of the Roan have been diapirically injected into the fault zones, squeezed up fault planes and over-thrust to lie above rocks of the younger Kundelungu. The over-thrust Roan lithologies occur as segments or “fragments” on surface. The fragments are intact units that preserve the original geological succession within each. A fragment could be hundreds of metres and aligned across the fault plane.
In the Katangan Copperbelt, mining for copper and cobalt occurs in these outcropping to sub-outcropping fragments.
1.4.2 Mineralisation Primary mineralisation, in the form of sulphides, within the Lower Roan is associated with the D Strat and RSF for the OBI and the SDB and SDS for the OBS and is thought to be syn-sedimentary in origin. Typical primary copper sulphide minerals are bornite, chalcopyrite, chalcosite and occasional native copper while cobalt is in the form of carrolite. The mineralization occurs as disseminations or in association with hydrothermal carbonate alteration and silicification.
Supergene mineralization is generally associated with the levels of oxidation in the sub-surface sometimes deeper than 100m below the surface. The most common secondary supergene minerals for copper and cobalt are malachite and heterogenite. Malachite is the main mineral mined within the confines of the current KOV Open Pit.
The RSC, a lithological unit stratigraphically intermediate between the upper and lower ore body host rocks, contains relatively less copper mineralization. The RSC contains appreciable copper mineralization near the contacts with the overlying SDB formation and the underlying RSF formations. The middle portion of the RSC, considered to be “sterile” by GCM, normally contains relatively less copper mineralization and is sometimes not sampled. The mineral potential of the RSC is less well known than that of other formations.
The RSC has been observed to be well mineralized in supergene cobalt hydroxide, heterogenite, which occurs as vug infillings, especially near the surface.
The mineralization at Tilwezembe Open Pit is atypical, being hosted by the Mwashya or R4 formation. The mineralization generally occurs as infilling of fissures and open fractures associated with the brecciation. The typical mineralization consists mainly of copper minerals (chalcopyrite, malachite and pseudomalachite), cobalt minerals (heterogenite, carrolite and spherocobaltite) and manganese (Mn) minerals (psilomelane and magnetite).
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1.5 Status of Material Assets Table 1 and Table 2 below provide certain details on the status of the Material Assets.
Table 1: Mining Assets
Property Type Status Licence
Comments Expiry Date Area
KTO and Mashamba East Open Pit
UG OP
Operating Development April 3, 2024 11,04 km2
KTO Operational Mashamba East in development, dewatering to commence in 2013
T-17 Open Pit OP Operating April 3, 2024 1,698 km2 Operational Mine KOV Open Pit OP Operating April 3, 2024 8,49 km2 Operational Mine
Tilwezembe Open Pit OP Dormant April 3, 2024 7,64 km2
Operations ceased in November 2008 due to lower copper/cobalt prices
Kananga Mine OP Dormant April 3, 2024 11,04 km2 Operations ceased due to pending relocation of railway
Kananga Extension OP Dormant May 7, 2022 0,849 km2 Operations ceased due to
pending relocation of railway UG – Underground Mine OP – Open Pit
Table 2: Mineral Processing Assets
Property Status
KTC Operating
Luilu Refinery Operating
1.6 Mineral Resources and Ore Reserves 1.6.1 Mineral Resources The consolidated mineral resources of the various areas of KCC as at 31 December 2011 are summarised in the table below. The actual mined out areas as on 31 December 2011 were used in the depletion of the mineral resources.
Table 3: KCC: Consolidated Mineral Resources as at 31 December 20111,2,3,4 Resource Classification Mt %TCu %TCo
Measured 40.5 4.14 0.54 Indicated 248.1 3.98 0.45 Measured and Indicated 288.6 4.00 0.46 Inferred 169.1 2.42 0.31 1) Mineral resources have been reported in accordance with the classification criteria of the Australian Code for
Reporting of Exploration Results, Mineral Resources and Ore Reserves, 2004 Edition (the JORC Code). If the definitions and classification standards adopted in NI 43-101 had been used instead of those of the JORC Code, the estimates of mineral resources would be substantially similar.
2) Mineral resources are inclusive of ore reserves.
3) Mineral resources are not ore reserves and do not have demonstrated economic viability.
4) The mineral resource estimates are for KCC's entire interest, whereas the Company owns 75% of KCC.
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1.6.1.1 Comparison of the 2011 and 2010 Mineral Resources In Table 4 below, the consolidated mineral resources of KCC as at December 31, 2011 are compared with those as at December 31, 2010.
Table 4: Comparison of the 2011 and 2010 Mineral Resources1,2,3,4,5
Classification Project Area 2011 2010
Mt %TCu %TCo Mt %TCu %TCo
Measured
KTO 32.1 4.33 0.58 30.7 4.54 0.54 T-17 Open Pit 4.5 2.71 0.54 0.0 0.00 0.00 KOV Open Pit 3.9 4.25 0.22 0.0 0.00 0.00 Subtotal 40.5 4.14 0.54 30.7 4.54 0.54
Indicated
KTO 32.9 4.63 0.57 35.7 4.69 0.60 Mashamba East Open Pit 75.0 1.80 0.38 75.0 1.80 0.38 T-17 Open Pit 9.4 4.44 0.65 8.5 2.75 0.87 KOV Open Pit 117.2 5.41 0.42 123.9 5.37 0.40 Kananga Mine 4.1 1.61 0.79 4.1 1.61 0.79 Tilwezembe Open Pit 9.5 1.89 0.60 9.5 1.89 0.60 Subtotal 248.1 3.98 0.45 256.7 3.95 0.45
Measured and Indicated
KTO 65.0 4.48 0.57 66.4 4.62 0.57 Mashamba East Open Pit 75.0 1.80 0.38 75.0 1.80 0.38 T-17 Open Pit 13.9 3.88 0.61 8.5 2.75 0.87 KOV Open Pit 121.1 5.37 0.41 123.9 5.37 0.40 Kananga Mine 4.1 1.61 0.79 4.1 1.61 0.79 Tilwezembe Open Pit 9.5 1.89 0.60 9.5 1.89 0.60 TOTAL 288.6 4.00 0.46 287.4 4.02 0.46
Inferred
KTO 11.0 5.00 0.59 10.6 5.11 0.59 Mashamba East Open Pit 65.3 0.76 0.10 65.3 0.76 0.10 T-17 Open Pit 5.2 4.21 0.98 15.3 1.91 0.61 KOV Open Pit 69.8 3.58 0.32 71.2 3.56 0.32 Kananga Mine 4.0 2.00 0.98 4.0 2.00 0.98 Tilwezembe Open Pit 13.8 1.75 0.60 13.8 1.75 0.60 TOTAL 169.1 2.42 0.31 180.2 2.32 0.32
1) Mineral resources have been reported in accordance with the classification criteria of the JORC Code. If the definitions and classification standards adopted in NI 43-101 had been used instead of the JORC Code, the estimates of mineral resources would be substantially similar.
2) Mineral resources are inclusive of ore reserves.
3) Mineral resources are not ore reserves and do not have demonstrated economic viability.
4) The mineral resource estimates are for KCC's entire interest, whereas the Company owns 75% of KCC.
5) Numbers may not add due to rounding.
There are no changes in the mineral resources reported for Mashamba East Open Pit, Kananga Mine and Tilwezembe Open Pit.
Overall, the mineral resources for Kamoto Copper Company (KCC) (in which the Company has a 75% interest) decreased by 9.9 million tonnes (2% of total mineral resources), while the TCu grade
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increased by 2% due to the increase in copper grade, which exists below the current pit floor in the T-17 Open Pit.
The material changes (more than 5% difference) in the mineral resources for KTO, T-17 Open Pit and KOV Open Pit are: a total increase of 5% in the measured mineral resource of KTO is based on the depletion of the resource due to mining (1.4 million tonnes) in 2011, an increase in the resource of 2.9 million tonnes due to the development of a new resource model for Etang North, which is based on new exploration data and updated density information, with subsequent movement of resources from Indicated mineral resources to measured mineral resources (increased confidence in the resource estimate) and a slight increase in the total tonnage due to a change in interpretation;
A decrease of 8% (2.8 million tonnes) in the indicated mineral resources of KTO is based on the movement of resources from indicated to measured mineral resources (increase in confidence in the resource estimate) based on the new interpretation;
The increase in measured mineral resources of 4.5 million tonnes for the T-17 Open Pit and 3.9 million tonnes for the KOV Open Pit, respectively, is due to increased confidence in the resource estimate based on recent exploration results, updated density information and new resource models based on new interpretations that are based on new drilling information; and
The RSC lithological component of the T-17 Open Pit below planned pit-bottom (8.9 million tonnes) was excluded from the resource statement to conform to the reporting of the mineral resources of KTO, where the RSC lithological component is excluded due to mining method limitations in the underground mines. All of these mineral resources are in the inferred category.
1.6.2 Ore Reserve Estimate Ukwazi Group visited the Material Assets of KCC during January 2012 and received geological and mineralogical data to assess and optimise mining plans developed for life of mine (LOM). The active operations for KCC can be classified into underground operations, namely KTO mine and surface mining operations, namely T-17 Open Pit and KOV Open Pit.
The KCC ore reserve estimate is set out below:
Table 5: KCC Ore Reserve Estimate1,2,3
Mining operation Proved Probable Total
Mt %TCu %TCo Mt %TCu %TCo Mt %TCu %TCo
Kamoto underground 13.0 3.43 0.51 19.4 3.70 0.53 32.4 3.59 0.52
T-17 underground 0.0 0.00 0.00 0.9 3.51 0.57 0.9 3.51 0.57
T-17 Open pit 0.0 0.00 0.00 1.6 3.52 0.56 1.6 3.52 0.56 Mashamba East Open pit 0.0 0.00 0.00 5.9 3.00 0.37 5.9 3.00 0.36
KOV Open pit 0.0 0.00 0.00 55.1 4.74 0.45 55.1 4.74 0.45
Total 13.0 3.43 0.51 83.0 4.33 0.46 96.0 4.21 0.47 1) The ore reserve estimates have been prepared in accordance with the classification criteria of the JORC Code.
If the definitions and classification standards of NI 43-101 had been used instead of those of the JORC Code, estimates of mineral reserves would be substantially similar to the estimates of ore reserves.
2) The ore reserve estimates are for KCC's entire interest in such ore reserves, whereas the Company owns 75% of KCC.
3) Numbers may not add due to rounding.
The current plan for KTO is to ramp yearly production up to 2.2 million tonnes of sulphide ore, to coincide with the completion of the Phase 4 processing plant expansion. Dilutions and mining over breaks applied
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vary from 4% to 13% and mining extractions vary from 3% – 5% inclusive of geological losses. The run-of-mine (ROM) head grade remains relatively constant at an average of 3.85% Cu over the life of 36.2 million tonnes ROM.
Two underground mines are planned for a production ramp up from 2018 onwards to replace and/or supplement production from the open pit operations. Conceptual work has been completed on T-17 underground and Kamoto East underground projects. These projects are planned on a total of 14.4 and 14.6 million tonnes ROM respectively at diluted grades of 3.63%Cu and 4.44%Cu.
T-17 Open Pit is approaching depletion and surface mining will conclude within the current open pit economic boundaries during 2012.
The KOV Open Pit delivers a ROM head grade of 4.33%Cu for a total of 83.8 million tonnes of ROM ore up to 2030. Ore production from the KOV Open Pit is primarily oxide material at 78% on average. The ore reserve is estimated at 55.1 million tonnes at 4.74%Cu, classified as probable ore reserves. KOV Open Pit is the major operational source of copper based on the current LOM plan.
For Mashamba East open pit, a total ROM production of 12.8 million tonnes at 2.72%Cu is planned at a production rate of up to 1.5 million tonnes per year. The pit would produce only oxide ore and the probable ore reserve is estimated at 5.9 million tonnes at 3.00%Cu.
KCC currently has an operational plan up to 2030 to mine a total of 169 million tonnes of copper ore, at an average copper grade of 3.90%Cu. The ore reserve estimate is 96.0 million tonnes at an average grade of 4.21%Cu.
1.6.3 Reconciliation with December 2010 Ore Reserve Estimate The outcome of the December 2011 ore reserve estimate is a decrease of only 1 million tonnes of ore reserve despite the fact that 4.5 million tonnes have been mined in 2011 due to the addition to the ore reserves of:
Ore reserves at KOV due to additional design work conducted,
Inclusion of the ore reserve up to 2014 at T-17 underground due to additional technical and design studies conducted during 2011; and
Inclusion of T-17 Extension due to additional study conducted.
Table 6: Ore Reserve Reconciliation1,2,3
Mining operation 2011 Ore Reserve
Estimate 2010 Ore Reserve
Estimate Notes Mt %TCu Mt %TCu
Kamoto underground 32.4 3.59 34.0 3.60 Mined out in 2011 and qualified to 2014
T-17 underground 0.9 3.51 0.0 0.00 Appropriates study to 2014 with increased resource Cu%
T-17 Open Pit 1.6 3.52 1.5 2.61 Approval and appropriate study of T-17 Extension
Mashamba East Open pit 5.9 3.00 5.9 3.00 Unchanged
KOV Open Pit 55.1 4.74 55.7 4.73 Mined out in 2012 and design adjustments
Total 96.0 4.21 97.0 4.20 Net reduction in ore reserves due to 2011 mining and other
1) The ore reserve estimates have been prepared in accordance with the classification criteria of the JORC Code. If the definitions and classification standards of NI 43-101 had been used instead of those of the JORC Code, estimates of mineral reserves would be substantially similar to the estimates of ore reserves.
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2) The ore reserve estimates are for KCC's entire interest in such ore reserves, whereas the Company owns 75% of KCC.
3) Numbers may not add due to rounding.
The ore reserves at KTO are qualified up to 2014. Appropriate technical design and scheduling study is required in the next 18 months to enable an ore reserve estimate from 2014 due to the material and strategic mine planning changes envisaged. The impact of these changes on the LOM Plan, mining infrastructure requirements and mining operational costs requires appropriate technical study.
1.7 Development and Operations The primary developments within the Material Assets during 2011 have been the following:
Exploration drilling continued in KTO, T-17 Open Pit and KOV Open Pit;
Re-interpretation of the geology of Etang North orebody in KTO, T-17 orebody and the orebodies in KOV;
KOV Open Pit was dewatered;
Milling capacity at KTC was increased to 7.68 million tonnes of ore per year; and
Luilu Refinery production capacity was increased to 150,000 tonnes of copper per annum due to the plant upgrade programme. Phase 3 was completed during the second quarter of 2011 and the New Phase 4 (inclusive of Phase 5) commenced during the fourth quarter of 2011.
Table 7: Plant and Processing Developments
Old Phase
Completion Increase in
Cu Capacity kt per annum
Increase in Co
Capacity kt per annum
New Phase
Completion Increase in
Cu Capacity kt per annum
Increase in Co
Capacity kt per annum
Refurbishment of existing facilities
1 2007 35 2 1 2007 35 2
2 2009 35 2 2 2009 35 2
3 2011 80 4 3 2011 80 4
Subtotal 150 8 150 8
Phase 4
4 2013 120 22.0 (as Cobalt
Hydroxide)
1.8 Interpretations and Conclusions The results of interpretations of developments of Material Assets are reported elsewhere in this report and have been relied upon to compile the mineral resource estimates included in section 14.1 and 14.2 and the ore reserve estimate in section 15.0.
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1.9 Recommendations The Qualified Person recommends the following actions be taken in respect of the Material Assets:
Further exploration of the operations which have underground mine potential such as the T-17 underground and KTE should be continued;
Dewatering strategy needs to be implemented particularly for open pits such as Mashamba East;
Plant and processing improvements and construction of new infrastructure such as the Solvent Extraction Plant and Electro-Winning Plant in terms of phase 4 need to be completed;
The capacity of the existing KTC tailings facility needs to be increased and the new tailings strategy needs to be implemented i.e. the use of a lime treatment plant to treat tailings prior to deposition and the authorisation and use of the Mupine Pit as a super tailings facility needs to be operationalised;
New deposition sites should be investigated since the existing tailings and waste facilities will require expansion in the future.
1.10 Economic Analysis Based on the free cash flows in Section 22.2, the NPV of KML’s holding in KCC is USD$5,890 million. This value includes the T-17 Underground and Kamoto East Underground mineral resources. These are not currently ore reserves, and do not have demonstrated economic viability. KCC intends to conduct further exploration of the properties including T-17 Underground and KTE. Table 8 shows the NPV of KCC at different discount rates.
Table 8: NPV of KML’s holding in KCC
NPV
(USD mil)
Discount Rate
7.5% 6,921
10.0% 5,890
12.5% 5,078
Because KCC is an existing operation and not a start-up, the internal rate of return and payback period are not meaningful calculations.
2.0 INTRODUCTION This ITR has been compiled for KML by GAA staff and other sub-consultants.
This ITR has been prepared to provide an insight into any material changes in the scientific and technical information concerning the Material Assets.
This ITR was prepared in compliance with the standards set out in the Canadian Securities Administrators’ NI 43-101, Companion Policy 43-101 CP and Form 43-101F1 and in conformity with the JORC Code of the Australasian Institute of Mining and Metallurgy, Australian Institute of Geoscientists and the Minerals Council of Australia.
2.1 Qualified Person This ITR was compiled by Willem van der Schyff, PriSciNat, (Registration Number 400176/05), a registered Professional Geologist with the South African Council for Natural Scientific Professions and a member of the Geological Society of South Africa. Each aspect of this ITR was prepared by or under the supervision of Mr.
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van der Schyff who is a Qualified Person as defined by NI 43-101. Mr. van der Schyff retains responsibility for his contribution described above.
Mr. van der Schyff has been a consultant with GAA for 7 years, prior to which he worked at a major mining house as a resource geologist.
2.2 Site Visits Mr. van der Schyff visited the Material Assets in January of 2012. During the site visits, Mr. van der Schyff inspected the Material Assets. Mr. van der Schyff verified the working protocols regarding the development of geological, mineral resource and ore reserve estimates, mine plans and processing plant development.
2.3 Basis of Technical Report This ITR was prepared as an update of the 2011 Technical Report. Many items presented in the 2011 Technical Report have served as a basis for this ITR specifically in regard to geology, mineralisation, plant and processing, legal tenure and financial estimates (such as operating expenditures, capital expenditures, and future cash flow).
This ITR is intended to update the 2011 Technical Report to reflect the following changes:
Phases 4 and 5 of the SX/EW refinery modular project are to be replaced by a New Phase 4 consisting of the conversion of the existing Luilu Copper Electro- Refinery to an EW plant fed by a new SX plant. This has resulted in $229 million of capital expenditure savings;
Restatement of the mineral resources to reflect the updated resource model of KTO, T-17 and KOV;
Update of the Life of Mine (LOM) plan, including confirmation of the intention to mine KTE from underground and to exploit additional T-17 mineral resources through underground mining techniques; and
Restatement of ore reserves to reflect the change in mining strategy with the exclusion of the KTE mineral resources from ore reserves pending finalization of the mine plan.
2.4 Terms of Reference The GAA Terms of Reference were to prepare the ITR to provide an insight into any material changes in the scientific and technical information concerning the Material Assets in accordance with the Canadian Securities Administrators’ NI 43-101, Companion Policy 43-101 CP and Form 43-101F1 and in conformity with the JORC Code of the Australasian Institute of Mining and Metallurgy, Australian Institute of Geoscientists and the Minerals Council of Australia.
Unless otherwise stated, all units of measurements in this ITR are metric; all costs are expressed in United States dollars ($); and the payable metals, copper and cobalt, are priced in $ per pound ($/lb). A detailed glossary of terms used in this ITR is attached as APPENDIX B.
The rehabilitation and expansion of the plant and facilities have been broken down into phases (each a Phase). Where these Phases have changed from the previous Technical Report, they are referred to as a new phase (New Phase). Phases 1, 2 and 3 and New Phase 4 are detailed in Table 7: Plant and Processing Developments.
3.0 RELIANCE ON OTHER EXPERTS Mr. van der Schyff, the Qualified Person for this ITR retains responsibility for this ITR as outlined in section 2.1 and as indicated on his certificate in section 30.0.
In the preparation of this ITR, the Qualified Person has relied on the technical contributions by GAA sub-consultants, SNC Lavalin (Robin Gardiner) in respect of sections 3.0 and 17.0, as well as the Ukwazi Group (Jaco Lotheringen) in respect of sections 3.0, 14.0, 15.0 and 16.0.
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This ITR is an update of the 2011 Technical Report as discussed in Section 2.3. The Qualified Person has also relied on the 2009 Technical Report prepared by SRK Consulting (South Africa) (Pty) Ltd (SRK). Table 6 provides details of the reports and information that have been relied upon by the Qualified Person in preparing this ITR:
Table 9: Reliance on Other Experts Organisation Report NI 43-101 Section
SRK
An Independent Technical Report on the Material Assets of Katanga Mining Limited, Katanga Province, Democratic Republic of Congo ( DRC ), dated March 17, 2009
Property Description and Location
Exploration and Drilling
Mineral Processing
Environmental Considerations
Economic Analysis
Risk Assessment
SRK Report on the proposed feasibility of Tailings Deposition in Mupine Pit, November 30, 2011 Tailings Deposition
GAA A Technical Report on the Material Assets of Katanga Mining Limited, Katanga Province, DRC, dated March 31, 2011
Ownership
Material Assets
Mineral Resources and Reserves
Drilling and Exploration
Mineral Processing
GAA Mineral Expert’s Report on the Material Assets of Katanga Mining Limited, Katanga Province, DRC, dated March 31, 2011
Ownership
Material Assets
Mineral Resources and Reserves
Drilling and Exploration
Mineral Processing
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4.0 PROPERTY DESCRIPTION AND LOCATION 4.1 KCC Rights: Summary of Permits Tabulated below is a summary of KCC’s permits and current legal tenure in regard to concession areas:
Table 10: Legal Tenure of KCC
Property Exploitation Permit Number Rights Granted Area of Title Valid Until
KTO and Mashamba East Open Pit
PE525 Cu, Co and associated minerals and metals
13 blocks, 11,04km2
April 3, 2024 Renewable
T-17 Open Pit PE11602 Cu, Co, nickel and gold 2 blocks, 1,698km2
April 3, 2024 Renewable
Extension of Kananga PE11601 Cu, Co, nickel and gold 1 block,
0,849km2 May 7, 2005 Renewable
KOV Open Pit PE4961 Cu, Co and associated minerals and the use of Surface rights
10 blocks, 8,49km2
April 3, 2024 Renewable
Tilwezembe Open Pit PE4963
Cu, Co and associated minerals and the use of Surface rights
9 blocks, 7,64km2
April 3, 2024 Renewable
Kananga Mine PE4960 Cu, Co and associated minerals and the use of Surface rights
13 blocks, 11,04km2
April 3, 2024 Renewable
PE’s under the DRC Mining Code are renewable in accordance with the terms of the DRC Mining Code for periods of 15 years.
A PE grants to its holder the exclusive right to carry out exploration and exploitation works for the minerals for which it has been granted. This right covers the construction of necessary facilities for mining exploration, the use of water and wood resources, and the free commercialisation of products for sale, in compliance with corresponding legislation.
4.2 KCC Rights: Joint Venture Agreements In April of 2007, a Commission of Enquiry (the Commission) was formed by the DRC government to review approximately 60 mining agreements entered into by government parastatal companies. The joint venture agreements that resulted in the establishment of both KCC and DCP were included in the mining agreements to be reviewed.
Following the release of the Commission’s report in November of 2007, KCC and DCP were notified on February 11, 2008 by the DRC Ministry of Mines (CAMI) of the objections and requirements regarding their partnerships with GCM.
In July of 2008, GCM and Katanga Finance Limited (KFL), a 100% subsidiary of KML, entered into a memorandum of understanding in terms of which certain amendments were agreed to be as reflected in an amended joint venture agreement. The parties agreed to the merger of KCC and DCP.
In August of 2008, the CAMI issued terms of reference for the renegotiation and/or termination of the mining contracts entered into by KCC and DCP.
Following a number of meetings during the course of the last quarter of 2008 and the first quarter of 2009, GCM, KFL and Global Enterprises Corporate Ltd. (GEC), in the presence of KCC, DCP, SIMCO, Katanga Mining Holdings Limited, Katanga Mining Finance Limited and KML (BVI) Holdco Ltd., entered into an
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amended joint venture agreement (AJVA) on July 25, 2009 (was also the effective date) which resulted in the termination of the original KCC joint venture agreement and the DCP joint venture agreement. The merger of KCC and DCP was ratified by Presidential Decree on April 27, 2010.
As a result of the AJVA:
Certain Permit d’Exploitations (each a PE) were transferred from DCP to KCC. The whole of PE525 (comprising 13 carrés) and part of PE4958 (i.e. the new PE11602 described below and comprising two carrés containing the T-17 deposit) were transferred to KCC. The Kamoto, Mashamba East and T-17 deposits and any extensions of these deposits which are within the perimeter of PE525 and the two carrés of PE4958 that have been transferred to KCC, shall be for the sole benefit of KCC. Such transfer was completed pursuant to a transfer deed dated July 27, 2009 and evidenced by the CAMI in its exploitation certificate no. CAMI/CE/5621/2009 dated November 27, 2009;
DCP PE’s were transferred to KCC following completion of the merger with DCP. In addition, one carré of PE 7044 (i.e. new PE11601 being an extension of the Kananga deposit) was transferred by GCM to KCC since the holder of PE652 released the carré to be transferred from its tailings area. Such transfer was completed pursuant to a transfer deed dated July 27, 2009 and evidenced by CAMI in its exploitation certificate no. CAMI/CE/5622/2009 dated November 27, 2009; and
The perimeter of the merged KCC/DCP concession area will contain the Necessary Surfaces (as defined in the AJVA.
Pursuant to the AJVA, the Necessary Surfaces will be sourced from PE8841 held by GCM and from one carré close to the T-17 deposit. Easements have been granted to enable KCC to establish and maintain operating facilities for the KOV Open Pit waste removal conveyor belt system. KCC has agreed to fund an independent contractor to determine whether the surfaces identified as potential Necessary Surfaces contain any mineral reserves. Provided no mineral reserves are discovered, the relevant surfaces shall be converted into multiple PEs (where required) and shall be leased to KCC. Should any mineral reserves be discovered in the identified surfaces, the mineral reserves shall be transferred to KCC and shall count as Replacement Reserves (as defined in 6.2 below) under the terms of the AJVA.
In addition, under the AJVA, KCC was granted an option for a period of three years following its merger with DCP to increase the Necessary Surfaces by the five carrés (to be leased) contained in PE8841, if such extension is required for the project. Beyond this three-year period, KCC shall have a pre-emptive right on these 5 carrés in the event that GCM is willing to transfer or make any part of them available to third parties.
The rent for the Necessary Surfaces (including the five additional carrés if the option is exercised within the three year period) amounts to $600,000/year. However, KCC, as the merged entity, will remain liable for the payment of the rental tax (22%) which will be in addition to the royalties owed by KCC to GCM.
As part of the AJVA, it has also been agreed that upon the winding up or liquidation of KCC, the mining rights and titles of KCC shall revert to GCM without further consideration.
Pursuant to the AJVA, GCM and KCC have signed an agreement relating to the lease by GCM to KCC of certain equipment and installations as described in an annex to the AJVA (the Equipment and Installations). The rent for the Equipment and Installations payable by KCC to GCM is $1,200,000/year to be deducted from the royalties owed by KCC to GCM. However, KCC will remain liable for the payment of the rental tax (22%) which will be in addition to these royalties.
KCC shall retrocede the Equipment and Installations free of charge to GCM upon lawful termination or final expiry of the AJVA.
As part of the AJVA, it has also been agreed that GCM grants and/or makes available to KCC, subject to payment of the reasonable maintenance costs, the following rights:
(i) The right to use roads, railways, rail routes, waterways; (ii) The right to avail itself of rights of way, easements, rights to water; and
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(iii) All the supplementary rights that can facilitate access to or use of the lands involved and the facilities located thereon, which GCM enjoys outside the perimeter of the KCC project insofar as they are necessary or desirable to carry out the project in the most cost effective manner.
Table 10 contains a list of the Material Assets and describes the:
(i) Area of each property; (ii) The location of each property; (iii) The PE corresponding to each property; (iv) The type of mineral tenure of each property; (v) Granted on or to each property; and (vi) The expiration date of each PE.
4.3 Replacement Reserves Pursuant to the AJVA, the mineral reserves to be replaced in exchange for the Dikuluwe and Mashamba West deposits surrendered to GCM pursuant to the release agreement (the Released Deposits) amount to approximately 4.0Mt of copper and 0.2Mt of cobalt.
No pas de porte (entry premium) shall be paid to GCM in relation to the transfer of the mineral reserves to be transferred to KCC as compensation for the Released Deposits (the Replacement Reserves).
Pursuant to the AJVA, GCM and KCC are also required to jointly scope, implement and manage an exploration programme (the Exploration Programme) with the object of identifying sufficient Replacement Reserves and transferring them to KCC by no later than July 1, 2015. The Exploration Programme can take place within the perimeters of:
(i) The KCC PEs (excluding the KTO, Mashamba East Open Pit, Tilwezembe Open Pit, Kananga Mine, T-17 Open Pit and KOV Open Pit deposits and any extensions of these deposits);
(ii) The Necessary Surfaces; or (iii) Other perimeters belonging to GCM.
The Exploration Programme is to be financed by way of a loan from KCC to GCM and refunded, without interest, by GCM through the set-off against the royalties and dividends payable by KCC.
If any Replacement Reserves are identified by GCM as a result of the Exploration Programme or otherwise, they shall be evaluated and certified in accordance with the JORC Code of the Australasian Institute of Mining and Metallurgy, Australian Institute of Geoscientists and Mineral Council of Australia, as amended.
Once GCM has satisfied KCC that it has good legal title to such Replacement Reserves and they are covered by valid PEs, KCC shall enter into a transfer deed or a lease, pursuant to which the Replacement Reserves shall be transferred or leased (amodié) to KCC.
If GCM does not replace these Released Deposits by July 1, 2015 it must pay $285,000,000 as financial compensation. KFL, GEC and GCM agreed that the financial compensation would be due from July 1, 2015 and that interest would be charged if the financial compensation is not paid within the two months following 1 July 2015. During the first 12 months following the two month grace period the interest rate applicable to the unpaid financial compensation amount would be limited to the London Interbank Offering Rate (Libor) (6 month) as opposed to Libor (6 month) + 300 basis points which will become applicable as of the end of the 12 month period.
GCM accepts that KCC may withhold any future revenues owed to GCM (i.e. royalties and dividends, with the exception of the pas de porte) until the financial compensation is fully paid.
4.4 Property Boundaries The property boundaries of the PEs of KCC are described in Figure 1. Figure 1 also shows the Necessary Surfaces.
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Figure 1: Property Boundaries of KCC's Exploitation Permits
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4.5 Royalties, Duties and Other Fees 4.5.1 Royalties payable to the State The holder of a mining exploitation title is subject to mining royalties which are calculated on the basis of the amount of sales minus the costs of transport, the costs of analysis concerning the quality control of the commercial product for sale, the cost of insurance and costs relating to the sale transaction. The royalties are due upon the sale of the product. The mining royalties are 2% for non ferrous metals.
4.5.2 Surface Rights Fees payable to the State Under Article 198 of the DRC Mining Code, KCC is required to pay surface rights fees of $5 per hectare per year or $424.78 per carré for exploitation permits.
Additional surface rights fees are payable by KCC as the merged entity holder of exploitation mining rights to the central government of the DRC pursuant to Article 238 of the DRC Mining Code at the rate of $0,08 per hectare.
4.5.3 Royalties payable to Gécamines Under the AJVA, it was agreed that the royalty rate for equipment and facilities provided by Gécamines as well as for ore reserve depletion was 2.5% of net revenues. Net revenues are to be determined on the same basis royalties are calculated under Article 240 of the DRC Mining Code, namely sales less transportation costs, quality control costs, insurance costs and marketing costs.
4.5.4 Pas de Porte payable to Gécamines A pas de porte payment is payable by KFL/GEC to Gécamines for access to the project. The total amount shall be $140 million, the payment of which will be completed as follows:
$5 million previously paid by GEC to Gécamines as a loan, was converted into a pas de porte payment.
$135 million to be paid by KFL. This will comprise:
$24.5 million which was paid by way of set-off against the amount of the advance granted by KFL to Gécamines for payment of the subscription price, and
$5 million was paid upon the transfer of PE525, PE11601 and PE11602 to KCC, described in Section 6.1 above; and
$10 million on an annual basis between 2009 and 2011 and $15 million on an annual basis between 2012 and 2015, with a final payment in 2016 of $15.5 million. The parties have agreed that these amounts shall be paid without any deductions or set off.
No further pas de porte will be payable in respect of the Replacement Reserves; however, any additional tonnage brought by Gécamines to KCC as the merged joint venture after the Released Deposits have been fully compensated will incur a new pas de porte payment of $35/t copper.
4.5.5 Customs Duties and Taxes Payable There is a requirement to pay customs duties and taxes in accordance DRC Legislation.
5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY
There has been no material change to the information contained in the Accessibility, Climate, Local Resources, Infrastructure and Physiography section of the 2011 Technical Report.
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6.0 HISTORY Other than as disclosed under section Property Description and Location or in the chart below, there has been no material change to the information contained in the History section of the 2009 Technical Report. The chronological developments in regards to the Material Assets are tabulated below.
Table 11: Chronological developments for Material Assets
Unit 2008 2009 2010 2011
KTO Mined ore Mt 0.6 1.1 1.3 1.6 Cu grade % 3.93 3.85 3.82 3.93 Co grade % 0.43 0.49 0.56 0.54
KOV Mined ore Mt 0.0 0.0 0.7 2.5 Cu grade % 0.00 0.00 4.43 4.97 Co grade % 0.00 0.00 0.30 0.24
T-17 Open Pit Mined ore Mt 0.4 1.6 1.9 0.4 Cu grade % 1.72 1.30 2.35 2.68 Co grade % 0.89 0.85 0.93 0.69
Tilwezembe Open Pit Mined ore Mt 0.6 0.0 0.0 0.0 Cu grade % 1.39 0.00 0.00 0.00 Co grade % 1.17 0.00 0.00 0.00
7.0 GEOLOGICAL SETTING 7.1 General Geology There has been no material change in the information contained in the General section of the 2010 Technical Report.
7.2 General Stratigraphy There has been no material change in the information contained in the General Stratigraphy section of the 2010 Technical Report.
7.3 Local Geology and Geological Models With the exception of Tilwezembe Open Pit all of the mineralized properties of KCC are localized within the Kolwezi Nappe, a Northeast striking synclinal basin with major and minor axes of approximately 20 km and 10 km respectively. Tilwezembe Open Pit is located about 12 km to the east of Kolwezi.
Within the Kolwezi Nappe, each of the project areas, T-17 Open Pit, KTO Mine, KOV Open Pit, Kananga Mine and Mashamba East Open Pit contain fragments with intact successions of Series Des Mines lithologies, which host the copper and cobalt mineralization. The fragments are often structurally complex, being tightly folded and exhibiting variable strikes and dips both within individual rafts and between neighbouring rafts.
7.3.1 KTO Mine The KTO operations extract mineralized copper ores from the Kamoto Principal fragment, which is differentiated from Kamoto East, mined in the KOV Open Pit, but contains the same lithologies. The morphology of the ore body is described as flat to gently dipping in the central parts, becoming steeper towards the flanks. Dips in the central parts vary between 0° and 20° increasing to about 45° towards the
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flanks. Dips in the flank regions are between 45° to 85°. The ore body is subdivided into three regions as follows and shown in the geological model, Figure 2:
The main central region, commonly referred to as the Principal;
Etang South; and
Etang North.
Figure 2: Plan view of KTO Mine Resource Model
Figure 3: Oblique view of KTO Mine with Resource Model
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7.3.2 T-17 Open Pit The T-17 Open Pit can be described as dismembered structurally complex packages, which belong to the southern flank of a synclinal fold that extends 2.6km and is overturned towards the north. Faulting is assumed to be the predominant process in the deformation and dismemberment of the deposit. The geological model is shown in Figure 4.
Figure 4: Oblique view of T-17 Open Pit Resource Model
7.3.3 KOV Open Pit There are three main individual “fragments” hosting mineralized Lower Roan lithologies within the KOV Open Pit area. These are Kamoto East, Oliveira and Virgule, from which the name KOV is derived. A fourth and smaller fragment, the FNSR, is a remnant of the Musonoi West fragment mined to the east of KOV Open Pit. The FNSR lies below and is sub-parallel to the Virgule orebody.
Other fragments within the area are OEUF and Variante. The OEUF consists mostly of hanging-wall lithologies occurring above the Virgule fragment, and the Variante lies below the Virgule and Oliveira fragments but outcrops towards the east in the Musonoi West area. Lower Roan lithologies have been identified in the Variante, but investigations indicate poor copper and cobalt mineralization within these lithologies. Within each of the mineralized fragments, the succession of lithologies is intact, although in the FNSR fragment the Lower Roan lithologies occur overturned.
The fragments that make up the KOV Open Pit orebody occur in an east-west-striking synclinal structure consisting of a steeply dipping southern limb and a shallow dipping northern limb, respectively named the Kamoto East and Virgule orebodies, while the Oliveira fragment is a shallower-dipping orebody in faulted contact with and below the Virgule orebody.
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Figure 5: Oblique view of KOV Open Pit Resource Model
7.3.4 Mashamba East Open Pit Mashamba East Open Pit operated from 1985 through 1988 and the open pit produced a total of 9.8 million tonnes of ore at an average grade of 4.96%Cu and 0.35%Co. By 1998, due to the lack of funds and increasing costs, the open pit was allowed to flood.
Structurally, the lithologies of the Mashamba East orebody strike to the north-east and dip gently to the north in the west and wraps around to strike almost north-south and dip to the east in the eastern portion of the property. The orebody consist of four units, RSF, SDB, BOMZ and RSC.
