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An Independent Technical Report on theMaterial Assets of

Katanga Mining Limited,Katanga Province,

Democratic Republic of Congo (“DRC”)

Katanga Mining LimitedKatanga Mining Limited

15 Golding RoadLondon

EnglandWW159JG

United Kingdom

17 March 2009

Principal Author: Roger Dixon Partner Pr.Eng, BSc (Hons) Mining, FSAIMM

ContributingAuthors:

Victor Simposya Partner (SRK) Pr.Sci.Nat, MSc (Mining), BSc (Geology), MSAIMM

Ebrahim Takolia Principal Consultant (SRK) MBA, BEconSc, MSAIMM, MSI (UK)

Wally Waldeck Partner (SRK) Pr.Eng, MBA, BSc Eng, FSAIMM, AMAMMSA

Henrietta Salter Principal Scientist (SRK) Pr. Sci. Nat, MSc, EngD

Anton von Wielligh Engineer (A&B Global) Pr.Eng, BEng (Hons)

Alan Naismith Partner (SRK) PrSciNat, MSc, MBA, FSAIMM, FSAIRE

Petrus CilliersConsultant (BatemanEngineering)

Pr.Eng, BEng (Chem Eng), MBA

Rob McNeill Partner (SRK) Pr.Tech (Eng) MSAICE, MSAPMI, MIWM

An Independent Technical Report on theMaterial Assets of

Katanga Mining Limited,Katanga Province,

DRC

Katanga Mining LimitedKatanga Mining Limited

15 Golding RoadLondon

EnglandWW159JG

United Kingdom

SRK Project Number 389772

SRK Consulting265 Oxford Road

Illovo2196

South Africa

P O Box 55291Northlands

2116South Africa

Tel: +27 11 441-1111Fax: +27 11 880-8086

[email protected]

17 March 2009

Explanatory Note

This document has been written in accordance with the requirements of the International System of

Units (SI Units) as applied in South Africa. The SI is the only system of units that is universally

recognized, so that it has a distinct advantage in establishing a dialogue globally. Even so, some

readers will be unfamiliar with the conventions of SI Units. For example, in this document, the

comma is used as the decimal marker and the space is used for the thousands separator (for numbers

larger than 9999).

In other words, 10 148,32 denotes ten thousand one hundred and forty-eight point three two. The

word ‘ton’ denotes a metric ton (1000 kg), unless otherwise stated. More information is at the

website of the Bureau International des Poids et Mesures, BIPM, at www.bipm.org. The website

offers a comprehensive, 88 page guide to SI Units in pdf format.

In some instances, non SI units are included. For instance, base-metal prices are commonly quoted in

US dollars a pound (USD/lb). In most such instances, the inclusion of the metric equivalent is

deemed unnecessary.

SRK ConsultingKML – Independent Technical Report (NI 43-101) Page 1

Compliance Cross-Reference

NI 43-101 SectionOverall

ResponsibilityResponsibility by

Section

Item 1 – Title Page Roger Dixon

Item 2 – Table of Contents Roger Dixon

Item 3 – Summary Roger Dixon

Item 4 – Introduction Roger Dixon

Item 5 – Reliance on Other Experts Roger Dixon

Item 6 – Property Description and Location Roger Dixon

Item 7 – Accesibility, Climate, Local Resources, Infrastructure and Physiography Roger Dixon

Item 8 – History Roger Dixon

Item 9 – Geological Setting Roger Dixon Victor Simposya

Item 10 – Deposit Types Roger Dixon Victor Simposya

Item 11- Mineralization Roger Dixon Victor Simposya

Item 12 – Exploration Roger Dixon Victor Simposya

Item 13 – Drilling Roger Dixon Victor Simposya

Item 14 – Sampling Method and Approach Roger Dixon Victor Simposya

Item 15 – Sample Preparation, Anlyses and Security Roger Dixon Victor Simposya

Item 16 – Data Verification Roger Dixon Victor Simposya

Item 17 – Adjacent Properties Roger Dixon Victor Simposya

Item 18 – Mineral Processing and Metallurgical Testing Roger Dixon Petrus Cilliers

Item 19 – Mineral Resource Estimates Roger Dixon Victor Simposya

Item 19 – KOV Mine Mineral Reserve Estimates Roger Dixon Wally Waldeck

Item 19 – T17, Kamoto and Mashamba East Mine Mineral Reserve Estimates Roger Dixon Anton von Wielligh

Item 20 – Other Relevant Data and Information Roger Dixon Victor Simposya

Item 20.1 – KOV Oxide / Sulphide Content Roger Dixon Victor Simposya

Item 20.2 – Dewatering Roger Dixon Henrietta Salter

Item 20.4 – Geotechnical Assessment Roger Dixon Allan Naismith

Item 20.5 – Tailings and Process Effluent Disposal Roger Dixon Rob McNeill

Item 21 – Interpretation and Conclusions Roger Dixon Victor Simposya

Item 22 – Recommendations Roger Dixon Victor Simposya

Item 23 – References Roger Dixon Victor Simposya

Item 24 – Date and Signature Page Roger Dixon Victor Simposya

Item 25 – Additional Requirements for Technical Reports on Development Properties andProduction Properties

Roger Dixon Victor Simposya

Item 25a3 – KOV Mine Mining Report Roger Dixon Wally Waldeck

Item 25a1,2&4 – T17, Kamoto and Mashamba East Mine Mining Reports Roger Dixon Anton von Wielligh

Item 25b – Recoverability Roger Dixon Petrus Cilliers

Item 25e – Environmental Considerations Roger Dixon Henrietta Salter

Item 25f,g,h,i&j – Mineral Economics Roger Dixon Ebrahim Takolia

Illustrations Roger Dixon Victor Simposya


Table of Contents3 Summary ......................................................................................................................................................................... 11

3.1 Property Description and Location............................................................................................................................... 11

3.2 Terms of Reference..................................................................................................................................................... 11

3.3 Ownership ................................................................................................................................................................... 12

3.4 Geology and Mineralization ......................................................................................................................................... 12

3.4.1 Geology ............................................................................................................................................................ 12

3.4.2 Mineralization.................................................................................................................................................... 13

3.5 Status of the Material Assets ....................................................................................................................................... 13

3.6 Mineral Resources and Reserves................................................................................................................................ 14

3.7 Interpretations and Conclusions .................................................................................................................................. 15

3.8 Recommendations ...................................................................................................................................................... 15

3.9 Economic Analysis ...................................................................................................................................................... 17

4 Introduction ......................................................................................................................................................................... 18

4.1 Description of Assets................................................................................................................................................... 18

4.2 Company Structure...................................................................................................................................................... 19

4.3 ITR – structure and compliance................................................................................................................................... 21

4.3.1 Structure ........................................................................................................................................................... 21

4.3.2 Compliance....................................................................................................................................................... 21

4.4 Terms of Reference..................................................................................................................................................... 21

4.5 Scope of Information and Site Visits ............................................................................................................................ 21

5 Reliance on Other Experts................................................................................................................................................... 23

6 Property Description and Location ..................................................................................................................................... 24

6.1 The DRC ..................................................................................................................................................................... 24

6.2 Regulatory Environment .............................................................................................................................................. 25

6.2.1 DRC Law in respect of Mining Title ................................................................................................................... 25

6.2.2 Area and Location of the Property..................................................................................................................... 26

6.2.3 Description of Legal Tenure .............................................................................................................................. 26

6.2.4 Proposed Amendments..................................................................................................................................... 30

6.2.5 DRC Mining Review.......................................................................................................................................... 31

6.2.6 Property Boundaries ............................................................................................ Error! Bookmark not defined.

6.2.7 Royalties Duties and Other Fees....................................................................................................................... 32

6.3 Environmental Liabilities.............................................................................................................................................. 33

7 Accessibility, Climate, Local Resources, Infrastructure and Physiography..................................................................... 37

7.1 Accessibility................................................................................................................................................................. 37

7.2 Climate........................................................................................................................................................................ 38

7.3 Local Resources.......................................................................................................................................................... 39

7.4 Infrastructure ............................................................................................................................................................... 39

7.5 Physiography .............................................................................................................................................................. 39

8 History ......................................................................................................................................................................... 40

8.1 Introduction ................................................................................................................................................................. 40

8.2 Prior Ownership of the Material Assets........................................................................................................................ 40

8.2.1 KCC Assets ...................................................................................................................................................... 40

8.2.2 DCP Assets ...................................................................................................................................................... 40

8.2.3 The Merger ....................................................................................................................................................... 40

8.3 Historical Development................................................................................................................................................ 40

8.4 Historical Exploration................................................................................................................................................... 41


8.5 Historical drilling .......................................................................................................................................................... 41

8.5.1 T17 Mine........................................................................................................................................................... 41

8.5.2 Tilwezembe Mine.............................................................................................................................................. 41

8.5.3 Kamoto Mine..................................................................................................................................................... 41

8.5.4 Kananga Mine................................................................................................................................................... 42

8.5.5 KOV Mine ......................................................................................................................................................... 42

8.5.6 Mashamba East Mine ....................................................................................................................................... 42

8.6 T17 Mine ..................................................................................................................................................................... 42

8.7 Tilwezembe Mine ........................................................................................................................................................ 43

8.8 Kamoto Mine ............................................................................................................................................................... 43

8.9 Kananga Mine ............................................................................................................................................................. 44

8.10 KOV Mine.................................................................................................................................................................... 44

8.11 Mashamba East Mine.................................................................................................................................................. 45

8.12 Kamoto Concentrator .................................................................................................................................................. 45

8.13 Luilu Metallurgical Plant............................................................................................................................................... 46

9 Geological Setting................................................................................................................................................................ 47

9.1 Regional Geology........................................................................................................................................................ 47

9.2 General Stratigraphy ................................................................................................................................................... 49

9.3 Project geology............................................................................................................................................................ 51

9.3.1 T17 Mine........................................................................................................................................................... 51

9.3.2 Tilwezembe Mine.............................................................................................................................................. 51

9.3.3 Kamoto Mine..................................................................................................................................................... 51

9.3.4 Kananga Mine................................................................................................................................................... 52

9.3.5 KOV Mine ......................................................................................................................................................... 52

9.3.6 Mashamba East Mine ....................................................................................................................................... 52

10 Deposit Types....................................................................................................................................................................... 53

11 Mineralization ....................................................................................................................................................................... 53

12 Exploration ......................................................................................................................................................................... 53

13 Drilling ......................................................................................................................................................................... 54

13.1 T17 Mine ..................................................................................................................................................................... 54

13.2 Tilwezembe Mine ........................................................................................................................................................ 54

13.3 Kamoto Mine ............................................................................................................................................................... 54

13.4 Kananga Mine ............................................................................................................................................................. 54

13.5 KOV Mine.................................................................................................................................................................... 54

13.6 Mashamba East Mine.................................................................................................................................................. 56

14 Sampling Method and Approach ......................................................................................................................................... 57

14.1 Historical Sampling...................................................................................................................................................... 57

14.2 T17 Mine, Kamoto Mine and Mashamba East Mine..................................................................................................... 58

14.3 Kananga Mine and Tilwezembe Mine .......................................................................................................................... 58

14.4 KOV Mine.................................................................................................................................................................... 58

15 Sample Preparation, Analyses and Security ...................................................................................................................... 59



15.3 KOV Mine.................................................................................................................................................................... 60

16 Data Verification ................................................................................................................................................................... 61




16.3 KOV Mine.................................................................................................................................................................... 62

17 Adjacent Properties.............................................................................................................................................................. 63

18 Mineral Processing and Metallurgical Testing ................................................................................................................... 63

18.1 Kamoto Mine Testwork................................................................................................................................................ 63

18.2 Previous KOV Testwork .............................................................................................................................................. 64

18.2.1 Samples Tested................................................................................................................................................ 64

18.2.2 Milling Testwork ................................................................................................................................................ 64

18.2.3 Hydrometallurgical Testwork ............................................................................................................................. 64

18.2.4 Other Testwork ................................................................................................................................................. 65

18.3 Recent KOV Mine Testwork ........................................................................................................................................ 65

18.3.1 Milling Testwork ................................................................................................................................................ 65

18.3.2 Flotation Testwork............................................................................................................................................. 66

18.3.3 Slurry Pumping Testwork .................................................................................................................................. 66

18.4 Oxide / Sulphide Content............................................................................................................................................. 66

18.5 Risks and Recommendations ...................................................................................................................................... 66

19 Mineral Resource and Mineral Reserve Estimates............................................................................................................. 67

19.1 Mineral Resource Estimates........................................................................................................................................ 67

19.1.1 Mineral Resource Estimation Methodology ....................................................................................................... 67

19.1.2 Data quality and quantity................................................................................................................................... 68

19.1.3 Core Recovery.................................................................................................................................................. 71

19.1.4 Data manipulation ............................................................................................................................................. 74

19.1.5 Grade distributions............................................................................................................................................ 76

19.1.6 Statistics ........................................................................................................................................................... 83

19.1.7 Variography ...................................................................................................................................................... 89

19.1.8 Grade estimation............................................................................................................................................... 97

19.2 Density Determinations ............................................................................................................................................. 100

19.2.1 T17 Mine......................................................................................................................................................... 101

19.2.2 Tilwezembe Mine............................................................................................................................................ 102

19.2.3 Kamoto Mine................................................................................................................................................... 102

19.2.4 Kananga Mine................................................................................................................................................. 103

19.2.5 KOV Mine ....................................................................................................................................................... 103

19.2.6 Mashamba East Mine ..................................................................................................................................... 104

19.2.7 Summary ........................................................................................................................................................ 105

19.3 Block model validation............................................................................................................................................... 105

19.4.1 T17 Mine......................................................................................................................................................... 106

19.4.2 Tilwezembe Mine............................................................................................................................................ 106

19.4.3 Kamoto Mine................................................................................................................................................... 108

19.4.4 Kananga Mine................................................................................................................................................. 109

19.4.5 KOV Mine ....................................................................................................................................................... 109

19.4.6 Mashamba East Mine ..................................................................................................................................... 111

19.5 Consolidated Mineral Resource Statement................................................................................................................ 112

19.6 Comparison of the 2008 and 2007 Mineral Resources .............................................................................................. 112

19.7 Consolidated Mineral Reserve Statement.................................................................................................................. 114

19.7.1 Comparison of the 2008 and 2007 Mineral Reserves...................................................................................... 114

20 Other Relevant Data and Information................................................................................................................................ 116

20.1 KOV Oxide / Sulphide Content .................................................................................................................................. 116

20.1.1 Description...................................................................................................................................................... 116


20.1.2 Assumptions ................................................................................................................................................... 117

20.2 Dewatering ................................................................................................................................................................ 117

20.2.1 Introduction..................................................................................................................................................... 117

20.2.2 Kamoto Underground...................................................................................................................................... 117

20.2.3 T17 ................................................................................................................................................................. 118

20.2.4 Mashamba East.............................................................................................................................................. 119

20.2.5 Dewatering of the KOV pit............................................................................................................................... 119

20.3 Infrastructure ............................................................................................................................................................. 122

20.3.1 General Infrastructure ..................................................................................................................................... 122

20.3.2 Power ............................................................................................................................................................. 125

20.4 Geotechnical Assessment ......................................................................................................................................... 126

20.4.1 Kamoto Mine................................................................................................................................................... 127

20.4.2 Open Pits........................................................................................................................................................ 128

20.4.3 Recommendation............................................................................................................................................ 129

20.5 Tailings and Process Effluent Disposal...................................................................................................................... 129

20.5.1 Scope of Study ............................................................................................................................................... 130

20.5.2 Design Assumptions ....................................................................................................................................... 130

20.5.3 Selection of Preferred Tailings Dams .............................................................................................................. 130

20.5.4 Description of Proposed Tailings Dams........................................................................................................... 131

20.5.5 Phasing of Tailings Dams ............................................................................................................................... 135

20.5.6 Risk Assessments........................................................................................................................................... 137

20.5.7 Hazardous Effluent Ponds .............................................................................................................................. 137

20.5.8 Closure Considerations................................................................................................................................... 137

20.5.9 Further Study .................................................................................................................................................. 137

20.5.10 Conclusions ................................................................................................................................................ 138

20.6 Risk Assessment....................................................................................................................................................... 139

20.6.1 Basis of the Risk Report.................................................................................................................................. 139

20.6.2 Risk Register .................................................................................................................................................. 139

20.6.3 Resources: Overstated Resource Estimate for T17 Mine ................................................................................ 139

20.6.4 Mining and Reserves: KOV Equipment ........................................................................................................... 139

20.6.5 Mining and Reserves: KOV Contract............................................................................................................... 139

20.6.6 Metallurgical Processing: Inability to Meet Schedule for Modules 1, 2 and 3 ................................................... 140

20.6.7 Metallurgical Processing: Unavailability and Quality of Key Reagents ............................................................. 140

20.6.8 Services: Poor Condition of Railway Line........................................................................................................ 140

20.6.9 Services: Availability of Rolling Stock.............................................................................................................. 141

20.6.10 Services: Under-developed in-country institutional infrastructure and capacity ............................................ 141

20.6.11 Services: Lack of Power Supply .................................................................................................................. 141

20.6.12 Environmental: Non-resolution of Liabilities ................................................................................................. 142

20.6.13 Environmental: Non-compliance with DRC Mining Code ............................................................................. 142

20.6.14 Human Resources: Senior Management and Technical Expertise............................................................... 142

20.6.15 Capital Costs: The Unpredictable Escalation of Costs ................................................................................. 142

20.6.16 Operating Costs: Deviation from Engineering Study Estimates.................................................................... 143

20.6.17 Risk Controls............................................................................................................................................... 143

21 Interpretations and Conclusions....................................................................................................................................... 144

22 Recommendations ............................................................................................................................................................. 144

23 References ....................................................................................................................................................................... 145

24 Date and Signature Pages ................................................................................................................................................. 145


24.1 Roger Dixon .............................................................................................................................................................. 146

24.2 Victor Simposya ........................................................................................................................................................ 147

24.3 Ebrahim Takolia ........................................................................................................................................................ 148

24.4 Herbert Gerald Waldeck ............................................................................................................................................ 149

24.5 Henrietta Salter ......................................................................................................................................................... 150

24.6 Anton von Wielligh..................................................................................................................................................... 151

24.7 Alan Naismith ............................................................................................................................................................ 152

24.8 Petrus Cilliers ............................................................................................................................................................ 153

24.9 Rob McNeill............................................................................................................................................................... 154

25 Additional Requirements for Production Properties........................................................................................................ 155

25a Mining Operations ............................................................................................................................................................... 155

25a.1 T17 Mine .................................................................................................................................................................... 155

25a.1.1 LoM Plan ...................................................................................................................................................... 155

25a.1.2 Mining Operations......................................................................................................................................... 155

25a.1.3 Risks ............................................................................................................................................................ 156

25a.2 Kamoto Mine .............................................................................................................................................................. 156

25a.2.1 LoM Plan ...................................................................................................................................................... 156


25a.2.3 Backfill .......................................................................................................................................................... 158

25a.2.4 Ventilation..................................................................................................................................................... 158

25a.2.5 Survey .......................................................................................................................................................... 158

25a.2.6 Opportunities ................................................................................................................................................ 158

25a.3 KOV Mine ................................................................................................................................................................... 158

25a.3.1 LoM Plan ...................................................................................................................................................... 158


25a.4 Mashamba East Mine ................................................................................................................................................. 160

25a.4.1 LoM Plan ...................................................................................................................................................... 160


25a.4.3 Risks ............................................................................................................................................................ 160

25b Recoverability...................................................................................................................................................................... 162

25b.1 Source of Information ................................................................................................................................................. 162

25b.2 Introduction................................................................................................................................................................. 162

25b.3 Ore Sources ............................................................................................................................................................... 162

25b.3.1 Ore Type and Mineralogy ............................................................................................................................. 162

25b.3.2 KOV Mine ..................................................................................................................................................... 163

25b.3.3 Kamoto Mine ................................................................................................................................................ 163

25b.4 Metallurgy................................................................................................................................................................... 163

25b.5 Processing Facilities ................................................................................................................................................... 164

25b.5.1 Kolwezi Concentrator.................................................................................................................................... 164

25b.5.2 Luilu Electro-Refinery.................................................................................................................................... 164

25b.5.3 Kamoto Concentrator.................................................................................................................................... 164

25b.5.4 Luilu Metallurgical Plant ................................................................................................................................ 165

25b.5.5 WOL/SX/EW Refinery Project....................................................................................................................... 166

25b.5.6 Process Plant Capacity................................................................................................................................. 168

25b.6 Copper and Cobalt Recovery...................................................................................................................................... 169

25b.7 Processing Schedule .................................................................................................................................................. 170

25b.7.1 Ore from KOV Mine ...................................................................................................................................... 170


25b.7.2 Ore from Kamoto Mine.................................................................................................................................. 170

25b.7.3 Other Ore Sources........................................................................................................................................ 170

25c Marketing Study................................................................................................................................................................... 173

25c.1 Introduction................................................................................................................................................................. 173

25c.2 Copper Marketing ....................................................................................................................................................... 173

25c.3 CRU Copper Price Forecasts...................................................................................................................................... 173

25c.4 Cobalt Marketing......................................................................................................................................................... 174

25c.5 CRU Cobalt Price Forecasts ....................................................................................................................................... 175

25d Contracts ....................................................................................................................................................................... 178

25e Environmental Considerations........................................................................................................................................... 178

25e.1 Legislation and compliance......................................................................................................................................... 178

25e.2 Environmental and social issues................................................................................................................................. 181

25e.2.1 General......................................................................................................................................................... 181

25e.2.2 Outstanding information................................................................................................................................ 182

25e.2.3 Ground and surface water ............................................................................................................................ 183

25e.2.4 Relocation of residents of Musonoi Village.................................................................................................... 185

25e.2.5 Relocation of residents of Ngonzo Village and other villages affected by Far West Tailings Dam.................. 186

25e.2.6 Air pollution................................................................................................................................................... 186

25e.2.7 Social issues................................................................................................................................................. 187

25e.2.8 Radiation ...................................................................................................................................................... 187

25e.3 Closure....................................................................................................................................................................... 188

25e.3.1 Closure planning........................................................................................................................................... 188

25e.3.2 Allocation of Closure Costs ........................................................................................................................... 189

25e.4 Material risks and potential opportunities to reduce liabilities ...................................................................................... 190

25e.4.1 Risks ............................................................................................................................................................ 190

25e.4.2 Opportunities ................................................................................................................................................ 191

25f Taxes and Key Business Operating Parameters................................................................................................................ 192

25g Capital and Operating Cost Estimates............................................................................................................................... 193

25g.1 Capital Cost Estimates................................................................................................................................................ 193

25g.2 Operating Cost Estimates ........................................................................................................................................... 194

25h Economic Analysis ............................................................................................................................................................. 195

25h Payback ....................................................................................................................................................................... 199

25i Mine Life ....................................................................................................................................................................... 199

Glossary of Terms, Abbreviations, Units and Chemical Elements ......................................................................................... 200

Glossary of Terms ................................................................................................................................................................ 200

Abbreviations ....................................................................................................................................................................... 205

Units ....................................................................................................................................................................... 210

Chemical Elements............................................................................................................................................................... 212


List of TablesTable 3.1: Summary Table of Material Mining Assets................................................................................................................... 13

Table 3.2: Summary Table of Material Processing Assets .............................................................................................................. 14

Table 3.3 KOL: Mineral Resources as at 31 December 2008 ........................................................................................................ 14

Table 3.4 KOL: Mineral Reserves as at 31 December 2008.......................................................................................................... 15

Table 3.5 KOL: Discount Rate Sensitivity ................................................................................................................................. 17

Table 3.6 Revenue and Grade Sensitivity................................................................................................................................... 17

Table 3.7 Revenue and Capital Cost Sensitivity.......................................................................................................................... 17

Table 3.8 Revenue and Operating Cost Sensitivity...................................................................................................................... 17

Table 4.1: Summary Table of Material Mining Assets................................................................................................................... 19

Table 4.2: Summary Table of Material Processing Assets .............................................................................................................. 19

Table 5.1: Reliance on Other Experts .......................................................................................................................................... 23

Table 6.1 Sequence of Key Socio-Political Events ...................................................................................................................... 24

Table 6.2: KCC: Mineral and Surface Rights ............................................................................................................................... 29

Table 6.3: DCP: Mineral and Surface Rights................................................................................................................................ 30

Table 6.4: Environmental Liabilities(1) ........................................................................................................................................ 34

Table 8.1: T17 Mine: Historical Production ................................................................................................................................. 42

Table 8.2: Tilwezembe Mine: Historical Production ..................................................................................................................... 43

Table 8.3: Kamoto Mine: Historical Production ........................................................................................................................... 44

Table 8.4: KOV: Historical Production ....................................................................................................................................... 45

Table 8.5: Kamoto Concentrator: Historical Production................................................................................................................. 45

Table 8.6: Luilu Metallurgical Plant: Historical Production............................................................................................................ 46

Table 19.1 T17 Mine: Core Recovery Data by Lithology............................................................................................................... 72

Table 19.2 Tilwezembe Mine: Recoveries within the Mineralized Zones ......................................................................................... 72

Table 19.3 Kamoto Mine: Core Recovery Data by Lithology ......................................................................................................... 72

Table 19.4 Kananga Mine: Recoveries within the Mineralized Zones.............................................................................................. 73

Table 19.5 KOV Mine: Recoveries within the Mineralized Zones................................................................................................... 73

Table 19.6 Mashamba East Mine: Core Recovery Data by Lithology .............................................................................................. 74

Table 19.7 T17 Mine: Statistics from the 2,5 m Composites by Lithology Type ............................................................................... 83

Table 19.8 Tilwezembe Mine: Statistics from the 1 m Composites by Lithology Type ...................................................................... 84

Table 19.9 Kamoto Mine: Kamoto Principal – Statistics per Unit ................................................................................................... 84

Table 19.10 Kamoto Mine: Etang South – Statistics per Unit............................................................................................................ 85

Table 19.11 Kamoto Mine: Etang North - Statistics per Unit ............................................................................................................ 85

Table 19.12 Kananga Mine: Statistics from the 1 m Composites by Lithology Type ............................................................................ 86

Table 19.13 KOV Mine: Virgule – Statistics from the 2,5 m Composites by Lithology Type ................................................................ 86

Table 19.14 KOV Mine: Oliveira – Statistics from the 2,5 m Composites by Lithology Type ............................................................... 87

Table 19.15 KOV Mine: FNSR – Statistics from the 2,5 m Composites by Lithology Type .................................................................. 87

Table 19.16 KOV Mine: Kamoto East – Statistics from the 2,5 m Composites by Lithology Type........................................................ 87

Table 19.17 KOV Mine: Statistics from Limited mid-RSC Sampling ................................................................................................ 89

Table 19.18 Mashamba East: Statistics from the 2,5m Composites by Lithology Type......................................................................... 89

Table 19.19 T17 Mine: Omni-directional variogram parameters by lithology..................................................................................... 90

Table 19.20 Tilwezembe Mine: Back transformed variogram parameters – Manganiferous Dolomites (Oxide and Sulphide).................... 91

Table 19.21 Tilwezembe Mine: Back transformed variogram parameters – Breccia (Oxide and Sulphide).............................................. 91

Table 19.22 Tilwezembe Mine: Back transformed variogram parameters – Tillites and Argillites (Oxide and Sulphide) .......................... 91

Table 19.23 Kamoto Mine: Copper variogram models ..................................................................................................................... 93

Table 19.24 Kamoto Mine: Cobalt variogram models ...................................................................................................................... 93

Table 19.25 Kananga Mine: Back transformed variogram parameters – upper orebody (Oxide and Sulphide) ......................................... 95


Table 19.26 Kananga Mine: Back transformed variogram parameters – internal/middle zone (Oxide and Sulphide) ................................ 95

Table 19.27 KOV Mine: Virgule- Omni-Directional Variogram Parameters by Lithology .................................................................... 96

Table 19.28 KOV Mine: Oliveira - Omni-Directional Variogram Parameters by Lithology ................................................................. 96

Table 19.29 Mashamba East Mine: Omni-directional Variogram Parameters by Lithology .................................................................. 97

Table 19.30 T17 Mine: Variogram parameters by lithology .............................................................................................................. 97

Table 19.31 Mashamba East Mine: Variogram parameters by lithology ........................................................................................... 100

Table 19.32 Gecamines criteria for assigning density values........................................................................................................... 101

Table 19.33 T17 Mine: Density Determinations on Various Lithologies........................................................................................... 101

Table 19.34 Tilwezembe Mine: Density Determinations on Various Lithologies ............................................................................... 102

Table 19.35 Kamoto Mine: Density Determinations on Various Lithologies ..................................................................................... 103

Table 19.36 Kananga Mine: Density Determinations on Various Lithologies .................................................................................... 103

Table 19.37 Mashamba East Mine: Density Determinations on Various Lithologies .......................................................................... 104

Table 19.38 Density Determinations on Various Lithologies .......................................................................................................... 105

Table 19.39 T17 Mine: Mineral Resources at 0% TCu cut-off (31 December 2008).......................................................................... 106

Table 19.40 Tilwezembe Mine: Mineral Resources at a 0,5% TCu cut-off (31 December 2008)......................................................... 107

Table 19.41 Kamoto Mine: Mineral Resources by zone (31 December 2008) .................................................................................. 108

Table 19.42 Kananga Mine: Mineral Resources at a 0,5% TCu cut-off (31 December 2008).............................................................. 109

Table 19.43 KOV Mine: Mineral Resources at a 0% TCu cut-off (31 December 2008) ..................................................................... 110

Table 19.44 Mashamba East Mine: Mineral Resources at a 0% TCu cut-off (31 December 2008) ....................................................... 111

Table 19.45 KOL: Mineral Resources as at 31 December 2008 ...................................................................................................... 112

Table 19.46 KOL: Comparison of the 2008 and 2007 Mineral Resource statements.......................................................................... 113

Table 19.47 KOL: Mineral Reserves as at 31 December 2008........................................................................................................ 114

Table 20.1 Tailings Dams: Key Dates ........................................................................................................................................ 136

Table 23.1: References ............................................................................................................................................................. 145

Table 25a.1 T17 Mine: LoM Production Profile........................................................................................................................... 155

Table 25a.2 Kamoto Mine: Mining Methods ................................................................................................................................ 156

Table 25a.3 Kamoto Mine: LoM Production Profile (2009-2023) ................................................................................................... 157

Table 25a.4 Kamoto Mine: LoM Production Profile (2024-2038) ................................................................................................... 157

Table 25a.5 KOV Mine: LoM Production Profile (2009-2023) ....................................................................................................... 159

Table 25a.6 KOV Mine: LoM Production Profile (2024-2038) ....................................................................................................... 159

Table 25a.7 Mashamba East Mine: LoM Production Profile (Base Case 2009-2027) ......................................................................... 161

Table 25b.1 Gecamines Historical Recoveries .............................................................................................................................. 165

Table 25b.2 Capacity Expansion: Refurbishment Phases................................................................................................................ 168

Table 25b.3 WOL/SX/EW Refinery Modules: Timing and Capacities ............................................................................................. 169

Table 25b.4 Assumed Copper and Cobalt Recovery by Ore Type.................................................................................................... 169

Table 25b.5 KOV Mine Design Feed Tonnage and Grade .............................................................................................................. 170

Table 25b.6 Kamoto Mine Design Feed Tonnage and Grade .......................................................................................................... 170

Table 25c.1 Marketing Study: CRU Strategies Prices and Applied Prices (2009-2023) ...................................................................... 177

Table 25e.1 Summary of key legislation and relevant compliance ................................................................................................... 180

Table 25e.2 Summary of project information gaps ........................................................................................................................ 182

Table 25e.3 Environmental Closure Costs.................................................................................................................................... 189

Table 25e.4 Allocation of Closure Costs ...................................................................................................................................... 190

Table 25f.1 Economic Analysis: Key Business Operating Parameters ............................................................................................. 192

Table 25g.1 Summary of Processing Capital Expenditure .............................................................................................................. 193

Table 25g.2 LoM Investment Capital Expenditure (2009-2018) ...................................................................................................... 194

Table 25g.3 LoM Operating Costs (2009-2018) ............................................................................................................................ 195

Table 25h.1 Discounted Cash Flow Model (2009-2023)................................................................................................................. 196


Table 25h.2 Discounted Cash Flow Model (2024-2038)................................................................................................................. 197

Table 25h.3 Discount Rate Sensitivity ......................................................................................................................................... 198

Table 25h.4 Revenue and Grade Sensitivity ................................................................................................................................. 198

Table 25h.5 Revenue and Capital Cost Sensitivity......................................................................................................................... 198

Table 25h.6 Revenue and Operating Cost Sensitivity..................................................................................................................... 198

Table 25h.7 Revenue Sensitivity using 17 March 2009 LME Forecast Contract Pricing ..................................................................... 199

List of FiguresFigure 4.1: Company Structure.................................................................................................................................................... 20

Figure 6.1: Geographic Location Map of the Material Assets .......................................................................................................... 35

Figure 6.2: General Location Map of the Material Assets ............................................................................................................... 36

Figure 9.1: Regional Geology ..................................................................................................................................................... 48

Figure 9.2: General Stratigraphy of the Katangan System ............................................................................................................... 50

Figure 13.1 KOV Mine: Comparisons of the Twin Hole Intersections, Thickness and %TCu grade by Lithology - SDB.......................... 55

Figure 13.2 KOV Mine: Comparisons of the Twin Hole Intersections, Thickness and %TCu grade by Lithology - RSF .......................... 56

Figure 13.3 KOV Mine: Comparisons of the Twin Hole Intersections, Thickness and %TCu grade by Lithology - DSTRAT ................... 56

Figure 19.1 Mashamba East: Drill-hole Location Plan, Geology and Pit Outline................................................................................ 68

Figure 19.2 T17 Mine: Drill-hole Location Plan and Geology ......................................................................................................... 69

Figure 19.3 Kananga Mine: Drill-hole Location Plan ..................................................................................................................... 69

Figure 19.4 Tilwezembe Mine: Drill-hole Location Plan, Geology and Pit Outline ............................................................................ 70

Figure 19.5 KOV Mine: Drill-hole Location Plan and Surface Topography....................................................................................... 71

Figure 19.6 T17 Mine: %TCu grade distribution in DST, RSF and DB respectively (longitudinal view) ................................................ 76

Figure 19.7 Tilwezembe Mine: %TCu grade distribution in the Breccia ............................................................................................ 77

Figure 19.8 Kamoto Mine: %TCu grade distribution in the OBI and OBS (plan view) ........................................................................ 78

Figure 19.9 Kananga Mine: Longitudinal sections showing UOB (top figure) and LOB (bottom figure) intersections ............................. 79

Figure 19.10 KOV Mine: %TCu grade distribution plots for Kamoto east and Virgule, and Oliveira and FNSR respectively ..................... 81

Figure 19.11 Mashamba East Mine: %TCu grade distribution in RSF and SDB respectively (plan view) ................................................ 82

Figure 20.1 Tailings Disposal: Probable Tailings Dams Sites .................................................................................................. 132

Figure 25b.1: Metallurgical Processing: Final Block Diagram........................................................................................................... 171

Figure 25b.2: Metallurgical Processing: Distribution by Ore Type..................................................................................................... 172

1

2

SRK House265 Oxford Road, Illovo2196 Johannesburg

PO Box 55291Northlands2116 South Africa

e-Mail: [email protected]: http://www.srk.co.za

Tel: +27 (11) 441 1111Fax: +27 (11) 880 8086

3 Summary

3.1 Property Description and Location

SRK Consulting (South Africa) (Proprietary) Limited has been commissioned by Katanga Mining

Limited (“KML”) to compile an Independent Technical Report (“ITR”) which complies with the

National Instrument 43-101: Standards of Disclosure for Mineral Companies (“NI 43-101”) on the

following operations / projects and associated infrastructure (the “Material Assets”) located near

Kolwezi in the Katanga Province of the DRC:

Mining Assets (the “Mining Complex”);

o T17, an operating open pit mine (“T17 Mine”);

o Tilwezembe, a recently closed open pit mine (“Tilwezembe Mine”);

o Kamoto, an operating underground mine (“Kamoto Mine”);

o Kananga, a dormant open pit mine (“Kananga Mine”);

o KOV open pit mine, a development project (“KOV Mine”);

o Mashamba East mine, a development project (“Mashamba East Mine”);

Processing Assets (the “Processing Complex”);

o Kamoto, an operating concentrator (“Kamoto Concentrator”);

o Luilu, an operating metallurgical plant (“Luilu Metallurgical Plant”);

o Additional WOL/SX/EW Refinery (“WOL/SX/EW Refinery Project); and

Infrastructure necessary for the production of the saleable metals.

3.2 Terms of Reference

Technical data used in this ITR has been derived using the revised Life-of-Mine (“LoM”) plan based

on the work done in the 2008 Kamoto Operating Limited Engineering Study (“2008 Study”)

compiled by SRK and Bateman Engineering.

As a result of the financial crisis, the impact of which was felt in the last quarter of 2008, and the

subsequent steep decline in commodity prices including copper and cobalt, KML changed its

strategy for the development of the mines and process plants, which necessitated a revised LoM

plan. The emphasis was to constrain capital expenditure in the initial years and to limit the

processing capacity to 310 ktpa Copper from the previous 400 ktpa Copper. Work based on the

revised LoM plan, which was completed in a two month period, is not to the level and standard

consistent with the work completed in the 2008 Study.


KOL NI43-101 TSX SUBMISSION (JN389772) - 31 MARCH 2009 (v1) February 2009

The Bateman Engineering 2008 Study and the subsequent work has been prepared for the exclusive

benefit of KOL and exclusively for the Phase 3,4 and 5 project, and is subject to a separate

agreement entered into between Bateman Engineering and KOL. The processing plant capacities as

well as the capital and operating cost estimates have been factored using the work from the 2008

Study. The capital and operating cost estimates given in 2009 work were based on Bateman

Engineering’s internal database and updated for scope changes only.

3.3 Ownership

The exploitation rights for Kamoto, Mashamba East and T17 Mines together with the Kamoto

Concentrator, Luilu Metallurgical Plant and WOL/SX/EW Refinery Project are held in a joint

venture vehicle KCC SARL (Kamoto Copper Company, “KCC”). The exploitation rights for KOV,

Tilwezembe and Kananga mines are held in a joint venture vehicle DCP SARL (DRC Copper and

Cobalt Project, “DCP”). Katanga Mining Limited (“KML”) has a 75% interest in both KCC and

DCP, with the remaining 25% of each entity held by Gecamines (La Générale des Carrières et des

Mines).

3.4 Geology and Mineralization

3.4.1 Geology

The mineralized zones are at the western end of the Katangan Copperbelt, one of the great

metallogenic provinces of the world, and which contains some of the world’s richest copper, cobalt

and uranium deposits. These deposits are hosted mainly by metasedimentary rocks of the late

proterozoic Katangan system, a 7 km thick succession of sediments with minor volcanics,

volcanoclastics and intrusives. Geochronological data indicate 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 extends for more than 300 km from Kolwezi in the north-west to Lubumbashi in

the south-east.

Stratigraphically, the rich copper and cobalt deposits found in Zambia and the DRC are localized in

the Roan Supergroup (“Roan”). The Roan occurs at the base of the Katanga succession,

unconformably overlying the basement rock of Kibaran age (mid-Proterozoic). The Roan is

separated from the overlying rocks of the Upper and Lower Kundelungu supergroups by a

conglomerate, the grand conglomerate. The Lower Kundelungu is composed of sandstones and

shales with a basal conglomerate, while the Upper Kundelungu consists essentially of sediments and

is separated from the Lower Kundelungu 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 of hundreds of metres aligned across the fault plane.



In the Katangan Copperbelt, mining for copper and cobalt occurs in these outcropping to sub-

outcropping fragments.

3.4.2 Mineralization

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, chalcocite 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 100 m 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 Mine pit.

The RSC, a lithological unit stratigraphically intermediate between the OBS and OBI 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 Gecamines, 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 Mine 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 manganite).

3.5 Status of the Material Assets

Tables 3.1 and 3.2 provide details on the staus of the assets.

Table 3.1: Summary Table of Material Mining Assets

Licence

Property Holder Type Status Expiry Date Area Comments

T17 Mine KCC op Operating 3 April 2024 3,40 km2 Mine operational

Tilwezembe Mine DCP op Dormant 3 April 2009 7,64 km2 Operations ceased in November 2008due to lower copper / cobalt prices

Kamoto Mine KCC ug Operating 3 April 2024 11,04 km2 Mine operational

Kananga Mine DCP op Dormant 3 April 2009 11,04 km2 Operations ceased due to pendingrelocation of rail line

KOV Mine DCP op Development 3 April 2009 8,49 km2 Pre-stripping and dewateringscheduled for 2009 / 2010

Mashamba EastMine

KCC op Development 3 April 2024 11,04* km2 Dewatering deferred to 2016

op = open pitug = underground* Part of the same mining licence



Table 3.2: Summary Table of Material Processing Assets

Property Holder Status Comments

Kamoto Concentrator KCC Operating Oxide 2008: 437,7 kt, Sulphide 2008: 562,8 kt

Luilu Metallurgical Plant KCC Operating Copper cathode 2008: 749 t

WOL/SX/EW RefineryProject

KCC Development 160 ktpa Cu

3.6 Mineral Resources and Reserves

As at 31December 2008, KML has Measured and Indicated Mineral Resources of 297,5 Mt of ore

with a grade of 4,02% Cu and 0,46% Co (Table 3.3 presents KML’s consolidated Mineral Resource

statement as of 31 December 2008), with Proved and Probable Mineral Reserves of 139,8 Mt of ore

with a grade of 4,50% Cu and 0,44% Co (Table 3.4 presents KML’s consolidated Mineral Reserve

statement as of 31 December 2008).

Table 3.3 KOL: Mineral Resources as at 31 December 2008

Resource Classification Project Area Mt %TCu %TCo

Measured Kamoto Mine 33,0 4,50% 0,58%

Subtotal 33,0 4,50% 0,58%

Kamoto Mine 35,7 4,69% 0,60%

Mashamba East Mine 75,0 1,80% 0,38%

Indicated T17 Mine 13,7 3,16% 0,64%

KOV Mine 126,9 5,33% 0,40%

Kananga Mine 4,1 1,61% 0,79%

Tilwezembe Mine 9,0 1,89% 0,60%

Subtotal 264,5 3,95% 0,45%

Kamoto Mine 68,7 4,60% 0,59%

Total Mashamba East Mine 75,0 1,80% 0,38%

Measured and T17 Mine 13,7 3,16% 0,64%

Indicated KOV Mine 126,9 5,33% 0,40%

Kananga Mine 4,1 1,61% 0,79%


TOTAL 297,5 4,02% 0,46%

Kamoto Mine 10,6 5,22% 0,53%


Inferred T17 Mine 16,7 1,77% 0,57%

KOV Mine 71,2 3,56% 0,32%

Kananga Mine 4,0 2,00% 0,98%


TOTAL 180,7 2,32% 0,31%

(1) Mineral Resources have been reported in accordance with the classification criteria of the South African Code for the Reporting ofMineral Resources and Mineral Reserves (the SAMREC Code).

(2) Mineral Resources are inclusive of Mineral Reserves.(3) Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability.



Table 3.4 KOL: Mineral Reserves as at 31 December 2008

Classification Project Area Mt %TCu %TCo

Proved Kamoto Mine 17,0 3,52% 0,51%

Subtotal 17,0 3,52% 0,51%

Kamoto Mine 19,4 3,70% 0,53%


Probable T17 Mine 3,1 2,67% 0,70%

KOV Mine 90,1 4,93% 0,38%

Kananga Mine 0,0 0,00% 0,00%


Subtotal 122,8 4,64% 0,43%

Kamoto Mine 36,4 3,62% 0,52%


Proved and T17 Mine 3,1 2,67% 0,70%

Probable KOV Mine 90,1 4,93% 0,38%

Kananga Mine 0,0 0,00% 0,00%


Total 139,8 4,50% 0,44%

(1) Mineral Reserves have been reported in accordance with the classification criteria of the South African Code for the Reporting ofMineral Resources and Mineral Reserves (the SAMREC Code).

(2) Mineral Resources are inclusive of Mineral Reserves.

3.7 Interpretations and Conclusions

The results and interpretations of exploration on the Material Assets are reported elsewhere in this

report and have been relied upon to compile the Mineral Resource statement included in Item 19.

3.8 Recommendations

Recommendations for future work required on the technical assets are included in other items in this

report. Specific action programs recommended are:

Dewatering;

o SRK believes that additional work is required to demonstrate correlation between

the model and field observations and data:

o A gap analysis should be undertaken to state clearly the information that is still

required and assumptions that have been used in the calibration of the model. For

example, how was the recharge value used in the model determined?

o It has been stated by AGES that there will now be a longer lead time for dewatering

to be effective for the southern part of the pit as cut 1 is now going to the north.

This scenario needs to be modelled to demonstrate that this is in fact the case.

o The aquifer parameters determined from pump testing and/or packer testing for

specific stratigraphical units should be built into the model on a more detailed level

and fed back into the detailed conceptual hydrogeological model.

o Using the more detailed conceptual hydrogeological model, the position of the

phreatic surface and potential heights of the seepage faces in the high walls should

be simulated within different geotechnical domains.



o The position of the phreatic surface (hydraulic head) at quarterly intervals for the

first 24 months needs to be modelled, followed by its position annually for the life of

mine. The probability of achieving these phreatic surfaces must also be determined.

o Additional drilling and pump testing should be undertaken to establish that the RAT

does in fact form an impermeable barrier for regional groundwater flow as this is a

critical assumption in the model.

o The model should be updated with the data from drilling currently being undertaken

to demonstrate that the aquifer on the eastern side of the Musonoi River is in fact

being impacted by the dewatering of KOV. This data will be critical in resolving the

issue of the source of flow into KOV and should be used to refine the dewatering

strategy in the future.

o Additional scenarios should be run to assess the risks if in-pit boreholes prove not to

be possible or only partially possible and pumping from the bottom of the pit

becomes the primary dewatering method. An assessment of how long it will take to

draw down the phreatic surface around the pit is required, assuming only passive

dewatering i.e. pumping from the pit sump only.

o The real and measured losses from the Musonoi catchment into the Kakifuluwe

River should be incorporated into the model.

o An assessment of the expected impacts on groundwater users needs to be undertaken

if part of the inflow into KOV comes from the east side of the Musonoi River.

Geotechnical;

o Further studies should be carried out on the rock chracteristics associated with the

Material Assets.

Tailings

o Further studies should be carried out to; fully characterise the physical and chemical

properties of the tailings streams, investigate the possibility of open end deposition,

further investigate the geology and hydrogeology at Mupine Pit and to characterise

potential construction materials.

In undertaking the study, Bateman Engineering has been provided with and has relied upon

records, documents and other information supplied by the client and other third parties. Save

as expressly stated in this report, Bateman Engineering has assumed and did not attempt to

verify the accuracy, reliability, sufficieny or validity of such information, data or records and

documents. Bateman therefore recommends:

Repeat testwork for all mineral properties;

Additional work the on the sizing of the WOL/SX/EW plant and configuration, as well as

capital and operating costs; and

Additional work on the capital and operating cost estimates.



3.9 Economic Analysis

This section presents a sensitivities for discount rates (Table 3.5) metal prices and grade

(Table 3.6); capital cost (Table 3.7) and operating costs (Table 3.8). The Net Present Values

(“NPVs”) should be considered only in relation to the risks mentioned in Item 20 of this report.

Table 3.5 KOL: Discount Rate Sensitivity

Discount Factor NPV

(%) (USDm)

8,00% 1027

10,00% 624

12,00% 324

14,00% 99

16,00% (70)

18,00% (197)

20,00% (292)

Table 3.6 Revenue and Grade Sensitivity

Sensitised Factor SensitivityRange (@14% discount rate)

Revenue / Commodity Price -15% -10% -5% 0% 10% 20% 30%

Grade (Cu %) 3,89% 4,12% 4,35% 4,50%* 5,03% 5,49% 5,95%

Grade (Co %) 0,37% 0,39% 0,41% 0,44%* 0,48% 0,52% 0,57%

(USDm) (USDm) (USDm) (USDm) (USDm) (USDm) (USDm)

Revenue / Grade (10) 27 63 99 171 242 312

* Reserve Grade

Table 3.7 Revenue and Capital Cost Sensitivity

NPV Revenue Sensitivity Range (@14% discount rate)

(USDm) -30% -20% -10% 0% 10% 20% 30%

-15% 521 558 594 630 702 773 842

Total -10% 345 382 418 454 526 597 667

Capital -5% 168 205 241 278 350 421 490

Costs 0% (10) 27 63 99 171 242 312

Sensitivity 10% (370) (333) (297) (261) (189) (118) (48)

Range 20% (738) (701) (665) (629) (557) (487) (417)

30% (1135) (1087) (1047) (1009) (937) (867) (797)

Table 3.8 Revenue and Operating Cost Sensitivity


(USDm) -30% -20% -10% 0% 10% 20% 30%

-15% 248 284 321 357 429 499 569

Total -10% 162 199 235 271 343 414 484

Operating -5% 76 113 149 185 257 328 398

Costs 0% (10) 27 63 99 171 242 312

Sensitivity 10% (182) (145) (109) (72) (1) 70 140

Range 20% (353) (317) (280) (244) (173) (102) (32)

30% (525) (489) (452) (416) (344) (274) (204)



4 IntroductionSRK Consulting (South Africa) (Proprietary) Limited has been commissioned by Katanga Mining

Limited (“KML”) to compile an Independent Technical Report (“ITR”) which complies with the

National Instrument 43-101: Standards of Disclosure for Mineral Companies (“NI 43-101”) on the

following operations / projects and associated infrastructure (the “Material Assets”) located near

Kolwezi in the Katanga Province of the DRC:

Mining Assets (the “Mining Complex”);

o T17, an operating open pit mine (“T17 Mine”);

o Tilwezembe, a recently closed open pit mine (“Tilwezembe Mine”);

o Kamoto, an operating underground mine (“Kamoto Mine”);

o Kananga, a dormant open pit mine (“Kananga Mine”);

o KOV open pit mine, a development project (“KOV Mine”);

o Mashamba East mine, a development project (“Mashamba East Mine”);

Processing Assets (the “Processing Complex”);

o Kamoto, an operating concentrator (“Kamoto Concentrator”);

o Luilu, an operating metallurgical plant (“Luilu Metallurgical Plant”);

o Additional WOL/SX/EW Refinery (“WOL/SX/EW Refinery Project); and

Infrastructure necessary for the production of the saleable metals.

