southern ashanti gold project ghana, west africa (21 august 2008)

8/19/2019 Southern Ashanti Gold Project Ghana, West Africa (21 August 2008)

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Southern Ashanti Gold Project

Ghana, West Africa

Technical Report

Effective Date: 21 August 2008

Prepared by:

Ron Heeks

Technical Manager

Adamus Resources Limited

On behalf of:




Adamus Resources LimitedSouthern Ashanti Gold Project, Ghana, Western Africa August 2008Technical Report

TABLE OF CONTENTS

1.0

Summary ....................................................................................................................1

1.1

General..........................................................................................................................1

1.2

Ownership .....................................................................................................................2 1.3

Mineral Resource Estimate............................................................................................4

1.4

Metallurgical Testing......................................................................................................6

1.4.1

General............................................................................................................6

1.4.2

Comminution Testing.......................................................................................6

1.4.3

Metallurgical Testing........................................................................................6

1.4.4

Testwork Summary..........................................................................................7

1.5

Mineral Processing........................................................................................................8

1.6

Open Pit Mining...........................................................................................................10

1.7

Ore Reserve Estimate .................................................................................................12

1.8

Financial Analysis Results ...........................................................................................15

1.9

Current Project Status .................................................................................................15

1.9.1

Mining Lease and License.............................................................................15

1.9.2

Project Implementation Plan..........................................................................15

1.10

Conclusion and Recommendations .............................................................................16

2.0 Introduction and Terms of Reference....................................................................17

2.1

Terms of Reference.....................................................................................................17

2.2

The Purpose of this Report..........................................................................................17

2.3

Qualifications and Experience .....................................................................................17

2.4

Principal Sources of Information..................................................................................18

3.0

Reliance on Other Experts .....................................................................................19

4.0 Property Description and Location........................................................................20

4.1

Project Area, Location and Access ..............................................................................20

4.2

Ownership ...................................................................................................................21

4.3

Description of Licences and Approvals........................................................................25

4.3.1

Anwia Deposit ...............................................................................................25

4.3.2

Salman Deposits ...........................................................................................25

4.3.3

Satellite Deposits...........................................................................................25

4.3.4

Royalties and Other Agreements...................................................................26

4.3.5

Environmental Liabilities................................................................................26

4.3.6

Extension Application ....................................................................................26

4.3.7

Grant of Mining Lease ...................................................................................26

4.3.8

Project Stability Agreement ...........................................................................27

5.0 Accessibi liTy, Cl imate, In frastructure and Physiography ................................... 27

5.1

Access.........................................................................................................................27

5.2

Climate ........................................................................................................................27

5.3

Topography, Elevation and Vegetation........................................................................27

5.4

Local Infrastructure......................................................................................................28

6.0 Exploration and Mining History .............................................................................28

6.1

Historical Mining Activity..............................................................................................28

6.2

Exploration and Ownership History of Salman Deposit Area .......................................29

6.3

Exploration and Ownership History of Anwia Deposit Area..........................................30

6.4

ARL Exploration...........................................................................................................31

6.5

Previous Resource Estimates......................................................................................31




13.5

Salman Assay Accuracy ..............................................................................................72

13.6

Salman Sampling and Assaying Precision...................................................................84

13.7

Conclusion...................................................................................................................87

14.0 ADjacent Propert ies ................................................................................................ 88

14.1

Adjacent properties......................................................................................................88

15.0

Metallurgi cal Testing ...............................................................................................88

15.1

Introduction..................................................................................................................88

15.2

Composite Samples and Sample Preparation .............................................................89

15.2.1

General..........................................................................................................89

15.2.2

Comminution Composites..............................................................................89

15.2.3

Leach Master Composites.............................................................................91

15.2.4

Leach Variability Composites ........................................................................92

15.2.5

Head Assays .................................................................................................95

15.3

Comminution ...............................................................................................................97

15.3.1

General..........................................................................................................97

15.3.2

Crushing Work Index .....................................................................................97

15.3.3

Unconfined Compressive Strength Tests ......................................................98

15.3.4

Advanced Media Competency Tests ...........................................................100

15.3.5

JK Drop Weight Tests..................................................................................102

15.3.6

SMC Testing................................................................................................110

15.3.7

Bond Comminution Tests ............................................................................111

15.4

Mineralogy.................................................................................................................112

15.4.1

General........................................................................................................112

15.4.2

Anwia Oxide Master ....................................................................................112

15.4.3

Anwia Transition Master ..............................................................................113

15.4.4

Anwia Sulphide Master................................................................................113

15.4.5

Salman Oxide Master..................................................................................113

15.4.6

Salman Transition Master............................................................................113

15.4.7

Salman Central Upper Transition (Variability Composite 12).......................113

15.4.8

Salman Central Lower Transition (Variability Composite 8).........................114

15.4.9

Salman North Upper Transition (Variability Composite 13)..........................114

15.4.10

Salman North Lower Transition (Variability Composite 9)............................114

15.4.11

Salman North Upper Transition (Variability Composite 23) – Granite..........114

15.4.12

Salman North Lower Transition (Variability Composite 21) – Granite..........114

15.4.13

Salman Sulphide AMC Comminution Composite.........................................115

15.5

Thickening .................................................................................................................115

15.5.1

General........................................................................................................115

15.5.2

Flocculant Screening Tests .........................................................................115

15.5.3

Dynamic Thickening Tests...........................................................................115

15.6

Viscosity ....................................................................................................................117

15.6.1

General........................................................................................................117

15.6.2

Anwia Oxide Master ....................................................................................117

15.6.3

Anwia Transition Master ..............................................................................119

15.6.4

Anwia Sulphide Master................................................................................120

15.6.5

Salman Oxide Master..................................................................................120

15.6.6

Salman Transition Master............................................................................121

15.7

Gravity Recovery .......................................................................................................123

15.7.1

General........................................................................................................123

15.7.2

3kg Batch Gravity Tests...............................................................................123

15.7.3

Bulk Gravity Tests .......................................................................................126

15.8

Leaching....................................................................................................................127

15.8.1

General........................................................................................................127

15.8.2

Leach Optimisation Testing .........................................................................127

15.8.3

Leach Variability Testing..............................................................................141

15.8.4

Effect of Oxygen Addition ............................................................................141

15.9

Oxygen Uptake..........................................................................................................146




15.9.1

General........................................................................................................146

15.10

Carbon Adsorption.....................................................................................................148

15.10.1

General........................................................................................................148

15.11

Cyanide Detoxification and Arsenic Precipitation.......................................................148

15.11.1

General........................................................................................................148

15.11.2

Cyanide Detoxification.................................................................................148

15.11.3

Arsenic Precipitation....................................................................................149

15.12

Flowsheet Selection ..................................................................................................154

15.13

Recovery Forecasts...................................................................................................155

16.0 Mineral Processing ...............................................................................................158

16.1

Engineering Design and Control Philosophy..............................................................158

16.2

Plant Configuration Options.......................................................................................158

16.3

Run of Mine (ROM) Pad ............................................................................................159

16.4

Crushing ....................................................................................................................161

16.5

Grinding and Classification ........................................................................................161

16.6

Gravity Concentration ................................................................................................162

16.7

Leach Feed Thickening .............................................................................................162

16.8

Leach and Adsorption Circuit.....................................................................................162

16.9

Elution and Gold Room Operations ...........................................................................163

16.9.1

Acid Wash ...................................................................................................164

16.9.2

Zadra Elution Circuit ....................................................................................164

16.9.3

Electrowinning and Gold Room ...................................................................164

16.9.4

Gold Barring ................................................................................................165

16.9.5

Gold Room Security ....................................................................................165

16.9.6

Carbon Regeneration ..................................................................................165

16.10

Cyanide Destruction and Tailings Disposal................................................................166

16.11

Reagents ...................................................................................................................167

16.11.1

Lime ............................................................................................................167

16.11.2

Cyanide .......................................................................................................167

16.11.3

Caustic ........................................................................................................167

16.11.4

Hydrochloric Acid.........................................................................................167

16.11.5

Activated Carbon.........................................................................................167

16.11.6

Sodium Metabisulphite ................................................................................168

16.11.7

Copper Sulphate..........................................................................................168

16.11.8

SAG Mill Balls..............................................................................................168

16.11.9

Flocculant....................................................................................................168

16.11.10

Ferric Sulphate ............................................................................................168

16.11.11

Sulphuric Acid .............................................................................................168

16.12

Services and Water ...................................................................................................169

16.12.1

Raw Water Supply.......................................................................................169

16.12.2

Process Water.............................................................................................169

16.12.3

Potable Water..............................................................................................169

16.12.4

Instrument Air ..............................................................................................169

16.12.5

Plant Air.......................................................................................................169

16.12.6

Low Pressure Air .........................................................................................169 16.12.7

Oxygen........................................................................................................170

16.12.8

Diesel Fuel ..................................................................................................170

16.12.9

Emergency Power Supply ...........................................................................170

17.0 Mineral Resource Estimate ..................................................................................171

17.1

Data Preparation and Treatment ...............................................................................171

17.1.1

Anwia Modelling Domains ...........................................................................171

17.1.2

Salman Modelling Domains .........................................................................173

17.1.3

Satellite Deposit Modelling Domains ...........................................................177

17.1.4

Mine Voids...................................................................................................181

17.1.5

Derivation of Preferred Assay Values ..........................................................181




17.1.6

Compositing ................................................................................................183

17.2

Exploratory Data Analysis..........................................................................................186

17.2.1

Anwia ..........................................................................................................186

17.2.2

Salman........................................................................................................186

17.2.3

Satellite Deposits.........................................................................................186

17.3

Spatial Continuity Analysis ........................................................................................187

17.3.1

Measures of Spatial Continuity ....................................................................187

17.3.2

Directional Controls on Gold Mineralisation.................................................187

17.4

Indicator Kriging.........................................................................................................196

17.4.1

Indicator Kriging for Recoverable Resource Estimation...............................196

17.4.2

Indicator Kriging Parameters .......................................................................197

17.5

Block Support Adjustment (Variance Adjustment) .....................................................200

17.5.1

General........................................................................................................200

17.5.2

The Variance Adjustment ............................................................................201

17.5.3

Shape of the Block Grade Distribution.........................................................201

17.5.4

The Information Effect .................................................................................201

17.5.5

Variance Adjustments Applied to the Resource Models ..............................202

17.6

Resource Classification .............................................................................................204

17.7

Anwia Resource Model..............................................................................................204

17.8

Salman Resource Model ...........................................................................................208 17.9

Satellite Deposits Resource Models ..........................................................................211

17.10

Mineral Resource Statement .....................................................................................216

17.11

Other .........................................................................................................................216

18.0

ORE Reserve Estimate..........................................................................................217

18.1

Introduction................................................................................................................217

18.2

Mining Study Scope...................................................................................................218

18.3

Parameters................................................................................................................218

18.3.1

Parameters Summary..................................................................................219

18.3.2

Gold Price and Royalty ................................................................................219

18.3.3

Throughput Costs........................................................................................219

18.3.4

Contract Mining Costs .................................................................................219

18.3.5

Resource Model and Surfaces ....................................................................222

18.3.6

Other Parameters........................................................................................222

18.4

Pit Limit Optimisations ...............................................................................................222

18.5

Mine Design...............................................................................................................227

18.6

Mining Quantities and Reserves................................................................................233

18.6.1

Cut-off Grades.............................................................................................233

18.6.2

Open Pit Quantities .....................................................................................234

18.6.3

Ore Reserve Estimate .................................................................................234

19.0 Other Relevant Data and Information ..................................................................237

19.1

Production Schedules................................................................................................237

20.0 Interpretation and Conclusions ...........................................................................238

21.0 Recommendations ................................................................................................239

22.0 References and Bibl iography...............................................................................240

23.0

Date and Signature Page ......................................................................................244

24.0 cert if icate of qualif ication .....................................................................................245




LIST OF TABLES

Table 1-1 Summary of Southern Ashanti Gold Project Resources at 0.8g/t cut-off ...........................5

Table 1-2 Key Process Design Parameters ......................................................................................8

Table 1-3 Metallurgical Recoveries...................................................................................................8 Table 1-4 Mineral Reserve Estimate...............................................................................................14

Table 2-1 Principal Sources of Information .....................................................................................18

Table 3-1 Reliance on Consultants in Addition to the Authors ........................................................19

Table 4-1 Tenure Summary – Southern Ashanti Gold Project ........................................................22

Table 6-1 Southern Ashanti Gold Resources at 1g/t cut-off estimated by Ravensgate ...................31

Table 6-2 Southern Ashanti gold resources at 1g/t cut-off estimated by SRK.................................31

Table 6-3 Southern Ashanti Gold Project resources at 1g/t cut-off estimate, ARL, January 2006...32

Table 6-4: Southern Ashanti Gold Project Resources at 1g/t cut-off estimate ARL, January 2007 ..32

Table 12-1

Bulk densities applied to the Anwia resource model.......................................................54

Table 12-2 Bulk densities applied to the Salman resource model...................................................56

Table 12-3: Bulk densities applied to the Satellite Deposits resource models..................................58

Table 15-1 Anwia Comminution Transition Master Composite Sample...........................................89

Table 15-2 Anwia Comminution Sulphide Master Composite Sample ............................................90

Table 15-3 Anwia Comminution Variability Composite Samples.....................................................90

Table 15-4 Salman Comminution Oxide Master Composite Sample...............................................90

Table 15-5 Salman Comminution Transition Master Composite Sample ........................................90

Table 15-6 Salman Comminution Sulphide Master Composite Sample ..........................................91

Table 15-7 Salman Comminution Variability Composite Samples ..................................................91

Table 15-8 Leach Master Composite Samples ...............................................................................92

Table 15-9 Salman Leach Variability Composite Samples............................................................93

Table 15-10 Anwia Leach Variability Composite Samples ..............................................................95

Table 15-11 Leach Master Composite Sample Head Assays .........................................................95

Table 15-12 Salman Leach Variability Composite Head Assays...................................................................96

Table 15-13 Anwia Leach Variability Composite Head Assays.....................................................................97

Table 15-14 Crushing Work Index Test Results...............................................................................98

Table 15-15 Anwia Variability Composite Sample Unconfined Compressive Strength ....................98

Table 15-16 Salman Variability Composite Sample Unconfined Compressive Strength.................99

Table 15-17 Master Composite JK Drop Weight Test Parameters................................................102

Table 15-18 SMC Test Results.....................................................................................................110

Table 15-19 Bond Test Results ....................................................................................................111

Table 15-20 Levin Test Results ....................................................................................................112

Table 15-21 Anwia Oxide Master Dynamic Thickening Tests .......................................................116

Table 15-22 Anwia Transition Master Dynamic Thickening Tests.................................................116

Table 15-23 Anwia Sulphide Master Dynamic Thickening Tests...................................................116

Table 15-24 Salman Oxide Master Dynamic Thickening Tests.....................................................117

Table 15-25 Salman Transition Master Dynamic Thickening Tests.............................................117

Table 15-26 Anwia Oxide Master Viscosity Test Summary...........................................................118

Table 15-27 Anwia Transition Master Viscosity Test Summary ....................................................119

Table 15-28 Anwia Sulphide Master Viscosity Test Summary ......................................................120

Table 15-29 Salman Oxide Viscosity Test Summary ....................................................................121

Table 15-30 Salman Transition Viscosity Test Summary..............................................................122

Table 15-31 Average 3kg Batch Gravity Test Results – Anwia .....................................................123

Table 15-32 3kg Batch Gravity Test Results – Anwia Oxide Variability Samples..........................124

Table 15-33 3kg Batch Gravity Test Results – Anwia Transition Variability Samples ...................124