Figure 6: Oblique view of Mashamba East Resource Model and Planned Pit Layout
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7.3.5 Kananga Mine The Kananga orebody outcrops and forms a ridge with a N-NE strike. The ridge falls quite rapidly towards the south and has been cut to form part of the embankment for the main railway line, which runs parallel to the ridge and 10m to 20m away from it for most of the strike length of the orebody. GCM’s interpretations indicate that Kananga is the northern limb of the Kananga-Dilala syncline, which plunges to the south. The geological model is shown in Figure 7.
Figure 7: Oblique view of the Kananga Resource Model
7.3.6 Tilwezembe Open Pit The mineralized zone of Tilwezembe is located in an NE-SW anticlinal structural lineament, which extends further to the east where there are known copper and cobalt deposits (Kisanfu, Myunga, Kalumbwe and Deziwa). Strongly brecciated siliceous dolomites and shales of the Mwashya Formation (or R4) dominate with interstitial bands of hematite and dolomites. The strata strike almost east-west and dips at about 45° to the south. The geological model is shown in Figure 8.
Figure 8: Oblique view of Tilwezembe Pit with Resource Model and Pit Layout
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7.3.7 Deposit Types The deposits are stratiform with supergene enrichment within the upper surface layers.
Stratiform deposits are hydrothermal deposits but the ore minerals are always confined within specific strata and are distributed in a manner that resembles particles in a sedimentary rock. Because stratiform deposits so closely resemble sedimentary rocks, controversy surrounds their origin. In certain cases, such as the White Pine copper deposits of Michigan, the historic Kupferschiefer deposits of Germany and Poland, and the important copper deposits of Zambia and DRC, research has demonstrated that the origin is similar to that of Mississippi Valley Type (MVT) deposits—that is, a hydrothermal solution moves through a porous aquifer at the base of a pile of sedimentary strata and, at certain places, deposits ore minerals in the overlying shales. The major difference between stratiform deposits and MVT deposits is that, in the case of stratiform deposits, the host rocks are generally shales (fine-grained, clastic sedimentary rocks) containing significant amounts of organic matter and fine-grained pyrite.
In ore deposit geology, supergene processes or enrichment occur relatively near the surface. Supergene processes include the predominance of meteoric water circulation with concomitant oxidation and chemical weathering. The descending meteoric waters oxidize the primary (hypogene) sulphide ore minerals and redistribute the metallic ore elements. Supergene enrichment occurs at the base of the oxidized portion of an ore deposit. Metals that have been leached from the oxidized ore are carried downward by percolating groundwater, and react with hypogene sulphides at the supergene-hypogene boundary. The reaction produces secondary sulphides with metal contents higher than those of the primary ore. This is particularly noted in copper ore deposits where the copper sulphide minerals chalcocite, covellite, digenite, and djurleite are deposited by the descending surface waters.
The copper deposits in the DRC are well known world class deposits.
8.0 MINERALISATION Primary mineralization, in the form of sulphides, within the Lower Roan is associated with the D Strat and RSF for the OBI and the SDB and SDS for the OBS and is thought to be syn-sedimentary in origin. Typical primary copper sulphide minerals are bornite, chalcopyrite, chalcosite and occasional native copper while cobalt is in the form of carrolite. The mineralization occurs as disseminations or in association with hydrothermal carbonate alteration and silicification. Supergene mineralization is generally associated with the levels of oxidation in the sub-surface sometimes deeper than 100m below surface. The most common secondary supergene minerals for copper and cobalt are malachite and heterogenite. Malachite is the main mineral mined within the confines of the current KOV Open Pit.
The RSC, a lithological unit stratigraphically intermediate between the upper and lower ore body host rocks, contains relatively less copper mineralization. The RSC contains appreciable copper mineralization near the contacts with the overlying SDB formation and the underlying RSF formations. The middle portion of the RSC, considered to be “sterile” by GCM, normally contains relatively less copper mineralization and is sometimes not sampled. The mineral potential of the RSC is less well known than that of other formations.
The RSC has been observed to be well mineralized in supergene cobalt hydroxide, heterogenite, which occurs as vug infillings, especially near the surface.
The mineralization at Tilwezembe Open Pit is atypical being hosted by the Mwashya or R4 Formation. The mineralization generally occurs as infilling of fissures and open fractures associated with the brecciation. The typical mineralization consists mainly of copper minerals (chalcopyrite, malachite and pseudomalachite), cobalt minerals (heterogenite, carrolite and spherocobaltite) and manganese minerals (psilomelane and magnetite).
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8.1 Mineral Resource Estimation Methodology Mineral Resource models from which mineral resources are quoted in this report were generated by the following independent consultants:
Table 12: Independent Consultants Name of Operation Name of Consultant
KTO Mine CCIC and KCC T-17 Open Pit KCC KOV Open Pit SRK and KCC Mashamba East Open Pit SRK Kananga Mine Snowden Tilwezembe Open Pit Snowden
GAA has the overall responsibility for the sign-off on the mineral resources and ore reserves and as part of that process has reviewed the methods and results adopted in the generation of the mineral resource models from the respective consultants and KCC staff. GAA’s review of the methods indicates similarities of approach between the various consultants, and the findings of the review are summarized below:
Collate all the GCM geological information in the form of plans and sections and drill-hole logs in hard-copy format;
Digitize the plans and sections for the generation of wireframe lithological models and capture drill-hole data;
Define the envelopes outlining the limits of the zones of mineralization within each fragment in the project. Generally, this is a lithological cut-off (or a 0%TCu cut-off) defining the OBI as mineralization in the Rat Grises (RATGR, an argillaceous dolomite), DTSRAT and RSF and the OBS mineralization within the SDB, the black ore mineral zone (BOMZ) and to a lesser extent, the shalle dolomitic (SD1a, the argillaceous part of the SDB). The RSC, which is intermediate between the OBI and OBS, is defined separately and split into a top, mid and bottom RSC. An exception was the work undertaken by Snowden on Kananga and Tilwezembe, where a 0.50%TCu cut-off was used;
Undertake statistical and geostatistical analyses of the sample data within the defined envelopes of mineralization and derive variogram parameters;
Estimate grades into the zones of mineralization using Kriging techniques with attendant geostatistical parameters, search neighbourhood and input composite data; and
Classify the mineral resources into the various categories defined by the JORC Code.
9.0 EXPLORATION AND DATA The project area contains mostly historical information from diamond drilling by the previous owners, GCM. Since 2009, KCC has conducted infill drilling in the 3 main production areas (KOV Open Pit, KTO and T-17 Open Pit) and the results of these drillings campaigns are referenced below and are included in the mineral resource or the ore reserve statements.
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9.1 Exploratory Data Analysis 9.1.1 KTO Mine Table 13 to Table 15 presents statistics for Kamoto Principal, Etang South and Etang North.
Table 13: Statistical Analysis of Kamoto Principal Lithology Variable No Samples Minimum Maximum Mean Std. dev Dstrat
%TCu
427 0.11 9.56 4.08 1.40 RSF 404 0.14 22.00 4.79 1.66 RSC_B 241 0.10 27.40 7.59 3.87 RSC_T 165 0.24 18.30 5.97 3.53 SD1A 326 0.40 22.40 5.87 2.19 Dstrat
%TCo
393 0.01 8.24 0.36 0.44 RSF 378 0.01 3.15 0.30 0.26 RSC_B 224 0.03 5.99 0.51 0.68 RSC_T 159 0.01 4.39 0.73 0.65 SD1A 307 0.01 6.40 0.63 0.51
Table 14: Statistical Analysis of Etang South Lithology Variable No Samples Minimum Maximum Mean Std. dev Dstrat
%TCu
89 0.30 7.62 2.89 1.33 RSF 103 0.20 16.16 3.44 1.44 RSC_B 50 0.18 12.05 2.98 2.35 RSC_T 39 0.25 12.00 2.46 1.87 SD1A 155 0.15 13.20 5.92 2.97 Dstrat
%TCo
91 0.08 17.61 0.75 1.84 RSF 104 0.10 2.55 0.60 0.42 RSC_B 49 0.22 5.94 1.22 1.06 RSC_T 40 0.17 2.96 0.96 0.61 SD1A 154 0.07 3.96 1.04 0.74 Table 15: Statistical Analysis of Etang North Lithology Variable No. Samples Minimum Maximum Mean Std. dev
OBS
%TCu
991 0.03 12.51 2.38 2.27
RSC 964 0.03 12.42 0.98 1.79
OBI 725 0.08 10.01 2.65 1.45
OBS
%TCo
815 0.01 7.41 0.49 0.63
RSC 918 0.01 7.44 0.49 0.62
OBI 622 0.01 3.22 0.36 0.38
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9.1.2 T-17 Open Pit The lithological wireframes were used to extract the sample data, and the data was composited to 2.5m lengths. Statistics from the composite files are given in Table 16. GAA could reproduce this table.
Table 16: Statistical Analysis of T-17 Open Pit Lithology Variable No Samples Minimum Maximum Mean Std Dev DSTRAT
%TCu
360 0.02 12.64 4.32 3.05 RSF 601 0.07 16.79 3.20 2.85 RSC 1059 0.02 35.03 1.44 2.91 SDB 1389 0.01 22.69 5.12 3.96 BOMZ 211 0.10 18.71 2.53 2.84 RATGRIS 163 0.09 10.00 3.47 2.79 DSTRAT
%TCo
360 0.00 4.45 0.41 0.53 RSF 601 0.03 7.60 0.88 1.05 RSC 1050 0.00 9.21 0.67 1.02 SDB 1389 0.00 8.55 0.80 1.17 BOMZ 211 0.05 2.23 0.47 0.48 RATGRIS 163 0.01 2.84 0.30 0.34
9.1.3 KOV Open Pit Assay data for each lithology were extracted from the database using the lithological wireframes. The lithological sample data were then composited separately at intervals of 2,5 m. The RSC was composited across the entire lithology unit, and the statistics reflect the sample intervals.
Statistics from the lithological composites within each of the four fragments are presented in Table 17 to
Table 20.
Table 17: Statistical Analysis of KOV Open Pit; Virgule Lithology Variable No Samples Minimum Maximum Mean Std dev
BOMZ
%TCu
77 0.10 12.00 3.54 3.10
SDB 296 0.02 12.92 5.97 3.60
RSC 384 0.08 23.14 4.39 3.47
RSF 171 0.14 12.00 6.20 3.03
DSTRAT 103 0.50 12.00 6.43 2.41
RATGR 49 0.80 14.35 6.36 3.42
BOMZ
%TCo
112 0.00 3.62 0.46 0.67
SDB 371 0.00 11.50 0.46 0.84
RSC 587 0.00 5.00 0.21 0.40
RSF 244 0.00 2.25 0.19 0.32
DSTRAT 128 0.00 1.52 0.22 0.31
RATGR 103 0.00 0.99 0.09 0.16
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Table 18: Statistical Analysis of KOV Open Pit; Oliviera Lithology Variable No Samples Minimum Maximum Mean Std dev BOMZ
%TCu
72 0.15 7.88 2.05 1.99 SDB 250 0.10 13.68 5.59 2.48 RSC 266 0.10 16.59 4.72 3.71 RSF 120 0.40 14.99 5.40 3.16 DSTRAT 73 0.91 12.00 4.92 2.31 RATGR 43 0.60 16.41 4.80 2.78 RATGR
%TCo
60 0.00 1.24 0.23 0.33 BOMZ 80 0.00 2.25 0.41 0.42 SDB 239 0.00 3.66 0.93 0.70 RSC 512 0.00 4.58 0.29 0.55 RSF 111 0.00 2.90 0.38 0.48 DSTRAT 71 0.00 1.61 0.32 0.33
Table 19: Statistical Analysis of KOV Open Pit; FNSR Lithology Variable No Samples Minimum Maximum Mean Std dev BOMZ
%TCu
9 0.53 12.00 5.24 3.41 SDB 71 1.11 12.00 7.93 3.16 RSC 69 0.15 12.00 5.27 4.28 RSF 25 1.66 12.47 5.76 2.40 DSTRAT 7 4.30 8.98 7.17 1.48 RATGR 2 4.06 8.02 6.04 1.98 BOMZ
%TCo
11 0.00 2.85 0.67 0.89 SDB 79 0.00 6.15 0.42 0.88 RSC 110 0.00 1.96 0.18 0.29 RSF 25 0.02 0.47 0.18 0.13 DSTRAT 9 0.00 0.16 0.04 0.05 RATGR 2 0.02 0.07 0.04 0.02
Table 20: Statistical Analysis of KOV Open Pit; Kamoto East Lithology Variable No Samples Minimum Maximum Mean Std dev BOMZ
%TCu
27 0.17 13.22 6.16 3.31 SDB 133 0.08 14.38 6.32 3.34 RSC 206 0.11 16.07 4.20 3.72 RSF 76 0.21 10.84 5.09 3.11 DSTRAT 43 0.21 11.55 5.53 2.89 DSTRAT 70 0.00 1.40 0.21 0.28 RATGR 15 3.02 9.02 6.37 1.71 BOMZ
%TCo
43 0.00 4.96 0.71 0.97 SDB 173 0.00 4.63 0.41 0.59 RSC 288 0.00 2.52 0.22 0.29 RSF 96 0.00 1.50 0.26 0.30 RATGR 25 0.00 0.86 0.13 0.19
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9.1.4 Mashamba East Open Pit The lithological wireframes were used to extract the sample data, and the data were composited to 2.5m lengths. Statistics from the composite files are shown in Table 21.
Table 21: Statistical Analysis of Mashamba East Open Pit Lithology Geozone Variable No Samples Minimum Maximum Mean Std dev
BOMZ WEST
%TCu 125 0.01 10.60 0.48 1.40 %TCo 125 0.01 0.44 0.09 0.11
EAST %TCu 66 0.01 12.00 0.70 1.78 %TCo 66 0.01 1.08 0.14 0.23
SDB WEST
%TCu 237 0.01 19.26 1.19 2.46 %TCo 237 0.01 4.14 0.36 0.61
EAST %TCu 64 0.01 9.15 1.82 2.51 %TCo 64 0.01 2.63 0.33 0.48
RSC WEST
%TCu 248 0.01 12.00 0.60 2.00 %TCo 248 0.00 2.01 0.07 0.23
EAST %TCu 112 0.01 12.00 0.69 2.06 %TCo 112 0.01 0.66 0.04 0.08
RSF WEST
%TCu 440 0.01 12.00 2.28 2.85 %TCo 440 0.00 2.80 0.37 0.54
EAST %TCu 119 0.01 11.37 3.83 2.96 %TCo 119 0.00 1.28 0.14 0.22
9.1.5 Kananga Mine The summary statistics of the declustered 1m composite data of the various rock types are presented in Table 22. The composite data was declustered using a cell size of 25m E by 25m N by 1m RL that approximates the drill-hole spacing in the closer spaced areas.
Table 22: Statistical Analysis of Kananga Mine Domain Variable No Samples Minimum Maximum Mean Std dev UOB_OX
%TCu
250 0.08 10.05 1.13 1.10 MID_OX 528 0.02 0.64 0.16 0.60 LOB_OX 297 0.03 9.28 1.93 0.80 UOB_SL 122 0.01 6.10 1.83 0.60 MID_SL 269 0.02 1.00 0.26 0.60 LOB_SL 234 0.13 6.75 2.18 0.60 UOB_OX
%AsCu
250 0.02 9.05 0.85 1.20 MID_OX 528 0.01 0.62 0.12 0.80 LOB_OX 297 0.01 9.26 1.26 1.10 UOB_SL 122 0.01 1.61 0.15 1.20 MID_SL 269 0.01 0.39 0.07 1.00 LOB_SL 234 0.01 2.99 0.22 1.60 UOB_OX
%TCo
250 0.06 4.51 0.64 1.20 MID_OX 528 0.02 2.73 0.27 0.80 LOB_OX 297 0.02 3.37 0.70 0.80 UOB_SL 122 0.01 4.79 0.94 1.10 MID_SL 269 0.06 2.03 0.53 0.60 LOB_SL 234 0.02 3.49 1.05 0.70
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9.1.6 Tilwezembe Open Pit The summary statistics of the declustered one-metre composite data of the various rock types are presented in Table 23. The composite data was declustered using a cell size of 25m E by 25m N by 1m RL that approximates the drill-hole spacing in the closer spaced areas.
Table 23: Statistical Analysis of Tilwezembe Open Pit Domain Variable Samples Minimum Maximum Mean Std dev OX_MNDOL
%TCu
1054 0.01 26.50 1.27 1.91 OX_BREC 485 0.10 19.84 3.78 0.89 OX_TILAR 674 0.05 4.92 0.56 1.05 SL_MNDOL 511 0.03 37.00 1.19 2.42 SL_BREC 339 0.02 21.38 3.29 0.97 SL_TILAR 406 0.01 6.27 0.53 1.28 OX_MNDOL
%ASCu
1054 0.01 14.12 0.91 1.85 OX_BREC 485 0.05 14.28 3.16 0.96 OX_TILAR 674 0.01 4.90 0.44 1.17 SL_MNDOL 511 0.01 6.88 0.24 2.97 SL_BREC 339 0.01 3.52 0.33 1.19 SL_TILAR 406 0.01 1.09 0.13 1.21 OX_MNDOL
%TCo
1054 0.01 11.15 0.50 1.47 OX_BREC 485 0.01 13.47 0.96 1.58 OX_TILAR 674 0.01 3.61 0.29 1.02 SL_MNDOL 511 0.01 7.94 0.38 1.52 SL_BREC 339 0.03 4.98 1.19 1.01 SL_TILAR 406 0.02 4.00 0.34 1.15
9.2 Variography Analysis The objectives of variography are to establish the major directions of continuity and to provide the variogram parameters required for geostatistical grade interpolation. The experimental semi-variogram (commonly referred to as the variogram) is the basic diagnostic tool of geostatistics. It is a mathematical function used to quantify the spatial variation and correlation of sample grades in various directions in a deposit. The variogram calculation is similar to the variance and it is arithmetically simple: the differences between pairs of sample values a particular distance apart are squared. This is repeated for increasing distances for all samples within a homogeneous zone. The variogram value is the sum of the squared differences divided by twice the number of pairs.
The experimental variogram can incorporate several important geological characteristics of a deposit in the estimation process. In order to use the experimental variogram in practical applications, the information it conveys must be quantified by fitting a smooth curve (model variogram) to the experimental variogram data points. The model variogram is based on a numerical equation and the numerical parameters are used to control various factors of geostatistical grade interpolation. There are a number of standard models that are used.
As the experimental variogram is based on a variance function and the variances must be positive, the model used must be such that all the values calculated from it are positive. It is best to use a model that has been found, from experience, to be representative of the spatial variation that exists in ore deposits. The spherical scheme model is most widely used. The variography parameters as reported by SRK and reproduced by GAA, is shown per area from Table 24 to Table 34.
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Table 24: KTO Mine except Etang North: Omni-directional Variography Parameters Lithology DSTRAT RSF RSC SDB BOMZ
Variable %TCu %TCo %TCu %TCo %TCu %TCo %TCu %TCo %TCu %TCo
C0 Nugget Variance 0.01 0.00 0.031 0.003 0.043 0.004 0.124 0.005 0.042 0.021
C1 Variance 1.26 0.07 1.97 0.12 2.29 0.11 3.26 0.17 1.83 0.18
Range 1 X Direction 9.96 11.87 3.92 3.05 7.83 7.52 6.98 10.96 7.93 56.65
Range 1 Y Direction 9.96 11.87 3.92 3.05 7.83 7.52 6.98 23.97 7.93 170.00
Range 1 Z Direction 5.40 8.40 3.92 8.50 6.70 7.20 6.98 4.96 7.80 11.68
C2 Variance 1.83 0.03 2.82 0.12 5.08 0.11 5.31 0.27 1.39 0.00
Range 2 X Direction 30.98 147.85 30.17 184.05 52.19 70.59 74.36 31.98 272.10 0.00
Range 2 Y Direction 30.98 147.85 30.17 184.05 52.19 70.59 74.36 324.92 272.10 0.00
Range 2 Z Direction 24.10 18.20 8.50 15.30 27.40 7.20 8.30 11.30 20.30 0.00
C3 Variance 1.44 0.00 1.82 0.00 0.00 0.00 2.43 0.00 0.00 0.00
Range 3 X Direction 324.96 0.00 296.61 0.00 0.00 0.00 189.72 0.00 0.00 0.00
Range 3 Y Direction 324.96 0.00 296.61 0.00 0.00 0.00 189.72 0.00 0.00 0.00
Range 3 Z Direction 24.10 0.00 20.30 0.00 0.00 0.00 27.40 0.00 0.00 0.00
Table 25: Etang North: Variography Parameters
Lithology Variable C0 Nugget C1 Variance Range C2
Variance Range2
OBS %TCu 2.390 2.53 32.00 0.00 0.00 RSC %TCu 0.418 3.30 37.00 0.00 0.00 OBI %TCu 1.263 0.09 40.00 0.57 80.00 OBS %TCo 0.078 0.24 7.50 0.18 30.00 RSC %TCo 0.138 0.13 28.00 0.00 0.00 OBI %TCo 0.085 0.07 46.00 0.00 0.00 Table 26: T-17 Open Pit: Omni-directional Variography Parameters
Zone Variable C0 Nugget Variance C1 Variance Range C2 Variance Range2
RATGRIS
%TCu
0.898 5.21 75.00 0.00 0.00 DSTRAT 3.951 4.54 75.00 0.00 0.00 RSF 4.574 2.15 75.00 0.00 0.00 RSC 7.200 4.56 58.00 1.23 73.00 SDB 9.000 6.70 150.00 0.00 0.00 BOMZ 9.200 1.00 100.00 5.29 150.00 RATGRIS/DS
%TCo
0.328 0.54 300.00 0.00 0.00 RSC 0.654 0.07 75.00 0.25 150.00 BOMZ 0.026 0.10 80.00 0.08 150.00 SDB 0.350 0.90 35.00 0.08 150.00
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Table 27: KOV Open Pit: Omni-directional Variography Parameters for Virgule Lithology Variable C0 Nugget Variance C1 Variance Range BOMZ
%TCu
1.65 3.38 309.38 SDB 8.74 3.78 380.12 RSC 8.59 3.44 209.69 RSF 2.44 0.68 245.78 DSTRAT 3.06 2.51 243.89 BOMZ
%TCo
0.03 0.07 204.17 SDB 0.05 0.13 204.08 RSC 0.08 0.07 266.89 RSF 0.00 0.01 217.35 DSTRAT 0.01 0.01 317.03
Table 28: KOV Open Pit: Omni-directional Variography Parameters for Oliviera Lithology Variable C0 Nugget Variance C1 Variance Range BOMZ
%TCu
0.00 2.68 289.22 SDB 3.15 1.17 303.40 RSC 3.19 1.56 261.88 RSF 1.01 2.90 265.60 DSTRAT 0.28 1.12 350.92 BOMZ
%TCo
0.04 0.06 284.57 SDB 0.01 0.03 192.65 RSC 0.03 0.01 221.57 RSF 0.00 0.02 178.66 DSTRAT 0.02 0.09 253.91
Table 29: Mashamba East Open Pit: Omni-directional Variography Parameters Lithology Variable C0 Nugget
Variance C1 Variance Range
BOMZ
%TCu
0.171 2.73 353.88 SDB 1.235 1.56 216.8 RSC 1.006 0.62 210.88 RSF 1.943 1.43 157.67 BOMZ
%TCo
0.022 0.03 467.49 SDB 0.051 0.11 291.64 RSC 0.077 0.04 256.14 RSF 0.042 0.09 332.59
Table 30: Tilwezembe Open Pit: Variogram Parameters for Manganiferous Dolomites
Variable (%)
C0 Nugget
Variance
Structure 1 Structure 2 Direction (Major, Semi-major, Minor)
C1 Variance
Range Dir 1 (m)
Range Dir 2 (m)
Range Dir 3 (m)
C2 Variance
Range Dir 1 (m)
Range Dir 2 (m)
Range Dir 3 (m)
%TCu 0.050 0.52 40 20 7 0.43 160 60 14 70,160,-40
%TCo 0.040 0.59 40 15 12 0.37 130 70 12 70,160,-40
%TMn 0.050 0.63 30 10 8 0.32 160 70 12 70,160,-40
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Table 31: Tilwezembe Open Pit: Variogram Parameters for Breccia
Variable (%)
C0 Nugget
Variance
Structure 1 Structure 2 Direction (Major, Semi-major, Minor)
C1 Variance
Range Dir 1 (m)
Range Dir 2 (m)
Range Dir 3 (m)
C2 Variance
Range Dir 1 (m)
Range Dir 2 (m)
Range Dir 3 (m)
%TCu 0.120 0.13 20 70 10 0.75 160 70 10 70,160,-60
%TCo 0.110 0.22 40 10 3 0.19 40 100 12 70,160,-60
%TMn 0.220 0.48 30 50 7 0.30 180 50 7 70,160,-60
Table 32: Tilwezembe Open Pit: Variogram Parameters for Tillites and Argillites
Variable (%)
C0 Nugget
Variance
Structure 1 Structure 2 Direction
(Major, Semi-major, Minor)
C1 Variance
Range Dir 1 (m)
Range Dir 2 (m)
Range Dir 3 (m)
C2 Variance
Range Dir 1 (m)
Range Dir 2 (m)
Range Dir 3 (m)
%TCu 0.150 0.4 110 80 15 0.44 125 110 15 80,170,-45
%TCo 0.090 0.18 180 50 2 0.73 180 50 20 80,170,-45
%TMn 0.120 0.53 35 90 12 0.36 160 90 12 80,170,-45
Table 33: Kananga Mine: Variogram Parameters for Upper Orebody
Variable (%)
C0 Nugget
Variance
Structure 1 Structure 2 Direction (Major, Semi-major, Minor)
C1 Variance
Range Dir 1 (m)
Range Dir 2 (m)
Range Dir 3 (m)
C2 Variance
Range Dir 1 (m)
Range Dir 2 (m)
Range Dir 3 (m)
%TCu 0.100 0.62 60 90 11 0.36 300 90 11 65,155,-40
%TCo 0.060 0.61 50 20 9 0.33 300 70 9 55,145,-35
%TMn 0.230 0.29 60 10 9 0.49 180 140 9 55,145,-35
Table 34: Kananga Mine: Variogram Parameters for Internal/Middle Zone
Variable (%)
C0 Nugget
Variance
Structure 1 Structure 2 Direction (Major, Semi-major, Minor)
C1 Variance
Range Dir 1 (m)
Range Dir 2 (m)
Range Dir 3 (m)
C2 Variance
Range Dir 1 (m)
Range Dir 2 (m)
Range Dir 3 (m)
%TCu 0.120 0.41 50 20 10 0.47 460 230 25 65,155,-40
%TCo 0.020 0.47 70 50 11 0.51 340 290 27 55,145,-35
%TMn 0.050 0.53 50 130 13 0.41 360 130 13 55,145,-35
9.3 Estimation Parameters 9.3.1 Estimation Plan There are a number of methods used in grade estimation. These methods may be based on the presence or absence of a variogram. For data that is spaced too far apart, or where a variogram cannot be generated, the inverse distance method of interpolation may be appropriate.
Kriging is another method that may be used. Kriging allocates weights based on the distance of the sample from the point or block being estimated, to the sample points surrounding the point or block for which grade is to be estimated. By allocating these weights to the samples, it ensures that the estimation error is
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minimised, ensuring the best estimation possible for that block. The error associated with estimation is called the kriging variance.
Ordinary Kriging (OK) was used for the purpose of this project, which is a specific algorithm that satisfies unbiased by ensuring that the Kriging weights in the local estimation are summed to 1. The results from the variography analysis were used during the estimation of each project area.
9.3.1.1 Density Assignment and Determinations Historically, GCM assigned density values based on the categorization of the ore type into dolomitic or non-dolomitic (siliceous) and its copper grade and based on an exhaustive dataset available from all GCM operations within the Katangan Copperbelt. A sample was considered dolomitic when the %TCu divided by the %CaO in the sample was less than or equal to 15 and non-dolomitic (siliceous) when it was greater than 15. GCM generalized empirical criterion defined three main categories of densities as shown in the table below.
Table 35: GCM criteria for assigning density values Definition and Criterion Density (t/m3)
Siliceous= (%TCu/%CaO) >=15 TCu>1<2% 2.00 TCu>2.0%. with <0.5% Cu Sulphite content 2.20 TCu>1<2%, with >1% TCo content 2.20 TCu >2.0%, with >0.5% Cu Sulphite content 2.40
Dolomitic = (%TCu/%CaO) <15 TCu >1<2% 2.40 TCu >2.0%, with <0.5% Cu Sulphite content 2.40 TCu >2.0%, with >0.5% Cu Sulphite content, >=0.5%Cu Oxide 2.40 TCu >2.0%, with >0.5% Cu Sulphite content, <=0.5%Cu Oxide 2.60
According to the GCM criterion, waste rock was generally assigned a density of 2.00t/m3 if it was siliceous and 2.40 t/m3 if the rock was considered dolomitic.
During the 2008 feasibility study, SRK reviewed the historical assayed dataset for all the projects in the application of these criteria and found that there were proportionately fewer assays for %CaO than the %TCu assays available for these criteria to be applied. However, SRK consider these values as guidelines for the possible ranges of density within the respective mineralized zones.
9.3.1.2 KTO Mine CCIC undertook limited density determinations of the various stratigraphic units to verify GCM empirical density values. The determinations were undertaken on selected lithological cores using Archimedes’ Principle, by which a sample is weighed in air and then in water using a Clover Scale. The measured masses then are entered into a simple formula to calculate the density. CCIC limited density determinations for KTO Mine are presented in the table below by stratigraphic unit.
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Table 36: KTO Mine: Density Determinations on various lithologies
Stratigraphic Unit Number of samples Minimum
(t/m3) Maximum
(t/m3) Average
(t/m3)
SD1a 9 2.69 2.90 2.80 BOMZ 8 2.74 2.92 2.86 RSC 8 2.51 2.96 2.69 RSF 6 2.57 3.03 2.81 DSTRAT 5 2.66 3.02 2.81 Grey RAT 3 2.64 2.77 2.70 Red RAT 3 2.63 2.75 2.67
On the basis of the limited density determinations, CCIC concluded that GCM approach was conservative and that upside potential existed with regard to the calculated resource tonnages, but recommended that bulk density determinations should be undertaken before higher density values can be used in the Resource Model.
CCIC used the average density values of 2.70t/m3 from the GCM table for the conversion of volume to tonnes for the KTO Mine model.
KCC personnel undertook a new density determination study on the new diamond drilling core from the recent exploration programme of Etang North. The results from this study are shown in Table 37.
These densities were applied in the new Etang North model.
Table 37: New Density Determinations for Etang North Stratigraphic Unit Number of samples Minimum (t/m3) Maximum (t/m3) Average (t/m3)
BOMZ 16 2.40 2.70 2.51 BOMZATRE 16 2.40 2.70 2.54 SDB 17 2.40 2.70 2.50 RSC 17 2.00 2.60 2.29 RSF 15 2.00 2.40 2.20 DSTRAT 15 2.20 2.60 2.48 RATGRIS 14 2.00 2.60 2.32
9.3.1.3 T-17 Open Pit CCIC undertook limited density determinations of the various stratigraphic units to verify GCM empirical densities. The determinations were undertaken on selected lithological cores using the Archimedes’ Principle by which a sample is weighed in air and then in water using a Clover Scale. The measured masses then are entered into a simple formula to calculate the density. CCIC limited density determinations for T-17 Open Pit are presented in the table below by stratigraphic unit.
Table 38: T-17 Open Pit: Density Determinations on Various Lithologies Stratigraphic Unit Number of samples Minimum (t/m3) Maximum (t/m3) Average (t/m3)
SDB 7 2.10 2.76 2.38 BOMZ 1 2.09 2.09 2.09 SDB 7 2.10 2.76 2.38 RSC 5 2.21 2.63 2.34 RSF 6 2.06 2.51 2.32 DSTRAT 5 1.88 2.40 2.13
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As a cross check and by way of a second method, CCIC also submitted samples for density checks to Set Point Laboratories where density determinations were undertaken using a multivolume gas pycnometer 1305 for helium displacement.
Two samples were also tested by the Mintek laboratory in Johannesburg and these provided figures of 2.84t/m3 for the siliceous material from Musonoie-T-17 West, and 2.74t/m3 for the dolomitic material from the same Resource Area. The method of density determination undertaken by Mintek has not been specified in the CCIC report.
On the basis of the limited density determinations, CCIC indicated that GCM approach was conservative and upside potential existed with regard to the calculated resource tonnages. However, CCIC recommended that bulk-density determinations should be undertaken before higher density values can be used in the Resource Model.
For the conversion of volume to tonnage in the T-17 Open Pit model, CCIC applied density values of 2.20t/m3 and 2.40t/m3 consistent with the GCM categories of oxide and mixed ore types.
KCC personnel undertook a new density determination study on the new diamond drilling core from the recent exploration programme of T-17. The results from this study are shown in Table 39.
These densities were applied in the updated T-17 model.
Table 39: New Density Determinations for T-17 Stratigraphic Unit Number of samples Minimum (t/m3) Maximum (t/m3) Average (t/m3)
BOMZ 8 1.20 2.80 2.09 SDB 18 2.00 2.80 2.32 RSC 18 2.10 2.50 2.33 RSF 17 2.00 2.50 2.29 DSTRAT 15 1.70 2.80 2.33 RATGRIS 11 2.10 2.70 2.36
9.3.1.4 KOV Open Pit In the models generated for KOV Open Pit, SRK used an inferred density of 2.20t/m3. The inference is based on the visual inspection of the mineralization in the cores and observations in the field, where the predominant copper mineral is malachite.
Intersections of mineralization from the drilling at KOV Open Pit confirm that the predominant mineralization is malachite, considered as an oxide, with minor sulphides at depth. There are limited density determinations from selected cores of the recent drilling. Although considered statistically inadequate to represent the sample dataset, indications from these determinations are that the density applied is appropriate.
9.3.1.5 Mashamba East Open Pit CCIC undertook limited density determinations of the various stratigraphic units to verify GCM empirical density values in the Mashamba East Open Pit. The method is as described above under Kamoto Mine. CCIC’s limited density determinations for Mashamba East Open Pit are presented in the table below, by stratigraphic unit.
Table 40: Mashamba East Open Pit: Density Determinations on Various Lithologies Stratigraphic Unit Number of samples Minimum (t/m3) Maximum (t/m3) Average (t/m3) SDB 17 2.34 2.76 2.52 RSC 10 2.40 2.61 2.51 RSF 5 2.28 2.50 2.39
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CCIC used the average density values of 2.20t/m3 and 2.40t/m3 from the GCM table for the conversion of volume to tonnes in siliceous and dolomitic mineralized zones respectively for the for the Mashamba East model.
9.3.1.6 Kananga Mine The procedures adopted for the density determinations at Kananga are the same as described for Tilwezembe and the values are listed in the table below.
Table 41: Kananga Mine: Density Determinations on Various Lithologies Domain Declustered mean (t/m3 ) Upper ore body oxides (UOB_OX) 1.80 Middle low-grade oxides (MID_OX) 1.80 Lower ore body oxides (LOB_OX) 2.00 Upper ore body sulphides (UOB_SL) 2.10 Middle low-grade sulphides (MID_SL) 2.00 Lower ore body sulphides (LOB_SL) 2.10
9.3.1.7 Tilwezembe Open Pit Snowden undertook density determinations on selected core samples using Archimedes’ Principle. The sample core pieces of approximately 100mm to 200mm length were wrapped in cling-film (Saran wrap) to prevent oxidation and weighed first in air and then when submerged in water. The difference in the weights is the weight of the water displaced.
No work has been done to determine the free moisture content of the samples. Resultantly, wet bulk densities were used during estimation.
Snowden indicate that no relationship exists between grade and density and therefore, bulk density factors were determined for each geological unit from the means of the specific gravity measurements after outliers were cut from the dataset. The de-clustered means were used and a maximum of 5% of the composites were cut from the dataset. The composite data was declustered using a cell size of 25m E by 25m N by 1m RL that approximates the drill-hole spacing. The bulk densities are presented in the table below.
Table 42: Tilwezembe Open Pit: Density Determinations on Various Lithologies
Domain Bottom Cut Top Cut Percentage cut
Declustered mean (before cut) (t/m3 )
Declustered mean (after cuts) (t/m3 )
Ox_MnDol 1.1 3.0 5 2.04 1.96 Ox_Brec 0.0 2.7 4 1.90 1.81 Ox_TillArg 0.0 3.0 5 2.09 1.98 Sl_MnDol 0.0 2.6 3 2.28 2.26 Sl_Brec 1.5 2.7 3 2.23 2.24 Sl_TillArg 1.8 2.5 3 2.18 2.18
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9.4 Summary Table 43 below indicates the densities that have been used in the conversion of volume to tonnes within the various project areas.
Table 43: Kananga Mine: Density Determinations on Various Lithologies Project Area Mineralized Zone Density (t/m3) T-17 Open Pit All See Table 39
Tilwezembe Open Pit
Ox_MnDol 1.96 Ox_Brec 1.81 Ox_TillArg 1.98 Sl_MnDol 2.26 Sl_Brec 2.24 Sl_TillArg 2.18
KTO Mine All except Etang North 2.70 Etang North See Table 37
Kananga Mine
Upper ore body oxides (UOB_OX) 1.80 Middle low-grade oxides (MID_OX) 1.80 Lower ore body oxides (LOB_OX) 2.00 Upper ore body sulphides (UOB_SL) 2.10 Middle low-grade sulphides (MID_SL) 2.00 Lower ore body sulphides (LOB_SL) 2.10
KOV Open Pit All 2.20
Mashamba East Open Pit Oxide mineralized zones 2.20 Mixed mineralized zones 2.40
10.0 DRILLING 10.1 KTO Mine GCM carried out both extensive surface and underground drilling to delineate the KTO orebodies. A total of 83 surface boreholes have been identified – drilled between 1952 and 1991. Underground holes were generally drilled as fans of 3 or more holes from especially mined-out cubbies. A total of 569 holes have been identified, drilled between 1972 and 2002. The upper parts of the Etang orebody – above 400 levels – are covered only by surface drilling. Borehole surveys appear to have been carried at regular intervals for surface boreholes, but deviations are rarely more than a few degrees. Historically, underground collared boreholes have not been surveyed.