4.1 Description of Assets

The exploitation rights for Kamoto, Mashamba East and T17 Mines together with the Kamoto

Concentrator, Luilu Metallurgical Plant and WOL/SX/EW Refinery Project are held in a joint

venture vehicle KCC SARL (Kamoto Copper Company, “KCC”). The exploitation rights for KOV,

Tilwezembe and Kananga mines are held in a joint venture vehicle DCP SARL (DRC Copper and

Cobalt Project, “DCP”). Katanga Mining Limited (“KML”) has a 75% interest in both KCC and

DCP, with the remaining 25% of each entity held by Gecamines (La Générale des Carrières et des

Mines). Refer to Figure 4.1 for further details.



Table 4.1: Summary Table of Material Mining Assets

Licence

Property Holder Type Status Expiry Date Area Comments

T17 Mine KCC op Operating 3 April 2024 3,40 km2 Mine operational

Tilwezembe Mine DCP op Dormant 3 April 2009 7,64 km2 Operations ceased in November 2008due to lower copper / cobalt prices

Kamoto Mine KCC ug Operating 3 April 2024 11,04 km2 Mine operational

Kananga Mine DCP op Dormant 3 April 2009 11,04 km2 Operations ceased due to pendingrelocation of rail line

KOV Mine DCP op Development 3 April 2009 8,49 km2 Pre-stripping and dewateringscheduled for 2009 / 2010

Mashamba EastMine

KCC op Development 3 April 2024 11,04* km2 Dewatering deferred to 2016

op = open pitug = underground* Part of the Kamoto mining licence

Table 4.2: Summary Table of Material Processing Assets

Property Holder Status Comments

Kamoto Concentrator KCC Operating Oxide 2008: 437,7 kt, Sulphide 2008: 562,8 kt

Luilu Metallurgical Plant KCC Operating Copper cathode 2008: 749 t

WOL/SX/EW Refinery Project KCC Early Development 160 ktpa Cu

4.2 Company Structure

KML, a Bermuda-based company, holds a 75% stake in two joint ventures with Gecamines, a State-

owned mining company in the DRC. These joint ventures, KCC and DCP, hold and operate adjacent

mining concessions.

Following the merger of KML with Nikanor Plc in January 2008 through which KML acquired its

stake in DCP, KML plan to consolidate the joint ventures into a single entity. This will require

integrating provisions of the DCP joint venture agreement into the KCC joint venture agreement, and

having the DRC Government issue a revised mining and exploitation concession to the consolidated

KCC joint venture.

Figure 4.1 illustrates the inter-corporate relationships between KML and its subsidiaries, including

the jurisdiction of incorporation.



Figure 4.1: Company Structure

SRK ConsultingKML – Independent Technical Report (NI43-101) Page 21


4.3 ITR – structure and compliance

4.3.1 Structure

This Independent Technical Report has been structured according to the Form 43-101F1 Technical

Report requirements.

4.3.2 Compliance

This report has been prepared under the direction of the Qualified Person (the “QP”) who assumes

overall professional responsibility for the document. The report, however, is published by SRK, the

commissioned entity, and accordingly SRK assumes responsibility for the views expressed herein.

Consequently all references to SRK mean the QP and vice-versa.

This technical report has been prepared in accordance with the National Instrument 43-101:

Standards of Disclosure for Mineral Companies.

4.4 Terms of Reference

The effective date (the “Effective Date”) of this ITR is deemed to be 1 January 2009, and is co-

incident with the valuation date and cash-flow projections as incorporated herein. The valuation of

the Material Assets is dependent upon the following:

Technical information as generated by KML at the Base Technical Information Date

(“BID”), which is 1 January 2009; and

Appropriate adjustments made by SRK to technical information which inter alia includes

depletion, historical performance and any additional material information provided by KML

to the Effective Date.

Technical data used in this ITR has been derived using the revised LoM plan given in the work done

in the 2008 Kamoto Operating Limited Engineering Study (“2008 Study”) compiled by SRK and

Bateman Engineering.

As a result of the financial crisis, the impact of which was felt in the last quarter of 2008, and the

subsequent steep decline in commodity prices including copper and cobalt, KML changed its

strategy for the development of the mines and process plant, which necessitated a revised LoM plan.

The emphasis was to constrain capital expenditure in the initial years and to limit the processing

capacity to 310 ktpa Copper from the previous 400 ktpa Copper. Work based on the revised LoM

plan, which was completed in a two month period, is not to the level and standard consistent with the

work completed in the 2008 Study.

The processing plant capacities as well as the capital and operating cost estimates have been factored

using the work from the 2008 Study. The capital and operating cost estimates based on the 2009

work were based on Bateman Engineering’s internal database and updated for scope changes only.

4.5 Scope of Information and Site Visits

This technical report is dependent upon technical, financial and legal input. The technical

information as provided to and taken in good faith by SRK has not been independently verified by

means of re-calculation. SRK has, however, conducted a review and assessment of all material



technical issues likely to influence the future performance of the Material Assets; the review and

assessment included the following:

Inspection visits to the Material Assets’ processing facilities, surface structures and

associated infrastructure undertaken between March and September 2008;

Discussion and enquiry following access to key on-mine and head office personnel between

March and September 2008;

A review and where considered appropriate by SRK, modification of the Material Assets

estimates and their classification of Mineral Resources and Mineral Reserves to reflect the

position as at 1 January 2009;

A review and where considered appropriate by SRK, modification of the Material Assets

production forecasts contained in the Life-of-Mine (“LoM”) plans;

Obtained independent forecasts for certain macro-economic parameters and commodity

prices and relied on these as inputs into the derivation of the cash flow projections of the

Material Assets;

Satisfied itself that such information is both appropriate and valid for valuation as reported

herein. SRK considers that with respect to all material technical-economic matters it has

undertaken all necessary investigations to ensure SAMREC compliance, in terms of the level

of disclosure; and

The Mineral Resources for all mines were estimated and are reported in accordance with the

classification criteria of the SAMREC Code. The disclosure of “Measured Mineral

Resources, “Indicated Mineral Resources” and “Inferred Mineral Resources” classifications

based on the SAMREC Code reconcile without variation to the “Measured Mineral

Resources”, “Indicated Mineral Resources” and “Inferred Mineral Resources” classifications

established by the Canadian Institute for Mining, Metallurgy and Petroleum as the CIM

Definition Standards on Mineral Resources and Mineral Reserves and referred to

NI 43-101..



5 Reliance on Other ExpertsTable 5.1 provides details of companies that provided specific information which SRK utilised in the

compilation of this technical report.

Table 5.1: Reliance on Other Experts

Organisation Area of Work Report/sNI 43-101

Section

A&B Global Mining

Mineral Reserves andMining: T17 Mine;Kamoto UndergroundMine; and Mashamba EastMine.

Mining Report on T17 Mine, Kamoto Mine and Mashamba EastMine, A&B Global Limited – 2009 Update.

19, 25a

Africon-MMCTraffic and transportationimpact study

Africon, September 2008. Draft Report: Katanga Miningoperations Kolwezi Traffic management and Road Safety plan.MMC Engineers, January 2008. Specialist Traffic andTransportation Study for proposed Copper and Cobalt MineOperations at KOV, Kananaga, Tilwezembe and Kanfukuma inKolwezi, Katanga Province DRC Final Draft Report.

25e

AGES South AfricaDewatering model andhydrogeological study

AGES, 2008. Technical Reports AS-R-2008-09-05 and AGES-R-08-01-28, AG-R-2008-11-24 KOL Geohydrology ReportVersion 3 DRAFT and AG-R-2008-10-02 KOV DewateringDFS V2 Final.

20.2

Bateman EngineeringMineral processing,metallurgical testing andon-mine infrastructure

Bateman Engineering Study November 2008 and EngineeringStudy Addendum February 2009.

18, 25b

CRU StrategiesCopper and Cobalt priceforecasts

Update of copper and cobalt price forecasts - January 23rd,2009.

25c

FM AcousticConsulting

Noise impact specialist

F le R Malherbe, March 2008. Noise Impact Study for theKamoto Project near Kolwezi in the DRC, Report no 07/1/2/Rev 2.F le R Malherbe, January 2007. Noise Impact Study for theNikanor DCP project near Kolwezi in the DRC, Revisions 1Report No 07/9/4.

25e

Foxfire Scientific Radiological assessmentFoxfire Scientific, January 2009, Katanga Mining Limited DRCCopper Mining Projects Phase 1 Radiation Survey and Samplingof Kolwezi and Tilwezembe Mining Concessions.

25e

Golder AssociatesAfrica

Aquatic ecology impactassessment

Golder Associates, March 2007. Kamoto joint Venture projectAquatic Ecology Report. Report No 10440-6031-2Golder Associates, May 2008. Baseline Assessment of AquaticEcosystems Associated with Nikanor Kolwezi Report no10473/08/2Golder Associates, March 2008, Baseline Assessment of AquaticEcosystems associated with Nikanor, Kolwezi Report No 10473-6028-1.

25e

Norton RoseLegal opinion – miningand minerals rights,permits

Legal opinion as indicated in Section 6.2 (6.2.1 to 6.2.7).Received by e-mail on 13 March 2009.

6.2

Snowden GroupMineral Resources:Kananga Mine andTilwezembe Mine

Katanga Mining Limited: Tilwezembe and Kananga Project –Mineral Reserve Estimate, July 2008.

19

Trans-Africa Projects(“TAP”)

Power requirementsSupply to Kamoto Installations in Kolwezi: Load Flows andCost Estimates – September 2008, including 2009 Addendum.

20.3.2

University ofGembloux

Ecology surveys

Nature Plus, April 2004, Ecological report on sites around thetown of Kolwezi, Gembloux Belgium.Nature Plus, April 2006, Ecological Report concerningTilwezembe copper hill, Gembloux, Belgium.Nature Plus, February 2007. Ecological Report concerningTailings dam at yenge and the Proposed Process Plant Site.

25e



6 Property Description and Location

6.1 The DRC

The Democratic Republic of the Congo (French: République démocratique du Congo), often referred

to as the DRC, and formerly known or referred to as Congo Free State, Belgian Congo, Congo-

Léopoldville, Congo-Kinshasa, and Zaire (or Zaïre in French), is the third largest country by area on

the African continent. Though it is located in the Central African UN sub-region, the nation is

economically and regionally affiliated with Southern Africa as a member of the Southern African

Development Community (SADC). It borders the Central African Republic and Sudan to the north,

Uganda, Rwanda, and Burundi to the east, Zambia and Angola to the south, and the Republic of the

Congo to the west, and it is separated from Tanzania by Lake Tanganyika to the east.

Kolwezi is the main administrative centre of the mineral-rich Kolwezi District. It is about 240 km

west of Lubumbashi, the capital of Katanga Province. The town is on one of the most significant

watersheds in Southern Africa: rivers flowing northward join the great Congo River system, and

those flowing south feed the Zambezi.

Since independence in 1960, the Democratic Republic of the Congo has endured a series of

disruptive political events, including several outbreaks of civil war. Since 2001, however, the overall

socio-political and economic climate has improved. In Katanga Province, the sequence of key socio-

political events is provided in Table 6.1.

Table 6.1 Sequence of Key Socio-Political Events

1960-1962 Secessionist war in Katanga in the aftermath of independence from Belgium

1966 Nationalization of Union Minière du Haut Katanga and the formation of Gecamines

1971 Katanga renamed Shaba

1977-1978 Civil war in Shaba. Pillage of Kolwezi in May 1978, and flooding of the mines

1990 Withdrawal of US economic aid

1991 Pillage of Kolwezi by disaffected armed forces

1996Tribal tensions in Katanga (and Kolwezi) and forced eviction of residents of Kasai

Province origin

1996 Rebel movement against the rule of President Mobutu. Little or no foreign investment

1997 End of Mobutu’s rule

1999 Civil war in the eastern DRC with ongoing impacts on international development

2002 Withdrawal of foreign troops

2002 onwards Economic reforms in cooperation with the World Bank and IMF

2002 Introduction of a new Mining Code

2003 Formation of a transitional unity government

Against this background, the local economy of Kolwezi has moved through phases of boom and

bust. The most serious downturn occurred from 1997 onwards, when Gecamines’ management and

financial problems led to drastic cuts in production (around 90%), and to widespread delays in the

payment of salaries. The decline of Gecamines precipitated a serious and ongoing economic



recession in Kolwezi. The World Bank is currently supporting efforts to restructure and recapitalize

Gecamines.

6.2 Regulatory Environment

6.2.1 DRC Law in respect of Mining Title

The DRC introduced the current Mining Code (Law No. 007/2002) (the “Code”), on 11 July 2002.

The Code was supplemented by the Mining Regulations (Decree No. 038/2003 of 26 March 2003)

(“MR”).

The right of ownership of the deposits of mineral substances constitutes in principle a right that is

separate and distinct from the rights resulting from the surface area. However, subject to any rights

of third parties over the surface, the holder of an exploitation licence has the right, pursuant to

Articles 64 and 283 of the Code, to use the land surface necessary for his activities and in particular

to build installations and infrastructures required for its mining exploitation, and to establish inside

or outside his demarcated perimeter means of communication and transport of any type. The

exploitation licence also entails the right to exploit artificial deposits (i.e. stockpiles and tailings)

located within the mining perimeter covered by the licence.

The period of validity of new exploitation permits granted under the Code (French: permis

d’exploitation) (“PE”) is 30 years. The term of validity of a PE that derives from a Concession issued

pursuant to the legal regime applicable prior to the enactment of the Code, however, expires on the

original expiry date (Articles 336 of the Code and 580(c) of the MR). However, it is renewable

several times for durations of fifteen years.

In terms of DRC property law (Law No. 73/020 of July 20, 1973), the soil and sub-soil are the

exclusive and inalienable property of the State. Rights to use the land can be obtained pursuant to a

grant of concession (French: concession ordinaire ou perpétuelle) by the State under the general

principles of property law; pursuant to a lease from the holder of a concession or pursuant to a grant

of rights to the minerals or timber located on the land.

In terms of the Code, any occupation of land depriving the rightful occupants of enjoyment of the

surface rights, any modification rendering the land unfit for cultivation, will cause the holder of the

mining rights, at the request of the rightful holders of the surface rights, to pay fair compensation,

corresponding to either the rent or the value of the land at the time of its occupation, plus 50%. Land

means the ground on which the individuals have always carried out or are effectively carrying out

any activity. However, the usual occupants of the land may, in agreement with the holder, continue

to exercise their right to cultivate the land provided that the work in the fields does not hinder the

mining activities. The owner of the surface rights may then no longer continue to construct buildings

on it. Simply passing through the land by the holder does not entitle the owner to any compensation

if no damage results there from (the Code, Article 281).

The PE entitles the holder to use the underground water and water courses within the permit area for

the requirements of the mining exploitation in compliance with the requirements set forth in the

environment plan to be submitted for the Project and approved by the Direction chargée de la

Protection de l’Environnement Minier (“DPEM”) and subject to the authorization of the Governor of

the province (Articles 64 and 283 of the Code).



The MR require the holder of a PE that is obtained pursuant to the transformation of a pre-existing

mining right to submit an Environmental Adjustment Plan (“EAP”) for approval (Article 408 of the

MR). Since the MR require that all exploitation activities should be undertaken in compliance with

the relevant approved plan for the protection of the environment (Article 404 of the MR), failure to

deliver the EAP may lead to suspension of works decided by the Minister in accordance with

Articles 292 of the Code and 570 of the MR.

Once an EAP is approved, the holder of the PE will be required to put in place a financial guarantee

as security for the performance of the rehabilitation obligations as determined in the EAP, which

must be acceptable to the DPEM. This security must be maintained until certification of satisfaction

of the obligations has been obtained. The amount of the security as well as any other sums that may

be provisioned by the titleholder for rehabilitation of the site are deductible in determining taxable

income up to 0,5% of the turnover for the tax year during which the provision is made.

In terms of the Code, a legal entity incorporated pursuant to Congolese law and that has its registered

administrative office in the DRC and whose corporate purpose is mining activity is eligible for

mining rights irrespective of the percentage equity interest held by an individual of foreign

nationality or a legal entity incorporated pursuant to foreign law (Code, Article 23).

The holder of a mining exploitation title will be subject to the mining royalties due to the Treasury

(at a rate of 2% for non-ferrous metals) on the amount of sales minus the costs of transport, analysis

concerning the quality control of the commercial product for sale, insurance, and costs relating to the

sale transaction (Code, Articles 240 and 241). Liability for mining royalties starts upon

commencement of exploitation. Such royalties are due upon sale of the product.

The transfer of a PE does not relieve the initial holder from its obligations regarding rehabilitation of

the environment (Article 186 of the Code). Liability for damages deriving from works prior to the

transfer is joint and several for both the former and the new title holder. The former holder is

required, however, to inform the new holder of any significant dangers or disadvantages resulting

from exploitation, insofar as it is aware of them. Failing which, in case of any environmental liability

arising prior to the transfer of the PE, the new holder will have the option to cancel or terminate the

transfer or to recoup a portion of the transfer price. The new holder can also request, at the expense

of the former title holder, the former title holder to eliminate the dangers or to suppress the

inconveniences that may be caused to third parties (Article 280 of the Code).

6.2.2 Area and Location of the Property

The KCC and DCP concession areas are located in the south-eastern part of the DRC near the

international border with Zambia (refer to Figure 6.1). The exact location of the KCC and DCP joint

venture concession areas is shown in Figure 6.2.

6.2.3 Description of Legal Tenure

SRK relied on the legal advisor to KML, Norton Rose LLC, for the legal tenure and mining rights

status of KML as it applies to all the Material Assets mentioned in this report.

KCC Rights

The mining rights from which KCC is currently benefiting originate from a Concession No. C23

granted by the DRC State to Gécamines.



Prior to the promulgation of the Mining Code, Gécamines’ mining rights for the exploration and

exploitation of copper, cobalt and associated mineral substances under Concession No. C23 were

granted under the regime of Order-Law No. 67/231 of May 11, 1967 relating to general legislation

for mines and hydrocarbons and renewed under the regime of Order-Law No. 81-013 of April 2,

1981 relating to general legislation for mines and hydrocarbons.

Following the entry into force of the Mining Code in 2002, Ministerial Order No.195/CAB/MINES-

HYDRO/01/2002 dated August 26, 2002 recognised Concession No. C23 as currently valid mining

rights belonging to Gécamines and transformed such rights into mining titles under the Mining Code.

As part of the transformation process, the areas covered by the Concessions under the former regime

were divided into exploitation permits and PE525, originally comprising 400 carrés was issued

having an expiration date of April 3, 2009, being the expiration date of Concession No. C23.

Exploitation permits are under the Mining Code renewable in accordance with the terms of the

Mining Code for periods of 15 years.

PE 525 covers copper, cobalt and associated mineral substances. PE525 was subsequently reduced to

297 carrés, on 30 December 2005. The land under this exploitation permit covers the area on which

the Kamoto Mine, the Kamoto, Dikuluwe, Mashamba East and West and T17 deposits are located,

as well as the facilities of the Kamoto Mine, the Kamoto concentrator, the DIMA concentrator and

the Luilu Hydro-metallurgical plants.

Exploitation permits grant to its holder the exclusive right to carry out exploration and exploitation

works of mineral matters 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.

Pursuant to the joint venture agreement No. 632/6711/SG/GC/2004 made on 4 February 2004

between Gécamines and KFL and ratified by Presidential Decree No 05/070 of 4 August 2005 (the

KCC Joint Venture Agreement) and a lease contract (“contrat d’amodiation”) No.

716/10518/SG/GC/05 dated 18 October 2005 (the KCC Lease Agreement), KCC has been granted

by Gécamines a lease authorising KCC to exercise the mining rights held by Gécamines under the

part of the PE525 covered by the KCC Lease Agreement (subject to the Mining Code and the KCC

Joint Venture Agreement). The KCC Lease Agreement is made for a term of 30 years renewable by

mutual agreement in accordance with the terms of the KCC Joint Venture Agreement.

By Minsterial Order No. 1020 dated 17 February 2006, the area covered by PE525 was reduced to

176 carrés and the balance of PE 525 converted into multiple PEs. The area covering the T17 deposit

is now situated on two carrés within PE4958.

By Ministerial Decree No.3187/CAB.MIN.MINES/01/2007 of 19 September 2007, PE525 was

further split at the request of Gécamines into two different permits, namely PE525 consisting of 20

carrés and PE8841containing most of the balance.

Gécamines has taken the view that the only areas to be leased to KCC were the mining zones of the

Kamoto, Dikuluwe, Mashamba East and West, and T17 deposits. In consequence of this, by

Ministerial Decree No. 3308/CAB.MIN.MINES/01/2007 of 28 December 2007, Gécamines has had

the area of PE525 further reduced, without KCC’s prior approval, to 13 carrés. Although the

perimeter of PE525, as reduced to 13 carrés, together with two carres in PE4958, covers the Kamoto,

Mashamba East and T17 deposits and mining zones, additional areas are required for dumps, storage



and tailings. This is a matter which has been the subject of discussion with Gécamines for some time

and has been agreed in principle to be resolved as set out below under “Proposed Amendments”.

Although it is clear under article 6.9 of the KCC Joint Venture Agreement that Gécamines is not

entitled to grant to a third party rights in the concession area granted to KCC without having

obtained KCC’s prior consent, KCC is aware that Gécamines has granted certain third parties rights

over areas which KCC maintains are covered by the KCC Lease Agreement but none of these grants

has to date interfered with KCC’s operations. Gécamines has now agreed in principle that the

exploitation permits covering the deposits be transferred to KCC. In addition it is proposed that the

Necessary Surfaces (as defined below) and installations and equipment rented from Gécamines by

KCC, following the merger with DCP, will be provided to KCC free from third party rights.

Given the changes made by Gécamines to PE525, and the disagreement with KCC which resulted

therefrom, the perimeter corresponding to the area covered by KCC Lease Agreement has not been

registered with the CAMI.

Pursuant to an agreement dated 7 February 2008, between Gécamines and KFL (the Release

Agreement), Gécamines and KFL agreed that KCC would release the Dikuluwe and Mashamba

West deposits covering an area of 7 carres contained in PE9681, which had been removed from

PE525 pursuant to the above mentioned Ministerial Decree No. 3308/CAB.MIN.MINES/01/2007 of

28 December 2007. Following this release, PE525, as reduced to 13 carrés, only covers the deposits

of Kamoto and Mashamba East.

As part of the Release Agreement, Gécamines agreed (i) to transfer to KCC the PEs covering the

areas leased under the KCC Lease Agreement and (ii) that KCC would receive an area sufficient for

the good functioning of its operations, including space for dams and the sites for tailings. The parties

also agreed that an amount shall be paid for this transfer, which amount is now agreed in principle to

be covered by the amount of the pas de porte payment.

PE525 has been renewed until 29 August 2022, pursuant to Ministerial Decree N°.

3180/CAB.MIN.MINES/01/2007 of 30 August 2007 and PE4958 has been renewed until 3 April

2024 pursuant to Ministerial Decree No. 3215/CAB.MIN.MINES/01/2007 of 21 September 2007.

Gécamines has a tailings exploitation permit No. PER 9683 covering the 13 carrés now covered by

the PE525 although there are no old tailings on this specific area which would interfere with

production from this area.

In addition to the KCC Lease Agreement, Gécamines has, pursuant to the KCC Joint Venture

Agreement, leased to KCC exclusive rights to use all processing facilities existing on the KCC–

concession areas (including the Kamoto and Dima concentrators, and Luilu plant facilities, together

with all their infrastructure and surface), and all mobile equipment. It has been agreed with

Gécamines in principle that all installations and infrastructures within the perimeter of the KCC-

concession areas, to the extent required, shall be rented by Gécamines to KCC (following its merger

with DCP) with rental being covered by the royalties agreed between Gécamines and KCC.



Table 6.2: KCC: Mineral and Surface Rights

PropertyExploitationPermitNumber

Rights Granted Location Held ByArea ofTitle

Valid Until

T17 Mine PE4958

Cu, Co and associated

minerals

+ Use of Surface

10°42'S,

25°25'EKCC

4 blocks,

1,70km2

03/04/2024;

renewable

Kamoto Mine and

Mashamba East MinePE525


minerals

+ Use of Surface

10°43'S,

25°24'EKCC

13 blocks,

11,04km2

03/04/2024;

renewable

DCP Rights

Pursuant to a joint venture agreement n° 656/6755/SG/GC/2004 made on 9 September 2004 between

Gécamines and Global Enterprises Corporate Ltd and ratified by Presidential Decree No. 05/070 of

13 October 2005 (the DCP Joint Venture Agreement), Gécamines agreed to transfer certain

exploitation permits to DCP and to grant DCP a lease and certain contractual rights over certain

facilities. These exploitation permits cover copper, cobalt and associated mineral substances. The

land under these exploitation permits comprised, at the time of execution, 32 carrés and covered the

copper and cobalt deposits of KOV, Kananga and Tilwezembe.

The ownership of the following exploitation permits have been assigned by Gécamines to DCP:

Exploitation Permit No. 4961 was assigned by Gécamines to DCP pursuant to a deed of

assignment dated 13 January 2006, registered with the CAMI on 2 March 2006. This

Exploitation Permit consists of 10 carrés, which comprise the KOV area;

Exploitation Permit No. 4960 was assigned by Gécamines to DCP pursuant to a deed of


Exploitation Permit consists of 13 carrés, which comprise the Kananga area;

Exploitation Permit No. 4963, was assigned by Gécamines to DCP pursuant a deed of


Exploitation Permit consists of 9 carrés, which comprise the Tilwezembe area;

(together the DCP Exploitation Permits).

All the DCP Exploitation Permits expire on 3 April 2009. Application has been made for their

renewal until 2024 and this is in progress. .

The application for renewal of a PE may only be refused for a limited number of reasons expressly

set out in the Code, including in particular:

Failure to pay surface charges,

Failure to demonstrate adequate remaining resource;

Insufficient financial capability of the titleholder; and

Failure to update environmental documentation.



The principal relevant ground for which the Exploitation Permits may be forfeited is due to failure

by the titleholder to pay the surface area fees per carré due pursuant to the Code. A mining

titleholder is liable to pay annual surface fees per carré.

Gécamines has, pursuant to the DCP Joint Venture Agreement, granted to DCP exclusive rights to

the rights attached to sites ancillary to the DCP Exploitation Permits, together with the processing

facilities existing on the DCP concession areas, as well as the Group West concentrator treatment

plant, the electro-refining plant known as “Luilu extension P2”, the installations known as “Siege” in

Group West, the waste sites and the Luilu hydro-metallic treatment plant, the KOV conveyor, the

other equipment at KZC, together with all their infrastructure and surface, and all mobile equipment.

It has been agreed with Gécamines in principle that all other installations currently used by DCP

within the DCP Exploitation permits or the Necessary Surfaces (as defined below), to the extent

required, shall be rented by Gécamines to KCC (following its merger with DCP) with rental being

covered by the royalties agreed between Gécamines, KCC and DCP. It has been agreed that the

Kolwezi Concentrator will be released for the benefit of Gécamines, and that Gécamines will re-

engage its former employees.

Table 6.3: DCP: Mineral and Surface Rights

PropertyExploitationPermitNumber

Rights Granted Location Held ByArea ofTitle

Valid Until

Tilwezembe Mine PE4963


minerals

+ Use of Surface

10°47'S,

25°42'EDCP

9 blocks,

7,64km2

03/04/2009;

renewable

Kananga Mine PE4960


minerals

+ Use of Surface

10°40'S,

25°28'EDCP

13 blocks,

11,04km2

03/04/2009;

renewable

KOV Mine PE4961


minerals

+ Use of Surface

10°42'S,

25°25'EDCP

10 blocks,

8,49km2

03/04/2009;

renewable

The Exploration Permits for DCP reflects the understanding from the due diligence carried out at the time of the merger between Katanga

and Nikanor.

6.2.4 Proposed Amendments

In connection with the DRC Commission Review as referred to below and in consequence of the

merger of Katanga and Nikanor PLC, Gécamines, KFL Limited, KCC, DCP and GEC are currently

finalizing the negotiations for an agreement (the Amended Joint Venture Agreement) to reflect the

proposed merger of DCP into KCC which will result in the amendment to the KCC Joint Venture

Agreement and termination of the DCP Joint Venture Agreement. As part of these discussions, it has

been agreed between Gécamines, KFL Limited and GEC (subject to final agreement in the Amended

Joint Venture Agreement) that:

(a) The whole of PE 525 (comprising 13 carrés) and part of PE 4958 (comprising 2 carrés

containing the T17 deposit) shall be transferred to KCC following completion of the merger

with DCP. The Kamoto, Mashemba East and T17 deposits and any extensions of these

deposits which are within the perimeter of PE 525 and the 2 carrés of PE 4958 to be

transferred, shall be for the sole benefit of KCC.



(b) The DCP Exploitation Permits shall be transferred to KCC following completion of the

merger with DCP. In addition, one carré of PE 7044 (being an extension of the Kananga

deposit) shall be transferred by Gécamines to KCC (following such merger) once the holder

of PE 652 has released the carré to be transferred from its tailings, or earlier if KCC has

agreed to grant an easement to the holder of PE 652.

(c) The perimeter of the merged KCC/DCP concession area will contain the surface

necessary for the proper operation of the current activities of KCC, including the spaces for

the dams, the future tailings produced by the current activities of KCC, the plants and other

necessary premises, as well as the storage areas (the Necessary Surfaces).

It has been agreed that the Necessary Surfaces will be sourced from PE 8841 held by Gecamines and

from one carré close to the T17 deposit. Easements shall be granted to enable KCC to establish and

maintain operating facilities for the KOV belt. An ad hoc commission, comprising technical

representations from Gécamines and KCC/DCP, shall consider and determine the source of the

Necessary Surfaces and easements. Once they have been determined, KCC shall fund an

independent contractor to determine whether the surfaces identified contain any mineral reserves.

Provided no reserves are discovered, the relevant surfaces shall be converted into multiple

exploitation permits (where required) and shall be transferred (or, in the case of the carré close to

T17, leased) to KCC, following its merger with DCP. Should any reserves be discovered in the

identified surfaces, the reserves shall be transferred to KCC and shall count as replacement reserves

under the terms of the Concession Release Agreement.

As part of the proposed amendment it has been agreed that upon the winding up or liquidation of

KCC the mining rights and titles of KCC shall revert to Gécamines without further consideration.

6.2.5 DRC Mining Review

In April 2007, a commission (the Commission) was formed by the DRC Government to review

approximately sixty (60) mining agreements entered into by para-statal companies of the Congolese

government. The KCC Joint Venture Agreement and the DCP Joint Venture Agreement were

included in the mining agreements to be reviewed.

The Commission provided its conclusions in its report made public in November 2007.

KCC and DCP were notified on 11 February 2008 by the DRC Ministry of Mines of the objections

and requirements regarding their partnerships with Gécamines further to the above-mentioned

November 2007 report.

In July 2008, Gécamines and KFL entered into a memorandum of understanding under which certain

amendments were agreed to be reflected in an amended joint venture agreement and the parties

agreed to the merger of KCC and DCP.

In August 2008, the DRC Ministry of Mines issued terms of reference for the renegotiations 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, the parties are currently negotiating the final terms of the Amended Joint Venture Agreement.

Until the Amended Joint Venture Agreement is finalised and becomes effective, the parties are

operating under the existing joint venture agreements.



6.2.6 Property Boundaries

The proposed property boundaries of the exploitation permits to be transferred to KCC and DCP are

as described in Figure 6.1 and 6.2.

The boundaries of the Necessary Surfaces will principally come from within PE8841, the boundaries

of which are shown in Figure 6.1 and 6.2. The precise areas from within PE8841 which are to be

included in Necessary Surfaces are to be agreed.

The boundaries have been taken from the maps at the CAMI relating to the boundaries of DCP’s

Exploitation Permits. Katanga has not separately surveyed the area. There are certain limited areas

outside land currently held by Gecamines where KCC will need to make application for licences to

operate infrastructure and tailings.

6.2.7 Royalties Duties and Other Fees

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, analysis concerning the quality control of

the commercial product for sale, 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.

Surface rights payable to the State

Under Article 198 of the Mining Code, KCC and DCP are required to pay surface rights fees of

USD5 per hectare per year or USD424,78 per carré for exploitation permits.

Additional surface fees are payable by KCC and DCP as holder of an exploitation mining right to the

central government of the DRC pursuant to Article 238 of the Mining Code at the rate of USD0,08

per hectare.

Royalties Payable to Gecamines

Under the KCC Joint Venture Agreement, KCC shall pay Gécamines for the use of the equipment

and facilities a sum equal to two percent (2%) of the net sales proceeds realized during the first three

(3) annual periods and one and a half percent (1.5%) of the net sales proceeds thereafter.

Under the DCP Joint Venture Agreement, DCP shall pay Gécamines for the transfer of the

Exploitation Permits and the use of the ancillary sites and processing installations a sum equal to two

percent (2%) of the net sales proceeds realized during the first four (4) years and one and a half

percent (1,5%) of the net sales proceeds thereafter.

Under the July MOU it was agreed that the royalty rate for equipment and facilities provided by

Gécamines as well as for ore reserve depletion will increase from 1,5% to 2,5% of net revenues. It

has been agreed in principle that these royalties will cover all use of the equipment and facilities and

the consumption and depletion of the deposits. “Net revenues” have been agreed to be defined as the

same basis as calculation of royalties under the DRC Mining Code, namely sales less transportation

costs, quality control costs, insurance costs and the marketing costs.



Pas de porte payable to Gecamines

A “pas de porte” (“entry premium”) payment shall be payable by KFL/GEC to Gécamines for the

access to the project. The total amount shall be USD140 million, the payment of which will be

completed as follows:

(i) USD5 million previously paid by GEC to Gecamines as a loan, being converted into a pas

de porte and therefore non-refundable;

(ii) USD135 million to be paid by KFL. This will comprise (a) USD24,5 million to be paid

by way of set-off against the amount of the advance to be granted by KFL to Gecamines for

payment of the subscription price (as described above); (b) USD5 million to be paid within

ten days of entry into the amended and restated joint venture agreement; and (c) USD10

million on an annual basis between 2009 and 2011 and USD15 million on an annual basis

between 2012 and 2015 , with a final payment in 2016 of USD15,5 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 to compensate for the

release of Dikuluwe and Mashamba West; however, any additional tonnage brought by Gécamines

to the Merged JV after the released deposits have been fully compensated will incur a new pas de

porte payment of USD35/t copper.

Customs and Duties Payable

In addition there is a requirement to pay customs duties and taxes in accordance with the law.

6.3 Environmental Liabilities

A preliminary liabilities assessment and closure costing exercise was undertaken in August 2008.

The methodology is described in the draft ESIA (SRK Report 390781/1). The project inherited a

number of existing environmental liabilities, and in addition the refurbishments, expansions and

operations have resulted in further liabilities and closure obligations. Contractually however,

Kamoto Copper Company (KCC) and Nikanor are indemnified from pre-existing liabilities provided

these have been quantified and apportioned. In light of the fact that this quantification and

apportionment did not take place prior to the commencement of refurbishments and operations,

liabilities have been split into three categories as follows:

Historic liabilities accruing to Gecamines. Historic liabilities that resulted from the

Gecamines operations will accrue to Gecamines. Where KML is not going to operate on

these areas, these liabilities can be clearly defined and would not be assumed by KML on the

basis of the joint venture agreement between KML and Gecamines. These include, but are

not limited to the Potopoto Tailings Dam and breach, the Kamoto Tailings Dam, uranium

stockpiles and demolition of existing buildings, plant and workshops.

Historic liabilities that will be assumed by KML. In certain areas, KML has commenced

with operations in areas already impacted by historic activities resulting in contributions

towards these historic liabilities. Examples in this category include the operation of the

Luilu Metallurgical Plant, Kolwezi Concentrator and Tilwezembe Mine, which while not

included in this project, have been worked by KOL and cannot be separated from historic

operations.



KOL’s current and future liabilities. The expansion and refurbishment of the project will

result in the operation of existing infrastructure and processes as well as the construction and

operation of new infrastructure that will result in impacts over the life of the mine, and

where these are not addressed through operational management will result in management at

closure, and therefore are allocated as closure obligations. These will include specifically,

but not exclusively, the Far West Tailings Dam, the Kamoto Interim Tailings Dam, new

waste rock dumps, new pipelines, and the expanded KOV Mine footprint area.

Table 6.4: Environmental Liabilities(1)

Area Unit Total KML GecaminesLuilu Metallurgical Plant (USDm) 16,7 1,5 15,2

Kolwezi Concentrator (USDm) 3,8 1,6 2,3

Kamoto Concentrator (USDm) 8,5 7,5 1,0

SKM (USDm) 0,1 0,1 0,0

Kamoto Underground Mine (USDm) 1,1 1,1 0,0

KOV Mine (USDm) 19,1 0,4 18,7

Tilwezembe Mine (USDm) 1,5 1,5 0,0

Mupine Pit (USDm) 16,8 0,0 16,8

T17 Mine (USDm) 5,1 5,1 0,0

Kamoto East (USDm) 6,8 0,0 6,8

Kamoto Tailings dam (USDm) 23,4 0,0 23,4

Poto-Poto Tailings dam (USDm) 137,7 0,0 137,7

Mashamba East Mine (USDm) 6,3 0,0 6,3

Subtotal (USDm) 246,9 18,7 228,1

Contingency @ 25% (USDm) 61,7 4,7 57,0

EPCM @ 10% (USDm) 24,7 1,9 22,8

Owners Cost @ 5% (USDm) 12,3 0,9 11,4

Subtotal (USDm) 98,7 7,5 91,2

Total (USDm) 345,6 26,2 319,4

(1) This assessment and the resultant liabilities are subject to negotiations with Gecamines.



Figure 6.1: Geographic Location Map of the Material Assets



Figure 6.2: General Location Map of the Material Assets


SRK G:\389772\ reports\KOL NI43-101 TSX SUBMISSION (JN389772) - 31 MARCH 2009 (v1) February 2009

7 Accessibility, Climate, Local Resources,Infrastructure and Physiography

7.1 Accessibility

Ground transport in the DRC has always been difficult. The terrain and climate of the Congo

Basin present serious barriers to road and rail construction, and the distances are enormous

across this vast country. Furthermore, chronic economic mismanagement and internal conflict

have led to serious under-investment over many years. On the other hand, the DRC has

thousands of kilometres of navigable waterways, and water transport has been the traditional,

dominant means of moving around approximately two-thirds of the country. The two civil

wars saw great destruction of transport infrastructure, from which the country has not yet

recovered. Many vehicles were destroyed or commandeered by militias, especially in the

north and east of the country, and the fuel-supply system was also badly affected.

Consequently, outside of Kinshasa, Matadi and Lubumbashi, private and commercial road

transport is almost non-existent, and traffic is scarce even where roads are in good condition.

The few vehicles in use outside these cities are run by the United Nations, aid agencies, the

DRC government, and a few larger companies such as those in the mining and energy sectors.

Air transport is the only effective means of moving between many places within the country.

The government, the United Nations, aid organizations and large companies use air rather

than ground transport to move personnel and freight. Compared to other African countries the

DRC has a large number of small domestic airlines and air charter companies.

Specific transportation relevant to the operations:

Rail Systems

The rail route between Likasi and Kolwezi is electrified (25 kV AC traction); the non-

electrified rail section from Likasi to the Zambian border is generally serviced by

diesel electric locomotives. The rail system operates at very low capacity as the

condition of the railway network is extremely poor.

There are restrictions on this line with the maximum speed limited to about 40 km/h

while the load is limited to 800 t or the train length is restricted to 14 wagons. There

are also width and load constraints where the railway line crosses the Lualaba river

(width approximately 3,0 m and load capacity of 65 t). Delivery time for rail transport

from Durban to Kolwezi is typically three weeks.

The Luilu railway station is equipped with sidings that will assist in the logistics of

the large volumes of imported reagents as well as the export of product.

Road Transport

Road access for equipment and material to be imported into the DRC from the south

is via Zimbabwe and Zambia, crossing the border into the DRC at the Kasumbalesa

border post. The road between Zambia and Likasi is in a fair condition. The 196 km

road between Likasi and Kolwezi is in exceptionally poor condition apart from the



final 30 km outside (east) of Kolwezi itself. Travel time for this distance averages

between four to five hours. There are three restrictions on the route, single lane

bridges at distances of 70 and 80 km from Likasi (these are not considered to be a

problem as there are bypass facilities at the bridge sites) and a bridge with a 20 t load

restriction where the road crosses the Nzilo lake approximately 30 km outside of

Kolwezi. Loads in excess of 20 t are rerouted to a pontoon to cross this section.

Air Transport

Lubumbashi

Lubumbashi is the main airport for the Katanga province and caters for international

flights. The airport has refuelling facilities, but there are occasional problems

obtaining fuel supplies. Maintenance facilities are available. If charter flights are

proceeding to other destinations in the province, customs and immigration must be

cleared at Lubumbashi.

Kolwezi

The air field at Kolwezi is an asphalt topped air strip 1750 m long. Site altitude is

1500 m above sea level. The condition of the strip is good, and the airfield is suitable

for medium sized aircraft. KOL has arrangements with a charter company that flies

on Wednesdays and Saturdays from Lanseria in Gauteng Province, South Africa, to

Kolwezi. Customs and immigration are cleared at Kolwezi.

7.2 Climate

Two broad climatic areas can be distinguished in the DRC. The Congo River basin, which lies

on the equator and forms around one-half of the country’s area, consists of low-lying rain

forest, which receives rainfall all year round. Temperatures are not as high as might be

expected at the equator, but humidity is generally high. The remainder of the country,

comprising the area around Kinshasa, and Kivu, Kasai and Katanga provinces, experiences

distinct rainy and dry seasons. Katanga province, lying largely at an elevation of 1000 m or

greater, experiences a climate with cooler, drier air than the majority of the country.

At only 10° latitude, daylight and night hours are almost equal, daylight lasting broadly from

06:00 to 18:00. Rapid temperature drops occur after sunset during the dry season as a result of

lack of cloud cover.

Five distinct seasons can be readily distinguished, namely:

1 Cool dry season May – July;

2 Hot dry season August – September;

3 Early rainy season October – November;

4 Full rainy season December – February; and

5 Late rainy season March - April.



7.3 Local Resources

The DRC has considerable hydroelectric power generating capacity, which is controlled and

distributed by the national power utility, Societe Nationale de Electricite (“SNEL”). Kolwezi

lies along the transcontinental railroad system and has access to both east and west coast ports

of Tanzania and Angola, as well as South Africa. Lubumbashi, some 300 km south east of

Kolwezi, is the commercial and industrial centre of the Katanga Province and hosts an

international airport.

The existing infrastructure around Kolwezi (e.g. buildings, water lines, workshops and roads)

is in a very bad state of repair. Cellular phones work in the area, although coverage is patchy.

7.4 Infrastructure

As a result of the mining activities (mainly under the control of the state-owned mining

company Gecamines) the area around Kolwezi has numerous open pits, waste rock, ore and

slag dumps, tailings dams, concentrators and other mining-related infrastructure. A substantial

urban infrastructure has also developed, including housing, roads, water supply and social

facilities. The processing plant sites are well established. Further details are provided in

Section 25b.

7.5 Physiography

Poor living conditions and civil war have endangered much of the biodiversity of the DRC.

The rising population has forced many people to become dependent on wild animals for either

their livelihood or their food. Extensive logging throughout the country by corporate loggers

and land clearing by farmers for agriculture or charcoal burners have caused destruction of the

forests throughout the country and resulted in significant degradation of habitat. Miombo

woodland is the predominant vegetation type in the Katanga Province. Such woodlands

extend over about 2,8 million km2 of the southern sub-humid tropical zone from Tanzania and

DRC in the north, through Zambia, Malawi and eastern Angola to Zimbabwe and

Mozambique in the south. Their distribution largely coincides with the flat to gently

undulating African (early Tertiary) and post African I (Miocene) planation surfaces that form

the Central African plateau. These woodlands constitute the largest more-or-less contiguous

block of deciduous tropical woodlands and dry forests in the world.

Miombo woodlands supply many goods and services that are essential to the well-being of

rural communities; some products acting as subsidies to agriculture while others provide for

basic needs, such as food, shelter and health (Clarke et al., 1996). Cavendish (2002) records

over one hundred different types of resource utilizations in a single miombo study area, with

many types having multiple species (e.g. 47 wild fruits, over 40 medicinal species, 40 wild

vegetables). Wood alone provides subsistence farmers and households with numerous

products, including poles and construction products, timber, materials for tool handles and

household utensils, foods, medicines, leaf litter, grazing and browse (Clarke et al., 1996). In

the Katanga Province, the population is heavily dependent on their supply of wood as a source

of energy through wood-fuel and charcoal production.



8 History

8.1 Introduction

This section gives a historical overview of the Material Assets including historical

development, infrastructure and operating results.

8.2 Prior Ownership of the Material Assets

From commencement of operations until 1967, all mining activities on the Property were

operated by Union Miniere du Haut Katanga (“UMHK”). However, following independence

in 1967, the mines were nationalized and incorporated as Gecamines.

8.2.1 KCC Assets

The joint venture is governed by the Kamoto Joint Venture Agreement. The parties to the

agreement are Gecamines, a DRC public enterprise incorporated under the laws of the DRC

and Kinross Forrest Limited (“KFL”), a private company incorporated under the laws of the

British Virgin Islands. Negotiations between Gecamines and KFL started in June 2001. The

Joint Venture Agreement between Gecamines and KFL was approved by Presidential Decree

dated August 2005 after all regulatory approvals were obtained.

8.2.2 DCP Assets

The joint venture is governed by the DCP SARL Joint Venture Agreement. The parties to the

agreement are Gecamines and Global Enterprises Corporate Limited (“GECL”), a private

company incorporated under the laws of the British Virgin Islands. Negotiations between

Gecamines and GECL started in May 2004. The Joint Venture Agreement between

Gecamines and GECL was approved by Presidential Decree in October 2005 after all

regulatory approvals were obtained.

8.2.3 The Merger

On 11 January 2008 KML announced that its shareholders approved the merger with Nikanor.

The merger brought together KCC and DCP assets to create a single operation.

8.3 Historical Development

Corporate mining activity in the province of Katanga began in 1906 with the formation of the

UMHK. In 1967, following national independence, the operations of UMHK were

nationalized and incorporated as Gecamines. At its peak in the late 1980s Gecamines

produced about 7% of the world’s copper and 62% of its cobalt. In 1986, Gecamines

produced 476 kt of copper and 14,5 kt of cobalt, 63,9 kt of zinc, 34,3 t of silver, and cadmium

and other minor metals (Source: Gecamines data).

The majority of this production came from the Kolwezi district. In 1995, production fell to

32,5 kt of copper 4 kt of cobalt, and 4,5 kt of zinc. The decline in metal production continued

to the point that primary production in the Kolwezi area virtually stopped.

Gecamines overall decline was due to a number of factors including:



The political isolation of the DRC (then Zaire) in 1991;

The loss of financial credit lines;

The lack of sustaining capital and maintenance improvements;

The social and political environment within the country during this period; and

The collapse of the Plateure in the central underground portion of the Kamoto Mine.

The Dikuluwe-Mashamba (“DIMA”) pit group operated from 1975 through 1998 during

which time a total of 57,7 Mt of ore grading 4,96% copper and 0,16% cobalt was mined

(Source: Gecamines). No significant production has come from T17.

8.4 Historical Exploration

Exploration work was undertaken by UMHK and Gecamines. The oldest hole on record still

retained by Gecamines is KTO2 (dated 13/07/1942), one of the original deep holes drilled for

the evaluation of the geology underlying the Kamoto Nord open pit. Numerous exploration

holes were drilled in the 1950s and 1960s for the underground operations of the Kamoto

Mine, with development beginning in the mid 1960s.

Prior to the exploitation of the DIMA pits, the surface area was drilled on a systematic grid

(usually 100 x 100 m or 100 x 50 m).

8.5 Historical drilling

8.5.1 T17 Mine

CCIC indicate that the T17 West deposit has been the subject of two diamond drilling

programs by Gecamines, with 3287,6 m drilled between 1938 and 1954, and 8011,3 m from

1986 to January 1988. The holes were drilled generally to a nominal 100 m x 100 m grid,

with certain areas being on 50 m x 50 m spacing.

8.5.2 Tilwezembe Mine

The historical drilling was undertaken by Gecamines, and the information was the basis for

the first Mineral Resources estimates for Tilwezembe Mine undertaken by SRK. The drilling

information included drill holes on a 25 m spacing within the operational Tilwezembe pit and

100 m x 50 m on Tilwezembe East.

8.5.3 Kamoto Mine

Gecamines carried out both extensive surface and underground drilling to delineate the

Kamoto ore bodies. 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 cubbies. A total of 569 holes have been identified, drilled between 1972 and

2002. The upper parts of the Etang – above 400 Level – 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. Underground collared boreholes have not been

surveyed.



8.5.4 Kananga Mine

The historical drilling at Kananga was limited to about eight holes drilled by Gecamines.

Details of the drilling, sample collection and sample analyses for the holes are not available.

8.5.5 KOV Mine

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 Mine pit.

In the 1980s, another drilling campaign was aimed at defining the KOV mineralized zones,

and this campaign continued into the early 1990s. The drilling was carried out along section

lines spaced about 100 m apart. Where feasible, drill holes were spaced about 100 m along

these section lines. The holes were prefixed KOV.

Most of the drill holes within the Kamoto East and the KOV Mine 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 drill hole intersections in Kamoto East were within the areas

that have been subsequently mined out, and there are very few orebody intersections below

the current pit bottom.

8.5.6 Mashamba East Mine

For the Mashamba East pit, drilling was undertaken from surface on a 100 m x 100 m spaced

grid on a local co-ordinate system parallel to the strike of the mineralized zone.

8.6 T17 Mine

T17 Mine is a new site, with no significant production history to date. Pre-stripping began in

May 2007 and the first blast was at the end of September 2007.

Table 8.1: T17 Mine: Historical Production

Production Units 2005 2006 2007 2008

Ore Mined (kt) - - 97,7 479,5

Cu Grade (%) - - 1,18% 1,72%

Co Grade (%) - - 0,47% 0,89%

Waste (kt) - - 4018,2 5405,8



8.7 Tilwezembe Mine

Mining has taken place intermittently since 1999. A rail siding and contractors yard were

established close to the site.

Predominant copper minerals are malachite and pseudo-malachite associated with the cobalt

mineral, heterogenite. The host rock is both dolomitic and siliceous. Copper and cobalt head

grades are reasonably well defined using both current and historical records from the

Gecamines geology database and from head grades of ore processed in the KZC concentrator.

The resulting oxide concentrate was leached and refined at the ‘Old’ Luilu and Shituru

refineries, located at Kolwezi and Likasi respectively, producing a “B” grade copper cathode

and cobalt metal. The mineralised zones in Tilwezembe also contain a high proportion of

manganese (which requires an adjustment to the processing for this ore).