Table 15-34 3kg Batch Gravity Test Results – Anwia Sulphide Variability Samples .....................124

Table 15-35 Average 3kg Batch Gravity Test Results – Salman...................................................125

Table 15-36 3kg Batch Gravity Test Results – Salman Oxide Variability Samples .......................125

Table 15-37 3kg Batch Gravity Test Results – Salman Transition Variability Samples.................125

Table 15-38 3kg Batch Gravity Test Results – Salman Sulphide Variability Samples...................126

Table 15-39 Anwia Bulk Gravity Test Results ...............................................................................126

Table 15-40 Salman Bulk Gravity Test Results.............................................................................127

Table 15-41 Anwia Oxide Leach Optimisation Test Results..........................................................128

Table 15-42 Anwia Transition Leach Optimisation Test Results ...................................................129

Table 15-43 Anwia Sulphide Leach Optimisation Test Results .....................................................130

Table 15-44 Salman Oxide Leach Optimisation Test Results .......................................................131

Table 15-45 Salman Transition Leach Optimisation Test Results.................................................132

Table 15-46 Gravity Circuit Economic Analysis Parameters .........................................................137

Table 15-47 Economic Analysis Parameters ................................................................................139

Table 15-48 Anwia Leach Variability Test Results ........................................................................142

Table 15-49 Salman Leach Variability Test Results – Oxide and Transition .................................143

Table 15-50 Salman Leach Variability Test Results – Sulphide ....................................................144

Table 15-51 Effect of Varying Oxygen/Aeration ............................................................................145

Table 15-52 Summary of Oxygen Uptake Test Results ................................................................147

Table 15-53 Summary of Carbon Adsorption Test Results ...........................................................148

Table 15-54 Summary of Cyanide Detoxification Test Results - Anwia.........................................150

Table 15-55 Summary of Cyanide Detoxification Test Results – Salman .....................................151

Table 15-56 Arsenic Precipitation Sighter Tests ...........................................................................152

Table 15-57 Summary of Arsenic Precipitation Optimisation Test Results....................................152

Table 15-58 Recommended Cyanide Detoxification and Arsenic Precipitation Parameters..........153

Table 15-59 Summary of Mill Feed and Operating Parameters ....................................................156

Table 15-60 Predicted Tailings Grades.........................................................................................157

Table 17-1: Anwia grid transformation parameters ........................................................................171

Table 17-2: Anwia resource modelling domains ............................................................................171

Table 17-3: Salman-Akanko resource modelling domains .............................................................173

Table 17-4: Satellite Deposits resource modelling domains...........................................................177

Table 17-5: Aliva grid transformation parameters ..........................................................................178

Table 17-6: Nfutu grid transformation parameters .........................................................................180

Table 17-7: Anwia drill holes preferred gold assay sources ...........................................................182

Table 17-8: Salman drill holes preferred gold assay sources.........................................................183

Table 17-9: Numbers of sample composites contained in Anwia modelling domains ....................185

Table 17-10: Numbers of sample composites contained in Salman modelling domains ................185

Table 17-11: Numbers of sample composites contained in Satellite Deposits modelling domains.185

Table 17-12: Anwia Model Framework & Kriging Search Parameters (Rotated Space).................198

Table 17-13: Salman Model Framework and Kriging Search Parameters......................................198

Table 17-14: Bokrobo Model Framework and Kriging Search Parameters ....................................199

Table 17-15: Aliva Model Framework and Kriging Search Parameters..........................................199 Table 17-16: Avrebo Model Framework and Kriging Search Parameters ......................................199

Table 17-17: Nfutu Model Framework and Kriging Search Parameters .........................................200

Table 17-18: Variance adjustments applied to the Anwia resource model .....................................202

Table 17-19: Variance adjustments applied to the Salman resource model...................................203

Table 17-20: Variance adjustments applied to the Satellite Deposits resource models..................203

Table 17-21: Summary of Southern Ashanti Gold Project Resources at 0.8g/t cut-off ...................216

Table 18-1 Salman Pit Optimisation Parameters Summary ..........................................................220

Table 18-2 Anwia Pit Optimisation Parameters Summary ............................................................220

Table 18-3 Resultant Average Process Recoveries for Resource Areas ......................................222




Table 18-4: Salman Pit Optimisation Results.................................................................................224

Table 18-5: Anwia pit Optimisation Results....................................................................................225

Table 18-6: Project Total Pit Optimisation Results.........................................................................226

Table 18-7: Cut-off Grades for Ore Reserves ................................................................................233

Table 18-8: Open Pit Quantities and Economics ...........................................................................235

Table 18-9: Ore Reserves by Mining Areas and Ore Types...........................................................236

Table 19-1 Production Schedule Results Summary......................................................................237




LIST OF FIGUR S

Figure 1-1 ARL Corporate Structure .................................................................................................3

Figure 1-2 Mineral Processing Flowsheet Summary.......................................................................11

Figure 1-3 Salman Mine Design .....................................................................................................13 Figure 1-4 Anwia Mine Design........................................................................................................14

Figure 4-1 Project Location.............................................................................................................20

Figure 4-2: ARL Corporate Structure ...............................................................................................21

Figure 4-3: Tenure Perimeter, Geology and Deposit Locations .......................................................24

Figure 7-1 Regional Geology ..........................................................................................................33

Figure 7-2 Salman Deposit Geology (Including Akanko).................................................................38

Figure 8-1 Anwia Deposit Geology and Mineralisation....................................................................43

Figure 12-1 Bulk densities of Anwia very weathered drill core ........................................................54

Figure 12-2 Bulk densities of Anwia moderately weathered drill core .............................................54

Figure 12-3 Bulk densities of Anwia weakly weathered drill core....................................................55

Figure 12-4 Bulk densities of Anwia fresh rock drill core.................................................................55

Figure 12-5 Bulk densities of Salman very weathered drill core......................................................56

Figure 12-6 Bulk densities of Salman moderately weathered drill core...........................................57

Figure 12-7 Bulk densities of Salman weakly weathered drill core..................................................57

Figure 12-8 Bulk densities of Salman fresh rock drill core ..............................................................58

Figure 13-1 Assays of Samax/AGC blanks submitted with Anwia drill samples..............................60

Figure 13-2 Assays of Adamus blanks submitted with Anwia drill samples.....................................60

Figure 13-3 Assays of STD4B submitted with Adamus Anwia samples ..........................................61






Figure 13-9 Assays of STD10B submitted with Adamus Anwia samples ........................................63

Figure 13-10 Scatter plot of nearest neighbour sample pairs..........................................................64

Figure 13-11 Q-Q plot of nearest neighbour sample pairs ..............................................................64

Figure 13-12 Scatter plot: Samax/AGC Anwia field re-splits ...........................................................65

Figure 13-13 Precision plot: Samax/AGC field re-splits ..................................................................65

Figure 13-14 Scatter plot: Adamus Anwia field re-splits ..................................................................67

Figure 13-15 Precision plot: Adamus Anwia field re-splits...............................................................67

Figure 13-16 Sample recoveries in 2002-2003 RC drilling, very weathered material ......................68

Figure 13-17 Sample recoveries in 2002-2003 RC drilling, moderately weathered material ...........69

Figure 13-18 Sample recoveries in 2002-2003 RC drilling, weakly weathered material..................69

Figure 13-19 Sample recoveries in 2002-2003 RC drilling, fresh rock ............................................70

Figure 13-20 Sample recoveries in 2006 RC drilling, very weathered material ...............................70

Figure 13-21 Sample recoveries in 2006 RC drilling, moderately weathered material ....................71

Figure 13-22 Sample recoveries in 2006 RC drilling, weakly weathered material...........................71

Figure 13-23 Sample recoveries in 2006 RC drilling, fresh rock .....................................................72

Figure 13-24 Adamus blanks submitted for SGS fire assay with Salman drill samples...................73

Figure 13-25 Adamus blanks submitted for Transworld CN leach assay with Salman drill samples73

Figure 13-26 Adamus blanks submitted for Transworld fire assay with 2002 drill samples .............73









Figure 13-32: Adamus blanks submitted for Genalysis fire assay with Salman drill samples...........75

Figure 13-33: Assays of STD1 submitted to SGS with Salman samples..........................................76

Figure 13-34: Assays of STD4B submitted to SGS with Salman samples .......................................76






Figure 13-40: Assays of STD10B submitted to SGS with Salman samples .....................................78

Figure 13-41: Assays of STD11 submitted to SGS with Salman samples........................................79

Figure 13-42: Assays of STD12 submitted to SGS with Salman samples........................................79

Figure 13-43: Assays of STD1 submitted to Transworld with Salman samples................................79

Figure 13-44: Assays of STD02 submitted to Transworld with Salman samples..............................80



Figure 13-47: Assays of STD4B submitted to Transworld with Salman samples .............................81








Figure 13-55: Assays of STD10B submitted to Transworld with Salman samples ...........................83

Figure 13-56: Assays of STD11 submitted to Transworld with Salman samples..............................84

Figure 13-57: Assays of STD13B submitted to Transworld with Salman samples ...........................84

Figure 13-58: Scatter plot: SGS fire assays of Salman field re-splits ...............................................85

Figure 13-59: Precision plot: SGS fire assays of Salman field re-splits............................................85

Figure 13-60: Scatter plot: Transworld fire assays of Salman field re-splits .....................................86

Figure 13-61: Precision plot: Transworld fire assays of Salman field re-splits..................................86

Figure 13-62: Scatter plot: Transworld CN leach assays of Salman field re-splits ...........................87

Figure 13-63: Precision plot: Transworld CN leach assays of Salman field re-splits ........................87

Figure 15-1 Unconfined Compressive Strength of Anwia Ore Variability Samples..........................99

Figure 15-2 Unconfined Compressive Strength of Salman Ore Variability Samples .....................100

Figure 15-3 Media Competency Test Product Sizing ....................................................................101

Figure 15-4 Impact Work Index Testing of AMC Test Survivors....................................................101

Figure 15-5 Variation of Impact Resistance with Particle Size – Anwia Transition........................103

Figure 15-6 Variation of Crushing Energy with Particle Size – Anwia Transition.........................103

Figure 15-7 Variation of Impact Resistance with Particle Size – Anwia Sulphide..........................105

Figure 15-8 Variation of Crushing Energy with Particle Size – Anwia Sulphide.............................105 Figure 15-9 Variation of Impact Resistance with Particle Size – Salman Oxide............................106

Figure 15-10 Variation of Crushing Energy with Particle Size – Salman Oxide.............................106

Figure 15-11 Variation of Impact Resistance with Particle Size – Salman Transition....................108

Figure 15-12 Variation of Crushing Energy with Particle Size – Salman Transition ......................108

Figure 15-13 Variation of Impact Resistance with Particle Size – Salman Sulphide .....................109

Figure 15-14 Variation of Crushing Energy with Particle Size – Salman Sulphide ........................109

Figure 15-15 Variation in Viscosity with Shear Rate and Pulp Density – Anwia Oxide................119

Figure 15-16 Variation in Viscosity with Shear Rate and Pulp Density – Salman Oxide ...............122

Figure 15-17 Effect of Cyanide Concentration on Cyanide Consumption – Salman .....................134




Figure 15-18 Effect of Cyanide Concentration on Cyanide Consumption – Anwia........................134

Figure 15-19 Effect of Gravity Gold Recovery – Salman Oxide ....................................................135

Figure 15-20 Effect of Gravity Gold Recovery – Anwia Oxide.......................................................136

Figure 15-21 Effect of Gravity Gold Recovery – Anwia Transition ................................................136

Figure 15-22 Effect of Gravity Gold Recovery – Anwia Sulphide ..................................................137

Figure 15-23 Effect of Varying Grind Size and Leach Time – Anwia Sulphide..............................140

Figure 15-24 Effect of Varying Grind Size and Leach Time – Salman Oxide................................140

Figure 15-25 Effect of Oxygen Addition – Anwia Sulphide ............................................................146

Figure 15-26 Oxygen Demand - Anwia and Salman Ores ............................................................147

Figure 15-27 Final Arsenic Concentration Variation with pH and Fe:As Molar Ratio.....................153

Figure 15-28 Recovery Variation with P80 and Leach Time .........................................................155

Figure 16-1 Summary Process Flowsheet Option 5.......................................................................160

Figure 17-1 Plan view of Anwia model domain wireframes...........................................................172

Figure 17-2 Interpreted geology, weathering and mineralisation, Anwia section 550350N ...........172

Figure 17-3: Plan view of the Salman-Akanko mineralisation wireframes......................................174

Figure 17-4: Interpreted geology and mineralisation on Salman South section 551850N ..............175

Figure 17-5: Interpreted geology and mineralisation on Salman Central section 552200N ............175

Figure 17-6: Interpreted geology and mineralisation on Teberu section 553800N .........................176

Figure 17-7: Interpreted geology and mineralisation on Salman North skewed section 554300N..176

Figure 17-8 Pseudo 3D view of Bokrobo model domain wireframes..............................................178

Figure 17-9: Pseudo 3D view of Aliva model domain wireframes in rotated grid............................179

Figure 17-10: Pseudo 3D view of Avrebo model domain wireframe...............................................180

Figure 17-11: Pseudo 3D view of Nfutu drill holes and trenches....................................................181

Figure 17-12: 3D variogram map,Indicator Threshold P 0.5, Anwia, Domain 1..............................188

Figure 17-13: 3D variogram map, Indicator Threshold P 0.5, Anwia, Domain 2.............................188

Figure 17-14 3D variogram map, Indicator Threshold P 0.5, Salman South Model, Domain 1 .....189

Figure 17-15 3D variogram map, Indicator Thresold P 0.5, Salman South Model, Domain 3 .......190

Figure 17-16 3D variogram map, Indicator Thresold P 0.5, Salman South Model, Domain 5 .......190

Figure 17-17 3D variogram map, Indicator Threshold P 0.5, Salman Central Model, Domain 1 ..191

Figure 17-18 3D variogram map, Indicator Threshold P 0.5, Salman Central Model, Domain 2 ..191

Figure 17-19 3D variogram map, Indicator Threshold P 0.5, Salman North Model, Domain 2......192

Figure 17-20: 3D variogram map, Indicator Threshold P 0.5, Salman North Model, Domain 3 ......192

Figure 17-21: 3D variogram map, Indicator Threshold P 0.5, Salman North Model, Domain 4, 5 and 6

193

Figure 17-22: 3D variogram map, Indicator Threshold P 0.5, Salman all models and all Primary

Domains, Sub Domain 1 ........................................................................................................193

Figure 17-23 3D variogram surface for the median indicator variogram model, Bokrobo Main Zone

(used for all domains) ............................................................................................................194

Figure 17-24: 3D variogram surface for the median indicator variogram model, Aliva Domain 1 ...194

Figure 17-25: 3D variogram surface for the median indicator variogram model, Aliva Domain 2 and 3

(also Domain 0)......................................................................................................................195

Figure 17-26: 3D variogram surface for the median indicator variogram model, ............................195 Figure 17-27: Anwia Panel Mean Grade Estimates, Section 550400N ..........................................205

Figure 17-28: Anwia Panel Recoverable Proportions at 1 g/t Cut-off, Section 550400N................205

Figure 17-29: Anwia Panel Confidence Categories, Section 550400N ..........................................206

Figure 17-30: Anwia Panel Mean Grade Estimates, Plan at 1.5RL................................................206

Figure 17-31: Anwia Panel Recoverable Proportions at 1 g/t Cut-off, Plan at 1.5RL......................207

Figure 17-32: Anwia Panel Confidence Categories, Plan at 1.5RL ................................................207

Figure 17-33: Salman Panel Mean Grade Estimates, Section 552137.5N.....................................208

Figure 17-34: Salman Panel Recoverable Proportions at 1 g/t Cut-off, Section 552137.5N...........209

Figure 17-35: Salman Panel Confidence Categories, Section 552137.5N.....................................209




Figure 17-36: Salman Central Panel Mean Grade Estimates, Plan at 1001.5RL ...........................210

Figure 17-37: Salman Central Panel Recoverable Proportions at 1g/t Cut-off, Plan at 1001.5RL..210

Figure 17-38: Salman Central Panel Confidence Categories, Plan at 1001.5RL ...........................211

Figure 17-39: Bokrobo MIK Model showing 1.0g/t cut-off resource (model panels scaled in the east

dimension by the proportion of contained resource)...............................................................212

Figure 17-40: Aliva MIK Model showing 1.0g/t cut-off resource (model panels scaled in the east


Figure 17-41: Avrebo MIK Model showing 1.0g/t cut-off resource (model panels scaled in the east


Figure 17-42: Nfutu MIK Model showing 1.0g/t cut-off resource (model panels scaled in the east


Figure 18-1: Salman Central and South Pit Designs......................................................................228

Figure 18-2: Salman North Pit Designs..........................................................................................229

Figure 18-3: Akanko Pit Designs ...................................................................................................230

Figure 18-4: Anwia Ultimate Pit Design (Local Grid)......................................................................231

Figure 18-5: Anwia Pit Stage Designs (Local Grid)........................................................................232



Adamus Resources LimitedSouthern Ashanti Gold Project, Ghana, Western Africa August 2008Technical Report Page 1

1

1.0 SUMMARY

1.1 General

The Southern Ashanti Gold Project (the Project) comprises the greenfields development of an opencut mining operation, a process plant and related infrastructure to mine and process ore from the

defined reserves of the Salman, Akanko and Anwia deposits. The Project is located in the Western

Region of Ghana, approximately 280 km west of the capital, Accra, and less than 20 kilometres from

the coast at Essiama. The Project area is accessed from Accra via the main coast highway to

Takoradi and from there by sealed road to the village of Teleku Bokazo and then by 10 kilometres of

gravel road.