KCC drilled 26 diamond drilling in the Kamoto underground mine in 2009 and 2010 to confirm the geological potential of the fragment of Etang North. The majority of these holes intercepted the copper-cobalt mineralization as originally planned by Gecamines in the various reports.
The copper grades of Bomzatre are relatively low compared to those announced by Gecamines.
The synthesis of these boreholes is shown in the Table 44.
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Table 44: Holes drilled in Etang North Section Hole_id Depth OBS RSC OBI %Cu %Co
ETANG NORTH
T4
F2513 87 34.94-45.21 45.21-71.17 71.17-80.75 1.23 0.35
F2514 99 37.68-48.37 48.37-79.77 79.77-94.34 1.34 0.41
F2516 150 47.45-57.94 57.94-100.72 100.72119.15/ 123.57-130.15
1.10/ 1.41
0.36/ 0.25
F2518 168 64.46-79.98 79.98-134.68 134.68-162 1.96 0.52
F2519 86 NI NI NI 0.00 0.00
F2520 63 NI NI NI 0.00 0.00
F2521 222 94.78-156.26 156.26-208.89 NI 2.11 0.39
T6
F2491 81 NI NI NI 0.00 0.00
F2490 132 78.00-86.95 86.95-117.97 117.97-131.23 1.47 0.34
F2489 147 82.71-92.06 92.06-115.29 115.29-137.16 1.76 0.35
F2488 168 97.70-110.79 110.79-140.50 140.5-159.32 2.32 0.46
T9
F2512 252 216.00-246.84/ 247.19-248.78 246.84-247.19 0-25.94 2.65/
3.34 0.42/ 0.19
F2510 220 138-154.81 154.81-187.24 0-15.22/ 187.24-210 1.80/3.08 0.36/
0.14
F2509 165 108.68-119.55 119.55-147.06 0-10.49/ 147.06-159.70 1.27/3.07 0.46/
0.12
F2508 150 108-122.58 122.58-144.95 0.00-8.00 1.35/ 3.12
0.47/ 0.73
F2511 91 59.05-83.94 1.49-59.05 0.00-1.49 1.83 0.74
T11
F2489 147 82.71-92.06 92.06-115.29 115.29-127.40/ 131.36-137.16
1.88/ 2.01
0.38/ 0.34
F2497 194 127.80-143.00 143.00-164.06 0.47-13.16/ 164.06-182.00
1.77/ 4.10
0.33/ 0.05
F2504 93 NI 19.95-58.15 0-19.95 1.77 0.29
F2499 237 165.97-186.92 38.02-39/ 186.92-220.45
0.80-38.02/ 220.45-230.27
4.19/ 2.34
0.05/ 0.54
F2517 273 237.69-268.52 43.39-49.59 1.95-43.39 4.00/ 0.78
0.56/ 0.32
F2495 168 111.56-126.30 126.30-148.65 148.65-155.97 1.43 0.44
T-17
F2501 113 NI NI NI 0.00 0.00
F2500 201 120.1-142.85 142.85-174 174-196 1.96 0.41
F2496 126 78.51-92.54 92.54-107.72 107.72-120 2.48 0.75
F2498 147 91.16-107.65 107.16-129.55 129.55-143.02 2.19 0.48 NI = No Intercept
10.2 T-17 Open Pit The T-17 orebody has been the subject of 2 diamond drilling programs by GCM, with 3,287m drilled between 1938 and 1954, and 8,011m drilled from 1986 to January 1988. The holes were drilled generally to a nominal 100m by 100m grid, with certain areas being on 50m by 50m spacing.
KCC drilled 20 holes in 2009 and 2010 within the current pit perimeter to confirm the continuation of the resource at depth. A total of 4,286m has been drilled, and the result from this drilling indicates that mineralization continues down-dip to a depth of more than 180m below the current pit design floor. Grades in these intersections have been encouraging, with the majority of intersect samples returning assay values of greater than 2.00%TCu. Further evaluation of this drilling data is currently being conducted to categorise this resource and allow for the conceptual mine plan to be developed to a level of confidence.
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In addition to the drilling referenced above, KCC have also recently drilled a total of 20 holes to the east of the current T-17 Open Pit perimeter to confirm an extension of the ore body along strike. The analysis of cores from this drilling campaign indicated significant intersections as shown in Table 45. This is leading to the delineation of additional mineral resources and a subsequent increase in the life of the current pit. Table 45: Summary of 2010 drill data for T-17
NI= No Intercept * N/A= Not Applicable
10.3 KOV Open Pit The drill-hole logs indicate that exploration drilling commenced in the early 1940s on the Kamoto East ore body, and initial holes were prefixed as KTO. Although the target was the Kamoto East ore body, a substantial number of KTO holes were later drilled into the present-day KOV Open Pit.
In the 1980s, another drilling campaign was aimed at defining the KOV Open Pit mineralized zones, and this campaign continued into the early 1990s. The drilling was carried out along section lines spaced about 100m apart. Where feasible, drill holes were spaced about 100m along these section lines. The holes were prefixed KOV Open Pit.
The KOV Open Pit drill-hole database contains a total of 214 drill-holes spaced on average about 100m. There are 100 intersections of the Virgule fragment, 75 for the Oliveira, 33 for Kamoto East and 19 for the FNSR. The demarcation between Virgule and Kamoto East is based on a boundary string file obtained from the GCM sections. However, for the modelling and grade estimation, certain drill holes overlap into the Virgule and Kamoto East and are therefore counted twice. Similarly, there are holes intersecting Virgule that also intersect FNSR and these are also counted twice.
T-17
Hole id Depth OBS RSC OBI %Cu %Co
T-17001 113.4 34.9-53.5 91-108.6 53.5-91.3 3.25 0.31
T-17002 136.8 45.7-70.7 110.2-129.55 70.7-110.2 3.10 0.11
T-17003 94.8 16.1-32 71.3-93 32.0-71.3 2.09 0.25
T-17004 88.9 1.2-14.3 56.1-77.4 14.3-56.1 2.27 0.57
T-17005 116.0 34.3-50 50-116 No Int 3.08 0.31
T-17006 200.7 4.8-108 172.6-194.7 108-172.6 3.79 0.38
T-17007 125.3 No Int No Int No Int N/A N/A
T-17008 68.7 No Int 44.7-64.5 12-44.7 2.43 0.48
T-17009 89.5 3-32.6 32.6-71.2 71.2-78.5 2.55 0.29
T-17010 151.7 23.1-38.1 38.1-56.6 56.6-83.5 3.20 0.54
T-17011 240.0 109.60-141 141-196.4 196.4-217 2.92 0.45
T-17012 265.6 176.35-202 202-232.4 232.4-254.30 3.49 0.67
T-17013 348.6 102.2-228.50 54.20-96.80 35.8-54.20 5.28 1.11
T-17014 333.1 167.50-236.70 129.45-167.50 93.50-126.50 4.76 1.75
T-17015 263.1 71.45-181.10 / 236.50-251.75
29.50-71.45 / 251.75-263.10 4-29.50 4.39 0.52
T-17016 275.0 107-125 61-107 35-61 2.50 0.51
T-17017 229.9 149.70-181.30 117.30-149.7 90.8-117.30 5.27 0.48
T-17018 392.3 269.80-352.20 215.20-269.80 184.50-215.20 3.54 0.70
T-17019 349.1 95.30-112.1 / 290.90-305.70
72.80-95.30 / 305.7-315.60
53.90-72.80 / 315.60-334.70 3.83 2.87
T-17020 355.4 123.40-146.70 / 297-315
102.60-123.40 / 315-329.90
97.60-102.6 / 329.9-346 1.95 0.52
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There is adequate drill-hole coverage for the Virgule and Oliveira while the FNSR and Kamoto East intersections are limited (excluding recent drilling in Kamoto East reported below). The extent of the FNSR is limited as it is a remnant of the fragment mined in the Musonoi Pit, to the east of the KOV Open Pit, and the data distribution is therefore relatively adequate. The data distribution compared to the extent and volume of the Kamoto East fragment is inadequate, especially considering that the bulk of the data are well above the current pit bottom and there are limited drilling intersections in this steeply dipping limb.
Most of the historical drill holes within the Kamoto East and the KOV Open Pit areas were drilled vertically, with only a few being inclined. Kamoto East drilling was problematic due to the steepness in the dip of the strata. As a result, the majority of the holes intersect the near surface expression of the Kamoto East ore body, and only the inclined holes provide intersections at depth.
In general, the majority of the historical drill hole intersections in Kamoto East were within the areas that have been subsequently mined out, and there are very few ore-body intersections below the current pit bottom.
During 2010, a total of 10 holes (3,461m) were drilled in the Kamoto East orebody by KCC. The majority of these holes intercepted the ore body either at or below the originally GCM drilled areas which were reported by SRK in the 2009 Technical Report and the data from this drilling campaign is within +/- 5% of the GCM resource model with regards to location and grade.
Table 46: Summary of 2010 drill data for Kamoto East
Kamoto East
Hole_id Depth (m) OBS RSC OBI %Cu %Co
KOV0601 319.9 208.4-225.80 225.8-243.60 243.6-247.6 4.49 0.32
KOV0602 250.5 NI. NI. NI. 0.00 0.00
KOV0603 317.1 221.9-236.8 236.8-252.8 252.8-263.1 6.61 0.31
KOV0604 325.0 216.13-236.48 236.48-252 252-274.14 4.22 0.22
KOV0617 430.9 353-429.1 NI. NI. 2.29 1.77
KOV0618 442.8 394.2-422 422-440.25 440.25-441.8 5.23 0.78
KOV0619 500.0 413.6-455 455-472.5 472.5-478.1 3.87 0.71
KOV0620 241.0 NI NI NI 0.00 0.00
KOV0621 402.3 NI NI NI 0.00 0.00
KOV0622 231.9 389.9-419.4 419.4-431.9 NI 4.42 0.32
Average 346.1 N/A N/A N/A 3.11 0.44
Total 3461.4 N/A N/A N/A N/A N/A
NI= No Intercept * N/A= Not Applicable
KCC have recently drilled 11 holes (1,577m) in the Virgule orebody (Cut 1a and Cut 2 of the scheduled mine plan) on the North West section of KOV Open Pit which confirmed location and grade of the resource model which was based on GCM data. Following this drilling, the mine plan was implemented and to date there has been good reconciliation between grade control assays and the resource model.
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Table 47: Summary of 2010 drill data for Virgule Hole_id Depth (m) OBS RSC OBI %Cu %Co
Virgule
KOV0605 140.0 92.8-101.2 101.2-117.85 117.85-135.1 5.17 0.22
KOV0606 312.6 NI NI. NI. 0.00 0.00
KOV0607 65.0 NI NI NI. 0.00 0.00
KOV0608 100.5 Leached Leached 55.1-82.4 1.45 0.07
KOV0609 99.0 Leached Leached 80.9-91.4 3.55 0.06
KOV0610 80.7 Leached Leached 55.2-71 3.91 0.23
KOV0611 122.5 Leached 82.4-101.9 101.9-119.8 3.42 0.34
KOV0612 160.5 106.5-125.8 125.8-144.8 144.8-158.7 5.97 0.14
KOV0614 90.2 34.6-40.1 40.1-58.9 58.9-84.9 2.62 0.07
KOV0615 100.5 Leached 46.8-64.9 76.5-100.2 5.86 0.12
KOV0616 306.0 273.3-248.4 248.4-264.6 264.6-281.7 3.38 0.48
Average 143.4 N/A N/A N/A 3.21 0.16
Total 1577.5 N/A N/A N/A N/A N/A
KCC has recently drilled 10 holes (2,262.3m) in the Oliveira Ore Body (Cut 2 of the schedule mine plan) on the north section of KOV Open Pit which confirmed the location and grade of the mineral resource model which was based on GCM data. Following this drilling, the mine plan was implemented and to date there has been good reconciliation between grade control assays and the mineral resource model.
Table 48: Summary of 2010 drill data for Oliviera Hole_id Depth (m) OBS RSC OBI %Cu %Co
KOV Oliviera
KOV0623 242.6 179.80-202.30 202.30-221 221-231.70 3.61 0.59
KOV0624 192.0 127.90-149.1 149.1-168.8 168.8-187.50 3.18 0.88
KOV0625 162.9 108-124.80 124.80-142.30 142.30-156.10 3.08 1.08
KOV0626 261.3 188.70-201.2 201.2-222.80 222.80-240.30 3.41 0.61
KOV0627 272.0 209-222.50 222.50-242.30 242.30-255.10 3.07 0.46
KOV0628 237.5 176.90-181.7 181.70-211.75 211.75-223.10 1.30 0.42
KOV0629 232.2 201.10-206.6 206.60-217.60 217.60-226.70 6.46 0.92
KOV0630 201.3 NI NI NI 0.00 0.00
KOV0631 211.5 156.50-167.20 167.2-196.80 196.80-210.50 1.93 0.54
KOV0632 249.0 193.52-207.17 207.17-227.90 227.90-241.74 3.63 0.67
Average 226.2 N/A N/A N/A 2.97 0.62
Total 2262.3 N/A N/A N/A N/A N/A
10.4 Mashamba East Open Pit For the Mashamba East Open Pit, drilling was undertaken from surface on a 100m by 100m spaced grid on a local co-ordinate system parallel to the strike of the mineralized zone. A total of 122 holes were drilled, comprising 20,466.4m drilled, with 1,788m intersecting copper mineralization of more than 1.00%TCu.
No new exploration drilling was done during 2011.
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10.5 Kananga Mine The historical drilling at Kananga was limited to about 8 holes drilled by GCM. Details of the drilling, sample collection and sample analyses for the holes are not available.
Currently there are 52 holes drilled in Kananga captured in the database.
No new exploration drilling was done during 2011.
10.6 Tilwezembe Open Pit The historical drilling was undertaken by GCM, and the information was the basis for the first pass mineral resource estimates for Tilwezembe Open Pit undertaken by SRK. The drilling information included drill holes on a 25m spacing within the operational Tilwezembe Open Pit and 100m by 50m on Tilwezembe East. A total of 157 holes have been drilled at Tilwezembe.
No new exploration drilling was done during 2011.
11.0 SAMPLING METHOD AND APPROACH 11.1 Historical Sampling Details of the historical sampling undertaken within each of the project areas are scant and based on personal communications with the respective consultant in each of the project areas. Inferences have been drawn from the observations from the sample database.
Cores from the ore body intersections were sampled for chemical analysis. The lengths of core sampled varied, and it is understood that this was a consequence of the sample recovered within each run. In the GCM logging sheet, there is a column for percentage recovery where values ranging from 1% to 100% are entered to describe the amount of core recovered in the sample length. Core recoveries are recorded only for cores that were sampled.
The lithologies sampled were the upper ore-body host rocks (lower SDS and SDB) and the lower ore-body rocks (RSF, DSTRAT and the RATGR) and portions of the RSC deemed to be mineralized. SRK understands that the visibility of copper mineralization in the core was used as the criterion for sampling the core. Core lengths deemed to be barren of copper were not sampled, and an entry was made in the sample log for that interval with the comment “steriles” or barren. It is possible, in SRK’s view, that the unsampled cores could contain finely disseminated copper mineralization not visible to the naked eye. There is a further possibility, especially in the RSC, that the “sterile” zones contain cobalt mineralization. In drill holes KOV 426 and KOV 427, the entire RSC is mineralized and returned good copper mineralization (2-3%) within the mid-RSC. In drill hole KOV 428, the mid-portion of the RSC was sampled. Partial or selective sampling, although common in the RSC, was also evident in the other Roan lithologies.
The assay database describes the sample in terms of the length, depths (from and to) of intersection and the amount of core recovered in that sample length. The sample database contains assay data for the following:
%TCu: the percentage total copper content of the sample;
%CuO: the percentage of the copper present as oxide. In the modelling, this is reported as %ASCu. Fewer than half of the samples were analyzed for %ASCu;
%Cu mal: the percentage of the copper as malachite. Only a few samples contain values on this column;
%TCo: the percentage total cobalt content of the sample; and
%CaO soluble: the relative proportion of soluble calcium oxide in the sample. Less than 30% of the total database was assayed for calcium oxide.
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11.2 Current Sampling Sampled zones are selected based on the visual observation of the lithological contacts. The geologist also marks on the core the direction along which the core should be split, after considering the attitude of the bedding or foliation relative to the core axes. The drill lengths and the recoveries are recorded in the sampling notebooks.
Sampling is carried out at a maximum of 1m drill length intervals and different stratigraphic units are sampled separately. The core samples are sawed into two halves. One half is broken up and bagged for assay while the other half is stored for future reference.
Core bags for a particular batch are pre-labelled and arranged in order from the first to the last sample. A tag with an identification or sample number is added to the bag containing the sample before the bag mouth is tied.
Split core sampling is done from the drill core. Prior to taking samples, the geologist examines the core and marks off the intervals to be sampled by drawing a line along the core with a marker pen. When the intervals have been selected, the core is split in half using a diamond saw or core splitter. Once the core is split; individual sample lengths are selected taking care to note stratigraphical and lithological boundaries. The whole width of mineralization and at least one metre of apparently barren or low grade hanging wall and foot wall material are covered.
The data is recorded as preliminary in the log sheets and is then transferred into the geological database (GDMS), logged in Lakefield’s sample-tracking system and stored on a shelf. The splits and resubmitted pulps are currently stored at SGS Lakefield and the check sample pulps at Set Point.
The analytical method used for the determination of total copper and total cobalt was X-ray fractionation. Acid-soluble copper and cobalt were determined by acid digestion (sulphuric acid) and analysis of the solution by Atomic Absorption Spectrometry (AAS).
The methods are described as:
For analysis of copper oxides each sample was weighed and mixed with an aliquot of dilute sulphuric acid enriched with sulphur dioxide. This mixture was agitated at room temperature for a set period and the sample residue filtered out of the solution. The solution was made up to volume and analyzed for copper and cobalt by AAS. This yielded acid-soluble results.
For analysis of copper sulphides the residue of the copper oxide preparation was placed in a beaker and mixed with multiple acids, with the residue being digested in the acid mixture. The solution was made up to volume and analyzed for copper and cobalt by AAS. This yielded an assay of acid-insoluble copper (AICu) and acid-insoluble cobalt (AICo) present as sulphides.
11.2.1 Data Quality and Quantity Historically, the core recoveries are in general low throughout the various project areas. This is mainly because of the fractured nature of the ore, as well as the drilling techniques used.
Table 49: Historical Core Recoveries per Area Area Average Core Recovery (%)
Kamoto 65 T-17 63 KOV 42 Mashamba East 71 Tilwezembe 80 Kananga 86
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The bulk of the data within the project areas is historical, with the exception of the recent drilling data reported above which has validated the GCM data. .
It is clear that modern drilling techniques improve core recovery.
The recent drilling campaigns in T-17 and KOV Open Pits and KTO provide confirmation of this fact with the average core recovered in T-17 Open Pit increasing to 82%, in KOV Open Pit increasing to 79% and in KTO to 91%. This is clearly an indication that current drilling techniques are much more focused towards sample recovery.
Using the historical low core recoveries in a resource model, there are two options to account for core loss:
Adjustment of the assay grades to account for core loss and regard the adjusted data as representative; and
Assume the assay grades of the recovered core represent the sample length, but account for the core loss in the classification on the premise of quality of data used in the estimation.
Adjustment of grade is considered preferable for recently acquired drill-hole information. As the bulk of the drilling information is historical for the project areas, SRK has considered the option of accounting for core loss in the classification. GAA agrees with this assessment.
11.2.2 Sampling Preparation and Analyses 11.2.2.1 Historical Sample Preparation and Analyses All historical sampling, sample preparation, analysis and security were undertaken by GCM over more than 50 years, and it is therefore difficult to comment on the quality of such work.
The historical core samples were cut along the longitudinal axis with one half of the core sent for laboratory analysis and the remaining half retained in the boxes. There was no systematic approach to sample lengths as indicated by the variations in the sample lengths in the database. The minimum sample taken was 0.5m and the maximum sample was 2.5m. The sample lengths were also a consequence of the sample recovered within the run.
The historical core samples were delivered to the laboratory for further sample preparation and analysis which was undertaken in-house by GCM.
11.2.2.2 Recent Sample Preparation and Analyses All the recent drilling for the feasibility study was done according to JORC standards.
The samples from Kananga and Tilwezembe were sent to two laboratories, Alfred H. Knight (Alfred Knight) in Kitwe and SGS in Ndola, for preparation and analysis. Sample preparation consisted of the following methodology:
Drying of sample;
Primary jaw and roll crushing of sample;
Splitting a sub sample of 250g using a riffle splitter; and
Pulverizing of the sub-sample to 75 μ and homogenizing.
The prepared samples were analyzed for the following variables, namely; %TCu, %TCo, %Mn, %ASCu and %ASCo.
Both Alfred Knight and SGS determined %TCu, %TCo, %Mn assays by multi-acid digestion (using hydrofluoric, nitric and perchloric acids) followed by dissolution in hydrochloric acid and AAS.
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For %ASCu and %ASCo, both laboratories used cold leaching with 5% sulphuric acid. However, Alfred Knight saturated with sulphur dioxide while SGS saturated with potassium sulphite before finishing with AAS. The laboratories may also have used different temperatures and digestion times.
11.3 Current Sample Preparation and Analyses All the current sampling is done by KCC personnel. KCC carries out the sample preparation at Luilu Laboratory which is owned and operated by KCC. Accreditation and certification of the Laboratory to ISO standards is planned for completion by 2014.
Each of the half core samples are crushed to ± 20 mm and then again crushed to ± 5mm. The crushed sample is split where necessary to produce a portion about 250g. The split is then pulverized to 50μ bagged and labelled. It is then submitted to the Laboratory for its respective analysis.
The analytical method used for the determination of total copper, total cobalt and copper oxide is AAS.
Acid soluble copper and cobalt is determined by acid digestion in a blend of nitric acid and hydrochloric acid. Analysis of the solution is by AAS.
For total copper and total cobalt, each sample is weighed and mixed with an aliquot of blended nitric acid and hydrochloric acid in a volumetric flask. The sample residue is filtered out of solution, made up to volume and analysed for copper and cobalt by AAS.
For copper oxide, each sample is weighed and mixed in a blend of ethanol, hydrated tin chloride and hydrofluoric acid. This mixture is agitated at the room temperature for a set period and the sample residue filtered out of the solution. The solution is made up to volume and analysed for copper oxide by AAS.
12.0 DATA VERIFICATION Data validation, which is a routine exercise for KCC, involves checking of hardcopies of printed data, 3D computer validation, and on screen checks in Datamine and Surpac. In addition, physical checks of collar and survey are done by plotting the processed data in section and on plan.
The database system employed at KCC is GDMS which has a built-in data security system. This prevents issues such as data overlaps, duplicates and gaps.
12.1 Quality Assurance and Quality Control Quality assurance and quality control (QAQC) is the methodology by which confidence levels are measured and maintained for assaying.
The main objectives of QAQC programs are to:
Minimize bias from sampling and assaying;
Ensure the accuracy and precision of assaying; and
Measure and demonstrate data integrity and validity for resource estimates and grade control.
12.2 QAQC Procedure The QAQC procedure involves insertion of blanks, duplicates and standard reference material at pre-determined and sequential intervals. Every 10th sample is a blank, every 20th is a duplicate and every 30th sample is a standard reference material with known mean and standard deviations. This cycle is repeated till the end of the hole.
KCC use 3 types of standard reference material; low, medium and high grade material. This ensures the full range of grade categories of both copper and cobalt are covered.
The reference material assays provide a method by which analytical accuracy is monitored and quantified. There are 2 parameters of interest when reporting analytical accuracy:
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The relative assay deviation from the expected value of the reference material; and
The average bias over time.
The deviation of a reference material's assay is measured and expressed as a relative standards deviation, thus making it possible to directly compare reference materials with different standard deviations. Acceptable limits are considered to be 95% of samples submitted to be within ± 2 standard deviations.
When acceptable limits for reference materials are not achieved the following course of actions are taken:
Cross-check KCC reference material assays with laboratory submitted reference material assays for the same period and/or batch;
In the case of bias, determine if it is 1 reference material type or all reference material types; and
After discussions with the Laboratory, an experiment may be undertaken to determine the reason for the variance.
13.0 MINERAL PROCESSING AND METALLURGICAL TESTING There have been significant changes to the KCC plant and processing infrastructure. These have been informed by an evaluation of existing processing techniques which have culminated in Phase 3 and Phase 4 plant and processing improvement project. More detailed description of this is included in paragraphs 17.0.
Iso therm testing was conducted by two independent parties for the SX plant on existing electrolyte suitability, insofar as metallurgical testing is concerned, modified to Phase 4 Project conditions. No testing was performed on electro-winning process as electro-winning downstream of SX is a well proven process.
Two independent testwork reports on milling from previous years were consulted as part of the Phase 4 Project development. Data from the operating plant was used in the flotation process development for Phase Project by SNC Lavalin. Investigations are in progress to confirm the expectation that returning the DIMA mills to semi-autogenous operation will increase the total installed capacity to ~ 9Mtpa.
14.0 MINERAL RESOURCE ESTIMATES As part of this report, GAA reviewed the mineral resource estimation methodologies and results from the estimations as produced by KCC and other independent consultancies.
The mineral resources for all the mines were estimated in accordance with the JORC Code. The following were considered when classifying the mineral resources:
The quantity, quality and age of the data used in the generation of the mineral resources;
The availability of assays in portions of the package due to selective sampling on the basis of visible copper mineralization;
The relatively incomplete assays for %ASCu and %CaO compared to the %TCu data in the historical data;
Limited density determinations undertaken on the various lithologies;
The risk in the data informing the mineral resource estimate;
The robustness of the geological model;
Historical mining activities and the reconciliation of tonnes and grade; and
The risk in the grade estimates.
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The classification of the mineral resources of the portions not updated in 2011 is considered to be conservative, due to the large bulk of historical data that had to be used during the mineral resource estimation for the 2008 feasibility study. As new data becomes available, using modern drilling and sampling techniques and controls, the confidence in the mineral resource estimate will increase, and therefore the mineral resource classification may be upgraded.
The classification of the mineral resources of the portions updated in 2011 is considered to be a true reflection of the status quo.
14.1 Mineral Resource Statement Table 50 represents the consolidated mineral resources of the various areas of KCC as at December 31, 2011.
Table 50: KCC: Consolidated Mineral Resources as at 31 December 20111,2,3,4 Classification Project Area Mt %TCu %TCo
Measured
KTO 32.1 4.33 0.58 T-17 Open Pit 4.5 2.71 0.54 KOV Open Pit 3.9 4.25 0.22 Subtotal 40.5 4.14 0.54
Indicated
KTO 32.9 4.63 0.57 Mashamba East Open Pit 75.0 1.80 0.38 T-17 Open Pit 9.4 4.44 0.65 KOV Open Pit 117.2 5.41 0.42 Kananga Mine 4.1 1.61 0.79 Tilwezembe Open Pit 9.5 1.89 0.60 Subtotal 248.1 3.98 0.45
Total Measured and Indicated
KTO 65.0 4.48 0.57 Mashamba East Open Pit 75.0 1.80 0.38 T-17 Open Pit 13.9 3.88 0.61 KOV Open Pit 121.1 5.37 0.41 Kananga Mine 4.1 1.61 0.79 Tilwezembe Open Pit 9.5 1.89 0.60 Total 288.6 4.00 0.46
Inferred
KTO 11.0 5.00 0.59 Mashamba East Open Pit 65.3 0.76 0.10 T-17 Open Pit 5.2 4.21 0.98 KOV Open Pit 69.8 3.58 0.32 Kananga Mine 4.0 2.00 0.98 Tilwezembe Open Pit 13.8 1.75 0.60 Total 169.1 2.42 0.31
1) Mineral resources have been reported in accordance with the classification criteria of the JORC Code. If the definitions and classification standards adopted in NI 43-101 had been used instead of the JORC Code, the estimates of mineral resources would be substantially similar.
2) Mineral resources are inclusive of ore reserves.
3) Mineral resources are not ore reserves and do not have demonstrated economic viability.
4) The mineral resource estimates are for KCC's entire interest, whereas the Company owns 75% of KCC.
Numbers may not add due to rounding.
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14.2 Comparison of the 2011 and 2010 Mineral Resources In Table 51 below, the consolidated mineral resources of KCC as at December 31, 2011 are compared with those as at December 31, 2010.
Table 51: Comparison of the 2011 and 2010 Mineral Resources as at December 31, 20111,2,3,4
Classification Project Area 2011 2010
Mt %TCu %TCo Mt %TCu TCo
Measured
KTO 32.1 4.33 0.58 30.7 4.54 0.54 T-17 Open Pit 4.5 2.71 0.54 0.0 0.00 0.00 KOV Open Pit 3.9 4.25 0.22 0.0 0.00 0.00 Subtotal 40.5 4.14 0.54 30.7 4.54 0.54
Indicated
KTO 32.9 4.63 0.57 35.7 4.69 0.60 Mashamba East Open Pit 75.0 1.80 0.38 75.0 1.80 0.38 T-17 Open Pit 9.4 4.44 0.65 8.5 2.75 0.87 KOV Open Pit 117.2 5.41 0.42 123.9 5.37 0.40 Kananga Mine 4.1 1.61 0.79 4.1 1.61 0.79 Tilwezembe Open Pit 9.5 1.89 0.60 9.5 1.89 0.60 Subtotal 248.1 3.98 0.45 256.7 3.95 0.45
Total Measured and Indicated
KTO 65.0 4.48 0.57 66.4 4.62 0.57 Mashamba East Open Pit 75.0 1.80 0.38 75.0 1.80 0.38 T-17 Open Pit 13.9 3.88 0.61 8.5 2.75 0.87 KOV Open Pit 121.1 5.37 0.41 123.9 5.37 0.40 Kananga Mine 4.1 1.61 0.79 4.1 1.61 0.79 Tilwezembe Open Pit 9.5 1.89 0.60 9.5 1.89 0.60 Total 288.6 4.00 0.46 287.4 4.02 0.46
Inferred
KTO 11.0 5.00 0.59 10.6 5.11 0.59 Mashamba East Open Pit 65.3 0.76 0.10 65.3 0.76 0.10 T-17 Open Pit 5.2 4.21 0.98 15.3 1.91 0.61 KOV Open Pit 69.8 3.58 0.32 71.2 3.56 0.32 Kananga Mine 4.0 2.00 0.98 4.0 2.00 0.98 Tilwezembe Open Pit 13.8 1.75 0.60 13.8 1.75 0.60 Total 169.1 2.42 0.31 180.2 2.32 0.32
1) Mineral resources have been reported in accordance with the classification criteria of theJORC Code. If the definitions and classification standards adopted in NI 43-101 had been used instead of those of the JORC Code, the estimates of mineral resources would be substantially similar.
2) Mineral resources are inclusive of ore reserves.
3) Mineral resources are not ore reserves and do not have demonstrated economic viability.
4) The mineral resource estimates are for KCC's entire interest, whereas the Company owns 75% of KCC.
There are no changes in the mineral resources reported for Mashamba East Open Pit, Kananga Mine and Tilwezembe Open Pit.
Overall, the mineral resources for KCC decreased by 2%, while the TCu grade increased by 2% due to the higher grade below current pit in T-17 Open Pit.
The material changes (more than 5% difference) in the mineral resources for KTO, T-17 Open Pit and KOV Open Pit are explained as follows:
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A total increase of 5% in the measured mineral resource of KTO is based on the depletion of the resource due to mining (1.5 million tonnes), an increase in mineral resources of 2.92 million tonnes due to the new resource model for Etang North, which is based on new exploration data and updated density information, which subsequent movement of resources from indicated mineral resources to measured mineral resources (increased confidence in the resource estimate) and a slight increase in the total tonnage;
A decrease of 8% (2.8 million tonnes) in the indicated mineral resources of KTO is based on the movement of resources from indicated to measured mineral resources (increase in confidence in the resource estimate);
The increase in measured mineral resources of 4.5 million tonnes for T-17 Open Pit and 3.9 million tonnes for KOV Open Pit respectively is due to the increased confidence in the resource estimate based on recent exploration results, updated density information and new resource models; and
The RSC component of T-17 Open Pit below planned pit-bottom (8.9 million tonnes) was excluded from the resource statement to conform to the reporting of the mineral resources of KTO, where the RSC is excluded due to mining method limitations in the underground mines. All of these mineral resources are in the inferred category.
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15.0 ORE RESERVE ESTIMATES Set up below are the key Mine planning and summary of technical issues related to the ore reserve estimates.
15.1 Overview of Mining Operations The active operations for KCC can be classified into underground operations, namely Kamoto underground mine (KTO) and surface mining operations, namely T-17 open pit and KOV open pit. The figure below is a graphical representation of the relative location of the current and planned operations at the operations of KCC.
Figure 9: KCC overall mining layout
KTO is an operational mine that produced 0.5 million tonnes Run of Mine (ROM) sulphide ore in 2008, 1.1 million tonnes at 3.90%Cu in 2009 and 1.3 million tonnes at 3.80%Cu in 2010 and 1.6 million tonnes at 3.7%Cu during 2011. The current plan is to ramp yearly production to 2.1 million tonnes sulphide ore, to coincide with the completion of the Phase 4 processing plant expansion.
The KTO Life of Mine (LOM) Plan and Reserve estimate stated in this report, is based on the feasibility study conducted by SRK Consulting in 2008 and depleted on a yearly basis to declare ore reserves.
This ore reserve estimate is qualified up to 2014, by which time a strategic revisit of the entire LOM Plan is required. This is due to the fact that the mining strategy followed in the previous technical studies is not aligned with the operational strategy in terms of overall mining sequence and mining methods.
The current KTO underground mine plan is based on proved and probable ore reserves with only 10% inferred mineral resources included in the LOM Plan. The preliminary economic assessment is in nature, it includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as ore reserves, and there is no certainty that the preliminary economic assessment will be realized.
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A range of mining methods are used and planned in previous technical studies and that include Room and Pillar Mining (RAP), Cut and Fill (CAF) and Sublevel Caving (SLC).
Mining related modifying factors applied are based on actual historic mining method performance. The dilutions and mining over breaks applied vary from 4% to 13% and mining extractions, including geological losses from 3% to 5%, depending on the mining methodology used. The qualified ore reserve estimate (up to 2014) for KTO is 32.4 million tonnes at 3.59%Cu.
Two new underground mines are planned to replace production from the open pit operations. Conceptual work has been completed on Kamoto East underground (KTE) project. This project is planned for ramp up from 2020 onwards and could deliver up to 90ktpa of recovered copper. Development at T-17 underground mine is planned during 2012 while first ore is delivered in 2014. An update on the mineral resource estimate lifted the mineral resource grade by 20%. This has a material positive impact on the feasibility of the project. Appropriate technical work has been conducted up to 2014 and is included in the ore reserve estimate. Appropriate technical work is required on the complete T-17 underground mine to enable an ore reserve estimate on the complete project.
Three surface operations are included in the LOM Plan, namely T-17 extension, KOV open pit and Mashamba East open pit. Tilwezembe open pit and Kananga open pit are not included and require further technical study.
The historical T-17 open pit is depleted and has delivered a total of 2.0 million ROM tonnes at 2.55%Cu in 2010 and 0.4 million tonnes at 3.32%Cu during 2011. A recent update of the mineral resource that increased the Cu% by 20% materially increased the T-17 extension opportunity. The plan is to commence production in 2013 after the water protection and civil works has been completed to enable the T-17 extension. A cut-off of 0.60%Cu has been applied which resulted in 20% mining losses of material below the specified cut-off grade. Mining dilutions of 10% is planned while a 5% geological loss is allowed for. The production scheduled is converted to probable ore reserves.
KOV open pit has been dewatered and is in the production ramp up phase. A total of 0.7million ROM tonnes were produced from KOV open pit in 2010 and 2.5 million tonnes ROM 2011. The operation is planned in two phases, namely cuts 1&2 followed by cuts 3&4. Mining related modifying factors include 1% mining losses below a cut-off grade of 0.60%Cu and mining dilutions of 9% with a 5% applied geological loss. The KOV open pit delivers a ROM head grade of 4.20%Cu for a total of 83.8 million tonnes of ROM ore up to the year 2030. Ore production from the KOV open pit is primarily Oxide material at 78% on average. The ore reserve is estimated at 55.1 million tonnes at 4.74% Cu, classified as probable ore reserve.
Mashamba East open pit is a dormant pit that requires dewatering. A total ROM production of 12.8 million tonnes at 2.72%Cu is planned at a production rate of up to 1.5 million tonnes ROM per year. Mining related modifying factors applied include 9% dilution and 5% geological losses. Due to the high portion of low grade material in this pit, the mining losses below the applied cut-off grade of 0.60%Cu is as high as 39% of the in-pit mineral resource. Mashamba East pit produces only oxide ore and the probable ore reserve is estimated at 5.9 million tonnes at 3.00%Cu.
The cumulative LOM production profile for the KCC operation ramps up to a peak of 10 million tonnes ROM per annum. Assuming that the processing plant is constrained by recovered copper, the production target of 310ktpa recovered metal copper for the Phase 4 expansion is achieved up to the year 2028. Cobalt production is limited to 30ktpa recovered cobalt. The cumulative waste stripping requirements for the three open pit operations are high at a peak of 45 million tonnes per year.
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Figure 10: KCC ROM Production profiles
The recovered copper contribution from each of the operations at KCC can be seen in the figure below. It can be seen that KOV open pit mine would produce the bulk of the recovered copper over the life of the operation based on the current LOM Plan.