Gecamines conducted open pit mining at Tilwezembe using contract mining on and off for a

period of about 7 years. Latterly, organised mining by Gecamines was replaced by artesinal

mining within the existing pit and along strike, until DCP recommenced mining in 2007. The

existing open pit is located at the western extremity of the ore body.

Table 8.2: Tilwezembe Mine: Historical Production

Production Units 2005 2006 2007 2008*

Ore Mined (kt) 52,0 109,5 480,3 609,8

Cu Grade (%) 3,90% 1,91% 1,72% 1,39%

Co Grade (%) 0,66% 0,81% 1,07% 1,17%

* Mining was discontinued during November 2008

8.8 Kamoto Mine

The official opening of the Kamoto Mine is given by Gecamines as 1942, with the beginning

of exploitation of the opencast resource in 1948 and the opening of the Kamoto underground

mineshaft in 1959.

Underground operations at the Kamoto Mine are accessed by twin declines, two primary

shafts and three secondary shafts. Primary access is through the declines and ore handling is

through the primary shafts from where crushed ore is transferred directly onto a conveyor to

the Kamoto concentrator.

Underground production, which began in 1969, used a variety of large-scale techniques

including cut and fill, room-and-pillar and sub-level caving. Production steadily increased to

reach the rate of 3 Mt/a by mid-1970. Production reached a peak in 1989, when the mine

produced 3,3 Mt of ore. In 1990, a major collapse in the central portion (the Plateure) of the

underground deposit resulted in the loss of approximately 15 Mt of Mineral Resource. Since

that time production from Kamoto Mine has steadily decreased to the point that primary

production essentially stopped. From 1969 throgh 2005 a total of 59,4 Mt of ore grading

4,21% Cu and 0,37% Co have been mined from Kamoto Mine.



Table 8.3: Kamoto Mine: Historical Production


Ore Hoisted (kt) 43,9 0 189,0 551,3

Cu Grade (%) 3,14% 0,00% 3,41% 3,93%

Co Grade (%) 0,30% 0,00% 0,44% 0,43%

8.9 Kananga Mine

The historical exploration by diamond drilling defined over a strike length of about 600 m

mineralized zones within the Kananga Hill. The ore is mainly oxide in nature with very little

sulphide material in the mineralogy. Predominant copper minerals are malachite and pseudo-

malachite associated with the cobalt mineral, heterogenite. The host rock is both dolomitic

and siliceous. Copper and cobalt head grades are reasonably well defined by current and

historical records from the Gecamines geology database and from head grades of ore

processed in the KZC concentrator. The resulting oxide concentrate was leached and refined

at the ‘Old’ Luilu and Shituru refineries, at Kolwezi and Likasi respectively, producing a “B”

grade copper cathode and cobalt metal.

The Kananga pit is close to the Dilala River and wetland and is also within 20 m of the

Lubumbashi-Lobito railway line.

Gecamines conducted open pit mining at Kananga using contract mining from around mid-

2004. Mining by Gecamines has been replaced by predominantly artisanal mining within the

existing pit and along strike. The existing open pit is at the western extremity of the ore body

and had been mined to a depth of approximately 20 m.

8.10 KOV Mine

KOV Mine comprises the four orebodies namely, Kamoto-East, Oliviera and Virgule (hence

the name KOV) and FNSR.

From the aerial photo of the KOV and Kamoto East pits (Figure 6.2), it can be seen that the

existing opencast mine workings are filled with water. A program to dewater the pits and the

surrounding slopes is under way. The KOV pit is in an area that is highly disturbed by past

mining activities. The old Musonoi pit, immediately east of KOV, has been partly back-filled.

Extensive accumulations of waste materials are present to the north and south of KOV.

Mining at the Musonoi pit commenced in 1943 and ceased production by 1984. The Kamoto

East pit started in 1959 and continued operations to 1985. Metallurgical records indicate that

the ore from the Kamoto East ore body was delivered to the plant from 1960 to 1985, while

ore production from the KOV pit commenced in early 1985 and carried on, at declining

production rates, through to 2000. During that time, some 38 Mt of ore was delivered to the

plant at an average grade of 5,8 % copper and 0,5 % cobalt.



Table 8.4: KOV: Historical Production

Period Source Plant Feed Head Grade (%) Contained Metal (kt)

(Mt) Cu Co Cu Co

1960-1985 Kamoto East 20 5,97 0,53 1194 106

1983-2000 KOV 18 5,56 0,48 1001 86

38 5,78 0,51 2195 192

8.11 Mashamba East Mine

Mashamba East Mine operated from 1985 through 1988 and the pit produced a total of 9,8 Mt

of ore at an average grade of 4,96% copper and 0,35% cobalt.

Production ceased in 1988 and by 1998, due to the lack of funds and increasing costs, the pit

was allowed to flood. A program to dewater the pits and the surrounding slopes is in place,

details of which are provided elsewhere in this report.

8.12 Kamoto Concentrator

The Kamoto concentrator consists of four sections: Kamoto 1 and 2 built in 1968 and 1972

respectively, and DIMA 1 and 2 built in 1981 and 1982. The Kamoto 1 and DIMA circuits

were designed to process mixed ore types, and Kamoto 2 was designed for sulphide ore. From

1969 through 2000, the Kamoto Concentrator processed over 126 Mt of ore at an average

grade of 4,33% copper and 0,28% cobalt. In its current configuration, the Kamoto

concentrator is capable of processing 7,5 Mt/a of ore. This throughput was exceeded from

1983 through 1987, with the production peaking at 7,6 Mt in 1985.

Table 8.5: Kamoto Concentrator: Historical Production


Oxide

Concentrate (t) (t) - - 1069 55 323

Cu Grade (%) - - 15,26% 16,90%

Co Grade (%) - - 0,64% 3,30%

Sulphide

Concentrate (t) (t) 4306 312 11 749 48 909

Cu Grade (%) 33,25% 32,26% 40,28% 40,80%

Co Grade (%) 2,22% 2,19% 5,86% 4,4%



8.13 Luilu Metallurgical Plant

The Luilu metallurgical plant is approximately 6 km north of the Kamoto Concentrator. It was

constructed in 1960. In 1972, it was expanded to an annual capacity of 175 kt of copper and

8 kt of cobalt. The plant has three roasters, a leaching circuit and electrolytic cells for copper

and cobalt production. From 1984 through 1989, annual production at Luilu averaged 173 kt

of copper and 5,9 kt of cobalt. Production peaked in 1986 at 177,5 kt of copper and 7,8 kt of

cobalt. By 1996, production had fallen to an estimated 27 kt of copper and 1,2 kt of cobalt.

Table 8.6: Luilu Metallurgical Plant: Historical Production


Cu Cathode (t) 12 229 6224 340 22 161

Co Cathode (t) - - - 749



9 Geological Setting

9.1 Regional Geology

The mineralized zones are at the western end of the Katangan Copperbelt, one of the great

metallogenic provinces of the world, and which contains some of the world’s richest copper,

cobalt and uranium deposits (Figure 9.1).

These deposits are hosted mainly by metasedimentary rocks of the late proterozoic Katangan

system, a 7 km thick succession of sediments with minor volcanics, volcanoclastics and

intrusives. Geochronological data indicate 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 extends for more than 300 km from Kolwezi in the north-

west to Lubumbashi in the south-east.

Stratigraphically, the rich copper and cobalt deposits found in Zambia and the DRC are

localized in the Roan Supergroup (“Roan”). The Roan occurs at the base of the Katanga

succession, unconformably overlying the basement rock of Kibaran age (mid-Proterozoic).

The Roan is separated from the overlying rocks of the Upper and Lower Kundelungu

supergroups by a conglomerate, the grand conglomerate. The Lower Kundelungu is

composed of sandstones and shales with a basal conglomerate, while the Upper Kundelungu

consists essentially of sediments and is separated from the Lower Kundelungu 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 of hundreds of

metres aligned across the fault plane.

In the Katangan Copperbelt, mining for copper and cobalt occurs in these outcropping to sub-

outcropping fragments.



Figure 9.1: Regional Geology



9.2 General Stratigraphy

The generalized stratigraphy of the Katangan System is shown in Figure 9.2. The Roan has

been correlated across the Katangan Copperbelt into four main formations or groupings, R1 to

R4. The divisions between each of the R series are often marked by an unconformity. The

main ore-body lithologies belong to the R2 Formation, but R3 and R4 Formations are also

known to contain mineralization. Within each of the R series are sub-divisions identifying the

different lithological units. Rocks belonging to the Roan Supergroup are described briefly

below from the oldest to the youngest:

Breche heterogene or heterogeneous breccia (BH): This breccia is composed of angular

and sometimes well rounded fragments of all the various rock types of the Roan. The

fragments vary in size from a few millimetres to several tens of millimetres in diameter, while

the matrix is made up of finer-grained sandy particles of the same material as the fragments.

Breche RAT or brecciated RAT (B RAT): A reddish-pink brecciated rock with calcite and

silica veinlets and is at times well mineralized with specular haematite, occurring as veinlets.

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). Both are

massive but sheared in places, silty or sandy, dolomitic rocks. Mineralization in the form of

malachite and black oxides occurs associated with the upper RAT.

Dolomie Stratifie or Stratified Dolomite (D Strat): This is a well-bedded to laminated,

argillaceous dolomite, which forms the base of the traditional “Lower Ore Zone” in

Gecamines’ nomenclature. The mineralization consists of copper and cobalt oxides.

Roches Siliceuses Feuilletées Foliated (Laminated) and Silicified Rocks (RSF): These are

grey to light-brown, thinly bedded laminated and highly silicified dolomites. The unit is

generally well mineralized with copper and cobalt oxides. Together with the D Strat, the RSF

comprise the Ore body Inferior (“OBI”).

Roches Silicieuses Cellulaires or Siliceous Rocks with Cavities (RSC): Vuggy and infilled

massive to stromatolitic silicified dolomites. Copper mineralization is almost absent in these

rocks, which were therefore regarded as barren. However, the infillings are enriched in wad

(manganese oxide) and heterogenite (cobalt oxide), and RSC is the target of artisanal activity.

Schistes De Base or Basal Schists (SDB): Reddish-brown to grey silty and nodular dolomite

to siltstone. This unit is well mineralized with copper and cobalt in varying amounts and

forms the Ore-body Superior (“OBS”).

Shales Dolomitiques Superieurs or Upper Dolomitic Shales (SDS): Yellowish, cream-to-

red, bedded laminated dolomitic siltstones and fine-grained sandstones. The rock is sparsely

mineralized with malachite.



Figure 9.2: General Stratigraphy of the Katangan System



Calcaire a Minerais Noirs or Calcareous Unit with Black Minerals (CMN): A slightly

banded and laminated light-grey to grey, silicified dolomite mineralized with black oxide of

iron, manganese and cobalt. The unit bears some similarities with the RSC.

Dipeta (R3): Greyish to dark red or brown stratified shales and micaceous schist.

Mwashya (R4): altered stratified greyish siliceous dolomitic rock with oolitic horizons and a

few bands of light-yellow, talcose schist. Nodules of haematite often occur.

9.3 Project geology

With the exception of Tilwezembe Mine, all of the mineralized properties of KOL 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 Mine is located about

20 km to the east of Kolwezi. Figure 9.1 shows the location of the deposits.

Within the Kolwezi Nappe, each of the project areas, T17 Mine, Kamoto Mine, KOV Mine

and Mashamba East Mine 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.

9.3.1 T17 Mine

The T17 West is described as dismembered structurally complex packages, which belong to

the southern flank of a synclinal fold that extends 2,6 km and is overturned towards the north.

Faulting is assumed to be the predominant process in the deformation and dismemberment of

the deposit.


The mineralized zone of Tilwezembe is located in an NE-SW anticlinal structural lineament,

which extends further to the east where 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 haematite and oolites. The strata strike

almost east-west and dips at about 45° to the south.

9.3.3 Kamoto Mine

The Kamoto underground operations extract mineralized copper ores from the Kamoto

Principal fragment, which is differentiated from the Kamoto East, mined in the KOV 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 flanks. Dips in the flank regions

are between 45° to 85°. The ore body is subdivided into four regions as follows:



The main central region, comprising Zones 1 to 8 and division 5. Commonly referred

to as the Principal;

Etang South;

Etang Nord; and

Ecaille Renverse.

9.3.4 Kananga Mine

The Kananga ore body outcrops and forms a ridge with a NNE strike. The ridge falls quite

rapidly towards the south and has been cut to form part of the embankment for the Lobito

railway line, which runs parallel to the ridge and 10 m to 20 m away from it for most of the

strike length of the ore body. Gecamines’ interpretations indicate that Kananga is the northern

limb of the Kananga-Dilala syncline, which plunges to the south.

9.3.5 KOV Mine

There are three main individual fragments hosting mineralized Lower Roan lithologies within

the KOV 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 pit. The FNSR lies below and is sub-parallel to the

Virgule ore body.

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 ore body 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 ore bodies, while the Oliveira fragment is a

shallower-dipping ore body in faulted contact with and below the Virgule ore body.


There is limited information on the geology of Mashamba East, except in the context of the

regional setting. Structurally, the lithologies of the Mashamba East 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.



10 Deposit TypesThe deposits fit in with the general description of stratiform with supergene enrichment

within the upper surface layers.

11 Mineralization

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,

chalcocite 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 100 m 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 Mine pit.

The RSC, a lithological unit stratigraphically intermediate between the OBS and OBI 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 Gecamines,

normally contains relatively less copper mineralization and was 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 Mine 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 manganite).

12 ExplorationNo exploration has been undertaken on behalf of the issuer other than in-fill drilling at

Tilwezembe Mine, Kananga Mine and KOV Mine.



13 Drilling

The project area contains mostly historical information from diamond drilling by the previous

owners, Gecamines. There has been limited drilling within the project area for the issuer and

only within Tilwezembe Mine, Kananga Mine and KOV Mine.

13.1 T17 Mine

There has been no recent drilling in T17 Mine.

13.2 Tilwezembe Mine

Snowden indicate that 152 diamond drill holes have been drilled since April 2006 on

Tilwezembe Mine. The drilling was undertaken by Remote Drilling Services (“RDS”) using

the wireline method with core sizes from PQ, HQ to NQ. A geologist from DCP monitored

the drilling daily.

The drill-hole spacing was about 50 m in the west and by 100 m in the east along strike and

along dip about 30 m the near the surface and 100 m at depth.

Only a portion of drill holes deeper than 100 m were down-hole surveyed by RDS, who used

an Eastman multi-shot down-the-hole camera. Snowden indicate that deviations were not

substantial and have little effect on the Mineral Resource estimation.

13.3 Kamoto Mine

There has been no recent drilling in Kamoto Mine.

13.4 Kananga Mine

At Kananga, 51 holes were drilled since April 2006, totalling 10 300 m. The drilling was

undertaken by RODIO (a South African drilling company) and supervised by Snowden. The

drill holes cover a strike length of about 550 m at a nominal spacing of about 50 m.

13.5 KOV Mine

There have been three phases of drilling within the KOV Mine pit, the first commencing in

late 2005 and the most recent being in 2007.

The 2005 drilling program was primarily to collect samples for metallurgical test work and

also confirm the general ore body intersections and grade. A program of 8 metallurgical and

15 confirmatory holes was initiated. The drilling program ran into difficulty; access was

difficult and the geotechnical conditions in the near surface rocks was poor. Only 8 holes

were completed, and they were used for metallurgical scoping test work.

The 2007 program was to provide information for geotechnical purposes both inside and to

the west and north-west of the KOV Mine pit. The program consisted of 25 vertical holes

penetrating the Virgule and Oliveira mineralized zones. The samples of the core from all rock

types were sent for geotechnical testing while the mineralized intersections were sent off to



Alfred H Knight Laboratory (accredited to SANAS, Facility Accreditation Number: T0141)

for analyses.

A third program of 15 holes was laid out to provide material for quality assurance and quality

control (“QA/QC”). The holes were planned to twin the Gecamines holes and thereby verify

the historical drilling information. However, due to problems of access, the holes were not

drilled at their intended positions.

A few holes were twinned, and the data available for the comparisons show good

correspondence in the thickness and %TCu grade in the SDB and less so for the RSF and

DSTRAT intersections. The differences in the latter could be ascribed to the geological

identification of the lithologies on which the selection is based for the comparisons. Figures

13.1 to 13.3 provide further details.

Figure 13.1 KOV Mine: Comparisons of the Twin Hole Intersections, Thickness and%TCu grade by Lithology - SDB

0.0

5.0

10.0

15.0

20.0

25.0

30.0

35.0

40.0

0.00

1.00

2.00

3.00

4.00

5.00

6.00

7.00

8.00

9.00

KOV461 KOV474 KOV475 KOV479 KOV484 KOV547 KOV558

THIC

KN

ESS,

m

%Tc

u

TWIN DDH

SDB

ORIGINAL TWIN THICK-ORIG THICK-TWIN

TH

ICK

NE

SS

,m



Figure 13.2 KOV Mine: Comparisons of the Twin Hole Intersections, Thickness and%TCu grade by Lithology - RSF

Figure 13.3 KOV Mine: Comparisons of the Twin Hole Intersections, Thickness and%TCu grade by Lithology - DSTRAT

13.6 Mashamba East Mine

There has been no recent drilling in Mashamba East Mine.

0.0

1.0

2.0

3.0

4.0

5.0

6.0

7.0

8.0

9.0

10.0

0.00

1.00

2.00

3.00

4.00

5.00

6.00

7.00

8.00

9.00

KOV461 KOV474 KOV475 KOV479 KOV484 KOV547

THIC

KN

ESS,

m

%Tc

u

TWIN DDH

RSF


0.0

1.0

2.0

3.0

4.0

5.0

6.0

0.00

1.00

2.00

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7.00

8.00

KV 461 KV 474 KV 475 KV 479 KV 484

THIO

CK

NES

S,m

%Tc

u

TWIN DDH

DSTRAT


TH

ICK

NE

SS

,m

TH

ICK

NE

SS

,m



14 Sampling Method and Approach

14.1 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.

SRK has described the observations on the historical sampling for KOV Mine, and the review

of the individual projects reports has indicated that the process described for KOV Mine was

applicable to the other projects. SRK’s review of the historical sampling for KOV Mine is

summarized below.

Cores from the ore body intersections were sampled for chemical analysis. The lengths of

core sampled varied, and it is SRK’s understanding that this was a consequence of the sample

recovered within each run. In the Gecamines 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

“sterile” 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.

Due to this pre-selection during sampling, the assay database is incomplete, and this affects

the accuracy of the Mineral Resource estimation.

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.

14.2 T17 Mine, Kamoto Mine and Mashamba East Mine

CCIC undertook a re-sampling of the existing cores for the T17 Mine, Kamoto Mine and

Mashamba East Mine. CCIC reported that 26 historical holes were re-sampled, and these

included the following:

DIK 171 (Mashamba East)

F2418 (Kamoto Principal OBS )

F2471 (Kamoto Principal OBI)

F2391 (Kamoto Etang)

MU321 (Musonoi-T17).

CCIC’s approach was to try as much as possible to replicate Gecamines sampling and assay

results on cores from each of the project resource areas to gain confidence in the reported

assay values. Details of the original Gecamines sampling protocol and methodology were

unavailable but, by inference, CCIC contend that samples were taken for assay based mainly

on lithology, and as such were of irregular lengths. CCIC studied the sample lengths and

found that the most popular sample length was between 1,5 and 2,0 m, although sample

lengths ranged from 0,02 m to 10,0 m.

CCIC replicated the sampling intervals of the original log sheets by quartering the existing

half-cores. The remaining quarter core of the sample was put back in the sample box and

remarked with the original sample number.

The quarter core samples were marked, cut and bagged under CCIC’s supervision at the

Kamoto Geological Department. Samples were cut by CCIC or Gecamines personnel on a

Wendt L18A B61936 saw with a 340 mm diamond blade.

A total of 654 samples were generated and dispatched to SGS Lakefield Research Africa (Pty)

Ltd (“SGS Lakefield”) (accredited to SANAS, Facility Accreditation Number: T0169) for

analysis.

14.3 Kananga Mine and Tilwezembe Mine

The procedures adopted during the sampling of the cores from Kananga and Tilwezembe are

similar and are described in Section 15.2.

After the drill-hole core was photographed and logged, the core was split with a diamond saw

and sampled at 1 m intervals within the mineralized units, honouring geological contacts.

14.4 KOV Mine

There have been three phases of drilling in the KOV pit since 2005. The initial drilling of 8

holes was undertaken over 6 months commencing in November 2005 to collect material for



metallurgical test work. A second phase of drilling was undertaken for geotechnical purposes

during 2006/07, and a third phase of confirmatory holes was drilled in 2007.

The phase 1 drilling was supervised by SRK. At the start of the drilling program, SRK

devised a core-sampling protocol. The protocol outlined the procedures to be followed, and

they included a review of the core to determine the contacts, which were to be used as a guide

during the sampling program. The protocol ensured that samples were taken at regular

intervals of approximately 1 m within the lithology as dictated by the core and the core

recovered. It also ensured that samples did not to straddle lithological contacts. Additional

samples were taken well outside the zones of visible mineralization.

The core from phase 2, phase 3, geotechnical, and confirmatory drilling was sampled under

DCP’s supervision.

15 Sample Preparation, Analyses and Security


CCIC indicate that the existing half samples were quartered using a diamond core-cutting

blade, washed, flagged with a unique sample number, and bagged. The samples were collated

into larger bags, per drill-hole, locked in a metal trunk and shipped to SGS Lakefield in South

Africa.

In all, 58 samples were taken from various boreholes for independent verification of the

Gecamines copper and cobalt assay figures for T17 Mine, Kamoto Mine and Mashamba East

Mine. These samples were sent to SGS Lakefield for preparation and analysis, with

renumbered pulps resubmitted to both SGS Lakefield and Set Point Laboratories (“Set Point”)

(accredited to SANAS, Facility Accreditation Number: T0223) as checks. The cut samples

were placed in metal containers and sealed under lock and key under the supervision of CCIC

personnel. The samples were then trucked to the client’s Lubumbashi offices before being

sent to SGS Lakefield.

CCIC indicate that SGS Lakefield staff entered each of the samples into their LIMS system,

which includes the client’s details, the list of samples and the analyses required. Each of the

quartered core samples was crushed to <2 mm, and the crushed sample was split, where

necessary, to produce a portion of about 250 g. The split (or entire crushed sample if less than

250g) was milled, bagged, labelled and stored in a box(es), logged in Lakefield’s sample-

tracking system and stored on a shelf.

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 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.

Appropriate QC checks were undertaken by CCIC on Kamoto Mine.


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. Preparation

activities consisted of:

Drying of sample;

Primary jaw and roll crushing of sample;

Splitting a sub sample of 250 g using a riffle splitter; and

Pulverizing of the sub-sample to 75 micrometres and homogenizing.

The samples were analyzed for %TCu, %TCo, %Mn, %AsCu and %AsCo.

Both Alfred H 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.

For %AsCu and %AsCo, both laboratories used cold leaching with 5% sulphuric acid.

However, Alfred H 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.

Appropriate QC checks were undertaken by CCIC on Kananga Mine and Tilwezembe Mine.

15.3 KOV Mine

All historical sampling, sample preparation, analysis and security were undertaken by

Gecamines over more than 50 years. SRK are unable to comment on the quality of such work.

The core sample was cut along the longitudinal axis with one half of the core sent for

laboratory analysis and the other 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 0,5 m and the maximum

sample was 2,5 m. The sample lengths were also a consequence of the sample recovered

within the run.

Analysis and QC were undertaken in-house by Gecamines.



Historical production from KOV was 38 Mt delivered to the plant at an average grade of 5,8%

copper and 0,5% cobalt. This compares favourably with the current Mineral Reserve grade of

4,93% Cu and 0,38% Co.

The phase 1 drilling was supervised by SRK, detailes of which are provided in Section 14.4.

The core from phase 2, phase 3, geotechnical, and confirmatory drilling was sampled under

DCP’s supervision.

16 Data Verification


CCIC captured the drill hole data from the hard-copy log sheets into Microsoft Excel

spreadsheets. A parallel entry system was set up by Maxwell GeoServices (“Maxwell”), using

a data-management system that allowed for the standardization of data capture and storage in

a database. The electronic copies were verified and validated by comparing with the original

hard-copy logs.

In all, 363 logs were supplied as scanned images of the original paper logs. These logs were

captured into the existing Excel database created during the Engineering Study. The

verification process involved comparisons of the electronic logs against the paper logs. Every

collar and survey data point was verified. The stratigraphic and sample logs were routinely

verified. During the validation process, however, any dubious or erroneous holes were

thoroughly verified. The validation process undertaken by CCIC involved the following:

Checking for any zero lengths, gaps, overlaps and duplicate entries in the Excel database

and verifying data against the hard-copy original logs;

3D visual validation, using Datamine™ Studio to validate collar and survey information;

Visual flagging of the lithology using surrounding drill hole to ensure consistency in the

logged stratigraphic succession across sections; and

Histograms plots of the copper and cobalt values were generated to identify and verify

any outliers.

During sampling for the 26 historically drilled holes, 654 samples were generated and

dispatched to SGS Lakefield for analysis. The analytical method used for the determination of

total copper and cobalt was X-ray fractionation. Copper and cobalt oxides were determined by

acid digestion (sulphuric acid) and analysis of the solution by AAS.

The core QC audit program consisted of diamond-drill core samples selected from each

resource area and stratigraphic interval. The program was designed to check the accuracy of

the recorded Gecamines copper and cobalt grades. Samples were prepared from the half core

remaining after preparation of the original samples assayed by Gecamines. Initial check

samples were run at SGS Lakefield with additional splits and checks completed by SGS

Lakefield and Set Point Laboratories. The work was carried out in accordance with

compliance procedures that addressed sample preparation, security, laboratory qualification,

and procedures.



The assay laboratories at Luilu were visited and copies of the procedures and protocols were

requested; however, they were not provided. It is therefore not clear if the replicate samples

exactly duplicated the methods by which most of the historical assaying was completed.


Snowden describe similar process for the validation of the database for Kananga and

Tilwezembe. The drill-hole data were stored in Excel spreadsheets, and Snowden imported

the data into a GEMS Access database where the following validation checks were

conducted:

Missing collar coordinates;

Interval errors (missing intervals, overlaps etc.) within drill-hole sample data;

Duplicate sample records;

Zero values within the sample data; and

Collar elevation errors.

Data validation included checking for cases where the %AsCu and %AsCo were greater than

the %TCu and %TCo values. At Tilwezembe Mine, cases were found where the %AsCu

values were greater than the %TCu and 30 cases for the %AsCo values being greater than the

%TCo out of a total database of 3000 samples. At Kananga Mine, the numbers were 20

samples and 14 samples respectively out of a total database of 2370.

Snowden indicate that difference in value between the acid soluble and the total grade were

small with about 18% of the copper and 17% of the cobalt datasets reporting differences

higher than 0,1% at Tilwezembe and at Kananga the number was 6% and 0% for the copper

and cobalt respectively reporting differences higher than 0,1%.

For the purpose of estimation, soluble assays greater than total assays were set equal to the

total assay.

16.3 KOV Mine

SRK has reviewed the database and the following issues with data quality are highlighted:

Poor core recovery within the ore body varying between 65 and 75%, the worst recovery

being in the RSC lithology;

The cutting of the total copper grade to 12%, if above 12%. However, not all the samples

with total copper grades above 12% in the database have been cut;

The RSC formation in contact with the SDB and the RSF was sampled, but near the

middle of the formation selectively sampling was on the basis of visible copper

mineralization;

The ore-body zones were sampled for total copper and total cobalt, but assays for acid

soluble copper and cobalt and for calcium oxide were limited; and



Samples with zero core recovery, mostly from the earlier drilling in Kamoto East, are

shown with high total copper values; presumably from the analyses of the sand collected

in the absence of core.

SRK examined selected drill-hole core but did not resample the core in any way. The KOV

pit was in production for over 20 years, and head grades during that period reflected the

grades in the database.

17 Adjacent Properties

There has been significant exploration and mining activity in the Kolwezi area. Copper

mining has been undertaken in the region for many decades.

Regional copper mining companies in the DRC include:

First Quantum Minerals Limited: Based on the Annual Information Form at 31

December 2008 using a 0,5% copper cut-off grade First Quantum Minerals Limited had

Measured Mineral Resources of 69,1 Mt at 1,39% Copper and Indicated Mineral

Resources of 175,1 Mt at 0,99 % Copper; and

Anvil Mining Limited: Based on the Annual Information Form at 31 December 2007

using a 0,5% copper cut-off grade Anvil Mining Limited had:

o Measured and Indicated Mineral Resources of 1,1 Mt at 7,01 % Copper at the

Dikulushi deposit;

o Indicated Mineral Resources of 8,0 Mt at 1,90 % Copper and 0,11% Cobalt at the

Kulu deposit; and

o Measured and Indicated Mineral Resources of 33,3 Mt at 3,68 % Copper at the

Kinsevere deposit.

18 Mineral Processing and MetallurgicalTesting

18.1 Kamoto Mine Testwork

The process design criteria assume that the ore from Kamoto underground will be the same as

that mined previously in terms of its milling, flotation, leach and other metallurgical

characteristics. Being a brownfield refurbishment this is not an unreasonable assumption

given the availability of 39 years’ historical records. However, in SRK’s view, it would have

been preferable to independently confirm the metallurgical characteristics of future ore (as

opposed to previously mined) in a programme of testwork. Prior to this study, some work

was done by Hatch to this end.



18.2 Previous KOV Testwork

18.2.1 Samples Tested

The process design criteria mostly assume that the ore mined from KOV Mine will be the

same as the sample used for the DCP testwork at Mintek in 2006, in terms of its milling,

flotation, leach and other metallurgical characteristics but different in terms of sulphide

copper content. The DCP testwork utilised bulk ore samples of the following five lithological

units identified at KOV Mine taken from a stockpiles of ore mined just before the mine closed

in 1999.

SDB surface material consisting mainly of oxidised material referred to as SDB Ox

but also a sulphide fraction called SDB Sulph;

RSC;

RSF;

D Strat; and

Rat Grise, the deepest material.

Such samples would presumably have been representative of the material being mined at the

time but, in the light of the anticipated change in the proportion of sulphides, such samples

are unlikely to be fully representative of future ore to be mined. In addition it is uncertain to

what extent the metallurgical characteristics of the stockpiled ore had changed since being

mined. In SRK’s view it would have been preferable to have conducted testwork on fresh

samples of ore to be mined over the life of project. At the time the difficulty of obtaining such

representative samples, due to the pit being flooded, precluded this.

18.2.2 Milling Testwork

JKTech (Pty) Ltd (“JKTech”) were commissioned to conduct drop weight testing of the five

ore types and to simulate single-stage SAG milling versus primary SAG mill and secondary

Ball milling. Generally all material types were characterised as being “soft to very soft” in

terms of resistance to impact breakage and abrasion breakage. Rat Grise however showed

evidence of bimodality in its relative density distribution, which strongly suggests that a

dense component could concentrate in the mill load and compromise mill performance. It will

therefore be important to ensure a good blend of feed materials to minimise such effects.

JKTech simulation results concluded that a two-stage SAG and ball mill circuit was more

efficient than a single stage SAG circuit. The two-stage circuit was accordingly incorporated

into the Nikanor flowsheet.

18.2.3 Hydrometallurgical Testwork

Mintek undertook a programme of testwork including mineralogy and copper and cobalt

hydrometallurgical processing. The test programme was conducted on a composite sample of

the five identified lithologies blended pro rata to the depth of the lithological units. Copper

and cobalt leaching, copper solvent extraction (“SX”) and copper electrowinning (“EW”)



were conducted at bench and pilot scale whilst cobalt purification and precipitation

investigations were conducted at bench scale on pilot-generated and synthetic solutions.

The leach was conducted in two steps, with pH being maintained with sulphuric acid addition

in the first part and the redox potential being controlled in the second part by introducing

sulphur dioxide gas. The pilot plant achieved fairly consistent copper and cobalt leaching

efficiencies up to 92% and 91% respectively. During pilot testing, the SX circuit was

simplified to comprise 2 extraction, 1 wash and 1 strip stages. Cathode produced in the pilot

plant achieved the desired LME Grade ‘A’ quality of >99.95% copper, although attention will

have to be given to certain impurities during full-scale operation.

Investigations into the purification and precipitation of cobalt bleed solution included the

following steps:

Fe/Mn removal with air/SO2;

Removal of aluminium and copper via precipitation with lime;

Calcination testwork on the final Co(OH)2 product with lime, and

Precipitation of Co(OH)2 salt using MgO.

The testwork identified optimum conditions for the removal of iron, manganese, aluminium

and copper. Properly controlled, cobalt losses should not exceed 1% in Fe/Mn precipitation

and 2% in Al/Cu precipitation.

Initial tests were conducted with an industry recognised MgO, whilst optimisation tests were

conducted with an alternate MgO that was preferred for the project. Complete cobalt

precipitation was achieved with MgO. Unfortunately MgO as a precipitant provided no

selectivity for cobalt over nickel, zinc and copper present in the feed, underlining the need to

remove these effectively during the purification steps. Furthermore it was not possible to

prepare solids with the desired composition of approximately 40% cobalt with <2% co-

precipitated magnesium. It is suspected that this was due to very slow kinetics displayed by

the alternate MgO and further tests using the industry recognised MgO were recommended.

Such tests have since been successfully completed.

18.2.4 Other Testwork

For all stages of precipitation, liquid/solid separation testwork was done on fresh

solutions/slurries to enable sizing specification of filters and thickeners. Mintek also

performed various corrosion tests under selected process conditions which allowed the

optimum material for the fabrication of vessels and equipment to be identified.

18.3 Recent KOV Mine Testwork

18.3.1 Milling Testwork

Mintek was commissioned to further investigate the amenability of the KOV Mine ore to

autogenous milling, with a view to using the two existing DIMA AG mills at the Kamoto

Concentrator rather than the new SAG and ball mills ordered by Nikanor. Testwork was

carried out on drill core samples from 4 ore zones, namely RSC, OBI, OBS-SDB and OBS-



SDS in the KOV Mine pit. Based on the ratio of Bond Rod Mill Work Index to Bond Ball

Mill Work Index, it was again concluded that the ore is not amenable to fully autogenous

milling. This was in line with the earlier JKTech investigations. However, it was also

concluded that the DIMA mills could achieve the required KOV ore throughput if converted

to SAG milling in series with the existing ball mills.

18.3.2 Flotation Testwork

Flotation testwork for the DCP/Nikanor project was originally done at Mintek on a leach

residue sample of a largely oxidised ore. It is now anticipated that future KOV Mine ore will

be increasingly sulphidic and a further programme of testwork has recently been carried out at

Mintek in order to finalise process design criteria for the new sulphide float ahead of whole

ore leach. This work has confirmed the preliminary flotation plant design used by Bateman

Engineering in the study and also confirmed the preliminary reagent suite.

18.3.3 Slurry Pumping Testwork

Paterson & Cooke, slurry pumping consultants, were commissioned to investigate the

pumping of KOV Mine milled oxide ore from the Kamoto Concentrator to the Luilu

Metallurgical Plant. They conducted testwork on a sample of ore extracted from the same

bulk material tested at Mintek. In addition they reviewed the tailings pumping system. A

system comprising running/standby pumps and two lines (running/standby) for the full flow

was identified as being the optimal system and was included in the 2008 Study.

18.4 Oxide / Sulphide Content

Oxides and sulphides require different processing routes. It has been assumed that the

oxide:sulphide ratio for the ore mine from KOV Mine is 60%:40%. Should this ratio be

significantly different to this assumption, it will affect the capacity of each process route and

may restrict production, which will affect sulphur consumption and costs.

18.5 Risks and Recommendations

The primary risk relates to the use of limited and old (1999) samples, which may influence

the reliability of the testwork.

It is recommended that further testing is undertaken, which should include samples from all

the mining properties in the appropriate ratios.



19 Mineral Resource and Mineral ReserveEstimates

19.1 Mineral Resource Estimates

19.1.1 Mineral Resource Estimation Methodology

Mineral Resource models from which Mineral Resources are quoted in this report were

generated by the following independent consultants:

T17 Mine CCIC

Tilwezembe Mine Snowden

Kamoto Mine CCIC

Kananga Mine Snowden

KOV Mine SRK

Mashamba East Mine CCIC.

T17 Mine and Mashamba East Mine were re-estimated by SRK as part of this study.

SRK has the overall responsibility for the sign-off on the Mineral Resources and Reserves

and as part of that process has reviewed the methods adopted in the generation of the Mineral

Resource models from the respective consultants. SRK’s review of the methods indicates

similarities of approach between the various consultants, and the processes are summarized

below:

Collate all the Gecamines 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 RATGR, DSRAT and RSF and the OBS

within the SDB, BOMZ and to a lesser extent, the SD1a. 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 0,5%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 SAMREC

Code.

19.1.2 Data quality and quantity

Mashamba East Mine and T17 Mine

The mineralized zones of the T17 Mine and the Mashamba East Mine have been drilled on a

nominal spacing of 100 m x 100 m. Additional holes in the T17 deposit in selected areas

reduced the spacing to a 50 m x 100 m grid. Refer to Figure 19.1 and 19.2 for drill-hole data.

Figure 19.1 Mashamba East: Drill-hole Location Plan, Geology and Pit Outline

432

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307000 N307000 N

307200 N307200 N

307400 N307400 N

307600 N307600 N

307800 N307800 N

308000 N308000 N

308200 N308200 N

308400 N308400 N

DIK 133

DIK 140

DIK 141

DIK 142

DIK 143DIK 144

DIK 145

DIK 146

DIK 147

DIK 148

DIK 150

DIK 151

DIK 153

DIK 154

DIK 161

DIK 162

DIK 163DIK 164

DIK 165

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DIK 167

DIK 168

DIK 172

DIK 631

DIK 632

DIK 639

DIK 640

DIK 645

DIK 646

DIK 647

DIK 649

DIK 650

DIK 652

DIK 654

DIK 656

DIK 658

DIK 660

DIK 662

DIK 663

DIK 664

DIK 665

DIK 667 DIK 681

DIK 708

DIK 709

DIK 713

DIK 716

DIK 717DIK 721

DIK 722 DIK 726

DIK 728

DIK 731

DIK 732

DIK 733

DIK 735

DIK 736

DIK 737

DIK 740

DIK 741

DIK 742

DIK 743

DIK 744

DIK 747

DIK 748

DIK 749

DIK 753

DIK 754

DIK 757

DIK 758

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DIK 765

DIK 766

DIK 768

DIK 769 DIK 770

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DIK 800

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DIK 805

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DIK 826

DIK 829

DIK 830

DIK 831

DIK 832

DIK 833

DIK 835

DIK 836

DIK 837 DIK 838

DIK 839

DIK 840

DIK 841

DIK 842

DIK 843

DIK 844

DIK 845

DIK 846

DIK 847



Figure 19.2 T17 Mine: Drill-hole Location Plan and Geology

Kananga Mine and Tilwezembe Mine

The drill-hole spacing intersecting the Kananga and Tilwezembe mineralized zones is on an

average spacing of 100 m along strike. Refer to Figure 19.3 and 19.4 for drill-hole data.

Figure 19.3 Kananga Mine: Drill-hole Location Plan

-1100

E

-110

0E

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-1200 N-1200 N

-1100 N-1100 N

-1000 N-1000 N

-900 N-900 N

-800 N-800 N

-700 N-700 N

-600 N-600 N

-500 N-500 N

MU

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8817500 N8817500 N

8817600 N8817600 N

8817700 N8817700 N

8817800 N8817800 N

8817900 N8817900 N

8818000 N8818000 N

8818100 N8818100 N

KNGW01KNGW02

KNGW03

KNGW04KNGW05

KNGW06

KNGW07

KNGW08

KNGW09

KNGW10

KNGW11

KNGW12

KNGW13KNGW14

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KNGW16

KNGW17KNGW18

KNGW19

KNGW20

KNGW20B

KNGW21

KNGW22 KNGW23

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KNGW27KNGW28

KNGW29KNGW30

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KNGW41 KNGW42

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KNGW48

KNGW49

KNGW50

KNGW51



Figure 19.4 Tilwezembe Mine: Drill-hole Location Plan, Geology and Pit Outline

KOV Mine

The KOV drill-hole database contains a total of 214 drill-holes spaced on average about

100 m x 100 m. 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 Gecamines 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.

There is adequate drill-hole coverage for the Virgule and Oliveira while the FNSR and

Kamoto East intersections are limited. 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 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.

356400 356600 356800 357000 357200 357400 357600 357800 358000 358200 358400 358600 358800 359000 359200 359400

Eastings

8805600

8805800

8806000

8806200

8806400

No

rth

ing

s



Figure 19.5 KOV Mine: Drill-hole Location Plan and Surface Topography

19.1.3 Core Recovery

Where possible and based on the data as captured for the various projects, the drill-hole data

include a column in which is recorded the sample length or percentage core recovery in

relation to the sampled interval. Core recovery describes the quality of the sample data being

used in the Mineral Resource estimates and has an effect on the quality of the Mineral

Resources reported. Low core recoveries can be due to a cavity or a consequence of bad

ground conditions or drilling practices. For the data to be representative, a core recovery in

the mineralized zones of at least 90% is often considered necessary.

Core recoveries within each of the project areas are indicated by project.

325800 326000 326200 326400 326600 326800 327000 327200 327400 327600 327800 328000 328200

Easting

8814200

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ort

hin

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Gecamines

2006 Met holes

2007 Evaluation

2007 Geotech

Drilling Campaigns



T17 Mine

Core recovery data by lithology for T17 Mine is contained in Table 19.1.

Table 19.1 T17 Mine: Core Recovery Data by Lithology

Variable RSF RSC SDB BOMZ

Minimum 1,25 12,16 10,81 18,18

Maximum 100,00 100,00 100,00 100,00

Average 56,92 61,66 62,69 69,22

Standard Deviation 24,07 27,23 27,71 26,66

Coefficient of Deviation 0,42 0,44 0,44 0,39

Tilwezembe Mine

Snowden report the recoveries in Table 19.2 in the mineralized zones at Tilwezembe.

Table 19.2 Tilwezembe Mine: Recoveries within the Mineralized Zones

Rock type Length Recovered Length Recovery

(m) (m) (%)

Manganiferous Dolomites (Oxides) 1435 910 63

Brecciated Material (Oxides) 639 536 84

Tillites and Argillites (Oxides) 867 754 87

Manganiferous Dolomites (Sulphides) 553 488 88

Brecciated Material (Sulphides) 360 343 95

Tillites and Argillites (Sulphides 428 412 96

Kamoto Mine

There are limited sample entries within the mineralized zones, with core-recovery data in the

Kamoto Mine database file presented to SRK. SRK extracted core-recovery data by region,

and the proportions of sample data with core recovery entries are indicated below:

Principal Region: 0,3% of sample data have core recovery captured;

Etang Nord: 11% of sample data have core recovery captured; and

Etang South: 64% of sample data have core recovery captured.

The Principal and Etang Nord regions were not studied, as the data density was low.

Approximately two thirds of the Etang South data contain core-recovery figures, and a study

of this region was made. Sample entries with core recovery within the mineralized zones were

extracted and analyzed statistically. The statistics are shown in Table 19.3. The CCIC reports

reviewed do not state whether core recovery was taken into account during the estimating and

classification of the Etang South Mineral Resource.

Table 19.3 Kamoto Mine: Core Recovery Data by Lithology

Minimum Maximum Average Standard Deviation

0 108,4 65,06 29,85



Kananga Mine

Snowden report the recoveries in Table 19.4 in the mineralized zones at Kananga.

Table 19.4 Kananga Mine: Recoveries within the Mineralized Zones

Rock type Length Recovered Length Recovery

(m) (m) (%)

Upper ore body oxides (UOB_OX) 363 312 86

Middle low-grade oxides (MID_OX) 679 492 73

Lower ore body oxides (LOB_OX) 380 318 84

Upper ore body sulphides (UOB_SL) 149 148 100

Middle low-grade sulphides (MID_SL) 287 286 100

Lower ore body sulphides (LOB_SL) 279 277 99

Total 2136 1834 86

KOV Mine

Core recovery data for KOV in each of the fragments are shown in Table 19.5.

Table 19.5 KOV Mine: Recoveries within the Mineralized Zones

Zone Lithology Minimum Maximum Average Standard Deviation

BOMZ 7,20 65,00 46,24 20,21

SDB 0,00 73,00 26,02 28,81

KMT RSC 0,00 73,00 14,96 22,62

RSF 0,00 100,00 19,06 25,49

DSTRAT 0,00 100,00 19,40 27,17

RATGR 0,00 95,56 19,46 26,75

BOMZ 0,00 105,00 29,40 39,21

SDB 0,00 115,00 47,07 40,37

VRG RSC 0,00 116,20 37,21 37,77

RSF 0,00 109,80 40,59 38,63

DSTRAT 0,00 105,00 45,07 38,56

RATGR 0,00 103,60 26,47 37,63

BOMZ 0,00 100,00 66,60 37,76

SDB 0,00 115,03 66,95 31,33

OLV RSC 0,00 108,44 31,24 35,84

RSF 0,00 103,00 53,80 32,60

DSTRAT 0,00 100,00 41,42 34,04

RATGR 0,00 95,00 27,78 29,85

BOMZ 52,00 112,50 83,52 19,22

SDB 0,00 129,52 47,73 36,84

FNSR RSC - - - -

RSF 0,00 108,40 56,57 40,20

DSTRAT 14,00 103,13 67,18 31,67

RATGR 83,00 94,12 88,56 5,56



Mashamba East Mine

Core recovery data by lithology for Mashamba East Mine are shown in Table 19.6.

Table 19.6 Mashamba East Mine: Core Recovery Data by Lithology

Variable RSF RSC SDB BOMZ

Minimum 13,33 29,79 29,17 19,15

Maximum 100,00 100,00 100,00 100,00

Average 67,43 69,63 72,81 73,58

Standard Deviation 19,90 21,75 20,46 22,67

Coefficient of Deviation 0,30 0,31 0,28 0,31

Summary

The bulk of the data within the project areas is historical, with the exception of Kananga Mine

and Tilwezembe Mine where drilling was undertaken recently and the historical data have not

been used.

The historical data are from the early 1940s at the earliest and the late 1980s at the latest.

SRK reviewed selected cores on site for KOV, Kananga and Tilwezembe and found that

sample recoveries within the mineralized zones were generally low and the core conditions

were variable.

SRK consider that 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.

19.1.4 Data manipulation

T17 Mine, Kamoto Mine and Mashamba East Mine

CCIC adopted the following criteria for the manipulation of the data at T17 Mine, Kamoto

Mine and Mashamba East Mine:

Drill-holes not sampled were removed from the database; and

Unsampled sections of the drill hole sandwiched between adjacent samples were assigned

a threshold value of 0,05%TCu and 0,01%TCo. This was more common in the RSC.



KOV Mine

The sample database contains sample intervals for which there are no sample values. These

intervals were not sampled by Gecamines as there was no visible copper mineralization and

the majority refer to the mid-portion of the RSC, which was considered barren. There are also

smaller intervals in between sampled intervals that were not sampled either because of core

loss or because they were considered to be not mineralized.

The non-sampled interval between two sampled intervals was assigned the weighted average

grade of the two samples adjacent to it. The criteria for this manipulation were the presence of

sampled intervals on either side and the fact that the non-sampled interval was less than 3 m

in length. Intervals greater than 3 m were not manipulated using the above criteria.

In the RSC, the mid-portion, which occupies about 10 to 15 m of its 25 m thickness, was

sampled only in very few instances. This was probably because of the absence of visible

copper mineralization.

The mid-RSC is differentiated from the rest of the RSC by a wireframe model defining the

extent of the unsampled section within each fragment. Where the drill hole was sampled

through the entire RSC lithology, the position of the mid-RSC wireframe was defined from

the surrounding holes with unsampled intervals. The mid-RSC wireframe was used to extract

the partial and complete sampled intervals within each of the drill holes inside the wireframe.

The drill-holes with complete sampling information in the mid-RSC were selected, and

weighted average grades for %TCu and %TCo were obtained for each fragment. These

weighted average values were assigned to the mid-RSC portion.



19.1.5 Grade distributions

T17 Mine

Figure 19.6 shows details of grade distributions at T17 Mine.

Figure 19.6 T17 Mine: %TCu grade distribution in DST, RSF and DB respectively(longitudinal view)

-1000 -950 -900 -850 -800 -750 -700 -650 -600 -550 -500 -450 -400 -350 -300 -250 -200 -150 -100 -50

X-Coordinate

1000

1050

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1250

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1350

1400

1450

1500

Z-C

oo

rdin

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%TCu

0.00 to 0.50

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1.00 to 2.00

2.00 to 3.00

3.00 to 4.00

4.00 to 5.00

5.00 to 7.50

7.50 to 20.00T17 PIT AT END MARCH 2008

SURFACE TOPO

LEGEND

-1000 -950 -900 -850 -800 -750 -700 -650 -600 -550 -500 -450 -400 -350 -300 -250 -200 -150 -100 -50

X-Coordinate

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1.00 to 2.00

2.00 to 3.00

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4.00 to 5.00

5.00 to 7.50


SURFACE TOPO

LEGEND

-1000 -950 -900 -850 -800 -750 -700 -650 -600 -550 -500 -450 -400 -350 -300 -250 -200 -150 -100 -50

X-Coordinate

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0.00 to 0.50

0.50 to 1.00

1.00 to 2.00

2.00 to 3.00

3.00 to 4.00

4.00 to 5.00

5.00 to 7.50


SURFACE TOPO

LEGEND



Tilwezembe Mine

Figure 19.7 depicts a longitudinal section along the axis of the Tilwezembe Mine ridge and

shows the positions of the drill-hole intersections and the relative %TCu grade distribution.

The mined-out portion of Tilwezembe is to the west of 35600m easting.

The drill-hole database indicates a quasi-25 to 30 m spacing near the surface to about 50 m to

elevations of about 1050 m above mean sea level (“mamsl”). Further to the east of 357200m

easting, the drill-hole spacing becomes much wider.

In terms of grade distribution, the intersections in the breccia indicate variability in the %TCu

grade within shorter distances.

Figure 19.7 Tilwezembe Mine: %TCu grade distribution in the Breccia

Kamoto Mine

Kamoto mine has been largely investigated by diamond drilling from underground mining

development, and the drill-hole distribution shows how the intensity of the development

varied from one area to another.

Overall, the %TCu grade distributions in either of the OBI and OBS mineralized zones

indicate minor variability where data exist, with the exception of the Etang, where the grades

are relatively lower. Refer to Figure 19.8 for details.