In 2006 Adamus Resources Limited (ARL or Adamus) commissioned a feasibility study (Study) as part

of the technical and economic evaluation of the Project from experienced consultants including

Lycopodium Engineering Pty Ltd, Mining Solutions Consultancy Pty Ltd, SGS Environment (Ghana)

and Knight Piesold Consulting.

The Study had as its basis a treatment plant with a processing capacity of 1.3 million tonnes per

annum (Mtpa), which gave an approximate six and half year mine life based on the current reserve.

Following release of the Study results in June 2007, including the initial ore reserve estimate for the

Project, a NI43-101 compliant Technical Report of the Study was released in November 2007 with a

further revised report released in December 2007.

This report follows the release in February 2008 of the updated ore reserve and mineral resource

estimates for the Project and incorporates new data into both the mining and resource sections of the

previously published Technical Report. It incorporates new resource calculations for the Anwia and

Salman deposits and the inclusion of resources from Bokrobo, Aliva, Avrebo and Nfutu (the “Satellite

Deposits”) not previously included in the Study. It also updates the previous mining study to reflect

changes in the economics of mining and processing parameters since preparation of the December

2007 Technical Report.

The main changes are:

• Gold price increase of approximately 35%

• Contract mining cost increase of approximately 15%

• Processing cost increase of approximately 15%

• A reduction of the cut off grade from 1.0 to 0.8 g\t as the lower figure better reflects results

from the previous pit optimisations.

The updated reserve estimate is based on a gold price of $800/oz, a royalty of 3.6%, the escalation in

input

costs from December 2006 and a change in cut-off grade to 0.8 g/t,

Changes to the resource estimate reflect the addition of data from drill programs undertaken between

March 2007 and December 2007. The main changes are:

• An update of the Salman and Anwia resource models by Hellman and Schofield (H & S,

March 2008) as a result of further drilling

• The inclusion of four satellite deposits contributing 2.5 Million tonnes @ 1.78 g/t measured

and indicated resources at 0.8g/t cut off.




2

A significant change has been the conversion of inferred resources to indicated resources allowing

this material to be incorporated into the pit optimisations.

With the change in parameters, new pit optimisation tables have been produced. Total reserve

estimate has increased to 10.46 M tonnes

Gold has been mined on a small scale from southwestern Ghana for centuries and the Gold Coast

Geological Survey recorded widespread bedrock and alluvial workings at many locations within the

area now covered by the Project. Historic production from the small Akanko gold mine (Salman north)

may have amounted to a few thousand ounces but otherwise there are no historic production

estimates. Small-scale artisanal gold mining activities (both alluvial and reef) continue to occur at a

few localities within the Project area.

The Project has a history of exploration going back to the late 1980’s. Exploration efforts in the 1990s

by various companies including Tropical Exploration and Mining Company, BHP Minerals, SEMAFO

Inc, SAMAX Gold Inc, and Ashanti Goldfields Corporation led to the identification of a significant

northerly trending zone of gold mineralisation in the eastern part of the Project, termed the Salman

Trend, and several discrete mineralised zones, including the Anwia Deposit, in the western part.Considerable drilling was undertaken at the Anwia Deposit by SEMAFO and SAMAX and a number of

resource estimates were subsequently made but the deposit was never progressed to mine.

ARL acquired the Salman deposit in 2002, the Anwia deposit in 2004 and the Akanko (Salman north)

deposit in 2005. Since acquisition ARL has continued exploration with both RC and diamond core

drilling campaigns, soil sampling and trenching over the three areas. The current resource has been

delineated by several phases of drilling from 1995 to 2007 by the various owners.

The Salman and Akanko deposits contain broad, near-surface zones of gold mineralisation extending

over several kilometres of strike. Approximately 8 kilometres of strike extent has been drill tested to

date indicating the presence of several discrete, multi-lode gold deposits, scattered along the shear

zone.

The Anwia deposit is located approximately 9 kilometres west of the Salman deposit. The surface

expression of the deposit consists of extensive, shallow dipping quartz veining, extending over several

hundred metres of strike. The deposit is hosted by a northeast dipping package of greywacke

(footwall) and interbedded greywacke-phyllite (hanging-wall).

Throughout this report resources and reserves are quoted relative to the JORC (2004) code, these

resource and reserve terms are equivalent to those referred to in the CIM Standards on Mineral

Resources and Reserves – Definitions and Guidelines, August 2000. Additionally all Mineral

Resources are quoted inclusive of Mineral Reserves.

This report has been compiled based on information available up to and including 31 December 2007.

All the currency reported in this document is in US dollars unless stated otherwise.

1.2 Ownership

ARL holds an interest in three granted prospecting licences, four granted mining licence and five

prospecting licence renewal applications currently covering a combined Project area of approximately

464 km2. A detailed description of these licenses is included in Table 4.1 of Section 4.2 below.




3


Adamus Investment

Enterprises Pty Ltd

African Gold

Properties Ltd

Castlegem

Pty Ltd

Adamus Holdings

Pty Ltd

Adamus Resources

Limited (Ghana) (1)

Semafo (Ghana)

Ltd(1)

Nkroful Mining

Limited(1)

Akanko Mining

Limited(2)

Ghana

BVI Australia

ARL owns those tenements comprising the Project area by way of the company structure shown on

Figure 1-1. The tenements comprising the Project area are subject to the statutory 10 per cent

interest retained by the Government of Ghana.

Figure 1-1 ARL Corporate Structure

Notes:

1. Ghanaian government holds statutory right to a 10% interest in the Ghanaian subsidiaries upon commencement of

production.

2. A third party holds 1,000 shares in Akanko Mining Limited representing 11% of the outstanding shares, so that ARL

has an 89% interest (8,000 shares) in Akanko Mining Limited which will be reduced to 80% on conversion of the

Ghanaian government’s 10% interest.

The Anwia deposit lies on the Ebi-Teleku Bokazo mining licence ML2/192 held by ARL. The currentlicence was granted on 11 April 2008 and has a ten year, extendable term. A royalty of 3 per cent of

net profit or 1 per cent of production (in each case, in relation to ore derived from the area of the

original Teleku Bokazo prospecting licence), whichever is greater, is payable by Adamus to Super

Paper Products Company, a previous holder of the original prospecting licence.

The Salman deposit lies on the Salman mining licencewhich was granted to Adamus Resources

Limited (Ghana) for a period of ten years commencing 11 April 2008 for a ten year, extendable term

The Akanko deposit lies on the Akanko mining licence granted under the same terms as the Salman

mining licence. Hereafter throughout this report, the Salman deposit and Akanko deposit are

collectively referred to as the Salman deposit.

The Akanko mining licence ML2/128 is held by Akanko Mining Limited, a company 89 per cent owned

by Adamus Holdings Pty Limited (a wholly owned subsidiary of ARL) and 11 per cent by Tropical

Mining and Exploration Ltd (TEMCO). Upon conversion of the Ghana government’s 10% statutory

interest, ARL’s interest will be reduced to 80 per cent.

Parts of both the Salman consolidated and Ebi-Teleku Bokazo prospecting licences are subject to

concessions in favour of Super Paper Products Company that permit the extraction of kaolin clays and

small-scale mining of kaolin near the village of New Aluku. To ARL’s knowledge the kaolin




4

concessions do not impinge on any area proposed for the development of the gold mine or its

associated infrastructure.

1.3 Mineral Resource Estimate

Several discrete gold deposits (including the Salman and Akanko deposits), collectively referred to asthe Salman Trend, have now been delineated along the Salman Shear Zone within the Project area.

The Salman Shear Zone comprises a ductile shear between 10 and 50 metres thick separating a

western hanging-wall of deformed, graphitic phyllite and thin-bedded greywacke from an eastern

footwall of thick-bedded greywacke. The shear zone dips moderately to steeply west over much of

the 8 kilometres drilled extent but locally it rolls over to dip steeply to the east. An altered biotite

granitoid intrudes the shear zone in the northern portion of the resource area. Gold mineralisation

occurs predominantly in vertical to west-dipping lodes approximately parallel to and splaying out into

the footwall of the main shear zone, in quartz-veined silica-sericite-carbonate-arsenopyrite altered

greywacke and/or granite. Below the base of oxidation, most gold is associated with fine-grained

arsenopyrite and is partially refractory.

The Anwia deposit is a discrete zone of gold mineralisation located approximately 9 kilometres west ofthe Salman Trend. The deposit is hosted by a northeast dipping package of greywacke (footwall) and

interbedded greywacke-phyllite (hanging-wall). In the western (footwall) part of the deposit gold

mineralisation is also hosted by a steeply northeast dipping granite dyke that gradually converges on

the hanging-wall in the northwest of the resource area. The few facing indicators apparent suggest

the metasedimentary package is overturned. Gold mineralisation is intimately associated with pyrite

disseminated within and around a complex array of deformed pale grey to dark blue-grey quartz-

carbonate-sericite±albite veins. A broad sericite-carbonate alteration zone about 200 metres thick and

450 metres long is developed in the footwall greywacke. The silica-sericite alteration zone is more

extensive than the gold-pyrite mineralisation. Unlike the Salman Shear Zone, there is no significant

component of refractory gold in the sulphide zone at Anwia.

Estimates of resources at the Salman deposit rely predominantly on sampling by ARL. Since

commencement of drilling at the Salman deposit, Adamus has maintained a quality control protocol

that allows routine monitoring of sampling precision and assay accuracy. An examination of QAQC

sample data indicates satisfactory performance of f ield sampling protocols and of assay laboratories.

Estimates of resources for the Anwia deposit rely substantially on drill sample assays generated by

previous owners for which little QAQC information is available. A nearest-neighbour comparison of

gold grades in ARL’s drill samples with those in samples from previous drilling demonstrates that the

two sample populations are equally representative of mineralisation grades at Anwia.

Two previous resource estimates have been completed for the Project at earlier stages of exploration

by independent resource consultants Ravensgate Pty Ltd (August 2004) and SRK Consulting

(February 2005). Additionally, to fulfil listing requirements for the TSXV, the Project was the subject of

an independent technical report by RSG Global Pty Ltd dated February 2004. An internal resource

estimate update was undertaken in January 2006 and a report detailing that work was filed with the

appropriate Canadian securities regulatory authorities in March 2006.

Recoverable resources at the Anwia and Salman deposits have been estimated using the method of

Multiple Indicator Kriging (MIK) with block support adjustment. Geological and weathering domains

were imposed to define domains of similar grade tenor and directional trends. The models estimate

resources into panels with dimensions of 20mE x 25mN x 3mRL. MIK of gold grades used indicator




5

variography based on the resource sample grades, with continuity of gold grades characterised by

indicator variograms at 14 indicator thresholds. A block support adjustment, incorporating an

adjustment for Information Effect, was used to estimate the recoverable gold resources assuming a

selective mining unit of 5mE x 8mN x 3mRL and grade control sampling at 5mE x 8mN x 1.5mRL.

The shape of the local block gold grade distribution has been assumed lognormal or normal

depending on the skewness of the local histogram of gold grades at sample support within each panelas estimated by Indicator Kriging. The estimates are considered recoverable by open pit mining and

application of ore loss and dilution factors in quantifying ore reserves is not recommended.

The recoverable resource estimates within each panel have been classified according to the

distribution of sampling in the kriging neighbourhood. This classification scheme takes into account

the uncertainty in the estimates related to the proximity and spatial distribution of the informing sample

composites. A summary of the mineral resource estimates for the Anwia, Salman and Satellite

deposits at 0.8g/t Au cut-off grade is set out below in Table 1-1.

Table 1-1 Summary of Southern Ashanti Gold Project Resources at 0.8g/t cut-off

Category Measured Indicated Inferred

Deposit Cut offgrade(g/t)

Mtonnes g/t Au

k oz Au

Mtonnes

g/t Au

k oz Au

Mtonnes g/t Au

k oz Au

Anwia 0.8 6.2 2.01 400 2.8 2.00 180 2.6 1.7 140

Salman 0.8 11.4 1.73 630 5.6 1.54 280 2.5 1.5 125

SatelliteDeposits 0.8 1.0 2.10 70 1.5 1.57 70 1.3 1.8 75

Total 0.8 18.6 1.84 1,100 9.8 1.67 530 6.4 1.6 340




6

Variograms at both the Anwia and Salman deposits show short ranges along strike and down-dip.

This is typical of hydrothermal gold deposits and there is nothing to indicate that a sensible grade

control sampling strategy in future open pit mining operations could not outline ore parcels at an

economic cut-off grade.

1.4 Metallurgical Testing

1.4.1 General

The Anwia and Salman orebodies have been the subject of previous metallurgical testwork campaigns

by the present and former owners of the Project.

The most recent testwork programme was carried out from June 2006 through to May 2007 under the

direction of ARL and Ozmet. This testwork was primarily performed at Ammtec laboratories in

Australia on samples selected to be representative of the Project deposits.

A conventional cyanidation treatment route was assumed for the ores, based on assessment of the

earlier testwork.

The testwork programme had four objectives:

• Establish (confirm) the processing route;

• Determine the optimum plant operating parameters for the ores to be processed;

• Evaluate the variability in metallurgical performance for the different deposits; and

• Define parameters required for the engineering and design of the plant.

1.4.2 Comminut ion Testing

A full series of comminution testwork was completed on the Anwia and Salman primary orebodies.

The primary ore samples are classified as having low to moderate competency behaviour and

compressive strength testing classified the samples as weak.

Rod mill and ball mill work indices are moderate. The rod:ball ratio is low (1.13) suggesting there is a

low potential the ore will form critical sized material in a SAG mill.

The abrasiveness of all ore types is classified as moderate and indicates that liner and media

consumption will not be excessive.