Figure 11: KCC recovered Cu profiles 2011
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The total ore reserve estimate at KCC is 96.0 million tonnes at 4.21% Cu of which 13.0 million tonnes is from the proved ore reserve category.
Table 52: KCC Ore Reserve Estimate1,2,3
Mining operation Proved Probable
Mt %TCu %TCo Mt %TCu %TCo
Kamoto underground 13.0 3.43 0.51 19.4 3.70 0.53
T-17 underground 0.0 0.00 0.00 0.9 3.51 0.57
T-17 Open pit 0.0 0.00 0.00 1.6 3.52 0.56
Mashamba East Open pit 0.0 0.00 0.00 5.9 3.00 0.37
KOV Open pit 0.0 0.00 0.00 55.1 4.74 0.45
Total 13.0 3.43 0.51 83.0 4.33 0.46
1. The ore reserve estimates have been prepared in accordance with the classification criteria of the JORC Code. If the definitions and classification standards of NI 43-101 had been used instead of those of the JORC Code, estimates of mineral reserves would be substantially similar to the estimates of ore reserves.
2. The ore reserve estimates are for KCC's entire interest in such ore reserves, whereas the Company owns 75% of KCC.
3. Numbers may not add due to rounding.
The major risks that could have a negative impact on the overall KCC planned production profile are dewatering, access and slope failures, available pit space, and available waste dumping space and grade control. These aspects could be mitigated by good operational management with specific reference to KOV mine due to the high required production rate.
15.2 Development and Operations The primary developments within the Material Assets during 2011 have been the following:
Exploration drilling continued in KTO, T-17 Open Pit and KOV Open Pit;
Re-interpretation of the geology of Etang North orebody in KTO, T-17 orebody and the orebodies in KOV;
KOV Open Pit was dewatered;
Milling capacity at KTC was increased to 7.68 million tonnes of ore per year; and
Luilu Refinery production capacity was increased to 150,000 tonnes of copper per annum due to the plant upgrade programme. Phase 3 was completed during the second quarter of 2011 and the New Phase 4 (inclusive of Phase 5) commenced during the fourth quarter of 2011.
15.3 Underground Previous design and feasibility studies undertaken for Kamoto underground include a feasibility study by Read, Swatman & Voigt (Pty) Ltd (RSV) in 2006, a detailed design by RSV in 2007 and a feasibility study by SRK Consulting (SRK) in 2008. The LOM Plan and Reserve estimate outlined in this report, is based on the feasibility study conducted by SRK in 2008. Extracts from the Kamoto underground feasibility are included below. Additional detail can be found in “Kamoto Operating Limited Feasibility Study” (SRK Consulting) dated November 2008.
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15.4 Kamoto Underground Life of Mine Plan - Previous Technical Study
15.4.1 Background An overall layout of KTO, illustrating the various mining zones is shown in Figure 12 below. This figure is a graphical representation of the design areas as discussed in this section.
Figure 12: General configuration of mining areas at Kamoto underground mine
Details of zone geometry and mining methods that have been considered during various investigations, together with an indication of ore tonnages available, to identify the relative importance of each zone, are presented in Table 53. Descriptions of different mining methods presented in previous reports, are described in detail in Section 7, Appendix D of the SRK feasibility study and are summarized below.
Table 53: Summary of mining methods and previous studies
Zone Geometry RSV 2006 Feasibility Study
RSV 2007 Detailed Design
2008 Feasibility Study
Z1 Bottom and Top OBS and OBI
Flat & steeply dipping portions.
Long Hole Retreat Stoping Cut and Fill Cut and Fill
Z2 OBS and OBI Flat & steeply dipping portions.
Long Hole Retreat Stoping Cut and Fill Cut and Fill
Z3 & Z4 OBS Flat Room & Pillar Room & Pillar Room & Pillar
Z3 & Z4 OBI Flat Long Hole Retreat Stoping
Long Hole Retreat Stoping Room & Pillar
Z5 OBS Flat Room & Pillar Room & Pillar Room & Pillar
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Zone Geometry RSV 2006 Feasibility Study
RSV 2007 Detailed Design
2008 Feasibility Study
Z5 OBI Flat Long Hole Retreat Stoping Room & Pillar Room & Pillar
Z6 OBI Flat & steeply dipping portions.
Long Hole Retreat Stoping Cut and Fill Post pillar cut and fill
& Cut and Fill
Z7 OBI Very steeply dipping N/A N/A Sub level caving
Z8 OBS Flat Room & Pillar Room & Pillar Room & Pillar
Z8 OBI Flat Long Hole Retreat Stoping Room & Pillar Room & Pillar
Z9 OBS Very steeply dipping Cut and Fill Cut and Fill Sub level caving
Z9 OBI Very steeply dipping Cut and Fill Cut and Fill Sub level caving
Etang OBS & OBI Steeply dipping Long Hole Retreat Stoping Cut and Fill Post pillar cut and fill
& Cut and Fill
Etang North OBS & OBI Steeply dipping Long Hole Retreat
Stoping Cut and Fill Post pillar cut and fill & Cut and Fill
Etang Middle & Bottom OBS & OBI Steeply dipping Not specifically
discussed Not specifically discussed
Post pillar cut and fill & Cut and Fill
15.4.2 Strategic Approach The bulk of the LOM Plan from which the current ore reserve estimate has been derived, is based on the previous technical studies. Strategic technical work commenced in 2011 to simplify the mining methodologies and apply historical, high extraction and appropriate methods in which the mining personnel are experienced and efficient. This study could have a material impact on the ore reserve estimate in terms of the sequence and approach to the various production zones, the extractions and dilutions achieved and cement and waste backfill requirements. Appropriate medium term technical work has been conducted up to the end of 2014. Due to the expected material change in mining strategy in Kamoto underground, the current ore reserve estimate is qualified up to the end of 2014. An appropriate and detailed strategic mine plan reflecting the current mining strategy should be developed to replace the previous technical studies.
15.4.3 Design Constraints Applicable to KTO Particular characteristics of the Kamoto ore body that influence mining practice and geotechnical conditions has been extensively discussed in the previous technical studies and includes:
Presence of a collapsed zone in the central “Plateure” portion of the ore body, which may influence stress distributions on adjacent areas;
Mining, and subsequent backfilling and waste dumping in the Kamoto North Open Pit directly overlying the underground workings may influence stress distributions underground;
Underground mining will take place in conjunction with development of the KOV Pit lying to the East, which may influence stress distributions;
With the exception of the Etang mining zones, the ore body generally has been extensively mined and new operations must interact with old workings;
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The dip of the ore body varies from flat to near vertical. For the purposes of mining method classification, a dip range from 0º to 12º is considered to be flat, from 13º to 55º to be steeply dipping and greater than 55º as vertical;
At shallow dip (< 45º), ore that is blasted will not run readily to gathering points. Mining methods within the ore body are required to maximize recovery of ore;
Hanging-wall strata immediately overlying the OBS are reported to be weak and friable and limit the extent of mining spans that can be developed;
Two ore bodies are present which are considered to be wide (between 8m and 16m) and are separated by a parting of variable width (5m to 15m);
Due to the width of the ore body, backfilling is an essential part of the mining strategy to increase overall recovery and protect against uncontrolled collapse of workings; and
There is a limited amount of rock generated from waste development for use as backfill and additional sources of fill are required.
15.4.4 RAP General Layout – Previous Technical Studies The RAP method is described in detail in Section 7, Appendix D.1 of the SRK feasibility study. RAP mining practised at KTO on OBS and OBI horizons, takes place in three phases:
1) Development of drives, 6m wide by 5m high on 25m centres to create square pillars 19m wide and 5m high;
2) Stripping of drives to a full width of 15m to create square pillars 10m wide and 5m high; and
3) Benching of drives to the full height of the ore body to create square pillars 10m wide and up to 15m high.
Note that benching operations take place under a hanging wall that is supported in the first two stages of mining. Following benching, stopes should be filled to at least two thirds of the pillar height with waste rock or hydraulic fill or both, to provide support for pillars. If fill is not used, much larger pillars would be required for a stable layout. As an indication, for a mining height of 15m with a 15m wide roadway, the required pillar width to provide a safety factor of 1.2 would be close to 50m. Overall volumetric extraction would reduce to approximately 40%. A typical RAP mining layout can be seen in Figure 13 below.
Figure 13: Room and Pillar mining layout (Zone 8)
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15.4.5 Long Hole Retreat Stoping General Layout – Previous Technical Studies As planned in the 2006 Feasibility Study, a mining block has dimensions of 150m on strike and 75m on dip. It contains 10 mining panels, each 15m wide and 75m long. The stope is bounded by barrier pillars that serve to provide regional support and protection for access ways and infrastructure contained within them.
Extensive application of Long Hole Retreat Stoping (LHRS) was proposed in the 2006 Feasibility Study. In the 2007 Detailed Design study, LHRS was restricted to OBI mining in Zones 3 and 4. Although applied to smaller areas, the principles governing the method remain valid and are summarized in this section. Should LHRS be applied in future operations, it probably will be on a smaller scale.
There are three phases of mining in each panel with the LHRS method:
1) Initially, a 4m by 5m drill drive is developed along the complete length of the panel to intersect access ways on either side;
2) A slot is created at one end of the panel. The panel is then retreated using conventional long hole drilling and blasting. Ore extraction is achieved using remotely operated loading equipment; and
3) On completion of ore extraction, barricades are constructed in the drill drives at each end of the panel and panels are backfilled using tailings-based fills.
15.4.6 Cut-and-Fill Mining Methods and Variations – Previous Technical Studies Cut-and-fill (CAF) mining methods envisioned for KTO include conventional up-dip retreat together with a post-pillar method for areas in which the hanging-wall span or stope span exceeds 15m. Where the dip ranges from 12º to 55º, preferred mining methods involve taking the ore body as a series of cuts, 5m in height, advancing up dip and placing fill consisting of waste rock, cyclone-classified tailings, cemented classified tailings, or a combination of waste rock and tailings fill, once stoping has been completed.
Because the horizontal width and exposed hanging wall spans of the two ore bodies may be considerable, particularly at shallow dip, two methods are considered:
1) Post Pillar Cut and Fill (PPCF); and
2) Longitudinal Cut and Fill (CAF).
The choice of method in any area will depend on the horizontal width of the ore body. The methods are considered in detail for use in sections of the three planned mining areas at Kamoto Mine:
1) Etang North;
2) Etang South; and
3) Zone 1 Upper.
A typical layout of the CAF mining method is shown in Figure 14 below.
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Figure 14: Cut and Fill longitudinal method
15.4.7 Sub-Level Caving General Layout – Previous Technical Studies To maximize the tonnage that can be recovered without using backfill, an additional mining method has been proposed for Zone 9 – this involves OS with subsequent caving of the overlying CAF mining zone.
General Considerations include:
The practice of sub level stoping is well established. Provided the appropriate mining practice is applied, the method will be applicable for OBS and OBI stopes alike;
It is not expected that there will be rock-mechanic constraints to scheduling; and
A potential benefit of this method lies in its ability to provide backfill to facilitate Zone 8 mining.
15.5 Kamoto Underground Mining Strategic Update The current LOM Plan relies on the technical studies conducted in 2008 on a range of new and current mining methods. This comprehensive list of the planned mining methods for the remainder of the life of the operation based on the previous technical studies is tabled under the “Kamoto underground Life of Mine schedule” section of this report.
An update of the LOM Plan is in progress to reflect the long term mining strategy. The strategy entails a limited number of mining methods with a focus on maximised extraction ratios and historical performance based on the appropriateness of the mining methods. The aim is to maximise efficiencies and productivity, and minimise risk through the experience of the mining teams with a small number of familiar mining methods. The mining methods envisaged includes Room and Pillar mining at dips of less than 12°, Transversal cut and Fill up to 55° and Longitudinal Cut and Fill at dips greater than 55°.
15.5.1 Room and Pillar This method is used in Zone 5 and 2 for dips below 12°. The 12.5m stopes with 12.5m pillars are mined as primary stopes and backfilled with cemented hydraulic fill. Pillars are mined after curing as Secondary stopes. The generic design of the room and pillar mining method (as applied by KCC at Kamoto underground) is shown in Figure 15 below.
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Figure 15: Room and Pillar with cemented backfill
The 6m x 5m access drives are developed on the hanging wall contact at the top of the block along strike. Cross drives are developed (6m x 5m) along the centre of stopes for the full length of thereof and connected by drifts for ventilation purposes. A strike gathering drive is developed at the bottom of the stope for cleaning the ore. Down long holes and slot are drilled from the sloped cut to the footwall of the stope blasted from the gathering drive end. The stopes are filled with cemented hydraulic. Pillars are developed and blasted similar to the primary stopes. These secondary stopes are filled with waste rock or hydraulic fill for regional stability. This is a high extraction mining method with limited lose and dilutions.
15.5.2 Cut and Fill Transverse This method is used in Zones 1 and 6 for dips between 12° and 55°. Stopes are 15m wide and pillars 8m wide along strike. A schematic representation of a general layout can be seen in Figure 16 below.
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Figure 16: Schematic representation of Transversal Cut and Fill
The mining method requires a 6m x 5m drive on the hanging wall contact at 15m vertical intervals. Crosscuts (6m x 5m) are developed to the foot wall contact of the specific ore body. Drifts are supported with 2.4m bolts at 1.2m centres while the 15m wide chambers are supported with additional 4m long anchors at 2m centres. Crosscuts are sloped to 15m wide chambers. A connection drive is developed to connect the chambers at the footwall contact. Slot raises and long hole blast rings are drilled from the top and bottom of the chambers. Rings are blasted sequentially towards the slot from the top down and bottom up. Stopes are backfilled depending on the availability of waste rock from other production area. Hydraulic fill is used if development waste is unavailable and appropriate backfill logistics exist in the area in question.
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15.5.3 Cut and Fill Longitudinal This method is used in Etang North and Esquil for dips from 55° and higher. Vertical level spacing remains unchanged at 15m. Stopes are designed at the width of the ore body, 60m in length and with 8m rib pillars. A schematic representation of the mining method design can be seen in Figure 17 below.
Figure 17: Schematic representation of longitudinal cut and fill
Level access is developed through a ramp system. Strike drives are developed at 6m x 5m and are located at the safest position in the ore body for that level based on geotechnical and geological parameters. These drives are developed to full width of the ore body. Down holes are drilled from the upper slyped drive for the slot and the stope. Slot and rings are blasted and cleaned from the lower level access. Stopes are backfilled with waste rock or hydraulic fill depending on availability. The initial development is 6m x 5m and is supported with 2.4m bolts at 1.2m centres. The 15m sloped development is supported with 4m bolts or anchors at 2m centres.
15.5.4 Future Reserve Updates Due to the strategic changes in the resource exploitation strategy, a material change in the ore reserve estimate for Kamoto underground is envisaged. Due to the relatively high extraction rates and low dilutions planned and achieved on the planned future three mining methods, a reduction in the ore reserve or ore reserve grades are not expected. It is expected that the full LOM Plan redesign and reschedule for all the underground production and project areas be completed in the next 18 months. The ore reserve estimates for Kamoto underground must be updated at this point.
Appropriate detail design and scheduling has been conducted for 2012 through 2014 based on the revised mining methodology. This has been updated in the Ore Reserve estimate and LOM Plan. The LOM Ore Reserve estimate for Kamoto underground as published in the 2012 Ore Reserve statement is therefore only qualified up to 2014. A full and appropriate update on the mineral resource exploitation strategy of Kamoto underground is required before the end of 2014. The current and complete Ore Reserve estimate
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based largely on the 2008 study will therefore not be appropriate or valid in the public domain by 2015 and should by fully replaced by appropriate technical mining engineering work.
15.5.5 Production Scheduling A total of 1.6 million tonnes were planned at 3.97%Cu during 2011 with actual production achieved at 1.6 million tonnes at 3.71%Cu.
Development metres per annum are shown in Table 54 below. The metres are split between capital and working cost development. The decision was that all main accesses (including inclines, declines and connecting crosscuts) and ventilation development (drives and ventilation shafts) were taken as capital development and the remainder (footwall drives, crosscuts and inter-level vent holes) as working-cost development.
Table 54: KTO Life of Mine schedule KTO Unit 2012 2013 2014 2015 2016 2017 2018 2019 2020 2021
Ore tonnes Mt 1.9 1.8 1.8 2.3 2.0 1.9 1.9 2.0 2.0 2.0
Cu grade % 3.88 3.94 3.89 3.67 4.12 4.07 3.95 3.82 3.78 3.89
Co grade % 0.57 0.58 0.57 0.57 0.57 0.57 0.59 0.54 0.52 0.51
Waste development kt 4.0 4.5 4.5 4.4 3.2 3.2 3.0 3.0 2.8 2.8
Recovered Cu kt 67.2 65.0 65.0 75.0 75.0 71.0 69.0 70.0 69.0 70.0
Recovered Co kt 8.4 8.1 8.1 9.9 8.8 8.4 8.7 8.4 8.1 7.8
KTO Unit 2022 2023 2024 2025 2026 2027 2028 2029 2030 Total
Ore tonnes Mt 2.0 2.1 2.1 2.0 2.0 1.8 2.0 1.5 1.1 36.2
Cu grade % 3.99 3.67 3.61 3.71 3.68 4.19 3.80 3.77 3.72 3.85
Co grade % 0.55 0.57 0.55 0.62 0.53 0.49 0.46 0.50 0.49 0.55
Waste development kt 2.9 2.7 2.4 2.3 2.2 2.2 2.0 1.5 0.8 54.0
Recovered Cu kt 73.0 68.0 67.1 67.9 66.9 67.3 68.3 50.0 35.9 1,260.6
Recovered Co kt 8.5 8.9 8.6 9.6 8.2 6.7 7.0 5.6 4.0 151.6
The production profile with ROM head grade and development requirements are shown in Figure 18 and Figure 19 below. The KTO expansion plan considers the development required to ramp up to 2.0 million tonnes per annum. The figure below shows the current LOM Plan to exploit the KTO mine mineral resources for a period of 19 years. The ROM head grade remains relatively constant over the life of the operation of 37 million tonnes ROM at a 3.85% Cu head grade.
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Figure 18: KTO LOM Plan profile
Figure 19: KTO development profile
15.5.6 KTO LOM Plan reconciliation A total of 1.6 million tonnes at 3.71% Cu have been mined during 2011. The complete variance between the 2011 LOM Plan and Ore Reserve estimate and 2012 LOM Plan and Ore Reserve estimate is the actual mining production carried out during 2011. Table 55 below summarises the variance between the 2011 and 2012 LOM Plans for KTO.
Table 55: KTO reconciliation Kamoto Underground 2012 LOMP 2011 LOMP Mined 2011
ROM (Mt) 36.2 37.8 1.6
Reserve (Mt) 32.4 34.0 1.6
Recovered Cu (kt) 1,260.0 1,320.0 50.0
Recovered Co (kt) 150.0 160.0 10.0
Waste development (km) 54.40 59.79 5.40
ROM Cu grade % 3.85 3.84 3.71
ROM Co grade % 0.55 0.54 0.53
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15.6 T-17 Underground 15.6.1 Background T-17 is an operational pit with a current design up to level 1,295m for a maximum depth of 130m. A 30m crown pillar is left between T-17 open pit and T-17 underground. An opportunity exists to develop an underground mine at T-17 to exploit the remaining classified mineral resources. Access from the current pit could be established. Figure 20 below is a graphical representation of the current pit and surface surveys indicating the T-17 open pit extension and the envisaged underground access and ventilation locations.
Figure 20: T-17 current pit, showing open pit extensions and underground access
15.6.2 Mining Strategy Current designs are on a conceptual level of detail. Further work is required to establish the technical, practical and economic viability of the project before the remaining T-17 mineral resources could be converted to ore reserves. Figure 21 below is a graphical representation of the conceptual design to exploit the resources at T-17 below an elevation of 1,300m. The known mineral resources are near vertical and extend approximately 200m below the planned pit design floor.
T-17 Open pit extension
N
T-17 Underground planned shafts
T-17 Underground planned access
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Figure 21: Conceptual design of T-17 underground design
The mining methods to be employed at T-17 Underground are Transverse and Longitudinal CAF mining. The Transverse CAF mining method consists of the main access drive longitudinally developed to the ore body. From the main access drive, stope drives are developed perpendicular to the main access drive. The stope drives are widened to create stopes, leaving a pillar between adjacent stopes.
In the case of the Longitudinal CAF mining method, the stope drives are placed longitudinally in the centre of the ore body and sloped towards both the hanging wall and footwall contacts for the top and bottom levels. Once the top and bottom levels have been sloped, parallel long hole drilling is utilised to extract the portion of reef in between. This mining method has a higher extraction ratio than the Transverse CAF mining method.
15.6.3 Future Ore Reserve Updates Appropriate technical planning is currently under way and scheduled for completion in the next 18 months. Ore reserves are only declared for the section where appropriate technical studies have been conducted (up to 2014). The ore reserve estimate could increase materially should the technical study be completed successfully for T-17 underground.
15.6.4 Production Scheduling An appropriate design and schedule was conducted on T-17 underground on primary development, secondary development, stoping and backfill requirements. The expansion of the appropriate detail over the LOM of the operation is currently in progress and should be completed before the end of the next planning cycle. Development is planned on level access systems from the western pit high wall on level 1270, 1165 and 1150. Development is planned to commence in 2012. A graphical representation of the design and schedule of these three levels are shown in Figure 22 below.
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Figure 22: T-17 underground appropriate design and schedule
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The LOM schedule based on a conceptual mining plan with appropriate detail up to 2015 can be seen in Table 56 below.
Table 56: T-17 Underground LOM production T-17 Underground Unit 2012 2013 2014 2015 2016 2017 2018 2019 2020 2021
Total ROM Ore tonnes Mt 0.0 0.0 0.3 0.6 1.4 1.5 1.7 1.5 1.5 1.5 ROM Stoping tonnes Mt 0.0 0.0 0.0 0.5 0.8 0.9 1.2 1.0 1.0 1.0 ROM Ore dev. meters km 0.0 0.0 2.4 4.3 4.1 4.7 4.2 3.7 3.7 3.7 Waste dev. meters km 1.0 5.6 6.5 4.5 4.0 3.0 1.0 1.0 1.0 1.0 ROM Cu grade % 0.00 0.00 4.49 3.02 3.02 3.65 3.65 3.65 3.65 3.65 ROM Co grade % 0.00 0.00 0.57 0.57 0.57 0.57 0.57 0.57 0.57 0.57 ROM Recovered Cu kt 0.0 0.0 11.6 15.9 35.5 46.1 51.7 46.7 46.7 46.1 ROM Recovered Co kt 0.0 0.0 1.1 2.3 5.1 5.5 6.2 5.6 5.6 5.5
T-17 Underground Unit 2022 2023 2024 2025 2026 2027 2028 2029 2030 Total
Total ROM Ore tonnes Mt 1.6 1.5 1.0 0.4 0.0 0.0 0.0 0.0 0.0 14.4
ROM Stoping tonnes Mt 1 1.5 1.0 0.4 0.0 0.0 0.0 0.0 0.0 10.3
ROM Ore dev. meters km 3.7 3.4 3.4 3.5 0.0 0.0 0.0 0.0 0.0 44.8
Waste dev. meters km 0.5 0.5 0.2 0.1 0.0 0.0 0.0 0.0 0.0 29.8
ROM Cu grade % 3.65 3.89 3.89 3.89 0.00 0.00 0.00 0.00 0.00 3.63
ROM Co grade % 0.57 0.57 0.57 0.57 0.00 0.00 0.00 0.00 0.00 0.57
ROM Recovered Cu kt 48.6 48.7 31.7 14.2 0.0 0.0 0.0 0.0 0.0 443.5
ROM Recovered Co kt 5.8 5.5 3.6 1.6 0.0 0.0 0.0 0.0 0.0 53.6
The ROM production profile and head grade for T-17 underground is shown in Figure 23 below.
Figure 23: T-17 Underground LOM profile
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An average ROM head grade of 3.60% Cu is achieved over the LOM. Based on average recoveries of 85% for copper and 65% for cobalt, the resulting recovered copper profile is shown in Figure 24 below.
Figure 24: T-17 Underground recovered Cu profile
Additional technical work is required to covert the bulk of the T-17 underground mineral resources to ore reserves. The conceptual designs created are reasonable in principle and the mining methodology is similar to existing mining methods at KTO. Mining related modifying factors were derived from the actual mining method efficiencies experienced at KTO.
Figure 25: T-17 underground development profile
15.6.5 T-17 underground LOM Plan reconciliation Additional resource technical work conducted resulted in a material change in the Cu% at the T-17 area. The mineral resource grade (Cu%) has increased by 20% from the 2011 mineral resource estimate to the 2012 mineral resource estimate. The variance in the 2011 and 2012 LOM Plan and the ore reserves is shown in Table 57 below and can be attributed to:
Resource upgrade; and
Additional design and mine planning on development.
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Table 57: T-17 underground reconciliation T-17 Underground 2012 LOMP 2011 LOMP Variance
ROM productions (Mt) 14.4 13.7 0.7
Ore Reserve (Mt) 0.9 0.0 0.9
Recovered Cu (kt) 443.5 300.0 143.5 Recovered Co (kt) 53.6 60 0.0
Ore development (km) 44.8 25.5 19.3
Waste development (km) 29.8 23.8 6.0
ROM Cu grade (%) 3.63 2.57 1.06
ROM Co grade (%) 0.57 0.69 0.12
15.7 Kamoto East underground 15.7.1 Background The mining method currently envisaged and allowed for at KTE is SLC. This mining method is suited for the extraction of the steep dipping portion of the ore body being targeted by KTE. The key advantages of the SLC is a rapid production ramp up due to the large number of loading points being created, a low intensity support requirement and a high percentage extraction of the ore body. A 30m crown pillar is left between KOV open pit and KTE.
The access development to KTE is illustrated in Figure 26 below.
Figure 26: KTE conceptual development layout
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The planned underground operations could deliver a ROM head grade of 4.44 %Cu for a total of 20.5 million tonnes of ROM ore from 2019 to 2030. For the time frame allowed for in the LOM Plan, only 15.7 million tonnes at 3.75 %Cu was included into the LOM Plan up to 2030. The peak planned production rate from underground operations is 2.4 million tonnes per annum of ore. The LOM production results are shown in Table 58 below.
Table 58: KTE LOM production Kamoto East Underground Unit 2012 2013 2014 2015 2016 2017 2018 2019 2020 2021
Total ROM tonnes Mt 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.6 1.2
Stoping tonnes Mt 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.2 0.5
Recovered Cu kt 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 21.7 45.3
Recovered Co kt 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 1.0 2.0
Ore development km 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.6 1.8
Waste development km 0.0 0.0 0.0 0.0 0.0 0.6 3.0 4.0 4.5 4.5
Cu grade % 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 4.57 4.57
Co grade % 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.28 0.27
Kamoto East Underground Unit 2022 2023 2024 2025 2026 2027 2028 2029 2030 Total
Total ROM tonnes Mt 1.0 1.2 1.6 1.8 1.1 1.3 1.1 1.2 2.6 14.7
Stoping tonnes Mt 0.8 1.1 1.4 1.5 0.9 1.1 0.9 1.0 2.2 11.6
Recovered Cu kt 39.7 44.6 63 69.7 42 47.7 38.9 44.8 96.6 553.8
Recovered Co kt 1.8 2.0 2.6 2.8 1.7 2.0 1.7 1.8 3.6 22.9
Ore development km 2.3 2.9 3.8 4.8 4.8 4.2 4.4 3.6 3.6 36.8
Waste development km 4.5 4.5 4.2 4.0 3.0 2.6 1.5 1.0 0.5 42.4
Cu grade % 4.57 4.34 4.53 4.53 4.53 4.32 4.32 4.36 4.36 4.44
Co grade % 0.27 0.25 0.24 0.24 0.24 0.24 0.24 0.23 0.21 0.24
The ROM production profile and head grade profile is indicated in Figure 27 below.
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Figure 27: Kamoto East Underground LOM profile
An average ROM head grade of 4.40% Cu is achieved over the LOM. Based on average recoveries of 85% for copper and 65% for cobalt, the resulting recovered copper profile is shown Figure 28 below.
Figure 28: KTE development requirements
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Figure 29: KTE recovered Cu profile
15.7.2 KTE LOM Plan Reconciliation Scheduling changes exist between the LOM Plan of 2011 versus the 2012 LOM Plan. No ore reserve was declared in 2011 and no ore reserve is declared in 2012. Additional technical work is required. A variance of approximately 6 million tonnes exists due to the fact that the LOM Plan is only consider up to 2030. No technical reason exists why these tonnes cannot be included in future.
Table 59: KTE Reconciliation
Description Waste (Mt) SR (t/t) Ore (Mt) Cu% Co% Recovered
Cu (kt) Recovered
Co (kt)
2011 LOMP 461.8 5.55 83.1 4.14 0.40 2,928.0 216.5 Optimisation and design adjustments 9.8 1.36 6.7 3.04 0.24 205.0 15.9
Mined/ booked 2011 -18.8 3.78 -2.5 4.81 0.38 -103.0 -6.2
Ore Not booked in 2011 -18.8 3.78 -2.4 2.59 0.20 -52.0 -3.1
2011_east_adj -6.8 5.29 -1.2 2.79 0.50 -28.0 -3.9
2012_lomp 471.0 5.62 83.8 4.14 0.40 2,950.0 219.2
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15.8 Modifying Factors 15.8.1 KTO Underground The applied modifying factors for the various mining methods are shown in Table 60 below.
Table 60: KTO Mine modifying factors Mining Method Description Metres %
Cut and Fill
Planned foot wall over break - OBI (12m wide stope) 1.0 8 Planned foot wall and hanging wall over break – OBI & OBS (12m wide stope) 0.5 4
Ore loss in stopes 0.0 3 Geological loss 0.0 5
Room-and-pillar
Development over break and Slyping over break stopes 0.1 2
Benching over break benching 0.0 0 Ore loss in stopes 0.0 3 Geological loss 0.0 5
LHRS
Side wall over break (initial) – 15m wide stopes 0.0 0 Side wall over break (secondary) – 15m wide stopes 1.5 5 Hanging wall over break (12m high stope) 0.5 4 Ore loss in stopes 0.0 3 Geological loss 0.0 5
Zone 7
Geological loss 0.0 5 Extraction after geological loss 0.0 60 Planned dilution (Foot wall and hanging wall over break - OBI & OBS respectively) 1.5 12
Zone 6 and Zone 3
Geological loss 0.0 5 Extraction after geological loss 0.0 50 Planned dilution (Foot wall and hanging wall over break - OBI & OBS respectively) 1.5 12
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Table 61: KTO Mine: Modifying factors for the PPCF Method Zone Ore Horizon Description Metres %
Etang North
OBI Mining dilution 0.0 11
Over break (hanging wall and footwall) 0.3 4
OBS
Mining dilution 0.0 17
Over break (hanging wall and footwall) 0.3 8
Ore loss in stopes 0.0 3
Geological loss 0.0 5
Etang South
OBI Mining dilution 0.0 13
Over break (hanging wall and footwall) 0.3 5
OBS
Mining dilution 0.0 13
Over break (hanging wall and footwall) 0.3 6
Ore loss in stopes 0.0 3
Geological loss 0.0 5
Zone 1 Top
OBI Mining dilution 0.0 13
Over break (hanging wall and footwall) 0.3 5
OBS
Mining dilution 0.0 10
Over break (hanging wall and footwall) 0.3 5
Ore loss in stopes 0.0 3
Geological loss 0.0 5
15.8.2 T-17 Underground Modifying factors applied to the CAF mining areas are shown in Table 62 below. Factors were applied globally level by level. A level spacing of 15m was assumed for the development layout.
Table 62: T-17 Underground modifying factors
Mining method Geological losses Mining losses Dilution
Longitudinal CAF 5% 5% 10%
Transverse CAF 5% 35% 5%
These factors were bench marked against factors obtained from areas where CAF mining was applied. 10% of the planned production is produced by applying Transverse CAF mining. The remaining 90% is subsequently produced by Longitudinal CAF mining. A 30m crown pillar is left between T-17 open pit and the underground workings. The planned underground operations deliver a ROM head grade of 3.63% Cu for a total of 14.4 million tonnes of Run of Mine ore from 2014 to 2030.
15.8.3 KTE Underground Modifying factors applied to the underground mining area are shown in Table 63 below. Factors were applied globally level by level as opposed to stope by stope. A level spacing of 15m was assumed for the development layout.
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Table 63: KTE modifying factors
Mining method Geological losses Mining losses Dilution
SLC 5% 15% 20%
These factors were bench marked against historical information obtained from areas where SLC was applied.
15.9 Ore Reserve Estimate The updated Ore Reserve Estimate is based on the mining methods described in the section above and is compliant with the JORC Code. The entire resource area was sub-divided into smaller portions per zone to aid the mine design and schedule process. The basis of conversion from mineral resource to ore reserve will be that measured mineral resources will convert to proved ore reserves and indicated mineral resources to probable ore reserves.
15.9.1 Ore Reserve Estimate – KCC Underground Operations The reserve estimate is partially based on the 2008 SRK feasibility study and updated based on the actual production performance achieved in 2008, 2009 and 2010. The Ore Reserve statement is shown in Table 64 below.
Table 64: KTO Ore Reserves as at 31 December 20111,2,3
Mining operation
Proved Probable Total Mt %TCu %TCo Mt %TCu %TCo Mt %TCu %TCo
Kamoto underground 13.0 3.43 0.51 19.4 3.70 0.53 32.4 3.59 0.52
T-17 underground 0.0 0.00 0.00 0.9 3.51 0.57 0.9 3.51 0.57
1) The ore reserve estimates have been prepared in accordance with the classification criteria of the JORC Code. If the definitions and classification standards of NI 43-101 had been used instead of those of the JORC Code, estimates of ore reserves would be substantially similar.
2) The ore reserve estimates are for KCC's entire interest in such reserves, whereas the Company owns 75% of KCC.
3) Numbers may not add due to rounding.
15.10 Ore Reserve Estimate 15.10.1 Ore Reserve Estimate as at December 31, 2011 In the medium term up to 2015 (4 years), a total of 12% ROM ore is currently planned in the LOM Plan from material that cannot be classified as ore reserves. More than 50% of the total ROM tonnes developed at KCC are planned from the KOV ore reserves over the medium term. This can be seen in the figures below. Over the long term, the KOV ore reserves account for 37% of the total ROM tonnes whereas 36% of the total LOM Plan ROM tonnages originate from material that can currently not be classified as an ore reserve. The economic assessment is preliminary in nature, it includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as ore reserves, and there is no certainty that the preliminary economic assessment will be realized.
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Figure 30: Reserve contributions over time
The ore reserve estimate for the KCC operation is shown in Table 65 below.
Table 65: KCC Ore Reserve table1,2,3
Mining operation
Proved Probable Total
Mt %TCu %TCo Mt %TCu %TCo Mt %TCu %TCo
Kamoto underground 13.0 3.43 0.51 19.4 3.7 0.53 32.4 3.59 0.52
T-17 underground 0.0 0.00 0.00 0.9 3.51 0.57 0.9 3.51 0.57
T-17 Open pit 0.0 0.00 0.00 1.6 3.52 0.56 1.6 3.52 0.56 Mashamba East Open pit 0.0 0.00 0.00 5.9 3.00 0.37 5.9 3.00 0.36
KOV Open pit 0.0 0.00 0.00 55.1 4.74 0.45 55.1 4.74 0.45
Total 13.0 3.43 0.51 83.0 4.33 0.46 96.0 4.21 0.47
1) The ore reserve estimates have been prepared in accordance with the classification criteria of the JORC Code. If the definitions and classification standards of NI 43-101 had been used instead of those of the JORC Code, estimates of mineral reserves would be substantially similar to the estimates of ore reserves.
2) The ore reserve estimates are for KCC's entire interest in such ore reserves, whereas the Company owns 75% of KCC.
3) Numbers may not add due to rounding.
15.11 Reconciliation with December 2010 Ore Reserve Estimate The outcome of the December 2011 ore reserve estimate is a decrease of only 1 million tonnes of ore reserve despite the fact that 4.5 million tonnes have been mined in 2011 due to the addition to the ore reserves of:
Ore reserves at KOV due to additional design work conducted,
Inclusion of the ore reserve up to 2014 at T-17 underground due to additional technical and design studies conducted during 2011, and
Inclusion of T-17 Extension due to additional study conducted.
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Table 66: Ore Reserve reconciliation1,2,3
Mining operation 2011 Reserve
Estimate 2010 Reserve
Estimate Notes Mt %TCu Mt %TCu
Kamoto underground 32.4 3.59 34.0 3.60 Mined out in 2011 and qualified to 2014
T-17 underground 0.9 3.51 0.0 0.00 Appropriates study to 2014 with increased resource Cu%
T-17 Open Pit 1.6 3.52 1.5 2.61 Approval and appropriate study of T-17 Extension
Mashamba East Open pit 5.9 3.00 5.9 3.00 Unchanged
KOV Open Pit 55.1 4.74 55.7 4.73 Mined out in 2012 and design adjustments
Total 96.0 4.21 97.0 4.20 Nett reduction in Reserves due to 2011 mining and other
1) The ore reserve estimates have been prepared in accordance with the classification criteria of the JORC Code. If the definitions and classification standards of NI 43-101 had been used instead of those of the JORC Code, estimates of mineral reserves would be substantially similar to the estimates of ore reserves.
2) The ore reserve estimates are for KCC's entire interest in such ore reserves, whereas the Company owns 75% of KCC
3) Numbers may not add due to rounding.
The ore reserves at KTO are qualified up to 2014. Appropriate technical design and scheduling study is required in the next 18 months to conduct an ore reserve estimate from 2014 due to the material and strategic mine planning changes envisaged. The impact of these changes on the LOM Plan, mining infrastructure requirements and mining operational costs requires appropriate technical study.
16.0 MINING METHODS Ore reserve estimations as part of the MER are based on LOM plans created from first principles. The first section below outlines the process followed.