356850 356900 356950 357000 357050 357100 357150 357200 357250 357300 357350 357400 357450 357500 357550 357600

Easting

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0.00 to 1.00

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3.50 to 4.50

4.50 to 4.00

6.00 to 9.00



Figure 19.8 Kamoto Mine: %TCu grade distribution in the OBI and OBS (plan view)

-1400 -1200 -1000 -800 -600 -400 -200 0 200 400 600 800 1000

X-Coordinate

200

400

600

800

1000

1200

1400

1600

1800

2000Y

-Co

ord

ina

te

%TCu

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3.0 to 4.0

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5.0 to 6.0

6.0 to 25.0

-1400 -1200 -1000 -800 -600 -400 -200 0 200 400 600 800 1000

X-Coordinate

200

400

600

800

1000

1200

1400

1600

1800

2000

Y-C

oord

inate

%TCu

0.2 to 1.0

1.0 to 2.0

2.0 to 3.0

3.0 to 4.0

4.0 to 5.0

5.0 to 6.0

6.0 to 25.0



Kananga Mine

Figure 19.9 Kananga Mine: Longitudinal sections showing UOB (top figure) and LOB(bottom figure) intersections

331250 331300 331350 331400 331450 331500 331550 331600 331650 331700 331750

Easting

1050

1100

1150

1200

1250

1300

1350

1400

Ele

vati

on

%TCu

0.00 to 1.00

1.00 to 1.50

1.50 to 2.00

2.00 to 3.50

3.50 to 4.50

4.50 to 4.00

6.00 to 9.00

331250 331300 331350 331400 331450 331500 331550 331600 331650 331700 331750

Easting

1050

1100

1150

1200

1250

1300

1350

1400

Ele

va

tio

n

%TCu

0.00 to 1.00

1.00 to 1.50

1.50 to 2.00

2.00 to 3.50

3.50 to 4.50

4.50 to 4.00

6.00 to 9.00



Kananga Mine

Kananga Mine is intersected by drill-holes on an average spacing of 50 m along strike and

less than 50 m across strike. There are more intersections in the LOB than the UOB,

especially further to the east.

The average grade of the intersections in Kananga are generally within the 2-to-3%TCu

bracket, with better %TCu intersections in the LOB compared with the UOB. The near-

surface intersections of the UOB appear to be depleted in copper. Refer to Figure 19.9 for

details.

KOV Mine

For KOV mine, plan-view plots have been generated showing the Virgule and Kamoto East

data and, in the second plot, the FNSR and Oliveira data. The drilling in KOV is on a nominal

spacing of 100 m along strike and is closer than 100 m across strike.

The %TCu grade distribution in Virgule indicates relatively high grades on the east, which

decrease with depth to the west. The drill-hole coverage for Kamoto east is very limited due

to its steep to near-vertical dip and the problems associated with surface drilling of vertical

zones of mineralization. The majority of the intersections in Kamoto East are above the

current surface topography and represent mined-out areas. There is a general consistency in

the %TCu grade distribution with lower grades to the west.

The drill-holes intersecting Oliveira show generally higher grades in the east and lower grades

to the north, the up-dip portion, and to the west. Figure 19.10 shows grade distributions at

KOV Mine.



Figure 19.10 KOV Mine: %TCu grade distribution plots for Kamoto east andVirgule, and Oliveira and FNSR respectively

325800 326000 326200 326400 326600 326800 327000 327200 327400 327600 327800 328000 328200 328400

Easting

8814200

8814400

8814600

8814800

8815000

8815200

8815400

8815600

8815800

8816000

8816200

8816400

Nort

hin

g

%TCu

1.0 to 2.0

2.0 to 3.0

3.0 to 4.0

4.0 to 5.0

5.0 to 6.0

6.0 to 8.0

8.0 to 12.0

VIRGULE

KAMOTO EAST

325800 326000 326200 326400 326600 326800 327000 327200 327400 327600 327800 328000 328200 328400

Easting

8814200

8814400

8814600

8814800

8815000

8815200

8815400

8815600

8815800

8816000

8816200

8816400

Nort

hin

g

%TCu

1.0 to 2.0

2.0 to 3.0

3.0 to 4.0

4.0 to 5.0

5.0 to 6.0

6.0 to 8.0

8.0 to 12.0



Mashamba East Mine

Figure 19.11 shows grade distributions at Mashamba East Mine.

Figure 19.11 Mashamba East Mine: %TCu grade distribution in RSF and SDBrespectively (plan view)

432400 432600 432800 433000 433200 433400 433600 433800 434000 434200 434400 434600

Easting

306800

307000

307200

307400

307600

307800

308000

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308400

Nort

hin

g

%TCu

0.01 to 0.50

0.50 to 1.00

1.00 to 2.00

2.00 to 3.00

3.00 to 4.00

4.00 to 5.00

5.00 to 7.50

7.50 to 20.00No assays

432400 432600 432800 433000 433200 433400 433600 433800 434000 434200 434400 434600

Easting

306800

307000

307200

307400

307600

307800

308000

308200

308400

Nort

hin

g

%TCu

0.01 to 0.50

0.50 to 1.00

1.00 to 2.00

2.00 to 3.00

3.00 to 4.00

4.00 to 5.00

5.00 to 7.50

7.50 to 20.00No assays



In Mashamba East, the drill-hole spacing is on average 100 m, but in the SDB the drill-hole

the spacing varies when considering the drill-holes with assays rather than the physical drill

position. From the point of view of the %TCu grade distribution, there are two main zones of

high-grade mineralization in the RSF, one in the west and the other in the turn-around in the

east. Otherwise, the RSF is of relatively lower grade. The SDB is generally low grade with

spotty high grade intersections.

19.1.6 Statistics

T17 Mine

The lithological wireframes were used to extract the sample data, and the data was

composited to 2,5 m lengths. Statistics from the composite files are given in Table 19.7.

Table 19.7 T17 Mine: Statistics from the 2,5 m Composites by Lithology Type

Lithology Variable No. Samples Minimum Maximum Mean Std. dev CoV

BOMZ %TCu 386 0,10 15,87 4,18 3,36 0,80

BOMZ %TCo 386 0,00 7,40 0,51 0,87 1,73

BOMZ %TCu 57 0,20 10,10 2,55 2,60 1,02

BOMZ %TCo 57 0,07 1,63 0,48 0,46 0,95

SDB %TCu 202 0,10 15,87 5,14 4,04 0,79

SDB %TCo 202 0,00 7,40 0,78 1,07 1,38

RSC %TCu 202 0,00 19,20 1,65 2,84 1,72

RSC %TCo 202 0,00 4,70 0,44 0,81 1,84

RSF %TCu 135 0,15 10,72 2,94 2,47 0,84

RSF %TCo 135 0,08 7,60 0,99 1,18 1,20

DSTRAT %TCu 101 0,27 9,85 3,93 2,60 0,66

DSTRAT %TCo 101 0,07 4,06 0,39 0,48 1,24

Tilwezembe Mine

The summary statistics of the declustered 1 m composite data of the various rock types are

presented in Tables 19.8. The composite data was declustered using a cell size of 25 mE by

25 mN by 1 mRL that approximates the drill-hole spacing in the closer spaced areas.

Generally, coefficients of variation (CV = standard deviation/mean) for copper and cobalt

(total and soluble) in the manganiferous dolomite zones are high (>1,5) and those in the other

zones are lower (<1,5). For manganese, CVs in the oxidized zones were generally high (<1,5)

and low in the sulphide zones (<1,5). All specific-gravity CVs were low and displayed low

variability.

Log histograms of the declustered 1 m composite data show that the distributions approached

log normality, although evidence of multiple populations was seen in some histograms.

Histograms are negatively skewed due to the inclusion of internal waste, and positive skew

reflects the large number of very high grade samples with a small number of relatively low-

grade samples.



Positively skewed histograms with high CVs are generally affected by small numbers of high

grade samples (outliers). Grade estimates made from positively skewed data are needed for

the control of these outliers and reduce the CVs to prevent them from overly influencing the

estimate and biasing the results.

Statistics from the composite files are indicated in Table 19.8.

Table 19.8 Tilwezembe Mine: Statistics from the 1 m Composites by Lithology Type

Domain Variable Samples Minimum Maximum Mean CV

OX_MNDOL %TCu 1054 0,01 26,5 1,27 1,91

OX_BREC %TCu 485 0,10 19,84 3,78 0,89

OX_TILAR %TCu 674 0,05 4,92 0,56 1,05

SL_MNDOL %TCu 511 0,03 37,00 1,19 2,42

SL_BREC %TCu 339 0,02 21,38 3,29 0,97

SL_TILAR %TCu 406 0,01 6,27 0,53 1,28

OX_MNDOL %AsCu 1054 0,01 14,12 0,91 1,85

OX_BREC %AsCu 485 0,05 14,28 3,16 0,96

OX_TILAR %AsCu 674 0,01 4,90 0,44 1,17

SL_MNDOL %AsCu 511 0,01 6,88 0,24 2,97

SL_BREC %AsCu 339 0,01 3,52 0,33 1,19

SL_TILAR %AsCu 406 0,01 1,09 0,13 1,21

OX_MNDOL %TCo 1054 0,01 11,15 0,5 1,47

OX_BREC %TCo 485 0,01 13,47 0,96 1,58

OX_TILAR %TCo 674 0,01 3,61 0,29 1,02

SL_MNDOL %TCo 511 0,01 7,94 0,38 1,52

SL_BREC %TCo 339 0,03 4,98 1,19 1,01

SL_TILAR %TCo 406 0,02 4,00 0,34 1,15

Kamoto Mine

Tables 19.9 to 19.11 present statistics for Region Principal, Etang North and Etang South.

Table 19.9 Kamoto Mine: Kamoto Principal – Statistics per Unit

%TCu %TCo

No.Samples

Minimum Maximum MeanStd.dev

No.Samples


Dstart 427 0,11 9,56 4,08 1,40 393 0,01 8,24 0,36 0,44

RSF 404 0,14 22,00 4,79 1,66 378 0,01 3,15 0,30 0,26

RSC_B 241 0,10 27,40 7,59 3,87 224 0,03 5,99 0,51 0,68

RSC_T 165 0,24 18,30 5,97 3,53 159 0,01 4,39 0,73 0,65

SD1A 326 0,40 22,40 5,87 2,19 307 0,01 6,40 0,63 0,51



Table 19.10 Kamoto Mine: Etang South – Statistics per Unit

%TCu %TCo

No.Samples


No.Samples


Dstart 89 0,30 7,62 2,89 1,33 91 0,08 17,61 0,75 1,84

RSF 103 0,20 16,16 3,44 1,44 104 0,10 2,55 0,60 0,42

RSC_B 50 0,18 12,05 2,98 2,35 49 0,22 5,94 1,22 1,06

RSC_T 39 0,25 12,00 2,46 1,87 40 0,17 2,96 0,96 0,61

SD1A 155 0,15 13,20 5,92 2,97 154 0,07 3,96 1,04 0,74

Table 19.11 Kamoto Mine: Etang North - Statistics per Unit

%TCu %TCo

No.Samples


No.Samples


Dstart 28 0,50 3,37 2,03 0,67 23 0,12 0,84 0,42 0,17

RSF 23 0,59 5,15 3,02 1,15 17 0,09 0,99 0,42 0,26

RSC_B 16 0,88 12,00 3,58 2,8 12 0,31 2,32 1,14 0,65

RSC_T 9 0,67 7,30 3,54 2,54 7 0,58 2,52 1,24 0,59

SD1A 45 0,22 8,07 3,36 1,47 31 0,32 2,39 0,82 0,54

Kananga Mine

The summary statistics of the declustered 1 m composite data of the various rock types are

presented in Table 19.12. The composite data was declustered using a cell size of 25 mE by

25 mN by 1 mRL that approximates the drill-hole spacing in the closer spaced areas.

The following have been observed from the statistical analysis:

The mean values of the internal or middle (MID) zone are lower in comparison with

those of the upper and lower orebodies (UOB and LOB) and the grades are generally

below 0,5% total copper. Consequently, mining most of the internal material may not

be economically viable.

The majority of the histograms approach log normality. Some are negatively skewed

due to the inclusion of small amounts of waste (low grades) and some are positively

skewed due the large number of low-grade samples with a small number of relatively

higher grade samples.

The coefficients of variation (CV = standard deviation/mean) are generally low (<1).

There are some exceptions, most notably for the UOB_OX domain, for which CV’s

were higher and between 1,1 and 1,9. The histograms with high CV’s are generally

affected by small numbers of outlier values. Grade estimates made from skewed data

need to control these outlier values and reduce the CVs to prevent them from overly

influencing the estimate and biasing the results.



Table 19.12 Kananga Mine: Statistics from the 1 m Composites by Lithology Type

Domain VariableNo.

SamplesMinimum Maximum Mean CV

UOB_OX %TCu 250 0,08 10,05 1,13 1,1

MID_OX %TCu 528 0,02 0,64 0,16 0,6

LOB_OX %TCu 297 0,03 9,28 1,93 0,8

UOB_SL %TCu 122 0,01 6,10 1,83 0,6

MID_SL %TCu 269 0,02 1,00 0,26 0,6

LOB_SL %TCu 234 0,13 6,75 2,18 0,6

UOB_OX %AsCu 250 0,02 9,05 0,85 1,2

MID_OX %AsCu 528 0,01 0,62 0,12 0,8

LOB_OX %AsCu 297 0,01 9,26 1,26 1,1

UOB_SL %AsCu 122 0,01 1,61 0,15 1,2

MID_SL %AsCu 269 0,01 0,39 0,07 1,2

LOB_SL %AsCu 234 0,01 2,99 0,22 1,6

UOB_OX %TCo 250 0,06 4,51 0,64 1,2

MID_OX %TCo 528 0,02 2,73 0,27 0,8

LOB_OX %TCo 297 0,02 3,37 0,70 0,8

UOB_SL %TCo 122 0,01 4,79 0,94 1,1

MID_SL %TCo 269 0,06 2,03 0,53 0,6

LOB_SL %TCo 234 0,02 3,49 1,05 0,7

KOV Mine

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

Tables 19.13 to 19.16.

Table 19.13 KOV Mine: Virgule – Statistics from the 2,5 m Composites by LithologyType

Lithology Variable No. Samples Minimum Maximum Mean Std dev

BOMZ %TCu 77 0,10 12,00 3,54 3,10

%TCo 112 0,00 3,62 0,46 0,67

SDB %TCu 296 0,02 12,92 5,97 3,60

%TCo 371 0,00 11,50 0,46 0.84

RSC %TCu 384 0,08 23,14 4,39 3,47

%TCo 587 0,00 5,00 0,21 0,40

RSF %TCu 171 0,14 12,00 6,20 3,03

%TCo 244 0,00 2,25 0,19 0,32

DSTRAT %TCu 103 0,50 12,00 6,43 2,41

%TCo 128 0,00 1,52 0,22 0,31

RATGR %TCu 49 0,80 14,35 6,36 3,42

%TCo 103 0,00 0,99 0,09 0,16



Table 19.14 KOV Mine: Oliveira – Statistics from the 2,5 m Composites by LithologyType


BOMZ %TCu 72 0,15 7,88 2,05 1,99

%TCo 80 0,00 2,25 0,41 0,42

SDB %TCu 250 0,10 13,68 5,59 2,48

%TCo 239 0,00 3,66 0,93 0,70

RSC %TCu 266 0,10 16,59 4,72 3,71

%TCo 512 0,00 4,58 0,29 0,55

RSF %TCu 120 0,40 14,99 5,40 3,16

%TCo 111 0,00 2,90 0,38 0,48

DSTRAT %TCu 73 0,91 12,00 4,92 2,31

%TCo 71 0,00 1,61 0,32 0,33

RATGR %TCu 43 0,60 16,41 4,80 2,78

%TCo 60 0,00 1,24 0,23 0,33

Table 19.15 KOV Mine: FNSR – Statistics from the 2,5 m Composites by LithologyType


BOMZ %TCu 9 0,53 12,00 5,24 3,41

%TCo 11 0,00 2,85 0,67 0,89

SDB %TCu 71 1,11 12,00 7,93 3,16

%TCo 79 0,00 6,15 0,42 0,88

RSC %TCu 69 0,15 12,00 5,27 4,28

%TCo 110 0,00 1,96 0,18 0,29

RSF %TCu 25 1,66 12,47 5,76 2,40

%TCo 25 0,02 0,47 0,18 0,13

DSTRAT %TCu 7 4,30 8,98 7,17 1,48

%TCo 9 0,00 0,16 0,04 0,05

RATGR %TCu 2 4,06 8,02 6,04 1,98

%TCo 2 0,02 0,07 0,04 0,02

Table 19.16 KOV Mine: Kamoto East – Statistics from the 2,5 m Composites byLithology Type


BOMZ %TCu 27 0,17 13,22 6,16 3,31

%TCo 43 0,00 4,96 0,71 0,97

SDB %TCu 133 0,08 14,38 6,32 3,34

%TCo 173 0,00 4,63 0,41 0,59

RSC %TCu 206 0,11 16,07 4,20 3,72

%TCo 288 0,00 2,52 0,22 0,29

RSF %TCu 76 0,21 10,84 5,09 3,11

%TCu 96 0,00 1,50 0,26 0,30

DSTRAT %TCo 43 0,21 11,55 5,53 2,89

%TCu 70 0,00 1,40 0,21 0,28

RATGR %TCo 15 3,02 9,02 6,37 1,71

%TCu 25 0,00 0,86 0,13 0,19



The statistics show that, in general, relatively high copper grades are associated with all the

lithologies from the BOMZ to the RATGR and within each fragment. The cobalt grades are

higher in the BOMZ and the SDB, which make up the upper mineralized zone, and lower in

the lower mineralized zone; RSF, DSTRAT and RATGR lithologies.

The RSC statistics reflect the grades associated with the upper RSC and lower RSC, which

occur near the contact with the well mineralized SDB and RSF respectively. Sampling of the

mid-RSC is limited, and these intervals are therefore not included in the statistics.

The mid-RSC, which occupies about 10 m to 15 m of the 25 m RSC thickness, was sampled

only in very few instances. This was probably because there was no visible copper

mineralization.

The mid-RSC is differentiated from the rest of the RSC by a wireframe model defining the

extent of the unsampled section within each fragment. Where the drill hole was sampled

through the entire RSC lithology, the position of the mid-RSC wireframe was defined from

the surrounding holes with unsampled intervals. The mid-RSC wireframe was used to extract

the partial and complete sampled intervals in each of the drill holes inside the wireframe.

The drillholes with complete sampling information in the mid-RSC were selected, and

weighted average grades for %TCu and %TCo were obtained for each fragment. These

weighted average values were assigned to the mid-RSC portion.

Statistics from the sampled intervals in the mid-RSC and respective fragments are shown in

Table 19.17.

Geological models of the lithology were based on the broad Gecamines’ lithological

interpretations presented on the 1:1000 sections and entries of lithologies in the drill-hole data

file. SRK refined the sectional lithological interpretations by snapping the strings onto the

drill-hole lithological contacts.

Three dimensional lithological wireframe models were generated for each of the four resource

areas: Kamoto East, Virgule, Oliveira and FNSR fragments.

The mineralized zones were defined on the basis of a 1.0%TCu cut-off grade. However, the

mineralization within each fragment extends from the BOMZ to RATGR and is well above

the cut-off grades.

Blocks of fundamental size 20 m x 20 m x 5m in the X, Y and Z directions have been used for

the modelling with sub-celling to 1/8th the block size in the X and Y directions and allowing

for the maximum sub-celling in the Z direction to allow for the reasonably accurate definition

of any lithological contact.



Table 19.17 KOV Mine: Statistics from Limited mid-RSC Sampling

Project Area Variable Minimum Maximum Average Weighted averageStd.dev

CV Count

Virgule

Virgule

%TCu 0,21 18,70 2,91 2,80 4,07 1,40 65

%TCo 0,03 9,15 0,53 0,22 1,46 2,75 65

Oliveira

Oliveira

%TCu 0,06 2,94 0,85 0,94 0,68 0,80 42

%TCo 0,06 0,96 0,42 0,45 0,26 0,61 42

FNSR

FNSR

%TCu 0,03 11,15 1,30 1,13 2,95 2,27 24

%TCo 0,00 0,27 0,08 0,08 0,06 0,67 25

Kamoto East

Kamoto East

%TCu 0,27 12,00 2,25 2,21 3,24 1,44 19

%TCo 0,00 1,04 1,04 0,23 0,23 0,22 19

The weighted average values for %TCu and %TCo have been assigned to the respective mid-

RSC volumes.

Mashamba East Mine

The lithological wireframes were used to extract the sample data, and the data were

composited to 2,5 m lengths. Statistics from the composite files are shown in Table 19.18.

Table 19.18 Mashamba East: Statistics from the 2,5m Composites by Lithology Type

Lithology Geozone VariableNo.

SamplesMinimum Maximum Mean Std. dev CoV

BOMZ WEST %TCu 125 0,01 10,60 0,48 1,40 2,94

BOMZ WEST %TCo 125 0,01 0,44 0,09 0,11 1,20

BOMZ EAST %TCu 66 0,01 12,00 0,70 1,78 2,53

BOMZ EAST %TCo 66 0,01 1,08 0,14 0,23 1,65

SDB WEST %TCu 237 0,01 19,26 1,19 2,46 2,06

SDB WEST %TCo 237 0,01 4,14 0,36 0,61 1,69

SDB EAST %TCu 64 0,01 9,15 1,82 2,51 1,38

SDB EAST %TCo 64 0,01 2,63 0,33 0,48 1,42

RSC WEST %TCu 248 0,01 12,00 0,60 2,00 3,34

RSC WEST %TCo 248 0,00 2,01 0,07 0,23 3,24

RSC EAST %TCu 112 0,01 12,00 0,69 2,06 2,98

RSC EAST %TCo 112 0,01 0,66 0,04 0,08 2,31

RSF WEST %TCu 440 0,01 12,00 2,28 2,85 1,25

RSF WEST %TCo 440 0,00 2,80 0,37 0,54 1,46

RSF EAST %TCu 119 0,01 11,37 3,83 2,96 0,77

RSF EAST %TCo 119 0,00 1,28 0,14 0,22 1,56

19.1.7 Variography

T17 Mine

For T17 Mine omni-directional variograms were generated in each of the lithologies. There

are limited sample data for the generation of directional variograms. The variogram

parameters are indicated in Table 19.19.



Table 19.19 T17 Mine: Omni-directional variogram parameters by lithology

Zone VariableNugget

Variance, C0C1 Variance Range C2 Variance Range2

DSTRAT %TCu 0,5135 1,775 232,79

%TCo 0,03565 0,04481 397,72

RSF %TCu 1,589 2,356 126,01 3,421 565,56

%TCo 0,08955 0,1464 321,98

RSC %TCu 0,1042 1,87 178,33

%TCo 0,006679 0,2475 180,73

SDB %TCu 0,3691 0,8281 256,58

%TCo 0,01728 0,05536 194,02

BOMZ %TCu 1,984 3,257 283,48

%TCo 0,03098 0,02172 283,32

Tilwezembe Mine

Traditional semi-variograms were calculated from the selected composite data using

Supervisor software. To improve the variogram structures, a normal scores transform was

performed on the composites before semi-variogram calculation.

Semi-variograms for the individual rock types were often not very robust, especially in the

sulphide zones that were not as extensively drilled as the oxide zones. The combination of the

oxide and sulphide zones resulted in an adequate amount of data to calculate reasonable and

robust variograms. As a result, variograms were modelled for the entire brecciated zone, the

entire manganiferous dolomite zone and the entire low-grade argillite/tillite zone.

Where two elements were highly correlated (correlation coefficient >0,7), the variogram

model of the most significant grade variable was applied direct to the second variable. As the

semi-variograms were predominantly calculated from data from the oxide zone that was more

extensively drilled than the sulphide zone, the oxide correlation matrices were considered.

High correlations existed between total copper and soluble copper and between total cobalt

and soluble cobalt. As a result, the total semi-variograms were directly applied to the soluble

elements to preserve the correlations. Therefore, semi-variograms were modelled only for

total copper, cobalt and manganese.

The directions of continuity for these elements were evaluated by making use of semi-

variogram contours on the horizontal, across-strike and dip planes. This allowed for the

determination of the strike, dip and plunge continuity. The calculated experimental semi-

variograms for each of the attributes were modelled using spherical models with two or three

structures. All variograms were standardized to a sill of one, representing all of the sample

variance. The nugget effect was determined by extrapolating the first two points of the down-

hole variogram to the Y-axis.

The variogram parameters are summarized in Tables 19.20 to 19.22.



Table 19.20 Tilwezembe Mine: Back transformed variogram parameters – Manganiferous Dolomites (Oxide and Sulphide)

Structure 1 Structure 2 Structure 3

Attribute

(%)

Nugget

effectSill

Range

Dir1 (m)

Range Dir2

(m)

Range

Dir3 (m)Sill

Range

Dir1 (m)

Range Dir2

(m)

Range

Dir3 (m)Sill

Range

Dir1 (m)

Range

Dir2 (m)

Range

Dir3 (m)

Total Cu 0,05 0,52 40 20 7 0,43 160 60 14 - - - -

Total Co 0,04 0,59 40 15 12 0,37 130 70 12 - - - -

Total Mn 0,05 0,63 30 10 8 0,32 160 70 12 - - - -

Table 19.21 Tilwezembe Mine: Back transformed variogram parameters – Breccia (Oxide and Sulphide)


Attribute

(%)

Nugget

effectSill

Range

Dir1 (m)

Range Dir2

(m)

Range

Dir3 (m)Sill

Range

Dir1 (m)

Range Dir2

(m)

Range

Dir3 (m)Sill

Range

Dir1 (m)

Range

Dir2 (m)

Range

Dir3 (m)

Total Cu 0,12 0,13 20 70 10 0,75 160 70 10 - - - -

Total Co 0,11 0,22 40 10 3 0,19 40 100 12 0,48 330 100 12

Total Mn 0,22 0,48 30 50 7 0,3 180 50 7 - - - -

Table 19.22 Tilwezembe Mine: Back transformed variogram parameters – Tillites and Argillites (Oxide and Sulphide)


Attribute

(%)

Nugget

effectSill

Range

Dir1 (m)

Range Dir2

(m)

Range

Dir3 (m)Sill

Range

Dir1 (m)

Range Dir2

(m)

Range

Dir3 (m)Sill

Range

Dir1 (m)

Range

Dir2 (m)

Range

Dir3 (m)

Total Cu 0,15 0,4 110 80 15 0,44 125 110 15 - - - -

Total Co 0,09 0,18 180 50 2 0,73 180 50 20 - - - -

Total Mn 0,12 0,53 35 90 12 0,36 160 90 12 - - - -



Kamoto Mine

Variogram calculation, analysis and modelling were done using Datamine™ Studio, on the

composited samples, within the Principal region only. The reason for excluding the other

regions from the variogram analysis was that each region is fault bounded, with the

magnitude of displacement unknown. Variogram parameters of the Principal region were

applied to the other regions due to the relatively small dataset of each region.

Due to the folded/warped and undulating nature of the orebody, the dataset was “unfolded”

prior to any variography. The objective of the “unfolding” exercise was to improve the

planarity of the dataset and hence improve the variogram structures. The method of unfolding

involved the generation of perimeters radiating from the centre, outwards. This was to ensure

that the central point became the “origin’ or pivot point.

All variograms were modelled in the “unfolded” space. Each unit was analysed individually

and the analysis was undertaken on %TCu and %TCo.

Variogram contours and spidergrams yielded no evidence of anisotropy in the dataset, hence

isotropic variogram models were used. The dataset of each unit was analysed using

histograms, cumulative histograms and probability plots to aid in identifying outliers.

Potential outliers were excluded from the variography and also used as a guideline for the

application of top-cut limits during grade estimation.

Up to six lag distances were used in the calculation of the experimental variograms. This was

done to assess the stability of the theoretical variogram models against fluctuating lags

distances because of the irregular sample/drilling spacing.

The following images summarise the variography of Cu and Co values per Unit, starting with

the DStrat, up to the SD1a.

The variogram parameters are summarized in Tables 19.23 and 19.24.



Table 19.23 Kamoto Mine: Copper variogram models

VARIABLE CO C1Range1 X-

dirRange1 Y-

dirRange1 Z-

dirC2

Range2 X-dir

Range2 Y-dir

Range2 Z-dir

C3Range3 X-

dirRange3 Y-

dirRange3 Z-

dir

DSTRAT 0,01 1,26 9,96 9,96 5,4 1,83 30,98 30,98 24,1 1,44 324,96 324,96 24,1

RSF 0,031 1,97 3,92 3,92 3,92 2,82 30,17 30,17 8,5 1,82 296,61 296,61 20,3

RSC 0,043 2,29 7,83 7,83 6,7 5,08 52,19 52,19 27,4 - - - -

SDB 0,124 3,26 6,98 6,98 6,98 5,31 74,36 74,36 8,3 2,43 189,72 189,72 27,4

BOMZ 0,042 1,83 7,93 7,93 7,8 1,39 272,1 272,1 20,3 - - - -

Table 19.24 Kamoto Mine: Cobalt variogram models

VARIABLE CO C1Range1 X-

dirRange1 Y-

dirRange1 Z-

dirC2

Range2 X-dir

Range2 Y-dir

Range2 Z-dir

DSTRAT 0 0,07 11,87 11,87 8,4 0,03 147,85 147,85 18,2

RSF 0,003 0,12 3,05 3,05 8,5 0,12 184,05 184,05 15,3

RSC 0,004 0,11 7,52 7,52 7,2 0,11 70,59 70,59 7,2

SDB 0,005 0,17 10,96 23,97 4,96 0,27 31,98 324,92 11,3

BOMZ 0,021 0,18 56,65 170 11,68 - - - -



Kananga Mine

Traditional semi-variograms were calculated from the selected composite data using

Supervisor software. To improve the variogram structures, normal scores transform was

performed on the composites before semi-variogram calculation.

Semi-variograms for the individual rock types were often not very robust, especially in the

sulphide zones that were not as extensively drilled as the oxide zones. The combination of the

oxide and sulphide zones resulted in an adequate amount of data to calculate reasonable and

robust variograms.

Where two elements were highly correlated (correlation coefficient >0,7), the variogram

model of the most significant grade variable was applied direct to the second variable. As the

semi-variograms were predominantly calculated from data from the oxide zone that was more

extensively drilled than the sulphide zone, the oxide correlation matrices were considered.

Strong correlations existed between total copper and soluble copper and between total cobalt

and soluble cobalt. As a result, the total semi-variograms were directly applied to the soluble

elements to preserve the correlations. Therefore, semi-variograms were only modelled for

total copper, cobalt and manganese.

The directions of continuity for these elements were evaluated by making use of semi-

variogram contours on the horizontal, across-strike and dip planes. This allowed for the

determination of the strike, dip and plunge continuity. The calculated experimental semi-

variograms for each of the attributes were modelled using spherical models with two or three

structures. All variograms were standardized to a sill of one, representing all of the sample

variance. The nugget effect was determined by extrapolating the first two points of the down-

hole variogram to the Y-axis. The variogram parameters are summarized in Table 19.25 and

19.26.



Table 19.25 Kananga Mine: Back transformed variogram parameters – upper orebody (Oxide and Sulphide)


Attribute Nugget SillRange

Dir1 (m)

Range

Dir2 (m)

Range

Dir3 (m)Sill

Range

Dir1 (m)

Range

Dir2 (m)

Range

Dir3 (m)Sill

Range

Dir1 (m)

Range

Dir2 (m)

Range

Dir3

(m)

Directions

(ZXZ axes

rotation)*

Total Cu (%) 0,01 0,62 60 90 11 0,36 300 90 11 - - - - (-30,115,0)

Total Co (%) 0,06 0,61 50 20 9 0,33 300 70 9 - - - - (-20,115,0)

Total Mn (%) 0,23 0,29 60 10 9 0,2 60 140 9 0,29 180 140 9 (-30,125,0)

* Rotations in a clockwise direction are positive, whilst those in an anti-clockwise direction are negative

Table 19.26 Kananga Mine: Back transformed variogram parameters – internal/middle zone (Oxide and Sulphide)


Attribute Nugget SillRange

Dir1 (m)

Range

Dir2 (m)

Range

Dir3 (m)Sill

Range

Dir1 (m)

Range

Dir2 (m)

Range

Dir3 (m)Sill

Range

Dir1 (m)

Range

Dir2 (m)

Range

Dir3

(m)

Directions

(ZXZ axes

rotation)*

Total Cu (%) 0,12 0,41 50 20 10 0,10 50 230 25 0.37 460 230 25 (-25,130,0)

Total Co (%) 0,02 0,47 70 50 11 0,51 340 290 27 - - - - (-35,125,0)

Total Mn (%) 0,05 0,53 50 130 13 0,41 360 130 13 - - - - (-40,140,0)

* Rotations in a clockwise direction are positive, whilst those in an anti-clockwise direction are negative



KOV Mine

Omni-directional variograms were generated for each of the lithologies, with the exception of

RATGR, in the Virgule and Oliveira fragments. There were insufficient composite data for the

generation of omni-directional variograms for the RATGR in both of these fragments.

Because of inadequate composite data, omni-directional variograms were not generated for the

lithologies in the Kamoto East and FNSR fragments.

The fitted omni-directional variograms for %TCu and %TCo within the various lithologies in the

Virgule and Oliveira fragments are shown in Tables 19.27and 19.28.

Table 19.27 KOV Mine: Virgule- Omni-Directional Variogram Parameters by Lithology

Lithology Variable Nugget C1 Range

BOMZ TCu 1,65 3,38 309,38

BOMZ TCo 0,03 0,07 204,17

SDB TCu 8,74 3,78 380,12

SDB TCo 0,05 0,13 204,08

RSC TCu 8,59 3,44 209,69

RSC TCo 0,08 0,07 266,89

RSF TCu 2,44 0,68 245,78

RSF TCo 0,00 0,01 217,35

DSTRAT TCu 3,06 2,51 243,89

DSTRAT TCo 0,01 0,01 317,03

Table 19.28 KOV Mine: Oliveira - Omni-Directional Variogram Parameters by Lithology

Lithology Variable Nugget C1 Range

BOMZ TCu 2,68 289,22

BOMZ TCo 0,04 0,06 284,57

SDB TCu 3,15 1,17 303,40

SDB TCo 0,01 0,03 192,65

RSC TCu 3,19 1,56 261,88

RSC TCo 0,03 0,01 221,57

RSF TCu 1,01 2,90 265,60

RSF TCo 0,00 0,02 178,66

DSTRAT TCu 0,28 1,12 350,92

DSTRAT TCo 0,02 0,09 253,91

Mashamba East Mine

For Mashamba East Mine, omni-directional variograms were generated for each of the lithologies.

There are limited sample data for the generation of directional variograms. The variogram

parameters are indicated in Table 19.29.



Table 19.29 Mashamba East Mine: Omni-directional Variogram Parameters by Lithology

Zone Variable Nugget Variance, C0 C1 Variance Range

RSF %TCu 1,943 1,425 157,67

%TCo 0,04157 0,08539 332,59

RSC %TCu 1,006 0,6158 210,88

%TCo 0,07737 0,03845 256,14

SDB %TCu 1,235 1,564 216,8

%TCo 0,05093 0,1141 291,64

BOMZ %TCu 0,1709 2,731 353,88

%TCo 0,02193 0,03127 467,49

19.1.8 Grade estimation

T17 Mine

SRK re-estimated the %TCu and %TCo grades into the T17 Mine block model as generated by

CCIC. SRK accepted the CCIC lithological model and used that for the selection of the samples to

use for the estimation.

Composites for the estimation were selected within each lithology to estimate the respective

lithology. The criterion adopted by CCIC to account for absent data, where 0,01%TCu was assigned

for unsampled sections in the drill-hole, was also applied

SRK estimated grades into the T17 Mine block model using the three-pass method as described

above with slight modification as to the parameters used for the search and orientation of the

ellipsoid. The first search neighbourhood was set equivalent to the variogram range. The variogram

parameters for the lithologies within T17 Mine are shown in Table 19.30.

For T17, “geozones” were defined to demarcate zones where there were changes in the strike and dip

of the strata. The search ellipsoid was oriented using the strike/dip parameters within these

“geozones”. The same lithological variogram parameters were used for the defined geozones.

Table 19.30 T17 Mine: Variogram parameters by lithology

Zone Variable Nugget Variance, C0 C1 Variance Range 1 (x,y) C2 Variance Range 2 (x,y)

DSTRAT %TCu 0,5135 1,775 232,79

%TCo 0,03565 0,04481 397,72

RSF %TCu 1,589 2,356 126,01 3,421 565,56

%TCo 0,08955 0,1464 321,98

BOMZ %TCu 0,1042 1,87 178,33

%TCo 0,006679 0,2475 180,73

SDS %TCu 0,3691 0,8281 256,58

%TCo 0,01728 0,05536 194,02

SDB %TCu 1,984 3,257 283,48

%TCo 0,03098 0,02172 283,32



Tilwezembe Mine

During the grade estimation process, Snowden applied various topcuts to the data to limit the

influence of high grades in he estimation into the various lithological domains and for the various

grade fields, %TCu, %AsCu %TCo, %AsCo and %Mn.

Ordinary kriging was used in the estimation process. The defined mineralized zones were subdivided

into two structural domains based on the strike. Samples used to estimate the respective blocks were

sourced within an ellipsoidal search oriented according to the structural domain in which a

mineralization block occurred.

Composites used in the estimation process were sourced within search criteria of

130 m x 60 m x 12 m in the strike, across strike and perpendicular to strike directions equivalent to

the variogram ranges.

Where blocks remained unestimated after the first pass, the search neighbourhood was doubled or

tripled.

Kamoto Mine

CCIC estimated the various geozones within Kamoto Mine using ordinary kriging and input

variogram and search neighbourhood parameters. The variogram parameters used are indicated in

Tables 19.23 and 19.24 and the same variogram range was used in all the directions.

Samples used in the estimation were sourced from search neighbourhoods equivalent to the second

variogram ranges for each of the variables and in each lithology, except for one case, where the only

first variogram range was modelled and this was used as the search criterion.

A second search, equivalent to either 1.25 or 1.5 times the first search was used to estimate blocks

that remained un-estimated after the first run. For bocks still not estimated after the two passes, a

third global search of 1000 x 1000 x 1000m was used.

The minimum number of samples used to interpolate into a block was set to 4 and the maximum 40.

Zonal control was applied per stratigraphic unit. Block model was coded using the stratigraphic

wireframes into the various stratigraphic units. Only samples having the same codes where used in

the interpolation of the various units.

The Kamoto Mine model was updated in 2008 by CCIC to address the concerns raised by SRK in

the review and also to deplete the model based on the findings of the project team on site with

respect to the previous mining areas which were not accounted for in the 2007 model. This affected

Zone 8 and 9.

CCIC indicate that the modelling and estimation methodology and used in the 2007 update was

applied in the 2008 update using the same estimation parameters.

The following key changes were made to the models in the 2008 updates:

Zone 8 and 9: The geological wireframes for Zones 8 and 9 were updated to account for

underground mapping information of the historical mining areas undertaken by Katanga



Mining under the supervision of the Chief Geologist during 2008. The updated underground

stope sheets were digitised by Katanga as 2-dimensional strings and presented to CCIC to

update the 3D stratigraphic and structural wireframes models. Assumptions were made

defining the extent of the stopes in the Z-direction to ensure that the mineralised zones were

depleted in the locations within the areas of the strings. This updated wireframe was then re-

estimated using the same estimation parameters as in the 2007 update, but with modified

search criteria to prevent the smearing of high grade samples from the well informed areas

into the poorly informed areas. An octant search criteria was applied, which allowed for at

least two octants to be filled for estimation.

Zone 9 above 415 level: based on the advice of Katanga personnel the wireframe models for

Zone 9 were restricted to below 415 level. Katanga personnel reviewed the historical mining

above 415 level and indicated that historical mining had depleted the mineralised zones

above this level. This restriction has resulted in a reduction of 637kt classified as Measured

in the 2007 update, 5,7Mt of Indicated and 1,59Mt of Inferred Mineral Resources.

Division 5 above the 415 level: Katanga personnel also reviewed the historical mining

above the 415 level in Division 5 and updated the stope outlines. Their investigations

indicated extensive mining above the 415 level in the Division 5 area with mineralised zones

only in remnant pillars, which were difficult to quantify. On the advice of Katanga

personnel, the Division 5 above 415 level was removed from the Mineral Resources,

amounting to 2.9Mt of Indicated Mineral Resources.

Kananga Mine

Similar procedures to those adopted for Tilwezembe Mine were applied for Kananga Mine.

Top cutting was applied;

The mineralized zones were split into two structural domains;

Composites used in the estimation process were sourced within search criteria of 180m x

60m x 10m in the strike, across strike and perpendicular to strike directions equivalent to the

variogram ranges;

Where blocks remained unestimated after the first pass, the search neighbourhood was

doubled or tripled; and

Ordinary kriging was used in the estimation process.

KOV Mine

Grades were estimated into the fundamental block size of 100 m x 100 m x 5 m in the X, Y and Z

directions respectively. The geological model remained on the 20m x 20m x 5m block size, but the

blocks were reorganized to fit into a larger block size for the estimation.

Grade estimation was by ordinary kriging for %TCu and %TCo using the omni-directional

variogram parameters and the composite data. Each lithology type was estimated separately using the

respective composite data and parameters. Samples estimating any given block were sourced from

the data set using a search ellipsoid oriented along the dip of the mineralized zone with the maximum



search radius being in the planar of the deposit. Maximum search radii for copper and cobalt were

400 m and 300 m respectively, while the search in the Z direction was limited to 20 m.

Estimation was undertaken in three passes: in the first pass, the minimum and maximum number of

samples to estimate a block were set at 10 and 20. In the second pass, the search neighbourhood was

expanded to twice the original search of 400 m and 300 m respectively for %TCu and %TCo, and the

minimum and maximum numbers of samples to estimate a block were reduced to 5 and 10. Blocks

that remained unestimated after the second pass were now estimated within a search neighbourhood

of 5 times the original search, and the minimum and maximum numbers of samples reduced to 3 and

5 respectively.

For the estimation of the RATGR lithology, the variogram parameters for the DSTRAT were used.

For the FNSR, variogram parameters for the respective lithologies from the Virgule were used in the

estimation process and under similar sample-search criteria. For Kamoto East, the inverse-distance-

squared method was used for the estimation of the grades into the lithologies. This is mainly due to

the insufficient sample coverage within each of the lithologies. After the estimation into the full RSC

model, the block model for the mid-RSC containing the assigned weighted average grade values was

then added to the full RSC model overprinting the estimated block values within the mid-RSC. This

was applicable to all the fragments.

Mashamba East Mine

The estimation procedure as described under T17 was also applied to Mashamba East. The

variogram parameters for the lithologies within Mashamba East Mine are shown in Table 19.31. The

first search neighbourhood was set equivalent to the variogram range.

Table 19.31 Mashamba East Mine: Variogram parameters by lithology

Zone Variable Nugget Variance, C0 C1 Variance Range (x,y) Range z Variable

RSF %TCu 1,943 1,425 157,67 33,33 %TCu

%TCo 0,04157 0,08539 332,59 28,03 %TCo

RSC %TCu 1,006 0,6158 210,88 33,33 %TCu

%TCo 0,07737 0,03845 256,14 28,03 %TCo

SDB %TCu 1,235 1,564 216,8 33,33 %TCu

%TCo 0,05093 0,1141 291,64 28,03 %TCo

BOMZ %TCu 0,1709 2,731 353,88 33,33 %TCu

%TCo 0,02193 0,03127 467,49 33,33 %TCo

19.2 Density Determinations

Historically, Gecamines 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 Gecamines 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. Gecamines generalized empirical criterion defined

three main categories of densities as shown in Table 19.32 below.



Table 19.32 Gecamines criteria for assigning density values

Definition and Criterion Density, t/m3

Siliceous= (%TCu/%CaO) >=15

TCu>1<2% 2,0

TCu>2,0%, with <0,5% Cu sulphite content 2,2

TCu>1<2%, with >1% TCo content 2,2

TCu >2,0%, with >0,5% Cu sulphite content 2,4

Dolomitic = (%TCu/%CaO) <15

TCu >1<2% 2,4

TCu >2,0%, with <0,5% Cu sulphite content 2,4

TCu >2,0%, with >0,5% Cu sulphite content, >=0,5%CuOx 2,4

TCu >2,0%, with >0,5% Cu sulphite content, <=0,5%CuOx 2,6

According to the Gecamines criterion, waste rock was generally assigned a density of 2,0 t/m3 if it

was siliceous and 2,4 t/ m3 if the rock was considered dolomitic.

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.

19.2.1 T17 Mine

CCIC undertook limited density determinations of the various stratigraphic units to verify

Gecamines 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 T17 Mine are presented in Table 19.33 by stratigraphic unit.

Table 19.33 T17 Mine: Density Determinations on Various Lithologies

Stratigraphic Unit Number of samples Minimum Maximum Average

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

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 2,84 for the siliceous material from T17 West, and 2,74 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 Gecamines approach was

conservative and upside potential existed with regard to the calculated resource tonnages. However,

CCIC recommended that in-situ 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 T17 model, CCIC applied density values of

2,2 t/m3 and 2,4 t/m3 consistent with the Gecamines categories of oxide and mixed ore types.


Snowden undertook density determinations on selected core samples using Archimedes’ Principle.

The sample core pieces of approximately 100 mm to 200 mm 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 25 mE by 25 mN by 1 mRL that approximates the drill-hole spacing. The bulk

densities are presented in Table 19.34.

Table 19.34 Tilwezembe Mine: Density Determinations on Various Lithologies

Domain Bottom Cut Top Cut Percentage cut Declustered mean Declustered mean

(before cut) (after cuts)

Ox_MnDol 1,1 3 5 2,04 1,96

Ox_Brec - 2,7 4 1,90 1,81

Ox_TillArg - 3 5 2,09 1,98

Sl_MnDol - 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

19.2.3 Kamoto Mine


Gecamines 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 Kamoto U/G are presented in Table 19.35 by stratigraphic

unit.



Table 19.35 Kamoto Mine: Density Determinations on Various Lithologies

Stratigraphic Unit Number of samples Minimum Maximum Average Stratigraphic Unit

SD1a 9 2,69 2,90 2,80 SD1a

BOMZ 8 2,74 2,92 2,86 BOMZ

RSC 8 2,51 2,96 2,69 RSC

RSF 6 2,57 3,03 2,81 RSF

DSTRAT 5 2,66 3,02 2,81 DSTRAT

Grey RAT 3 2,64 2,77 2,70 Grey RAT

Red RAT 3 2,63 2,75 2,67 Red RAT

On the basis of the limited density determinations, CCIC concluded that Gecamines approach was

conservative and that upside potential existed with regard to the calculated resource tonnages, but

recommended that in situ bulk density determinations should be undertaken before higher density

values can be used in the Resource Model.

CCIC used the average density values of 2,7 t/m3 from the Gecamines table for the conversion of

volume to tons for the for the Kamoto Mine model.

19.2.4 Kananga Mine

The procedures adopted for the density determinations at Kananga are the same as described for

Tilwezembe and the values are listed in Table 19.36.

Table 19.36 Kananga Mine: Density Determinations on Various Lithologies

Domain Declustered mean

Upper ore body oxides (UOB_OX) 1,8

Middle low-grade oxides (MID_OX) 1,8

Lower ore body oxides (LOB_OX) 2,0

Upper ore body sulphides (UOB_SL) 2,1

Middle low-grade sulphides (MID_SL) 2,0

Lower ore body sulphides (LOB_SL) 2,1

19.2.5 KOV Mine

In the models generated for KOV, SRK used a density of 2,2 t/m3. This 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 Mine 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.



Gecamines’ empirical density values in the Mashamba East Mine. The method is as described above

under Kamoto Mine. CCIC’s limited density determinations for Mashamba East are presented in

Table 19.37, by stratigraphic unit.

Table 19.37 Mashamba East Mine: Density Determinations on Various Lithologies

Stratigraphic Unit Number of samples Minimum Maximum Average

SDB 17 2,34 2,76 2,52

RSC 10 2,40 2,61 2,51

RSF 5 2,28 2,50 2,39

CCIC used the average density values of 2,2 t/m3 and 2,4 t/m3 from the Gecamines table for the

conversion of volume to tons in siliceous and dolomitic mineralized zones respectively for the

Mashamba Mine model.



19.2.7 Summary

Table 19.38 indicates the densities that have been used in the conversion of volume to tons within

the various project areas.

Table 19.38 Density Determinations on Various Lithologies

Project Area Mineralized Zone Density, t/m3

T17 Mine Oxide mineralized zones 2,2

Tilwezembe Mine Ox_MnDol 1,96

Ox_Brec 1,81

Ox_TillArg 1,98

Sl_MnDol 2,26

Sl_Brec 2,24

Sl_TillArg 2,18

Kamoto Mine 2,7

Kananga Mine Upper ore body oxides (UOB_OX) 1,8

Middle low-grade oxides (MID_OX) 1,8

Lower ore body oxides (LOB_OX) 2

Upper ore body sulphides (UOB_SL) 2,1

Middle low-grade sulphides (MID_SL) 2,0

Lower ore body sulphides (LOB_SL) 2,1

KOV 2,2

Mashamba East Mine Oxide mineralized zones 2,2

Mixed mineralized zones 2,4

19.3 Block model validation

The block models from the respective projects were validated by comparing the mean values of the

data against the mean values of the estimates within a given search distance. The validation is mostly

specified along an easting, and composites and block model data are selected within search limits on

either side of the easting for the comparisons.



19.4 Mineral Resources and Classification

The Mineral Resources for all mines were estimated in accordance with the 2007 SAMREC Code.

Should the Mineral Resources be stated in accordance with the CIM standards on Mineral Resources

and Mineral Reserves, there would be no significant difference.

19.4.1 T17 MineMineral Resources for T17 Mine were estimated and reported by lithology. In classifying the

Mineral Resources, SRK considered the following:

The quantity and quality of the data used in the generation of the mineral resources, of

particular importance is the core recovery;

The unavailability of assays in portions of the package due to selective sampling on the basis

of visible copper mineralization;

The inconsistent cutting of the copper grades;

The relatively incomplete assays for %ASCu and %CaO compared to the %TCu data; and

Limited density determinations undertaken on the various lithologies.

SRK classified the mineralized zones of T17 Mine as Indicated and Inferred Mineral Resources. The

mineralized zones with limited drill-hole intersections were classified as Inferred Mineral Resource.

The Mineral Resources within the RSC lithology were also classified as Inferred.