Pebble crushing is not included although an allowance has been made in layouts and electrical design

for installation should any significant quantity of Salman primary ore be treated.

1.4.3 Metallurgical Testing

The approach taken to the metallurgical testwork was to compile master composite samples,

representing the majority of the known in pit resource established in the previously prepared Scoping




7

Study (2006). Variability composite samples were also compiled from appropriate combinations of

sample reserves used to form the master composite samples.

Each of the master composite samples was subjected to gravity/leach testing to assess the potential

for gravity concentration and determine the optimum leach conditions. The variability composite

samples were then subjected to testwork using the optimum conditions established from the mastercomposite sample testwork. The rationale was that the results obtained on the master composite

samples, rather than any of the minor components of the total resource, should dictate the process

design parameters.

The complete testwork programme comprised the following:

• Unconfined compressive strength (UCS) determinations.

• SMC Drop-Weight Testwork.

• JK Drop Weight Testwork (for SAG mill amenability).

• Bond Abrasion Index (Ai) Determinations.

• Bond Rod Work Index (RWi) Determinations.

• Bond Ball Work Index (BWi) Determinations.

• Head Assay Analysis.

• Cyanidation Optimisation Testwork.

• Carbon Adsorption Testwork.

• Cyanide Destruction Testwork.

• Thickening and Viscosity Testwork on Slurries.

• Arsenic precipitation.

• Tailings Consolidation.

•

Tailings Geochemistry.

1.4.4 Testwork Summary

The metallurgical treatment route selected based on the results of the testwork can be summarised as

follows:

• Primary crushing using a jaw crusher;

• Single stage SAG milling;

• Gravity concentration of a portion of cyclone feed using a centrifugal concentrator;

•

A thickening stage capable of being run in a leach feed or leach tail configuration;• Carbon in leach (CIL); and

• Zadra stripping plant.

The key process design parameters derived from the testwork are set out below in Table 1-2. Plant

design has been based on design maximum recoveries for each of the ore deposits based on the

selected treatment route, as set out in Table 1-3.




8

Table 1-2 Key Process Design Parameters

Mill Circuit Design

UCS MPa <180

Bond Rod Mill Work Index kWh/t 14.7

Bond Ball Mill Work Index kWh/t 13

Optimum Mill Product Size, P80 microns 106

Leach Feed Thickening

Settling Rate t/m2/h 1.00

Flocculant Consumption g/t ore 20.0

Underflow Density % solids 50

Leach/CIP Design

Residence Time hours 30

Pulp Density (feed) % solids 50

Cyanide Addition Rate kg/t 2.0

pH adjusted with lime 10 - 10.5

Lime Requirement (>90% CaO) kg/t 2.0

Table 1-3 Metallurgical Recoveries

Deposit Ore Type Option 5

Recoveries %

Anwia Oxide 95.8

Transitional 93.3

Primary 91.8

Salman Oxide 86.8

Transitional 74.7

Primary 54.7

1.5 Mineral Processing

Since the Anwia primary ore will be the hardest ore type encountered it was initially agreed that the

front end plant design would be configured to achieve a 1.3 Mtpa throughput, at a grind of 80%

passing 75 micron, of this type of ore.

Project optimisation following the development of capital and operating costs for the Project resulted in

a series of plant options being considered which examined the effect of coarsening the target grind

size and determining the impact on Project capital costs, operating costs and gold recovery.

The five options reviewed are as follows:

Option 1: Deletion of the two stage grinding circuit and insertion of a 5.5m dia. x 6.0m EGL SAG mill

(based on Golden Pride mill size such that design time can be minimised), deletion of the

Intensive Cyanidation Reactor and insertion of a Gemini Table to treat gravity circuit

concentrates, deletion of leach feed thickening, deletion of one CIL tank, deletion of the

PSA plant, treatment of tails slurry via decant return dilution to meet <50ppm CN WAD

target and Arsenic precipitation only on supernatant solutions prior to discharge.




9

Option 2: Deletion of the surge bin and dead stockpile and insertion of direct feed from the jaw

crusher to the SAG mill deletion of the two stage grinding circuit and insertion of a 5.5m dia

x 7.32m EGL SAG mill ( correct size to meet required 1.3 Mtpa throughput), deletion of the

Intensive Cyanidation Reactor and insertion of a Gemini Table to treat gravity circuit

concentrates, deletion of leach feed thickening, deletion of one CIL tank, deletion of the

PSA plant, treatment of tails slurry via thickening and decant return dilution to meet<50ppm CN WAD target and Arsenic precipitation only on supernatant solutions prior to

discharge.


crusher to the SAG mill deletion of the two stage grinding circuit and insertion of a 5.5m

DIA * 6.0m EGL SAG mill, deletion of the Intensive Cyanidation Reactor and insertion of a

Gemini Table to treat gravity circuit concentrates, deletion of leach feed thickening, deletion

of one CIL tank, deletion of the PSA plant, treatment of tails slurry via thickening and

decant return dilution to meet <50ppm CN WAD target and Arsenic precipitation only on

supernatant solutions prior to discharge and use of a single column for both acid wash and

elution cycles.

Option 4: Deletion of the Intensive Cyanidation Reactor and insertion of a Gemini Table to treat

gravity circuit concentrates, and deletion of one CIL tank.

Option 5: Deletion of the two stage grinding circuit and insertion of a 5.5m DIA * 6.0m EGL SAG mill,

deletion of the Intensive Cyanidation Reactor and insertion of a Gemini Table to treat


Following a review of all options, ARL concluded that Option 5 provided the best return for the Project

in terms of capital and operating costs and gold recovery. Option 5 includes a front end plant design

configured to achieve 1.3 Mtpa on a mill feed blend as designated by the mining schedule at a grind of

80% passing 106 micron.

The treatment plant flowsheet is based on single stage crushing, single stage SAG milling, gravity

recovery of free gold from a portion of cyclone feed, pre-leach thickening, a single stage of leaching

and a five stage CIL circuit. Gold will be recovered by a 5 tonne Zadra elution circuit with

electrowinning of the gold onto stainless steel cathodes. The electro-deposited gold will be removed

with high pressure water sprays and smelted to a final bullion product.

The design of the treatment plant will reflect:

• A simple and robust process flowsheet based on the testwork completed and analysed;

•

Sturdy, well proven equipment;

• A control philosophy for a plant with an appropriate level of automation and remote control

facilities, supplemented by sufficient alarming and diagnostics to facilitate troubleshooting; and

• A proposed flowsheet which has been selected to suit the various orebodies associated with

the Project.

The critical characteristics of the plant design are:




10

• Inclusion of single stage SAG milling to achieve 80% passing 106 micron in leach feeds for the

average mill feed blend at 1.3 Mtpa.

• Inclusion of a gravity circuit based on testwork results indicating high gravity gold recoveries

for Anwia ores.

• Inclusion of a cyanide detoxification circuit to meet International Cyanide Code standards.

• Inclusion of an arsenic precipitation stage due to elevated arsenic levels in Salman transitional

and Anwia transition and primary ores.

The summary process flow sheet is set out in Figure 1-2. The general control philosophy is for a plant

with minimal automation. The plant will have a crusher control panel and a central mill control room

from which the status of the major electrical equipment can be monitored, and from which some of the

regulatory control loops can be monitored and adjusted. The starting and stopping of most electrical

drives will be performed at the stop/start control stations located adjacent to each drive or in the case

of major equipment, started locally or by remote from the control room.

1.6 Open Pit Mining

The selected mining method for the Project is conventional open pit mining including drilling, blasting,

loading and hauling operations carried out by a qualified mining contractor with experience in Ghana.

The Salman (including Akanko) and Anwia deposits were the subjects of an open pit mining study.

The construction of the processing plant is currently planned close to the Salman deposit. The mining

contractor will also be responsible for hauling the ore from the Anwia pit to the plant site by way of a

public road.




11

Tail

Conc

Underflow

Overflow

ROM Ore Stockpile

Primary Crusher

Surge BinCrushed OreStockpile

SAG Mill

Cyclones

Gravity Concentrator

Cyanide Leach

Leach/Carbon Adsorption (CIL)

Carbon Stripping

Carbon Regeneration

Thickener

Cyanide Destruction

Arsenic Precipitation

Tailings StorageFacility

Gold Room

Gold Bullion

Figure 1-2 Mineral Processing Flowsheet Summary




12

In the mining study update, a gold price of $800/oz was used in the pit optimisations and the

calculation of the economic cut-off grades for reserves reporting. No pit optimisations were

undertaken for the Satellite Deposits and therefore no reserves are reported for these deposits.

The plan views of the mine design and site layout for Salman and Anwia deposits are provided in

Figure 1-3 and Figure 1-4 . The depths of the proposed pits at the Salman deposit will generally varybetween 30 and 70m depending on the variable topography over a strike length of 7km. The final

open pit at Anwia will be developed in three major stages to a depth of 180m, some 30m deeper than

the previous design

The conventional excavator and truck mining will be performed on 3 m mining benches while drilling

and blasting operations will be based on 6 m benches. The grade control operations in the pits will be

based on RC drilling and sampling practice well ahead of the mining front to allow short and medium

term production planning. The ore from the pits will be hauled to the ROM pad and re-handled by

front-end loader into the ROM bin. The low grade ore will be stockpiled at the designated areas for

treatment at the end of the mine life. The final production schedule allows for the filling of 30% of the

waste cost-effectively back into the mined Salman pits.

1.7 Ore Reserve Estimate

The Project ore reserves are classified within the confidence categories in Table 1-4. The reserve

statement in the table complies with the AusIMM JORC guidelines and satisfies the requirements of

National Instrument 43-101 of the Canadian Securities Administrators as set forth in Section 2 of this

report.

The cut-off grades for the mineral reserves vary depending on the degree of oxidisation degree, host

rock type and deposit areas as specified in Section 18.7. The cut-off grades generally vary between

0.7g/t and 1.1g/t gold for majority of the reserves.




13

Figure 1-3 Salman Mine Design




14

Figure 1-4 Anwia Mine Design

Table 1-4 Ore Reserve Estimate

Proven Reserve Probabl e Reserve Total ReservePit Stage

T*1000 Au g/t T*1000 Au g/t T*1000 Au g/t Anwia 1 1,381 2.01 35 1.60 1,416 2.00

Anwia 2 1,645 2.15 267 2.44 1,912 2.19

Anwia 3 2,327 2.10 1,238 2.32 3,565 2.18

Anwia North 144 1.48 39 1.56 183 1.50

ANWIA TOTAL 5,497 2.08 1,579 2.31 7,076 2.13

Akango N2 112 1.41 7 1.69 119 1.43

Akango N 92 2.09 1 1.17 93 2.08

Akango 539 1.54 74 1.70 613 1.56

Akango South 181 1.66 23 1.77 204 1.67

Salman North 719 2.20 58 2.20 777 2.20

Teberu Footwall 138 2.20 28 2.23 166 2.21

Nugget Hill 193 2.45 69 2.11 262 2.36

Salman Central 1 973 2.73 10 2.06 983 2.72

Salman Central 2 681 2.66 8 1.70 689 2.65

Salman South 1 664 1.30 7 0.96 671 1.03

Salman South 2 197 1.12 24 0.98 221 1.10

Salman SW 133 2.07 11 2.83 144 2.13

SALMAN TOTAL 4,622 2.09 320 1.91 4,942 2.08

PROJECT TOTAL 10,119 2.08 1,899 2.24 12,018 2.11




15

1.8 Financial Analysis Results

Only the changes in the parameters and the study results have been included in this revised technicalreport.The details of the change in parameters, pit optimisation results, pit designs and production

schedules have been provided in the appendices. The reader should refer to the November 2007

and December 2007 technical reports for further details of the mining study and the financial analyses.

1.9 Current Project Status

1.9.1 Mining Lease and License

The development of the Project mine requires the grant of one or more mining leases. Application for

a mining lease in Ghana requires completion of a feasibility study to the satisfaction of the GhanaMinerals Commission. ARL was granted Mining Leases over the Aniwa and Salman Deposits by the

Minerals Commission in April 2008.

In conjunction with the lodgement of a feasibility study with the Minerals Commission, ARL is also

required to lodge an environmental and social impact statement (ESIS) and Resettlement Action Plan

(RAP) for the relocation of Salman Village, to the Ghana Environmental Protection Agency (EPA).

EPA approval of this documentation must be received in order for the Minerals Commission to grant

the mining lease. The ESIS and RAP have been submitted in relation to the Salman Deposits. In

relation to the Anwia Deposit, the ESIS has been submitted and the RAP is being finalised for

submission.

ARL are also in the process of negotiating a Project Stability Agreement with the Ghana Minister ofFinance which will set out guidelines for the payment of royalties, taxes and duties etc, during the

development and operational phases of the Project. Initial discussions have already been held with

the Finance Department with regard to establishing the terms upon which ARL will draft the proposed

agreement.

1.9.2 Project Implementation Plan

The recommended development methodology for the design and construction management of the

Project is the EPCM approach, thus allowing ARL, as owner, to maintain control of the budget,

schedule and quality of the end product through all stages of project development. The Project capital

cost estimate has been developed on the basis that a single organisation (the Engineer) will providethe EPCM services with the assistance of specialist sub-consultants as required.

ARL will establish a management team to manage all aspects of Project development. The ARL team

will manage and liaise with the engineer delivery of the EPCM services during construction, and will

also be responsible for the various preproduction activities not included in the EPCM services

including:

• Implementation of the contract mining operation;




16

• recruitment;

• establishment of operating systems;

• training; and

• permitting and statutory liaison.

The project schedule indicates that 22 months will be required from the commencement of EPCM

services until “practical completion” (ie, completion of pre- and wet commissioning of the plant). A

further one month has been allocated to the completion of ore commissioning. The sourcing and

employment of the ARL operations personnel and the tender and award of the mining contract and the

mobilisation of the selected contractor will be undertaken during the 23 month schedule period. The

schedule is based on specific design requirements, vendor nominated manufacturing and delivery

periods and in-house experience with similar projects.

1.10 Conclusion and Recommendations

It is currently expected that the implementation of the Project schedule will commence followingconfirmation of the availability of grid power. First gold production for the Project is expected to occur

within 2 years of commencement of construction.

The Project financial analysis supports the proposed future development of the Project, subject to

availability of grid power. However, exploration and development activities in 2008 should focus on

further enhancing the economics of the Project by endeavouring to add additional ore reserves and

reducing, where possible, the estimated capital costs, by further focused exploration and the

examination of alternative plant options where possible.




17

2.0 INTRODUCTION AND TERMS OF REFERENCE

2.1 Terms of Reference

In October 2007, Mining Solutions Consultancy Pty Ltd (MSC) prepared a Technical Report on the

Project for ARL which complied with ARL’s disclosure and reporting requirements set forth in the

National Instrument 43-101, Companion Policy 43-101CP, and Form 43-101F1. The

Technical Report was filed in Canada with the British Columbia Securities Commission and the

Alberta Securities Commission in November 2007. In December 2007 MSC prepared a revised

Technical Report on the Project for ARL which complied with ARL’s disclosure and reporting

requirements set forth in the National Instrument 43-101, Companion Policy 43-101CP, and Form 43-

101F1. The revised Technical Report was filed in Canada with the British Columbia Securities

Commission and the Alberta Securities Commission in December 2007.

This update to the December 2007 Technical Report incorporates new data into the mining and

resource sections of the previously lodged Technical Report. The principal components of the Updateare an Update Study of the open pit mining components of the feasibility study completed in April

2008 (Dincer; 2008) and a Project Resource Update completed in January 2008 (Hellman and

Schofield; 2008a) and the Summary Resource Report for the Satellite Deposits (Hellman and

Schofield; 2008b) .