16.1 LOM Planning Process 16.1.1 Overview The Life of Mine planning process for surface mining operations can be summarised in the following steps:
Selective Mining Unit (SMU) selection;
Pit optimisation;
Pit design;
Scheduling unit selection and design; and
Production planning.
A brief description of each step is given in this section. The pits considered for the LOM Plan and ore reserve conversion process are T-17 open pit, KOV open pit and Mashamba East open pit of which T-17 and KOV are active.
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16.1.2 Mining Model The smallest mining unit (SMU) is defined as the smallest unit that that can be mined as a complete unit. SMU’s applied to the various resource block models are shown in Table 67 below.
Table 67: Selected SMU dimensions per pit
Mining Operation Unit SMU
T-17 m 5 x 5 x 5
Mashamba East m 10 x 10 x 5
KOV m 10 x 10 x 5
16.1.2.1 Dilution Dilution is defined as the waste material intentionally added during mining block modelling to what has been site-specifically defined as in-situ mineral resources, in order to make it practically mineable. The methodology applied in determining the dilution is as follows:
On the ore contacts (where the in-situ resource block consists of a percentage ore material and a percentage waste material) the tonnage and grade of the reserve block is defined as the weighted average tonnage and grade of the materials contained in the original resource block;
In cases where the total in-situ resource block is ore, the corresponding reserve block is defined as a 100% ROM block with the same grade attributes as the in-situ blocks.
16.1.2.2 Mining Loss Mining loss is defined as those reported mineral resources which are contained in planned blocks that are not defined as ore reserve type blocks, in other words if scheduled, these blocks are destined for waste dumps. The methodology in determining mining loss is as follows:
Mining loss is addressed through the application of a copper cut-off to the diluted ore material.
The ore blocks that originally had a high percentage of in-situ ore will normally fall above the cut-off grade while ore blocks that originally had a low percentage of in-situ ore will fall below the cut-off.
16.1.3 Pit Optimisation One of the outputs of the pit optimisation process is to determine the position and extent of the final pit boundary. The GEMCOM Whittle pit optimisation software is employed for this purpose. For brevity the software will be referred to as “Whittle”.
Whittle uses the Lerchs-Grossmann algorithm to determine the optimal shape for an open pit in three dimensions. The method is applied to a block model of the ore body, and progressively constructs lists of related blocks that should, or should not, be mined. The final lists define a pit outline that has the highest total relative value, subject to the required pit slopes. This outline includes every block that "adds value" when waste stripping is taken into account and excludes every block that "destroys value". It takes into account all revenues and costs as well as mining and processing parameters.
Although a detail description of the Whittle methodology is beyond the scope of this report, the following provides a brief summary. The optimisation process can be divided into two processes:
1) Creation of a range of nested pit shells of increasing sizes. This is done by varying the product price and generating a pit shell at each price point;
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2) Selection of the optimal pit shell. This is achieved by generating various production schedules for each pit shell and calculating the net present value for each schedule. The output of this process is a series of “pit-versus-value” curves.
16.1.4 Pit Design The mining method applied is conventional open pit mining, consisting of drilling, blasting, loading and hauling. As part of the mine planning process, a pit design is undertaken once an optimal pit shell has been selected. The pit design process considers:
Safe operations;
Continuous access to individual blocks and the working benches;
Equipment units and movement requirements;
Geotechnical recommendations;
Water handling;
Backfill opportunities; and
The phasing of operations or pre-stripping.
Design work was performed in GEMCOM Surpac mine planning software. The selected optimum pit shell is used as the design limits. All the input parameters are incorporated to create a three dimensional pit design. The pit design is used to evaluate the tonnage and grades of the different ore types. Pit designs were created based on the current mining methodology that includes mining at 5m or 10m benches. Ramp and pit access designs considered the largest expected hauler dimension specifications, ensuring safe and practical execution.
The current T-17 open pit is virtually depleted. The T-17 open pit extension that extends the pit laterally has been approved. Some civil and engineering work is required to enable the extension. This is planned for 2012 with production commencing in 2013. All pit designs adhere to current geotechnical requirements.
16.1.5 Scheduling Units Block designs are conducted based on typical blast block or practical bench and production block dimensions. Ramps are designed and scheduled separately at appropriate rates. The block designs simulate the scheduling units. Each block could contain a range of material types that could be selectively loaded to separate locations (ROM stockpile, various stockpiles or waste dumps).
16.1.6 Production Scheduling Schedules were produced in RUNGE Xpac. Operating slope angles for each pit are maintained and increased to the final pit slopes when the ultimate pit is reached. Schedules consider the available pit space, number and size of excavators required and the practical constraints of each pit.
16.2 T-17 Open Pit The original and optimised T-17 pit has virtually been depleted during 2011. This pit delivered a total of 2.0 million ROM tonnes at 2.53%Cu during 2010 and 0.4 million ROM tonnes at 3.32%Cu during 2011. A cut-off grade of 0.60%Cu is applied and a mining dilution of 10% is achieved with mining losses estimated at 20%. Geological losses of 5% are allowed for.
The T-17 open pit extension that extends the pit laterally has been approved. Civil and engineering work is required to enable the extension. This entails the diversion of an access road and diversion and protection of the water way (through culverts) to the east of T-17. The total approved cost is US$2.4 million and is planned for completion in 2012. The road diversion and bridge construction with a three stage river diversion relative to the updated optimised pit outlines can be seen in the Figure 31 below.
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Figure 31: T-17 Extension prerequisite construction work
Additional dewatering of the eastern high wall of the pit is required to protect the crusher installation and pit against failure. Through analyses KCC estimated that the pit walls will not fail to the crusher. A procedure has been drafted and approved to cover both the authorisation by KCC to allow mining contractor, Enterprise Generale Malta Forrest (EGMF) to proceed with blasting operations and the blast security and blast firing procedure to be used by EGMF.
Four diamond drill holes were drilled between the pit slope and the crusher to assess the ground conditions. Additional Precaution for long term stability requires this portion on the pit will be backfilled with waste rock to ensure long term stability of the crusher. The Figure 32 below illustrates the relative locations of the crusher, roads and river culverts and the T-17 pit extension.
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Figure 32: T-17 extension with crusher protection and road and river diversions
All prerequisite work to enable the mining of T-17 extension is planned for 2012 with production commencing in 2013.
16.2.1 Modifying Factors A total of 5% geological losses have been applied. This implies that 5% of the material modelled as ore reserves are mined as waste due to structural resource losses. This is a tonnage loss that does not impact the ROM head grade. At a cut-off grade of 0.60%Cu, losses of material below the SMU cut-off grade are expected at 20% while mining dilutions allowed for are 10% on average as seen from the figure below.
16.2.2 Pit Optimisation KCC conducted a pit optimisation for the T-17 extension based on historical and known stripping ratios of the T-17 pit.
16.2.3 Pit Design The figure below is a view of T-17 in an eastern direction from the lookout point. The pit extension towards the crusher installation, access road and river is indicated in Figure 33 below.
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Figure 33: T-17 open pit
The current pit design used as basis for the LOMP and ore reserve estimate is shown in Figure 34 below. The final pit depth is 130m to an elevation of 1,300m. A graphical representation of the final pit design with current topography is shown in Figure 35 below.
Figure 34: T-17 current pit survey showing section line A-A
A
A N
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Figure 35: Section line A-A of the T-17 pit extension relative to the 2011 T-17 pit survey
The pit design criteria are based on current practice and can be seen in Table 68 below.
Table 68: T-17 pit design criteria Pit Design Criteria Unit T-17
Bench height m 5.0
Berm width m 4.3
Batter angle degrees 78.0
Ramp width m 16.0
Ramp gradient degrees 5.0 (1 in 12)
16.2.4 Production Scheduling The production plan is to achieve a total of 1.6 million tonnes of ROM oxide ore from 2012 to 2015, peaking at 0.6 million tonnes per year during 2014 and 2015. The LOMP profile is shown in Table 69 below.
Table 69: T-17 extension LOM production profile Year Unit 2012 2013 2014 2015 2016 2017 2018 2019 2020 Total
Ore Mt 0.0 0.4 0.6 0.6 0.0 0.0 0.0 0.0 0.0 1.6
Recovered Cu Kt 0.0 9.2 20.9 20.9 0.0 0.0 0.0 0.0 0.0 51.0
Recovered Co Kt 0.0 1.0 2.6 2.6 0.0 0.0 0.0 0.0 0.0 6.3
Waste Kt 0.0 1.7 1.6 0.9 0.0 0.0 0.0 0.0 0.0 4.2
Cu grade % 0.00 2.93 3.91 3.91 0.00 0.00 0.00 0.00 0.00 3.69
Co grade % 0.00 0.44 0.64 0.64 0.00 0.00 0.00 0.00 0.00 0.59
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The LOM profiles profile is shown in Figure 36 below.
Figure 36: T-17 extension tonnage and copper grade profile
Ore from the T-17 Extension pit is exclusively oxide ore as shown in Figure 37 and Figure 38 below.
Figure 37: T-17 ore type profile excluding T-17 extension
Figure 38: T-17 reserve profile excluding T-17 extension
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16.2.5 T-17 Extension LOM Plan Reconciliation A material change in the resource grade of the model resulted in higher ROM Cu% grades. Specialist engineering work is planned for 2012 with operations commencing in 2013.
Table 70: T-17 extension reconciliation Ore type classification 2012 LOMP 2011 LOMP Variance
ROM Oxide(Mt) 1.6 1.5 0.2 ROM Oxide Ore (Mt) 1.6 1.5 0.2 Recovered Cu tonnes (kt) 51.0 33.22 17.8 Recovered Co tonnes (Mt) 6.3 4.3 1.9 Waste tonnes (million) 4.2 2.6 1.6 ROM Cu grade % 3.69 2.61 1.08 ROM Co grade % 0.59 0.46 0.13
16.3 KOV Open Pit 16.3.1 Background A total of 0.7 million ROM tonnes were produced from KOV pit in 2010 and 2.5 million tonnes during 2011. The operation is planned in 2 phases, namely cuts 1 & 2 followed by cuts 3 & 4. Mining related modifying factors include 1% mining losses below a cut-off grade of 0.60%Cu and mining dilutions of 9% with a 5% applied geological loss. The KOV pit delivers a ROM head grade of 4.22%Cu for a total of 83.0 million tonnes of ROM ore up to the year 2030. Ore production from the KOV pit is primarily Oxide material at 78% on average.
16.3.2 Modifying Factors A total of 5% geological losses have been applied. This implies that 5% of the material modelled as ore reserves are mined as waste due to structural resource losses. This is a tonnage loss that does not impact the ROM head grade.
A cut-off grade of 0.60%Cu was applied at the KOV pit. The basis of the cut-off grade calculation is to determine the break even cost based on selling, processing and royalty cost. The cut-off grade considers revenues generated from copper and cobalt with the appropriate processing recoveries applied. The costs, revenues and recoveries allowed for are tabulated in the table below. A total of 1% mineral resource losses of material below (Figure 39) the SMU cut-off grade are estimated while mining dilutions are 9% on average.
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Figure 39: KOV dilution curve for a 10x10x5 SMU
16.3.3 Pit Optimisation The Whittle software package was used to conduct an open pit optimisation. A final overall pit slope angle of 26° was assumed for the northern slopes and 35° for the remaining slopes. The pit optimisation parameters assumed for KOV are shown in Table 71 below.
Table 71: KOV optimisation parameters
Optimisation Parameters Unit KOV Parameter
Reference mining cost US$/t 2.40
Processing cost US$/t 31.81
Selling cost – Cu US$/t Cu 750
Selling cost – Co US$/t Co 750
Discount rate % 12
Cu recovery % 85
Co recovery % 65
Cu Price US$/t 4 790
Co Price US$/t 25 228
16.3.4 Pit Design The current pit is active and dewatered up to pit bottom towards the west where mining activities are focussed. A large slope failure on the northern slope has been re-established with a smaller slope failure that was visible during the site visit early in of 2012. These slope failures are due to a known geotechnical risk zone. Indications are that this zone does not extend towards the pit limits which imply that less geotechnical constraints could apply once the geotechnical zone was mined through. The overall operational face angle allowed for in the production plan should allow access on multiple levels and minimise the risk of large slope failures. The final slope angles used for the pit design process assume lower final slope angles to the north of the pit. The Figure 40 below is viewed in a northern direction from the lookout point with the floor cleaning and small slope failure indicated.
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Figure 40: KOV open pit
Cuts 3&4 are based on the optimised pit shell for the LOM Plan. A graphical representation of the final pit design with current topography is shown in Figure 41 and Figure 42 below.
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Figure 41: KOV pit design relative to the end of 2011 pit survey
Figure 42: Section B-B through KOV indicating the ultimate pit versus the current pit
The pit design criteria are based on current practice and can be seen in Table 72 below.
Table 72: KOV pit design criteria Pit Design Criteria Unit KOV
Bench height (< depth:335m) m 10.0
Bench height (> depth:335m) m 5.0
Berm width (10m benches) m 13.5
Berm width (5m benches) m 6.0
Batter angle (10m benches) degrees 75.0
Batter angle (5m benches) degrees 65.0
Ramp width m 35.0
Ramp gradient degrees 5.7 (1 in 10)
N B
B
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16.3.5 Production Scheduling Production from the KOV pit is based on the available pit space and to maintain the production requirement. An overall production profile for the total KCC operation of 310 000 tonnes of recovered copper per annum is targeted. The LOMP profile is tabulated (Table 73) below.
Table 73: KOV LOM production profile KOV Unit 2012 2013 2014 2015 2016 2017 2018 2019 2020 2021
Ore Mt 3.6 3.7 5.2 5.6 5.1 5.0 4.4 4.2 3.7 3.0 Recovered Cu Kt 151.1 148.9 173.8 180.4 201.8 184.7 160.5 158.4 145.9 124.7 Recovered Co Kt 9.5 8.5 7.3 7.2 10.3 13.3 10.2 10.8 10.6 8.6 Waste Mt 36.4 41.7 39.8 34.4 37.1 37.0 33.4 33.9 34.2 33.0 Cu grade % 4.97 4.76 3.96 3.82 4.68 4.33 4.29 4.45 4.62 4.97 Co grade % 0.41 0.35 0.22 0.20 0.31 0.41 0.36 0.40 0.44 0.41
KOV Unit 2022 2023 2024 2025 2026 2027 2028 2029 2030 Total
Ore Mt 3.3 3.9 4.0 4.4 5.5 5.4 5.4 4.4 4.1 83.8 Recovered Cu Kt 122.7 122.9 121.2 128.3 175.3 166.3 196.5 206.5 124.5 2,994.4 Recovered Co Kt 13.7 13.9 12.5 15.3 17.5 16.0 16.6 12.9 8.2 222.9 Waste Mt 31.3 25.6 22.3 14.1 10.2 4.4 1.4 0.5 0.5 471.0 Cu grade % 4.33 3.70 3.59 3.40 3.72 3.62 4.32 5.57 3.54 4.20 Co grade % 0.63 0.55 0.48 0.53 0.48 0.45 0.48 0.45 0.31 0.41
Figure 43: KOV LOM profile
It can be seen from Figure 44 below that the bulk of the ex pit tonnes generated throughout the LOMP from the KOV pit can be classified as oxide ore.
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Figure 44: KOV ore type profiles
The Figure 45 below indicates that a material portion of the ROM tonnes generated throughout the LOM Plan is from the inferred mineral resource category and cannot be classified as ore reserves as of the date of this report. The economic assessment is preliminary in nature, it includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as ore reserves, and there is no certainty that the preliminary economic assessment will be realized.
Figure 45: KOV Reserve profile
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16.3.6 KOV LOMP Reconciliation The Figure 46 below is an isometric graphical block model representation of the ultimate pit (in red) with the area mined during 2011 indicated in blue.
Figure 46: KOV ultimate pit with mined out area and pit block designs
Figure 47: Isometric view showing mined out and adjusted pit designs
The variance in the LOM Plan tonnages and ore reserve estimates can be seen in Table 74 below. The variance is defined as a combination of factors including:
Mined out quantities during 2011 as delivered to the plant;
East Design Adjustment
LOM Plan 2012 to 2030
Block designs
LEGEND
East Design Adjustment
Mined 2011
LOM Plan 2012 to 2030
Mined 2011
LEGEND
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Low grade mined out quantities above the cut-off grade (0.60%Cu) that has been modelled geologically not mined as ore;
Design changes on the eastern perimeter of the concession to prevent waste stripping outside of the concession area. No current agreement exists to strip waste outside of the concession area or relocate infrastructure outside of the KCC concession area;
Various pit and block design updates and additional detail including an update on the mud horizon previously not in the survey model; and
Practical design changes.
A total of 4.2 million tonnes of waste is added to the model due to practical design changes for access and infrastructure protection, and due to the additional mud deposited on the pit floor after the dewatering phase. A total of 22.3 million tonnes of waste and 2.5 million tonnes of ore have been booked on KOV pit during 2011. Selective high grade mining has taken place that implies that 1.8 million tonnes of the 2.5 million tonnes of ore has been mined as waste.
Table 74: KOV LOM Plan reconciliation Description Ore (Mt) Cu% Co%
2011 LOMP 83.15 4.14 0.40
Optimisation and design adjustments 6.73 3.67 0.30
Mined/ booked 2011 -2.52 4.81 0.38
Ore Not booked in 2011 -2.37 2.59 0.20
2011_east_adj -1.19 2.79 0.50
2012_lomp 83.80 4.20 0.41
1) No material mineral resource changes have been applied to the KOV pit.
16.4 Mashamba East 16.4.1 Background Mashamba East is a dormant pit that requires dewatering. This pit is included in the LOM to replace the lower recovered copper production from the KOV pit from 2018 onwards. A total ROM production of 12.8 million tonnes at 2.80%Cu is planned at a production rate of 1.8 million tonnes per year. Mining related modifying factors applied include 9% dilution and 5% geological losses. Due to the high portion of low grade material in this pit, the mining losses below the applied cut-off grade of 0.6 %Cu is as high as 39% of the Resource. Mashamba East pit produce only oxide ore.
16.4.2 Modifying factors A total of 5% geological losses have been applied. This implies that 5% of the material modelled as ore reserves are mined as waste due to structural mineral resource losses. This is a tonnage loss that does not impact the ROM head grade.
A cut-off grade of 0.60%Cu was applied at the Mashamba East pit. The basis of the cut-off grade calculation is to determine the break even cost based on selling, processing and royalty cost. The cut-off grade considers revenues generated from copper and cobalt with the appropriate processing recoveries applied. The costs, revenues and recoveries allowed for are discussed later in this section. A total of 39% resource losses of material below the SMU cut-off grade are estimated due to the high sensitivity of material at low Cu grades. An average of 9% additional tonnes are applied as dilution as seen from Figure 48 below.
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Figure 48: Mashamba East losses and dilutions
16.4.3 Pit Optimisation The pit optimisation parameters assumed for Mashamba East are tabled below. An overall final pit slope angle of 35° was applied.
Table 75: Mashamba East pit optimisation parameters
Optimisation Parameters Unit Mashamba East
Reference mining cost US$/t 3.40
Processing cost US$/t 48.44
Selling cost – Cu US$/t Cu 750
Selling cost – Co US$/t Co 750
Discount rate % 12
Cu recovery % 85
Co recovery % 65
Cu Price US$/t 4 790
Co Price US$/t 25 228
The processing cost is applied in the optimisation software, but not limited to, as the costs associated with the mining, handling and processing of specifically ore where the reference mining costs applies to ore and waste. The higher relative processing cost (relative to the KOV US$31) can be attributed to the lower volumes mined and the additional cost of hauling ore to the plant, due to the pit location.
16.4.4 Pit design The current pit survey of this dormant operation and optimised pit shell was used as basis of the pit design as shown below. A graphical representation of the final pit design with current topography is shown in Figure 49 below.
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Figure 49: Mashamba East pit design
Mashamba East pit design parameters are shown in Table 76 below.
Table 76: Mashamba East Pit design criteria Pit Design Criteria Unit Mashamba East
Bench height m 10.0
Berm width m 8.0
Batter angle degrees 65.0
Ramp width m 25.0
Ramp gradient degrees 5.2 (1 in 11)
16.4.5 Production scheduling Production from the Mashamba East open pit is planned at a maximum of 1.5mtpa. The mine planning criteria is tabulated below. The LOMP profile is shown in Table 77 below.
Table 77: Mashamba East LOM production profile
Mashamba East Unit 2012 2013 2014 2015 2016 2017 2018 2019 2020 2021
Ore Mt 0.0 0.0 0.0 0.0 0.0 0.0 0.8 0.9 1.5 1.5
Recovered Cu kt 0.0 0.0 0.0 0.0 0.0 0.0 29.9 36.1 24.7 24.7
Recovered Co kt 0.0 0.0 0.0 0.0 0.0 0.0 0.5 1.1 3.7 3.7
Waste Mt 0.0 0.0 0.0 0.0 0.0 0.0 7.0 9.8 11.3 12.5
Cu grade % 0.00 0.00 0.00 0.00 0.00 0.00 4.43 4.64 1.93 1.93
Co grade % 0.00 0.00 0.00 0.00 0.00 0.00 0.11 0.19 0.38 0.38
N
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Mashamba East Units 2022 2023 2024 2025 2026 2027 2028 2029 2030 Total
Ore Mt 1.5 1.2 1.2 1.2 1.2 1.2 0.3 0.3 0.0 12.8
Recovered Cu kt 24.7 27.9 27.9 27.9 27.9 27.9 8.6 7.7 0.0 296.0
Recovered Co kt 3.7 2.9 2.9 2.9 2.9 2.9 0.8 0.2 0.0 28.0
Waste Mt 13.0 13.3 13.0 4.5 4.9 0.1 0.0 0.0 0.0 89.4
Cu grade % 1.93 2.77 2.77 2.77 2.77 2.77 2.93 3.02 0.00 2.72
Co grade % 0.38 0.37 0.37 0.37 0.37 0.37 0.34 0.09 0.00 0.34
The LOM production profile for Mashamba East open pit is indicated in Figure 50 below.
Figure 50: Mashamba East LOM profile
It can be seen from Figure 51 below that all the ex pit tonnes generated throughout the LOM Plan from the Mashamba East open pit can be classified as oxide ore.
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Figure 51: Mashamba East ore type profiles
Figure 52 below indicates that a material portion of the ROM tonnes generated throughout the LOMP is from the inferred mineral resource category and cannot be classified as ore reserves as of the date of this report. The economic assessment is preliminary in nature, it includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as ore reserves, and there is no certainty that the preliminary economic assessment will be realized.
Figure 52: Mashamba East Reserve profile
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16.4.6 Mashamba East LOM reconciliation No changes occurred in the Mashamba East LOM Plan totals. A reschedule was conducted.
Table 78: Mashamba East LOM reconciliation Mashamba East Open pit 2012 LOMP 2011 LOMP
Ore tonnes (Mt) 12.8 12.8
Recovered Cu tonnes (kt) 296.0 299.0
Recovered Co tonnes (kt) 28.0 28.2
Waste tonnes (Mt) 89.4 89.4
Cu grade % 2.72 2.75
Co grade % 0.34 0.34
16.5 KCC Open Pit Ore Reserve Statement The ore reserve of the surface mining operations is based on pit optimisations, pit designs and production schedules generated with a December 31, 2011 base date. The ore reserve estimated and constrained to the surface mining operations is tabulated (Table 79) below.
Table 79: Surface mining operations Ore Reserve table as at December 31, 20111,2,3
Surface Mining operation
Proved Probable Total
Mt %TCu %TCo Mt %TCu %TCo Mt %TCu %TCo
T-17 Open pit 0.0 0.00 0.00 1.6 3.52 0.56 1.6 3.52 0.56
Mashamba East Open pit 0.0 0.00 0.00 5.9 3.00 0.36 5.9 3.00 0.36
KOV Open pit 0.0 0.00 0.00 55.1 4.74 0.45 55.1 4.74 0.45
Total 0.0 0.00 0.00 62.7 4.54 0.45 62.7 4.54 0.45
1) The ore reserve estimates have been prepared in accordance with the classification criteria of the JORC Code. If the definitions and classification standards of NI 43-101 had been used instead of those of the JORC Code, estimates of mineral reserves would be substantially similar to the estimates of ore reserves.
2) The ore reserve estimates are for KCC's entire interest in such ore reserves, whereas the Company owns 75% of KCC.
3) Numbers may not add due to rounding.
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16.6 Life of Mine Plan The combined KCC ROM production profile per operation is shown in Figure 53 below.
Figure 53: Life of Mine plan ROM ore profile for the combined KCC operation
The combined KCC waste stripping profile per operation is shown in Figure 54 below.
Figure 54: LOM Plan waste profile for the combined KCC operation
Due to the planned processing capacity of 310 000 tonnes of recovered copper per annum, the production profile was targeted and constrained to achieve this.
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Figure 55: LOM recovered copper profile for the combined KCC operation
The recovered cobalt per annum per operation is indicated in Figure 56 below. Due to the 30 ktpa constraint on the production of cobalt, beneficiation of ore from KOV pit for cobalt has been constrained to prevent cobalt over delivery.
Figure 56: LOM Plan recovered cobalt profile for the combined KCC operation
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16.7 Risk 16.7.1 KTO
The high levels of development required for the production ramp up, poses a risk to the achievement of the LOM profile. Any delay in the development requirements would result in a delayed ramp up profile from the KTO section. Since 2010, all planned development targets have been achieved.
A strategic revisit of the KTO LOM Plan could have a material impact on the KTO Reserve estimate. The KTO Reserve estimate is therefore valid until 2014, by which time a complete and appropriate LOM Plan based on the current mining strategy must be completed and approved.
16.7.2 KTE Although KTE is included in the LOM Plan, additional work is required to convert the ROM tonnes from Resources to Reserves. The access and mining strategies on which the LOM Plan is based are operational at KTO while mining related modifying factors are based on actual efficiencies at KTO. The risk could be mitigated by a comprehensive engineering and design study.
16.7.3 T-17 Detail implementation studies on the T-17 underground mine only exist on a medium term basis up to the end of 2014. Appropriate work is highly recommended to define the technical mining plan for T-17 and understand the role T-17 underground would play within the composite KCC operational plan.
16.8 Surface Operations Dewatering: Ongoing dewatering of KOV open pit and initial and continuous dewatering of Mashamba
East open pit is required. The dewatering of the T-17 extension area and protection of the river in culverts is a prerequisite for production to commence at T-17 extension. The current dewatering strategy has recently been revised to accommodate the pit designs referenced in the 2010 Technical Report and historical hydrogeological data which is difficult to validate. As such, the current dewatering strategy could potentially prove to be not as effective as designed to maintain dry slope conditions. This potentially could introduce risks associated with potential impacts on both the economics of the pit, pit stability and the production schedule. This risk could be mitigated by performing further studies to gain a better understanding of the hydrogeology so that an appropriate dewatering strategy is defined that enables safe mining as cost effectively and efficiently as possible.
Access and slope failures: KOV open pit will develop into a large operation up to 400m deep. Production rates are high with up to 45 million tonnes of material that should be moved from the pit per annum. Small slope failures could have a negative impact on access to the production benches. A strategy should be developed to establish an alternative ramp access to production areas. Slope failures could negatively affect safety and production performance at the KCC surface operations. This is a not an unknown risk and is monitored by on site personnel on a continual basis. Dual access to working areas should be established to decrease the risk of production losses.
Available pit space: A high production rate requires sufficient working areas or face length. Additional face length should be established and maintained on the southern section of the KOV open pit.
Available waste dumping space: The KOV open pit produce 462 million tonnes of waste up to 2030. The available waste dumping space close to the KOV open pit is insufficient although as part of the ARJVA additional surface rights have been allocated to KCC to the north of the current KOV open pit concession. However, for improved cost efficiency, a detailed technical plan to back fill mining waste into depleted pits should be developed. With the appropriate mining sequence in KOV open pit, a total of 150 million tonnes could be back filled while the T-17 open pit is available for back filling from 2013 onwards should the appropriate technical studies be completed.
Grade control: An operational cut-off grade of 0.6% Cu has been applied to the mining models. Grade control practices for a high volume operation should be developed. Inefficient grade control
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systems could result in Resource losses or that uneconomic production tonnes are processed where the planned revenues does not cover the processing and selling costs.
16.9 General The processing plant design assumes feed copper and cobalt grades within a predefined range. Should the grades vary materially from the design parameters, a higher ROM production rate is required to maintain the recovered metal profiles. Should the required increased plant feed exceed the plant feed design, the ability to maintain the recovered copper profile could be compromised.
17.0 RECOVERY METHODS: PLANT AND PROCESSING 17.1 General Process Commentary The KTC and the Luilu Refinery constitute the process plants of KCC with KTC being approximately 4 km from Kolwezi and the Luilu Refinery being 6 km distance from KTC.
KCC processing plants produce copper cathode metal, cobalt metal and copper concentrate as products from ores received from KOV Open Pit, T-17 Open Pit and the KTO. Future production may be derived from Mashamba East Open Pit and potentially T-17 Underground Mine. As described in the section 1.6.2, ore sourced from the open pits are predominantly oxide, disseminated with sulphide mineralization in certain locations and depths within the pits. This has the potential to produce a mixed ore feed to KTC. Ore sourced from KTO is almost exclusively sulphide.
The metallurgical processing is currently divided into two areas, namely crushing and milling and flotation at KTC and copper and cobalt metal recovery through a leach and electro winning process at the Luilu Refinery.
The value of the plant and equipment has been taken into account in the economic evaluation as set out in section 21.0 and section 22.0.
17.2 Status Update Between November 2009 and August 2011, KCC undertook the Kamoto Phase 3 project which was a continuation of the earlier rehabilitation of the existing process plant facilities. The design output of Phase 3 was 150,000 tpa of copper and 8,000 tpa of cobalt metal.
In April 2011, KCC commenced the Kamoto Phase 4, which will increase output to 270,000 tpa of copper and 8,000 tpa of cobalt metal, with additional cobalt in the form of hydroxide.
Phase 4 will generally comprise modifications to the existing plant with a limited number of additional major items of plant and equipment such as a ROM crusher and conveyor, additional flotation units, a new roaster, an SX plant and conversion of the existing unused Luilu electro-refinery to an electro-winning plant. Certain elements of the Phase 3 Project have been carried over to Phase 4. The estimated cost of Phase 4 is USD$635 million and is due for completion in July 2013.
17.3 Processing Facilities 17.3.1 KTC Original Facilities The original KTC consisted of Kamoto 1 and 2 sections built in 1968 and 1972 respectively and DIMA 1 and 2 sections built in 1981 and 1982 respectively.
Kamoto 1 treated mixed ore and oxides. The circuit comprised the following unit processes:
Autogenous milling operating in closed circuit with hydro cyclones;
Sulphide flotation including roughing, cleaning and middlings regrind to produce a sulphide concentrate;
Sulphidisation of oxide minerals; and
Oxide flotation including roughing and cleaning to produce an oxide concentrate.
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Kamoto 2 primarily treated sulphide ore from the KTO. The circuit comprised the following unit processes:
Autogenous milling operating in closed circuit with hydro cyclones; and
Sulphide flotation including roughing, cleaning and middlings regrind.
The DIMA 1 circuit primarily treated oxides and mixed oxide/sulphide ore feeds. The circuit comprised the following unit processes:
Primary autogenous milling and secondary ball milling operating in closed circuit with hydro cyclones;
Sulphide flotation including roughing, cleaning, re-cleaning and middlings re-grind to produce a sulphide concentrate;
Sulphidisation of oxide minerals; and
Oxide flotation including roughing and cleaning to produce an oxide concentrate.
The DIMA 2 circuit treated oxide ore. The circuit comprised the following unit processes:
Primary autogenous milling and secondary ball milling operating in closed circuit with hydro cyclones;
Sulphidisation of oxide minerals; and
Oxide flotation including roughing and cleaning to produce an oxide concentrate.
17.3.2 KTC Current Operations 17.3.2.1 Oxide Ore Crushing and Milling Mixed ore from KOV Open Pit is transported by truck and stockpiled near the B3 jaw crusher at KTC. The B3 jaw crusher has a nameplate capacity of 500 tph. Mixed ore is blended, crushed to minus 250 mm and conveyed to stockpiles. Ore is milled in two 28 feet cascade mills with nameplate capacities of 150tph each, both in closed circuit with cyclones. Cyclone underflow is milled in ball mills. Final milled product is nominally 70% to 75% minus 75µm.
Mixed ore from KOV Open Pit is also transported by truck and stockpiled near the B4 jaw crusher at KTC. The B4 jaw crusher has a nameplate capacity of 700tph. Mixed ore is blended, crushed to minus 250 mm and conveyed to stockpiles. Ore is milled in two 32 feet cascade mills with a capacity of 300tph each, both in closed circuit with cyclones. Cyclone underflow is milled in ball mills. Final milled product is nominally 70% to 75% minus 75µm.
17.3.2.2 Sulphide Ore Crushing and Milling Sulphide ore from KTO is crushed underground by gyratory crushing to minus 400 mm and hoisted to surface. From there it is conveyed to stockpiles ahead of KTC. The ore is milled in two 28 feet cascade mills with a nameplate capacity of 150tph each in closed circuit with cyclones to 90% minus 75µm.
17.3.2.3 Oxide Flotation Milled oxide ore is subjected to a 3 phase flotation process, namely roughing, scavenging and cleaning. Reagent addition in the oxide flotation section is more complex than in the sulphide flotation section: namely a collector, potassium normal butyl xanthate (PNBX), a dispersant (Na2SiO3 - water glass), a sulphidiser (NaHS), a frother (G41) and a blend of fatty acid collectors (Rinkalore).
Copper recovery in the oxide circuit is consistent with industry standards. Final oxide concentrate is pumped via pipelines to the Luilu Refinery.
Tailings are classified through a bank of cyclones and either pumped to the tailings dam or underground for back-fill
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17.3.2.4 Sulphide Flotation The milled sulphide ore is subjected to a 5 phase flotation process of roughing, scavenging, cleaning, re-cleaning and final re-cleaning. Final sulphide concentrate is pumped via pipelines to the Luilu Refinery. In the sulphide flotation section, only G41 and a collector, sodium iso-butyl xanthate (SIBX) are used.
Tailings are cycloned and either pumped to the tailings dam or underground for back-fill.
17.3.3 KTC Project Development As part of the implementation of Phase 3, KTC was rehabilitated to its original installed milling capacity of 7.6Mtpa, equivalent to 230ktpa finished copper production at standard process efficiencies and resource model grades.
In conjunction with the mills rehabilitation program, KCC site manufactured flotation cells, consistent with the original design, to add more flotation capacity.
Oxide concentrate that is surplus to the Luilu Refinery Phase 3 capacity is forwarded to a filtration, bagging and storage plant at KTC which has a design plant capacity of 10ktpm concentrate (grading > 22%Cu). The oxide concentrate is exported as a final product.
Figure 57: KCC Filtration, Bagging and Storage Plant
The following plant and process description outlines Phase 4 which is currently in implementation at KTC.
17.3.3.1 Crushing, Materials Handling and Ore Storage A new crusher will be located between the KOV Open Pit and KTC to treat the KOV ROM mixed ore. The crusher product will be conveyed via an overland conveying system to the CM1 and CM2 stockpiles or the CM6 and CM7 silos and stockpiles.
The T-17 Open Pit ore will be transported by mine haul trucks to a new crusher adjacent to the new KOV Open Pit ROM ore crusher. The crusher product will be conveyed to the CM1 and CM2 stockpiles or the CM6 and CM7 silos and stockpiles.
The ROM ore from the Mashamba East Open Pit will be transported by mine haul trucks to the B4 crusher circuit. The crusher product will report to the CM6 and CM7 silos and stockpiles.
17.3.3.2 Milling Circuits Sulphide ore milling will consist of CM3, CM4 and BM3 milling circuits and will be used for sulphide milling exclusively.
The mixed ore milling will consist of CM1/BM1, CM2/BM2, CM6/BM6 and CM7/BM7 milling circuits.
A review of the existing DIMA mills identified the potential to return these mills from autogenous operation to a semi-autogenous operation. This had the potential to increase total installed milling capacity from 7.6Mtpa to ~9Mtpa, equivalent to 270ktpa finished copper production. Investigations are in progress to
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confirm the expectation that returning the DIMA mills to semi-autogenous operation will increase the total installed milling capacity to ~9Mtpa. A further increase to 310ktpa copper will necessitate the installation of additional crushing, milling and flotation capacity.
17.3.3.3 Flotation Circuits The sulphide ore flotation circuit will handle CM3 and CM4 milling circuits’ products and will consist of roughing, scavenging and three stage cleaning circuits. There will be minor modifications to the existing circuits which use 8.5m3 self aerated cells.
The mixed ore flotation circuits will comprise 3 areas, namely; CM1/CM2, CM6 and CM7. Each area will have dedicated sulphide rougher, oxide rougher, sulphide cleaner and oxide cleaner circuits. Existing 8.5m3 self aerated cells will be used where possible and additional flotation capacity will comprise new 50m3 and 20m3 tank cells. Concentrate re-grind circuit will be followed by sulphide and oxide flotation to further liberate the locked sulphides and reduce the oxides in the sulphide concentrate.
17.3.3.4 Concentrate Handling The sulphide concentrate will be transferred by 2 pump trains (1 duty, 1 standby) to the sulphide concentrate receiving section at the Luilu Refinery. The oxide concentrate will be transferred by 2 pump trains (1 duty, 1 standby) to the oxide concentrate receiving section at the Luilu Refinery.