Table 19.39 T17 Mine: Mineral Resources at 0% TCu cut-off (31 December 2008)

Indicated Inferred

Zone Mt %TCu %TCo Mt %TCu %TCo

OBI DSTRAT 2,8 3,67 0,33 1,1 4,09 0,76

RSF 3,3 2,64 1,04 1,5 3,57 1,84

OBS SDB 4,5 4,20 0,75 1,8 5,32 1,10

BOMZ 1,4 2,78 0,59 0,5 1,52 0,43

SD 1,7 0,92 0,14 0,6 0,78 0,09

RSC 11,2 0,80 0,32

13,7 3,16 0,64 16,7 1,77 0,57




The Mineral Resource for Tilwezembe Mine was estimated at 30 Mt at a total copper grade of 1,5%

and a total cobalt grade of 0,5%.

The following criteria were assessed in the classification as stated in the Snowden report:

The risk in the data informing the Mineral Resource estimate;

The robustness of the geological model; and

The risk in the grade estimates.



Snowden classified portions of the Tilwezembe Mineral Resources as: Indicated where the estimates

were within first search, equivalent to the variogram range; and Inferred for the remainder.

West of an easting of 357 300 mE, blocks above an elevation of 1200 mRL were classified as

Indicated. Between eastings of 357 300 mE and 357 525 mE, blocks above an elevation of 1245

mRL were classified as Indicated. All remaining blocks were classified as Inferred. The confidence

in manganese and soluble copper and cobalt assays was low, and these elements were therefore not

included in the resource statement. The Mineral Resource for Tilwezembe is reported above a %TCu

cut-off grade of 0,5%.

Table 19.40 Tilwezembe Mine: Mineral Resources at a 0,5% TCu cut-off (31 December2008)

Classification Mt %TCu %TCo

Total Indicated 9,0 1,89 0,60

Total Inferred 13,1 1,80 0,62





19.4.3 Kamoto Mine

Table 19.41 Kamoto Mine: Mineral Resources by zone (31 December 2008)

Classification Zone Mt %TCu %TCo

Measured 1 7,0 4,63 0,61

2 0,9 4,32 0,31

3 3,0 4,91 0,46

4 0,5 5,02 0,24

5 3,3 5,14 0,39

6 1,8 5,74 0,36

7 0,1 5,65 0,16

8 1,1 5,44 0,44

9 1,1 5,85 0,31

11 0,1 4,84 0,66

Etang N 2,6 2,85 0,61

Etang S 11,5 4,05 0,76

Sub-total 33,0 4,50 0,58

Indicated 1 3,9 5,35 0,82

2 1,9 4,79 0,59

3 2,4 5,62 0,5

4 1,6 5,27 0,35

5 1,8 6,03 0,45

6 2,1 6,02 0,27

7 7,2 5,65 0,31

8 0,5 4,16 0,39

9 0,5 5,64 0,3

11 0,8 5,14 0,69

Etang N 4,4 3,21 0,7

Etang S 8,7 3,28 0,89

Sub-total 35,7 4,69 0,60

Total Measured 1 10,9 4,89 0,69

and 2 2,8 4,64 0,5

Indicated 3 5,4 5,22 0,48

4 2,1 5,21 0,32

5 5,1 5,45 0,41

6 4,0 5,89 0,31

7 7,2 5,65 0,31

8 1,6 5,05 0,42

9 1,6 5,78 0,31

11 0,9 5,11 0,69

Etang N 7,0 3,08 0,67

Etang S 20,2 3,72 0,82

Total 68,7 4,60 0,59

Inferred 1 1,8 4,52 0,83

2 1,0 4,44 0,69

3 0,1 5,76 0,52

4 1,3 4,74 0,41

8 0,0 5,71 0,7

11 6,5 5,45 0,55

Sub-total 10,6 5,22 0,53





19.4.4 Kananga Mine

The Mineral Resource for Kananga Mine was estimated at 15,2 Mt at a total copper grade of 1% and

a total cobalt grade of 0,7%. The following criteria were assessed in the classification as stated in the

Snowden report:

The risk in the data informing the Mineral Resource estimate;

The robustness of the geological model; and

The risk in the grade estimates.

After consideration of all other items such as geological continuity and data quality, areas

predominantly estimated during the first search, equivalent to the variogram range, were classified as

Indicated and those estimated during the second search as Inferred.

Snowden classified portions of the Kananga Mineral Resources as Indicated where the estimates

where within first search, equivalent to the variogram range and the remainder as Inferred Mineral

Resource. West of 331 600 mE, blocks above an elevation of 1160 mRL were classified as Indicated.

Table 19.42 Kananga Mine: Mineral Resources at a 0,5% TCu cut-off (31 December 2008)

Type Classification Mt %TCu %TCo

Oxide Indicated 2,9 1,54 0,70

Sulphide Indicated 1,2 1,77 1,02

Total Indicated 4,1 1,61 0,79

Oxide Inferred 0,4 1,25 0,53

Sulphide Inferred 3,6 2,08 1,03

Total Inferred 4,0 2,00 0,98



19.4.5 KOV Mine

The bases for the classification include:

The quantity and quality of the data used in the generation of the Mineral Resource

estimates;




The relatively incomplete assays for %ASCu and %CaO compared with the %TCu data;

The unavailability of density determinations for any of the lithologies within each of the

fragments;

That there was a history of mining within each of the pits; and

The unavailability of historical production records for reconciliation.

The Mineral Resources for KOV have been classified as Indicated and Inferred.



The Mineral Resources for Virgule and Oliveira have been classified as Indicated on the basis of the

continuity of the geology and mineralization and the adequate drill-hole coverage.

The Mineral Resource for the Kamoto East has been classified as Inferred due to the paucity of data

intersecting the fragment.

Similarly, the mid-RSC in all the fragments has been classified as Inferred for the same reason, viz.

paucity of data.

Table 19.43 KOV Mine: Mineral Resources at a 0% TCu cut-off (31 December 2008)

Classification Indicated Inferred

Fragment Lithology Mt %TCu %TCo Mt %TCu %TCo

Virgule

BOMZ 6,1 3,92 0,59

SDB 17,8 6,13 0,52

RSC 14,5 4,43 0,22

RSF 11,4 5,91 0,21

DSTRAT 9,1 6,03 0,24

RATGR 6,9 5,84 0,11

RSC-MID 10,8 2,8 0,22

Sub-total 66 5,47 0,32 11 2,8 0,22

Oliveira

BOMZ 4,8 2,07 0,41

SDB 14,4 5,63 0,94

RSC 11,6 4,57 0,27

RSF 7,4 5,16 0,36

DSTRAT 5,3 5,13 0,3

RATGR 3,6 4,89 0,26

RSC-MID 16,5 0,94 0,47

Sub-total 47 4,82 0,51 16 0,94 0,47

FNSR

BOMZ 2,1 5,98 0,78

SDB 5,5 7,27 0,55

RSC 4,1 5,55 0,18

RSF 1,4 5,99 0,19

DSTRAT 0,8 6,96 0,04

RATGR 0,2 6,04 0,04

RSC-MID 3,6 1,13 0,08

Sub-total 14 6,41 0,41 4 1,13 0,08

Kamoto East

BOMZ 2,8 6,65 0,77

SDB 9,2 6,31 0,47

RSC 8,7 3,53 0,29

RSF 7,8 6,58 0,17

DSTRAT 4,4 4,85 0,13

RATGR 2 5,9 0,08

RSC-MID 5,6 2,2 0,23

Sub-total 40 5,04 0,31

TOTAL 126,9 5,33 0,40 71,2 3,56 0,32





19.4.6 Mashamba East MineThe Mineral Resources for Mashamba East Mine were estimated and reported by lithology. In

classifying the Mineral Resources, SRK considered the following:

The quantity and quality of the data used in the generation of the Mineral Resource

estimates; of particular importance is the core recovery;




The relatively incomplete assays for %ASCu compared with the %TCu data;

Limited density determinations undertaken on the various lithologies; and

That there was a history of mining Mashamba East, but historical reconciliation production

records are unavailable.

SRK classified the mineralized zones of Mashamba East (with the exception of the RSC)as Indicated

Mineral Resource.

Table 19.44 Mashamba East Mine: Mineral Resources at a 0% TCu cut-off (31 December2008)

Zone Mt %TCu %TCo

Indicated RSF 9 31,0 2,77 0,49

SDB 25 24,5 1,46 0,44

BOMZ 21 19,5 0,68 0,13

Total 75,0 1,80 0,38

Inferred RSC 43 65,3 0,76 0,10





19.5 Consolidated Mineral Resource Statement

Table 19.45 presents KOL’s consolidated Mineral Resource statement as of 31 December 2008.

Table 19.45 KOL: Mineral Resources as at 31 December 2008


Measured Kamoto Mine 33,0 4,50% 0,58%

Subtotal 33,0 4,50% 0,58%

Kamoto Mine 35,7 4,69% 0,60%


Indicated T17 Mine 13,7 3,16% 0,64%

KOV Mine 126,9 5,33% 0,40%

Kananga Mine 4,1 1,61% 0,79%


Subtotal 264,5 3,95% 0,45%

Kamoto Mine 68,7 4,60% 0,59%


Measured and T17 Mine 13,7 3,16% 0,64%

Indicated KOV Mine 126,9 5,33% 0,40%

Kananga Mine 4,1 1,61% 0,79%


TOTAL 297,5 4,02% 0,46%

Kamoto Mine 10,6 5,22% 0,53%


Inferred T17 Mine 16,7 1,77% 0,57%

KOV Mine 71,2 3,56% 0,32%

Kananga Mine 4,0 2,00% 0,98%


Total 180,7 2,32% 0,31%



19.6 Comparison of the 2008 and 2007 Mineral Resources

In Table 19.46, the 2008 Mineral Resources for the KOL assets are compared with those declared in

2007 with notes attached to explain the difference. In most cases the difference is due to the change

in the reporting adopted for 2008, where the Mineral Resources are declared inclusive of Mineral

Reserves. In 2008, an optimistic pit model has not been run to determine the declared Mineral

Resources as this method excludes any material below the optimistic pit from the Mineral Resource

base.

The Mashamba West and Dikuluwe properties were returned to Gecamines and therefore do not

form part of the 2008 Mineral Resource and Reserve Statement.



Table 19.46 KOL: Comparison of the 2008 and 2007 Mineral Resource statements

2008 2007

Classification Project Area Mt %TCu %TCo Mt %TCu %TCo

Kamoto Mine(1) 33,0 4,50% 0,58% 35,3 4,56% 0,58%

Measured Mashamba East Mine(2), (4) - - - 29,1 2,30% 0,47%

T17 Mine(2), (4) - - - 8,7 3,69% 0,73%

Mashamba West (4), (5) - - - 13,6 2,97% 0,18%

Dikuluwe (4), (5) - - - 24,1 3,89% 0,10%

Sub-total 33,0 4,50% 0,58% 110,9 3,56% 0,41%

Kamoto Mine(1) 35,7 4,69% 0,60% 48,9 5,03% 0,58%

Mashamba East Mine(2), (4) 75,0 1,80% 0,38% 11,0 1,89% 0,43%

Indicated T17 Mine(2), (4) 13,7 3,16% 0,64% 2,9 3,55% 0,82%

KOV Mine 126,9 5,33% 0,40% 126,9 5,33% 0,40%

Kananga Mine(3) 4,1 1,61% 0,79% 0,0 0,00% 0,00%

Tilwezembe Mine(3) 9,0 1,89% 0,60% 0,0 0,00% 0,00%

Mashamba West (4), (5) - - - 3,1 2,87% 0,13%

Dikuluwe (4), (5) - - - 8,8 4,19% 0,09%

Sub-total 264,5 3,95% 0,45% 201,7 4,96% 0,43%

Kamoto Mine(1) 68,7 4,60% 0,59% 84,3 4,83% 0,58%

Total Mashamba East Mine(2), (4) 75,0 1,80% 0,38% 40,1 2,19% 0,46%

Measured and T17 Mine(2), (4) 13,7 3,16% 0,64% 11,6 3,65% 0,75%

Indicated KOV Mine 126,9 5,33% 0,40% 126,9 5,33% 0,40%

Kananga Mine(3) 4,1 1,61% 0,79% 0,0 0,00% 0,00%

Tilwezembe Mine(3) 9,0 1,89% 0,60% 0,0 0,00% 0,00%

Mashamba West (4), (5) - - - 16,7 2,95% 0,17%

Dikuluwe (4), (5) - - - 32,9 3,97% 0,09%

TOTAL 297,5 4,02% 0,46% 312,6 4,46% 0,42%

Kamoto Mine(1) 10,6 5,22% 0,53% 12,2 5,30% 0,57%

Mashamba East Mine(2), (4) 65,3 0,76% 0,10% 5,3 2,14% 0,58%

Inferred T17 Mine(2), (4) 16,7 1,77% 0,57% 0,0 0,00% 0,00%

KOV Mine 71,2 3,56% 0,32% 71,2 3,56% 0,32%

Kananga Mine(3) 4,0 2,00% 0,98% 4,0 1,44% 0,74%

Tilwezembe Mine(3) 13,1 1,80% 0,62% 13,1 1,59% 0,65%

Mashamba West (4), (5) - - - 0,0 0,00% 0,00%

Dikuluwe (4), (5) - - - 9,8 4,29% 0,08%

Total 180,7 2,32% 0,31% 115,7 3,44% 0,39%

2008 Mineral Resources are quoted inclusive of Mineral Reserves

(1) Re-modelling and depletions by CCIC.

(2) Original CCIC model incomplete, SRK re-estimated and re-classified.

(3) Additional drilling and re-modelling.

(4) 2007 Resources constrained by optimistic pit outline.

(5) Properties returned to Gecamines.



19.7 Consolidated Mineral Reserve Statement

Table 19.47 presents KOL’s consolidated Mineral Reserve statement as at 31 December 2008. A

discussion on the LoM plans and mining assumptions in support of the Mineral Reserves is included

in Section 25a.

Table 19.47 KOL: Mineral Reserves as at 31 December 2008


Proved Kamoto Mine 17,0 3,52% 0,51%

Subtotal 17,0 3,52% 0,51%

Kamoto Mine 19,4 3,70% 0,53%


Probable T17 Mine 3,1 2,67% 0,70%

KOV Mine 90,1 4,93% 0,38%

Kananga Mine 0,0 0,00% 0,00%


Subtotal 122,8 4,64% 0,43%

Kamoto Mine 36,4 3,62% 0,52%


Proved and T17 Mine 3,1 2,67% 0,70%

Probable KOV Mine 90,1 4,93% 0,38%

Kananga Mine 0,0 0,00% 0,00%


Total 139,8 4,50% 0,44%


(2) Mineral Resources are inclusive of Mineral Reserves.

19.7.1 Comparison of the 2008 and 2007 Mineral Reserves

Table 19.48 indicates the reconciliation between the Reserve Statement for 31 December 2008 as

presented in this document to the Reserve Statement issued by KOL during February 2007. Notes

on reconciliation between 2008 and 2007 Reserve Statements:

Kamoto Mine

o The Resource model had been updated and had an effect on the Reserve Statement

per zone. This comment is applicable to all the grade values per mining area;

o All geological losses were reduced from a 10% loss to 5%. The reduction is

considered to be valid, since the nature of the resource is consistent and no major

discontinuities are expected;

o Three areas had been redesigned, where the previously proposed mining method

(cut-and-fill) would not be practical. The zones were redesigned for PPCF;

o Zone 9 had been depleted significantly (by about 1,8 Mt);

o Zone 10 and Division 5 had been removed in totality from the resource

(approximately 6 Mt);

o Zone 8 Reserves had been reduced by 53% due to a change in the geological model

(Zones 8 and 9 were split); and



o Probable Reserves all increased by 5% due to the change in the geological loss

value.

T17 Mine

o New geological model changed the Mineral Resource classification and thus

changed the Mineral Reserves; and

o RSC was classified as Inferred in the 31 December 2008 Statement.

Mashamba East Mine

o New geological model changed the Mineral Resource classification and thus

changed the Mineral Reserves; and

o RSC was classified as Inferred Resources in the 2008 Statement.

Table 19.48 KOL: Comparison of the 31 December 2008 and 2007 Mineral Reservestatements

2008 2007

Classification Project Area Mt %TCu %TCo Mt %TCu %TCo

Kamoto Mine 17,2 3,53 0,51 17,1 3,63 0,41

Proved Mashamba East Mine - - - 15,6 2,82 0,46

T17 Mine - - - 1,1 3,38 0,36

Mashamba West(2) - - - 4,6 3,31% 0,12%

Dikuluwe(2) - - - 15,9 3,59% 0,10%

Sub-total 17,2 3,53 0,51 54,3 3,35% 0,31%

Kamoto Mine 19,4 3,70% 0,53% 28,4 3,99% 0,52%

Mashamba East Mine 10,2 4,39% 0,52% 3,7 2,64% 0,54%

Probable T17 Mine 3,1 2,67% 0,70% 0,5 2,96% 0,39%

KOV Mine 90,1 4,93% 0,38% - - -

Kananga Mine - - - - - -

Tilwezembe Mine - - - - - -

Mashamba West(2) - - - 1,3 3,00% 0,09%

Dikuluwe(2) - - - 4,9 3,46% 0,10%

Sub-total 122,8 4,64% 0,43% 38,8 3,75% 0,45%

Kamoto Mine 36,4 3,62% 0,52% 45,5 3,85% 0,48%

Total Mashamba East Mine 10,2 4,39% 0,52% 19,3 2,79% 0,48%

Proved and T17 Mine 3,1 2,67% 0,70% 1,6 3,24% 0,37%

Probable KOV Mine 90,1 4,93% 0,38% - - -

Kananga Mine - - - - - -

Tilwezembe Mine - - - - - -

Mashamba West(2) - - - 5,9 3,24% 0,11%

Dikuluwe(2) - - - 20,8 3,56% 0,10%

TOTAL 139,8 4,50% 0,44% 93,1 3,52% 0,37%


(2) Properties returned to Gecamines.



20 Other Relevant Data and Information

20.1 KOV Oxide / Sulphide Content

20.1.1 Description

The KOV database contains a total of 275 holes, of which about 215 were drilled during the

Gecamines era, spanning 40 years from 1940s to the 1980s. During the Gecamines era, not every

sample analyzed for %TCu, was analyzed for %AsCu: the combined KOV assay database contains a

total of 7180 samples analyzed for %TCu, out of which 3737 or 52% were analyzed for %AsCu.

Several reasons have been put forward by various workers in the DRC for the inconsistent sampling

practice of which the most common is that the sampling was based on visual inspection of the

minerals, implying oxide mineralization in the samples not analyzed for %AsCu. There is no

consistency in the distribution of the %AsCu assays compared with those for %TCu. In some cases

there is only one sample in the entire drill-hole analyzed for %AsCu. This has made it difficult to use

the assay database to determine the proportions of oxide or sulphides or both in the mineralized

zones.

SRK selected samples that have assays for both %TCu and %AsCu (the AsCu database) and used

these for the determination of the proportions of oxide or sulphide or both, expressing this as a ratio

of %AsCu to %TCu. Four main fragments have been identified in the KOV project area: Kamoto

East, Virgule, Oliveira and FNSR. This oxide/sulphide study was based on the distribution of the

ratios in each of the four individual fragments.

Samples falling within a 15 m bench slice were extracted from the AsCu database in each of the four

project areas. The bench centre was consistent with the block model centres, and the sample

selection criterion was 7,5 m on either side of the centre. Statistics were computed for the samples

falling within a particular bench and within a particular fragment for %TCu, %AsCu and Ratio.

Comparative statistics were also computed on the entire KOV database for %TCu.

The initial study of splitting the samples by ore type into OBS, RSC and OBI for each of the

fragments indicated insufficient OBS and RSC sample counts for any meaningful conclusions to be

drawn. Higher sample counts were observed in the OBI, which contributes immensely to the study.

For the final study, no sub-selections were done. Instead, samples were selected across the entire

stratigraphic succession from the BOMZ to the RATGRIS, disregarding the ore-type boundaries

completely.

Ratio distribution plots by bench inclusive of %AsCu and %TCu grades and number of samples

within the bench slice were generated. A fourth-order polynomial was fitted to the ratio data for each

fragment to describe the best-fit equation.

The polynomial best-fit equations have been applied to the each block in the model to provide an

estimate of the ratio. SRK caution against using the ratio as absolute indication of the proportions of

oxide or sulphides or both in the mineralized zones because the quantity and quality of the data used

to derive these expectations were low.



KML have indicated that they will devise a drilling and evaluation plan to improve the

understanding of the oxide/sulphide ratios.

20.1.2 Assumptions

Oxides and sulphides require different processing routes. It has been assumed that the oxide:sulphide

ratio for the ore mine from KOV Mine is 60%:40%, based on the study outlined in Section 20.1.1.

Should this ratio be significantly different to this assumption, it will affect the capacity of each

process route and may restrict production, which will affect sulphur consumption and costs.

20.2 Dewatering

20.2.1 Introduction

This section is based on information provided to SRK in AGES Technical Reports AS-R-2008-09-05

and AGES-R-08-01-28, AG-R-2008-11-24 KOL Geohydrology Report Version 3 DRAFT and AG-

R-2008-10-02 KOV Dewatering DFS V2 Final.

AGES and AfriCon were appointed by KML to conduct a study for the dewatering of the KOV pit.

AGES was responsible for the mine dewatering concept and pit slope dewatering borehole layout,

while AfriCon undertook the pit lake dewatering design and implementation and engineering layout.

It is understood that the purpose of the investigation was to determine the geohydrological and

engineering requirements and options for the pre-operational and life of mine dewatering of KOV.

Based on hydrogeological data generated through a programme of drilling and sampling, AGES

constructed a hydrogeological model which was used to develop the dewatering strategy.

20.2.2 Kamoto Underground

The hydrogeological characteristics of the water-bearing mine series formations (CMN, RAC and

SD) are variable depending on the degree of alteration. Current water inflows into the underground

workings have been estimated by KOL to be 2 350 m3/hr. KOL is making provision to increase the

volume pumped to 6 900 m3/hr based on historic pumping capacity.

A geophysical survey in the area between the mine workings and the Mupine pit has indicated a zone

of high conductivity that could potentially be intercepted. A number of preferential flow paths are

believed to exist, which could connect Mupine to the underground workings. With the deposition of

tailings into the Mupine pit, it is estimated that the head of water in the pit will rise by approximately

40m. AGES has estimated an additional 500-800 m3/hr will flow into Kamoto from the Mupine Pit

due to the increased head. However, in the longer term, the deposition of the fine tailings material is

expected to reduce the permeability of the base of the pit, reducing the potential for seepage, with

predicted inflows decreasing to 300 m3/hr. It is recommended that the borehole MUP05 proposed by

AGES be drilled and pump tested and additional isotope samples be taken from inflows into the

underground workings and from the new boreholes MUP1, MUP2, MUP3 so that a more detailed

understanding of the consequence of tailings disposal in the Mupine pit can be gained.



20.2.3 T17

The mine plan put forward for the 2008 Study has been changed significantly in the Addendum

study to allow for greater tonnages to be mined over the first two years. With regard to the

geotechnical validation of the new design put forward by KOL, a number of parameters have

changed, based on the drawings presented in December 2008 (SRK has been provided with a mining

plan, but more detail is required for geotechnical analysis). The changes include:

The floor elevation of the pit has dropped from approximately 1340m to approximately

1 300 m. This means that the pit walls will be higher than those analysed in the 2008 Study

and there is a greater chance that the lower portion of the pit will intersect the water table.

The access ramp positions at the end of 2010 will lie predominantly in the hanging wall and

not in the footwall as with the design assessed in the 2008 Study. This means that the overall

angle of the footwall slope in RAT will be greater than that analysed previously and the

hanging wall will be flatter.

The northern wall of the pit will intersect the (possibly saturated) backfill in the Musonoi pit,

possibly to a greater depth than envisaged in the 2008 study. The overall slope angle in this

portion of the pit should be less than in intact hanging wall strata. No details of the slope

design adjacent to the Musonoi pit have been provided yet for review.

The pit appears to extend further to the west than in the previous design. There appears to be

a more complex geological structure towards the west with significant faulting being shown

on geological sections X +850 and X +950. No detailed geological plans of this area have

been provided. Should this faulting be adversely orientated, the security of the access ramp

for Cut 4 could be in question.

The potential for water pressure development in both footwall and hanging wall slopes has

not been established. Limited water level information has been presented by AGES. A

provision for dewatering using horizontal drains has also been suggested by AGES but no

technical basis for the dewatering design has been provided for review by SRK. It is SRK’s

opinion that horizontal drain holes will not effectively address operational problems such as

wet working areas and water in blastholes.

The pit bottom is already below the water level in the Musonoi River. As a result, there is

the risk that if there is any failure on the south-eastern high wall the Musonoi River could

inundate the pit.

To validate the design SRK would require the following work to be carried out, based on the

assumption that the geotechnical parameters determined for KOV will also be applicable to T17 as

no geotechnical information is available for T17:

It is understood that a pit design is available Although the 2008 design indicated that the

slopes carried safety factors in the order of 2, because of the changes in pit geometry these

slopes must be re-analysed and the risks re-quantified for the new design.

The water pressures prevailing in both hanging wall (including Musonoi backfill) and, in

particular, the footwall RAT must be established as a matter of urgency. Should water at a

pressure that could impact on slope stability be identified, the permeability of the associated



strata must be established as a basis for designing a depressurisation system. The alternative

of implementing a system of vertical pumping holes should be investigated. As the pit

appears to be relatively dry to an elevation of approximately 1345m, a system of vertical

holes need not necessarily be implemented from surface but could be established on a lower,

in-pit elevation to reduce implementation time and cost.

A definitive structural geological model for the western portion of the pit must be created

and reviewed to assess the risk of structurally bound instability occurring in the Cut 4 access

ramp and consequent risk to 2010 and 2011 production profiles.

In view of the current uncertainty regarding geotechnical design parameters and water pressures,

SRK has recommended that a geotechnical monitoring programme is implemented immediately,

incorporating regular (weekly) pit rim and bench inspections, establishment of a network of

monitoring points that are surveyed monthly and installation of a network of piezometers that are

monitored weekly. Although this will not prevent a failure, early recognition may allow a redesign

to reduce the impact of failure and will also reduce the risk of injury to personnel and the loss of

equipment.

20.2.4 Mashamba East

This is located on the southern edge of the Roan Basin outcrop and is characterised by several

southwest to northeast trending synclinal/anticlinal structures and southeast to northwest trending

faults which reportedly have a compartmentalising effect on the groundwater in the area. As mining

is not scheduled to commence until 2018, a discussion of the dewatering of this pit is not included as

the dewatering design will need to be updated based on the findings of the KOV dewatering exercise

discussed in detail below.

20.2.5 Dewatering of the KOV pit

AGES has stated that the dewatering design has been based on the 2006 mine plan, although the

costing has been updated using the information made available for this study in July 2008. It is

AGES opinion that the original dewatering design will be appropriate for the 2008 mine plan.

It is understood that the AGES dewatering strategy is to reduce the groundwater head and capture

most of the water before it flows into the open pit using both in-pit and ex-pit boreholes, i.e. it is

aimed at intercepting the maximum volume of water prior to entering the pit but not necessarily

drawing down the phreatic surface on the high walls uniformly around the pit void.

SRK believes that the objective of mine dewatering must be to depressurize the pit slopes and ensure

that the flows into the pits are manageable to enable safe mining as cost effectively and efficiently as

possible. However, the strategy proposed by AGES is focused on maximising dewatering volumes

and if this does not result in the lowering of the phreatic surface as envisaged, the pit slopes may not

be sufficiently depressurised.

The key potential risks that are therefore associated with the proposed dewatering strategy are:



As some of the assumptions in the hydrogeological model have not yet been validated, the

dewatering strategy could potentially prove to be inefficient as the boreholes may not be

correctly located and/or an insufficient number planned.

If the dewatering strategy proves to be inappropriate, the assumption used in the mine

planning that dry slope conditions will prevail could be incorrect, which could potentially

impact on both the economics of the pit and pit stability, and thus on the production

schedule.

There are other potential risks that could impact on the pit dewatering, such as the availability of

electricity, site security and changes in mine plan. Although important, the significance of these

risks is considered to be lower than the key potential risks cited above and thus they are not

discussed further in this section.

A detailed list of recommendations for additional work that must be undertaken to resolve the

areas of uncertainty has been compiled and is presented below. Based on the outcomes of this

additional work, suitable mitigation measures to adequately manage the risks can be developed.

Dewatering Strategy

The implications for the mine plan if the drilling method/equipment does not achieve the

anticipated timeframes for borehole installation need to be assessed and an alternative

dewatering strategy proposed.

The geohydrological characteristics of the individual strata making up the KOV rock mass

must be identified using both laboratory and field test work. Packer testing and falling head

tests must be done to identify the characteristics of specific horizons that may require

depressurization.

Based on the outcome of the packer testing, point piezometers and stand pipes in open

boreholes should be installed to understand the pore pressures and phreatic surface acting in

the different stratigraphical horizons on the pit high walls, both during pre-operational and

operational stages. This will allow assessment of the degree of success of the dewatering

strategy in lowering the hydraulic head.

A more detailed conceptual hydrogeological model should then be developed showing cross

sections north-south and east-west showing the simulated hydraulic head within the range of

possible hydraulic conductivities.

Based on the above information, an assessment should be made on the adequacy of the

number and location of the vertical dewatering boreholes. This information can also be used

to optimise the placement of in-pit horizontal boreholes.

In terms of the timeframes for dewatering, sensitivity analyses should be undertaken for

specific parameters such as recharge, hydraulic conductivities (both vertical and horizontal)

for different predictive simulations (i.e. varying the number of boreholes, pumps, passive

drainage to underground ring tunnel, drainholes etc) with the numerical flow model. This

will enable a more accurate assessment of the risks related to the proposed dewatering

strategy.



The timing of the KOV and Kamoto East dewatering needs to be presented in more detail as

the management of the water levels in the pits relative to each other is critical in maintaining

pit stability along the south wall of KOV.

Estimates of drain spacing must be undertaken on the south wall of the pit to ensure that this

formation is depressurised and, once constructed, shut in pressure tests should be used to

measure the head and hydraulic conductivity to optimise the planned drain spacing.

Additional drilling and pump testing is required to demonstrate the efficacy for dewatering

of the pit of pumping more than 1 km from KOV in the Variante, as well as to establish if

the water to be pumped from dewatering boreholes DCP04 and DCP05 are linked to seepage

water from the Kingamyambo tailings toe dam rather than intercepting water from the

Musonoi River.

Verification of the appropriateness of the location of the boreholes between KOV and the

Musonoi Pit should be undertaken.

Different options should be assessed and recommendations presented should it be found that

the condition of the benches for in-pit boreholes is unsuitable, taking into consideration that

although additional pumps will drain the pit lake, they will not lower the surrounding

phreatic surface as effectively as in-pit boreholes.

Alternative strategies should be investigated for dewatering if it is found not to be possible

to achieve the desired drawdown once the current strategy is implemented. For example, an

assessment of whether dewatering galleries or boreholes drilled from the Kamoto

Underground into KOV may be more effective in dewatering than the planned borehole

dewatering scheme should be evaluated in more detail as it may well prove to be the better

option technically/financially/environmentally.

The investigation proposed by AGES into suitable methods for controlling the assumed flow

from the Musonoi River towards KOV during the operational phase should be carried out to

establish if any of the options will be possible and how cost effective they are likely to be.

Groundwater Model

SRK believes that additional work is required to demonstrate correlation between the model and

field observations and data.

A gap analysis should be undertaken to state clearly the information that is still required and

assumptions that have been used in the calibration of the model. For example, how was the

recharge value used in the model determined?

It has been stated by AGES that there will now be a longer lead time for dewatering to be

effective for the southern part of the pit as cut 1 is now going to the north. This scenario

needs to be modelled to demonstrate that this is in fact the case.

The aquifer parameters determined from pump testing and/or packer testing for specific

stratigraphical units should be built into the model on a more detailed level and fed back into

the detailed conceptual hydrogeological model.



Using the more detailed conceptual hydrogeological model, the position of the phreatic

surface and potential heights of the seepage faces in the high walls should be simulated

within different geotechnical domains.

The position of the phreatic surface (hydraulic head) at quarterly intervals for the first 24

months needs to be modelled, followed by its position annually for the life of mine. The

probability of achieving these phreatic surfaces must also be determined.

Additional drilling and pump testing should be undertaken to establish that the RAT does in

fact form an impermeable barrier for regional groundwater flow as this is a critical

assumption in the model.

The model should be updated with the data from drilling currently being undertaken to

demonstrate that the aquifer on the eastern side of the Musonoi River is in fact being

impacted by the dewatering of KOV. This data will be critical in resolving the issue of the

source of flow into KOV and should be used to refine the dewatering strategy in the future.

Additional scenarios should be run to assess the risks if in-pit boreholes prove not to be

possible or only partially possible and pumping from the bottom of the pit becomes the

primary dewatering method. An assessment of how long it will take to draw down the

phreatic surface around the pit is required, assuming only passive dewatering i.e. pumping

from the pit sump only.

The real and measured losses from the Musonoi catchment into the Kakifuluwe River should

be incorporated into the model.

An assessment of the expected impacts on groundwater users needs to be undertaken if part

of the inflow into KOV comes from the east side of the Musonoi River.

20.3 Infrastructure

This section is separated into two parts: general infrastructure and power.

20.3.1 General Infrastructure

This section summarises new infrastructure proposed as part of the Bateman Engineering 2008 Study

to support the merged process facility and particularly the new WOL/SX/EW plant. Bateman,

following agreement with KML, considered the following proposed new infrastructure as part of its

scope of work. The revised work based on the 2008 Study contains many new assumptions and

reductions in scope for general infrastructure to reflect the reduced capacities. These areas require

further detailed assessment, especially where existing buildings may be intended for use:

Plant Laboratory: The plant laboratory footprint is 94,5 x 19,4 m. Space has been allocated

for future expansion. An external bulk store with a footprint of 12,2 x 11,1 m has been

included to cater for all geological coarse reject samples. The area will be terraced and

surrounded by a security fence and associated gates. All necessary benches, basins with

splash backs, fume hoods, shelves and cupboards, glass ware, utensils, consumables, acids,

buffer solutions, reagents and reagent standards, spares, laboratory furniture, trolleys, office

furniture and stationery have been included. The building will be brick clad with steel rafters

and sheeting. Ceilings are included at 2,6m elevation over general areas and concrete slabs

over fire risk areas.



Environmental Laboratory: The following is included: A brick clad building with steel

rafters and IBR sheeting with a footprint of 20,55 x 13,0 m. The area will be made up of a

main entrance with air lock, including a reception area; an office; a sample receipt area with

air lock; an effluent preparation room; a water preparation room; an environmental

instrument room; an ICP instrument room; a balance room with air lock; a wash up room; a

store room and a change room /ablutions. The selection of additional equipment subject to

options selection is yet to be finalised.

Workshop: Provision of sheeted structural steelwork frame covering a footprint of 900 m2

with the eaves height set at 5 m from final floor level. An allowance has been made for a

10 t overhead electric travelling crane. Roller shutter doors will all be chain operated.

Stores: Provision of sheeted structural steelwork frame covering a footprint of 1150 m2

with the eaves height set at 5 m from final floor level. Roller shutter doors will all be chain

operated.

Offices: Provision for a single storey prefab type building with a footprint of 2000 m2. No

allowance has been made for the refurbishment of the existing administration office. No

allowance has been made for general furniture and fittings as all these will be supplied by

KOL.

Workshop/Stores Offices: The existing administration office will be used and thus no

allowance has been made.

Change House: The provision for a single storey prefab type building with an approximate

footprint of 600 m2. This footprint includes the changing facilities for male and females,

ablutions, showers and laundry services. Provision of associated 400 no. off lockers, benches

and laundry equipment is included.

Production Control Room: Provision for a single storey prefab type building with an

approximate footprint of 600 m2. This footprint includes the offices, meeting room,

ablutions, kitchen, store and main control room.

Fencing and Security: A provision to build a new 2300 m long masonry boundary wall

(3 m high) has been made on the new proposed boundary line around the lime, limestone

plant, the collection pods, SX / EW and HV yard on the northern side. A provision for

internal fencing of a 3,6 km long, 2,4 m high, welded mesh fence and associated gates have

been made.

Earthworks and Terracing: Provision of all site clearance, bulk earthworks, terracing and

earthworks. The limits of clearing will be the toe-line of the terrace required for the

construction of the plant (approximately 28 ha footprint in total). Included are storm water

ponds, fire water pond, SX ponds and SX tank farm, excavations and slope formations.

Bitumen Surfaced Roads: An allowance has been made to patch and reseal the existing

sealed roads and construct new sealed roads for construction vehicles.

Wearing Coarse Treated Roads: An allowance has been made to construct new gravel

roads.



Storm water Channels and Berms: An allowance has been made to construct and line

open channels with grouted stone pitch and concrete.

Storm Water Pipes: Included is approximately 635 m of 600 mm diameter storm water

pipes for the divergence of existing storm water drainage within the Luilu plant.

Ponds: An allowance has been made for excavations to receive HDPE linings, excavations

for trench anchorage, (supply and installation of HDPE linings is allowed for under Sub

Area K21), spillways, scour pipes, valves and overflows.

Landfill Site: An allowance has been made for landscaping.

Sewerage Reticulation: Provision of sewerage reticulation and manholes for the change

house only. Note that the sewerage treatment plant will be provided and operated by KOL.

All reticulation is limited to within the process plant boundary on the plant site.

Construction Services, Offices, Store and Vehicles: A provision for a 24 x 48 m

construction store including an office, toilets and associated services to be erected in the

laydown area. The provision for an office and ablutions with a set footprint of approximately

80 m2.

Sulphur Storage and Cathode Loading: The track condition is such that the 740 m of

track on existing alignment can be refurbished through fettling. A further 1440 m of new

track, including 3 turnouts must be constructed. This is for the separate acid loading facility

as well as repositioning the existing crossover towards the operational plant. The latter is

required to provide continued access to the operational plant. All of the above have been

allowed for. For electrowinning, the condition of the track infrastructure over the entire

length is such that all track components will have to be scrapped and replaced. A total of

3360 m of new track and 10 turnouts must be constructed. All of the above have been

allowed for including scrapping and diesel off loading.

In-motion Weighbridge: A provision has been made for a high accuracy in-motion train

weighing system complete with lightning protection. Included is a 6 m air-conditioner

insulated container type trackside hut. The calibration of the weighbridge is included.

Storm water Collection Ponds: These are included with anchor trenches and associated

concrete.

Laydown Area: A provision has been made for a hard stand laydown area with a set

footprint of 404 m long x 280 m wide. The laydown area will be fenced off with a main gate

and provision for a security kiosk at the entrance.

Construction Camp: Accommodation units for 124 people (total of 1346 m2) have been

allowed for. Provisions for internal water and sewer reticulation of the ENFI camp have

been allowed for.

Accommodation units: 65 Accommodation units (total of 30 745 m2) and associated

services have been allowed for. Each unit comprises of 8 rooms.

Recreation: Three recreation buildings (total of 1242 m2) have been allowed for.



Kitchen, Dinning Rooms and Stores: Two pre fabricated kitchens and dinning rooms (total

of 2528 m2) have been allowed for.

Offices for Camp Management: Provision of 112 m2 for offices. The reception, clinic,

store and security office are 56 m2 each. Provision for a gatehouse with a footprint of 30 m2.

Screened Area for Kitchens: Provision for 78 m long x 1,8 m high court yard walls and a

100 mm thick concrete slab set at 380 m2 with a sump and grid.

Sewer Reticulation: A provision for all sewerage reticulation and manholes has been made

with a sewer line which is 3,2km long with a total of 73 manholes.

Water Reticulation: The water line is 4,7km long with a total of 21 valves and 10 hydrants.

Drainage for Storm Water: 5250 m2 of concrete lined V-drains and channels have been

included.

Roads: 25 300 m2 of gravel roads and associated walkways located within the plant have

been included.

Fencing: Provision of a 5 km long, 2,4 m high, 4 strand welded mesh fence has been made.

Bateman Village: This includes accommodation units for 150 staff and 30 guests (total of

2636 m2) have been allowed for. Each unit comprises of a 1 bedroom with en-suite bathroom

complete with shower, wash hand basin, toilet and associated services. Transportation to site

is included. Also included are recreation facilities; a kitchen; dinning rooms and stores; a

laundry room; offices for camp management and catering; a screened area for kitchens;

sewer and water reticulation; drainage for storm water; roads; fencing; and camp

management.

20.3.2 Power

The consumption of electricity by the various facilities of KOL in Kolwezi required to produce up to

310 ktpa of saleable copper peaks at 245 MW from Repartiteur Ouest (“RO”) substation from 2012

onwards.

At present the current operations are supplied at 120 kV. The bulk of the additional load requirement

will be from the WOL/SX/EW Refinery Project (136 MW) and it is proposed to supply that load at

220 kV from the “Station de Conversion de Kolwezi” (“SCK”) at 220 kV. For this part of the study,

KOL contracted the services of Trans-Africa Projects (“TAP”), an independent consultant who are

currently involved as EPCM Consultants in several projects for mining companies in Katanga

Province. Their area of expertise is high voltage transmission (up to 765 kV), mostly gained in

South Africa.

This section is based on a report prepared by TAP for KOL and on an assessment of LoM power

requirements by Bernie Cyr, an electrical engineer employed by KOL. The reliability of the SNEL

network for the supply of electricity to the facilities of KOL situated west of Kolwezi in the DRC is

investigated. TAP is also contracted to provide the design services for the high voltage infrastructure

for the mine.

The loads will consist of:



The Luilu Metallurgical Plant, which will take up to 74 MW through 2013, reducing to 60

MW after 2013;

The Kamoto Mine, Kamoto Concentrator and Mashamba East Mine dewatering supplied

from Kadi substation which will require 49 MW from 2014 onwards;

The KOV mine which will take initially require 30 MW and stabilise at 32 MW; and

The WOL/SX/EW Refinery Project, where the load will progressively increase from 61 MW

in 2014 to 105 MW in 2015 and stabilise at that level afterwards.

These loads could be adjusted at a later stage to guarantee a production of 310 ktpa of copper per

year. It is also assumed that demand side management will maintain the combined load of KOL to

approximately 85% of the sum of the individual peaks.

The HV Katanga network of SNEL is made of 220 kV and 120 kV lines and a number of

substations. Most of the Katanga power stations are situated in the vicinity of Kolwezi and export

their output either to SCK or RO. The electricity will be provided by SNEL at 220 kV from SCK or

at 120 kV from the RO which is connected to SCK via a single 100 MVA transformer.

A 10 year load forecast was prepared for the whole of the DRC and for the Katanga Province, based

on information provided by SNEL and the mine operators. Compared to early 2008, the total

demand has increased as new loads have emerged, but SNEL warned that most mine operators are

exaggerating their requirements and have advised KOL to use a ‘certainty factor’ of 85%.

In Kinshasa, some 100 MW of load cannot be supplied at present because of transmission limitations

but will materialize as soon as new lines are commissioned between Inga and Kinshasa. The events

of November 2008 caused many projects in the DRC to be put on hold for an indefinite period or to

reduce production capacity. This has resulted in a much reduced demand for the Kolwezi area. The

TAP Report has been adjusted accordingly.

20.4 Geotechnical Assessment

The geotechnical aspects of the following Material Assets have been evaluated:

T17 Mine;

Tilwezembe Mine (included in the 2008 Study, excluded from the current plan);

Kamoto Mine;

KOV Mine; and

Mashamba East Mine.

Mupine Pit is not intended for further production, but a preliminary geotechnical assessment of the

pit walls (with a view to assessing their capability for supporting tailings-disposal infrastructure) has

been conducted.

Kananga Pit also lies within the Katanga Mining lease area but is not incorporated in the current ore

resource and has not been investigated in this study.

Information presented in this section has been obtained from Gecamines plans, sections and reports,

SRK reports, and three on-site investigations in November 2007 and April and May 2008.



Geological descriptions of key rock types present on the site have been obtained from previous

Gecamines reports and a previous investigation by SRK.

The conclusions outlined in this section are based on the 2008 Study. The objective of this study was

to provide design criteria for use in mine design. For underground mining at Kamoto Mine, this

involved generating acceptable stope and pillar geometries. For surface mining at T17, Tilwezembe,

KOV and Mashamba East Mines slope angles for individual stacks and overall slopes were

recommended.

20.4.1 Kamoto Mine

The stratigraphy of the mining area is comprised essentially of dolomitic strata interbedded with

shales. Two ore bodies are encountered: the upper ore body (OBS) is between 8 m and 13 m in

thickness and is hosted in the lower SDS series of dolomitic shales. The lower ore body (OBI) is

between 10 m and 15 m in thickness and is hosted mainly in the RSF series of siliceous dolomites.

The ore bodies are separated by a siliceous dolomitic parting, the RSC, between 8 m and 15 m in

thickness. The ore body is bowl shaped with a flat central area, the Plateure, and surrounding areas

gradually increasing in dip, locally becoming almost vertical.

With the exception of information obtained for the KOV Mine design in 2006, geotechnical

information has been obtained from Gecamines reports produced between 1982 and 1991. Site visits

were conducted between 2006 and 2008 during which the overall rock mass quality in both surface

and underground environments was assessed and used in conjunction with laboratory test

information presented in the Gecamines reports to generate rock mass strength parameters for use in

design. The design rock mass strength allocated to OBS is 48 MPa and to the OBI is 56 MPa. These

values are considered to be acceptable for this level of study. It is recommended that further

sampling and testing is conducted to provide specific information for detailed design of mining

areas, particularly Etang where effects of weathering may become evident at higher elevations and

rock mass strength may deteriorate.

Mining method selection has been strongly influenced by a combination of ore body dip and ore

body width. In general, multi cut methods incorporating backfill as a working platform have been

recommended. At dips of less than 12º the room and pillar method used by Gecamines is

recommended. This involves development, sliping and benching phases of extraction to create a

slender pillar which is supported by backfill. It is noted that a pillar collapse occurred in September

1990 which resulted in partial closure of the underground mine. Documentation relating to the cause

of the collapse has not been reviewed but it is surmised that a combination of over mining, non-

superimposition of pillars and non-placement of backfill were contributing factors. The design

recommended is considered appropriate for the Kamoto environment and the need for adherence to

design during implementation is emphasized. Detailed design work should take account of un-mined

areas when estimating pillar loading.

The effect of the Plateure collapse on stress distribution in surrounding areas is unknown. The effects

of stress concentration have not been explicitly considered in this study and the major stress

component is assumed to act vertically. It is recommended that in-situ stresses are measured and

their effect on the stability of adjacent areas is assessed during detailed mine design.

The long hole retreat stoping method recommended in a previous study requires placement of

cemented fill in primary panels to allow complete extraction of secondary panels. It is considered



that this method is experimental in the Kamoto Mine environment and should not be incorporated

into this study.

Open stoping methods are subject to restricted hanging wall spans to maintain stability and minimize

dilution. Consequently, the overall extraction achievable is relatively low. Design spans are based on

a combination of those achieved in existing operations and rock mass quality considerations. Narrow

(10 m wide) crown pillars are likely to be unstable and will not provide the required protection to

stoping operations. The operational practicality of this method requires further investigation,

particularly if single drive access is planned. Incorporation of this method into the study is not

recommended.

At dips of between 13º and 55º, longitudinal cut and fill mining methods are preferred. Where stopes

are wide and at the lower range of dip, post pillar cut and fill mining can be employed with slender

post pillars acting as span breakers. Either waste rock or tailings backfill can be utilized as fill.

Incorporation of regional support pillars at centres of approximately 140 m has been recommended

where mining can become laterally and vertically extensive. It is recommended that pillar stresses

are monitored during mining to determine the precise location of regional support.

In specific areas with dips exceeding 55º the possibility of using cave mining exists. Detailed design

of a caving layout for Zone 9 has been prepared by Kamoto personnel. The method has been

reviewed and risks arising from remnant pillars and waste rock fill have been identified but are not

considered insurmountable. Due to the limited amount of ore that will be extracted, it is not expected

that caving effects will be experienced on surface.

Backfill is an essential element of the mining systems to provide both a working platform for multi

lift mining methods and long term confinement to enhance the support capability of pillars. Although

backfill will not exert sufficient resistance to prevent hanging wall deformation, it is recommended

that stopes are filled as tightly as practicable to minimize deformation and consequent load transfer

to pillars.

In the short term, waste rock from development or surface dumps is considered acceptable as a

source of fill material. In view of potential cost savings it is recommended that tailings based fill

systems are investigated to replace some, or most of the waste rock fill.

A schematic modularized backfill system design has been presented that will deliver 100 m³ per

module of either cyclone classified tailings or thickened tailings. Illustrative capital and operating

costs have been provided. It is recommended that a detailed backfill reticulation system is designed

taking cognizance of routing logistics and mining schedules.

The mining methods recommended are “entry” methods and therefore stope hanging walls will

require short term support. Current support systems are based on 2,4 m long mechanical roof bolts,

fully grouted rebars or Swellex installed in a closely spaced pattern. Additional cable anchor support

has been recommended for multiple lift methods. Support systems have not been optimized in this

study and there remains the potential to do so.

20.4.2 Open Pits

Information from a geotechnical investigation programme conducted in 2006-2007 was used to

generate Mohr Coulomb parameters for KOV Mine. No geotechnical information relevant to the

other surface operations was available and slopes were based on geometries previously achieved in



the open pits. It is noted that there is a considerable amount of variability in the KOV Mine data due

to the nature of the rock mass and it is recommended that further sampling and testing is conducted

to increase the level of confidence in the values used.

Slope angles on Mashamba East and T17 Mines have been based on geometric information derived

from previous open pit mining in the DIMA pits. It is recognised that these are not optimum slopes

and there is the potential for increasing slope angles if further geotechnical investigations are carried

out. Analyses of designed final slopes using rock mass parameters derived for KOV as surrogate

parameters have generated safety factors in the range between 1,6 and 2,1 assuming that slope

drainage is poor. These values are considered to be high and there is the potential for increasing

slope angles if further geotechnical investigations are carried out.

Should Mashamba East be mined to its planned limit, there is the possibility that the high wall will

intersect the Kamoto Interim Tailings Dam. Notwithstanding the risk of mudrush if tailings are not

fully drained, the safety factor is reduced significantly (assuming that the slope is subject to a water

surcharge). It is recommended that further detailed design work is carried out to fully evaluate this

scenario.

Tilwezembe is located in a different geological environment to the other pits with the footwall and

ore body rock mass comprising highly fractured argillite and the hanging wall comprising tillite. In

the absence of any other information except a previously mined slope approximately 50 m high, an

overall slope angle of 45º was used for design. Although safety factors were estimated to lie in the

range of 1,7 to 1,9, the confidence in these results is low. It is strongly recommended that a

comprehensive geotechnical investigation is carried out to confirm acceptability of designed slope

angles.