This updated Technical Report complies with Canadian National Instrument 43-101, for the ‘Standards

of Disclosure for Mineral Projects’ (the Instrument) and the resource and reserve classifications

adopted by CIM Council in August 2000. The report is also consistent with the ‘Australasian Code for

Reporting of Mineral Resources and Ore Reserves’ of September 2004 (the Code) as prepared by the

Joint Ore Reserves Committee of the Australasian Institute of Mining and Metallurgy, Australian

Institute of Geoscientists and Mineral Council of Australia (JORC).

2.2 The Purpose of this Report

This report has been prepared to comply with ARL’s disclosure and reporting requirements set forth in

the National Instrument 43-101, Companion Policy 43-101CP, and Form 43-101F1.

2.3 Qualifications and Experience

This Technical Report has been authored by Mr Ron Heeks B.App Sc. (Geology) a member of the Australian Institute for Mining & Metallurgy (AusIMM). Mr Heeks has over 25 years experience in the

mining industry with more than 15 years experience in the calculation of mineral resources and ore

reserves. Mr Heeks visited the Southern Ashanti Gold Project on 14 February 2008 for 13 days for

the purpose of reviewing deposit geology and resource drilling data

.




18

2.4 Principal Sources of Information

The listing of the principal sources of information of this Report is provided in Table 2-1. This Report

has been compiled based on information available up to and including 31 December 2007.

Table 2-1 Principal Sources of Information

Activi ty Company

Pit Optimisation, Pit Design, Mining Schedule andReserve estimation

Mining Solutions Consultancy Pty Ltd (MSC)

Drilling, Data Verification, Geology and ResourceEstimation

Hellman & Schofield Pty Ltd (H&S)

Metallurgical Testwork Ammtec Pty Ltd (Ammtec)

Metallurgical testwork supervision and assessmentof results.

Kentgold Holdings Pty Ltd t/a Ozmet (Ozmet)

Plant and Infrastructure Engineering, Capital andOperating Cost Estimates Lycopodium Engineering Pty Ltd (Lycopodium)

Environmental Baseline Studies and preparation ofESIS and Resettlement Action Plan

SGS Environment (SGS)

Tailings Storage Facility Design, Plant Geotechnicalassessment and overall Water Balance assessment

Knight Piesold Consulting (Ghana and Australia) (Knight Piesold)

Pit Geotechnical Assessment George, Orr and Associates Pty Ltd (GOA)

Hyrogeological Assessment Knight Piesold Consulting (RSA) (KPRSA)

Project Management and Financial Analysis Adamus Resources Limited (ARL)




19

3.0 RELIANCE ON OTHER EXPERTS

In addition to the authors of this report, the updated ore reserve and mineral resource estimates for

the Project rely on the data, information and recommendations provided by the consultants listed inTable 3-1.

Table 3-1 Reliance on Consultants

Activit y Company Secti ons

Pit Optimisation, Pit Design, Mining

Schedule and Reserve estimation

Mining Solutions Consultancy Pty Ltd 1.1, 1.6-1.10,

17, 18.1, 18.2,

19 and 20

Drilling, Data Verification, Geology and

Resource Estimation

Hellman & Schofield Pty Ltd 1.3, 6-13 and

17

Plant and Infrastructure Engineering,

Capital and Operating Cost Estimates

Lycopodium Engineering Pty Ltd 1.5, 16 and

19.1Metallurgical testwork supervision and

assessment of results.

Kentgold Holdings Pty Ltd t/a Ozmet 1.4 and 15

Metallurgical Testwork Ammtec Pty Ltd 15

Hyrogeological Assessment Knight Piesold Consulting (RSA) 16.10

Comminution Circuit Design Oreway Mineral Consultants Pty Ltd 16

The author of this Report is not qualified to provide comment on the legal issues set out in Sections

1.2, 4.2 and 4.3 of this Report. The information contained in Sections 1.2, 4.2 and 4.3 of this Report

is based on reports prepared by ARL’s Ghanaian legal counsel, Bentsi-Enchill Letsa & Ankomah, to

April 2007 and subsequent correspondence between ARL and the Ghanaian Minerals Commission.

The author of this Report is also not qualified to provide comment on the environmental matters set

out in Section 18.3 of this Report. The information contained in Section 18.3 is based on reports

prepared by Mr Andrew Hester BSc(Hons) MPhil (Environmental Management) Pr Nat Sci (South

Africa). Mr Heeks does not undertake or accept any responsibility or liability in any way whatsoever

to any person or entity in respect of those parts of this document, or any errors in or omissions from it,

whether arising from negligence or any other basis in law whatsoever.




20

4.0 PROPERTY DESCRIPTION AND LOCATION

4.1 Project Area, Location and Access

The Project is centred on Latitude 5º00’N and Longitude 2º14’W in the Western Region of Ghana,

West Africa. The Project site is located in the south-west corner of the Western Region of Ghana and

is approximately 280 km west of the capital, Accra, and less than 20 km from the coast at Essiama

(Figure 4-1). The Project area is accessed from Accra via the main coast highway to Takoradi and

from there by sealed road to the village of Teleku Bokazo and then by 10 kms of all weather gravel

road. The Project area is approximately 464 km2.

Figure 4-1 Project Location




21


Adamus Investment

Enterprises Pty Ltd

African Gold

Properties Ltd

Castlegem

Pty Ltd

Adamus Holdings

Pty Ltd

Adamus Resources

Limited (Ghana) (1)

Semafo (Ghana)

Ltd(1)

Nkroful Mining

Limited(1)

Akanko Mining

Limited(2)

Ghana

BVI Australia

4.2 Ownership

ARL owns those tenements comprising the Project area by way of the company structure shown on

Figure 4-2. The tenements comprising the Project area are subject to a statutory 10 per cent interest

retained by the Government of Ghana upon commencement of production.

Figure 4-2: ARL Corporate Structure

Notes:

1. Ghanaian government holds statutory right to a 10% interest in the Ghanaian subsidiaries upon commencement of

production.

2. A third party holds 1,000 shares in Akanko Mining Limited representing 11% of the outstanding shares, so that ARL

has an 89% interest (8,000 shares) in Akanko Mining Limited which will be reduced to 80% on conversion of the

Ghanaian government’s 10% interest.

ARL holds interests in three granted prospecting licences, four granted mining lease and five

prospecting licence renewal applications covering a combined area of approximately 464 km2 (Table

4-1). ARL holds 100 per cent interest in the tenements set out in Table 4-1, subject to the right of the

Ghanaian government to retain a 10 per cent interest upon commencement of production, except for

the Akanko prospecting licence which is held by Akanko Mining Limited which is owned as to 89 per

cent by ARL and 11 per cent by a third party.

The Anwia and Salman gold deposits are located in the central part of the project area (Figure 4-3).

Property boundaries are located by description using latitudes and longitudes.




22

Table 4-1 Tenure Summary – Southern Ashanti Gold Project

Name Holder Area

km2

Granted Comments

Salman

Prospecting

Licence

Adamus Resources

Ltd (Ghana)

48 30/06/2005 Granted to BHP Minerals in 1994, transferred to AGR in

1999 and reduced by 50% in 2002. In 2002 the licence

was transferred to AGP which subsequently transferred it

to Adamus Resources Ltd (Ghana) in 2004. This

prospecting licence was issued to Adamus Resources Ltd

(Ghana) in 2005 following consolidation of the then existing

Salman PL with Adamus' Tumentu PL and parts of

Adamus' Ankobra and Ankobra River PLs.

Salman Mining Licence area covering 26 square kilometres

was excised in April 2008.

Salman

Mining

Licence

Adamus Resources

Ltd (Ghana)

26 11/04/2008 Granted to Adamus Resources Ltd (Ghana) in 2008.

Excised from the Salman Prospecting Licence.

AnkobraProspecting

Licence

Adamus ResourcesLtd (Ghana) 10 14/10/1994 Initially granted to BHP Minerals in 1994, transferred to AGR in 1999 and reduced by 50% in 2001. In 2002, AGR

transferred the license to AGP which subsequently

transferred it to Adamus Resources Ltd (Ghana) in 2004.

Part of this licence was subsequently assigned to and

consolidated with the Salman PL in 2005.

Ankobra River

Prospecting

Licence

Adamus Resources

Ltd (Ghana)

28 29/11/2004 Granted to Adamus Resources Ltd (Ghana) in 2004. Part

of the licence was subsequently assigned to and

consolidated with the Salman PL in 2005.

Apa Tam

Prospecting

Licence

Adamus Resources

Ltd (Ghana)


AsantaProspecting

Licence

Adamus ResourcesLtd (Ghana)


Akanko

Prospecting

LIcence

Akanko Mining Ltd 26 24/02/1995 Granted to Tropical Exploration and Mining Co Ltd

("TEMCO") in 1995. In 2002 TEMCO entered into a joint

venture with Hightime Investments Pty Ltd (Hightime),In

2003 Hightime assigned its interest in the joint venture to

Adamus Holdings Pty Ltd (Adamus Holdings).. The joint

venture agreement between TEMCO and Adamus

Holdings provided for (i) an initial option fee of US$15,000

followed by payment of US$10,000 annually by Adamus

Holdings to TEMCO until a decision to mine was made: (ii)

minimum expenditures of US$100,000 in year one andUS$200,000 over years 2 and 3 whereupon Adamus

Holdings would earn an 80% interest TEMCO would have

a 10% free carried to decision to mine and Adamus

Holdings would have an exclusive option to purchase

remaining 10% interest from TEMCO for USD$1 per

Proven Resource ounce, 2.5% net smelter return or

exploration expenditure over 3 years commencing

December 12, 2002. Adamus Holdings was acquired by

ARL in 2004. Adamus Holdings earned 80% under the




23

11/04/2008

joint venture agreement in 2004, and the Akanko

Prospecting Licence was assigned to Akanko Mining Ltd (a

company owned 89% by Adamus Holdings and 11% by

TEMCO) in 2005.

Akanko Mining Licence area covering 17 square kilometres

was excised in April 2008. Amended PL application lodgedfor area remaining after excision of ML

Akanko Mining

Licence

Akanko Mining Ltd 17 11/04/2008 Granted to Akanko Mining Ltd in 2008. Excised from the

Akanko PL.

Name Holder Area

km2

Granted Comments

Ebi - Teleku

Bokazo

Prospecting

Licence

Semafo (Ghana) Ltd 31 4/01/1996 The Ebi - Teleku Bokazo prospecting licence initially

comprised two licences, the Teleku Bokazo prospecting

licence issued to Super Paper Products Limited (SPPC) in

1995, and the Ebi prospecting licence granted to Amuanyi

Co Ltd in 1996. Both licences were acquired by Semafo(Ghana) Ltd in 1997 and approval to merge the two

prospecting licences to form the Ebi - Teleku Bokazo

prospecting licence was granted by the Minerals

Commission in 1998. In 1998 Semafo (Ghana) Ltd entered

into a joint venture with Samax Gold Inc.in respect of the

area covered by the licence.. Samax withdrew from the

joint venture in 2000 and ARL acquired Semafo (Ghana)

Ltd in 2004.

Ebi-Teleku Bokazo Mining Licence area covering 31 square

kilometres was excised in April 2008. Amended PL

application lodged for area remaining after excision of ML

Ebi - TelekuBokazo Mining

LIcence

Adamus ResourcesLtd (Ghana)

50 11/04/2008 Granted to Adamus Resources Limited (Ghana) in 2008.Incorporates parts of the E-TB and Asanta Prospecting

Licences.

Anwia South

Mining Licence

Nkroful Mining

Limited

49 29/03/2006 Granted to Nkroful Mining Limited for a period of 10 years.

Nkroful Mining was acquired by ARL in 2006.

Mfuma

Prospecting

LIcence

Nkroful Mining

Limited

30 03/04/2007 Granted to Nkroful Mining Limited for a period of 2 years.

Nkroful Mining was acquired by ARL in 2006.




24

Figure 4-3: Tenure Perimeter, Geology and Deposit Locations




25

4.3 Description of Licences and Approvals

4.3.1 Anwia Deposit

The Anwia deposit lies on the Ebi-Teleku Bokazo prospecting licence PL2/192 held by Semafo

(Ghana) Limited, a wholly owned subsidiary of ARL. Application was made for renewal of the

prospecting licence prior to its expiry on 2 January 2008. Subsequent to this PL application, ARL

applied for a Mining Licence over an area within the existing PL. The Ebi-Teluku Bokazo Mining

Licence was granted on 11 April 2008. ARL have lodged an amended PL application for a renewal of

the Ebi-Teluku Bokazo Prospecting Licence. The application is for a two year renewal of the area

remaining outside the Mining Licence

4.3.2 Salman Deposi ts

The Salman and Akanko gold deposits lie on the Salman Consolidated prospecting licence and the

Akanko prospecting licence, respectively.

The Salman consolidated prospecting licence PL2/193 was granted to Adamus Resources Limited

(Ghana) for a period of two years commencing 30 June 2005 replacing and consolidating several

fragmented prospecting licences previously held by ARL. ARL lodged an application for an extension

to the license and a one year extension was granted in March 2008. In March 2008, ARL applied for a

mining licence over an area within the existing PL. The Salman Mining Licence was granted on 11

April 2008. After expiry of the Salman PL in March 2009, ARL will lodge a two year renewal application

over the area remaining outside the Mining Licence.

The Akanko prospecting licence PL2/128 is held by Akanko Mining Limited, a joint venture company

89 per cent owned by Adamus Holdings Pty Limited (a wholly owned subsidiary of ARL) and 11 percent by Tropical Mining and Exploration Ltd (TEMCO). Application was made to renew the Akanko

Prospecting Licence prior to its expiry on 10 January 2008. Subsequent to this PL application, ARL

applied for a Mining Licence over an area within the existing PL. The Akanko Mining Licence was

granted on 11 April 2008 and Adamus lodged an amended PL application. The application is for a two

year renewal of the remaining area outside the Mining Licence.

4.3.3 Satellite Deposi ts

Bokrobo and Aliva deposits lie in the Anwia South Mining Licence. The licence was granted to Nkroful

Mining Limited for a period of 10 years in March 2006. Nkroful Mining was acquired by ARL in 2006.

Avrebo deposit lies in the Apa Tam Prospecting Licence granted to Adamus Resources Ltd (Ghana) in2005.

Nfutu deposit lies in both the Ebi-Teluku Bokazo Mining Licence granted on 11 April 2008 and the

Asanta Prospecting Licence granted to Adamus Resources Ltd (Ghana) in 2004.




26

4.3.4 Royalties and Other Agreements

Parts of both the Salman consolidated and Ebi-Teleku Bokazo prospecting licences are subject to

concessions in favour of Super Paper Products Company that permit the extraction of kaolin clays and

small-scale mining of kaolin is taking place near the village of New Aluku. To ARL’s knowledge the

kaolin concessions do not impinge on any area proposed for the development of the Project mine or

its associated infrastructure.

A royalty of between 3-6% of the Project revenue is payable to the Ghanaian government. However,

historically no operating mine has paid in excess of 3%. In addition, a royalty of 1 per cent of gold

recovered or 3 per cent of net profits, (in each case, in relation to ore derived from the area of the

original Teleku Bokazo prospecting licence) whichever is greater is payable to Super Paper Products

Company, a previous holder of the original prospecting licence.

The Ghanian government retains the right to retain 10 per cent interest in any or all tenements upon

commencement of production.

4.3.5 Environmental Liabilit ies

There are no environmental liabilities associated with any tenements within the SAGP at this time.

4.3.6 Extension Appli cation

In order to renew the granted prospecting licences an application which includes a technical report on

the work undertaken during the previous licence period, an application fee for renewal, advance

payment of the annual rent for the licence area, and a programme of exploration work for the

extension period must be submitted.