The Kamoto Concentrator Phase 4 Block Flow Diagram set out in Figure 58 below illustrates the:-
Existing Facility (no modifications)
Existing Facility (minor modifications)
Existing facility (major modifications)
New Facilities
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Figure 58: The Kamoto Concentrator Phase 4 Block Flow Diagram
One TwoSulphide Ore Mixed Ore
Crusher Crushers
ThreeSulphide Mill Mixed Ore Mill
Stockpile Stockpile
Five Six NineteenSulphide Milling Mixed Ore Milling Oxide Concentrate
Transfer
Fifteen Thirteen Seven Eight Twenty OneSulphide Re-cleaner Sulphide Cleaner Sulphide Rougher Mixed Ore Sulphide Concentrate Eighteen
Flotation Flotation Flotation Rougher Flotation Regrind Oxide CleanerFlotation
Sixteen Fourteen Nine Eleven Twenty TwoSulphide Re-recleaner Sulphide Cleaner Sulphide Scavenger Mixed Ore Oxide Regrind Sulphide
Flotation Scavenger Flotation Rougher Flotation Flotation
NotesTwenty Three 1. Tailings Thickening and Disposal to beRegrind Oxide done by SRK and not SNC-Lavalin.
FlotationLegend
Twenty FourMixed Ore Sulphide - New Facility
Seventeen - Existing Facility (major mods)Sulphide Concentrate Twenty
Transfer Note 1 Tailings Thickening - Existing Facility (minor mods) and Disposal
- Existing Facility (no mods)
Twenty FiveFlotation Reagents
2011/06/06 PB Client Approval ND MM JD30/05/2011 PA Internal Review ND MM JD Document Number Rev
Date Rev. Description Des. Chk. App. 150063A-0000-49EC-0002 PB
Kamoto Redevelopment Project Phase 4
Kamoto Concentrator Block Flow Diagram
Four
Cleaner Flotation
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17.3.4 Luilu Refinery Original Facilities Production at the Luilu Refinery, located approximately 6km north of KTC, commenced in 1960. The process route employed was roast leach-electro-winning which was typical of other contemporary DRC and Zambian Copperbelt operations. The circuit comprised the following unit processes:
Sulphide and oxide concentrate receipt, dewatering and storage;
Sulphide concentrate roasting;
Sulphuric acid copper leach of roaster calcine and oxide concentrate (oxidising leach assisted by air injection);
Secondary leach using high acid-consuming (dolomitic) concentrates;
Counter-current decantation and clarification;
Leach tailings filtration and residual sulphide flotation;
Tailings neutralisation and disposal;
Selenium removal via up-flow reactor containing copper granules;
Copper EW onto copper starter sheets;
De-copperising of cobalt bleed solution – two-stage EW;
Cobalt bleed solution purification including the following steps:
Iron removal by controlled pH precipitation using milk of lime;
Copper removal by two-stage controlled pH precipitation using milk of lime;
Nickel removal by controlled pH precipitation using NaHS and cobalt chips;
Zinc removal by the addition of hydrogen sulphide (H2S) and neutralisation with sodium carbonate solution;
Controlled pH precipitation of cobalt with milk of lime;
Cobalt re-leaching with spent electrolyte and sulphuric acid under controlled pH;
Cobalt EW; and
Cobalt vacuum degassing and burnishing.
The Luilu Refinery was designed to process sulphide and oxide concentrates with an initial capacity of 80ktpa copper cathode. During the 1970’s, capacity was expanded to 175ktpa copper cathode and 8ktpa cobalt cathode. The grade of cathode copper produced in the first EW stage never met LME Grade 'A' quality, while most of the cathode and copper sponge produced in the secondary EW was not of commercial quality, and, was recycled to the Shituru smelter at Likasi. Cobalt recovery across the plant was <65%, with the majority of the cobalt losses occurring at nickel and zinc sulphide precipitation, with some also at iron removal and cobalt precipitation.
The condition of the plant in 2004, when taken over by KCC, was extremely poor and almost totally run down. A progressive renewal programme was planned, to match the increasing throughput. Considerable progress has been made to-date in the phased rehabilitation exercise. Completion of Phase 1 was in December of 2007 and completion of Phase 2 in December of 2009. The new roaster was commissioned in late 2009.
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Phase 3 undertaken by SNC-Lavalin (South Africa) has essentially completed the majority of the rehabilitation of KTC and the Luilu Refinery, although certain elements have been incorporated into Phase 4. A process simulation model was developed which indicated that on completion of Phase 3, the KTC and Luilu Refinery will produce 150ktpa of copper and 8ktpa of cobalt.
17.3.5 Luilu Refinery: Current Operations 17.3.5.1 Concentrate Reception The primary purpose of concentrate thickening and filtration is to create a storage buffer capacity at the Luilu Refinery, prior to sulphide roasting and oxide leaching. In addition it removes water from the concentrates to lower reagent consumption in the subsequent, leaching circuit.
Concentrates received from KTC are dewatered in 2 oxide and 2 sulphide thickeners prior to being pumped to a set of 5 drum filters. In Phase 3 an oxide thickener and 2 drum filters were installed.
The filtered oxide and sulphide concentrate remain in separate circuits and are conveyed to a storage shed. They are reclaimed for feeding to the roasting and leaching circuits respectively.
17.3.5.2 Sulphide Concentrate Sulphide concentrate is reclaimed from a bay in the storage shed and conveyed to a repulp mill where it is slurried at high pulp density. The slurry is injected into 2 or 3 of 3 roasters; one of 450tpd nameplate capacity and 2 of 150tpd nameplate capacity. The roasters operate at 650°C. Slurry is injected with excess fluidizing air. The sulphide concentrate is roasted to a calcine which comprises acid soluble compounds of copper and cobalt. The hot calcine is cooled to below 300°C and quenched in spent electrolyte from the copper EW tank-house. Hot gases are passed through cyclones for dust capture, cooled with spent electrolyte, scrubbed and discharged through stacks. The calcine/spent electrolyte slurry is pumped to the leaching section.
17.3.5.3 Oxide Concentrate Oxide concentrate is reclaimed from a bay in the storage shed and repulped with spent electrolyte.
17.3.5.4 Leach Circuits In the atmospheric leach section, the repulped oxide slurry is combined with the calcine/spent electrolyte slurry from the roasters and with sulphuric acid. In order to improve the dissolution of cobalt, controlled amounts of sodium meta-bisulphite (Na2S2O5) is added to the leach.
After leaching, the slurry is thickened and the overflow clarified. The underflow is subjected to counter current decantation and residue filtration to separate liquid and residue solids. The residue solids contain most of the sulphides entering the circuit with the oxide concentrate which bypasses the sulphation roast and remain in a form which is not acid soluble.
The residue solids are subjected to flotation where most of the remaining sulphides are recovered and pumped to the sulphide thickeners for recovering the metal values. Residue flotation tailings are pumped to ponds after neutralization with lime.
17.3.5.5 Copper Circuits The clarified overflow is pumped to the copper EW tank-house, where commercial copper is electro-won in cells with lead-antimony anodes and stainless steel cathodes. A small stream of the clarified overflow goes to a starter sheet section where starter sheets are made for the de-coppering section. A bleed stream of spent electrolyte is taken off and subjected to iron removal by oxidation and by adjustment of the pH with copper hydroxide from a de-coppering step in the cobalt plant. The precipitated ferric-hydroxide is thickened, filtered and pumped to residue tailings. Another bleed stream of spent electrolyte is subjected to secondary copper electro winning in two stages to remove the bulk of the copper before pumping it to the cobalt plant. The bulk of the spent electrolyte is pumped to leaching and to cool the off gas and calcine in the roasters.
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17.3.5.6 Cobalt Circuit The feed to the cobalt recovery circuit is a bleed stream taken from the copper EW tank-house.
In the purification part of the cobalt plant, impurities are removed by precipitation, thickening and filtration in the following order:
Iron and aluminium by oxidation with air and pH adjustment with milk of lime;
Copper precipitation by pH adjustment with milk of lime (filter cake recycled back to ferric precipitation);
Nickel removal with gaseous H2S and recycled cobalt granules and pH adjusted with sulphuric acid. This unit is currently not in operation as nickel levels are very low; and
Zinc removal with gaseous H2S and sodium carbonate for pH adjustment.
After zinc removal, the cobalt is precipitated from solution with milk of lime at pH of 7.8 to 8.4 and filtered. The washed cobalt hydroxide is then dissolved with spent cobalt electrolyte and sulphuric acid, subjected to thickening, clarification and polishing filtration before pumping as advance electrolyte to the cobalt EW tank-house.
In the cobalt EW tank-house, the electrolyte is heated to 70°C and cobalt is electro-won in cells with lead-antimony anodes and stainless steel cathodes. After stripping, the cobalt cathode pieces are subject to heat treatment and polishing for production of final product.
The stripped electrodes are returned back into the electro winning circuit to repeat the process.
17.3.6 Luilu Copper Electro-Refinery The following plant and process description outlines Phase 4 which is in implementation at the Luilu Refinery.
It should be noted that the sulphide and oxide reception area at Luilu is under process development and the description below, may be superseded.
17.3.6.1 Sulphide Concentrate Sulphide concentrate will be thickened in the existing sulphide concentrate thickeners and filtered in new pressure filters to produce a dry concentrate cake for stockpiling in existing facilities. The concentrate will be reclaimed, milled in an existing ball mill and fed to repulping tanks prior to the repulped slurry being fed to roasters.
17.3.6.2 Oxide Concentrates Oxide concentrate will be thickened in the existing sulphide concentrate thickeners and filtered in new pressure filters to produce a dry concentrate cake for stockpiling in existing facilities. The concentrate will be reclaimed and repulped with raffinate from the SX plant in oxide repulping tanks. The repulped slurry will become a feed to the primary leach plant
17.3.6.3 Sulphide Concentrate Roasting The concentrate from the sulphide repulp will be treated in sulphating roasters (2 old roasters, 1 existing Hatch roaster and a new Phase 4 roaster) to produce an acid soluble calcine. The calcine will be repulped with raffinate from the SX plant before being fed to the primary leach circuit.
17.3.6.4 Primary Leaching The calcine slurry and oxide slurry will be pumped to the primary leach circuit consisting 5 primary leach trains. The primary leach circuit will leach copper in dilute sulphuric acid under atmospheric pressure to produce a copper rich liquor and residual solids. The sulphuric acid will arise from the raffinate recycle and any additional acid requirements will be made up with concentrated sulphuric acid.
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17.3.6.5 Thickening and Residue Handling The leach slurry will be fed to the solid liquid separation circuit ahead of the SX circuit. The slurry will be thickened in 9 primary thickeners and 9 clarifiers to produce a clean overflow with low solids content. This clean overflow will be the pregnant liquor solution (PLS) feed to the SX plant. The primary thickener underflow will be treated by two stages of residue washing. The first stage will be counter current decantation and the second stage is by residue filtration belt filters. The cake from the residue belt filters will be repulped and then discharged to residue flotation where undissolved sulphide copper will be recovered. The flotation residue will be discharged to a tailings dam.
17.3.6.6 Copper Solvent Extraction The clarified PLS will be pumped to a new 3 train copper SX plant. In each train the copper from the clarified PLS will be loaded onto the organic in the extraction stages to produce a loaded organic stream and a raffinate stream with a low copper concentration. The raffinate will be recycled back to leach, whereas the loaded organic will be treated in the stripping stages to produce a loaded electrolyte and a stripped organic stream. The stripped organic will then be recycled to the extraction stages. The loaded electrolyte will be treated in filters to remove any entrained organic prior to being fed as advance electrolyte to the electro winning plants.
Figure 59: Earthworks underway for the construction of the SX Plant
17.3.6.7 Electro winning The existing electro winning plant at the Luilu Refinery will be upgraded to accept advance electrolyte from 1 of the SX plant trains. A cathode washing and stripping machine will be incorporated in the upgrade. This electrowinning plant will be run as a variable top up facility with an output up to ~100,000tpa of LME Grade A copper.
The existing unused electro-refinery will be converted to an electrowinning plant to accept advance electrolyte from 2 of the SX plant trains. 2 cathode washing and stripping machines will be incorporated in the conversion. This electro winning plant will be run on base load with an output of ~200,000tpa LME Grade A copper.
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In the electrowinning plants, the advance electrolyte will be treated by plating pure copper from the copper sulphate solution onto stainless steel blanks. The pure LME Grade A copper will then be stripped off the cathode blanks and stockpiled before being sold. The spent electrolyte will be recycled back to the SX plant to be used as strip liquor.
Figure 60: Electrowinning plant under conversion
17.3.6.8 Iron and Aluminium Precipitation In order to prevent the build up of cobalt and impurities in the processing circuits, it will be necessary to bleed a small portion of raffinate out of the process for recovery of cobalt and copper. The first impurity removal circuit will be the precipitation of iron and aluminium out of solution, by adjusting the pH with lime.
17.3.6.9 Copper Precipitation The solution from the iron and aluminium precipitation circuit will be treated with lime to further raise the pH and precipitate the copper. The precipitate will be fed back to the leach circuit to re-dissolve the copper. The overflow will be fed to the zinc SX circuit.
17.3.6.10 Zinc Solvent Extraction The solution from the copper precipitation circuit will be fed to the zinc SX circuit. The raffinate stream from the zinc SX plant will become the feed to the cobalt precipitation plant. The zinc rich loaded electrolyte will be bled from the circuit to residue disposal, where it will be neutralised.
17.3.6.11 Cobalt Recovery Cobalt will be precipitated from solution by raising the pH with lime to produce a cobalt free solution and a cobalt rich precipitate. The slurry will be thickened and filtered. Approximately half the cake will be bagged and sold. The remainder will be re-dissolved in spent cobalt electrolyte. The slurry after dissolution will then be thickened and filtered to remove impurities and the clean cobalt rich solution will be fed to the existing cobalt electro-winning circuit to plate the cobalt onto the cathode blanks.
The Luilu Phase 4 Block Flow Diagram in Figure 61 set out below illustrates this:
Existing Facility (no modifications);
Existing Facility (minor modifications);
Existing facility (major modifications); and
New Facilities.
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one two thirty three thirty two thirty oneSulphide concentrate Oxide concentrate Cobalt catholyte Gypsum Gypsum
Thicken and filter Thicken and filter Clarification Filtration Thickening
three four thirty four thirtySulphide concentrate Oxide concentrate Cobalt Cobalt hydroxide
Stockpile Stockpile electrowinning Redissolution
five six twenty eight twenty nineSulphide concentrate Oxide concentrate Cobalt hydroxide Cobalt hydroxide
Repulping Repulping Thickening Filtrationsixteen
Fe/Al Ppt.twenty seven
Cobalt hydroxide Cobalt cakePrecipitation to toll refining
seventeen & eighteenFe/Al Thickening
seven and Filtration twenty sixSulphide concentrate Zinc SX
Roastingthirty five
Copper Removaleight eleven by Ppt
Calcine quench Primary leach Legend
23 24 - New FacilityThickening
and clarification twenty two - Existing Facility (major mods)Primary Copper
twenty New EW - Existing Facility (minor mods)HGSX
- Existing Facility (no mods)thirteen twelve nineteen twenty one
CCD Primary thickener Primary Primary copperClarification Existing EW
fourteenResidue filtration
fifteen Luilu Block Flow DiagramResidue flotation Residue disposal 2011/06/06 PB Client Approval ND MS JD
30/05/2011 PA Internal Review ND MS JD Document Number RevDate Rev. Description Des. Chk. App. 150063A-0000-49EC-0001 PB
Kamoto Redevelopment Project Phase 4
Figure 61: The Luilu Phase 4 Block Flow Diagram
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18.0 PROJECT INFRASTRUCTURE The project infrastructure is addressed in paragraph 1.5 which refers to the Material Assets and in paragraph 17.0 dealing with the Plant and Processing. The major change in infrastructure relate to Phase 4 of the upgrade to the Luilu refinery.
19.0 MARKET STUDIES AND CONTRACTS 19.1 Markets 19.1.1 Copper Copper is a major industrial metal (ranking third after iron and aluminium by consumption) because it is highly conductive (electrically and thermally), highly ductile and malleable and resistant to corrosion. Electrical applications of copper include power transmission and generation; building wiring; motors; transformers; telecommunications; electronics and electronic products; and renewable energy production systems. Copper and brass (an alloy of copper) are the primary metals used in plumbing pipes, taps, valves and fittings. Further applications of copper include decorative features; roofing; marine applications; heat exchangers; and in alloys used for gears, bearings and turbine blades.
Global copper mine production was 16.0Mt in 2010, with 7.0Mt (or 44%) produced in Chile, by far the largest producer. Africa produced 1.2Mt (7.5%). Global refinery production in 2010 was 19.0Mt. Global consumption was slightly higher at 19.4Mt. The International Copper Study Group (4 October 2011) estimates global mine production for 2012 at 17.6Mt, with global consumption at 20.4Mt. Table 80 shows the historical and 2012 forecast global refined copper market balance.
Table 80: Global refined copper market balance (Source: USGS 2006-2009, ICSG 2010-2012)
Production 2006 2007 2008 2009 2010 2011 2012
Mine Production (Mt) 15.0 15.4 15.5 16.0 16.0 16.0 17.6
Refined Production (Mt) 17.2 17.9 18.2 18.3 19.0 19.4 20.1
Consumption (Mt) 17.0 18.2 18.0 18.1 19.3 19.7 20.3
LME Copper Price ($’000/t avg) 6.7 7.1 7.0 6.1 7.5 8.8 8.2
*Forecast †1 Jan - 16 Feb 2012
The copper price has demonstrated significant volatility since 2008, as shown in Figure 62. The price of copper reached a new high of $8,900 /tonne in July 2008. Thereafter, as the financial crisis took effect on the global economy, the price declined to $2,810 /tonne in December 2008, the lowest level in almost 5 years. From that low, the price generally trended upward, reaching a new record high of above $10,000 /tonne. After trading in a band from $10,000 to $9,000 /tonne, the price sharply declined in August 2011 to around $6,700 /tonne. It has recovered since then, with a bid price on 16 February 2012 of $8,209 /tonne.
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Figure 62: The London Metal Exchange copper price from January 2007 to date (Source: LME)
The copper price forecast used in the economic evaluation of the project is shown in Table 81. The forecast is based on published LME monthly futures prices. These publicly available prices are quoted in nominal terms. The financial model used for the economic evaluation is in real terms (2012 United States dollars), and the real copper price forecast is derived from the nominal prices using the US CPI estimates in Table 81.
Table 81: Copper price forecast Copper price ($/tonne) 2012 2013 2014 2015 2016 2017 2018 2019 2020 Long
Term
Real 8,500 8,416 8,823 8,735 8,649 8,087 7,536 6,757 6,213 6,000
US CPI 1.0% 1.0% 1.0% 1.0% 1.0% 1.5% 1.5% 1.5% 1.5% 1.5%
19.1.2 Cobalt Cobalt has many commercial, industrial and military applications. The leading use of cobalt is in rechargeable battery electrodes. The temperature stability and heat- and corrosion-resistance of cobalt-based superalloys makes them suitable for use in turbine blades for jet turbines and gas turbine engines. Other uses of cobalt include vehicle airbags; catalysts for the petroleum and chemical industries; cemented carbides and diamond cutting and abrasion tools; drying agents for paints, varnishes, and inks; dyes and pigments; ground coats for porcelain enamels; high-speed steels; magnetic recording media; magnets; and steel-belted radial tyres.
Global mine production of cobalt was 88,000 tonnes in 2010, with 45,000 tonnes (or 51%) produced in the DRC, the largest producer. Zambia is the second largest producer with 11,000 tonnes (12.5%). Global refinery production in 2010 was 76,000 tonnes. Table 82 shows the historical global refined cobalt market balance.
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Table 82: Global refined cobalt market balance Metric tonnes 2006 2007 2008 2009 2010
Global Mine Production (Mt) 0.7 0.7 0.8 0.8 0.8
Total World Refined Production (Mt) 0.5 0.5 0.5 0.6 0.7
Cobalt Price ($/lb avg) 15.35 28.31 36.16 15.89 18.75
Source Cobalt News (Oct 2005 – Jan 2012) Published by the Cobalt Development Institute
The cobalt price reached a record of $48.63 /pound in March 2008, falling in line with other commodities to a 5-year low of $11 /pound in December 2008. The price recovered, reaching a maximum of $19.64 /pound in January 2011. Cobalt started trading on the LME in May 2010, and the LME cobalt price is shown in Figure 63 in $/tonne ($10/lb = $22 046/tonne). The cobalt price declined during 2011, reaching a low of $12.25 in October. The LME closing price was $32 250/tonne ($14.63 /lb) on 16 February 2012.
Figure 63: The London Metal Exchange cobalt price from April 2010 to date (Source: LME)
The cobalt price forecast used in the economic evaluation of the project is shown in Table 83. The forecast is based on the Metal Bulletin 99.8%Co $/pound price (in nominal terms) available for the next spot delivery. The forward curve is assumed to gradually decline for the next three years, before falling to its long term value. The financial model used for the economic evaluation is in real terms (2012 USD), and the real cobalt price forecast is derived from the nominal prices using the US CPI estimates in Table 83.
Table 83: Cobalt price forecast Cobalt price ($/pound) 2012 2013 2014 2015 2016 2017 2018 2019 2020 Long
Term Real 14.00 13.86 11.76 11.65 11.53 11.42 11.30 10.81 10.65 10.50
US CPI 1.0% 1.0% 1.0% 1.0% 1.0% 1.5% 1.5% 1.5% 1.5% 1.5%
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19.2 Contracts KML has entered into offtake agreements with Glencore International AG, pursuant to which Glencore will buy 100% of the quantities of Cu and Co produced by KCC over the life of the mine. The offtake agreements are negotiated at arm’s length at standard market terms.
20.0 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL COMMUNITY IMPACT
20.1 Terms of Reference The scope of the environmental and social review focused on the key environmental, social risks and liabilities associated with the mining and processing assets owned by KCC. The review specifically focused on:
Compliance with DRC Legislation and the Equator Principles;
Emissions and releases (air, water, noise and vibration); and
Social risks (such as artisanal miners, health, reputational and human rights).
The review was conducted at KCC on January 18 - 19 2012 and included:
Interviews with the acting environmental manager, health and safety manager and process managers for KCC;
Collection and review of environmental permits, reports, management plans and monitoring data; and
A site visit to surface infrastructure and processing facilities.
The review was of current operations and, where possible, it included environmental and social performance of all KCC operations.
20.1.1 Regulatory Context Set out below is a summary of the legal and policy aspects that define the compliance criteria utilised during the review process.
The key environmental and social legislation utilised for the review is summarised in Table 84 below.
Table 84: Summary of relevant Environmental and Social DRC legislation relating to KCC Legislation Summary
Law No. 007/2002 of July 11, 2002 Relating to the Mining Code
General - Provides the regulations relating to prospecting, exploration, exploitation, processing, transportation and sale of mineral substances. It includes mining restrictions, tax regimes, environmental impact study, site rehabilitation and compensation requirements.
Decree No 038/2003 of 26 March 2003 of the Mining Code
Annexe I – Indicates which authorities are competent to determine restricted areas. Annexe II – Provides guidelines for the provision of a financial guarantee for environmental rehabilitation. Annexe IV - Regulations on storage sites for mining and quarrying products. Annexe IX - Guidelines for Environmental Impact Studies. Annexe X – Minimum standards for rehabilitation and closure of mining facilities. Annexe XI – relates to classification of mine waste and subsequent
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Legislation Summary
protection measures. Annexe XII – Defines restricted and sensitive environments Annexe XIII – Defines noise measurement methodology Annexe XIV – Defines the minimum standards for ensuring the structural stability of mining waste storage areas
Forest Code (Law 11 of 2002) Restricts the development or exploitation of forestry areas and defines the requirements for applying for the development or exploitation of forestry.
Decree of 6 May 1952 on Water Establishes provincial commission to prevent misuse of water resources.
Regulations on lake and water course contamination and pollution 1 July 1914
Determines protection areas for streams, lakes, and other water resources which may constitute a source of drinkable water.
Regulation 79-244 of 16 October 1997 Defines the tax regime for exploitation of forest products, hunting and fishing licences, operating permits for dangerous, unhealthy, or nuisance establishments.
Regulation no 69-041 of 22 August 1969
Sets a framework for the general conservation of wild life in general by preventing and controlling hunting, trapping, transporting, disturbing or illegally causing animals to flee from their natural habitat.
20.1.2 The Equator Principles The Equator Principles are a set of principles that have been adopted by 73 international finance institutions (known as the Equator Principles Financial Institutions or EPFIs). The principles have been adopted to ensure that EPFIs fund only projects which are or will be developed in a socially responsible manner reflecting sound environmental management practices. The EPFIs commit not to provide loans to project’s where the borrower will not or is unable to comply with the respective environmental policies and procedures that implement the Equator Principles (www.equator-principles.com).
Specifically the Equator Principles include:
Principle 1: Review and Categorisation;
Principle 2: Social and Environmental Assessment;
Principle 3: Applicable Social and Environmental Standards;
Principle 4: Action Plan and Management System;
Principle 5: Consultation and Disclosure;
Principle 6: Grievance Mechanism;
Principle 7: Independent Review;
Principle 8: Covenants;
Principle 9: Independent Monitoring and Reporting; and
Principle 10: EPFI Reporting.
The Equator Principles primarily include and directly reference the International Finance Corporation (IFC) environmental and social performance standards and World Bank general and industry specific environmental, health and safety (EHS) guidelines (see below).
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This review references the Equator Principles and IFC Performance standards as part of the compliance criteria.
20.1.3 The International Finance Corporation Performance Standards The IFC performance standards on social and environmental sustainability were developed by the IFC and were last updated on 1 January 2012. They comprise of eight performance standards namely:
Performance Standard 1: Assessment and Management of Environmental and Social Risks and Impacts;
Performance Standard 2: Labour and Working Conditions;
Performance Standard 3: Resource Efficiency and Pollution Prevention – linked to the World Bank Environmental Health and Safety (EHS) Guidelines;
Performance Standard 4: Community Health, Safety, and Security;
Performance Standard 5: Land Acquisition and Involuntary Resettlement;
Performance Standard 6: Biodiversity Conservation and Sustainable Management of Living Natural Resources;
Performance Standard 7: Indigenous Peoples; and
Performance Standard 8: Cultural Heritage.
Performance Standard 1 establishes the importance of:
(i) Integrated assessment to identify the social and environmental impacts, risks, and opportunities of projects;
(ii) Effective community engagement through disclosure of project-related information and consultation with local communities on matters that directly affect them; and
(iii) The management of social and environmental performance throughout the life of the project.
Performance Standards 2 through 8 establish specific requirements to avoid, reduce, mitigate or compensate for impacts on people and the environment, and to improve conditions where appropriate. Performance Standards 2 through 8 describe potential social and environmental impacts that require particular attention in emerging markets. Where social or environmental impacts are anticipated, the developer is required to manage them through its Social and Environmental Management System consistent with Performance Standard 1.
20.1.4 The World Bank Group Environmental Health and Safety (EHS) Guidelines The EHS Guidelines (April 30, 2007) are technical reference documents with general and industry specific (i.e., mining) examples of Good International Industry Practice (GIIP). Reference to the EHS guidelines is required under Performance Standard 3.
The EHS Guidelines contain the performance levels and measures normally acceptable to the IFC and are generally considered to be achievable in new facilities at reasonable cost. When host country regulations differ from the levels and measures presented in the EHS Guidelines, projects are expected to achieve whichever standard is more stringent.
20.1.5 KCC Environmental Policy As an historic mining district, the region has a suite of environmental legacies. KCC regards environmental stewardship as an integral part of its business and will manage the impact of its activities in an environmentally and socially responsible manner as follows:
Operating with a continual awareness of the actual and potential impacts of everything on the environment.
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Preventing pollution from operations, by adopting good international practice to minimize and mitigate risks to the environment.
Complying with local laws and regulations and, where relevant, adhering to other more relevant stringent standards.
Developing, implementing and maintaining a comprehensive Environmental Management System and striving to establish a management system that mirrors the ISO 14000 standard series.
Reviewing and evaluating environmental performance.
Ensuring that adequate resources are available to meet environmental obligations and responsibilities.
Reporting environmental incidents and implementing measures to prevent reoccurrences.
Monitoring environmental impacts and regularly reporting these results.
Recognizing the importance of communicating our environmental policy internally and externally and providing the necessary education and training to employees and contractors to ensure adherence to our environmental policy.
Considering sustainable development from planning to eventual implementation of all projects.
20.1.6 Information Sources Reviewed Information sources reviewed included the following:
KCC Environmental Impact Assessments (EIA’s) and Environmental Management Plans (EMP’s) submitted to DRC regulators;
Approvals, permits and associated conditions;
Annual environmental and social monitoring reports;
Closure plans;
Environmental budgets for 2011;
Environmental staff structure;
Internal audit reports;
Due diligence; and
Community development and public relations plans.
20.1.6.1 Site Areas visited Figure 64 below illustrates the route of the site visit. The facilities visited included:
Kamoto Underground Mine (KTO) surface entrance only;
T-17 Open Pit and underground extension;
KOV Open Pit and underground extension Kamoto East Underground Mine (KTE) ;
Mupine Open Pit;
The Kamoto Concentrator (KTC) ; and
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The Luilu Metallurgical Plant Luilu Refinery.
Figure 64: Site visit route
20.1.7 Limitations of the Review The review was limited to KCC material assets and did not include any review of adjacent operations. The review was conducted by GAA Environmental Consultant, Mr. Nyundo Armitage, and focused only on key environmental and social risks as previously identified in reports provided prior to the review and identified from a site visit and interviews with KCC personnel. No detailed inspections of operations or independent environmental sampling were undertaken.
20.1.8 Results of Review 20.1.8.1 Authorisations and Operating Licences KCC have the following authorisations (inclusive of associated conditions) in place as issued by the DPEM:
Environmental Authorisation for Exploitation Permits – 4960, 4961 and 4963 issued in 2007;
Exploitation Permits 11601, 11602 and 525 issued in 2011 covering KCC material assets.
20.1.8.2 EIS and EMP KCC have an approved Environmental Impact Study (EIS) and Environmental Management Plan (EMP) compiled by DRC Green, which was largely based on the EIS and EMP developed by SRK in 2010. DRC Green, a locally registered environmental consultant, was appointed to update the EIS and EMP in 2010 in line with the DRC mining code. The EIS and EMP were approved in March 2011.
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KCC are currently implementing an Environmental Management System by the use of electronic software called Isometrix with the goal of becoming ISO14001 compliant within the next 2 years. KCC has established an Environmental Steering Committee to address environmental risks.
20.1.9 Key Environmental Risks Relating to KCC Operations Based on the review protocol and compliance criteria the following prioritization has been given to the identified environmental and social risks:
Critical Risk (risks identified that could result in project cessation and/or risk to human health);
Moderate Risk; and
Low Risk.
Only critical and moderate risks are discussed below.
Critical risks identified relate to:
Tailings slurry discharge to the Luilu River; and
Kamoto Interim Tailings Dam nearing full capacity by June / July 2012
Moderate Risks identified relate to:
Community Health and Safety risks due to communities accessing the haul roads – (Site Access along haul roads);
20.1.9.1 Critical Risk Tailings Slurry Discharge to the Luilu River Of concern, is the discharge of slurry arising from the Luilu Refinery which ultimately flows into the Luilu River. This is a legacy issue and KCC has implemented remedial measures to address this discharge.
20.1.10 Overview The Luilu Refinery currently treats copper oxide concentrate and sulphide concentrate from the Kamoto Concentrator. Concentrate treatment at the Luilu refinery includes:
Following crushing and milling at KTC – the oxide concentrate is thickened and filtered, leached through the addition of raffinate solution (low pH solution), prior to being processed through the solvent extraction and electrowinning circuit to produce copper cathodes for export;
The sulphide concentrate originating from underground operations, processed through the Kamoto Concentrator is roasted and added to the oxide leach circuit.
Cobalt is extracted from the leach circuit through cobalt hydroxide precipitation, thickening and electrowinning to produce cobalt cathodes; and
During the process aluminium and iron are precipitated.
Effluent and tailings slurry generated by the Luilu refinery (comprising low pH raffinate solution, precipitated aluminium and iron, suspended and dissolved solids, heavy metals is discharged and ultimately flows into the Luilu River via drainage channels.
Effluent discharge into the Luilu River:
Gécamines – operated the Luilu refinery and discharged this slurry into the river for approximately 20 years from the mid 1970’s to the mid 1990’s; and
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Since its operations commenced in 2008, KCC has been in regular contact with local, provincial and national authorities on developing plans to eliminate the discharge.
20.1.11 Impact of the Tailings Discharge Aquatic and surface water assessments of the Luilu River were undertaken prior to commencement of operations in 2007. Results of these assessments indicated that the Luilu River had already been severely impacted and unable to sustain forms of aquatic life. The SRK EIS in 2010 indicated that historical and continued discharges of slurry from different sources into the Luilu River were still impacting on the Luilu River water quality.
Discharge of slurry coupled with existing contamination arising from an upstream tailings storage facility owned by Gécamines, contributed to the pollution of the river in the form of surface water quality deterioration and sedimentation.
20.1.12 KCC Remediation Measures Actions taken to address effluent discharge into the Lulilu river:
Tailings Strategy: Slurry neutralization by treatment with milk of lime and operation of the lined capture facilities. Approach will ensure the cessation of the tailings discharge info to Luilu river.
Treatment System: Provisions to contain the slurry from the refinery have been identified. Following detailed design and engineering works with SNC Lavalin, construction of the treatment facilities commenced in 2011 (neutralisation tanks, pipelines, lime addition circuits, and pumps). The system was commissioned in March 2012.
Kamoto Interim Tailings Dam Nearing Full Capacity The Kamoto Interim Tailings Storage Facility (KITSF), which accepts concentrator tailings from the Kamoto concentrator, has approximately 8 months of capacity remaining.
The proposed replacement tailings facility at the unused Mupine open pit remains subject to an environmental approval process from the DPEM.
20.2 Moderate Risk The following moderate risk has been identified.
20.2.1 Community Health and Safety – Site Access Community members from the Musonoi village were observed utilising haul roads as a thoroughfare from their village to Kolwezi town. Due to the presence of heavy vehicles (such as haul trucks), accidents pose a risk to pedestrians using the haul roads.. It should be noted that alternative safe routes have been provided for the Musonoi village community to access Kolwezi.
20.2.2 General Aspects such as dust and air quality mitigation and monitoring, improved health and safety behaviour, management of waste and hazardous materials, historical soil contamination management, vibration monitoring and blasting management, and provision of adequate resources and training measures have been identified by KCC through internal auditing as additional areas requiring focused management. Review of internal audit reports indicates that measures are in place to improve performance in these identified areas.
21.0 CAPITAL AND OPERATING COSTS 21.1 Capital Cost Estimates Capital expenditure is required by KCC to develop new and existing mining areas; acquire and replace mining equipment; develop new and repair existing processing facilities; develop and repair general infrastructure, buildings, facilities, ponds and tailings facilities; and expenditure of a general nature to expand
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and maintain the operations of KCC. Table 85 shows the nature of the capital expenditure required for each operational area.
Table 85: Capital Expenditure by operational area Area Capital expenditure requirement
KTO Mine Purchase of underground mining equipment to meet LOM plans and ramp up to 2mtpa, development costs to access mining areas and ventilation infrastructure.
KOV Open Pit Waste stripping to access the ore body.
Kamoto East Underground Mine
Initially for the development from the KTO Mine through to the new mine. Thereafter, for the purchase of mining equipment required to meet LOM plans, development to access mining areas and ventilation infrastructure. Also technical work to convert resources into reserves.
T-17 Underground Mine Initially for the development of a portal from surface. Thereafter, for the purchase of mining equipment required to meet LOM plans, development to access mining areas and ventilation infrastructure.
T-17 Open pit Development work to enable the mining of T-17 extension. Civil and engineering work for the diversion of an access road and diversion and protection of the water way to the east of T-17.
Mashamba East Mine Purchase of mining equipment required to meet LOM plans.
Processing plants Kamoto Phase 4 project to increase Cu and Co output.
Effluent Ponds and Tailings Development of the ponds and tailings facilities, raise the earth containment embankment.
Power Supply Development and refurbishment of power supply infrastructure.
Environmental and Social Far West Tailings Dam stakeholder engagement, jobs and economic opportunities; tarring of roads (to reduce dust and road safety hazards); dust monitoring equipment; equipment for sulphur dioxide emission reductions/monitoring; surface water management (containment and management); general and hazardous waste management (trenches and buildings); ad hoc equipment for ground water; water settlement facilities for suspended solids radiation monitoring and survey equipment; and emergency response equipment and vehicles.
General Unallocated infrastructure of a general nature required to sustain the operations of KCC, including stay-in-business capital of USD 55 million per annum.
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Table 86 presents the expected capital expenditure annually from 2012 – 2016, the total from 2017 – 2030, and the total over the life of mine.
Table 86: Capital Expenditure by year
USD million 2012 2013 2014 2015 2016 2017-2030
LOM Total
Mining
Kamoto 52.0 30.9 23.8 11.8 18.0 73.0 209.5
KOV 10.0 5.0 20.8 16.8 0.0 0.0 52.6
Kamoto East Underground 0.0 0.0 0.0 0.0 0.0 343.2 343.2
Mashamba East 8.0 12.4 21.5 31.2 3.0 8.6 84.7
T-17 Underground 5.4 29.6 16.3 16.3 16.3 147.0 231.0
Mining subtotal 75.4 77.8 82.5 76.1 37.3 571.8 920.9
Processing
Phase 4 440.9 135.2 0.0 0.0 0.0 0.0 576.1
Additional processing 0.0 0.0 250.0 0.0 0.0 0.0 250.0
Processing subtotal 440.9 135.2 250.0 0.0 0.0 0.0 826.1
Other Cost Centres
Tailings 15.1 15.1 15.0 12.7 12.7 100.8 171.3
Social & Environmental 10.0 10.0 10.0 10.0 10.0 137.5 187.5
Power 30.0 28.0 17.6 7.8 6.5 51.6 141.5
General capital expenditure 54.4 55.0 55.0 55.0 55.0 756.3 1 029.3
Other subtotal 109.4 108.1 97.6 85.5 84.1 1 046.2 1 530.9
Total capital expenditure 625.8 321.1 430.1 161.5 121.5 1 618.0 3 277.9
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21.2 Operating Cost Estimates The operating costs for each area of KCC are shown in Table 87.