A comprehensive geotechnical study was conducted in 2006 and 2007 to generate design parameters

for KOV Mine that were applied to the Cut 3 and Cut 7 mining layouts. Overall slope angles of 25º

and 22º were recommended for the northern slopes of Cut 3 and Cut 7 respectively. These angles are

considered to be low but reflect the uncertainty associated with adversely dipping shear structures

that may exist in those slopes. Angles of 40º and 30º are recommended for all other slopes.

A preliminary survey of slopes surrounding the flooded Mupine pit was carried out to establish the

possibility of siting tailings disposal infrastructure around its perimeter. Southern and western slopes

were observed to have undergone extensive collapse presumably associated with major faulting. The

northern slopes located in RAT appear to be stable although localized collapse has occurred

generally in association with artisanal workings. Access to the current water level is possible along a

stable ramp located on the northern wall.

20.4.3 Recommendation

Further studies should be carried out on the rock chracteristics associated with the Material Assets.

20.5 Tailings and Process Effluent Disposal

This section of the report documents the feasibility design and costing of new tailings disposal dams

and related infrastructure.



20.5.1 Scope of Study

A tailings dam site selection study was carried out for which nine sites were considered. The sites

were subsequently ranked and the preferred sites considered for further study. Due to the anticipated

high tailings volume discharge over the life of the mine, it was necessary to consider more than one

dam to accommodate the expected quantity.

The study for the individual dams was limited to the assessment and design of the dams, return water

dams, drainage and access infrastructure inside the perimeter fence. The slurry delivery to the dams

and return water pipelines from the return water dams has been excluded, as has the mechanical and

electrical design, which have been covered by others in this report..

In addition to the tailings dams, consideration has also been given to the provision of lined ponds to

accommodate any hazardous process effluents discharging from the Luilu Plant.

20.5.2 Design Assumptions

The tailings dams were designed to accommodate the life of mine production plan, dated 2nd

February 2009 running from 2009 to 2032, which indicates a total tailings capacity of 155 Mt is

required, increasing from around 215 ktpm in 2009 to 738 ktpm in 2022, then decreasing towards the

end of the mine life. It has been assumed that the tailings will be produced by two plants, namely the

Luilu and Kamoto Concentrator Plants. As the Kamoto Concentrator Plant is already operating,

limited test work has been carried out on the physical tailings characteristics currently emanating

from the plant, however, no information is currently available for the Luilu Plant. Based on available

information, it is anticipated that the tailings will be too fine to be used for the building of the outer

walls of the dam during operation.

The preliminary test work (TCLP and acid base accounting) indicates that the tailings currently

being generated at Kamoto Concentrator exceed the Low Risk limits for lead and copper in terms of

the DRC regulations. However, the regulations do not set out limits for High Risk tailings for

copper and lead. The geochemical test work carried out by SRK on the copper tailings from the

Luilu Metallurgical Plant currently classifies the material as High Risk due to its acid generation

potential. This tailings stream also contains significant concentrations of Cu (and Co), which

currently exceed the DRC effluent discharge limit. However, it is anticipated that an appropriate

alkaline pH will be maintained in the tailings stream and that process improvements will reduce the

concentrations of Cu and Co in the Luilu Plant tailings, therefore SRK has recommended that it will

not be necessary to synthetically line the tailings dams and that excess decant water from the dams

may be allowed to spill into the environment at times of high rainfall. The maintenance of alkaline

pH will require effective monitoring and management by KOL.

20.5.3 Selection of Preferred Tailings Dams

The nine sites considered are shown on Figure 20.1 and comprise the following:

a) Kamoto Interim Tailings Dam;

b) Far West Tailings Dam;

c) Mupiné Super Pit;

d) South West Tailings Dam;

e) Mashamba East Pit;



f) Potopoto Super Dam;

g) Luilu River Super Dam;

h) Potopoto Interim Dam;

i) Kamoto Tailings Dam.

The sites were each assessed in terms of various environmental, public acceptance, engineering and

economic criteria. The criteria assessments were combined to produce an aggregate ranking for each

site. Based on this selection process three sites were chosen as the preferred sites, namely; the

Kamoto Interim, Mupine Super Pit and Far West tailings dam sites.

20.5.4 Description of Proposed Tailings Dams

Kamoto Interim Tailings Dam

The location is a brownfield site situated approximately 300 m south of the existing Kamoto

Concentrator Plant and 200 m north of the northern rim of the Mashamba East Pit.

The site is currently partly occupied by a newly built tailings dam which was commissioned in

November 2008. The planned site covers approximately 88 ha. Prior to construction of the existing

dam the site was largely covered by previously deposited tailings. The site is bounded by high

ground to the north. The area is underlain by natural colluvial and residual clayey and silty sands,

passing into Roan dolomitic series rocks in the southern part and the Roan Dipeta sandstones in the

northern part. Available borehole information in the area indicates that the groundwater is present at

approximately 14 to 19 m below ground level. Prior to commissioning of the dam in November

2008, analyses of groundwater samples indicated concentrations of iron and selenium above the

World Health Organisation drinking water guidelines. This is common in the Kolwezi area and is

likely to be due to the mineralogy as these elevated levels are seen across the region. The existing

impoundment wall is approximately 2,3 km long and varies from 1,5 to. 3,5 m in height, constructed

from compacted soil. It will be raised and extended, using nearby overburden and development

waste, to attain the desired capacity, with the final wall height increasing to a total height of between

1,5 and 11 m and with a length of 2,4 km. No wall has been provided on the northern side of the

dam, where the natural ground rises.

Additional features of the final dam will include:

Toe catchment paddocks, to contain tailings spills and eroded slope runoff from the wall;

Underdrainage, to drain water from inside the wall of the dam;

A penstock to drain process and stormwater from the surface of the dam;

A solution trench to drain water discharging from the underdrains and penstock, and direct it

to the return water dam;

A return water dam to contain water discharge from the dam prior to pumping to the plant;



Figure 20.1 Tailings Disposal: Probable Tailings Dams Sites



A diversion bund will be constructed down stream of the dam, to deflect a tailings flow slide

away from the settlement downstream of the dam, in the unlikely event of a dam wall

failure; and

A tailings deposition pipeline to distribute tailings around the inner wall of the dam, with

deposition carried out via the spigotting method.

Tests carried out on the soils below the site indicate that they have a low permeability. The

hydrogeological information available for the area, together with the information from geochemical

tests carried out on tailings from the Kamoto Concentrator Plant indicate that contamination of the

groundwater due to seepage of leachate from the tailings is likely to be of low significance. This

assessment includes the understanding that an appropriate alkaline pH will be maintained in the

tailings stream and that process improvements will reduce the concentrations of Cu and Co in the

Luilu Plant tailings.

It should be noted that at times of high rainfall it is expected that the return water dam will spill into

the environment. This is deemed to be acceptable based on the pH balance and plant improvement

anticipated.

The envisaged maximum capacity is 5,8 Million m3 or 7 Mt at an assumed in-situ dry density of

1,2 t/m3.

Mupine Super Pit

The location is a brownfield site situated approximately 1km southeast of the existing Luilu Plant

and 3 km northeast of the Kamoto Concentrator Plant.

Part of the site is currently occupied by the Mupine Pit, a previously mined pit approximately 1km

long on a east-northeast/west-southwest axis (1,5km including an access ramp on the eastern end)

and approximately 500 m wide. The pit is approximately 135 m deep at its deepest point, but it is

believed that the pit is partially infilled, reducing its current depth to approximately 80 m at the

location of the highest elevation on the pit walls to 65 m below the lowest wall elevation. The pit is

partially infilled with water, which is at a depth of 60 m at the location of the highest elevation on

the pit walls to 45 m below the lowest wall elevation.

A number of drainage channels are incised into the natural ground and waste dumps in the area,

including a number which drain into the pit.

No residential development is present in the vicinity of the site.

The geological sections indicate a steeply dipping strata with a number of thrust faults, including the

Roan RAT, CMN and RGS formations. The whole known orebody has been mined out and further

mining is not envisaged.

The site encloses an approximate maximum surface area of 176 hectares. The envisaged maximum

capacity is 47 Million m3 or 52 Mt at an assumed in-situ dry density of 1,1 t/m3.

The dam will comprise two main phases. Initially the tailings will be deposited into the existing

Mupine pit via a ring main tailings distribution pipe laid on the surface around the perimeter of the

dam, with deposition carried out using the open-ending method. When the pit approaches capacity a

wall will be constructed around the perimeter of the dam, to create what has been named the Mupine



Super Pit. The wall will only surround the west, north and east sides of the dam as a waste rock

dump forms an existing barrier to the south. The proposed wall will be approximately 4 km long and

will be of variable height due to the irregularity of the existing ground, however, the maximum

height will be approximately 30 m.

Water will be decanted from the super dam and sent to a return water transfer station and ponds

located to the east of the pit. Some water will be returned to the plant, but during periods of high

rainfall, some water will be discharged into the environment.

An allowance has been made in the design and costing to relocate services, including a pipeline,

railway trucks and roads, located to the west of the pit.

Rock formations comprising some dolomitic material are anticipated between the Mupine Pit and the

underground workings. Dolomitic formations have the potential for dissolution and the formation of

high permeability groundwater flow paths when exposed to acidic groundwater. The water currently

in the pit is slightly alkaline, as is the tailings slurry that has been tested from the Kamoto

Concentrator. Based on information made available, the tailings slurry that will be deposited in the

pit will have the same, or higher, pH than the existing water in the pit.

Far West Tailings Dam

The location is a greenfield site approximately 4 km southwest of the existing Luilu Plant and 5 km

northwest of the existing Kamoto Concentrator Plant. The site is bordered by the Luilu River to the

east. The site has a gently sloping profile towards the Poto Poto and Luilu Rivers.

The Far West site is located off the Roan Basin, on the Kundelungu tillite, to the west of the Luilu

River. The underlying geology comprises mainly shale and dolomite formations. The rocks are

overlain by an average of more than 5 m of fine residual soils.

The groundwater levels range between 1 m below surface close to the Luilu River, down to

approximately 5 m below surface on the western side of the dam. Analyses of groundwater samples

has indicated that the area has relatively low concentrations of total dissolved solids, with some

elevated lead, selenium and iron, as expected from the natural mineralogy of the area. Groundwater

flow, based on the deep boreholes, is predominantly from west to east, towards the Luilu River.

A number of small villages are located in the vicinity of the site, which it will be necessary to

relocate prior to construction.

The site encloses an approximate maximum surface area of 420 ha. The envisaged maximum

capacity is 83 Million m3 or 99 Mt at an assumed in-situ dry density of 1,2 t/m3. However, based on

the existing mine plan only 79 Million m3 or 95 Mt will be required.

The proposed impoundment wall is approximately 9,8 km long and will reach a height of

approximately 50 m at its highest point. No wall has been provided on the western side of the dam,

where the natural ground rises. The wall will mainly be constructed using poorly sorted waste taken

from the overburden and development waste dumps situated around 3 to 7 km from the site, across

the Luilu River to the east.

It is proposed that an initial starter impoundment wall (Phase 1) be constructed prior to the

commencement of deposition, and thereafter the wall height will be gradually raised in advance of

the deposited tailings throughout the operational life of the dam.



Additional features of the final dam will include:

Toe catchment paddocks, to contain tailings spills and eroded slope runoff from the wall;

Underdrainage, to drain water from inside the wall of the dam;

A penstock to drain process and stormwater from the surface of the dam;

A solution trench to drain water discharging from the underdrains and penstock, and direct it

to the return water dam;

A return water dam to contain water discharging from the dam, prior to pumping back to the

plant;

A clean water dam, to store clean water up stream of the dam for the use of the local

inhabitants; and

A tailings deposition pipeline to distribute tailings around the inner wall of the dam, with

deposition carried out via the spigotting method.

Tests carried out on the soils below the site indicate that they have a low permeability. The

hydrogeological information available for the area, together with the information from geochemical

tests carried out on tailings from the Kamoto Concentrator Plant indicate that contamination of the

groundwater, and the river via the groundwater, due to seepage of leachate from the tailings is likely

to be of low significance.

It should be noted that at times of high rainfall it is expected that the return water dam will spill into

the environment. This is deemed to be acceptable based on the pH balance and plant improvement

anticipated.

Part of the dam site falls outside of KOL’s current concession area.

20.5.5 Phasing of Tailings DamsThe Kamoto Interim dam has already been constructed. Consequently it is proposed to use this dam,

and to raise and extend the wall to accommodate the initial tailings from the life of mine plan.

Subsequently the Mupine pit will be commissioned, when the Kamoto Interim Dam is nearing

capacity, followed by the Far West Dam. The key dates have been summarised below in Table 20.1.



Table 20.1 Tailings Dams: Key Dates

Date (ending)Existing Kamoto Interim

Tailings DamKamoto Interim Dam Extension Mupine Pit (Pit only) Mupine Super Pit (with new Wall

around pit)Far West Dam

2009 - Mar Start Construction

2009 - May End Deposition Start Deposition

2010 - May Start Construction

2011 - Aug End Deposition Start Deposition

2012 - Apr Start Construction (Phase 1 only)

2014 - Mar Start Deposition (including wallextensions)

2014 - May Start Construction

2016 - Apr End Deposition Start Deposition

2022 - Mar End Deposition

2032 - Dec End Deposition (including wallextensions)



20.5.6 Risk Assessments

A risk assessment was carried out to determine the risk to life and property in the event of a

catastrophic failure of the dams. The results indicate that the risks associated with failure of the dams

are lower than published acceptable norms for the Far West and Kamoto Interim tailings dams.

Before the risk assessment can be completed for the Mupine Super Pit, additional investigation is

required.

20.5.7 Hazardous Effluent Ponds

It is anticipated that hazardous process effluent, unsuitable for deposition in the tailings dams, will

be generated by the Luilu Plant. To accommodate this effluent an allowance has been made in the

design and costings to construct twenty two lined ponds to the west of the Luilu Plant. The location

of the ponds is shown on Figure 20.1. It should be noted that one pond has previously been

constructed at the site and is currently unused.

20.5.8 Closure Considerations

The overall closure approach in relation to the tailings dams and ponds will be to ensure physical and

chemical stability of the dams as far as is practicable at the time of closure and to minimise the post

closure maintenance required.

20.5.9 Further Study

Tailings Characteristics

Based on available information, it is understood that the tailings will be too fine to allow them to be

used for wall building. Consequently, the design allowed for herein consists of the tailings behind

walls constructed from mine overburden and development waste to the full closure height. This

decision has significant implications for the cost of the dam as the material required to build the

walls will need to be excavated and transported to the site.

It is recommended that additional studies be carried out to fully characterise the physical

characteristics of the tailings, particularly the Luilu Plant tailings when samples become available, to

determine if a portion can be used for wall building.

Further chemical testing should be carried out on the tailings slurry to establish the potential

contamination risk to the environment. SRK will than be able to form a better understanding of

whether or not some effluent streams will need to be handled separately.

Overburden and Development Waste Dumps

It has been assumed that the majority of the material used to construct the impoundment walls will

be taken from the overburden and development waste dumps in the area.

It is recommended that a geotechnical fieldwork and laboratory testing programme be carried out to

more fully characterise the material within the dumps. This will assist in the optimisation of the wall

construction and allow haulage distances for suitable material to be minimised.



Tailings Deposition

Should the tailings prove to have a particularly fine grading it may be possible to dispense with

spigotting for the Far West and Kamoto Interim dams. As an alternative, open ended pipe deposition

could be considered, using pipe off-takes along the perimeter tailings distribution pipeline. The open

ending approach would result in savings in the cost of spigot piping.

Mupine Pit and Super Pit

It is recommended that the borehole MUP05 proposed by AGES be drilled and pump tested and

additional isotope samples be taken from inflows into the underground workings and from the new

boreholes MUP1, MUP2, MUP3 so that a more detailed understanding of the risk of tailings disposal

in the Mupine pit can be gained.

20.5.10 Conclusions

The main conclusions are as follows:

A site selection study was carried out and the three top ranked sites chosen to accommodate

the anticipated tailings quantities, namely the Kamoto Interim Dam, Far West Dam and

Mupine Super Pit;

The three dams can accommodate the 155 Mt of tailings anticipated over the LoM;

Currently available information indicates that the tailings will not be sufficiently coarse to be

used for wall building, consequently it has been planned for walls to be constructed from

waste dump material;

It is understood that an appropriate alkaline pH will be maintained in the tailings stream and

that process improvements will reduce the concentrations of Cu and Co in the Luilu tailings.

Based on this, no tailings dam liners have been allowed for in the design, the risk of

preferential pathways due to dissolution of dolomite between the Mupine Pit and Kamoto

Mine underground workings is deemed acceptable, and excess decant and overflow water

from the dams is deemed acceptable for release to the environment;

The total capital cost for the tailings dams is estimated to be USD196,7 million and the

operating cost USD324,5 million, with a combined total of USD521,2 million; and

An allowance has been made for 22 additional hazardous effluent ponds. The total capital

cost for the ponds is estimated to be USD137,4 million.

Further studies should be carried out to; fully characterise the physical and chemical properties of the

tailings streams, investigate the possibility of open end deposition, further investigate the geology

and hydrogeology at Mupine Pit and to characterise potential construction materials.



20.6 Risk Assessment

To assess the Project risk, the 2008 Engineering Study was subject to a high level risk assessment

facilitated by CorProfit Systems Africa (“CorProfit”), which complies with the AS/NZS 4360

Standard used to identify risks and their controls measures.

20.6.1 Basis of the Risk Report

This report presents the outcome of risk assessments undertaken at two workshops on 28 August and

5 September 2008 involving members from all the disciplines in the Engineering Study.

20.6.2 Risk Register

20.6.3 Resources: Overstated Resource Estimate for T17 Mine

The risk: The grade estimated for the current workings of the T17 Mine may

be overstated as current grade differs significantly from estimates

based on the block model.

Residual Risk Rating: High

The Consequence: Reduced production of copper and cobalt.

Risk Mitigation Measure/s: No risk mitigation measures identified.

20.6.4 Mining and Reserves: KOV Equipment

The risk: Late delivery of equipment. The late delivery of equipment, as

experienced by the mining industry world-wide due to the shortages

and long lead times for equipment will negatively impact

production.



Risk Mitigation Measure/s: Expedite delivery schedules.

20.6.5 Mining and Reserves: KOV Contract

The risk: Higher cash mining costs. At the time of publication of the

Engineering Study, the KOV mining contract had not been signed.

As a result, SRK has estimated the cash mining costs, which may

deviate from the costs that will eventually be realised as a result of

the mining contract.



Risk Mitigation Measure/s: Finalize contract.



20.6.6 Metallurgical Processing: Inability to Meet Schedule for Modules 1, 2 and 3

The risk: The inability to meet the development schedules for the production

of Modules 1,2 and 3. Delays may occur as a result of many factors

including the delivery of equipment, the availability of manpower,

and other infrastructure and in-country issues which may be beyond

the control of management.



Risk Mitigation Measure/s: Realistic scheduling and budget.

20.6.7 Metallurgical Processing: Unavailability and Quality of Key Reagents

The risk: The unavailability and quality of key reagents (like sulphur and

lime) that are required in large quantities and are critical to the

metallurgical process.



Risk Mitigation Measure/s: Detailed supply management plan.

20.6.8 Services: Poor Condition of Railway Line

The risk: The poor condition of the railway line will impede efficient

production by not allowing the efficient, on-time delivery of

finished products or the supply of key input materials on time.

Residual Risk Rating: Very High

The Consequence: (i) Reduced production of copper and cobalt; and

(ii) Higher logistics costs.

Risk Mitigation Measure/s: (i) Reschedule plans to match rail capacity;

(ii) Engage with governments and railway operators; and

(iii) Engage with other potential rail users.



20.6.9 Services: Availability of Rolling Stock

The risk: Rolling stock (locomotives and wagons) will not be available on

time to transport the scheduled increases in production of finished

products and key input materials.

Residual Risk Rating: Very High

The Consequence: (i) Reduced production of copper and cobalt; and

(ii) Higher logistics costs.

Risk Mitigation Measure/s: (i) Establish capacity; and

(ii) Negotiate with SNCC (the rail operator) and other railway

groups.

20.6.10 Services: Under-developed in-country institutional infrastructure andcapacity

The risk: The DRC national and local governments and their agencies will not

have the ability to deliver on the infrastructure requirements of the

Project.


The Consequence: (i) Project viability (NPV and IRR); and

(ii) Project delay.

Risk Mitigation Measure/s: (i) Develop relationships with other stakeholders, governments and

agencies; and

(ii) Support capacity development initiatives.

20.6.11 Services: Lack of Power Supply

The risk: Power generation will not meet the requirements of the Project as

power requirements beyond 2011 require large-scale investment by

the DRC government and SNEL.


The Consequence: Project viability (NPV and IRR).

Risk Mitigation Measure/s: Invest with SNEL in additional power generation.



20.6.12 Environmental: Non-resolution of Liabilities

The risk: KOL may be liable for previous liabilities incurred before

operations commenced under current management.


The Consequence: KOL will inherit the liabilities.

Risk Mitigation Measure/s: Quantify current liabilities and negotiate.

20.6.13 Environmental: Non-compliance with DRC Mining Code

The risk: The risk that the mining and/or environmental licences will be

revoked as a result of non-compliance with the DRC mining code.


The Consequence: Loss of licence.

Risk Mitigation Measure/s: Expedite and adhere to environmental-management plan.

20.6.14 Human Resources: Senior Management and Technical Expertise

The risk: The inability to recruit and retain senior management and operation-

critical technical expertise to manage and operate the mines and

processing plants.


The Consequence: (i) Project viability (NPV and IRR); and

(ii) Ability to comply with legislation necessary to ensure the

optimal operation of the business.

Risk Mitigation Measure/s: (i) Review the company’s strategy;

(ii) Review the recruitment and retention plan; and

(iii) Facilitate the provision of services with Government and other

service providers.

20.6.15 Capital Costs: The Unpredictable Escalation of Costs

The risk: Recent projects in the mining industry world-wide have experienced

unpredictable capital cost overruns due to various macroeconomic

and microeconomic factors that cannot be predicted with any degree

of confidence.


The Consequence: Inaccurate capital cost estimate.

Risk Mitigation Measure/s: Regular review of estimates.



20.6.16 Operating Costs: Deviation from Engineering Study Estimates

The risk: Recent projects in the mining industry world-wide have experienced

unpredictable operating cost overruns due to various

macroeconomic and microeconomic factors which cannot be

predicted with any degree of confidence.


The Consequence: Project viability (NPV and IRR).

Risk Mitigation Measure/s: Include adequate contingency and run sensitivity models.

20.6.17 Risk Controls

As indicated previously, this report presents the outcome of risk assessments undertaken at two

workshops on 28 August and 5 September 2008. KML had indicated that the following risk control

measures will be implemented, and since these mitigation measures were based on the 2008 Study,

some of these may no longer apply:

KML has recently hired some experienced environmental professionals with technical

expertise in mine operations. In addition, KML employs some 100 expatriates and has made,

and will continue to make, significant progress on infrastructure development and

expansion, which will greatly assist with the recruitment and retention of skilled staff;

Medical facilities have been expanded and upgraded with further activities in the process of

being implemented. A formal malaria-control initiative has been launched and so has a

community inoculation program. KML recently hired a Director of Medical Services, and he

is now coordinating the healthcare program for employees and their dependents. A

community health coordinator has also been recruited as has a small team led to implement

the malaria-control initiative. Crusader Health have been contracted to provide health care

insurance cover for both occupational health and the primary healthcare of employees and

their dependents. This will result in the construction and running of three community clinics

and a new containerized hospital in town. The Director of Medical Services is engaging

local stakeholders on this aspect;

Multiple potable water wells and lines have been installed or rehabilitated and rebuilding of

the community road network is under way. Donations to local schools will continue, and the

construction of an expatriate’s school is planned for 2009.

A dedicated senior manager of KML’s health and safety program is now in place at the

corporate level. An ongoing audit has elevated health-and-safety awareness and made

significant progress in improving working conditions at site. An ISO-based management

system is under development and is being progressively introduced to site operations; and a

major incident/crisis response plan has been developed under the guidance of recognized

experts;

KML has developed systems and procedures to liaise effectively with the community and

stakeholders. In the past year KML has consulted with local stakeholders, invested in social

programs and rehabilitated local infrastructure. The Governor of Katanga region has been



supportive of KML’s activities. The opening of community liaison offices is planned over

the next 6 months and will further strengthen links to local communities.

Management information systems have been upgraded. To this end, microwave capacity has

been upgraded and is now reliable; the reinstallation of an expanded fibre-optic network is

about to begin and the construction of a secure and stable IT facility (Phoenix) is nearing

completion; Mincom (software) implementation is progressing and being rapidly debugged;

power-source issues have been stabilized; professional IT staff has been hired and is being

progressively expanded; working relationships with contractors knowledgeable of and

experienced in the Congo are rapidly improving; and DR (management information system)

planning and implementation is to begin in the last quarter of 2008.

21 Interpretations and ConclusionsThe results and interpretations of exploration on the Material Assets are reported elsewhere in this

report and have been relied upon to compile the Mineral Resource statement included in Item 19.

22 Recommendations

Recommendations by SRK for future work required on the technical assets are included in other

items in this report. Specific action programs recommended are:

Dewatering Item 20.2;

Geotechnical Item 20.4; and

Tailings Item 20.5.

In undertaking the study, Bateman Engineering has been provided with and has relied upon records,

documents and other information supplied by the client and other third parties. Save as expressly

stated in this report, Bateman Engineering has assumed and did not attempt to verify the accuracy,

reliability, sufficiency or validity of such information, data or records and documents. Bateman

therefore recommends:

Repeat testwork for all mineral properties –Item 18;

Additional work the on the sizing of the WOL/SX/EW plant and configuration, as well as

capital and operating costs – Item 25b; and

Additional work on the capital and operating cost estimates.



23 ReferencesTable 23.1 provides details of companies that provided specific information which SRK utilised in

the compilation of this technical report.

Table 23.1: References

Organisation Report/sNI 43-101

Section

A&B Global MiningMining Report on T17 Mine, Kamoto Mine and Mashamba East Mine, A&B Global Limited –2009 Update.

19, 25a

Africon-MMC

Africon, September 2008. Draft Report: Katanga Mining operations Kolwezi Trafficmanagement and Road Safety plan.MMC Engineers, January 2008. Specialist Traffic and Transportation Study for proposedCopper and Cobalt Mine Operations at KOV, Kananaga, Tilwezembe and Kanfukuma inKolwezi, Katanga Province DRC Final Draft Report.

25e

AGES South AfricaAGES, 2008. Technical Reports AS-R-2008-09-05 and AGES-R-08-01-28, AG-R-2008-11-24KOL Geohydrology Report Version 3 DRAFT and AG-R-2008-10-02 KOV Dewatering DFSV2 Final.

20.2

Bateman Engineering Bateman Engineering Study November 2008 and Engineering Study Addendum February 2009. 18, 25b

CRU Strategies Update of copper and cobalt price forecasts - January 23rd, 2009. 25c

FM AcousticConsulting

F le R Malherbe, March 2008. Noise Impact Study for the Kamoto Project near Kolwezi in theDRC, Report no 07/1/2/ Rev 2.F le R Malherbe, January 2007. Noise Impact Study for the Nikanor DCP project near Kolweziin the DRC, Revisions 1 Report No 07/9/4.

25e

Foxfire ScientificFoxfire Scientific, January 2009, Katanga Mining Limited DRC Copper Mining Projects Phase1 Radiation Survey and Sampling of Kolwezi and Tilwezembe Mining Concessions.

25e

Golder AssociatesAfrica

Golder Associates, March 2007. Kamoto joint Venture project Aquatic Ecology Report. ReportNo 10440-6031-2Golder Associates, May 2008. Baseline Assessment of Aquatic Ecosystems Associated withNikanor Kolwezi Report no 10473/08/2Golder Associates, March 2008, Baseline Assessment of Aquatic Ecosystems associated withNikanor, Kolwezi Report No 10473-6028-1.

25e

Norton RoseLegal opinion as indicated in Section 6.2 (6.2.1 to 6.2.7). Received by e-mail on 13 March2009.

6.2

Snowden GroupKatanga Mining Limited: Tilwezembe and Kananga Project – Mineral Reserve Estimate, July2008.

19

Trans-Africa Projects(“TAP”)

Supply to Kamoto Installations in Kolwezi: Load Flows and Cost Estimates – September 2008,including 2009 Addendum.

20.3.2

University ofGembloux

Nature Plus, April 2004, Ecological report on sites around the town of Kolwezi, GemblouxBelgium.Nature Plus, April 2006, Ecological Report concerning Tilwezembe copper hill, Gembloux,Belgium.Nature Plus, February 2007. Ecological Report concerning Tailings dam at yenge and theProposed Process Plant Site.

25e

24 Date and Signature Pages



24.1 Roger Dixon

To accompany the report dated 17 March 2009 and entitled “An Independent Technical Report on

the Material Assets of Katanga Mining Limited, Katanga Province, DRC.”

I, Roger Dixon, hereby certify that:

i. I am a Partner and Director with the firm SRK Consulting (South Africa) (Pty) Limited

(SRK) with an office at SRK House, 265 Oxford Road, Illovo, Johannesburg 2196, South

Africa;

ii. I am a graduate of the Royal School of Mines, Imperial College London with a BSc (Hons) in

Mining Engineering in 1971. I have practiced my profession continuously since 1972;

iii. I am registered as a Professional Engineer with the Engineering Council of South Africa since

2000. I am a Honorary Fellow of the Southern African Institute of Mining and Metallurgy;

iv. I have not received, nor do I expect to receive, any interest, directly or indirectly, in KML;

v. As of the date of this certificate, to the best of my knowledge, information and belief, this

technical report contains all scientific and technical information that is required to be

disclosed to make the technical report not misleading;

vi. I have read National Instrument 43-101 and Form 43-101F1 and by reason of my education

and past relevant work experience, I fulfil the requirements to be a “Qualified Person” for the

purposes of National Instrument 43-101. This technical report has been prepared in

compliance with National Instrument 43-101 and Form 43-101F1;

vii. I am independent of the issuer as defined in Section 1.4 of National Instrument 43-101;

viii. I am responsible for the preparation of the overall report;

ix. I have visited the properties of the Material Assets and my last visit was in from 4-8 August

2008;

x. SRK was retained by KML to prepare an Independent Technical Report on the Material

Assets in accordance with National Instrument 43-101. The preceding report is based on my

review of project files and information provided by KML and other relevant professional

consultants.

___________________________________

Johannesburg, South Africa Roger Dixon, Pr.Eng, BSc (Hons) Mining, FSAIMM

17 March 2009 Partner and Director

SRK Consulting



24.2 Victor Simposya



I, Victor Simposya, hereby certify that:

i. I am a Principal Geologist and Partner with the firm SRK Consulting (South Africa) (Pty)

Limited with an office at SRK House, 265 Oxford Road, Illovo, Johannesburg 2196, South

Africa;

ii. I am a graduate of the University of Zambia with a BSc Sci (Geology) in 1979. I obtained an

MSc (Mining) at Montana Tech, USA in 1990. I have practiced my profession continuously

since 1980;

iii. I am registered as a Professional Natural Scientist with the Engineering Council of South

Africa since 2003. I am a member of the Southern African Institute of Mining and

Metallurgy










viii. I am responsible for the preparation of the geology section and mineral resource estimates in

the technical report for all the assets included in Items 9, 10, 11, 12, 13, 14, 15, 16, and 19;

ix. I have visited the properties of the Material Assets and my last visit was in from 4-8

December 2006;

x. I have prior experience at the properties and acted as the Qualified Person with overall

responsibility for the reporting of Mineral Resources for the “Independent Competent

Persons’ Report on the Material Properties of Global Enterprise Corporate Limited”

published on 26 June 2006.

___________________________________

Johannesburg, South Africa Victor Simposya, Pr.Sci.Nat, MSc (Mining), BSc (Geology),

MSAIMM

17 March 2009 Principal Geologist and Partner

SRK Consulting



24.3 Ebrahim Takolia



I, Ebrahim Takolia, hereby certify that:

i. I am a Principal Consultant with the firm SRK Consulting (South Africa) (Pty) Limited with

an office at SRK House, 265 Oxford Road, Illovo, Johannesburg 2196, South Africa;

ii. I am a graduate of the University of the Witwatersrand with a BEconSc in 1995. I obtained a

MBA from Henley Management College (UK) in 2005. I have practised my profession

continuously since 1994;

iii. I am a member of the Southern African Institute of Mining and Metallurgy and the UK

Securities and Investment Institute;










viii. I have not visited the properties of the Material Assets;

ix. I am responsible for the preparation of the assessment of the mineral economics in Items 25f,

25g, 25h, 25i and 25j of the technical report.

___________________________________

Johannesburg, South Africa Ebrahim Takolia, MBA, BEconSc, MSAIMM, MSI (UK)

17 March 2009 Principal Consultant

SRK Consulting



24.4 Herbert Gerald Waldeck



I, HG Waldeck, hereby certify that:

i. I am a Partner with the firm SRK Consulting (South Africa) (Pty) Limited (“SRK’) with an

office at SRK House, 265 Oxford Road, Illovo, Johannesburg 2196, South Africa;

ii. I am a graduate of the University of Pretoria with a BSc in Mining Engineering obtained in

1971. I obtained a MBA from the University of Potchefstroom in 1975. I have practised my

profession continuously since 1972.

iii. I was awarded a Mine Manager's Certificate of Competency (Metalliferous) in 1974 and have

been registered as a Professional Engineer with the Engineering Council of South Africa

(Registration No 910077) since 1991. I am a Fellow in good standing of the Southern

African Institute of Mining and Metallurgy and an Associate Member of the Association of

Mine Managers of South Africa.










viii. I am responsible for the preparation of the mining schedules and reserve statements in respect

of KOV Mine included in items 19 and 25a3 of the technical report;

ix. I have visited the properties of the Material Assets and my last visit was from 1-5 May 2006;

x. I have prior experience at the properties and acted as the Qualified Person with overall

responsibility for the CPR and the reporting of Mineral Reserve for the “Independent

Competent Persons’ Report on the Material Properties of Global Enterprise Corporate

Limited” published on 26 June 2006.

___________________________________

Johannesburg, South Africa Herbert Gerald Waldeck, Pr.Eng, MBA, BSc Eng, FSAIMM,

AMAMMSA

17 March 2009 Partner and Principal Mining Engineer

SRK Consulting



24.5 Henrietta Salter



I, Henrietta Salter, hereby certify that:

i. I am a Principal Scientist with the firm SRK Consulting (South Africa) (Pty) Limited

(“SRK’) with an office at SRK House, 265 Oxford Road, Illovo, Johannesburg 2196, South

Africa;

ii. I graduated from the University of Surrey with a BSc (Hons) in Chemistry in 1992. I

obtained an MSc in Water Management from the University of Surrey. In 1999, I was

awarded EngD in Environmental Technology from the University of Surrey. I have practiced

my profession continuously since 1993;

iii. I am registered as a Professional Natural Scientist with the South African Council for Natural

Scientific Professions (Registration No 400131/05);










viii. I am responsible for the preparation of the Dewatering and Environment assessment as

described in items 20.2 and 25 (e) of the report;

ix. I have not visited the properties of the Material Assets;

x. I have visited the properties of the Material Assets and my last visit was in 2-6 July 2007.

___________________________________

Johannesburg, South Africa Henrietta Salter, Pr. Sci. Nat, MSc, EngD

17 March 2009 Principal Scientist

SRK Consulting



24.6 Anton von Wielligh



I, Anton von Wielligh, hereby certify that:

i. I am a mining engineer with the firm A&B Global Mining Consultants with an office at Eco

Fusion 6, Block B, Witch Hazel Ave, Highveld Techno Park, Centurion, which are sub-

contracted to SRK Consulting (South Africa) (Pty) Limited (“SRK”) with an office at SRK

House, 265 Oxford Road, Illovo, Johannesburg, 2196, South Africa;

ii. I am a graduate of the University of Pretoria with a BEng in Mining in 2000. I obtained BEng

(Hons) from the University of Pretoria in 2003. I have practiced my profession continuously

since 2000 (ECSA registration no. 20080084);

iii. I have not received, nor do I respect to receive, any interest, directly or indirectly, in KML;

iv. As of the date of this certificate, to the best of my knowledge, information and belief, this



v. I have read National Instrument 43-101 and Form 43-101F1 and by reason of my education




vi. I, as a Qualified Person, am independent of the issuer as defined in Section 1.4 of National

Instrument 43-101;

vii. I am responsible for the preparation of Reserve Estimates of T17 Mine, Mashamba East Mine

and Kamoto Mine described in Items 19, 25a1, 25a2 and 25a4.

___________________________________

Johannesburg, South Africa Anton von Wielligh, Pr.Eng, BEng (Hons)

17 March 2009 Mining Engineer

A&B Global Mining Consultants



24.7 Alan Naismith



I, Alan Naismith, hereby certify that:

i. I am a partner with the firm SRK Consulting (South Africa) (Pty) Limited (SRK) with an

office at SRK House, 265 Oxford Road, Illovo, Johannesburg 2196, South Africa;

ii. I am a graduate of the Portsmouth Polytechnic UK with a BSc (Hons) Engineering Geology. I

obtained an MSC in Rock Mechanics and Excavation Engineering from the University of

Newcastle UK and an MBA from the University of Witwatersrand;

iii. I am registered as a Professional Natural Scientist with the South African Council for Natural

Scientific Professions. I am a registered fellow member of the South African Institute of

Mining an Metallurgy, and fellow member of the South African Institute of Rock Engineers;










viii. I am responsible for the preparation of the Geotechnical report described in item 20.4 of this

report;

ix. I have visited the properties of the Material Assets and my last visit was from May 12th to

16th, 2008;

x. I have prior experience at the properties having visited surface and underground mining

operations in November 2007 and March 2008.

___________________________________

Johannesburg, South Africa Alan Naismith, PrSciNat, MSc, MBA, FSAIMM, FSAIRE

17 March 2009 Associate Mining Consultant

SRK Consulting



24.8 Petrus Cilliers



I, Petrus Cilliers, hereby certify that:

i. I am a Process Manager with the firm Bateman Engineering with an office at Bartlett Road,

Boksburg, P O Box 25937, East Rand 1462, South Africa;

ii. I am a graduate of the University of Pretoria with BEng (Chemical Eng) 1988, and an MBA

(University of Pretoria) 1995. I have practiced my profession continuously since 1989;

iii. I am registered as a Professional Engineer with the Engineering Council of South Africa since

1993.










viii. I am responsible for the preparation of the sections describing the Mineral Processing and

Metallurgical testing described in items 18 and 25 (b) of the technical report;

ix. I have visited the properties of the Material Assets and my last visit was from 17-21

November 2008;

x. I have prior experience at the properties, including the period August 2006 to March 2008

when I worked for Hatch, which acted as the engineering contactor for Katanga Mining.

___________________________________

Johannesburg, South Africa Petrus Cilliers, Pr.Eng, BEng (Chem Eng), MBA

17 March 2009 Process Manager

Bateman Engineering



24.9 Rob McNeill



I, Rob McNeill, hereby certify that:

i. I am a Structural and Civil Engineer with the firm SRK Consulting (South Africa) (Pty)

Limited (SRK) with an office at SRK Pietermaritzburg, Suite 201, Sinodale Centre, 245

Burger Street 3201 Pietermaritzburg, South Africa;

ii. I am a graduate of the Zimbabwe (Rhodesia) Polytechnic with an Intermediate Diploma in

Civil Engineering T3 in 1977 and a National Diploma Civil Engineering T4 in 1978. I have

practiced my profession continuously since 1979;

iii. I am an associate member of the South African Institute of Civil Engineers, member of the

South African Project Management Institute and a member of the Institute of Waste

Management;










viii. I have visited the properties of the Material Assets and my last visit was on 13 and 14

November 2008;

ix. I am responsible for the preparation of the Tailings and Process Effluent Disposal report

described in item 20.5 of this report.

___________________________________

Johannesburg, South Africa Rob McNeill, Pr.Tech (Eng) MSAICE, MSAPMI, MIWM

17 March 2009 Structural/Civil Engineer

SRK Consulting



25 Additional Requirements for ProductionProperties

25a Mining Operations

25a.1 T17 Mine

25a.1.1 LoM PlanTables 25a.1 provides details of the LoM production schedule.

Table 25a.1 T17 Mine: LoM Production Profile

Unit Total 2009 2010 2011

1 2 3

Waste (kt) 22 433 18 248 3774 410

Strip Ratio (kt) 7,3 12,4 3,0 1,1

RoM Ore (kt) 3085 1476 1245 363

Cu (%) 2,67 1,71 3,35 4,27

(kt) 82 25 42 16

Co (%) 0,70 0,82 0,57 0,67

(kt) 22 12 7 2

25a.1.2 Mining Operations

The mining operations will be undertaken by contractors, Enterprise Generale Malta Forrest

(EGMF), which will be responsible for mining the pit as scheduled in the optimal (practical) design.

Pre-stripping will not be a separate phase of the schedule as stripping of overburden is planned to

coincide with the ore production.

The operation will be a conventional truck and shovel operation with drill and blast. The material is

loaded in-pit by 6 m3 hydraulic face shovels and transported by Terex TR45 t dump trucks to the

respective stockpile (ore to the run-of-mine pad and waste to the waste dump). The material will be

loaded by the face shovel from 5 m benches.

The work will be organized into 3 shifts of 8 hours each, 25 days a month (Sundays and public

holidays being days of rest). The performance of the maintenance and the organization of the mining

operations will ensure an availability factor of 80% for mining equipment and a utilization factor of

83% to 85%. The utilization factor includes the loss of time associated with the general organization

of the mine, in other words the operations associated with shift changes, blasting operations, and the

transfer of machinery and equipment. Production is therefore guaranteed by an absolute equipment

utilization rate of 68%. Sundays constitute a safety margin for maintenance as well as mining.



25a.1.3 Risks

The following risks were identified by SRK:

There is a risk that the grade in the upper levels of T17 Mine pit has been overstated, as the

actual grade mined from the current operations is not reconciling with the block model; and

A dewatering plan exists; however, no borehole studies have been made on the hydrology

around the pit and stability of the pit slopes below the water table.

Action plans to mitigate these risks have been developed by site management.

25a.2 Kamoto Mine

25a.2.1 LoM PlanTables 25a.3 and 25a.4 provides details of the LoM production schedule.


Kamoto Mine is an underground mine and various mining methods will be applied to the different

areas in the mine:

Room-and-Pillar: This method was proposed for the OBS in zones; 3, 4, 5 and 8 and for the

OBI on Zones 5 and 8;

Cut and Fill (“CAF”): This method was proposed for zones 1, 2, 6, 7, 9, 10, Etang North

and South and Division 5. Following a review of the flatter areas within these zones, it was

found that this particular method would not suit areas that are flatter than 55º, since the

blasted rock would not flow naturally to the level below (dip of mineralization is lower than

the natural angle of repose). These zones included Zone 1 Top and Etang North and South;

and

Long-hole retreat stoping (“LHRS”): This was proposed for areas on the OBI in Zones 3

and 4.

Table 25a.2 Kamoto Mine: Mining Methods

Zone 2008 Study

Zone 1 PPCF

Zone 2 CAF

Zone 3 OBS Room & Pillar

Zone 3 OBI LHRS


Zone 4 OBI LHRS


Zone 5 OBI Room & Pillar

Zone 6 CAF

Zone 7 CAF


Zone 8 OBI Room & Pillar

Zone 9 Sub-level caving

Etang South PPCF

Etang North PPCF

Zone 10 Depleted

Division 5 Excluded



Table 25a.3 Kamoto Mine: LoM Production Profile (2009-2023)

Unit Total 2009 2010 2011 2012 2013 2014 2015 2016 2017 2018 2019 2020 2021 2022 2023

1 2 3 4 5 6 7 8 9 10 11 12 13 14 15

Development (m) 19 402 4502 4044 3154 2518 1276 407 406 407 406 407 406 407 406 407 250

Ore tons (kt) 36 424 1177 1753 2102 2108 2169 2193 2121 2176 2094 2106 2088 2068 1964 1646 1453

Cu (%) 3,63 3,54 3,94 3,11 3,34 3,48 3,54 3,85 3,74 3,64 3,69 3,85 3,74 3,72 3,73 3,63

(kt) 1321 42 69 65 70 76 78 82 81 76 78 80 77 73 61 53

Co (%) 0,52 0,47 0,44 0,43 0,49 0,57 0,61 0,63 0,61 0,60 0,54 0,52 0,53 0,49 0,51 0,47

(kt) 190 6 8 9 10 12 13 13 13 13 11 11 11 10 8 7

Table 25a.4 Kamoto Mine: LoM Production Profile (2024-2038)

Unit 2024 2025 2026 2027 2028 2029 2030 2031 2032 2033 2034 2035 2036 2037 2038

16 17 18 19 20 21 22 23 24 25 26 27 28 29 30

Development (m) - - - - - - - - - - - - - - -

Ore tons (kt) 1349 1093 984 928 926 782 507 443 195 - - - - - -

Cu (%) 3,60 3,69 3,76 3,77 3,64 3,48 3,63 3,17 2,89 - - - - - -

(kt) 49 40 37 35 34 27 18 14 6 - - - - - -

Co (%) 0,50 0,48 0,47 0,48 0,45 0,44 0,40 0,48 0,51 - - - - - -

(kt) 7 5 5 4 4 3 2 2 1 - - - - - -


25a.2.3 Backfill

The type of fill used will be rockfill, which will be trucked from surface for the first five years of the

LoM plan. During this five year period backfill from tailings will be investigated. Provision has been

made in capital estimates for a backfill plant and distribution system.

25a.2.4 Ventilation

Ventilation is a constraining factor in the build-up phase of the project. The intake airways provide

sufficient quantities of fresh air, but the unavailability of return airways is a risk. The applied design

and sequence mitigates this risk, and each area has a ventilation system returning air from the

workings to a return air way that exhausts out of the mine.

25a.2.5 Survey

This design and schedule are highly dependent on the survey information used to generate the digital

data required by Mine 2-4D. Inaccurate survey information could result in a flawed Mineral Reserve

Statement. It is suggested that a survey audit is carried out to verify the validity of the information

and reconcile it with the digital data.

25a.2.6 Opportunities

The following opportunities were identified:

Backfill from tailings could be introduced at a later stage, and there may be an opportunity

to increase the extraction percentage of Mineral Resources by optimizing pillar sizes; and

Exploration drilling to increase confidence in the Inferred resources and their reclassification

as Indicated or Measured Resources. This would allow for conversion to Reserves and

inclusion in the LoM plan.

25a.3 KOV Mine

25a.3.1 LoM PlanTables 25a.5 and 25a.6 provides details of the LoM production schedule.


Table 25a.5 KOV Mine: LoM Production Profile (2009-2023)

Unit Total 2009 2010 2011 2012 2013 2014 2015 2016 2017 2018 2019 2020 2021 2022 2023

1 2 3 4 5 6 7 8 9 10 11 12 13 14 15

Waste (kt) 982 368 5 910 25 343 25 886 27 127 51 942 53 969 53 149 53 817 53 601 53 819 54 130 54 407 54 617 54 484 52 674

Strip Ratio (kt) 10,8 - 49,6 25,3 13,5 20,0 15,4 11,7 11,2 11,2 11,2 11,3 11,3 11,4 11,3 11,0

RoM Ore (kt) 90,149 - 511 1022 2013 2591 3504 4526 4817 4803 4803 4803 4817 4803 4803 4803

Cu (%) 4,93 - 1,83 2,66 3,95 5,16 6,05 5,20 5,74 6,29 5,00 4,58 5,32 3,73 4,71 5,26

(kt) 4449 - 9 27 79 134 212 235 276 302 240 220 256 179 226 253

Co (%) 0,38 - 0,19 0,16 0,29 0,45 0,40 0,26 0,45 0,38 0,34 0,54 0,63 0,55 0,61 0,35

(kt) 346 - 1 2 6 12 14 12 22 18 16 26 30 26 29 17

Table 25a.6 KOV Mine: LoM Production Profile (2024-2038)

Unit 2024 2025 2026 2027 2028 2029 2030 2031 2032 2033 2034 2035 2036 2037 2038

16 17 18 19 20 21 22 23 24 25 26 27 28 29 30

Waste (kt) 51 085 52 407 49 672 49 674 44 998 37 764 17 547 4345 - - - - - - -

Strip Ratio (kt) 10,6 10,9 10,3 10,3 9,3 7,9 3,7 1,1 - - - - - - -

RoM Ore (kt) 4817 4803 4803 4803 4817 4803 4803 3878 - - - - - - -

Cu (%) 4,98 5,35 4,23 4,12 4,59 4,88 5,03 5,24 - - - - - - -

(kt) 240 257 203 198 221 234 242 203 - - - - - - -

Co (%) 0,37 0,31 0,42 0,35 0,32 0,24 0,26 0,17 - - - - - - -

(kt) 18 15 20 17 16 11 12 7 - - - - - - -


25a.3.2 Mining OperationsThe operation will be a conventional truck and shovel operation with drill and blast. The activities

considered include; pre-stripping of free-dig rock, blast-hole and smooth blasting drilling, load and

haul to and from stockpile, load and haul to the ore and waste crushers and to the waste dumps.

Allowance for mining equipment was made by others for the rehabilitation of the waste dumps, pit

dewatering, crushing and conveying of waste rock, workshop facilities, and electricity supply and

distribution.

The currently purchased equipment that was considered is as follows:

RH 340 Face shovel;

CAT 994 Front-end loader;

CAT 793 Dump truck;

PV-271 Drill rigs;

CAT D10 Trackdozers;

CAT 834 Wheeldozers;

CAT 16 Grader;

Cat 777 Watercart and construction truck; and

CAT 992 construction loader.

25a.4 Mashamba East Mine

25a.4.1 LoM PlanTables 25a.7 provides details of the LoM production schedule.


The operation will be a conventional truck and shovel operation with drill and blast. The material

will be loaded in-pit with hydraulic face shovels with 21 m3 buckets and transported by haul trucks

to the respective stockpile (ore to the RoM pad and waste to the dump). The broken material will be

loaded by the face shovel from 10 m benches.

25a.4.3 Risks

Slimes in the bottom of Mashamba East pit could be underestimated. This could influence the

production schedule in a manner that the start and build-up could not be achieved, since additional

time would be required to clean the pit bottom. There are currently no preventive measures available

to minimize this risk. Volume estimations can be done using sonar techniques in the mean time.