Submitted applications are assessed by a government committee. In general, and provided the

application is lodged with all required documentation, fees and other payments, the renewal is

accepted and the formal notification follows. ARL has lodged the technical reports and application

documentation in respect of those licences requiring renewal or partial renewal for submission by the

applicable due dates.

4.3.7 Grant of Mining Lease

The development of the Project mine requires the grant of several mining leases. Application for a

mining lease in Ghana requires completion of a feasibility study to the satisfaction of the Ghana

Minerals Commission.

In conjunction with the lodgement of the feasibility study with the Minerals Commission, ARL lodged

an environmental and social impact statement (ESIS) and resettlement action plan (RAP) for the

relocation of Salman Village, to the Ghana Environmental Protection Agency (EPA). EPA approval of

this documentation must be received in order for the Minerals Commission to grant the mining lease.

The Salman, Akanko and Anwia Mining Licences were granted by the Minister of Lands, Forestry and

Mines on 11 April 2008. ARL has the exclusive rights to work, develop and produce within the Mining

Licence areas for an initial ten years. The initial term is renewable.




27

4.3.8 Project Stability Agreement

ARL is also in the process of negotiating a Project Stability Agreement with the Ghana Minister of

Finance which will set guidelines for royalties, tax rates, and duties etc, during the development and

operational phases of the Project. Initial discussions have already been held with the Finance

Department with regard establishing the terms on which ARL will prepare a draft Project stability

agreement.

5.0 ACCESSIBILITY, CLIMATE, INFRASTRUCTURE AND

PHYSIOGRAPHY

5.1 Access

The Project site is accessed from Accra via a sealed road to Teleku Bokazo via Takoradi, and then by

10 kilometres of all weather roads. The sealed coastal highway linking Ghana and Cote d’Ivoire

passes through the southern edge of the Project, and a series of sealed and formed gravel roads

linking the coastal highway at Essiama with the regional mining centre Tarkwa, pass through the

centre of the Project area. Numerous unformed gravel roads link villages throughout the project

concessions. Access within more remote portions of the Project is restricted to footpaths and cut-

lines, and can become difficult during the peak of the monsoonal seasons (May to July and October to

November).

5.2 Climate

Ghana lies just north of the equator and the climate is tropical, particularly in the southern half of the

country. Seasonal temperature variations are minor. Daytime temperatures are high throughout the

year, reaching about 30º C on most days. Diurnal variation is about 6 to 10º C in the humid south and

somewhat larger in the drier northern areas.

In the Project area the climate is fairly typical of that for south-west Ghana, with wet seasons from

March to July and from September through October with a dry season between December and

February. The Project area has an average annual rainfall of 2,023 mm with an average humidity of

80 per cent. Annual evaporation is approximately 2,850 mm.

5.3 Topography, Elevation and Vegetation

The Project lies in the Nzema East District of the Western Region of Ghana. The Project is located in

hilly terrain dissected by broad, flat drainages that typically form swamps in the wet season between

May and late October. Hill tops are generally at very similar elevations, reflecting the elevation of a

previous erosional peneplane that is now extensively eroded. Drainages in the Anwia and Salman

areas are between 10 and 15 metres above sea level. Maximum elevations are around 80 metres

above sea level but the areas impacted by the Project generally lie at less than 50 metres elevation.




28

Despite the subdued topography, hill slopes are typically steep. Ecologically, the Project area is

situated in the Wet Evergreen Forest Zone.

Vegetation cover is dominantly secondary forest and fallowed farming areas rapidly return to scrub

cover in the slash and burn farming cycle typically employed for subsistence crops. Dominant

subsistence crops are cassava and maize, grown in small allotments with no formal boundaries. Thedominant commercial crops are coconuts and oil palm and to a lesser extent, cocoa is grown in

isolated areas.

The Project site is largely transformed and has experienced extensive degradation in recent years.

The main land uses include secondary forest, subsistence and cash crop farming and artisinal mining.

There are several local villages near the Project site, the closest being the Salman and Anwia villages.

5.4 Local Infrastructure

The Salman Trend (containing the Salman deposit and the Akanko deposit) lies immediately west ofSalman and Akanko villages, and a periodically graded gravel road linking Teleku Bokazo with the

sealed Agona-Tarkwa road cuts across the mineralised trend between the Salman Central and Nugget

Hill deposits. A formed gravel road links the Salman and Akanko villages and runs parallel to the

Salman Trend providing ready access over approximately 3.5 kilometres. The Anwia deposit is

approximately 500 metres north of the town of Teleku Bokazo and is within 1 kilometre of the bitumen

road connecting Teleku Bokazo with Essiama on the southwest Ghana coastal highway. Several

kilometres of bulldozed exploration tracks have been established to provide access to the remainder

of the Salman Trend, Anwia, Nfutu and various less advanced prospects.

The Project is centred 80 kilometres west-northwest of Ghana’s major export port of Takoradi. An

electricity power relay station connected to the national grid is located between the townships of

Nkroful and Essiama, within 14 kilometres of the Salman Trend and 500 metres of the Anwia deposit.

Abundant processing water is likely to be available from both the Ankobra River and ground-water

resources.

6.0 EXPLORATION AND MINING HISTORY

6.1 Historical Mining Activity

Small-scale colonial and artisanal (galamsay) gold workings were widespread throughout the Project

area, the most being significant being at Akanko and along the Salman Trend, Anwia and Nkroful.There were several gold dredging operations in the Ankobra River between 1900 and 1920, and

reference is made to numerous small alluvial and hard rock gold workings in various Gold Coast

Geological Survey annual reports between 1930 and 1940, and in Junner (1935). Sporadic hard rock

mining commenced at Akanko in 1881, culminating with the efforts of Finsbury Pavement Financial

Trust Ltd. in 1934-1935.

Several shafts were sunk into the crest of the low ridge northwest of Akanko village and at least three

adits driven from the foot of the Akanko ridge. Some of the shafts reached at least 56 metres depth

and there was more than 250 metres of underground development along the mineralised quartz reef




29

(Tropical Exploration and Mining Co Ltd, 1997). Gold production has been estimated at several

thousand ounces although there is no official record (Tropical Exploration and Mining Co Ltd, 1997).

Artisanal alluvial and reef mining activity continues intermittently at various locations including Nkroful

(hard rock), Anwia (hard rock), Bokrobo (hard rock) and Salman (alluvial) to the present day.

Secondary manganese mineralisation was identified east of Salman village during 1952-1953 andinvestigated via pitting and trenching by the African Manganese Company Limited, which was then

mining the Nsuta manganese deposits near Tarkwa. The Salman manganese deposits are relatively

small and have not been developed.

6.2 Exploration and Ownership History of Salman Deposit Area

The Akanko area of the Salman deposit was held by Ghanorcan Resources Ltd in the 1987-1989

period. Exploration activity included geophysics, soil and rock chip sampling, and trenching, but was

limited to the immediate vicinity of the old Akanko mine (Tropical Exploration and Mining Co Ltd,

1997).

Tropical Exploration and Mining Company Limited (TEMCO) took up a prospecting licence, the

Akanko PL, covering the old Akanko mine in 1995. Ghana Manganese Corporation (“GMC”) held a

concession over the area of the Salman deposit (the Salman-Aboaji prospecting licence) in the early

1990’s and engaged TEMCO to explore the area for manganese and gold. Between 1992 and 1997

TEMCO completed an extensive soil sampling program (2,716 samples assayed for gold, arsenic,

copper, antimony and manganese) and minor stream sediment and rock chip sampling over the

Salman area, identifying a strong (>200 ppb Au, peaking at c. 9 g/t Au) north-northeast trending gold-

arsenic anomalous zone extending over at least 8 kilometres and including scattered colonial and

artisanal gold workings which is now referred to as the Salman Trend.

Three parallel but lower level and less continuous gold-arsenic anomalous zones were also

recognized. TEMCO excavated and sampled 97 trenches across the Salman soil anomaly and

selected adjacent anomalies, intersecting numerous broad intervals of moderate grade gold

mineralisation (up to 86m at 5.60 g/t Au) associated with quartz veins in saprolitic and lateritic regolith

over a strike length of at least 8 kilometres of the Salman Trend. Additional activities included pitting

(51 pits) and banka drilling (14 holes) for alluvial gold, and pitting of the manganese deposits east of

Salman.

BHP Minerals (BHP) acquired and joint ventured into a number of concessions in the mid-1990s and

undertook a variety of activities including soil sampling, acquisition and processing of Landsat

imagery, and magnetic and GeoTEM surveying focusing on the previously identified Salman Trend

and parallel features. The tenor of mineralisation at surface encountered by TEMCO along the

Salman Trend was confirmed and some further lower level anomalies identified along strike to the

north of the Ankobra River and on parallel features.

BHP then completed 75 drill holes, including 4 HQ diameter diamond core holes for c. 571m and 71

reverse circulation (RC) holes for 6,961 m, on 12 northwest trending traverses across the Salman

Trend. The drill traverses were spaced between 200 and 500 metres apart over a total strike length of

c. 4 kilometres. Almost all drill holes were oriented at -45° towards 300° UTM (BHP local grid west).

Significant mineralisation was encountered on all but two of the traverses and confirmed the presence

of significant gold mineralisation (such as 34m at 3.50 g/t Au from 2m in SRCH018, and 25m at 4.52

g/t Au from 34m in SRCH082) to a vertical depth of at least 80 metres along the Salman Trend. A




30

multi-lode system was implied and in some places appeared to dip west, almost parallel to the drill

holes, making sectional interpretation difficult. BHP interpreted a series of narrow, near-vertical

primary lodes which mushroomed out in the oxide zone, and estimated a resource of 1.1 Moz (Bolton

and Amegashi, 1996). The BHP estimate did not comply with currently applicable standards, is not

regarded as reliable and is not relevant to the Project study. Relevant mineral resource and ore

reserve estimates are provided in Section 16 of this Report.

The Salman prospect was acquired from BHP by African Gold Resources Limited (AGR) in 1999.

AGR completed a validation of the BHP database, undertook some additional soil sampling, trenching

and ground-based magnetic surveying, and carried out a program of mobile metal ion (MMI) soil

sampling. Hightime Investments Pty Ltd joint ventured into TEMCO’s Akanko prospecting licence in

2002 and followed up some of the previously identified soil gold anomalies with more detailed soil

sampling and channel sampling of bulldozer traverses. Hightime Investments assigned its interest in

the joint venture to Adamus Holdings in 2003. Adamus Holdings was acquired by ARL in 2004. No

further holes were drilled into the Salman Trend until ARL became involved in 2002.

6.3 Exploration and Ownership History of Anwia Deposit Area

The exploration history of the western prospects including Anwia and Nfutu involved a different group

of companies. Between 1995 and 1998 the Teleku Bokazo and Ebi Prospecting Licences were held

by Canadian mineral exploration company SEMAFO Inc. (Semafo). During this period the company

completed a systematic soil sampling survey over the entire concession area. The majority of the

survey was conducted on a 30m by 120m pattern, and four prominent gold anomalies were identified

for infill sampling, down to 15m by 60m grid in the case of the Anwia deposit. A small trenching

program was carried out, mainly over the area of the Anwia deposit where 7 trenches were completed

in October 1995. All trenches returned some intercepts over 1g/t Au. Semafo went on to drill a total of

322 RC holes (22,448 m) and 75 diamond holes (12,911.5 m) at the Anwia deposit. Results were very

encouraging but a high density of drilling in four main orientations was applied in an attempt to clarify

the apparent poor continuity within and between mineralised zones, a distribution pattern now

considered to largely reflect a poor understanding of the controls on mineralisation.

SAMAX Gold Inc. (Samax) entered into a joint venture arrangement with Semafo on the Ebi – Teleku

Bokazo property in 1998, with management of the project passing to Samax. Immediately following

the establishment of the Anwia joint venture, Samax was acquired by Ashanti Goldfields Company

Limited (AGC) and the joint venture continued to operate under the Samax name as a wholly owned

subsidiary of AGC. Two vertical RC drilling campaigns were undertaken by Samax to validate

proposed interpretations, refine the geometry and limits to mineralisation, and provide data to assist in

preparation of resource estimates. A total of 153 RC holes (9,002 m) were drilled, of which two were

completed with PQ-diameter diamond core tails of 29.3m and 48.5m length respectively. One further

diamond hole was drilled from surface with PQ core (70 m) to provide oriented geotechnical data and

representative samples for metallurgical testwork.

Semafo, Samax and AGC made several resource estimates for Anwia, ranging from 3.6 Mt at 1.29 g/t

for 147,000 oz Au (Semafo, 1997) to 2.94 Mt at 3.09 g/t for 292,000 oz Au. None of these estimates

complied with either the JORC Code or CNI 43-101 requirements.

Samax withdrew from the joint venture in 2000 and ARL acquired Semafo in 2004.




31

6.4 ARL Exploration

ARL progressively acquired its current tenements between 2002 and 2006 pursuant to the acquisitions

and joint ventures described in Table 4.1 of Section 4.2 of this Report.

Adamus initially focused its attention on delineating and quantifying near-surface mineralisation alongthe Salman Trend and adjacent structures. At the Ebi -Teleku Bokazo PL containing the Anwia

deposit and purchased by ARL in 2004, work focused on verifying and improving drill hole surveying,

geological logging and drill testing down plunge extensions to the known mineralisation. ARL’s

activities on the combined tenure from 2002 includes geological mapping, soil sampling c. 10,000

samples), trenching and channel sampling of bulldozer traverses (11,200m), heliborne radiometric,

magnetic and electromagnetic surveying (c. 340 km2), and several exploration and resource

delineation RC and diamond core drilling campaigns (c. 150,000m combined RC and diamond core

drilling ).

In early 2006 ARL commenced a full feasibility study into the development of the Project following the

results of a detailed scoping study based on earlier resource estimates. Further in-fill drilling

programmes were undertaken as part of the feasibility study completed in June 2007 and are ongoing.

Results of those programmes completed before 31 December 2007, are incorporated into the mineralresource and reserve estimates in this Revised Technical Report.

ARL commenced detailed exploration of the Satellite Deposits in 2006. At the time of the Feasibility

Study, insufficient data was available for incorporation of the Satellite Deposits into resource

categories. Results of programmes completed before 31 December 2007, are incorporated into the

mineral resource estimates in this Revised Technical Report.

6.5 Previous Resource Estimates

In addition to the resource estimates by previous property owners, Adamus commissioned consultants

to undertake resource estimates at Anwia and Salman. In August 2004 Ravensgate Pty Ltd undertook

estimates using ordinary kriging into blocks measuring 4mE x 10mN x 4mRL (Anwia) or 5mE x 5mN x

2.5mRL (Salman) constrained by mineralisation wireframes interpreted at approximately 0.5g/t Au cut-

off grade (Ravensgate, 2004). The spatial influence of extreme grades was limited by application of a

filter in the kriging process. Resultant estimated resources are listed in Table 6-1

Table 6-1 Southern Ashanti Gold Resources at 1g/t cut-off estimated by Ravensgate

Measured Indicated Inferred

Cut-off g/t Mtonnes g/t Au k oz Au Mtonnes g/t Au k oz Au Mtonnes g/t Au k oz Au

Anwia - - - 2.350 2.22 167.7 1.825 2.43 142.6

Salman - - - 5.067 2.20 358.3 0.767 1.94 47.9

Total - - - 7.417 2.21 526.0 2.592 2.29 1905.0

In February 2005 SRK Consulting undertook an update of Anwia and Salman resource estimates

using the Uniform Conditioning method to estimate recoverable resources (SRK Consulting, 2005).

SRK’s estimates are listed in Table 6-2..