Table 87: Operating costs by area Area Description/basis Cost applied
KTO Mine Based on current and budgeted costs as an owner operation.
Weighted average cost over LOM: USD 45.57 /t ore mined.
KOV Open Pit Based on contractual rates charged by mining contractor Enterprise Generale Malta Forrest.
Costs include $7.06/bcm for mining and $1.80/t for ore haulage.
T-17 Open Pit Based on contractual rates charged by mining contractor Enterprise Generale Malta Forrest.
Costs include $7.70/bcm for mining and $2.80 for ore haulage.
Kamoto East Underground Mine
Based on estimated costs as an owner operation. Weighted average cost over LOM: USD 29.69 /t ore mined.
T-17 Underground Mine Based on estimated costs as an owner operation. Weighted average cost over LOM: USD 33.15 /t ore mined.
Mashamba East Open Pit
Based on a contractor performing the works. Weighted average cost over LOM: USD 28.70 /t ore mined.
Kamoto Concentrator Plant costs for reagents, consumables and electricity. Sulphides: USD 7.00 /t ore feed Oxides: USD 16.00 /t ore feed.
Luilu Refinery Excluding acid and lime costs. USD 0.185 /lb for finished Cu USD 0.114/lb for finished Co.
General and Administration
Head office and other centralised costs. USD 87.6 million per annum.
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The actual costs of major operating items are detailed on an annual basis in Table 88.
Table 88: Major Operational Expenditure
Operating Costs 2012 2013 2014 2015 2016 2017 - 2031
USD million
Mining 249.1 245.3 253.6 256.4 249.3 3,416.4
Luilu Refinery 134.0 122.0 164.1 178.7 186.3 2,916.7
Kamoto Concentrator 70.9 77.3 110.5 124.7 117.3 1,788.9
Total Operating Costs 454.0 444.6 528.2 559.8 552.9 8,122.0
General and Administrative Costs 87.6 87.6 87.6 87.6 87.6 1,246.1
22.0 ECONOMIC ANALYSIS 22.1 Principal Assumptions The discounted cash flow (DCF) methodology, which determines the value of an asset by calculating the net present value (NPV) of the future cash flows over the useful life of that asset, was used for the economic analysis and valuation of KCC. Because KCC is an operational mining company with several active mines, its mineral resources and ore reserves are well-defined, and a comprehensive body of technical and financial information on its current and planned operations is available. This information allows the future cash flows of KCC throughout the life of the mine to be projected, making DCF analysis an appropriate methodology.
The economic analysis is primarily based on proven and probable ore reserves, but includes mineral resources from T-17 underground and Kamoto East underground that have not been converted into ore reserves. Section 3.4 (e) of the Rules and Policies of National Instrument 43-101 states that if a disclosure includes the results of an economic analysis of mineral resources, an equally prominent statement that mineral resources that are not ore reserves do not have demonstrated economic viability must be included. The ore and recovered metal tonnages over the life of mine are shown in Table 89, showing the contribution of ore reserves (Kamoto underground, T-17 underground, T-17 Open pit, Mashamba East Open pit, KOV Open pit) and mineral resources that are not ore reserves (T-17 underground and Kamoto East underground). KCC intends to conduct further exploration of the properties including T-17 underground and Kamoto East underground.
Table 89: LOM Ore and Metals Volumes
Classification Ore mined (mt)
Cu recovered (kt)
Co recovered (kt)
Ore Reserves (Kamoto UG, T-17 OP, T-17 UG, Mashamba OP & KOV OP) 96.0 4,441.8 376.2
Mineral Resources (T-17 UG and Kamoto East UG) 29.1 1,000.9 79.7 TOTAL 125.1 5,442.7 455.9
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Other principal assumptions used in the analysis are:
Mining and processing production rates, head grades and recoveries are discussed in Section 13.0 (Mineral Processing and Metallurgical Testing), Section 16.8 (Surface operations), Section 16.6 (Life Of Mine Plan) and Section 17.0 (Plant And Processing).
The valuation date is December 31, 2011;
The discount rate is set at 10% in real terms, which is the discount rate used by Glencore across its portfolio of mining assets;
Copper and cobalt prices are discussed in Section 19.0 (Market Studies and Contracts);
Capital expenditure is discussed in Section 21.2 (Capital Cost Estimates). Capital expenditure is subject to a DRC Capital Allowance of 60% in the first year and is depreciated on a reducing balance each year thereafter;
Operating expenditure is in Section 21.2 (Operating Cost Estimates);
Royalties, tax, capital allowances and exchange rates are discussed in Section 22.4;
KML’s equity share in KCC is 75%; and
KML’s attributable economic interest in KCC is derived after the deduction of cash flows attributable to GCM from KCC’s free cash flow.
22.2 Cash flow forecasts The projected cash flow for KCC for 2012 to 2021 is shown in Table 90, and for 2022 to 2030 in Table 91. Table 91 also shows the cumulative cash flows over the life of KCC.
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Table 90: KCC cash flow for 2012 to 2021 Cash Flow Unit 2012 2013 2014 2015 2016 2017 2018 2019 2020 2021
Revenue MUSD 1,225 2,007 2,726 2,932 3,047 2,957 2,728 2,477 2,358 2,306
Freight, Insurance and Sales Costs MUSD (189) (214) (300) (332) (351) (372) (355) (353) (366) (371)
Royalties
(47) (81) (109) (84) (54) (68) (107) (96) (90) (87)
Net Revenue MUSD 989 1,713 2,316 2,515 2,642 2,518 2,266 2,028 1,902 1,848
Operating Costs MUSD (478) (467) (549) (580) (573) (587) (614) (617) (655) (668)
Other Costs MUSD (89) (89) (89) (89) (89) (89) (89) (89) (89) (89)
Net change in working capital MUSD (31) (129) 2 (17) (12) 7 2 16 (8) (3)
Total Expenses MUSD (599) (686) (636) (686) (674) (669) (701) (691) (752) (760)
Taxation MUSD (1) (2) (3) (634) (556) (590) (473) (354) (293) (264)
Capital Expenditure MUSD (626) (321) (430) (162) (121) (123) (144) (130) (143) (137)
Gécamines Dividends MUSD (16) (19) (64) (26) (16) (17) (156) (219) (181) (169)
Net Free Cash MUSD (252) 685 1,184 1,008 1,275 1,118 792 636 533 517
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Table 91: KCC cash flow for 2022 to 2030 and LOM total Cash Flow Unit 2022 2023 2024 2025 2026 2027 2028 2029 2030 LOM total
Revenue MUSD 2,318 2,318 2,315 2,318 2,315 2,286 2,252 2,180 2,155 45,219
Freight, Insurance and Sales Costs MUSD (374) (374) (374) (374) (374) (365) (355) (335) (328) (6,457)
Royalties
(87) (87) (87) (87) (87) (86) (85) (83) (82) (1,596)
Net Revenue MUSD 1,856 1,856 1,854 1,856 1,854 1,834 1,811 1,762 1,746 37,165
Operating Costs MUSD (677) (710) (693) (631) (591) (552) (522) (470) (505) (11,138)
Other Costs MUSD (89) (89) (89) (89) (89) (89) (89) (89) (89) (1,693)
Net change in working capital MUSD (4) (1) 9 29 19 20 14 25 (15) (79)
Total Expenses MUSD (770) (800) (773) (692) (661) (621) (598) (534) (609) (12,911)
Taxation MUSD (263) (285) (269) (289) (323) (332) (334) (331) (332) (5,929)
Capital Expenditure MUSD (137) (128) (120) (109) (92) (92) (100) (94) (69) (3,278)
Gécamines Dividends MUSD (169) (161) (170) (187) (192) (197) (195) (200) (180) (2,532)
Net Free Cash MUSD 517 481 523 580 586 592 584 603 555 12,516
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22.3 Valuation of KCC Based on the free cash flows in Section 22.2, the NPV of KML’s holding in KCC is USD 5,890 million. Table 92 shows the total value of KML’s holding in KCC and the contribution made by the ore reserves and mineral resources to overall value.
Table 92: Project Valuation
Total NPV of KML’s holding in KCC
(USD million)
Value contributed by ore reserves (USD million)
Value contributed by mineral resources
(USD million)
5,890 5,099 791
The T-17 Underground and Kamoto East Underground mineral resources contribute USD791 million to the valuation. These are not ore reserves, and do not have demonstrated economic viability. KCC intends to conduct further exploration and technical studies of the properties in order to convert the mineral resources to ore reserves.
Table 93 presents the NPV of KML’s holding in KCC at different discount rates.
Table 93: NPV of KML’s holding in KCC
NPV
(USD million)
Discount Rate
7.5% 6,921
10.0% 5,890
12.5% 5,078
Because KCC is an existing operation and not a start-up, the internal rate of return and payback period are not meaningful calculations.
22.4 Taxation, Royalties and Other Interests The royalties, taxes, import duties and pas de porte payments applicable to KCC are shown in Table 94.
Table 94: Royalty, tax and import duty assumptions Description Application Rate
DRC Royalty % of revenue less selling expenses 2.0%
GCM Royalty % of revenue less selling expenses 2.5%
DRC Corporate Tax % of taxable income in period 30%
Import Duty Charged on certain imported items 3% to 5%
Pas de porte 2012 – 2015 per annum USD$15.0 million
2016 USD$15.5 million
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22.5 Sensitivity Analysis The sensitivity of the value of KML’s interest in KCC to changes in metal prices, operating costs and capital expenditure is shown in Table 95, Table 96 and Table 97 respectively.
Table 95: Sensitivity of the value of KML’s interest in KCC to changes in metal prices NPV Change in metal prices
(USD million) -20% -10% 0% 10% 20%
Discount Rate
7.5% 4,450 5,701 6,921 8,151 9,401
10.0% 3,760 4,842 5,890 6,943 8,011
12.5% 3,214 4,163 5,078 5,993 6,920
Table 96: Sensitivity of the value of KML’s interest in KCC to changes in operating costs NPV Change in operating costs
(USD million) -20% -10% 0% 10% 20%
Discount Rate
7.5% 7,650 7,318 6,921 6,563 6,191
10.0% 6,513 6,230 5,890 5,583 5,264
12.5% 5,620 5,374 5,078 4,810 4,532
Table 97: Sensitivity of the value of KML’s interest in KCC to changes in capital expenditure NPV Change in capital expenditure
(USD million) -20% -10% 0% 10% 20%
Discount Rate
7.5% 7,160 7,060 6,921 6,823 6,684
10.0% 6,115 6,019 5,890 5,796 5,667
12.5% 5,291 5,198 5,078 4,987 4,866
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Figure 65: Sensitivity of the value of KML’s holding in KCC to changes in key variables
The value is most sensitive to metal prices, with a positive 1% change in metal prices causing a positive change of USD 120 million in project value. KCC is less sensitive to changes in operating costs and capital expenditure, with a positive 1% change causing negative changes of USD 31 million and USD 16 million respectively.
22.6 Risk Analysis The risk analysis in this ITR comes from the following sources:
The Competent Persons, as defined in the JORC Code, involved in the technical analysis of the Material Assets were interviewed to identify and record project risks; and
The risks identified by SRK in the 2009 Technical Report were reviewed and included in this analysis if they were determined to still be valid.
22.7 Mining Risks The major risks that could have a negative impact on the planned production profile are:
Dewatering: Initial and continuous dewatering of the KOV Open Pit and Mashamba East Open Pit is required. The current dewatering strategy has recently been revised to accommodate the pit designs referenced in the 2010 Technical Report and historical hydrogeological data which is difficult to validate. As such, the current dewatering strategy could potentially prove to be less effective at maintaining dry slope conditions than anticipated. This potentially could introduce risks associated with potential impacts on both the economics of the pit, pit stability and the production schedule. This risk is being mitigated by performing further studies to gain a better understanding of the hydrogeology so that an appropriate dewatering strategy is defined that enables safe mining as cost effectively and efficiently as possible;
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Access and slope failures: KOV Open Pit will develop into a large operation up to 400m deep. Production rates are high with up to 4.5Mt of material that should be moved from the KOV Open Pit per annum. Small slope failures could have a negative impact on access system to the production benches. A strategy should be developed to establish an alternative ramp access to production areas;
Available pit space: A high production rate requires sufficient working areas or face length. Additional face length should be established and maintained on the southern section of the KOV Open Pit;
Available waste dumping space: The KOV Open Pit will produce 46.2Mt of waste up to 2030. The available waste dumping space close to the KOV Open Pit is insufficient for the LOM although as part of the AJVA additional surface rights have been allocated to KCC to the north of the current KOV Open Pit concession. However, for improved cost efficiency, a detailed technical plan to back fill mining waste into depleted pits should be developed. With the appropriate mining sequence in KOV Open Pit, a total of 15.0Mt could be back filled while the T-17 Open Pit is available for back filling from 2013 onwards should the appropriate technical studies be completed;
Grade control: An operational cut-off grade of 0.6% Cu has been applied to the mining models. Grade control practices for a high volume operation should be developed. Inefficient grade control systems may result in mineral resource losses or cause uneconomic production tonnes to be processed in situations where the planned revenues does not cover the processing and selling costs;
Waste Rock: Site personnel interviewed, indicated radio-active waste rock occurring within the T-17 Open Pit is handled and stored separately from non radio-active waste rock. KCC has capped this radio-active waste rock dump with non radio-active waste rock in compliance with international accepted standards; and
Collapsed Zone: The presence of a collapsed zone in the central plateau portion of the KTO ore body may influence stress distribution in adjacent areas.
22.8 Processing risks Unavailability and Quality of Key Reagents for Metallurgical Processing Potential risk that critical process reagents may not be available in the required quantities or quality, leading to reduced production of copper and cobalt. Availability of re-agents is less of a problem than previously as there have been improvements in supply chain management and inventory control.
Power Availability and Supply Fluctuations Power requirements to operate at the scheduled production profile are approximately 230-250MW and there are risks that this power may not be available through the national grid and may lead to power disruptions or supply fluctuation. KCC has entered into an agreement with the state utility, Société Nationale d’Electricité (SNEL), to refurbish the DC link between Kinshasa and Kolwezi SCK/RO stations to increase power availability to KCC to a minimum 160MW in 2011. KCC is also in advanced negotiation with SNEL to provide capital to refurbish additional power infrastructure within the DRC to increase the available power supply from the Inga hydroelectricity facility to 450MW to the Katanga Province by 2015.
The work with SNEL is progressing and power supply constraints seem less of risk; however the new electro winning plant will require as much as 60-75MW and variable supply to the electro-winning plant would be a major production issue.
22.9 Capital risks 22.9.1 Escalation of Costs Projects in the mining industry world-wide have recently experienced unpredictable capital cost overruns due to various macroeconomic and microeconomic factors that cannot be predicted with any reliable degree of certainty. Capital cost overruns require more funding and reduce project returns. This risk is rated as high but is being mitigated by management through regular reviews of capital cost estimates by the KCC project team
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and their appointed independent engineers, who provide certified project control software and an extensive an up-to-date database of capital costs for many aspects of the development.
22.10 Operating Risks 22.10.1 Poor Condition of Railways The poor condition of the railway line may impede efficient production by not allowing the efficient, on-time delivery of finished products or the supply of key input materials on time, leading to reduced production of copper and cobalt and higher logistics costs. This risk is rated very high. Possible mitigation measures include:
Rescheduling production plans to match rail capacity;
Engaging with governments and railway operators;
Engaging with other potential rail users; and
Resorting to road transportation (at a higher cost) for logistics, although road costs have been factored into the financial model.
22.10.2 Underdeveloped In-Country Institutional Infrastructure and Capacity The DRC’s national and local governments and their agencies may not have the ability to deliver on the infrastructure requirements of the project, reducing the project feasibility or causing delays. This risk is rated high, and may be mitigated by developing relationships with other stakeholders, governments and agencies; and supporting capacity development initiatives.
22.10.3 Senior Management and Technical Expertise Recruiting and retaining senior management and critical technical expertise to manage and operate the mines and processing plants is an issue, rated as a high risk. It potentially affects the ability of the project to run optimally and comply with legislation. Mitigation measures include reviewing KML’s and KCC’s employment strategy, recruitment and retention plan; and facilitating the provision of contractor's services with Government and other service providers. KML has appointed a significant proportion of expatriate employees amongst its management level. There are currently no critical vacancies.
22.10.4 Artisanal Miners There are a large number of artisanal miners working on the adjacent tailings dam and waste rock dumps. Artisanal mining activities are mostly on adjacent mining concessions that are not under the control of KCC including parts of Mupine pit which belongs partially to Gécamines.
22.11 Sovereign risk There was an increase in sovereign risk given the allegation of electoral irregularities, following the November 2011 elections and post election violence in December 2011. However, such incidents of unrest, have historically taken place in parts of the country far away from KML/KCC's operations. The situation has returned to normal and the Kabila government remains in power with no major changes expected in mining legislation or rules applicable to private ownership of mineral rights.
The DRC (as a whole) continues to be at risk of being affected by varying degrees of political and economic instability in the future, which is outside of KML/KCC's control.
Changes in the DRC legal system may expose KML/KCC's operations in this region to increased operational risks and/or compliance costs.
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22.12 Economic and Market Risks 22.12.1 Commodity Prices: Copper and cobalt market prices are significant drivers of the profitability for KCC and the value of KML’s interest in KCC. These prices are subject to wide fluctuations beyond the control of KCC or KML due to factors such as demand for the commodities caused by global economic conditions and prospects, supply from various sources, currency and interest rate changes, and speculative activities. Sustained commodity prices below the costs of production may cause the curtailment or suspension of operations. There is some scope to manage market risk through hedging, but this may lead to loss of upside during periods of high commodity prices.
22.12.2 Operating Costs: Project operating costs also affect the profitability of KML and the value of the KCC project. These are subject to a wide range of pressures such as energy prices, oil prices, chemical prices, labour costs and inflation.
22.12.3 Currency Risk: Project revenues are in United States dollars; however, input costs may be in other currencies, specifically the South African Rand. Variations in currency exchange rates can affect production costs and affect project profitability.
22.13 Environmental and Social Risks These risks were identified during an environmental and social audit conducted by GAA at KCC on 23-27 January 2012.
22.13.1 Tailings Slurry Discharge to the Luilu River: Aquatic and surface water assessment of the Luilu River was undertaken prior to commencing operations in 2007. Results of these assessments indicated that the Luilu River had already been severely impacted and unable to sustain forms of aquatic life. The SRK EIS in 2010 indicated that historical and continued discharges of slurry from different sources into the Luilu River were still impacting on the Luilu River water quality.
Discharge of slurry coupled with existing contamination arising from an upstream tailings storage facility owned by Gécamines, contributed to the pollution of the river in the form of surface water quality deterioration and sedimentation.
22.13.2 Community Health and Safety – Site Access: Community members from the Musonoi village were observed utilising haul roads as a thoroughfare from the village to Kolwezi town. Due to the presence of heavy vehicles (such as haul trucks), accidents pose a risk to pedestrians using haul road.. Alternative access points have been provided as discussed in environmental section.
22.13.3 Storm Water Management: It was observed that there is a lack of sufficient storm water drainage channels in and around the mine site, resulting in significant erosion of haul roads and tracks as well as ponding of water at low lying areas. Storm water management plans have been discussed in environmental section above.
22.13.4 Other Risks: Other risks include:
Tailings capacity;
Dewatering of the Mupine Pit.
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23.0 ADJACENT PROPERTIES In respect of adjacent concessions KCC has entered into various joint venture, lease agreements or servitudes, allowing KCC to access these properties and/or to locate relevant infrastructure on these sites. Specific issues concerning adjacent properties owned by GCM has already been described in paragraph 4.2 above.
24.0 OTHER RELEVANT DATA AND INFORMATION 25.0 TAILINGS AND WASTE 25.1 Location Description The Kamoto Interim Tailings Storage Facility (KITSF) is the only operational tailings disposal facility at Kamoto Copper Mine (KCC). The TSF was commissioned in 2008 for the safe disposal of copper and cobalt tailings from the KTC plant, which is located 500m northwest of the TSF.
25.2 Description The TSF is a hill-side impoundment with an earth starter embankment located on the western, southern and eastern perimeter. The TSF currently has a top surface area of 71.9ha and its tailings storage capacity approximately is 1.3 million m³.
The tailings delivery pipeline is a 350 mm diameter rubber lined, steel pipeline spigot which is positioned on top of the earth starter wall and extends along the western, southern and eastern flanks of the TSF, with 1 valve positioned in the north-west corner. A number of spigot outlets are set into this delivery pipe.
A spigot deposition system is used to distribute tailings around the TSF.
The TSF has 4 penstock intake structures i.e. the main penstock structure is located in the northern portion of the TSF and 3 intermediate intakes which are located on the same outlet pipe to the south and at progressively lower elevations. The supernatant water flows out through the penstock pipeline to the outfall trench, located south west of the western embankment, which discharges into the return water dam, situated further west.
A toe drain constructed along the inner toe of the main embankment also discharges into the outfall trench at the same location as the penstock outlet. The return water dam is equipped with a pumping and piping arrangement to pump water back to the KTC plant for reuse.
The design life of the TSF was initially envisaged to be 11 months in its current form, but due to an increase in tailings production, a revised design increased the life of the TSF for a further 6 to 8 months during 2012.
KCC is currently waiting for the DRC’s Department for the Protection of the Mining Environment (DPEM) to grant authorisation to allow the deposition of tailings into the Mupine Pit. Until the authorisation is granted, the TSF will need to be utilised. There will be a need to raise the earth containment embankment of the TSF by 1 - 5m to create sufficient deposition space, before the pit becomes available in September 2012.
26.0 CLOSURE 26.1 Approach and Limitations to Closure Cost Update 26.1.1 Approach An indicative closure cost estimate based on information available at the time was conducted by GAA for December 2010 (hereafter referred to as “indicative closure costs”). This indicative closure cost was subsequently updated for December 2011 (hereafter referred to as “updated closure costs”) using current unit rates and incorporating new information. The updated closure cost was conducted for unscheduled and scheduled closure, taking into consideration a number of aspects that were not considered in the indicative closure cost. This information was not available when the indicative closure cost was compiled. These aspects are detailed in Sections 26.1.4 and 26.1.5.2.
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The approach followed with the updated closure costs was largely similar to that for the indicative closure costs as follows:
Verification and updating of the relevant mining areas and associated infrastructure, primarily from Google Earth imagery and limited available plans;
Identification of infrastructure and land use sub-categories within the above mining operations area characterised by similar conditions, for example light, medium or heavy infrastructural areas, waste rock and spoils stockpiles, and moderately or severely disturbed surface conditions;
Interpretation of the type, nature and sizes of structures from available information and measurement of the delineated areas in AutoCAD;
Review of the closure approach and criteria adopted for the indicative closure cost by GAA and verifying and updating of the respective measures;
Updating and verification of unit rates used for the indicative closure cost for plant dismantling and demolition, as well as associated rehabilitation. Benchmarking of the new unit rates was done as per recent tenders available to GAA, similar work conducted recently in Africa, as well as consultation with demolition and earthworks practitioners;
Application of the above unit rates and associated quantities in spreadsheets arranged into sub-categories to illuminate the respective closure cost components for the cost update; and
Compilation of a report reflecting the approach applied by GAA in determining the updated closure costs. Matters requiring attention to ensure that future closure costing is improved and more realistic are also listed.
26.1.2 Limitations
The updated closure cost was conducted primarily as a desktop assessment, based on limited additional information and photographs of the mine area. As a result a number of aspects must either still be verified or further refined to improve the accuracy of the closure costs. Hence, this updated closure costs must also regarded as indicative only and as such provides the basis for future closure cost updates;
The updated closure cost has not been compiled for the purposes of purchase of the mine; and serves only to inform internal financial and accounting requirements;
A one-day site visit to the mine site in support of the overall project, also addressing closure cost aspects, was conducted. This time on site was not sufficient to gain a full understanding of the closure related site aspects;
The updated closure cost estimation was conducted from both an “end of life-of-mine” (scheduled) scenario, as well as an immediate (unscheduled) scenario. Scheduled closure incorporates a number of planned or recently initiated activities, such as the construction of the new electro-winning and solvent extraction plant infrastructure, planned expanded open pit mining and tailings deposition in the Mupine open pit. Unscheduled closure essentially considers closing the mine in its current state,
The costing assumes that closure of the operations occurs with no remedial work having been done over the remaining operational life, If material work will be conducted this would influence the closure costs;
The cost assessment does not include any operating or capital costs that are likely to be incurred due to management of the environment during operations;
Although the costs the metallurgical plants have been included, the Joint Venture agreements with Gécamines indicate that once KCC has finished utilising them, they will revert back to Gécamines.
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These facilities are currently leased from GCM on terms defined in the Amended and Restarted Joint Venture Agreement for a period to 2024, with two 10 year renewable options;
The cost assessment does not include infrastructure that falls within KCC concession area, but which has not been operated by KCC. Examples include but are not limited to: Poto Poto Tailings dam, Kamoto Tailings dam and any waste dumps from KOV, various infrastructures recently been constructed due to GCM related activities within the KCC concessions.; and
The assessment does not include infrastructure previously operated by KCC, but which has now been transferred out of the concession area, such as the Kolwezi Concentrator.
26.1.3 Available Information The sources of information used for the closure cost estimate were as follows:
Map of the Tilwezembe Concession PE 4963 DCP. Infrastructures on perimeters of the Convention JVACR. July 2009. KCC – Kolwezi DRC;
Map of the KCC Infrastructures dated June 01, 2011 (datum WGS84_35S);
The December 2010 Review of Unscheduled Closure Costs (Decommissioning and Restoration Costs) for KCC report dated January 2011, compiled by GAA;
Emailed correspondence from Robin Gardiner of SNC Lavalin to GAA , dated January 31, 2012; and
Photographs of various infrastructure elements and mine features, taken during the site visit conducted by GAA.
26.1.4 Battery Limits The battery limits for this closure cost update were based on those established for the 2010 closure cost review and indicative closure; and have been updated to include a number of items listed below. The infrastructure description obtained from the available information is listed separately from those aspects inferred from the imagery and/or added by GAA.
26.1.4.1 Available Information The following battery limits and their respective footprint areas were previously deduced from the available maps and measured using Google Earth imagery. Please note that these items have been used as per the indicative closure cost, as insufficient information was available to verify or refine any of these items:
Overall mine site and plant complexes, including fugitive disturbed areas (but excluding areas defined as GCM’s liability): 2323 ha;
Luilu plant and related areas: 51.6ha;
Kamoto concentrator and related areas: 36.4ha;
SKM and related areas: 5.4ha;
Kamoto underground and related areas: 25ha;
KOV pit and related areas: 50ha;
T-17 pit: and related areas: 35.7 ha; and
Riverine areas requiring rehabilitation and reinstatement: 4ha.
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A number of tailings storage facilities evident on aerial imagery are understood to be located outside of the overall battery limit of this closure cost update. Hence GAA has excluded the following components from this cost update:
Kamoto TSF;
Poto Poto TSF;
Kingamyambe TSF; and
Sulphide TSF.
The unnamed TSF to the south of the Luilu plant was also included in this cost update. It is noted that the full rehabilitation costs for the latter has not been included, as this TSF is pre-existing.
Furthermore the following planned future or recently initiated components / activities were included in the updated closure cost, as described below:
The planned expanded open pit mining of the Mashamba East Pit;
The planned expanded open pit mining of the KTO East Pit;
The planned expanded open pit mining of the KOV Pit;
The planned deposition of tailings material in the Mupine Pit; and
The construction of the new electro-winning and solvent extraction plant and associated infrastructure.
In addition, the rehabilitation of the planned KOV stockpiles footprint indicated on the 2011 KCC infrastructures drawing were also included.
26.1.4.2 General Additions by GAA The following general closure costing components and related activities were also considered by GAA in the determination of their indicative closure cost update:
Collection, handling and disposal of demolitions waste;
The rehabilitation and reinstatement of affected streams and drainage lines;
Rehabilitation of disturbed areas, including the collection, handling and disposal of contaminated soil as well as the removal and disposal of fugitive concrete; and
Additional allowances, including preliminary and general (P&Gs) and contingencies.
It is understood that the mine will pump water from the existing open pits that are to be mined further. Once mining of the various pits has ceased it is expected that the pits will eventually fill in with precipitation and groundwater ingress. It is expected that ongoing management of contaminated excess mine water arising from the reclaimed mine workings may be required; involving collection, handling, treatment and safe disposal of the treated mine water. However the need, nature and extent of this aspect are unknown and hence have been omitted from the closure cost estimate. If required, this could add a notable additional capital expenditure and ongoing operational costs.
The battery areas indicated in Figure 66 were assumed and considered. This includes the PE525 and PE4961 concession areas. An area described as Area No.2 includes the Luilu plant, and PE11602 which includes the T-17 open pit and surrounding infrastructure. The battery limits also includes the TSF to the south of the Luilu plant as is described above.
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Figure 66: Map of Infrastructure on perimeters of the convention AJVA - KCC
26.1.5 Assumptions and Qualifications The assumptions and qualifications listed below have been made with respect to the closure cost update.
26.1.5.1 General
The closure costs for the plant site could comprise a number of cost components. This report only addresses the decommissioning and reclamation/restoration costs, equating to an outside (third party) contractor establishing on-site and conducting the decommissioning and reclamation-related work. Other components such as staffing of the plant site following decommissioning, the infrastructure and support services (e.g. power supply.) for the staff, as well as workforce matters such as separation packages and re-training/re-skilling, are outside the scope of this report;
Based on the above, dedicated contractors would be commissioned to conduct the demolition and work on the mining site and associated areas. This would inter alia require establishment costs for the
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demolition and rehabilitation contractors and hence, the allowance of preliminary and general (P&Gs) in the cost estimate. Allowance has also been made for third party contractors and consultants to conduct post closure care and maintenance work, as well as compliance monitoring;
Steel and related material from the plant demolition which has salvage value will remain on-site for sale to third parties with any salvage value from demolition waste material for GCM’s benefit;
As there would be no salvage value to KCC, no cost off-sets due to salvage values were considered in terms of accepted practice and thus only gross closure costs are reported;
Concrete footings and bases would be demolished to a maximum of 1 000 mm below the final surface topography;
All useable stockpiles of raw and/or saleable material would have been processed and removed off-site at closure and none of these would remain on site, thus requiring reclamation; and
The existing villages would not be demolished, but would be transferred to third parties. This also applies to the services related to the village such as water supply and sewage treatment.
26.1.5.2 Site Specific It has been assumed that at mine closure the mine site and associated disturbed areas will be reclaimed to a sustainable predetermined final land use. This will not only require the dismantling of the physical infrastructure and addressing the aesthetic effects of the reclaimed mine site, but also addressing the residual impacts of the operations on the receiving environment. Therefore, the closure cost update addresses, as far as reasonable, the possible latent and residual effects. In this regard the following site-specific closure measures have also been included in the cost estimate:
The dismantling and rehabilitation of the expanded electro-winning and copper solvent extraction plant infrastructure, construction of which has recently been initiated, has been included in this closure cost update, as follows:
The new solvent extraction plant footprint area is approximately 13.4 ha in extent, of which 50% has been assumed to be heavy infrastructure and the remainder medium infrastructure;
The electro-winning plant footprint is approximately 1.5 ha in extent and is assumed to consist of medium infrastructure, with an extra over allowance for removal of concrete not exceeding 250mm in thickness;
The footprint of the new substation to be constructed for the electro-winning plant will be approximately 0.5 ha; and
The electro-winning plant storage and export yard will be approximately 4 ha.
It is foreseen that demolition waste, such as concrete and building rubble, would be largely inert and that a dedicated waste disposal facility will be licensed and constructed for the purpose of disposal of demolition waste. Provision for establishing a 2.5 ha waste disposal facility within the mining rights area for the disposal of all demolition waste, as it has been assumed that no suitable facility is available within a close enough distance. Provision is made for the facility once decommissioned, to be covered with material sourced from site and for the establishment of vegetation directly onto the cover material;
Covering of the relevant TSF’s with a soil cover will not be done, as it is understood that vegetation establishes naturally on the tailings material and that covering of the TSFs is not normally done in this region. Hence, allowance for the establishment of grassland vegetation only has been made. However it is anticipated that this approach will not prevent ingress of rainfall water and hence will not prevent contaminated net percolation (waste load), and must be verified and updated as required to conform to best practice requirements;
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The rehabilitation of evaporation/pollution control ponds will include the following:
Removal of sediment up to a depth of 400 mm;
Removal of synthetic liner;
Removal of contaminated soil that could have occurred in those places where the liner has leaked;
Collection, transport and disposal of the contaminated sediment and soil; and
Establishment of a suitable woodland vegetation cover.
The remaining waste rock and/or over burden dumps will be shaped and vegetated;
Rehabilitation of the planned KOV stockpiles for scheduled closure only, of which the footprint area has been assumed to be approximately 13.4 ha in extent, by removing contaminated material to a depth of 300 to 500 mm, shaping the footprint area and re-establishment of woodland vegetation. Note that placement of a growth medium was not provided for as it is understood that acceptable levels of natural re-vegetation of disturbed areas occurs in time;
Different shaping, levelling and re-vegetation methods will apply for disturbed areas based on the nature, extent and severity of disturbance. The following categories have been assumed:
Moderately to severely disturbed general areas, over which suitable woodland vegetation will be established;
Waste rock dumps, overburden stockpiles and other similar areas onto which grassland vegetation will be established; and
Tailings material, over which grassland material will be established.
Dedicated rates for the shaping, levelling and rehabilitation have been applied for the above categories.
The Kamoto underground workings are assumed to be accessed by one incline shaft with associated portal;
In addition to the above underground mining, the rest of the mining is and/or will be conducted from four open pits on surface;
The final mining voids or remaining open pits will not be in-filled and allowed to become open lakes over time with the required access control whilst these are re-watering (flooding). In order to limit access, an open rock enviro-bund to a height of at least 3 meters and its inside toe 20 m from the long term break-back line of the pit/void will be constructed . The bund will also serve the following purposes:
Safety measure to isolate the pit from people and animals by restricting access to the pit and voids;
Visual screening; and
Divert surface water runoff away from and around the pit, preventing erosion of pit or void lip/edge.
As sufficient vector information on the extent to which the various open pits will be expanded was not available it was assumed that for scheduled closure the enviro-bund will double in length, although it was assumed that only 20% of this cost would be accounted for during closure and that the rest of the cost associated with creating the bunds will be operational;
It was assumed that the entire volume of material that will be required to create the bunds for scheduled closure can be sourced from the pits during operations, however this must be confirmed;
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It is understood that the Mupine pit will be used for the future ongoing disposal of tailings material. It is assumed that the pit has sufficient capacity to contain all tailings material emanating from future mining and no provision for the expansion, deepening or any further preparation of the pit has been made in this regard. This aspect must be verified in future closure cost updates as the potential requirement to expand the pit and/or create additional tailings disposal facilities will have significant cost implications. Allowance for the following has been made:
For scheduled closure, establishment of grassland vegetation directly on the tailings material once deposition has ceased; and
For unscheduled closure, establishment of an enviro-bund as described above around the extent of the current pit, as it was assumed that no further tailings material would require deposition.
Load and haul of waste rock material from site over a maximum distance of 2 km for infilling of the shaft portal and 3 km for creating the various enviro-bunds respectively, has been assumed;
Removal of contaminated soil from disturbed areas as part of general surface rehabilitation is required for approximately 20 percent of the reclaimed infrastructural footprint areas;
Allowance has been made for a nominal amount of fugitive concrete to be removed and disposed of; and
Allowance has been made for care and maintenance as well as surface and groundwater quality monitoring to be conducted for a minimum period of 5 years to ensure and assess success of the implemented rehabilitation and closure measures. However no allowance for ongoing collection, handling, treatment and safe disposal of treated mine water following closure has been made and the possible need and extent of this requirement must be determined and incorporated in future closure cost updates.
26.1.5.3 Additional Allowances Fixed ratios for P&Gs (6%) and contingencies (10%) have been applied.
26.1.6 Unit Rates Please note that all rates are expressed in US dollars.
26.1.6.1 General Surface Shaping It has been assumed that general surface shaping would be required over most of the areas where surface infrastructure has been removed as part of the overall surface rehabilitation. This includes the stockpiling of building/demolition rubble to be removed for disposal, as well as the subsequent shaping and profiling of these surfaces. It has been assumed that shaping and profiling would involve the dozing of material at approximately 500 to 750 mm depth, the rate for which that was used being $8 615.00/ha.
It was assumed that all stockpiles, waste rock, overburden dumps and other relatively loose material would be at an angle of repose which has been estimated at 36o and that the average height of the dumps is approximately 20 m. The dumps will be dozed and reworked to 1:5 slopes for which a rate of $1.55/m3 was used.
26.1.6.2 Rehabilitation of Evaporation Ponds It was assumed that all evaporation ponds are lined. The rehabilitation of the sediment ponds was assumed to include the removal of contaminated sediment to a depth of approximately 400mm; removal of a synthetic liner; contaminated soil removal under the liner as a result of potential leaks to a depth of 250mm; and the load and haul of contaminated sediment. The total rate for the rehabilitation of evaporation ponds was therefore estimated at approximately $6.50/m².
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26.1.6.3 Establishing of Vegetation (General) Woodland vegetation is being established on the majority of disturbed areas including areas from which plant and surface infrastructure is removed, generally disturbed areas. If vegetation has to be established on un-compacted growth medium/topsoil, soil amelioration will most likely be required. This will depend on the nature of the soil, whether the topsoil was stockpiled and the period of stockpiling. In order to determine a unit rate for re-vegetation, allowance has been made to apply 0.5 ton/ha fertiliser, 5 ton/ha lime and 15 ton/ha organic material such as well-cured cattle manure. If cultivation and seeding are also included, but ripping to alleviate compaction excluded, this rate equates to approximately $2 500.00/ha.