Table 25a.7 Mashamba East Mine: LoM Production Profile (Base Case 2009-2027)

Unit Total 2009 2010 2011 2012 2013 2014 2015 2016 2017 2018 2019 2020 2021 2022 2023 2024 2025 2026 2027

1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19

Waste (kt) 138 270 - - - - - - - - - 9021 9021 9295 9295 19 057 19 029 18 700 25 350 18 541 960

Strip Ratio (kt) 13,6 - - - - - - - - - 14,1 14,1 17,5 17,5 18,7 15,3 15,3 14,6 10,4 1,1

RoM Ore (kt) 10 190 - - - - - - - - - 638 638 532 532 1021 1241 1220 1738 1789 839

Cu (%) 4,39 - - - - - - - - - 4,53 4,53 4,77 4,77 4,35 4,92 4,50 4,37 4,07 3,50

(kt) 447 - - - - - - - - - 29 29 25 25 44 61 55 76 73 29

Co (%) 0,52 - - - - - - - - - 0,45 0,45 0,55 0,55 0,54 0,50 0,59 0,57 0,52 0,41

(kt) 53 - - - - - - - - - 3 3 3 3 6 6 7 10 9 3



25b Recoverability

25b.1 Source of Information

The data contained in this section pertaining to the Kamoto Concentrator and Luilu Metallurgical

Plant has been sourced from KML and that for the WOL/SX/EW Refinery Project has been

compiled using data sourced from the Bateman Engineering study.

25b.2 Introduction

As mentioned in Section 5, prior to January 2008, Nikanor and KML pursued separate projects on

adjacent concessions.

Processing assets awarded to Nikanor included the Kolwezi Concentrator and Luilu Electro-refinery.

Following a review of these facilities in 2005, it was concluded that it was not viable to rehabilitate

them. In June 2006, Global Enterprises Corporate Ltd (“GEC”) completed a Definitive Feasibility

Study (“DFS”) on the DRC Copper/ Cobalt Project SARL (“DCP”) on behalf of Nikanor. Processing

aspects of the study were sub-contracted to Bateman. The DFS proposed a new hydrometallurgical

facility with a capacity of 250 ktpa annum of LME Grade ‘A’ copper cathode and 30 ktpa of cobalt

as a hydroxide salt. Bateman subsequently continued with front-end engineering services for the

DCP project including the placement of orders on long-delivery items such as mills, thickeners and

an acid plant.

Processing assets awarded to Katanga included the Kamoto Concentrator with a historical capacity

of 7,5 Mtpa ore and the Luilu Metallurgical Plant with a nameplate capacity of 175 ktpa copper and

8 ktpa cobalt cathode. In July 2006, Katanga embarked on Phase 1 of a four-phase rehabilitation

programme to return the Kamoto Concentrator and the Luilu Metallurgical Plant to close to the

nameplate capacity. Hatch was commissioned to undertake the engineering design and Phase 1 (the

phases are explained in detail later in the Section) was commissioned in December 2007.

In January 2008, Katanga merged with Nikanor and Katanga assumed management responsibility of

the combined mining assets through KOL. Bateman Engineering was appointed to undertake an

Engineering Study, including the rationalisation of the processing projects for the merged assets

where synergies existed. This was completed initially in October 2008 but revised in January 2009

due to adverse developments in the world financial markets and is the subject of this report.

25b.3 Ore Sources

Ore will principally be derived from the KOV and Kamoto Mines. Other minor sources of ore will

include the T17 Mine in the near term and the Mashamba East Mine in later years, which will

primarily provide oxide ores.

25b.3.1 Ore Type and Mineralogy

The most important ore types that make up the bulk of the KOL ore resources are sedimentary or

stratiform copper-cobalt type. The copper-cobalt deposits comprise both oxide and sulphide ores.



The principal copper and cobalt sulphide minerals are chalcocite [Cu2S] and carrolite [(Co,Cu)2S4]

respectively. Subordinate sulphide minerals of importance include bornite [Cu5FeS4], covellite [CuS]

and chalcopyrite [CuFeS2], which is generally present in small quantities.

Supergene mineralisation is generally associated with the levels of subsurface oxidation, sometimes

more than 100 m below surface, where the sulphide minerals have been altered to carbonates,

silicates and phosphates. The most common secondary copper mineral is malachite [Cu2CO3(OH)2]

with lesser minerals including cuprite [Cu2O], cornetite [Cu3(PO4)(OH)3], liberthenite [Cu2

(OH)PO4] and pseudomalachite [Cu5(PO4)2(OH)4.H2O]. Small quantities of copper silicates such as

chrysocolla [CuSiO3.nH2O] and copper carbonates such as azurite [Cu3(OH)2(CO3)2] also exist.

Heterogenite [(Co,Cu,Mn,Fe)O(OH)] and kolwezite [(Cu,Co)2(CO3)(OH)2] are the principal

secondary cobalt minerals, with some goethite[(Fe,Co)O(OH)].

The predominant gangue mineral is quartz [SiO2], but dolomite [Ca,Mg(CO3)2] is present in varying

and very significant amounts. Other notable gangue components are mica [KMg3Si3AlO10(F,OH)2],

clay [K-Al-Mg-Fe silicate hydroxides] and chlorite [(Mg,Fe)5Al(Si3Al)O10(OH)8].

25b.3.2 KOV Mine

Historically, KOV Mine has been treated as an oxide deposit. However, the latest information

suggests there is 30-50% sulphide in the remaining ore.

Typical primary copper sulphide minerals are bornite, chalcocite and minor chalcopyrite, while

cobalt is in the form of carrolite. The mineralization occurs as disseminations or in association with

hydrothermal carbonate alteration and silicification.

The most common secondary supergene minerals for copper and cobalt are malachite and

heterogenite. Quartz, mica and dolomite are the predominant gangue minerals. KOV Mine ore is

best treated by sulphide flotation followed by sulphuric acid leaching of flotation tails.

25b.3.3 Kamoto Mine

Kamoto Mine’s mineralization is primarily sulphidic. Chalcocite and bornite are the major copper

containing minerals occurring roughly in the ratio 1.5:1 predominantly as liberated grains. Trace

amounts of covellite and chalcopyrite are also present.

Kamoto Mine ore is amenable to sulphide flotation.

25b.4 Metallurgy

The oxide and sulphide ores are typically treated by flotation to recover a higher grade concentrate

that is then further treated in the refinery to recover a proportion of the copper and cobalt contained

in the flotation concentrate. However, the oxide ores are very weathered and display poorer flotation

response, which results in lower metal recoveries, generally lower copper and cobalt grades in

concentrates and higher operating costs in comparison to sulphide ores. Acid consumption in

leaching the oxide concentrate is also higher and a mix of oxide and sulphide concentrates is

considered optimal feed.

A key distinction between the KOV Mine ore and the Kamoto Mine ore is that, whilst the latter is

almost entirely sulphidic, the former contains approximately 60% oxide ore underlain by sulphide

ore. In terms of processing, early KOV Mine ores will be net consumers of sulphuric acid in



leaching, whilst the Kamoto Mine ore will generate sulphuric acid via roasting ahead of leaching.

The merger thus presents significant synergies by interlinking the two process facilities.

25b.5 Processing Facilities

25b.5.1 Kolwezi Concentrator

The Kolwezi Concentrator was originally commissioned in 1940. After years of neglect and the

onset of war, production through this concentrator steadily declined. The plant was recently

refurbished by DCP to the extent that would allow ongoing treatment of ore from Tilwezembe mine

(which has recently been stopped). Generally however, the condition of the Kolwezi Concentrator is

very poor and following the merger it was considered preferable to refurbish the Kamoto

Concentrator due to its better condition and more appropriate sulphide processing facilities.

Operations at Kolwezi Concentrator were discontinued in November 2008.

25b.5.2 Luilu Electro-Refinery

In a refurbishment programme initiated by UMHK in 1987 an electro-refining circuit consisting of a

copper melting and anode casting facility and a 100 ktpa electro-refining cell house was constructed

at Luilu. The melting and casting facility was apparently only operated for a few weeks before a

runaway halted production and only one of the three sections of the cell house were ever operated.

The apparently obvious option of converting the electro-refining cell house to an electro-winning

cell house was considered but was shown not to be viable when compared to constructing a new 250

ktpa electro-winning plant. This could possibly be reconsidered since the new plant capacity has

recently been reduced to 160 ktpa.

25b.5.3 Kamoto Concentrator

The original Kamoto concentrator consists 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 hydrocyclones;

Sulphide flotation including roughing, cleaning and middlings regrind to produce a sulphide

concentrate;

Sulphidisation with NaHS; and

Oxide flotation including roughing and cleaning to produce an oxide concentrate.

Kamoto 2 primarily treated sulphide ore from Kamoto Mine. The circuit comprised the following

unit processes:

Autogenous milling operating in closed circuit with hydrocyclones; and

Sulphide flotation including roughing, cleaning and middlings regrind.

The DIMA 1 circuit primarily treated oxides and mixed banded oxide / sulphide ore feeds. The

circuit comprised the following unit processes:

Primary autogenous milling and secondary ball milling operating in closed circuit with

hydrocyclones;



Sulphide flotation including roughing, cleaning, re-cleaning and middlings re-grind to

produce a sulphide 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

hydrocyclones;



The Bateman Engineering Study for Phase 5 reports that, from 1969 to 2000, over 126 Mt of ore at

an average grade of 4,33% copper and 0,28% cobalt was processed through Kamoto Concentrator.

Typical flotation recoveries achieved by Gecamines for oxides and sulphides in this period are

summarised in Table 25b.1 below.

Table 25b.1 Gecamines Historical Recoveries

Ore Type Copper Recovery Cobalt Recovery

Sulphides 85% 75%

Oxides 75% 45%

It is reasonable to expect that recoveries in excess of these historical figures can be achieved, given

the use of new improved equipment (e.g. larger flotation cells), improved reagents and improved

plant control. Higher values have been assumed in the study report and in KOL forecasts. Prior to

being taken over by KCC, the condition of the Kamoto Concentrator had been allowed to deteriorate

badly and only the portion of the plant processing sulphides was operating. Significant progress has

been made recently by KOL in refurbishing sections of the plant as required for the phased ramp-up

of the plant.

Based on the mining schedule assumed in the study, the refurbishment will continue during Phases 3

and 4, including much of the DIMA section, primarily to process KOV ore.

25b.5.4 Luilu Metallurgical Plant

Production at the Luilu Metallurgical Plant, located approximately 6 km north of the Kamoto

Concentrator, commenced in 1960. The process route employed was roast-leach-electro-winning

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 electro-winning (“EW”) onto copper starter sheets (being converted to stainless steel

blanks);

De-copperising of cobalt bleed solution – two-stage EW;

Cobalt bleed solution purification including the following steps;

o Iron removal by controlled pH precipitation using milk of lime;

o Copper removal by two-stage controlled pH precipitation using milk of lime;

o Nickel removal via precipitation with sodium hydrogen sulphide (NaHS) and cobalt

chips under controlled pH;

o Zinc removal by the addition of hydrogen sulphide (H2S) and neutralisation with

sodium carbonate solution;

o Controlled pH precipitation of cobalt with milk of lime;

o Cobalt re-leaching with spent electrolyte and sulphuric acid under controlled pH;

Cobalt EW; and

Cobalt vacuum degassing and burnishing.

The Luilu Metallurgical Plant was designed to process sulphide and oxide concentrates with an

initial capacity of 80 ktpa copper cathode. During the 1970s capacity was expanded to 175 ktpa

copper cathode and 8 ktpa 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 only 45 to 60%, with the majority of the cobalt losses occurring

at nickel and zinc sulphide precipitation with some also at iron removal and cobalt precipitation. By

means of improved pH control throughout the circuit and elimination of sulphide precipitation, it

should be possible to improve cobalt recovery up to 85%. The Nikanor testwork confirms this.

The condition of the plant in 2006, when taken over by KCC, was extremely poor and almost totally

run down. For this project, a progressive renewal programme was planned, to match the increasing

throughput. Considerable progress has been made to-date by KOL in their phased rehabilitation

exercise. Completion of Phase 1 was December 2007, with the bulk of Phase 2 planned to be

completed by March 2009. A new roaster is being built as part of Phase 2 with completion expected

in Q2 2009.

25b.5.5 WOL/SX/EW Refinery Project

The proposed new circuit will be installed in phased “modules” (see below) and will ultimately

comprise the following unit processes:

Primary crushing;

Milling in closed circuit with hydrocyclones;



Sulphide flotation comprising roughing, scavenging, cleaning and re-cleaning stages;

Pre-leach dewatering (thickeners);

Sulphuric acid copper leach of flotation tailings;

Sulphuric acid cobalt leach of flotation tailings (reducing leach assisted by addition of SO2);

Post-leach thickening and counter-current decantation;

Leach residue disposal;

Clarification of High-Grade and Low-Grade pregnant leach solutions;

HG and LG copper solvent extraction (“SX”);

Electrolyte filtration;

Copper EW;

Cobalt bleed solution purification including the following steps;

o Iron removal by controlled pH precipitation using milled limestone slurry;

o Manganese precipitation via contact with an air/ SO2 mixture;

o Aluminium and copper removal by two-stage controlled pH precipitation using milk

of lime;

o Cobalt hydroxide precipitation with milk of magnesia slurry;

Cobalt hydroxide filtration, drying and packaging; and

Effluent treatment by precipitation with milk of lime at pH 10.3.

The new and old circuits will be linked by means of transferring the old leach residue and a spent

electrolyte bleed to the new WOL section, with a corresponding volume bleed of SX raffinate back

to the old leach. This will eliminate the need for the old secondary leach and secondary EW.

In the early years (up to 2014) tailings and leach residue will be pumped to an interim dam; then to

the old Mupine Pit (in years 2015 to 2019). Thereafter they will be disposed of in a new,

conventional, unlined tailings dam with the slurry being distributed around the perimeter by a system

of moveable pipes. Surplus clear water will be allowed to overflow a penstock pipe into a return

water dam ahead of recycle to the process. Surplus water, if any, will overflow to the Luilu River,

after being appropriately treated. A sulphur-burning acid plant will be constructed to support the

whole ore leach and a sulphur dioxide liquefaction plant will also be installed to provide

concentrated SO2 gas to the following consumers:

Reductive leach; and

Fe/Mn removal in the Cobalt section.

Crushed limestone (CaCO3) and burnt lime (CaO) will be sourced from the DRC supplier, the CCC

Lime Plant at Likasi or from sources outside the DRC, which may include Turkey, and transported

200 km by rail to site. The crushed limestone will be milled to produce a finely ground limestone

slurry. Burnt lime will be slaked in a slowly rotating ball mill to produce a fine milk of lime slurry.



The process design criteria assume that all reagents, including particularly lime, limestone, sulphur,

magnesia, flocculants, SX extractant and SX diluent, will be available in the amounts and qualities

specified in the design. In SRK’s view, the supply of key reagents such as lime, magnesia and

sulphur need to be further reviewed and confirmed contractually as soon as practical.

Utilities will include the generation of steam with electrode boilers, compressed air generation,

production of demineralised water with reverse osmosis and a fire water system. Raw water for the

plant will be sourced from the KOV Mine pit dewatering operation. Chlorinated potable water will

be produced on site utilizing filtered raw water.

25b.5.6 Process Plant Capacity

Copper Capacity

The target capacity of the merged facility is 310 ktpa cathode copper, which will be produced in both

the Luilu Metallurgical Plant and the proposed WOL/SX/EW Refinery Project. In November 2008,

the capacity of the new copper circuit was reduced from its original design capacity of 250 ktpa

cathode to 160 ktpa, with the balance of up to 150 ktpa to be produced in the refurbished old refinery

section. It is important to note that only the new copper circuit will produce LME grade ‘A’ cathode.

It is also important to note that appropriate redesign work has not been possible in the time available

since the decision to reduce capacity.

Cobalt Capacity

The target capacity of the merged facility will be 27 ktpa cobalt. It was initially intended that all

cobalt would be produced as cobalt hydroxide in the new cobalt circuit and that the old cobalt plant

would be closed down. However, it is now intended to continue with the production of cobalt

cathode in the refurbished old plant, increasing to the original nameplate capacity of 8 ktpa. This will

require further study.

Phased Capacity Expansion

The capacity expansion of processing facilities has been planned using a phased approach. The

capacities and timing of the phased refurbishment of the old Luilu Metallurgical Plant and the

development of the proposed new SX/EW Refinery project are included in Tables 25b.2 and 25b.3.

Table 25b.2 Capacity Expansion: Refurbishment Phases

Capacity Capacity Timing Comment

Cu (ktpa) Co (ktpa)

Phase 1 35 2.0 End 2007 Complete

Phase 2 +35 +4,0 March 2009 In progress - Incl. 1st roaster

Phase 3 +40 +3,0 1st Quarter 2012 110 ktpa total Cu capacity

Phase 4 +40 +1,0 1st Quarter 2013 Incl. 2nd roaster

Total 150 10,0 * 8ktpa metal plus 2 ktpa hydroxide

The original DCP WOL/SX/EW Refinery design was modified so that construction could be

scheduled in three separate modules, at approximately yearly intervals consistent with the mining

plans. However, recent changes to the mining plans to better balance mine production and operating

capacity have resulted in only two distinct new plant modules now being required. The timing and

capacities of the new modules as finally proposed are summarised in Table 25b.3.



Table 25b.3 WOL/SX/EW Refinery Modules: Timing and Capacities

Capacity Capacity Timing Comment

Cu (ktpa) Co (ktpa)

Modules 1& 2 80 10 1st Quarter 2014 Modules 1 & 2 combined, only 1 EW

Module 3 80 9 1st Quarter 2015 Extra Leach, SX & EW

Total 160 19

The phased refurbishment of old plant will be complete by year 2013, with two new plant modules

coming on line in 2014 and 2015. Careful planning will be required to achieve this as outlined

elsewhere in this study.

The majority of the process flow sheet outlined will be installed in combined Module 1/2, including:

Whole ore leach (1 train);

High-grade SX (1 train);

Low-grade SX (1 train); and

Copper EW (one 80 ktpa cell house).

Module 3 will double copper capacity by means of:

Second whole ore leach train;

Second high-grade SX train; and

Second 80 ktpa cell house module.

25b.6 Copper and Cobalt Recovery

Mined ore will be directed to specific processing routes on the basis of the five ore types as shown

schematically in Figure 25b.2. The recovery of copper and cobalt attributable to each ore type has

been based on historical performance in the case of Kamoto Mine ore and test results in the case of

KOV Mine ore. No account has been taken of recovery dependence on head grade. Overall

recoveries, including refining losses, that were assumed for planning purposes in the Engineering

Study and operating model are summarised in Table 25b.4 below.

Table 25b.4 Assumed Copper and Cobalt Recovery by Ore Type

Ore Type Proportion Copper Cobalt

In Feed (%) Recovery (%Cu) Recovery (%Co)

A - Sulphide 30% 79,9% 68,4%

B – Mixed Oxide/Sulphide* 60% 87,1% 78,4%

C – High acid-consuming mixed* 5% 75,6% 49,4%

D – Oxide only (low-grade) 5% 72,8% 36,6%

E – Oxide only (high-grade) 0% 92,0% 85,0%

* These recoveries are based on 40% sulphides in mixed ore and should be better defined. It is recommended that further work be

conducted to ascertain the accuracy of these ratios.

NOTE: Recoveries actually being achieved and budgeted currently differ from the above due to a number of reasons including the

incomplete status of the plant and non-standard operating conditions.



25b.7 Processing Schedule

25b.7.1 Ore from KOV Mine

RoM ore will be delivered from the KOV pit to B3 primary crusher by haul trucks, from where the

crushed ore will be conveyed to the Kamoto Concentrator. Planned design ore feed tonnages and

grades once all modules are completed are included in Table 25b.5 based on the current mine plan.

Table 25b.5 KOV Mine Design Feed Tonnage and Grade

Description Unit Value

Mean feed capacity ktpa 4 800

Mean copper grade %Cu 4,93

Mean cobalt grade %Co 0,38

KOV ore will initially be processed through an oxide flotation line at the Kamoto Concentrator to

generate an oxide concentrate that will be pumped overland to the Phase 2, 3 and 4 plant. Thereafter,

from Module 1 and 2 onwards, the ore will be sulphide floated in a new section at KTC. The

sulphide concentrate will join the Kamoto sulphide concentrate stream and the flotation tails (oxides)

will not be floated but will be pumped direct to the new whole ore leach to recover acid-soluble

oxide copper and cobalt.

25b.7.2 Ore from Kamoto Mine

RoM from the Kamoto Mine is crushed underground and conveyed to the Kamoto Concentrator.

Planned design ore feed tonnages and grades are summarised in Table 25b.6. Kamoto Mine ore will

be processed through the Kamoto Concentrator sulphide flotation and the sulphide concentrate will

be pumped to Luilu Metallurgical Plant (roasters). It will be important to control mass pulls to

achieve a concentrate grade greater than 12,5% sulphur, to ensure autothermal operation of the

roasters. Historically, Hatch did limited testwork on the current concentrate samples to design the

roaster currently being constructed. Further testwork is recommended to comfirm 12,5% sulphur in

concentrate can be achieved from all mixed ore types, as well as to assess the impact a variation on

this specification will have. The leach receovery of copper and cobalt post roaster was based on

historical data, obtained and evaluated by Hatch.

Table 25b.6 Kamoto Mine Design Feed Tonnage and Grade

Description Unit Value

Mean feed capacity ktpa 2100

Mean copper grade %Cu 3,59

Mean cobalt grade %Co 0,53

25b.7.3 Other Ore Sources

In addition to the KOV Mine and Kamoto Mine ores, the following oxide and mixed oxide/sulphide

ores are available: T17 Mine, which is a source of oxide ore (type ‘D’ in Figure 25b.2) to be utilised

during years 2009 - 2011; and Mashamba East Mine, which is a source of mixed ore (80% oxide -

type ‘B’ in Figure 25b2) to be utilised from year 2018 on.



Figure 25b.1: Metallurgical Processing: Final Block Diagram



Figure 25b.2: Metallurgical Processing: Distribution by Ore Type



25c Marketing Study

25c.1 Introduction

The Luilu Metallurgical Plant will produce copper cathode and cobalt in the form of cobalt cathode

plus some hydroxide salt. The WOL/SX/EW Refinery Project will produce LME grade ‘A’

electrowinning cathode and cobalt hydroxide. The old refinery will produce a lower grade

“commercial” copper cathode and cobalt cathode approximately 99,3% pure. This section has been

prepared from information supplied to KML by LN Metals Advisory Services and CRU Strategies,

part of the CRU Group.

25c.2 Copper Marketing

It is anticipated that the marketing will follow the same structure as that employed for Luilu copper

cathodes. KOL will therefore be a producer of copper and cobalt from two different plants, the

refurbished Luilu Metallurgical Plant and the new WOL/SX/EW Refinery Project.

Being new production from a new refinery, the WOL/SX/EW Refinery Project copper will have

none of the quality issues associated with the original Luilu Metallurgical Plant cathodes. The

marketing, however, will follow the same structure as that employed for Luilu Metallurgical Plant

cathodes. The marketing will include:

Marketing the initial production during the start-up phase;

Producing and supplying Grade A quality material to the international market place.; and

LME registration and long-term marketing.

The marketing of the initial production will be governed by the same criteria as those for the Luilu

cathodes regarding quality standards. The objective of KOL is to produce a cathode suitable for rod

mills and registration as LME Grade A copper. Impurity levels must be kept well below even the

maximum levels specified in the LME contract.

25c.3 CRU Copper Price Forecasts

Over the second half of 2008, the copper price fell from a high in early July of almost USD9000/t to

a low in December of USD2902/t – a decline of more than 60%. Even the recent rebalancing of

commodity indices has had little effect, lifting copper prices to the USD3000-3400/t range in early

January. The LME 3-month copper price averaged USD6893/t in 2008, down 2,9% on the 2007

average price.

World copper demand is expected to fall by 4,0%, or more than 700 kt in 2009, to reach 17,4 Mt.

The contraction will be particularly pronounced in the developed regions of North America, Japan

and Western Europe. The supply-side response from the copper industry (in contrast to other

commodities) is expected to be modest, with production set to fall by just 0,3% year-on-year.

Significant volumes of refined production capacity are expected to come on stream, particularly in

China and Chile, which will offset cutbacks at existing smelters. As a result, a market surplus of

572 kt is forecast for 2009, and prices are expected to average just USD2800/t, down 59% on the

2008 average. Another surplus, of 463 kt is forecast for 2010, despite some recovery in demand.



Prices will continue to slide, to a low of USD2694/t in 2010, as a result of a cumulative surplus of

just over 1Mt building up over this two year period.

The struggle for financing is likely to go on for some time, which will limit the entry of new supply

to the market in the medium term. Potential lenders are becoming increasingly risk averse and

companies and projects are coming under heightened scrutiny, especially in light of the current low

prices. With prices forecast to fall further, additional output cuts are likely, as some producers will

be unable to cover even their cash costs.

However, stocks will have built up considerably, and renewed consumption growth will be a

prerequisite for higher prices to emerge. CRU expect that supply disruptions will continue, so some

short term price rallies cannot be ruled out. However, without the support of robust demand growth,

these will be unsustainable. CRU do not anticipate an economic recovery in 2009, although demand

growth in the last quarter will be positive when compared to a weak Q4 2008. In 2010, CRU

anticipate that demand growth will pick up, and will strengthen in 2011 as the effects of

infrastructure packages start to boost copper demand and the world economy revives.

Demand growth thereafter will be more moderate, returning to trend by 2013. Also, the period of

lower copper prices will take the heat out of substitution pressures which may otherwise have

permanently dented growth prospects.

As the effects of the above flow through, CRU Strategies forecasts that the copper market will be in

balance by 2011, returning to deficit in both 2012 and 2013; as a result prices will rise to reach an

average of USD3280/t in 2013.

On the production side, the financing delays will limit refined production growth over the medium

term, with a risk that low prices and tight credit availability will result in more production cutbacks

than CRU have forecast. However, in the period 2014 to 2020, CRU expect the impact of the

financial crisis and economic downturn to have abated, and a recovery to be underway.

China will continue to account for the majority of copper demand growth to 2020, while India will

also become a significant player, and other developing regions, such as Africa, the Middle East and

South East Asia will see strong growth over the period. In contrast, the industralised countries of

North America and Western Europe and Japan will experience very moderate growth or a

contraction as their economies mature and their manufacturing base shifts towards the emerging

countries.

On the supply side, the effects of medium term project delays will be felt, as the market emerges

from the shadow of the financial crisis. As a result, copper prices are expected to move higher over

the 2014 to 2020 period, moving above USD4000/t, in nominal terms, in 2014, USD5000/t in 2015

and then finally USD6000/t in 2020.

25c.4 Cobalt Marketing

The planned production levels of cobalt are bound to have an effect on the market and subsequently

on the price. Since it is the metal price that sets the price base for the whole market, an expansion in

cathode output largely would have a much greater effect on the price than an increase in cobalt

carbonate or hydroxide would.

In a metal market of about 37 kt, an increase of 8 kt in 2009 is very significant. As with copper, the

quality of cobalt metal is important and, in the early years of production, is unlikely to be high



enough to achieve sales to the super-alloy industry. It will be more suitable for the chemical industry,

in which the Chinese and Umicore are the largest participants. This automatically excludes a

significant annual quantity of metal from a very important sector of the market.

Therefore, it is recommended that KOL should produce a range of products from metal through to

carbonate and hydroxide with at least 50% in the latter two forms. The effect on the metal market

and price will thus be reduced. Even so, the price in 2009 could fall to around USD 15/lb.

25c.5 CRU Cobalt Price Forecasts

In 2008, the High Grade (99,8%) cobalt price averaged USD39/lb, an increase of 39% year-on-year.

During late Q1 and early Q2, prices above USD50/lb were sustained, and at the time of CRUs

previous report, in August 2008, the cobalt price stood at approximately USD30/lb. However,

deteriorating macroeconomic conditions, and a collapse in cobalt demand, led prices to plummet

during the latter months of 2008. The cobalt price dropped below USD20/lb in mid-November, and

reached a low of USD13/lb in December. In January 2009, the price has stabilised in the

USD16-19/lb range, though trading volumes have remained low.

All the major demand sectors have seen a considerable slump in business, and this has resulted in a

decline in world cobalt consumption of 4% in 2008, to approximately 57 kt. The US auto industry is

still reeling from the impact of the global crisis, and Boeing has recently announced a sharp drop in

2008 orders (as well as deliveries which were impacted by a long-running strike at the company’s

main production facility). Furthermore, key electronics manufacturers, such as Nokia and Motorola,

have announced downwardly revised earnings and market forecasts. This is further evidence of a

demand slowdown, and will put additional pressure on battery manufacturers. Cobalt consumption is

expected to fall by a further 6% in 2009, to less than 54 kt. As a result, the cobalt price is forecast to

average USD19/lb in 2009, a fall of more than 50% from 2008.

CRU Strategies expects that a recovery in cobalt consumption will begin in 2010, with global

demand expected to increase by 10%, to more than 59 kt – or approximately the level of

consumption in 2007. Cobalt demand growth is expected to average 4,6% per year between 2008

and 2015, notwithstanding the forecast contraction in 2009.

The supply-side has responded rapidly to the recent collapse in cobalt demand and prices. In 2008,

cobalt production fell by 2,4% year-on-year, to less than 55,2 kt. Further falls in production are

expected in 2009. Little or no new supply will come on stream, as a result of the decreased

availability of finance, as well as the weak market fundamentals. Sharp production cuts have also

been implemented by existing producers. The DRC has forecast that cobalt production will be halved

during Q1 2009, due to factory closures and decreasing demand. CMSK, which was expected to

produce 3500 t in 2008, announced that it stopped operating on 20 December 2008, citing low prices

and a lack of demand. Other suspended operations include Anvil Mining’s Dikilushi facility, while

Luanshya Copper Mine (LCM), Zambia’s largest cobalt producer, has suspended its Chambesi

operations. 2009 could still see a small market surplus, but there is considerable uncertainty

surrounding the magnitude and duration of the industry’s production cuts.

As a result of production cutbacks, and the expected delays to a raft of new projects, which will limit

any market surplus, the current slump in cobalt prices is forecast to be less prolonged than for other

metals. CRU Strategies expects prices to increase to USD24/lb, on average, in 2010, as a result of a

recovery in demand, coupled with the limited availability of material as a result of supply cuts.



Cobalt prices are not expected to return to levels seen in 2007 and the first half of 2008, unless the

cutbacks in supply and concerns about future production volumes are able to offset falling demand,

which is unlikely. Beyond 2010, market surpluses are expected to build up, as current production

cutbacks are reversed and new supply begins to come on stream, which will result in a decline in

prices, to USD18/lb in 2011 and USD14/lb in 2012.

In the longer term, CRU Strategies still expects the cobalt market to be significantly over-supplied,

which will lead the cobalt price to bottom out at USD9/lb in nominal terms in 2015. Beyond the

medium term, the cobalt price is expected to converge towards the long run marginal cost of

USD10/lb, in real USD2008 terms.



Table 25c.1 Marketing Study: CRU Strategies Prices and Applied Prices (2009-2023)

2009 2010) 2011 2012 2013 2014 2015 2016 2017 2018 2019 2020 2021 2022

January 2009 MoneyTerms Prices

LME PayFactor

Unit 1 2 3 4 5 6 7 8 9 10 11 12 13 14 LOM

CRU Strategies Base CasePrices

Copper (USD/lb) 1,24 1,17 1,24 1,31 1,36 1,65 2,05 2,07 2,10 2,11 2,12 2,14 2,14 1,84 1,84

Cobalt (USD/lb) 18,00 23,00 17,00 13,00 9,00 9,00 8,00 9,00 9,00 11,00 10,00 10,00 10,00 10,00 10,00

Applied Prices

Luilu Metallurgical Plant-copper cathode

(USD/lb) 1,19 1,12 1,20 1,26 1,32 1,61 2,01 2,03 2,04 2,06 2,08 2,09 2,09 1,82 1,82

Luilu Metallurgical Plant-cobalt metal

(USD/lb) 18,00 23,00 17,00 13,00 9,00 9,00 8,00 9,00 9,00 11,00 10,00 10,00 10,00 10,00 10,00

WOL/SX/EW RefineryProject - copper cathode

(USD/lb) 1,24 1,17 1,24 1,31 1,36 1,65 2,05 2,07 2,10 2,11 2,12 2,14 2,14 1,84 1,84

WOL/SX/EW RefineryProject - cobalt salt

(USD/lb) 14,40 18,40 13,60 10,40 7,20 7,20 6,40 7,20 7,20 8,80 8,00 8,00 8,00 8,00 8,00

Kolwezi concentrate - copperin concentrate

(USD/lb) 0,78 0,74 0,78 0,83 0,86 1,04 1,29 1,30 1,32 1,33 1,34 1,35 1,35 1,16 1,16

Kolwezi concentrate - cobaltin concentrate

(USD/lb) 9,00 11,50 8,50 6,50 4,50 4,50 4,00 4,50 4,50 5,50 5,00 5,00 5,00 5,00 5,00

Pay Factors reflect terms of off-take agreements with Glencore, a Swiss-based commodity trading company.



25d ContractsSRK reviewed the following contracts, which are within industry norms:

The contract with the mining contractor for the T17 Mine. The terms of this contract are

typical of mining contractors operating in central Africa;

The transportation agreement for the transport of finished product from Kolwezi to the port

at Durban, South Africa; and

The SNEL power agreement for the repair of the Zongo, Mwandingusha and Koni Groups,

the repair of the associated transmission networks and the compensation of reactive energy

at Kimwenza (relates to the joint upgrade of power infrastructure).

SRK relied on the legal advisor to KML, Norton Rose LLC, for the legal tenure and mining rights

status of KML as it applies to all the Material Assets mentioned in this report.

25e Environmental ConsiderationsThe following section includes discussion and comment on the environmental management aspects

of the Mine. Comment is included on the status of environmental legislation applicable to the mine,

compliance with legislation, key liabilities and risks, and opportunities to overcome several of these

liabilities and risks.

This section of the report has been prepared on the basis of:

The Draft Environmental and Social Impact Assessment (ESIA) for the KOL Mine project,

completed and submitted to Katanga Mining Limited (KML) in December 2008;

The Environmental Impact Statements (EIS) completed for KOL1 and DRC Copper and

Cobalt project (DCP);

SRK’s experience in the Kolwezi area - SRK is approved as a consultant (by the Ministry of

Mines in the DRC) to carry out environmental studies and to submit environmental reports

locally (Ministerial Order No. 334 / CAB.MIN / MINES / 01 / 2004) and has been involved

in the key mining projects in Kolwezi since 2002.

For the purpose of this section liability and risk are defined as follows: a liability can be assigned a

monetary value to be included in the financial model, while a risk involves too much uncertainty to

enable cost predictions to be made. Liabilities are further split into historic liabilities accruing to

Gecamines, historic liabilities that will be assumed by KML and KML’s current and future

liabilities, and these are defined later in this section.

25e.1 Legislation and compliance

The following comments should not be interpreted as an exhaustive legal review by legal specialists

but rather the identification of significant legislation affecting the environmental planning. Reference

1 It is noted that the SRK-prepared EIS for KOL is not the report which supports the licence.



should be made to Chapter 2 of the draft ESIA which presents a summary of DRC laws pertaining to

the management of the environment.

The primary piece of legislation governing mining environmental issues in the Democratic Republic

of the Congo (DRC) is the Mining (Code) Act 007/2002 of 11 July 2002. The Mining Code is

supported by the Mining Regulations (Decree No 038/2003 of 26 March 2003), which implement the

provisions of the Code including the environmental and social obligations relating to mining

projects. The Regulations contain a number of Annexures, with those pertinent to the environment

listed below:

Annex II: Financial surety for rehabilitation of the environment;

Annex III: Environmental Code of Conduct for Prospectors;

Annex VII: Mitigation and Rehabilitation Plan (MRP);

Annex VIII: Guidelines for preparing an MRP;

Annex IX: Guidelines for preparing an environmental impact study (EIS) and environmental

management plan for a project (EMPP);

Annex X: Closure measures;

Annex XI: Classification of mining wastes and their characteristics (standards for effluents);

Annex XII: Sensitive environments; and

Annex XIII: Method for the measurement of noise.

In addition to compliance with the in-country legislation, KML seeks to comply with the

requirements of the Equator Principles which are a set of guidelines adopted by Equator Principles

Finance Institutions (EPFI) including the International Finance Corporation (IFC). As a Category A

project the Mine is required to:

Produce a comprehensive Environmental and Social Impact Assessment (ESIA) and

Environmental and Social management Plan (ESMP) to a standard that meets the EPFI’s

satisfaction,

Ensure compliance with the IFC’s Performance Standards;

Ensure compliance with the in-country legislation, standards and regulations;

Undertake consultation in a structured and culturally appropriate manner; and

Establish a grievance mechanism.

Table 25e1 summaries key legislation and actions that have been taken to comply with this

legislation.



Table 25e.1 Summary of key legislation and relevant compliance

Legislation Compliance

Exploitation (mining) permit

In terms of Title V of Decree no 038/2003, the

company must apply for an Exploitation

(mining) Permit. In order to apply for such a

permit the company must be the title holder of

a valid Exploration Permit(s)

DCP holds exploitation permits as follows:

PE4960 – Kananga Mine

PE4961 – KOV Mine

PE4963 - Tilwezembe Mine

KCC holds exploitation permits as follows:

PE4948 - T17 Mine

PE525 – Kamoto Underground Mine

PE525 – Mashamba East Mine

Environmental Impact Study (EIS) and Environmental Management Plan (EMP)

The environmental obligations are set out in

Title XVIII of decree no 038/2003. With the

exception of temporary quarrying, any mining

operation requires an approved Environmental

Impact Study (EIS) and an Environmental

Management Plan (EMP).

Schedule IX (Contents of EIA and EMP) sets

out the contents of the EIS and the EMP and

provides detail regarding specific management

measures and standards that are required.

Environmental impact assessments have been undertaken for previous project

descriptions associated with this mine complex namely KOL and DCP. DCP has

been granted an environmental license by the Directorate for the Protection of the

Mining Environment (DPEM) in terms of the DCP EIS and EMP submitted in 2006.

KOL prepared and submitted an EIS to DPEM in 2006. Thereafter it was confirmed

that Gecamines as the titleholder had been granted an environmental license which

KOL is obliged to uphold.

In 2008 an ESIA was undertaken for the combined project description. A draft ESIA

report has been prepared but has not yet been released for public comment. In

addition the following management plans have been prepared by KML:

Framework Resettlement Action Plan

Stakeholder Engagement Plan

Social Development Plan

Community Health and Safety Plan

Occupational Health and Safety Plan

Labour and Human Resources Plan

Security Plan

Influx Management Strategy

Artisanal and Small Scale Miners Plan

Integrated Waste Management Plan

Emergency Response and Preparedness Plan

Rehabilitation and Closure Plan

Transportation and Traffic Management Plan

The ESMP and public disclosure meetings are to be completed in the first half of

2009.

Financial security obligation

Articles 410 to 414 set out the obligations in

respect of a financial security which must be

provided in terms of Article 294 of the Mining

Code. The financial security must provide for

the rehabilitation of the environment. The

funds of the financial security shall be made

available to the State and managed for the

purpose of rehabilitating the site.

The environmental costs as well as liabilities and closure costs for the project are

outlined in the relevant section of this report. It is understood that KML plans to set

aside the base percentage required by the Code on a yearly basis, starting in 2009.



Legislation Compliance

Schedule II (Financial security)

This Schedule deals with the financial security

required in respect of rehabilitation of the

environment. The amount of the financial

security shall be determined in accordance

with the approved Environmental Plan. The

financial security shall be maintained until the

Titleholder has been issued with an

Environmental Clearance Certificate. There are

a number of alternative means by which such

security can be provided. The total amount of

the financial security shall be paid in

accordance with a timetable taking into

consideration the duration of the mining

activities capped at 15 years. The first

instalment is only required in year 4 and the

quantum of the instalment increases with time

such that the major part is only required in the

last five years of the 15 year period.

The environmental costs as well as liabilities and closure costs for the project are

outlined in the relevant section of this report. It is understood that KML plan to set

aside the base percentage required by the Code on a yearly basis, starting in 2009.

25e.2 Environmental and social issues

25e.2.1 General

Following almost a century of mining and metallurgical activities, predominantly for copper and

cobalt minerals, the area around Kolwezi has numerous open pits, waste rock dumps, tailings dams,

concentrators and other mining-related infrastructure. This has led to extensive environmental

impacts. Rivers in the area, the Luilu and Musonoi, are contaminated by tailings disposed directly

into them. The failure of a large upstream tailings dam, the Poto-poto dam has also caused

significant contamination of the Luilu River. The worked areas remain largely unvegetated and in

combination with the numerous tailings dams and entrainment of dust from the unpaved roads,

contribute to high levels of dust in the area.

The Kolwezi area is characterised by wide spread poverty and a legacy of social challenges

including high unemployment, low levels of skills, dilapidated social infrastructure and services, and

low levels of functionality and delivery within all levels of government. The refurbishment and

operation of KOL, has already resulted in positive impacts on the socio-economic environment

although the impact of reducing the scale and/or cancelling these initiatives is as yet unidentified.

Reference should be made to the Draft ESIA (SRK Report 390781/1, November 2008) which

outlines in detail, the environmental and social baseline for the project area and describes the

anticipated impacts associated with the construction, operation and decommissioning of the project.

As a refurbishment and expansion project, operations, and construction, and the environmental and

social impact assessments thereof have been virtually concurrent. In some cases operational

decisions have been made before the appropriate environmental and social assessments have been

undertaken and the management of the impacts implemented. The refurbishment has however, in

some instances, led to visible improvements to the site as well as implementation of social

programmes and environmental efficiencies. During the course of 2008, organisational capacity was

ramped up and the development of various company directives, policies and management plans



ensued. The downscaling of operations in the last months of 2008 have however resulted in reduced

management capacity and concomitant reductions in the implementation of social and environmental

programmes.

25e.2.2 Outstanding information

There are significant gaps in the information that has been provided to SRK and a summary of

outstanding information is supplied in Table 25e.2, together with the assumptions that have been

made, based on available information, to allow the impact assessment to be completed. The

implication is that the assessment of the environmental impact of these items will still be required in

future.

Table 25e.2 Summary of project information gaps

Subject Project information gaps Approach in ESIA

Land rights DRC government approval of proposed concession boundaries It is assumed that proposed boundaries

will be approved

Waste rock

dumps

Waste rock dump designs for T17 Mine, Tilwezembe Mine, KOV

northern waste rock dump, Mashamba East Mine

Waste rock dump positions have been

assumed

Waste rock

analyses

Samples of some waste rock lithologies not available for

geochemical test work

Assess when the samples and analyses

are available

KOV dewatering

sludge

Chemical composition of pit water and sludge Assess when the samples and analyses

are available

KOV pit sludge Method of removal and disposal Assess when information is available

Processing Chemical balance

Mass balance

Water balance

Assess when information is available

Processing Composition of the solid residues and liquid fractions of all waste

streams (Kolwezi and Kamoto concentrators and Old and New

Luilu Refineries) and to where these will report

Assess qualitatively

Processing The reagents balance for the process Assess qualitatively

Processing Composition and volumes of return water from the tailings dam to

be used in the process


Processing Composition (and metal speciation) of the sludge arising from the

metal precipitation process in the cobalt circuit

Assume that the sludge reports to the

Luilu Effluent Ponds and assess

qualitatively

Processing Composition, volumes and disposal route of the “crud” from the

solvent extraction

Assume that the Crud reports to the

Luilu Effluent Ponds and assess

qualitatively

Processing Process streams that are likely to require blowdown or bleed during

operations and where the blowdown or bleed will be disposed of

and at what rate/frequency


Energy Heavy fuel oil plant design Assess when information is available



Subject Project information gaps Approach in ESIA

Route of all linear infrastructure and sub-stations for power

Associated

infrastructure

Infrastructure associated with subcontractors during construction:

sewage, water supply, water treatment, accommodation, solid

waste disposal


Associated

infrastructure

Infrastructure associated with operations: sewage, water supply,

water treatment, accommodation, solid waste disposal, medical

waste disposal


Transportation Off-site transportation during construction and operations: modes,

routes, volumes, frequency


Associated

infrastructure

Aggregate quarries Assess when information is available

Associated

infrastructure

Fuel storage Assess when information is available

Associated

infrastructure

Pipelines between the Luilu Metallurgical Plant, Kamoto

Concentrator and the Far West Tailings dam


Employment Number of people and contractors to be employed during

construction


Alternatives Alternatives as they relate to processing, mining, waste rock

management, associated infrastructure, energy use, water

Describe alternatives considered

Overall project Changes to the project as described in the ESIA document which

have resulted from KML’s November 2008 review of the project

and operational realities

Assess when changes are finalised

25e.2.3 Ground and surface water

Surface and groundwater specialist studies were undertaken for the Draft ESIA compiled by SRK

(SRK Report No. 390781/1, November 2008) to understand the hydrological and geohydrological

environment in the project area.

The Luilu and Musonoi rivers are already heavily impacted on by the historic mining activities in the

area. As a result of the numerous discharges, the quality of the rivers is not suitable for domestic or

agricultural use. However, groundwater quality in the area is surprisingly good and is the source

generally used by local communities. It is thought possible that the high rainfall provides sufficient

groundwater recharge to dilute the seepage from the numerous tailings dams. Discharge and seepage

from the project activities are likely to impact further on ground and surface water quality.

Water management in the Kolwezi area is complicated by the fact that it is a net water positive

environment, with significant seasonal fluctuations. It is thus envisaged that there will be significant

discharges from project infrastructure during the wet season, predominantly into the Luilu River.

The KOV dewatering system will also discharge into the Luilu throughout the life of the project.



The design of the Far West and Kamoto Interim tailings dams has allowed for settlement of

suspended solids. The details of the management measures for other aspects of the project have not

yet been provided to SRK.

KOV pit-lake dewatering

The descriptions of the sludge removal process were reviewed in the Africon report “Preliminary

functional design of decanting dams to support dewatering of KOV and Kamoto East pits, Kolwezi:

DRC (Africon Project No. 102996, October 2007) and in AGES AG-R-2008-11-24 KOL

Geohydrology Report Version 3 DRAFT and AGES AG-R-2008-10-02 KOV Dewatering DFS V2

Final.

Based on the current information available, it is understood from the description that the silty water

will be pumped from the pit at approximately 5 percent solids by mass. In ppm or mg/l terms a

concentration of 5 percent TSS is 50 000 mg/l (ppm). This water will be settled in the decant dams

where Africon has estimated that 70 percent of the solids will be removed, leaving 30 percent to

decant to the Musonoi River, i.e. 15 000 mg/l. However, the concentration of TSS in the discharge

is likely to be slightly higher than 15 000 mg/l, since some of the water sent to the decant dam will

be used to slurry the 70 percent of the solids that did settle successfully to the silt dam.

The DRC discharge limit for suspended solids is 100 mg/l and the IFC guideline value is 50 mg/l.

Clearly the proposed discharge is far in excess of either of these values. At the concentrations

currently envisaged, the impact on both water quality and flow in the Musonoi River is expected to

be highly significant.

The alternatives to the management measures currently planned for dewatering and desludging of the

KOV pit-lake must be reassessed to ensure compliance with DRC law.

Geochemical characterisation and management of tailings effluent

Samples of the current tailings streams at Kamoto Concentrator and the Luilu Refinery were

analysed and assessed as required by the DRC Mining Code. Test work was conducted on simulated

tailings from the oxide ore from KOV produced by Mintek in 2007. The test work on these samples

is unlikely to be fully representative of future ore to be mined at KOV, as the process simulation was

carried out on material obtained from stockpiles dating back to 1999, and thus it is uncertain to what

extent the metallurgical characteristics of the stockpiled ore had changed since being mined.

On the basis of the preliminary test work (TCLP and acid base accounting) the tailings currently

being generated at Kamoto Concentrator exceed the Low Risk limits for lead and copper in terms of

the DRC regulations. However, the regulations do not set out limits for High Risk tailings for

copper and lead.

Based on the geochemical test work carried out by SRK, the copper tailings from the Luilu Refinery

currently classifies as High Risk due to its acid generation potential and significant concentrations of

copper and cobalt. Process monitoring information provided by KOL indicates that the discharge

from the current operation at Luilu contains up to 600 tonnes of copper and 200 – 300 tonnes of

cobalt per month. It should be noted that the DRC effluent discharge regulation limits copper to 1.5

mg/l and the International Finance Corporation (IFC) guideline to 0.3 mg/l. Unless an appropriate

alkaline pH is maintained, a significant proportion of these metals will be in solution, resulting in

significant contamination of ground and surface water. The maintenance of an alkaline pH will

require effective management by KOL. It is understood that plans are being put in place to improve



the copper and cobalt recovery at the Luilu Refinery, however, no information has been provided to

SRK.

With a DRC High Risk classification, the Luilu copper tailings would be expected to require dam

design and operation to minimize potential for the leachate to seep or spill to ground or surface

water, not just in terms of lining configuration, but also in terms of overall water management to

prevent excess contaminated water release from the tailings and/or process system.

It must be noted that all the geochemical test work described above has been carried out on only two

sets of the three tailings streams (of which two are from Kamoto Concentrator and one from the

Luilu copper tailings), and the results will need verification on further samples, as will the acid

generation potential of the composite tailings from Kamoto Concentrator.

Based on the current information available, it is understood that an appropriate alkaline pH will be

maintained in the tailings stream, and that process improvements will reduce the concentrations of

copper and cobalt, SRK has recommended that it will not be necessary to line the storage facilities.

The maintenance of alkaline pH will require effective monitoring and management by KOL.

Geochemical characterisation and management of process effluents

As part of the copper and cobalt processing, a number of impurities need to be removed. It is

currently understood that these could include iron, manganese, aluminium, zinc, nickel and

selenium. At the concentrations likely to be produced by the refinery, it is possible that this process

effluent will be a hazardous waste. It is also possible that the semi-solid “crud” formed in the

solvent extraction process will be a hazardous waste. Although it is understood that simulations of

the process have been undertaken, this information has not been provided to SRK.

Based on the information currently available to SRK and the February 2006 Nikanor Feasibility

Study (SRK Report 356391, February 2006) which identified and evaluated sites for the disposal of

tailings and solid wastes from the Kamoto Concentrator and existing Luilu Refinery, it has been

recommended that hazardous waste streams from the Luilu Refinery must be disposed of in a

hazardous waste facility involving the construction and operation of a number of engineered lined

ponds at a site to the south-west of the Luilu Refinery.