Table 6-2 Southern Ashanti gold resources at 1g/t cut-of f estimated by SRK




32


Cut-off g/t Mtonnes g/t Au k oz Au Mtonnes g/t Au k oz Au Mtonnes g/t Au k oz Au

Anwia 3.045 2.39 234 2.482 2.20 176 0.794 1.74 44

Salman 2.758 2.19 194 6.871 1.90 419 2.210 1.72 122

Total 5.803 2.29 428 9.353 1.98 595 3.004 1.73 166

Updated resource estimates for, for both Anwia and Salman deposits, were undertaken by ARL in

January 2006 using multiple indicator kriging to estimate recoverable resources. Anwia estimates

used updated interpretations of mineralisation domains but informing data were nearly identical to

those available to SRK. January 2006 Salman estimates were based on revised interpretations of

mineralised domains and used additional drill hole data available to November 2005. ARL’s estimates

are listed in Table 6-3

Table 6-3 Southern Ashanti Gold Project resources at 1g/t cut-of f estimate, ARL, January 2006

Measured Indicated InferredCut-off g/t Mtonnes g/t Au k oz Au Mtonnes g/t Au k oz Au Mtonnes g/t Au k oz Au

Anwia 4.5 2.50 360 1.8 1.97 110 2.4 1.95 150

Salman 3.8 2.24 270 4.6 1.97 290 5.2 2.03 340

Total 8.3 2.38 630 6.4 1.97 400 7.7 2.01 490

In January 2007 ARL updated resource estimates for the Salman deposit also employing multiple

indicator kriging to estimate recoverable resources. Anwia estimates were reported for completeness

but were unchanged from those reported in January 2006. Salman estimates were based on revised

interpretations of mineralised domains and used additional drill hole data available to 30 November

2006. The January 2007 ARL estimates are listed in Table 6-4..

Table 6-4: Southern Ashanti Gold Project Resources at 1g/t cut -off est imate ARL, January 2007


Cut-off

g/t

Mtonnes g/t Au k oz Au Mtonnes g/t Au k oz Au Mtonnes g/t Au k oz Au

Anwia 4.5 2.50 360 1.8 1.97 110 2.4 1.95 150

Salman 7.3 2.14 500 3.2 1.89 200 3.8 1.90 240

Total 11.8 2.28 860 5.0 1.92 310 6.3 1.92 390




33

7.0 GEOLOGICAL SETTING

7.1 Regional and Local Geology

The major gold deposits in Ghana are hosted by Palaeoproterozoic rocks of the West African Craton,

which includes the Birimian Supergroup, a series of metavolcanic and metasedimentary rocks, the

Tarkwaian Group, comprising fluvial metasedimentary rocks, and various gabbroic to granitic

intrusives Figure 7-1. Gold mineralisation within the Birimian Supergroup is associated with

mesothermal quartz veins and structurally controlled, while both mesothermal shear-hosted and

palaeoplacer gold deposits occur in the Tarkwaian Group. The Project is underlain by Birimian

Supergroup rocks with minor granitic intrusions, bounded by large granitoid bodies to the west and

east, and poorly defined areas of Tarkwaian Group in the east.

Figure 7-1 Regional Geology




34

In Ghana, the Birimian Supergroup (c. 2.1-2.2 Ga) is divided into a series of narrow (typically 20-60

kilometres wide), northeast striking, laterally extensive volcanic “belts” separated by broader

sedimentary “basins” (formerly termed the Upper and Lower Birimian; Griffis et al. 2002). The volcanic

belts are dominated by massive andesitic to basaltic (tholeiitic) flows, coarse andesitic to dacitic

volcaniclastics, and locally abundant pillow lavas. Phyllite and greywacke dominate the basins, andwidely exhibit primary sedimentary structures indicative of deposition by subaqueous sediment gravity

flows. Redeposited tuffs of andesitic to dacitic composition are a feature of the basin margins, forming

packages a few metres to hundreds of metres thick intercalated with the epiclastic greywackes and

phyllites. Thin packages of chert, graphitic phyllite and fine-grained manganese-rich sediments often

mark the transitional zone between Birimian belt and basin.

The Tarkwaian Group (c. 2.1 Ga), comprising a sequence of conglomerate, arkosic sandstone and

quartzite, siltstone and shale, is in extent largely confined to fault-bound slivers and unconformable

packages within the Birimian volcanic belts. Detrital zircons from the Tarkwaian have returned ages

up to 2240 Ma but mostly in the range 2190 to 2130 Ma (Davis et al., 1994, Allibone et al., 2002a,

Griffis et al., 2002). Initially thought to be much younger than the Birimian units, regional mapping

suggests a common structural history (Eisenlohr and Hirdes, 1992). At the north end of the AshantiBelt the Tarkwaian Group is apparently intruded by the 2081±25 Ma Banso granitoid (Eisenlohr and

Hirdes, 1992, Davis et al, 1994), providing an upper age limit and bracketing deposition between c.

2081-2130 Ma.

Two main granitoid suites are recognized in Ghana, Dixcove-type and Cape Coast-type. The Dixcove

suite is generally confined to the volcanic belts and includes a range of small plutons to large

batholiths of mafic to intermediate composition. Hornblende is the dominant mafic phase and many

bodies are not foliated. U-Pb zircon dating of Dixcove granitoids indicates crystallization between c.

2135 and 2185 Ma (Hirdes et al., 1992, Boher et al., 1992, Oberthür et al., 1998). The larger Cape

Coast bodies are typically foliated (often gneissic) biotite granitioids, commonly with migmatitic

margins and prominent contact metamorphic aureoles. Cape Coast type granitoids are most

widespread in the sedimentary basins and U-Pb zircon dating indicates they are younger than theDixcove suite, crystallization occurring between c. 2116 and 2088 Ma (Hirdes et al., 1992, Boher et al.,

1992, Davis et al., 1994, Oberthür et al., 1998).

Intrusion of the Birimian mafic volcanics by Dixcove suite granitoids suggests eruption of the Ashanti

Belt before 2185 Ma, while a Sm-Nd isochron suggests metamorphism of the belt c. 2153±13 Ma

(Hirdes et al., 1992, Boher et al., 1992, Oberthür et al., 1998, Allibone et al., 2002a). Allibone et al.

(2002a) refer to this early metamorphism of the mafic belt volcanics and intrusion of the Dixcove suite

as the Eburnian I Orogeny. Detrital zircons from Birimian sedimentary rocks in the Ashanti and Sefwi

belts and Kumasi Basin have generally returned U-Pb dates in the 2130 to 2180 Ma range (Davis et

al., 1994), and in the southern Ashanti Belt up to 2260 Ma (Loh and Hirdes, 1996). Intrusion of the

Cape Coast suite granitoids provides an upper age limit of c. 2116 Ma for deposition of the Birimian

metasediments. The deformation and metamorphism of the entire Birimian Supergroup and

Tarkwaian Group and intrusion of the Cape Coast granitoid suite between c. 2116 and 2088 Ma is

widely referred to as the Eburnian Orogeny (e.g. Griffis et al., 2002), or more specifically, the Eburnian

II Orogeny (Allibone et al.,2002a).




35

The detrital zircon age range for the Birimian metasediments is almost exactly coincident with those

from the Tarkwaian Group, leading to suggestions that the Tarkwaian and Birimian sediments were

deposited contemporaneously along and adjacent to an emergent Birimian volcanic chain, the

Tarkwaian within narrow, fault bounded terrestrial basins on the Birimian volcanic belts and the

Birimian sediments in the flanking marine basins (e.g. Griffis et al., 2002). Compositionally, the

Birimian sedimentary rocks could easily have been (and most likely were) derived from Birimianvolcanic belts and the presence of redeposited tuffs within the basins indicates active intermediate to

felsic volcanism during deposition. The youngest detrital zircons indicate Birimian metasediments

were still being deposited at least 55 m.y. after eruption of some of the Birimian metabasalts (before

2185 Ma, above), and the basalts could represent the older parts of the volcanic chains which

subsequently erupted more felsic volcaniclastic material into the adjacent basins. With an age range

of 2135 - 2185 Ma the Dixcove granitoid suite could represent the metamorphosed and eroded roots

of these andesite-dacite volcanoes. However, the quartz-rich composition and presence of foliated

clasts of Birimian sedimentary rocks within “contemporaneous” Tarkwaian conglomerates (Milési et al.,

1991) indicates that many aspects of Birimian and Tarkwaian tectonic development have yet to be

satisfactorily resolved.

Metamorphic grade of the Birimian rocks is greenschist facies, with local amphibolite facies aureolesaround granitoid plutons. Recent work in the southern Ashanti region (John et al., 1999) suggests that

the greenschist facies is widely retrograde after amphibolite facies conditions. Both belt and basin

packages are highly deformed with widespread isoclinal folding and regional bedding-parallel

cleavage attributed to regional northwest–southeast compression during the peak of the Eburnian

Orogeny c. 2100 Ma. Regional northeast striking shear zones parallel to the belt margins are also

assumed to have developed during peak Eburnian and appear to be fundamentally important in the

development of the Birimian gold deposits for which Ghana is well known such as Ashanti, Prestea-

Bogosu, Konongo, and Bibiani. Adamus’ Southern Ashanti Gold Project covers the south-western

margin of the famous Ashanti Belt. The Salman Trend of gold deposits is believed to be associated

with the same belt-margin shear zones that host the Prestea, Bogosu, and Obuasi-Ashanti gold

deposits (Figure 7-1) and has many characteristics typical of these deposits. The Anwia Deposit is

located within the adjacent Birimian basin rocks, several kilometres west of the belt margin fault zone,and has a contrasting mineralisation style.




36

7.2 Project Geology

7.2.1 General

Basement exposure is generally poor within the Project and largely restricted to road cuttings, a few

stream beds, prospecting pits and trenches, and drill pads. Laterite and mottled clay zones are locally

developed on ridges, and saprolite typically extends to 10-30 metres beneath surface and locally as

much as 80 m. The eastern part of the project is largely underlain by Birimian volcanic and

volcaniclastic rocks assigned to the Ashanti Belt, the western part mainly by Birimian metasedimentary

rocks of basin and basin margin affinity in the southeastern corner of the Kumasi Basin. The Birimian

volcanics are thought to be faulted against the Tarkwaian Group immediately northeast of the current

tenure, and a small area of quartz-rich fluvial rocks immediately east of Axim may also belong to the

Tarkwaian (see also Loh and Hirdes, 1996, Griffis et al., 2002). A large biotite granite body is exposed

in the western part of the project area and probably belongs to the Cape Coast suite. Two large,

magnetically zoned probably Dixcove-type granitoid batholiths intrude the volcanics at the easternedge of the project, and curved magnetic ridges adjacent to these intrusives could represent contact

aureoles. Several narrow granitoid dykes and fault slivers up to 13 kilometres long and 700 metres

thick of uncertain affinity are scattered through the project area, and some near-circular geophysical

features (electromagnetic resistors with weakly magnetic edges) between 1.5 and 2 kilometres

diameter northeast of Anwia may represent subsurface plutons. Two north-striking dolerite dykes are

known from geophysics and drilling in the Nkroful-Anwia area.

There is no formal subdivision of the Birimian Supergroup in the Southern Ashanti area but several

lithologically and geophysically distinct units can be identified and three litho-structural domains are

recognised: Avrebo, Salman and Anwia (Figure 4-3).

7.2.2 Avrebo Domain

The Avrebo Domain encompasses the eastern part of the project area underlain by Birimian volcanic

and volcaniclastic rocks, minor Birimian greywacke and phyllite packages, and Dixcove-type intrusive

bodies. Primary layering is generally steep and strikes north-northeast to northeast. Cleavages are

not particularly well developed in the volcanic lithologies (cf. phyllite and greywacke packages) but at

least two or three weak foliations are evident in most exposures and are of similar orientation to those

of the Salman Domain (below). Scattered lenses of greywacke and phyllite within the volcanic rocks

are probably fault bounded, and a large north to northeast striking shear zone is identified within the

volcanics in the Avrebo area. The Avrebo Domain covers the southwestern edge Ashanti Belt, and

the volcanics appear to be faulted against a package of basin margin metasediments to the west

(Salman Domain). The eastern margin is undefined.




37

7.2.3 Salman Domain

The Salman Domain comprises a zone 4-5 kilometres wide immediately west the Ashanti volcanic belt

comprising near-vertical north to northeast striking metasedimentary packages separated a series of

similarly oriented mylonitic shear zones informally termed the Ankobra Fault Zone. Greywacke and

phyllite packages dominate the Salman Domain, minor lithologies include redeposited andesitic to

dacitic crystal-lithic tuffs, pebbly volcanogenic greywackes and conglomerates, and rare andesitic

dykes or flows. The Adamanso Shear Zone separates the Salman Domain from the Avrebo Domain

(belt volcanics) to the east, and the Aluku Shear Zone from metasediments of the Anwia Domain

(basin) to the west.

Four distinct cleavages and fold generations recognised. S1 is a strong graphitic cleavage parallel to

S0 and presumably developed during regional isoclinal folding, the main architectural event. Facing is

rarely identifiable in saprolitic exposures but is generally westwards. Centimetre-scale isoclinal F1

folds in S0 were observed at one locality. S 2, a thick graphitic crenulation, is generally the most

conspicuous cleavage and is axial planar to widespread metre-scale, south to southwest plunging,open to close folds. Varying asymmetry suggests the presence of large scale F2 folding, and some

100-200 metre wavelength open folds mapped along the Salman Shear zone are thought to be F2

structures (Figure 7-2). S2 generally strikes between east and northeast, dips between 50° and 90° to

the southeast, and the average F2 axis is orientated c. 65→220. S3 is a thinner and typically finer

crenulation than S2. Strike is northwest and dip is steep (average c. 80° to northeast). Metre-scale

open F3 folds in S 0 and S 2 can be observed at a few localities but are not as well developed and

widespread as F2 folds. Some of the larger scale (100-1000 m) rotation of S 2 within the Salman

Domain is attributed to medium scale F3 folds. S 4 is a weak sub-horizontal cleavage only locally

discernable associated with gentle open folding and warping of S0, S1 and S 2.

The Salman Shear Zone, host to the Salman Trend gold deposits, is the best known and explored

fault within the Ankobra Fault Zone. While the Salman Shear Zone appears the main locus of goldmineralisation, pockets of gold mineralisation have been identified on or adjacent to other faults within

the Ankobra fault set, including the Mamposo and Adamanso shear zones.




38

Figure 7-2 Salman Deposi t Geology (Including Akanko)




39

The Salman and Mamposo shear zones are defined by both geological mapping and geophysics,

while the other members of the Ankobra fault set, including the Adamanso and Aluku shear zones, are

based on S0 discontinuities defined by geophysics. Detailed mapping and geophysical interpretation

shows that Salman Shear Zone extends from the Gulf of Guinea coast between Asanta and Sawoma

for at least 20 kilometres through Salman and Akanko and north of the Ankobra River to the Bansoarea. It is then interpreted to continue for a further 30 kilometres north, along with several of the other

Ankobra faults, to merge with the Central Fault Zone of Allibone et al. (2002b) at Prestea-Bogosu:

fabrics and lithologies within the Salman Shear and Central Fault Zone are closely comparable.

Exposures of the Salman and Mamposo shear zones are characterised by the presence of tightly

folded and boudinaged greywacke beds and quartz veins within a highly deformed zone of graphitic

mylonite and phyllite up to c. 125 metres thick. The mylonitic fabric and S1 are both crenulated by S 2

indicating a D1 origin for the Ankobra fault set and development was presumably associated with

regional D1 isoclinal folding during west-northwest – east-southeast compression. Outcrop

observations indicate dextral reactivation of the Salman Shear during D2 and sinistral reactivation

during D3, but the bulk of the strain appears to be D1. Several slices of altered, S1-foliated biotite-

ilmenite granitoid, termed the Akanko Granitoid, are included within the Salman Shear: all observedcontacts are faulted (typically mylonitic) and the bodies were either structurally modified after intrusion

within the shear zone or entirely structurally emplaced.