26.1.6.4 Vegetation Establishment on Tailings It has been assumed that no topsoil/growth medium is required for tailing storage facility rehabilitation after shaping to the required outer slopes and upper surface has been completed. Grassland vegetation would thus be established on the modified tailings. The rate includes the amelioration of the tailings by applying fertiliser and 80 ton/ha of organic material such as well-cured cattle manure. The rate also includes the cultivation actions to work the ameliorants into the tailings and the seeding of the ameliorated tailings. With outer slopes at gradients not steeper than 11o (1:5), a rate of $3 500.00/ha was adopted. This excludes the shaping of the outer slopes and upper surfaces of the facilities.
26.1.6.5 Vegetation Establishment on Waste Rock Dumps It has been assumed that vegetation will be established directly on the ameliorated waste rock after shaping. This rate includes the amelioration of the cover layer by applying fertiliser and 40 ton/ha of organic material such as well-cured cattle manure. With shaping it will be ensured that in those places where the shaped surface would be too rocky for grassland vegetation establishment, that finer waste rock material would be exposed on the surface to facilitate vegetation establishment. This rate also includes the cultivation actions to work the ameliorants into the waste rock and the seeding of the ameliorated waste rock. With outer slopes at gradients no steeper than 11o (1:5), a rate of $2 900.00/ha was adopted. This excludes the shaping of the outer slopes and upper surfaces of the waste rock dumps.
26.1.6.6 Creation of Enviro-Bund A number of final voids will remain once mining has been completed and it has been assumed that for economic reasons, these will remain open pit lakes after closure. The rock enviro-bunds will be built to a height of at least 3 meters and it’s inside toe 20 m from the long term break-back line of the pit/void. It has been assumed that the enviro-bunds will be created from material sourced from site with load and haul over a distance of 2 to 3 km, and placing and shaping of the material. The rate applied for the construction of the enviro-bunds as described above is $5.03/m3.
26.1.6.7 Groundwater Monitoring It has been assumed that groundwater quality monitoring has to continue at least four times a year for at least five years post-closure. Allowance has therefore been made to conduct the above monitoring at sixteen monitoring points (boreholes), consisting of two monitoring points at the four active pits, two monitoring points each at the interim tailings dam and Mupine pit tailings dam, two points at the shaft complex and two points at the plant complex.
If assumed that it would take at least five man-days of an independent specialist to conduct the sampling at these points at an hourly rate of $80.00/hour, this would equate to about $3 300.00 per sampling event for professional fees and associated disbursements. If an additional allowance is made for sample analysis of $225.00 per sample, this equates to an additional amount of $3 600.00, totalling to $6 900.00 per event. Taking other disbursements (approximately 15 %) into account this amount could be rounded to $7 950.00 per sampling event, or $31 800.00 per year for the mine.
A dedicated hydro-geological assessment must however be conducted to determine the actual post-closure groundwater monitoring requirements and incorporated into future closure cost updates.
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26.1.6.8 Surface Water Monitoring As for groundwater monitoring, it has similarly been assumed that surface water monitoring will have to continue for at least five years post-closure. For the purposes of this costing it has therefore been assumed that sampling has to be conducted at six monitoring points, namely four operational pits, downstream of the KTO interim tailings dam, and downstream of the main plant area.
Similar professional and analytical fees, requiring a minimum of two days on site per sampling event, as well as disbursements (costs for borehole pumping costs substituted by boat renting) would apply. It is also foreseen that surface water monitoring would be conducted monthly. Annual surface water sampling costs are therefore estimated to amount to approximately $32 400.00.
However, as for groundwater monitoring, the actual surface water monitoring requirements must be determined based on a dedicated study and incorporated into future closure cost determinations.
26.1.6.9 Rehabilitation Monitoring Based on the extent of existing and possible future mining, corrective action on the rehabilitated areas may be required and as such post-closure monitoring of rehabilitated areas will be required. However the intensity of monitoring will vary, depending on the nature of the activity or infrastructure that had previously occurred in an area. Allowance was therefore made for high intensity as well as low intensity monitoring to reflect site specific conditions, as discussed below.
26.1.6.9.1 High Intensity Rehabilitation Monitoring The high intensity rehabilitation areas are deemed to include all areas from which plant and buildings have been removed, as well as all tailings dams, where vegetation is established directly on the tailings material. As a result of the more challenging conditions that are expected in these areas it is likely that rehabilitation monitoring will also be required for a longer period, i.e. five years. It has been assumed that fifteen man-days would be required to conduct rehabilitation monitoring over the approximately 279 ha area. Assuming a consultant rate of $80.00/hr this would equate to $9 600.00 per event. If it is assumed that this has to be conducted twice a year, the annual costs would amount to $19 200.00 or roughly $69.00/ha. If an additional $10/ha is added for travelling and accommodation the overall rate is $79.00/ha/year; or $395.00/ha over the required 5 year period.
26.1.6.9.2 Low Intensity Rehabilitation Monitoring It has been assumed that the remaining surface rehabilitation will, due to the high rainfall and rate of natural re-vegetation, constitute low intensity rehabilitation. Subsequently, rehabilitation monitoring will also only be conducted for three years following closure in these areas. It was assumed that it would require approximately twenty days to monitor the remaining 2 044 ha low maintenance rehabilitation areas. Applying the same approach as above, each monitoring event would cost approximately $12 800.00. If conducted twice a year this equates to approximately $12.50/ha. If an additional $10/hour is added for travelling and accommodation, the overall rate is $22.50/ha/year, or $67.50/ha over the required 3 year period.
26.1.6.10 Rehabilitation Care and Maintenance As mentioned, permanent care and maintenance could be required to ensure sufficient vegetation establishment, repair erosion damage and avoid ponding in subsided areas. As with monitoring the intensity of care and maintenance will vary, depending on the nature of the rehabilitated area. Different rates for the aftercare and maintenance of high and low intensity rehabilitation areas have therefore also been applied as discussed below.
26.1.6.10.1 High Intensity Care and Maintenance It is assumed that this would require 15 weeks per year of a team of 20 workers and two JCBs as supporting equipment to conduct the corrective measures over 279 ha. It has been assumed that the hourly rate of the workers is $3.50 and the equipment $525.00/day (per machine). If accommodation and travelling of $10.00/ha is also added, the overall rate is about $445.00/ha/year, or $2,225.00/ha over the required 5 year period.
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It has been assumed that the workers and equipment could be sourced locally.
26.1.6.10.2 Low Intensity Care and Maintenance Low intensity care and maintenance would be required over generally disturbed (extensive) areas. It is assumed that this would require twenty weeks per year of a team of 20 workers and two JCB as supporting equipment to conduct the corrective measures over the remaining 2 044ha. It has been assumed that the hourly rate of the workers is $3.50 and the equipment $525.00/day (per machine). If accommodation and travelling of $10.00/ha is also added, the overall rate is about $110.00/ha/year, or $330.00/ha over the required 3 year period.
Once again, it has been assumed that the workers and equipment could be sourced locally.
26.1.7 Closure Cost Assessment An assessment of the updated closure cost is presented in Table 98 in a format routinely used for closure cost determinations, addressing the following categories:
Infrastructural areas;
Mining areas;
General surface reclamation;
Water management;
Post closure aspects; and
Additional allowances.
The updated closure cost determined for scheduled and unscheduled closure is reflected in Table 99.
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Table 98: Closure cost assessment
Closure component Closure cost assessment
Unscheduled closure (December 2011) Scheduled closure (Undetermined) 26.1.7.1 Infrastructural Areas
Processing plants, steel structures, reinforced concrete structures, offices, workshops, pump stations, residential buildings and related structures and infrastructure
This area includes the processing plant and related structures assumed to consist of both heavy and medium infrastructure for Luilu Plant to a total of 45.9 ha; and Kamoto Concentrator to a total of 13 ha, at an assumed plant coverage of 40% of the total area;
All processing plant and related structures assumed to consist of light infrastructure for Kamoto Underground, totalling 250 ha, at coverage of 40% of the total area;
Additional steel buildings and similar structures at the Luilu Plant totalling 3.9 ha, at a coverage of 40%;
All processing plant and related light structures for all other infrastructural areas totalling 20.8 ha, at a coverage of 40%;
Storage and warehousing structures at the Kamoto Concentrator to a total of 5.9 ha, at a coverage of 40%;
All substations and associated electrical installations within the lease area;
It was assumed that inert demolition waste will be disposed of at a newly constructed landfill site that is positioned within the mine lease area. Allowance was made for an average haul distance of 3 km for the disposal of the waste. It was assumed a 900mm thick compacted cover of in-situ material would cap the site with allowance for compaction of the lower 600mm and establishment of grassland cover; and
General surface rehabilitation will be implemented on footprint areas.
As for unscheduled closure, but with the additional inclusion of the following:
The new solvent extraction plant footprint area of approximately 26 ha in extent, of which 50% has been assumed to be heavy infrastructure and the remainder medium infrastructure at an assumed plant coverage of 40% of the total area;
The electro-winning plant footprint is approximately 1.5 ha in extent and is assumed to consist of medium infrastructure, with an extra over allowance for removal of concrete not exceeding 250mm in thickness;
The footprint of the new substation to be constructed for the electro-winning plant will be approximately 0.5 hectares; and
The solvent extraction plant storage and export yard will be approximately 4 ha, of which approximately 40% will be covered with steel structures 12 m high.
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Closure component Closure cost assessment
Unscheduled closure (December 2011) Scheduled closure (Undetermined)
Roads, power lines, railways and pipelines.
Roads within the mine lease area are limited and the cost associated with their removal is included in infrastructural areas above; and
Similarly, the removal of power lines, railways and pipelines is included in the infrastructural areas above.
As for unscheduled closure.
26.1.7.2 Mining Areas
Open cast mining areas, including final voids and ramps
Rock enviro-bunds will be built around all open pits in their current condition, to a height of at least 3 meters and with the inside toe 20 m from the long term break-back line of the pit/void;
Rock enviro-bunds will be entirely created from existing waste rock material already dumped within the mine lease area and will be transported via truck and shovel to the open pits;
Creation of an enviro-bund around Mupine pit, which will not be used for tailings disposal in unscheduled closure, has been included;
Provision of an enviro-bund around the KTO East Pit has in included in the closure cost update which had been omitted in the indicative costing;
It was assumed the mining voids with standing water will be left as is; and
No cost has been allowed for ongoing water treatment following closure as the need and scope of this aspect is not known at present; however it is foreseen that this could amount to significant capital and ongoing operational costs.
As for unscheduled closure, but assuming that the open pit expansion as a result of future mining will require the doubling of the total length of enviro-bund to be created;
Rock enviro-bunds will be entirely created from material excavated from the pits during future mining and will largely form part of operational costs. However closure cost provision has been made for 20% of the total volume of rock, to account for movement of material and repair of bunds required during closure;
Creation of an enviro-bund around Mupine pit has not been included as it was assumed that the pit will be completely filled with tailings material;
It was assumed that the pits would be de-watered for ongoing mining operations as required, but that the final mining voids will be left with standing water; and
No cost has been allowed for ongoing water treatment following closure.
Shafts The incline shaft will receive a reinforced concrete plug and the portal will be backfilled with waste rock. A shaft As for unscheduled closure.
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Closure component Closure cost assessment
Unscheduled closure (December 2011) Scheduled closure (Undetermined) portal opening of 5x5 m has been assumed.
Tailings disposal facilities
It was assumed that the tailings dams are on average 20 m high;
In-situ rehabilitation of the tailings facilities were anticipated without any capping requirements;
Allowance was made to re-shape the side slopes to a more sustainable slope of at least 1:5;
The penstock will be plugged and sealed; and
Vegetation will be established directly on the tailings material. A high application rate of organic material (80 ton/ha) was assumed to assist with the vegetation establishment.
As for unscheduled closure, however in addition it was assumed the Mupine pit will be completely filled with tailings material and that grassland vegetation will be established on the final landform.
Waste rock dumps
Large waste rock dumps will be rehabilitated in-situ;
Small dumps will be used to backfill the incline shaft portal or will be used for the construction of enviro-bunds;
Dumps that will be rehabilitated in-situ will be reshaped to decrease slope angles to at least 1:5;
In-situ rehabilitation of the large rock dumps were anticipated without any capping requirements;
It was anticipated that no topsoil will be required on the rock dumps; and
Allowance was made to establish grassland vegetation on the waste rock dumps.
As per unscheduled closure, however for scheduled closure it was assumed that the enviro-bunds will be created from material originating from the pits as a result of future mining. The result is that approximately 740 000m3 of additional waste rock material will be shaped in-situ during scheduled closure.
Evaporation ponds The rehabilitation of the sediment ponds was assumed
to include the removal of contaminated sediment to a depth of approximately 400mm; removal of a synthetic liner and contaminated soil under the liner as a result
As per unscheduled closure.
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Closure component Closure cost assessment
Unscheduled closure (December 2011) Scheduled closure (Undetermined) of potential leaks to a depth of 250mm; and the load and haul of contaminated sediment and disposal onto one of the tailings dams., and establishment of woodland vegetation on the final levelled surface.
26.1.7.3 General Surface Reclamation
Topsoil / growth medium No allowance was made to import topsoil except where
specifically indicated. It was assumed that the underlying soil can be effectively rehabilitated and ameliorated to sustain vegetation.
As for unscheduled closure.
Removal of contaminated material.
An allowance was made to remove 500 mm of contaminated material over 20%of all infrastructural and generally disturbed areas but excluding all water management infrastructure and drainage areas, and to dispose of this material on a tailings dam.
As for unscheduled closure.
Shaping and levelling of footprint areas
Allowance was made to stockpile demolition waste for removal, fill excavations through a cut to fill action and re-profile the area to allow free drainage.
As for unscheduled closure.
Establish vegetation An allowance was made for soil amelioration,
cultivation and seeding with suitable indigenous woodland species mixture.
As for unscheduled closure.
26.1.7.4 Water Management
Re-instatement of drainage lines
Allowance has been made to re-instate natural drainage lines over the Allowance for full reclaimed area; and
A nominal allowance was made to rehabilitate polluted sediment washout areas in watercourses downstream of the tailings dams.
As for unscheduled closure.
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Closure component Closure cost assessment
Unscheduled closure (December 2011) Scheduled closure (Undetermined) 26.1.7.5 Post-closure Aspects
Surface water and groundwater monitoring.
Surface water monthly monitoring over a 5 year period at 6 surface water monitoring points.
Groundwater monitoring on a quarterly basis over a 5 year period at 16 monitoring points.
Surface water monthly monitoring s for unscheduled closure.
Groundwater monitoring on a quarterly basis over a 5 year period at 14 monitoring points.
Rehabilitation monitoring An allowance has been included for both high intensity
monitoring (all infrastructure and tailings dam areas) over a five year period; and low intensity rehabilitation monitoring (remaining areas) over a three year period.
As for unscheduled closure, however the additional electro-winning and solvent extraction plant as well as Mupine tailings disposal facility have been included as high intensity rehabilitation areas.
Care and maintenance
An allowance has been included for both high intensity care and maintenance (all infrastructure and tailings dam areas) over a five year period; and low intensity care and maintenance (remaining areas) over a three year period.
As for unscheduled closure, however the additional electro-winning and solvent extraction plant as well as Mupine tailings disposal facility have been included as high intensity rehabilitation areas.
26.1.7.6 Additional Allowances
Preliminary and general Additional allowance of 6% of the total for
infrastructural and related aspects has been made, which is aligned to best practice guidelines.
As for unscheduled closure.
Contingencies Additional allowance of 10%of the total for
infrastructure and related aspects, which is aligned to best practice guidelines.
As for unscheduled closure.
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Table 99: Overall cost comparison Kamoto Dima Closure Costs
INFRASTRUCTURE AND RELATED ASPECTS Previous indicative cost (Dec 2010)
UNSCHEDULED Closure cost update (December 2011)
SCHEDULED Closure cost update (Undetermined)
1 Infrastructural aspects 37.5 43.5 56.7 2 Mining aspects 8.8 13.6 13.0 3 General surface reclamation 45.2 45.7 45.3 4 Water management 0.9 0.9 117.8
SUB-TOTAL 1 (Infrastructure and related aspects) 94.1 105.2 117.8
5 Post closure aspects 1.8 1.6 1.9
SUB-TOTAL 2 (Post-closure aspects) 1.8 1.6 1.9
6 ADDITIONAL ALLOWANCES
6.1 Preliminary and general 5.6 6.3 7.1 6.2 Contingencies 9.4 10.5 11.8
SUB-TOTAL 3 (Additional allowances) 15.1 16.8 18.8
GRAND TOTAL (Sub-total 1+2+3) 110.9 123.6 138.5
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26.1.8 Summary of Update Closure Costs All unit rates previously utilised for the indicative closure cost of 2010 were benchmarked against those of a number of recent and comparable closure costings for mines in other Africa countries. The rates were, where deemed appropriate, also cross-checked against current rates for South African-based closure projects and adjusted where relevant.
In summary, the scheduled closure cost of approximately $138.5 million is approximately 25% higher than the previous indicative closure costs of $110.9 million as determined by GAA in December 2010. Of this, the largest percentage is attributed to the inclusion of the costs associated with the closure and rehabilitation of a substantial amount of additional plant and infrastructure, namely the electro-winning and solvent extraction plant. It is estimated that the costs associated with dismantling and removing the existing plant infrastructure from site amounts to approximately $39.6 million, whilst the additional plant will cost approximately $13.2 million extra.
Furthermore, the inclusion of the rehabilitation of the KOV stockpiles, KTO East Pit and the Mupine Pit tailings disposal facility in scheduled closure, which was not incorporated into the previous indicative closure costing, accounts for a further estimated $1 million. The remaining difference is largely attributable to a refinement in the closure approach in terms of the mining areas and the various applicable unit rates.
The unscheduled closure cost of approximately $123.6 million, is approximately $12.7 million higher than the escalated indicative closure cost of 2010. This difference is largely attributable to a refinement in the closure approach in terms of the mining areas and the various applicable unit rates, as well as the inclusion the rehabilitation of the KTO East Pit.
For both scheduled and unscheduled closure, the totals for general surface reclamation, water management and post-closure aspects have not changed significantly as several rates applied during the 2010 cost estimate were risk averse and upon revision did not require significant escalation. Refinement of the post-closure monitoring and aftercare approaches have also resulted in the provisions for these aspects remaining largely unaltered.
26.1.9 Matters requiring further attention The following matters require attention to arrive at a definitive closure cost estimate:
Limited new information was available to inform the closure cost update, specifically regarding the extent of future open pit mining and the construction of additional plant and infrastructure. As a result, potential additional costs may be associated with the decommissioning and rehabilitation of these areas. More accurate information for these activities must therefore be obtained and incorporated in future cost updates, in order to inform future financial decision making;
Accurate and more detailed vector data of the existing battery limits were not available. Hence, it was not possible to refine the existing battery limits for the purposes of compiling the bill of quantities. More detailed information could also influence future costs;
Confirmation and documentation of battery limits for the closure costing and providing the motivations/reasons for the inclusion and exclusion of areas must be obtained;
Compilation of proper inventories of infrastructure and mining activities within the respective battery limits must be compiled and signed-off by the mine;
On-site quantification and measurement of those closure cost components with uncertainty with respect to the closure measures required must be done;
Confirmation, through on-site assessments, of potential effect of any mining-related spillages on the local stream/river. In the event that contamination that can safely be removed without causing excessive damage to the streambeds is identified, allowance should be made for at least the de-silting and re-instatement of contaminated stream beds and banks. Other measures may also be required in order to allow the natural aquatic ecosystems to return as far as possible. This could have a notable
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cost and has been excluded from this cost estimate, mainly due to uncertainty on responsibility for this environmental liability and the fact that such areas could not be detected from aerial imagery;
Actual positions for and number of surface and groundwater sampling / monitoring points required must be determined, based on dedicated geo-hydrological and water quality assessments, and incorporated in updated closure cost; and
The possible need for ongoing water treatment must be established and quantified, to ensure that contaminated decant from the flooded pits and underground working do not pose a long term threat to ground and surface water quality or to associated aquatic health. As previously mentioned this aspect may result in significant capital and ongoing post-closure operating costs if required.
26.1.10 Conclusion The findings as reflected in this report have primarily been based on the interpretation of Google Earth images of the respective sites, used for the previous indicative closure cost estimate compiled in 2010. In those instances where the required information was not available, estimates were made based on experience. Limited new information comprising a number of on-site photographs and approximate footprint size estimates for the new plant infrastructure was available. This constitutes a shortcoming in terms of this assessment and ideally must be addressed for future updates.
Unit rates for the purpose of the review were obtained from GAA’s existing data base and/or from demolition practitioners. Where required, these were adapted to reflect site-specific conditions. As such, a risk-averse perspective was adopted and mainly errs on the side of caution. This approach allows for the costs to be refined further as information becomes available, as opposed to possible under-estimation and associated provision that could lead to liability shortfalls. Nevertheless due to the severely disturbed and possibly contaminated nature of the mining area, even small differences in battery limits could have a notable effect on the computed closure costs.
More work is required to arrive at improved closure costs, also giving attention to those aspects highlighted in this report.
27.0 INTERPRETATIONS AND CONCLUSIONS The results of interpretations of developments of Material Assets are reported elsewhere in this report and have been relied upon to compile the mineral resource estimates included in section 14.1 and 14.2 and the ore reserve estimate in section 15.0.
28.0 RECOMMENDATIONS The Qualified Person recommends the following actions be taken in respect of the Material Assets:
further exploration of the operations which have underground mine potential such as the T-17 underground and KTE should be continued;
dewatering strategy needs to be implemented particularly for open pits such as Mashamba East;
plant and processing improvements and construction of new infrastructure such as the Solvent Extraction Plant and Electro-Winning Plant in terms of phase 4 need to be completed;
the capacity of the existing KTC tailings facility needs to be increased and the new tailings strategy needs to be implemented i.e. the use of a lime treatment plant to treat tailings prior to deposition and the authorisation and use of the Mupine Pit as a super tailings facility needs to be operationalised;
new deposition sites should be investigated since the existing tailings and waste facilities will require expansion in the future.
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28.1 Dewatering The element dewatering strategy needs to continue to avoid flooding of surface and/or underground mine workings.
28.2 Tailings and Waste A formal surveillance program should be initiated whereby data such as under drain flows, peizometer
and freeboard readings can be measured on a monthly basis and monitored for trends.
Monthly meetings should be set up for the year.
Deposition must be planned in advance for the month. This must be checked on a daily basis regarding the actual deposition taking place on the TSF (Planned vs Actual).
Pool walls should be extended up to the penstock inlets for safe access to decant water off the TSF.
A catwalk structure of either timber or metal is to be constructed around each penstock inlet to assure safe and easy access while removing penstock rings to decant water off the TSF.
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29.0 REFERENCES Metago (2009). KML Due Diligence Environmental and Mine Waste Review. Report Number C001-02-
01, February 2009.
SRK (2010). Kamoto Project: EIS and EMP for the Project. SRK Report 411167/1, March 2010.
C.G.E.A./A.P.KAT (2010). Letter: Evaluation de la Radioactivité et Programme de Radioprotection dans les Installations Minières et Radiologiques de KAMOTO COOPER COMPANY. République Démocratique du Congo, Commissariat Général à l’Energie Atomique / Antenne Provinciale du Katanga (C.G.E.A./A.P.KAT), 20 Mai 2010.
KCC (October 2009). Site environmental incident register: November 2008 to October 2009.
KCC (October 2010). Site environmental incident register: November 2009 to October 2010.
KCC Luilu Metallurgical Laboratory Excel spreadsheet: Water results up to March 2010 (Water sample quality laboratory data from various locations for the period May of 2009 to March of 2010).
KCC Luilu Metallurgical Laboratory Excel spreadsheet: Dust samples 2010 (Monthly KCC dust fall monitoring data for the period January to September of 2010).
SRK (2009). An ITR on the Material Assets of KML, Katanga Province, DRC. SRK Project: 389772, March 17, 2009.
Talbot & Talbot Laboratories water quality analytical data reports dated as follows:
Report: Katanga 8636/10 R1, dated June 24, 2010 (samples collected 27 May 2010);
Report: Katanga 11806/10, dated August 16, 2010 (samples collected 14 to 20 July 2010); and
Report: Katanga 13881/10, dated September 16, 2010 (samples collected September 1 – 2, 2010).
GOLDER ASSOCIATES AFRICA (PTY) LTD.
/s/ Willem van der Schyff /s/ Spencer Eckstein
Willem van der Schyff Spencer Eckstein Geologist IRP
WvdS/SE/ndw
Reg. No. 2002/007104/07 Directors: FR Sutherland, AM van Niekerk, SAP Brown, L Greyling Golder, Golder Associates and the GA globe design are trademarks of Golder Associates Corporation.
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30.0 DATE AND SIGNATURE PAGES
CERTIFICATE OF A QUALIFIED PERSON
1) I, Willem van der Schyff, Professional Natural Scientist, am a Geologist at Golder Associates Africa. I reside at 11 Marauder Street, Pierre van Ryneveld, Centurion, South Africa.
2) This certificate applies to the technical report of Katanga Mining Limited (Katanga) entitled. An Independent Technical Report on the Material Assets of Katanga Mining Limited, Katanga Province, Democratic Republic of Congo dated March 31, 2012 (the Technical Report).
3) I am a graduate of the University of Pretoria, South Africa with a BSc (Hons) in Geology obtained in 1990 and a Graduate Diploma in Engineering (Mining) obtained in 2000 from the University of Witwatersrand, South Africa. I am a registered Professional Geologist with the South African Council for Natural Scientific Professions and a member of the Geological Society of South Africa. I am a qualified person as that term is defined in National Instrument 43-101 – Standards of Disclosure for Mineral Projects (NI 43-101).
4) I completed site visits to Katanga's Material Assets (as defined in the Technical Report) between January 16 – 20, 2012.
5) I am responsible for preparing and supervising the preparation of the Technical Report.
6) I am independent of Katanga as described by Section 1.4 of NI 43-101.
7) I previously assisted in the preparation of a report on the properties subject to the Technical Report and I had prior involvement with the Material Assets from 1997 to 1999 before Katanga acquired in the Material Assets, and in each case was independent of Katanga.
8) I have read NI 43-101 and the Technical Report has been prepared in compliance with NI 43-101.
9) As of the date of this certificate, 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 March 30, 2012 /s/ Willem van der Schyff ___________________ Willem van der Schyff
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APPENDIX A Document Limitations
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This Document has been provided by Golder Associates Africa Pty Ltd (Golder) subject to the following limitations:
i) This Document has been prepared for the particular purpose outlined in Golder’s proposal and no responsibility is accepted for the use of this Document, in whole or in part, in other contexts or for any other purpose.
ii) The scope and the period of Golder’s Services are as described in Golder’s proposal, and are subject to restrictions and limitations. Golder did not perform a complete assessment of all possible conditions or circumstances that may exist at the site referenced in the Document. If a service is not expressly indicated, do not assume it has been provided. If a matter is not addressed, do not assume that any determination has been made by Golder in regards to it.
iii) Conditions may exist which were undetectable given the limited nature of the enquiry Golder was retained to undertake with respect to the site. Variations in conditions may occur between investigatory locations, and there may be special conditions pertaining to the site which have not been revealed by the investigation and which have not therefore been taken into account in the Document. Accordingly, additional studies and actions may be required.
iv) In addition, it is recognised that the passage of time affects the information and assessment provided in this Document. Golder’s opinions are based upon information that existed at the time of the production of the Document. It is understood that the Services provided allowed Golder to form no more than an opinion of the actual conditions of the site at the time the site was visited and cannot be used to assess the effect of any subsequent changes in the quality of the site, or its surroundings, or any laws or regulations.
v) Any assessments made in this Document are based on the conditions indicated from published sources and the investigation described. No warranty is included, either express or implied, that the actual conditions will conform exactly to the assessments contained in this Document.
vi) Where data supplied by the client or other external sources, including previous site investigation data, have been used, it has been assumed that the information is correct unless otherwise stated. No responsibility is accepted by Golder for incomplete or inaccurate data supplied by others.
vii) The Client acknowledges that Golder may have retained sub-consultants affiliated with Golder to provide Services for the benefit of Golder. Golder will be fully responsible to the Client for the Services and work done by all of its sub-consultants and subcontractors. The Client agrees that it will only assert claims against and seek to recover losses, damages or other liabilities from Golder and not Golder’s affiliated companies. To the maximum extent allowed by law, the Client acknowledges and agrees it will not have any legal recourse, and waives any expense, loss, claim, demand, or cause of action, against Golder’s affiliated companies, and their employees, officers and directors.
viii) This Document is provided for sole use by the Client and is confidential to it and its professional advisers. No responsibility whatsoever for the contents of this Document will be accepted to any person other than the Client. Any use which a third party makes of this Document, or any reliance on or decisions to be made based on it, is the responsibility of such third parties. Golder accepts no responsibility for damages, if any, suffered by any third party as a result of decisions made or actions based on this Document.
GOLDER ASSOCIATES AFRICA (PTY) LTD
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APPENDIX B Abbreviations and Glossary of Terms
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GLOSSARY OF TECHNICAL TERMS AND DEFINITIONS 2008 Feasibility Study Feasibility Study on Kamoto Operating Limited compiled by SRK
Consulting (South Africa) (Pty) Limited dated November 2008 2009 Technical Report SRK Consulting (South Africa) (Pty) Limited Technical Report dated 17
March 2009 2010 Technical Report Technical report prepared by KML, entitled A Technical Report on the
Material Assets of KML, Katanga Province, DRC dated 31 March 2010 AJVA Has the meaning ascribed to it in Section 6.1 KCC Rights Argillaceous Term describing sedimentary rock with modal grain size in the silt fraction Assay The chemical analysis of mineral samples to determine the metal content Basal conglomerate A conglomerate formed at the earliest portion of a stratigraphical unit Bond Rod Mill Work Index Refers to the traditional measurements of ore grindability for
large/intermediate particle sizes or course rock sizes (75mm – 50mm) Bond Ball Mill Work Index Refers to the tradition measurements of ore grindability for small particle
sizes (<3mm) Care and maintenance This involves the maintaining and corrective action as requires as well as
conducting the required inspection and monitoring to demonstrate achievement of success of the implemented measures
Closure This involves the application for closure certificate and initiation of transfer of ongoing care and maintenance to third parties
Contingencies This allows for making reasonable allowance for possible oversights/omissions and possible work not foreseen at the time of compilation of the closure costs. Allowance of between 10 percent and 20 percent would usually be made based on the accuracy of the estimations.
Decommissioning This relates to the situation after cessation of operations involving the deconstruction/removal and/or transfer of surface infrastructure and the initiation of general site reclamation
Dip Angle of inclination of a geological feature/rock from the horizontal Diamictite A sedimentary rock with particle sizes ranging from clay to boulder size Drill-hole Method of sampling rock that has not been exposed D Strat (Stratified Dolomite or Dolomie Stratfie)
This is a well bedded to laminated, argillaceous dolomite, which forms the base of the traditional Lower Ore Zone in GCM' nomenclature
Fault The surface of a fracture along which movement has occurred Gecamines Generale de Carrieres et des Mines, State-owned Mining Company in the
DRC - principally concerned with Copper Mining in Katanga province Grade The measure of concentration of copper or cobalt within mineralized rock Indicated mineral resource That part of a mineral resource for which tonnage, densities, shape,
physical characteristics, grade and mineral content can be estimated with a reasonable level of confidence. It is based on exploration, sampling and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes. The locations are too widely or inappropriately spaced to confirm geological and/or grade continuity but are spaced closely enough for continuity to be assumed (JORC Code)
Inferred mineral resource That part of a mineral resource for which tonnage, grade and mineral
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content can be estimated with a low level of confidence. It is inferred from geological evidence and assumed but not verified geological and/or grade continuity. It is based on information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes which may be limited or of uncertain quality and reliability (JORC Code)
Katangan Copperbelt Refers to the copper mineral complex which stretches between Zambia and the DRC and in this instance refers to the copper found in the Katangan region
KOV Open Pit KOV open pit mine, an operating open pit mine Lithology or lithogical Geological description pertaining to different rock types Lower Roan Subgroup of the Roan Group consisting of shales with grit; dolomites;
Argillaceous dolomites; arenites and argillites; main Cu stratiform mineralization, also referred to as R2
Mashamba East Open Pit Mashamba East mine, a dormant open pit mine Material Assets Means the properties listed under Section 4.2 of the ITR. measured mineral resource That part of a mineral resource for which tonnage, densities, shape,
physical characteristics, grade and mineral content can be estimated with a high level of confidence. It is based on detailed and reliable exploration, sampling and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes. The locations are spaced closely enough to confirm geological and grade continuity (JORC Code)
Metasedimentary Metamorphosed sedimentary rock mineral resource A concentration or occurrence of material of economic interest in or on
the earth's crust in such form, quality and quantity that there are reasonable and realistic prospects for eventual economic extraction. The location, quantity, grade geological characteristics and continuity of a mineral resource are known, estimated or interpreted from specific geological evidence and knowledge. Mineral resources are sub-divided in order of increasing geological confidence, into inferred, indicated and measured categories (JORC Code)
Musonoie-T-17 West The Musonoie-T-17 open pit mine is the initial oxide source for the mine complex. It is a new site, with no significant production history to date. Pre-stripping began in May 2007 and the first blast was fired at the end of September 2007
Mwashya or R4 Altered stratified greyish siliceous dolomitic rock with oolitic horizons and a few bands of light yellow talcose schist
Nappe
A highly folded body of rock which has suffered considerable tectonic transport on an orogenic belt
Ore reserve The economically mineable part of a measured and/or indicated mineral resource. It includes diluting materials and allowances for losses, which may occur when the material is mined. Appropriate assessments and studies have been carried out, and include consideration of and modification by realistically assumed mining, metallurgical, economic,
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marketing legal, environmental, social and governmental factors. These assessments demonstrate at the time of reporting that extraction could reasonably be justified. Ore reserves are sub-divided in order of increasing confidence into probable ore reserves and proved ore reserves (JORC Code)
Orogeny An orogeny is a period of mountain building leading to the intensely deformed belts which constitute mountain ranges
Post-closure The period of ongoing care and maintenance, as per arrangement with third parties
Preliminary and Generals (P&Gs)
This is a key cost item which is directly related to whether third party contractors are applied for site reclamation. This cost item comprises both fixed and time-related charges. The former makes allowance for establishment (and de-establishment) of contractors on site, as well as covering their operational requirements for their offices (electricity/water/communications) and latrines. Time-related items make allowance for the running costs of the fixed charged items for the contract period.
preliminary economic assessment A study other than a pre-feasibility or feasibility study (as defined in NI 43-101), that includes an economic analysis of the potential viability of mineral resources (as defined in NI 43-101)
probable ore reserve The economically mineable part of an indicated, and in some circumstances, a measured mineral resource. It includes diluting materials and allowances for losses which may occur when the material is mined. Appropriate assessments and studies have been carried out, and include consideration of and modification by realistically assumed mining, metallurgical, economic, marketing, legal, environmental, social and governmental factors. These assessments demonstrate at the time of reporting that extraction could reasonably be justified (JORC Code)
Proterozoic Era of geological time between 2,5x109 and 570x106 years ago proved ore reserve The economically mineable part of a measured mineral resource. It
includes diluting materials and allowances for losses which may occur when the material is mined. Appropriate assessments and studies have been carried out, and include consideration of and modification by realistically assumed mining, metallurgical, economic, marketing, legal, environmental, social and governmental factors. These assessments demonstrate at the time of reporting that extraction could reasonably be justified (JORC Code)
Pseudomorph A secondary mineral which has replaced another but maintained its shape
Reclamation The re-instatement of a disturbed area into a usable state (not necessarily its pre-mining state) as defined by broad land use and related performance objectives
Remediation To assist in the rehabilitation process by enhancing the quality of an area through specific actions to improve especially bio-physical site conditions
Rehabilitation The return of a disturbed area to its original state, or as close as possible to this state
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Roches Argilleuses Talceuse (RAT)
The RAT is considered the boundary between the R2 and R1 units and consists of an upper RAT Grises (R2) and a lower RAT Lilas (R1)
Roches Siliceuses Feuilletées Foliated (Laminated) and Silicified Rocks (RSF)
This is a grey to light brown thinly bedded laminated and highly silicified dolomites
Roches Silicieuses Cellulaires or Siliceous Rocks with Cavities (RSC)
Vuggy and infilled massive to stromatolitic silicified dolomites
Sandstone Clastic sedimentary rock with more than 25% clasts of sand Scheduled closure Closure that happens at the planned date and/or time horizon
Schist/s A regionally metamorphosed rock characterised by a parallel arrangement of the bulk of the constituent minerals
Schistes De Base or Basal Schists (SDB)
Reddish-brown to grey silty and nodular dolomite to siltstone
Sedimentary Rocks formed by the accumulation of sediments, formed by the erosion of other rocks
Shale Argillaceous rock with closely spaced, well defined laminae Shales Dolomitiques Superieurs or Upper Dolomitic Shales (SDS)
Yellowish, cream to red bedded laminated dolomitic siltstones and fine-grained sandstones.
Standard deviation Is the square root of the variance Stratigraphy Study of stratified rocks in terms of time and space Silicilastics Clastic rock where the clasts within the rock are mostly silicate minerals
Syn Together with Tectonic Relating to a major earth structure and its deformation T-17 Open Pit T-17 Musonoi Open Pit Mine Tilwezembe Open Pit Tilwezembe, a recently closed open pit mine Transgression An incursion of the sea over land area or over a shallow sea Unscheduled closure Immediate closure of a site, representing decommissioning and
rehabilitation of the site in its present state
Variance Is the difference between the sample value and the mean grade Volcanics One of three groups into which rocks have been divided. The volcanic
assemblage includes all extrusive rocks and associated intrusive ones
Volcaniclastics One of the three groups into which rocks have been divided. The volcanic assemblage includes all extrusive rocks and associated intrusive ones
Golder Associates Africa (Pty) Ltd. PO Box 6001 Halfway House, 1685 Thandanani Park Matuka Close Halfway Gardens Midrand South Africa T: [+27] (11) 254 4800
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