The design proposed by SRK includes the construction and operation of a number of ponds with a

composite HDPE liner, leak detection system and stormwater management systems. The overall

pond layout suitable for the expected life of the project has allowed for an initial four ponds,

assuming that two ponds would be built immediately and thereafter further ponds as the tonnage

build up dictated. KML have made a commitment to undertake this work. However, although the

first pond has been constructed, no pump or pipeline system has been installed so the main waste

stream is being discharged to the Kamoto Interim Tailings Dam at a pH corrected to 8.0, but others

currently spill into the Luilu River via drains/canals.

25e.2.4 Relocation of residents of Musonoi Village

The development and expansion of the KOV pit and associated activities and infrastructure will

result in a combination of impacts including elevated noise levels, increased risk of traffic related

accidents, blasting and vibrations and increased levels of dust. While mitigation of individual

impacts may be possible to some extent, the combination of impacts will more appropriately be



resolved through the resettlement of the entire Musonoi Village, which has an estimated population

of up to 30 000 people.

The resettlement process is outlined in KML’s Framework Resettlement Action Plan (“FRAP”). The

resettlement process is expected to take a minimum of five years to implement. While no

requirement to resettle the village has been communicated to residents, it is expected that the

villagers are already aware and/or expect to be resettled. Based on the current information available,

it is understood that KML is committed to following due process and entering into suitable

negotiations with the affected community well in advance of any impending move.

25e.2.5 Relocation of residents of Ngonzo Village and other villages affected by FarWest Tailings Dam

The hamlets of Machine, Nana and Yenge are located within the footprint of the Far West Tailings

Dam, with several other hamlets, Kampemba, Kyapamoka, Masikini, Dique Kalemba, Ferme Jina

and the village of Ngonzo on the periphery. The construction of the Far West Tailings Dam and the

associated activities would require the resettlement of at minimum the hamlets of Machine, Nana

and Yenge and would result in a combination of impacts to the remaining hamlets and villages

including increased noise levels, decreased safety, loss of access to land, increased dust, loss of

social cohesion and reduced water quality and access to water. Resettlement of these villages will be

required to effectively address the combination of impacts. A resettlement site has already been

identified and resettlement actions, which are expected to take at least five to eight months to

implement, have been outlined in the KML FRAP. There has been some interaction with the

households of the directly affected villages (including some compensation payments), and with

leaders in Ngonzo village. No negotiations have commenced with the other affected villagers. It is

understood that some of the villagers have been requested not to build any further structures or to

plant any further crops. As a result, there are expectations around resettlement, especially in Ngonzo

and the directly affected communities. To minimise uncertainty, to address any hardship related to

reduced farming activity, and to align all villages involved, the resettlement process should not be

delayed.

25e.2.6 Air pollution

According to dust fallout modelling the baseline dust fallout conditions are high with large parts of

the concession experiencing peak concentrations above the South African Industrial Guidelines2 and

areas along roads and close to tailings dams and waste rock dumps experiencing double this and

exceeding the South African Alert Guidelines. Modelling of operational conditions indicates that

vehicle-entrained dust and windblown dust from the tailings dams are overwhelmingly the

predominant sources of nuisance and respirable dust. Improvement of the roads, through dust

suppression and surfacing and progressive vegetation of the tailings dams will significantly improve

this situation. Provisions have been made for this in the environmental capital and operating costs.

2 Used in the absence of relevant DRC guidelines



25e.2.7 Social issues

Extensive stakeholder engagement has taken place as part of the environmental and social

assessments, and further disclosure meetings are planned for early 2009. Social assessments have

also been conducted to support the identification and description of social impacts. Issues that have

been raised by the communities relate largely to the perceived benefits that the mine should bring to

the communities including job creation, improvements to social infrastructure and services.

A combination of positive and negative social changes is likely. In some cases these are already

occurring as a result of this refurbishment and expansion project. The key drivers of change and the

impacts that will result from these include: foreign investment in the Kolwezi District and the

associated stimulation of the local economy (through associated multiplier effects), the construction

and operations of the mine and associated infrastructure which will change the movement of people

and displace people in some instances, the creation of jobs as well as the expectation for local

employment and investment in community development, the disruptions to existing artisanal and

small scale mining operations and the consequent economic displacement, improvements to local

infrastructure such as roads and clinics as well as the increased pressure on infrastructure, capacity

building, the payment of taxes and royalties and the implementation of KML’s social policies and

programmes.

The following management plans which were prepared as part of the ESIA outline the actions that

KML would need to take to comply with IFC requirements:

Stakeholder engagement plan;

Social development plan;

Influx management plan;

Artisanal and small scale miners plan;

Community health and safety plan;

Framework resettlement action plan; and

Security plan.

While the Kolwezi Concentrator and Tilwezembe operations were expected to be short term

operations, their recent closure and the expected associated retrenchment of staff is likely to reduce

the anticipated positive impacts associated with job creation to some extent. In addition, the

temporary suspension of social programmes in the communities may have further negative impacts

with respect to development of trust and goodwill between the project and the community, and may

impact on the mine’s social licence to operate.

25e.2.8 Radiation

A preliminary radiological assessment was undertaken in 2008. This study was undertaken to

respond to a decree from the Department for the Protection of the Mining Environment (DPEM) and

investigate preliminary indications of extensive distribution of radioactive material across the project

site. The surveys and sampling performed during the radiological assessment were sufficient to

demonstrate that there is not a widespread hazard from radioactive materials, although there are

certain locations that merit the implementation of controls and mitigation measures. The areas



which have been identified where risk mitigation measures are deemed warranted are the disused

uranium ore storage locations adjacent to Kingamyambo Tailings Dam and north of the KOV Pit,

Katapula-Luilu tailings area, Kolwezi concentrator ore concentrate storage area, Sulphide Tailings

Dam, and the Luilu River. While the uranium storage locations are not a KML liability, they are

reported for purposes of transparency. Management measures, including the communication of the

results of the study to Gecamines and other mines, and the implementation of a radiation safety

programme are included in the capital and operating costs.

It must be noted that the radiological assessment did not evaluate ground water risks and only

conducted a limited sampling of surface water sources, nor has it evaluated sources of airborne

contamination.

25e.3 Closure

25e.3.1 Closure planning

A Rehabilitation and Closure Plan was prepared in 2008 to provide guidance and commitments with

respect to the environmental and social elements of decommissioning and closure. The plan

addresses the management of biophysical and social impacts at closure. It is understood that KML

has committed to implementing operational efficiencies throughout the operations phase to

progressively manage and reduce possible environmental impacts as far as practicable, thereby

reducing the liability at closure.

With respect to the socio-economic impacts of closure, management commitments have been

outlined with respect to management of direct impacts on mine employees and their families, loss of

economic benefits to local businesses and the economy, loss of support to the local government and

other government levels with respect to taxes and royalty incomes, increased pressure on

maintenance of infrastructure and services, loss of community benefits in the form of mine-funded

community development programmes, loss of access to environmental resources as a result of

residual impacts to groundwater, surface water and soils. KML will need to reschedule the

management measures contained within the Rehabilitation and Closure Plan with respect to the

closure of the Kolwezi Concentrator and Tilwezembe Mine once a decision has been taken on the

future of these facilities.

In terms of Article 100 of the DRC Mining Environmental Regulations, all buildings and surface

infrastructure will have to be demolished on closure unless a representative of the local communities

submits a written request to the Minister and demonstrates that such buildings and infrastructure are

necessary for the socio-economic development of the area.

The closure costs have been prepared on the basis that:

Land will be returned to its pre-mining state, or where this is not practicable or beneficial will be

made safe for humans and animals;

All infrastructure other than residue deposits and that infrastructure returned to Gecamines will

be demolished and appropriately removed, disposed of, buried or filled in;

Where it is inappropriate to leave roads, pipelines and boreholes intact for community use, these

will be ripped up and rehabilitated;

Waste rock dumps will be reprofiled using cut and fill techniques to a maximum slope angle of

18 degrees and revegetate;



Tailings dams surfaces will be revegetated;

Engineered waste disposal sites will be capped;

Waste will either be removed to engineered waste facilities or covered to minimise

environmental exposure; and

Backfilling of pits is not practicable, however access control berms will be constructed from

waste rock to limit humans and livestock gaining entry.

25e.3.2 Allocation of Closure Costs

The closure costs provided for the Kolwezi Concentrator and Tilwezembe Mine, originally

scheduled for Year 5, remain relevant as these were based on operations in 2008 and will need to be

rescheduled to Year 1 to address the recent closure of these two facilities.

The assessment evaluated the environmental obligations in terms of current liabilities, which are

current liabilities that exist at 1 April 2008 and are estimated at USD 26,2 million, and closure costs,

which can be reasonably expected to be payable at the end of LoM. The closure costs are liabilities

which KOL are likely to incur during operations and includes new infrastructure that they will

construct to support their activities. Refer to Table 25.e3 for further details.

Table 25e.3 Environmental Closure Costs

Area Unit Total

Luilu Plant (USDm) 16,7

Kolwezi Concentrator (USDm) 3,8

Kamoto Concentrator (USDm) 8,5

SKM (USDm) 0,1

Kamoto Underground (USDm) 1,1

KOV Pit (USDm) 33,9

Tilwezembe (USDm) 1,5

Mupine Pit (USDm) -

Pit T17 (USDm) 5,1

Kamoto East (USDm) 6,9

Far West (USDm) 55,3

Subtotal (USDm) 132,9

Contingency @ 25% (USDm) 33,2

ECMP @ 10% (USDm) 13,3

Owners cost @ 5% (USDm) 6,6

Subtotal (USDm) 53,2

Total (USDm) 186,1

Based on DRC legislation annual expenditure for rehabilitation is planned with the aim of reducing

the final costs at the end of LoM (Annexure II from the DRC Mining Code indicates that full

payments must be made by year 15 of the operation). However, SRK’s experience is that a payment

schedule longer than 15 years is acceptable. Accordingly, the Tilwezembe Mine portion and the

Kolwezi Concentrator portion closure costs amounting to USD 5,3 million are allocated in year 2009

as mining operations stopped in November 2008 and the remainder of USD 180,1 million is

allocated as indicated in Table 25.e4.



Table 25e.4 Allocation of Closure Costs

Allocation of Costs

Yr 1 to 18 (USD 2,1 million) Yr 19 to Yr 21 (USD 18,1 million) Yr 22 and Yr 23 (USD 46,7 million)

25e.4 Material risks and potential opportunities to reduce liabilities

25e.4.1 Risks

Non-compliance with commitments in the existing authorised EMPs. Non compliances relate to

water management and some social commitments. Clause 12.2 of the joint venture agreement

(JVA) between DCP and Gecamines also includes an undertaking to comply with the

requirements of the Mining Code and internationally acceptable standards – non-compliance

with the commitments made in the EMP would be viewed as non compliance with this clause.

Influx of job seekers into the Kolwezi area with associated negative risks where migrants are not

assimilated, placing stress on services and disrupting the status quo in communities. The Influx

Management Strategy addresses actions to address this risk to international standards.

Artisanal and small scale mining. At present KOL is cohabiting with artisanal and small scale

miners. This situation will change when small scale mining is displaced by commercial mining,

with possible social and economic risks. The Artisanal Mining Strategy and Social Development

Plan outline initiatives that would address this to the level of international standards.

Resettlement of villages could be a risk to the project if not undertaken to internationally

acceptable levels. The company’s resettlement practices are being reviewed and the Framework

Resettlement Action Plan outline actions that would meet international requirements.

Social licence to operate is influenced by the management of impacts, management of

community expectations, management of community, health and safety programmes, social

development programmes, transparency and communication. The recent abrupt suspension of

social programmes in the communities may contribute towards increasing this risk and it is

further understood that retrenchments are shortly to be announced which is likely to result in

further disruptions and social unrest. A communication strategy upholding internationally

accepted principles for stakeholder engagement is necessary as transparency will be required to

manage this risk.

Operational issues with respect to tailings management. The maintenance of an appropriately

alkaline pH will require careful management by KOL to ensure that potentially toxic metals

remain as precipitates and prevent generation of acidity and thus significant impacts on ground

and surface water quality.

Management of process effluents. Potentially hazardous waste streams will need to be disposed

of in a hazardous waste facility until it they have been adequately characterised. This is to

ensure impacts on ground and surface water can be adequately managed if the material is

disposed of to the unlined tailings facilities.

KOV pit-lake dewatering. The measures proposed to manage the dewatering of the KOV pit-

lake currently will result in non-compliance with the DRC effluent limit and IFC guideline value

for suspended solids and cause significant impacts in the Musonoi River. Alternative

management measures will need to be investigated to ensure compliance with DRC regulations

and prevent major contamination of the Musonoi River.



It is understood that Gecamines has provided permission for the use of the GH pit for disposal of

tailings from the Kolwezi concentrator, however it is unclear if Gecamines has obtained

environmental permits from DPEM. SRK is also not aware that an environmental permit has yet

been sought or obtained for the disposal of tailings from the Kamoto Concentrator and Luilu

Refinery into the Mupine Pit.

Concession boundaries for the project have not been finalised with the DRC government and

part of the Far West Tailings Dam footprint is not included in the existing concession.

KML may be exposed to third party risks associated with other mining operations in the Kolwezi

area. Examples include:

o the potential for future dewatering of the DIMA Pits upstream of the concession

affecting the groundwater regime through impacts on water quality and/or aquifer

drawdown;

o social tensions that may escalate as a result of the combination of existing social

instability, recruitment of foreign skilled workers and limited increase in employment of

local labour or worse, the retrenchment of local labour;

o lack of uniformity by the various mines in delivery on social goals, environmental

management and pollution prevention measures;

o increased pressure on infrastructure and services specifically medical services, roads

infrastructure, sanitation and housing.

It is understood that recent reductions in professional staffing have taken place within the KML

organisation. SRK is concerned that this places KML at risk with respect to the necessary

organisational capacity to implement the social and environmental management programmes to

Equator Principle standards.

25e.4.2 Opportunities

It is understood that the First Quantum Minerals (FQM) project adjacent to the KOL operations

requires additional water supply. It may be possible to negotiate with FQM for it to pay part of

the costs of the pipeline and pumping of the dewatering water to the Luilu River in return for

access to the water.

The Far West Tailings Dam, Luilu Refinery and Kamoto Concentrator will generate a number of

effluent streams that are expected to cause deterioration in the quality of water in the Luilu. The

discharge of the expected high volumes of dewatering water could provide dilution of these

effluents and thus mitigate the impact of the discharges.



25f Taxes and Key Business OperatingParametersThe operating parameters, which govern the royalty, tax, rehabilitation, depreciation and import duty

rates applicable to the business, are provided in Table 25f.1:

Table 25f.1 Economic Analysis: Key Business Operating Parameters

Description Explanation Rate

DRC royalty (deduct from turnover) % of revenue, less selling expenses 2.0%

Gecamines royalty (deduct from turnover) % of revenue, less selling expenses and debt redemption 2.5 %

DRC corporate tax rate 30%

Depreciation (year 1) 60%

Depreciation (year 2 to 5) 10%

Import duties Charged on certain imported items 3%

A royalty of 2% will be payable to the DRC, while a royalty of 2,5% will be payable to Gecamines.

The DRC corporate tax rate of 30% has been applied. According to DRC legislation, Taxation can be

offset against capital and deferred. All capital expenditure has been depreciated at 60% in the first

year and 10% for each year thereafter. An import duty of 3% is applied to imported goods, if

applicable.

In addition, USD120,5 million has been budgeted and included in the cash flow model up to 2014

for payment to Gecamines for its Pas de porte payments.



25g Capital and Operating Cost Estimates

25g.1 Capital Cost Estimates

The major capital cost items are provided in Table 25g.2. These include:

Kamoto Mine: capital expendure includes provisions for purchase of the mining fleet

requirement required by the LoM plan, development costs to access new mining areas and

USD10 million in 2009 and 2010 for ventilation infrastructure.

KOV Mine: capital expendure includes provisions for purchase of the mining fleet requirement

required by the LoM plan. There is no provision for equipment in the first five years as all of the

equipment required during this period had been purchased by 31 December 2008 The pre-

stripping required to access the ore body has been capitalised in 2009 and 2010.

Mashamba East Mine: capital expendure includes provisions for purchase of the mining fleet

required in the LoM plan.

Processing Capital Expenditure: the capital expendure includes provisions for the

development of the process plants as described in Item 25b and set out in details in Table 25g.1.

As noted elsewhere in this report, further study is recommended to improve the accuracy of the

cost estimates, as these have been factorised from the 2008 Study and care should be exercised

in their use.

Table 25g.1 Summary of Processing Capital Expenditure

2009 2010 2011 2012 2013 2014 Total

Unit 1 2 3 4 5 6

Phase 2 (USDm) 38,0 0,0 0,0 0,0 0,0 0,0 38,0

Phase 3 (USDm) 1,4 75,5 63,0 0,0 0,0 0,0 140,0

Phase 4 (USDm) 0,0 0,9 45,7 35,8 0,0 0,0 82,4

Phase 5- MergedModule1 and 2

(USDm) 0,0 38,5 435,9 455,8 60,7 0,0 990,8

Phase 5- Module 3 (USDm) 0,0 0,0 0,0 2,3 121,5 101,4 225,2

Total (USDm) 39,4 114,9 544,6 493,8 182,1 101,4 1476,3

Effluent Ponds and Tailings: this refers to the capital expenditure provisions for the

development of the ponds and tailings facilities described in Item 20.5.

Environmental: this includes Musonoi Village, Far West Tailings Dam and ad-hoc

resettlement; artisanal mining; stakeholder engagement; jobs and economic opportunities; tarring

of roads (to reduce dust and for risks associated with road safety); dust-monitoring

equipment;equipment for sulphur-dioxide-emission reductions / monitoring; surface-water

management (containment and management);general and hazardous waste (trenches and

building); ad-hoc equipment for groundwater (equipment);water settlement facilities (for

suspended solids); radiation monitoring and survey equipment;emergency response equipment;

and vehicles and equipment.

Dewatering: this refers to the costs associated with the dewatering of the KOV Mine amd

Mashamba East Mine, which includes dewatering boreholes, borehole pumps, monitoring,

design and control systems, software, geophysical surveys, water sampling and analysis, the

regular updating of the dewatering model and the purchase and operation of equipment.

Power: this refers to the capital expenditure for the provision of power as described in Item

20.3.2.



General Infrastructure: this refers to unallocated infrastructure spending of a general nature

that is required to sustain mining operation in the DRC and cancellation costs (USD62 million)

arising out of the changes to the plant requirements and revised business strategy.

Table 25g.2 LoM Investment Capital Expenditure (2009-2018)

2009 2010 2011 2012 2013 2014 2015 2016 2017 2018

Unit 1 2 3 4 5 6 7 8 9 10

Mining

Kamoto underground (USDm) 24,6 20,1 7,9 5,6 2,9 14,8 2,8 9,0 0,9 14,8

KOV mobile mining fleet (USDm) 0,0 0,0 0,0 0,0 0,0 168,0 0,0 0,0 0,0 0,0

KOV pre-stripping (USDm) 10,1 54,9 0,0 0,0 0,0 0,0 0,0 0,0 0,0 0,0

Mashamba East (USDm) 0,0 0,0 0,0 0,0 0,0 0,0 0,0 0,0 0,0 32,7

Subtotal (USDm) 34,7 75,0 7,9 5,6 2,9 182,8 2,8 9,0 0,9 47,5

Processing

All phases and modules (USDm) 39,4 114,9 544,6 493,8 182,1 101,4 0,0 0,0 0,0 0,0

Subtotal (USDm) 39,4 114,9 544,6 493,8 182,1 101,4 0,0 0,0 0,0 0,0

Other Cost Centres

Tailings (USDm) 18,5 30,9 5,7 37,1 37,1 24,9 24,9 24,9 24,9 24,9

Environmental and socialcosts

(USDm) 24,4 11,2 10,9 13,2 31,9 33,1 1,8 4,1 1,8 1,9

Dewatering (USDm) 33,2 7,9 0,0 0,0 0,0 0,0 0,0 38,7 3,0 0,0

Power (USDm) 3,2 0,3 0,0 8,2 52,7 16,3 0,0 0,0 0,0 0,0

General capitalexpenditure

(USDm) 81,5 15,3 37,1 17,9 3,5 9,9 0,0 0,0 9,5 1,6

Subtotal (USDm) 160,8 65,6 53,7 76,4 125,2 84,2 26,7 67,6 39,2 28,5

Total capitalexpenditure

(USDm) 234,9 255,5 606,2 575,9 310,2 368,5 29,5 76,6 40,1 76,0

25g.2 Operating Cost Estimates

The major operating cost items are provided in Table 25g.3. These include:

Open Pit and Underground Mining: this includes the mining cash costs from:

o T17 Mine: the cost estimate is based on a mining contractor EGMF and is forecast at

USD8,50/bcm for mining and USD1,70/t ore mined for haulage.

o Kamoto Mine: the operating cost was estimated from first principles as an owner

operation. The cost estimate is based on the mining method employed at each zone as

decribed in Item 19. Cost data for the various mining methods was gathered from

various sources, including InfoMine USA and similar projects undertaken by SRK. The

InfoMine data were the most up-to-date (July 2008) but are related to North American

mines. The costs were compared to similar sized mines in Africa and South America and

adjusted to account for the higher fuel, steel, power and shipping costs. The weighted

average cost applied over the LoM is USD33,83/t ore mined.

o KOV Mine: the operating cost was estimated from first principles as an owner

operation. This was because the mining contract was under internal review and not

finalized in time for this study.

The mining consumables cost was estimated using the estimated costs for: operating the

mining and ancillary equipment using the hourly consumption rates for fuel and

lubricants; drilling and blasting based on the requirements of the mine plan; manning of

the mine operation together with the mine workshop and mine technical office; similar

sized equipment as that which would be required for KOV Mine; drilling using



assumptions on the drill steel life for drilling; explosives delivered to the mine and

provided by AEL. The weighted average cost applied over the LoM is USD2,35/t ore

mined and USD1,88/t waste mined.

o Mashamba East Mine: The operating cost was estimated from first principles as an

owner operation. The mining consumables estimates was carried out using the estimated

mining and ancillary equipment, hourly fuel consumption and provision for consumed

lubricants, and drilling and blasting requirement. Manning of the mine operation

together with the mine workshop and mine technical office was also estimated. Cost

information for similar size equipment was gathered from other projects and was used

for the mine operating cost estimation. The weighted average cost applied over the LoM

is USD5,06/t ore mined (estimate includes waste mining).

Kamoto Concentrator Costs: this includes plant costs for reagents, consumables and power

and is based on fixed costs of USD4 million per annum and a variable cost of USD3,03/t ore

feed for the oxide circuit and USD7,88/t ore feed for the sulphide circuit.

Luilu Metallurgical Plant: this includes plant costs for reagents, consumables and power and is

based on fixed costs of USD9 million per annum and a variable cost of USD0,46/lb Cu from

2009 to 2016) and USD0,41/lb Cu from 2016 onwards.

WOL/SX/EW Refinery Project: USD0,34/lb for the Cu circuit and USD3,40/lb for the Co

circuit.

General and Administrative Costs: this refers to head office and other centralised costs.

Freight, Insurance and Sales Costs: It is understood that all finished products (copper, cobalt)

will be transported through Durban (the FOB point) either to Europe or to the Far East. Finished

product CIF Rotterdam costs applied for Cu: USD675/t and Co: USD828/t.

Table 25g.3 LoM Operating Costs (2009-2018)

2009 2010 2011 2012 2013 2014 2015 2016 2017 2018

Unit 1 2 3 4 5 6 7 8 9 10

OPERATING COSTS

Open Pit andUnderground Mining

(USDm) (136) (93) (129) (145) (197) (185) (166) (174) (165) (221)

Old KTC and LuiluMetallurgical Plant

(USDm) (64) (82) (82) (162) (173) (141) (134) (141) (136) (132)

Kamoto ConcentratorCosts

(USDm) (19) (23) (20) (22) (25) (16) (17) (15) (16) (18)

SX/EW Refinery (USDm) - - - (45) (63) (205) (148) (258) (226) (197)

Tailings (USDm) (1) (1) (1) (1) (1) (16) (16) (17) (16) (16)

Total Operating Costs (USDm) (219) (199) (231) (375) (458) (563) (481) (604) (559) (583)

General andAdministrative Costs

(USDm) (150) (70) (80) (105) (105) (105) (90.5) (80) (80) (80)

25h Economic AnalysisThis section presents a cash flow model (Tables 25h.1 and 25h.2) and sensitivities for discount rates

(Table 25h.3) metal prices and grade (Table 25h.4); capital cost (Table 25h.5) and operating costs

(Table 25h.6). The Net Present Values (“NPVs”) should be considered only in relation to the risks

mentioned in Item 20 of this report.



Table 25h.1 Discounted Cash Flow Model (2009-2023)

2009 2010 2011 2012 2013 2014 2015 2016 2017 2018 2019 2020 2021 2022 2023

January 2009 Money Terms Prices Unit 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15

Revenue (USDm) 424 484 444 754 735 1176 1270 1727 1720 1636 1909 2050 1635 1645 1607

Less: Freight, Insurance and Sales

Costs(USDm) (42) (53) (54) (113) (123) (179) (172) (221) (223) (199) (231) (250) (189) (216) (223)

Net Revenue (USDm) 383 431 390 641 612 996 1098 1506 1497 1436 1679 1800 1447 1429 1384

Expenses

Operating Costs (USDm) (219) (199) (231) (375) (458) (563) (481) (604) (559) (583) (674) (691) (605) (685) (647)

Other Costs (USDm) (175) (95) (112) (151) (149) (166) (162) (170) (170) (175) (188) (199) (177) (187) (176)

Net change in working capital (USDm) 14 (0) (0) (13) (2) (19) 0 (20) 2 2 (13) (4) 15 (3) 4

Total Expenses (USDm) (381) (293) (343) (539) (609) (748) (643) (794) (728) (756) (874) (894) (767) (875) (820)

Taxation (USDm) - - - - - - - - - - (137) (208) (216) (174) (148)

Capital Expenditure (USDm) (236) (259) (609) (579) (313) (370) (33) (80) (43) (79) (183) (264) (57) (20) (56)

Net Free Cash (USDm) (234) (121) (562) (477) (311) (122) 422 632 726 601 485 435 406 360 360



Table 25h.2 Discounted Cash Flow Model (2024-2038)

2024 2025 2026 2027 2028 2029 2030 2031 2032 2033 2034 2035 2036 2037 2038

January 2009 Money Terms Prices Unit 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30

Revenue (USDm) 1584 1640 1522 1189 1119 1069 1038 840 - - - - - - -

Less: Freight, Insurance and Sales

Costs(USDm) (216) (229) (201) (162) (156) (155) (152) (126) - - - - - - -

Net Revenue (USDm) 1368 1411 1321 1027 963 914 886 714 - - - - - - -

Expenses

Operating Costs (USDm) (651) (689) (623) (439) (396) (335) (298) (226) - - - - - - -

Other Costs (USDm) (184) (177) (186) (176) (174) (171) (188) (209) - - - - - - -

Net change in working capital (USDm) 0 (3) 5 18 5 5 4 8 31 - - - - - -

Total Expenses (USDm) (835) (868) (803) (597) (565) (502) (483) (427) 31 - - - - - -

Taxation (USDm) (144) (111) (153) (144) (115) (110) (121) (119) (84) - - - - - -

Capital Expenditure (USDm) (214) (10) (10) (10) (9) (6) (6) - - - - - - - -

Net Free Cash (USDm) 176 421 354 275 274 296 276 168 (53) - - - - - -



Table 25h.3 Discount Rate Sensitivity

Discount Factor NPV

(%) (USDm)

8,00% 1027

10,00% 624

12,00% 324

14,00% 99

16,00% (70)

18,00% (197)

20,00% (292)

Table 25h.4 Revenue and Grade Sensitivity

SensitisedFactor SensitivityRange (@14% discount rate)


Grade (Cu %) 3,89% 4,12% 4,35% 4,50%* 5,03% 5,49% 5,95%

Grade (Co %) 0,37% 0,39% 0,41% 0,44%* 0,48% 0,52% 0,57%


Revenue / Grade (10) 27 63 99 171 242 312

* Reserve Grade

Table 25h.5 Revenue and Capital Cost Sensitivity


(USDm) -30% -20% -10% 0% 10% 20% 30%

-15% 521 558 594 630 702 773 842

Total -10% 345 382 418 454 526 597 667

Capital -5% 168 205 241 278 350 421 490

Costs 0% (10) 27 63 99 171 242 312

Sensitivity 10% (370) (333) (297) (261) (189) (118) (48)

Range 20% (738) (701) (665) (629) (557) (487) (417)

30% (1135) (1087) (1047) (1009) (937) (867) (797)

Table 25h.6 Revenue and Operating Cost Sensitivity


(USDm) -30% -20% -10% 0% 10% 20% 30%

-15% 248 284 321 357 429 499 569

Total -10% 162 199 235 271 343 414 484

Operating -5% 76 113 149 185 257 328 398

Costs 0% (10) 27 63 99 171 242 312

Sensitivity 10% (182) (145) (109) (72) (1) 70 140

Range 20% (353) (317) (280) (244) (173) (102) (32)

30% (525) (489) (452) (416) (344) (274) (204)



Table 25h.7 Revenue Sensitivity using 17 March 2009 LME Forecast Contract Pricing

SensitisedFactor SensitivityRange (@14% discount rate)



Revenue / Grade 143 179 215 251 322 392 461

Table 25h.7 reflects the NPVs using the LME 3-Month copper price data at 17 March 2009 with

applicable contract prices up to 2018.

25h PaybackThe payback for the initial six years of capital invested, USD2,35 billion, excluding interest is 10

years. The payback period for the invested capital of USD2,13 billion at an interest rate of 7,52%

(the sum of the US Federal Reserve five-year interest rate swap of 2,52% and a 5% risk premium) is

18 years.

25i Mine LifeBased on the assumptions at 31 December 2008, KML has Proven and Probable Mineral Reserves of

139,8 Mt of ore with a grade of 4,93% Cu and 0,38% Co, which support the LoM plans provided in

Item 25a.



Glossary of Terms, Abbreviations, Units andChemical Elements

Glossary of Terms

Aeolian erosion, transport, and deposition of material due to theaction of wind at or near the earth’s surface

Arenaceous term describing sedimentary rocks with a modal grain sizein the sand fraction

Argillaceous term describing sedimentary rocks with a modal grain sizein the silt fraction

Assay the chemical analysis of mineral samples to determine themetal content

Assaying the chemical analysis of mineral samples to determine themetal content

Basal conglomerate a conglomerate formed at the earliest portion of astratigraphical unit

Basement complex the widespread association of igneous and metamorphicrocks which are covered uncomfortably byunmetamorphosed sediments

Basinal a basin like depression that may be erosional or structuralin origin

Bateman or BatemanEngineering

means Bateman Projects Limited.

Biotite/titic trioctahedral

Breche Heterogene orHeterogeneous Breccia (BH

this breccia is composed of angular and sometimes wellrounded rock fragments of all the various rock types ofthe Roan Group

Breche RAT or BrecciatedRAT (B RAT

a reddish-pink brecciated rock with calcite and silicaveinlets and is at times well mineralised with specularhaematite, occurring as veinlets

Calcaire a Minerais Noirs orCalcareous Unit with BlackMinerals (CMN

a slightly banded and laminated light grey to greysilicified dolomite mineralised with black oxide of iron,manganese and cobalt

Capital expenditure all other expenditures not classified as operating costs

Concentrate a metal-rich product resulting from a mineral enrichmentprocess such as gravity concentration or flotation, inwhich most of the desired mineral has been separatedfrom the waste material in the ore.

Crushing initial process of reducing ore particle size to render itmore amenable for further processing

Cut-off grade the grade of mineralized rock which determines as towhether or not it is economic to



Decline a surface or sub-surface excavation in the form of a tunnelwhich is developed from the uppermost point downwards

Detrital a term applied to any particles of minerals, or, morerarely, rocks, which have derived from pre-existing rockby processes of weathering and erosion

Dilution waste which is unavoidably mined with ore

Dilution waste which is unavoidably mined with ore

Dip angle of inclination of a geological feature/rock from thehorizontal

Dipeta (R3) greyish to dark red or brown stratified shales andmicaceous schist

Dolomie Stratifie orStratified Dolomite (D Strat)

this is a well bedded to laminated, argillaceous dolomite,which forms the base of the traditional “Lower Ore Zone”in Gecamines’ nomenclature

Dolomites the name of a sedimentary carbonate rock and a mineral,both composed of calcium magnesium carbonate

Drill-hole method of sampling rock that has not been exposed

Effective Date effective date of Engineering study

Fault the surface of a fracture along which movement hasoccurred

Feldspar/s the most important single group of rock forming silicateminerals

Ferricrete a conlgomerate consisting of surfical sand and gravelcemented into a hard mass by iron oxide derived fromoxidation of percolating solutions of iron salts

Ferruginised containing iron

Filtration process of separating solid material from a liquid

Flotation the process by which the surface chemistry of the desiredmineral particles is chemically modified such that theypreferentially attach themselves to bubbles and float to thepulp surface in specially designed machines. the gangueor waste minerals are chemically depressed and do notfloat, thus allowing the valuable minerals to beconcentrated and separated from the undesired material.

Footwall the underlying side of an ore body or stope

Geochronological the measurement of time intervals on a geological scale

Grade the measure of concentration of copper or cobalt withinmineralised rock

Hangingwall the overlying side of an ore body or slope

Haulage a horizontal underground excavation which is used totransport mined ore

Hydrogeology a science that deals with sub-surface water and withrelated geologic aspects of surface water



Indicated Mineral Resource that part of a mineral resource for which tonnage,densities, shape, physical characteristics, grade andmineral content can be estimated with a reasonable levelof confidence. it is based on exploration, sampling andtesting information gathered through appropriatetechniques from locations such as outcrops, trenches, pits,workings and drill holes. the locations are too widely orinappropriately spaced to confirm geological and/or gradecontinuity but are spaced closely enough for continuity tobe assumed

Inferred Mineral Resource that part of a mineral resource for which tonnage, gradeand mineral content can be estimated with a low level ofconfidence. it is inferred from geological evidence andassumed but not verified geological and/or gradecontinuity. it is based on information gathered throughappropriate techniques from locations such as outcrops,trenches, pits, workings and drill holes which may belimited or of uncertain quality and reliability

Intrusives a body of igneous rock which has forced itself onto pre-existing rocks, either along some definite structuralfeature or by defamation or cross-cutting of the invadedrocks

Kamoto Concentrator Kamoto, an operating concentrator

Kamoto Mine Kamoto Underground Mine

Kananga Mine Kananga Mine

Kolwezi Concentrator Kolwezi, an operating concentrator

KOV Mine KOV Mine

Laterite red residual soil developed in humid, tropical andsubtropical regions of good drainage

Lithology/ical geological description pertaining to different rock types

LoM plans life-of-mine plans

Luilu Metallurgical Plant Luilu, an operating metallurgical plant

Mashamba East Mine Mashamba East Mine

Material Assets operations/projects and associated infrastructure

Measured Mineral Resource that part of a mineral resource for which tonnage,densities, shape, physical characteristics, grade andmineral content can be estimated with a high level ofconfidence. it is based on detailed and reliableexploration, sampling and testing information gatheredthrough appropriate techniques from locations such asoutcrops, trenches, pits, workings and drill holes. thelocations are spaced closely enough to confirm geologicaland grade continuity

Metasediments/tory metamorphosed sedimentary rock

Mica/micaceous layer-lattice minerals of the three-layer type, and may bedivided into the dioctahedral muscovite group and thetrioctahedral phlogopite-biotite group



Milling a general term used to describe the process in which theore is crushed and ground and subjected to physical orchemical treatment to extract the valuable metals to aconcentrate or finished product.

Mineral / mining lease a lease area for which mineral rights are held

Mineral Reserve the economically mineable material derived from ameasured and/or indicated mineral resource. it is inclusiveof diluting materials and allows for losses that may occurwhen the material is mined. appropriate assessments,which may include feasibility studies, have been carriedout, including consideration of, and modification by,realistically assumed mining, metallurgical, economic,marketing, legal, environmental, social and governmentalfactors. these assessments demonstrate at the time ofreporting that extraction is reasonably justified. mineralreserves are sub-divided in order of increasing confidenceinto probable mineral reserves and proved mineral reserve

Mineral Resource a concentration [or occurrence] of material of economicinterest in or on the earth’s crust in such form, quality andquantity that there are reasonable and realistic prospectsfor eventual economic extraction. the location, quantity,grade, continuity and other geological characteristics of amineral resource are known, estimated from specificgeological evidence and knowledge, or interpreted from awell constrained and portrayed geological model. mineralresources are sub-divided in order of increasingconfidence, in respect of geoscientific evidence, intoinferred, indicated and measured categories

Mineral rights a right or any share therein acquired, in terms of theminerals act to any right to dig or mine

Mwashya (R4) altered stratified greyish siliceous dolomitic rock withoolitic horizons and a few bands of light yellow talcoseschist.

On-going capital capital estimates of a routine nature which are necessaryfor sustaining operations (also known as sustainingcapital)

Orogeny an orogeny is a period of mountain building leading to theintensely deformed belts which constitute mountainranges.

Probable Mineral Reserve the economically mineable material derived from ameasured and/or indicated mineral resource. It isestimated with a lower level of confidence than a provedmineral reserve. It is inclusive of diluting materials andallows for losses that may occur when the material ismined. Appropriate assessments, which may includefeasibility studies, have been carried out, and includingconsideration of, and modification by, realisticallyassumed mining, metallurgical, economic, marketing,legal, environmental, social and governmental factors.These assessments demonstrate at the time of reportingthat extraction is reasonably justified.



Processing Complex processing assets

Project capital capital expenditure which is associated with specificprojects of a non-routine nature

Proterozoic Era of geological time between 2,5x109 and 570x106 yearsago

Proved Mineral Reserve the economically mineable material derived from ameasured mineral resource. it is estimated with a highlevel of confidence. it is inclusive of diluting materialsand allows for losses that may occur when the material ismined. appropriate assessments, which may includefeasibility studies, have been carried out, includingconsideration of and modification by realistically assumedmining, metallurgical, economic, marketing, legal,environmental, social and governmental factors. theseassessments demonstrate at the time of reporting thatextraction is reasonably justified

Roches Argilleuses Talceuse(RAT

the RAT is considered the boundary between the R2 andR1 units and consists of an upper RAT Grises (R2) and alower RAT lilas (R1

Roches Siliceuses FeuilletéesFoliated (Laminated) andSilicified Rocks (RSF

this is a grey to light brown thinly bedded laminated andhighly silicified dolomites

Roches SilicieusesCellulaires or SiliceousRocks with Cavities (RSC

Vuggy and infilled massive to stromatolitic silicifieddolomites

SAMREC code South African code for reporting of Mineral Resourcesand Mineral Reserves

Saprolite a soft, earthy, typically clay-rich, thoroughly decomposedrock, formed in place by chemical weathering of igneous,sedimentary and metarmophic rocks

Schist/s a regionally metamorphasised rock characterised by aparallel arrangement of the bulk of the constituentminerals

Schistes De Base or BasalSchists (SDB

reddish-brown to grey silty and nodular dolomite tosiltstone

Sedimentary pertaining to rocks formed by the accumulation ofsediments, formed by the erosion of other rocks

Shaft an opening cut downwards from the surface fortransporting personnel, equipment, supplies, ore andwaste

Shales DolomitiquesSuperieurs or UpperDolomitic Shales (SDS

yellowish, cream to red bedded laminated dolomiticsiltstones and fine-grained sandstones.

Smelting a high temperature pyrometallurgical operation conductedin a furnace, in which the valuable metal is collected to amolten matte or doré phase and separated from the ganguecomponents that accumulate in a less dense molten slagphase.



SRK Group SRK Global limited

Stope underground void created by mining

Stope an excavation from which ore has been removed in aseries of steps. It is any excavation in a mine, other thandevelopment workings made for the purpose of extractingore

Stratigraphy study of stratified rocks in terms of time and space

Strike direction of line formed by the intersection of stratasurfaces with the horizontal plane, always perpendicularto the dip direction

Sub-vertical shaft an opening cut below the surface downwards from anestablished surface shaft

Sulphide sulphur bearing mineral

WOL/SX/EW RefineryProject

SX/EW Refinery

T17 Mine Musonoi – T17 Mine

Tailings finely ground waste rock from which valuable minerals ormetals have been extracted

The Code Mining Code (Law No. 007/2002)

Tilwezembe Mine Tilwezembe Mine

Volcanics one of three groups into which rocks have been divided.The vocanic assemblage includes all extrusive rocks andassociated intrusive ones

Volcanoclastics one of the three groups into which rocks have beendivided. The volcanic assemblage includes all extrusiverocks and associated intrusive ones

Abbreviations

3D Three dimensional

AAS Atomic Absorption Spectroscopy

ABA Acid Base Accounting

AEL African Explosive Limited

AGES Africa Geo-Environmental Services

AICo Acid Insoluble Co

AICu Acid Insoluble Cu

ANFO Ammonium nitrate and fuel oil

APELL Awareness and Preparedness for Emergencies at Local Level

ASCo Acid Soluble cobalt

ASCu acid Soluble Copper

BBR Beitbridge, Bulawayo Railway

BID Base Technical Information Date



CAT Caterpillar

CBO Community-based organisation

CCD Counter-Current Decantation

CCIC Caracle Creek International Consultancy

CCT Cyclone Classified Tailings

CDT Confederation Democratique du Traivailleurs

CAF cut and fill (longitudinal)

CGS Confederation Syndicale du Congo

CGTC Centrale Generale du Travail au Congo

CorProfit CorProfit Systems Africa

CSC Confederation Syndicale du Congo

CTP Conscience des Travailleurs & Paysans

CV coefficients of variation

DC Direct current

DCF Discounted cashflow

DCP DRC Copper and Cobalt Project

DCP SARL DRC Copper and Cobalt Project, “DCP”

DEM Developpements et Exploitations Minières

DIMA The Dikuluwe-Mashamba

DPEM Direction 206harge de la Protection de l’Environnement Minier

DPEM Department for the Protection of the Mining Environment

DRC Democratic Republic of Congo

DRMS Design Rock Mass Strength

DRO Diesel Range of Organics

E Young’s Modulus

EAP Environmental Adjustment Plan

EBIT Earnings Before Interest and Tax

EIS Environmental Impact Study

EIS Environmental Impact Statements

EMP Environmental Management Plan

EPCM Engineering Procurement Construction Management

ESIA Environmental and Social Impact Assessment

ESMP Environmental and Social Management Plan

Etang Nth Etang North

Etang Sth Etang South

EW Electro-winning

FEL Front End Loader



FNSR Rock fragment

FOB Free On Board

FOS factors of safety

FQM First Quantum Minerals

FRAP Framework Resettlement Action Plan

FS factors of safety

FSAIMM Fellow of SAIMM

FWTD Far West Tailings Dam

G&A General and Administrative

GDP Gross Domestic Product

GEC Global Enterprises Corporate Limited

GECL Gecamines and Global Enterprises Corporate Limited

GIIP Good international industry practice

GNI Gross National Income

GSI Geological Strength Index

H Horizontal

HG High Grade

HG SX high-grade copper solvent extraction

HR Human Resources

HSSE Health Safety Social and Environment

HV High voltage

HVDC High voltage Direct Curent

IAW Intake Airway

IFC International Finance Corporation

IMF International Monetary Fund

INPP Institut National de Préparation Professionnelle

INSS Institut National de Sécurité Sociale

IRR Internal Rate of Return

JKTech JKTech (Pty) Ltd

JVA Joint Venture Agreement

KCC Kamoto Copper Company

KCC SARL Kamoto Copper Company, “KCC”

KFL KFL Limited

KFL Kinross Forrest Limited

KITD Kamoto Interim Tailings Dam

KML Katanga Mining Limited

KMT Kinganyambo Musonoi Tailings



KOL Kamoto Operating Limited

KOV Kamoto Oliviera and Virgule

KZC Kolwezi concentrator

LCF Longitudinal Cut and fill

LEM limit equilibrium method

LG Low Grade

LG SX low-grade copper solvent extraction

LHD Load Haul and Dump

LHRS Long Hole Retreat Stoping

LME London Metal Exchange

LoM Life-of-Mine

LRMC Long Run Marginal Cost

MAR mean annual rainfall

Maxwell Maxwell GeoServices

MR Mining Regulations (Decree No. 038/2003 of 26 March 2003)

MRMR Mining Rock Mass Rating

MSI The UK Securities and Investment Institute

NGO Non-governmental organisations

NPA National Ports Authority

NPV Net Present Value

NPVS Mine Flow Optimiser

NRZ National Railway of Zimbabwe

OS open stoping

OTUC Organisation des Travailleurs Unis on Congo

PCB Polychlorinated Biphenyl

PE permis d’exploitation

PF probability of failure

PPCFpost pillar cut and fill (pillar may be eliminated under certainconditions)

PPE safety equipment

PPIA Primary Project Impacted Area

Pr. Eng Professional Engineer

PrSciNatProfessional Member of South African Council for NaturalScientists

QA/QC Quality Assurance and Quality Control

QC Quality control

RAP Room and Pillar

RAW Return Airway



RDS Remote Drilling Services

REGIDESO State parastatal

RO Repartiteur Ouest

RoM Run of Mine

ROP Roll-Over-Protection

RSZ Railway systems of Zambia

SACNASP South African Council for Natural Scientific Professions

SADC Southern African Development Community

SAG Semi-Autogenous Grinding

SAIMM The Southern African Institute of Mining and Metallurgy

SAMREC South African Minerals and Resources Committee

SANS South African National Standards

SANAS South African National System

SCK Station de Conversion de Kolwezi

SD Standard Deviation

SDP Social Development Plan

SEIA Social and Economic Impact Assessment

SEP Stakeholder Engagement Plan

Set Point SGS Lakefield and Set Point Laboratories

SG specific gravity

SGS SGS Inspection Service Ltd

SGS Lakefield SGS Lakefield Research Africa (Pty) Ltd

SI International System of Units

SNCC Société Nationale des Chemins de Fer du Congo

SNEL Société National d’Electricité

SPIA Secondary Project Impacted Area

SRK SRK Consulting (South Africa) (Proprietary) Limited

SSL Soil Screening Levels

SX/EW Solvent Extraction/Electro-winning

TAP Trans-Africa Projects

TCLP Toxicity Characterization Leach Protocol

TDS Total Dissolved Solids

TEM Technical Economic Model

The SAMREC CodeSouth African Code for the Reporting of Mineral Resources andReserves

The SAMVAL Code South African Code for the Reporting of Mineral Asset Valuation

TPH Total Petroleum Hydrocarbon

TT Thickened Tailings



UCS Uniaxial Compressive Strength

UG underground

UMHK Union Miniere du Haut Katanga

UN United Nations

UNTC Union Nationale des Travailleurs du Congo

USD United States Dollar

V Vertical

WACC weighted average cost of capital

WGS84 World Geodetic System of 1984

WHO World Health Organisation

XRF X-Ray Fluorescence

ZAR South African Rand

ZRL Zambian Railways

Units

% percentage

%ASCu percentage Acid Soluble copper

%CaO percentage calcium oxide

%Cu percentage copper

%CuO percentage copper as oxide

%TCo percentage total cobalt

%TCu percentage total copper

± plus or minus

bcm bank cubic meter

bn billion

c/lb cents per pond

dBA decibels

GPa Giga Pascal

ha hectare

ha/yr hectare per year

kg Kilogram

kg/t kilogram per tonne

km kilometre

km/h kilometres per hour

km/hr kilometres per hour

km2 square kilometres

kPa kilo Pascal

kt kilo tonne

ktpa kilo tonnes per annum

ktpm kilo tonnes per month

kV kilo Volt

kV AC kilo Volt Alternating Current

kWh Kilowatt-hour



kWh Kilowatt-hour

l litre

l/hr litres per hour

l/sec litres per second

lb pound

m metre

m/d metres per day

m/s metre per second

m² square metre

m2/day square metre per day

m³ cubic metres

m3/d cubic metres per day

m3/ha/d cubic metre per hectare per day

m3/hr cubic metres per hour

m³/s cubis metres per second

mamsl metres above mean sea level

mbgl metres below ground level

mg/l milligram per litre

mm milli metre

mm/year millimetre per year

Mm3 Million cubic metres

MPa Mega Pascal

Mt Million tonnes

Mt Million tonnes

Mtpa Million tonnes per annum

MVA Mega Volt Ampere

MW Mega Watt

MWh Mega Watt hour

º Degrees

pH Measure of the acidity or alkalinity of a solution

sec second

sq. km square kilometres

t tonne (1000 kg)

t/m3 tonnes per cubic metre

tpa tonnes per annum

tpd tonnes per day

tph tones per hour

tphr tones per hour

tpvm tonnes per vertical metre

USD/bcm United States Dollars per bank cubic metre

USD/h United States Dollars per hour

USD/t United States Dollars per tonne

USD/t/km United States Dollars per tonne per kilometre

USDm United States Dollar million

vmpa vertical meter per annum



Chemical Elements

(Co,Cu)2S4 carrolite

(Co,Cu,Mn,Fe)O(OH) heterogenite

(Cu,Co)2(CO3)(OH)2 kolwezite

(Fe,Co)O(OH) goethite

(Mg,Fe)5Al(Si3Al)O10(OH)8 chlorite

As arsenic

Ca,Mg(CO3)2 dolomite

CaCO3 limestone

CuO copper oxide

CaO lime

Co cobalt

Co(OH)2 cobalt hydroxide

Cr chrome

Cu copper

Cu2 (OH)PO4 liberthenite

Cu2CO3(OH)2 malachite

Cu2O cuprite

Cu2S chalcocite

Cu3(PO4)(OH)3 cornetite

Cu5(PO4)2(OH)4.H2O pseudomalachite

Cu5FeS4 bornite

CuS covellite

Fe iron

H2S hydrogen sulphide

H2SO4 sulphuric acid

K-Al-Mg-Fe silicate hydroxides clay

KMg3Si3AlO10(F,OH)2 mica

MgO magnesium oxide

Mn manganese

NaHS sodium hydrogen sulphide

Ni nickel

NO2 nitrogen dioxide

Pb lead

Se selenium

SiO2 Silica / quartz

SO2 sulphur dioxide



SRK Consulting

REPORT DISTRIBUTION RECORD

REPORT

NO. :

kol ni43-101 draft (jn389772) - 2009 (v32)

COPY NO. :REV(0)

Name/Title Company Copy Date Authorized by

Roger Dixon

Ebrahim Takolia

Victor Simposya

Ally Burger

Henrietta Salter

Petrus Cilliers

APPROVAL

SIGNATURE :

COPYRIGHT

These technical reports (a) enjoy copyright protection and the copyright vests in SRK unless

otherwise agreed to in writing; (b) may not be reproduced or transmitted in any form or by any

means whatsoever to any person without the written permission of the copyright holder.

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