7.2.4 Anwia Domain

The Anwia Domain is characterised by modest S0 dips over large areas and large-scale open folding

of S0. Porphyroblastic greywackes, phyllite, and kaolinitic (ex-vitric?) redeposited tuff are the dominant

lithologies: the thick kaolinitic tuff packages, widespread porphyroblasts, and generally low graphite

content distinguish the Anwia Domain lithologically from the adjacent Salman Domain. There is no

evidence yet for the presence of the large graphitic phyllite and mylonite shear zones which

characterise the Salman Domain.

The S1, S2, S3 and S 4 cleavages are of the same style as for the Salman Domain, although S 2 and S 3

are of much more variable orientation (rotated by up to 90°) indicating large-scale, open post-F 3

folding. Geophysical interpretation (especially EM) supported by a few field observations suggests a

kilometre-scale dome and basin geometry within the Anwia Domain (Figure 4-3): open, metre-scale

dome and basin folding produced by F2-F3 interference was observed in outcrop at the Anwia Deposit.

The few facings apparent suggest the sedimentary sequence is extensively overturned.

The western margin of the Anwia Domain is not defined; the eastern margin with the Salman Domain

is placed along the geophysically inferred Aluku Shear Zone. By analogy with thin-skinned fold-thrust

belts the Aluku Shear Zone could represent the oldest, basin-ward thrust in an imbricate fault zone,

and the Mamposo, Salman and Adamanso shear zones represent progressively younger thrustsformed as the Birimian basin (Anwia Domain and west) was pushed eastwards over the Birimian

volcanic belt during regional D1 compression. Similar character of S 2 to S 4 in both Salman and Anwia

domains suggests a common post-D1 deformational history, comprising sinistral (D 2) then dextral (D3)

modification and reactivation of the D1 architecture, followed by post-orogenic relaxation (D 4). The

same sequence of compression with isoclinal folding and thrust fault development, followed by

sinistral then late dextral wrenching has been proposed for the Obuasi-Ashanti area c. 150 kilometres

along strike to the north-northeast (Allibone et al. 2002a).




40

8.0 DEPOSIT TYPE AND MINERALISATION

8.1 Salman Deposit Type and Mineralisation

The Salman Trend is defined by a semi-continuous +200ppb gold in soil anomaly extending for at

least 9 kilometres along the Salman Shear Zone. Approximately 8 kilometres of strike extent has

been drill tested to date with several discrete, multi-lode gold deposits being found along the shear

zone (Figure 7-2). Significant gold deposits within the Salman Trend include, from south to north,

Salman South, Salman Central, Teberu Footwall, Salman North, North Hill, Akanko Central, and

Akanko North. Salman Central, Nugget Hill and Salman North deposits are located on conspicuous

north to north-northwest bends (left-hand flexures) up to 1000 metres long within the overall north-

northeast strike of the shear zone. The smaller (100 m) flexures in the Salman Central – Nugget Hill

area appear to be gentle F2 folds, while the broader bends may be a combination of D 1 irregularities

and F2 folds.

The Salman Shear Zone is made up of a western hanging-wall c. 10-125 metres thick of deformedphyllite and thin bedded greywacke with S-C bands and graphitic mylonite zones, and an eastern

footwall of thick bedded greywacke with minor sheared phyllite zones up to a few metres thick.

Boudinaged S0-parallel quartz veins and greywacke beds are a characteristic feature of the hanging

wall zone. Numerous metre-scale open to tight F2 folds are locally evident in both hanging-wall and

footwall, and quartz veins in a variety of orientations (below) are locally conspicuous. The shear zone

dips modestly to steeply west over much of the 8 kilometres drilled extent, but locally, such as at

Nugget Hill, North Hill and Salman South it rolls over to dip c. 60° to the east. Narrow slivers of altered

biotite granitoid are locally included within the shear zone south of Salman North. The granitoid

bodies exhibit S1 cleavage and exposures at Nugget Hill and drill core from Salman North show the

presence of graphitic mylonite faults on both western and eastern margins. The granitoid body

becomes continuous north of Salman North.

At Salman South, Central and Nugget Hill gold mineralisation occurs in vertical to west dipping lodes

approximately parallel to and splaying out from the main footwall-hanging-wall contact. Both the shear

zone and gold mineralisation roll over to dip steeply eastwards at about 100 metres depth within the

Salman South deposit. Most of the gold lodes are within the immediate footwall within quartz-veined

silica-sericite-carbonate-arsenopyrite altered greywacke and/or granite. Some narrow, shear zone-

parallel zones of gold mineralisation are present in the hanging-wall. The mineralised zone at Salman

Central gradually transgresses the shear zone from mostly footwall-hosted at the southern end to

mainly hanging-wall hosted at the northern end. At Teberu the main gold lodes are within the footwall

greywacke, approximately 100-300 metres east of the footwall–-hanging-wall contact, and presumably

associated with west dipping subsidiary shears. Both west and east dipping lodes are defined at

Salman North: west-dipping along the main shear zone on the western granite margin (hanging-wallshear zone); east-dipping parallel to the footwall shears along and adjacent to the eastern granitoid

margin (footwall faults). The highly fractured granitoid beneath the intersection of west and east

dipping hanging- and footwall is extensively mineralised. Similar west and east dipping lodes are

present adjacent to and within the granitoid body at Akanko.




41

Gold mineralisation is associated with a complex array of deformed quartz veins and arsenopyrite-

silica-sericite-carbonate alteration. Petrography, metallurgical work and field observations suggest

there are three principal styles of gold mineralisation, as follows:

• Nuggety free gold within quartz veins;

• Gold associated with fine acicular arsenopyrite disseminated (typically <5 per cent) in silica-sericite-carbonate altered wallrocks adjacent to mineralised quartz veins, and probably within

some of the grey quartz veins;

• Free particulate gold within the oxide zone, derived from the weathering of the former two

primary mineralisation types.

Most quartz veins are small (<2m thick and <10m long), and locally make up to c. 20 per cent of the

footwall and hanging-wall. At least five types of quartz vein sets are identified, all to some degree

gold-mineralised:

• S 1-parallel quartz veins, typically white, less commonly grey, extensively boudinaged parallel

to S1 and locally reduced to dislocated fragments within graphitic mylonite. These veins areparticularly characteristic of the hanging-wall, but also occur in the footwall and parallel to S1 in

the granitoid bodies. Individual boudins reach up to c. 2m thick and several metres long.

Accessory carbonate and sericite are widely present, less commonly also tourmaline. Niche

samples have returned up to 25 g/t Au and free gold observed in RC cuttings of white quartz

veins probably belonged to this set.

• Sub-vertical S 1-perpendicular quartz lenses, occurring immediately adjacent to the hanging-

wall – footwall contact, and especially within the sheared granitoid bodies. Texture and colour

range from white with S1 fracture cleavage to breccias in which white to pale grey quartz

fragments are cemented by smoky grey quartz. Accessory sericite and/or tourmaline are

widespread, the latter becoming visually more abundant, along with granitoid bodies, north of

Salman Central. Niche samples have returned up to 9.5 g/t Au.

• Steeply southwest dipping, north to northwest striking fractured grey quartz veins cross-cutting

S1, typically <0.5 metre thick and up to 10-20 metres long with widespread open F2 folds. This

set is particularly characteristic of the footwall greywacke zone. Sericite and carbonate are

widespread accessory components, locally tourmaline. Niche samples have returned up to

3.5 g/t Au.

• Southeast dipping smoky grey quartz veins, cross-cutting S 1, up to one metre thick and 220

metres long. The most conspicuous examples are in the footwall greywacke zone at the

historic Akanko mine where free gold was observed and niche samples returned up to 8.5 g/t

Au. Tourmaline alteration is particularly strong (in both veins and wall rock) at Akanko and

unpublished historic mining reports note an association between free gold and tourmaline-

bearing zones within the Akanko reef.

• Sub-horizontal, fractured white and grey quartz veins, only observed in the footwall and mostly

within the granitoid fault slivers. Sericite and carbonate are widespread accessory

components. Niche samples have returned up to 0.95 g/t Au.




42

Gold mineralisation does not appear restricted to a particular vein set or sets, and there are zones

where type 1 and 3 appear barren. The relative timing of gold mineralisation event or events has not

been satisfactorily established at Salman but a relatively late syn-D 3 origin is currently preferred. At

Salman Central the S1-parallel veins (type 1) are widespread throughout all of the hanging-wall but

only at the northern end of the deposit are they significantly mineralised. This relationship suggests

the main gold mineralisation phase occurred after formation of the D1 type 1 S 1-parallel veins. Thepresence of gold within type 3 veins suggests mineralisation during or after D 2 sinistral reactivation of

the Salman Shear Zone. Allibone et al. (2002b) proposed that mineralisation at Bogosu (50-70

kilometres along strike to the north-northeast of Salman) occurred during sinistral wrenching

(equivalent to Salman D2), with gold mineralisation concentrated principally in steeply plunging lodes

in dilational left-hand jogs and to lesser extent shallow dipping lodes in right-hand restraining bends

within the Central Fault Zone. The occurrence of the major Salman Central and North deposits on left-

hand flexures agrees, at first glance, with the D2 sinistral wrench mineralisation model. However, the

shallow dipping footwall faults and lodes at Salman Central and North are more geometrically

compatible with mineralisation on restraining-bend thrusts. The left-hand bends along the Salman

Shear would have been restraining during dextral wrenching and mineralisation may have occurred

during D3.

8.2 Anwia Deposit Type and Mineralisation

The Anwia gold deposit is hosted by a northeast dipping package of greywacke (footwall) and

interbedded greywacke-phyllite (hanging-wall). In the western (footwall) part of the deposit gold

mineralisation is also hosted by a steeply northeast dipping granite dyke that gradually converges on

the hanging-wall to the northwest (Figure 8-1). Three cleavages are present: north-northeast striking

S2, east striking S3, and locally a sub-horizontal S4. Gentle to open, metre-scale F2 folds are

widespread, and small-scale open dome and basin F2-F3 interference patterns were locally observed

in outcrop. The few facing indicators apparent suggest the metasedimentary package is overturned.

Gold mineralisation is intimately associated with pyrite disseminated within and around a complex

array of deformed pale grey to dark blue grey quartz-carbonate-sericite±albite veins. A broad silica-

sericite alteration zone about 200 metres thick and 450 metres long is developed in the footwall

greywacke sequence and in some areas obliterates primary sedimentary structure. The silica-sericite

alteration zone is more extensive than the gold-pyrite mineralisation.

The surface projection of identified mineralisation trends northwest for approximately 900 metres and

is up to 400 metres wide (Figure 8-1). Within this zone seven distinct domains of varying orientation

and style were used for the resource estimation. Most of the gold mineralisation is located in the

southeastern part of the deposit where a very broad, modestly northwesterly plunging (c. 35°) zone

transgresses the hanging-wall greywacke-phyllite sequence into the intensely silica-sericite altered

footwall greywacke unit. This broad zone passes upwards into an extensive horizontal mineralisation

zone around 50 metres beneath surface. Mineralisation becomes sporadic along trend to the

northwest until the northern end of the granite dyke is encountered. Limited drilling along the granite

dyke also indicates the presence of northeast dipping lodes parallel to the granite margins (Figure

8-1).




43

Figure 8-1 Anwia Deposi t Geology and Mineralisation




44

8.3 Satellite Deposit Types and Mineralisation

The Bokrobo deposit on the Anwia South property, 3.2 km south-southeast of Anwia (Figure 4-3)

comprises generally north-south trending, steeply west dipping auriferous quartz veins hosted by

strongly silica-iron carbonate altered, medium to coarse grained, carbonaceous greywacke. A north-

south trending dolerite dyke, dipping sub-vertically to the west cuts the depth extension of the mainvein. In the southern portion of the deposit, an east-west trending, steeply dipping, steeply south-

southeast plunging ‘dyke-like’ granitic intrusion is cut by numerous auriferous quarz veins. In the north

of the deposit, mineralisation generally occurs in a single lode. In the south, the mineralisation the

mineralisation occurs as two main lodes and a series of narrow stacked lodes around the granite

intrusion, Current interpretation has the main mineralisation occurring post-granite intrusion and pre-

dolerite intrusion. Some remobilisation occurred in favourable structural sites probably syn-dolerite

intrusion.

At Aliva, 1.6km southeast of Bokrobo, the gold mineralization occurs as a series of stacked,

moderately east-dipping lenses in the southern section of the deposit and in two narrow, steeper east-

dipping lodes separated by a barren zone in the northern section.

At Avrebo, 12km southeast of Salman deposit, the gold mineralization occurs in north-northwest to

south-southwest trending, subvertical to steeply east-dipping, strongly sericite-iron carbonate altered

lodes within metabasalt.

At Nfutu, 2.7 km east of Anwia, interpretation of the limited drilling and trenching completed has not

yet adequately defined the style or constraints on the distribution of the mineralization.

9.0 EXPLORATION

9.1 Trenching Methods

Approximately 7040m of trenches were undertaken in several campaigns by TEMCO, AGR and ARL

for inclusion in the Salman Trend resource dataset. The breakdown by company is approximately

3934m of TEMCO trenches, 1263m of AGR trenches, and 1843m of ARL trenches and bulldozer

channel traverses.

Most of the trenches (5752m) were excavated manually to depths of 1 to 3 metres, generally reaching

the mottled clays around the base of the laterite gravel and, particularly along the ridges at Salman

Central and North, the top of the saprolite. BHP and AGR deepened and resampled some of the

original TEMCO trenches, returning results of similar magnitude and confirming the validity of the

TEMCO work. Most of these manual trenches are oriented northwest-southeast (c. 300°) almostperpendicular to the overall strike of the Salman Shear Zone and mineralisation.

ARL also channel sampled 23 selected bulldozed drill lines for 1288m at Salman Central, Salman

North and Akanko where saprolitic and mottled clay zone material was exposed. The channel lines at

Salman Central and Akanko trend UTM east-west, those at Salman North approx. northwest (c. 300°).

Spacing between the trenches ranges from 40 metres to 100 metres, with a few gaps up to 1300

metres in swampy or alluvial areas such as between North Hill and Akanko South.




45

9.2 Pits

A recommendation arising from the February 2006 Salman resource estimate was to gain vertical

channel samples to better define gold grades in the near-surface, deflated laterite profile and the

uppermost saprolite. Through early 2006, 357 pits were manually excavated to depths of 3-4m on a

nominal 25mE x 50mN spacing between 551100N and 552700N (Salman South and Salman Centralareas) and between 554250N and 555550N (Salman North and North Hill areas). A total of 1316m

was sampled with vertical channel samples with lengths of individual samples between 0.5 and 1.5m,

the majority being 1m. Gold assays are available for 1319 samples.

9.3 Drilling Methods

The current Project resource estimate is based on a combined total of approximately 150 kilometres of

reverse circulation percussion (“RC”) and diamond core drilling. This breaks down to 40,120m RC and

24089.6m diamond drill core for a combined 64,209m at Anwia, and 81,084m RC and 4,740.7m

diamond drill core for a combined 85,824.8m within the Salman Trend. Almost all of the drilling within

the Salman Trend was conducted by ARL, while at the Anwia deposit the larger proportion was carriedout by Semafo and Samax. Drill spacing over the Salman resource area varies from 50m x 25m to

25m x 25m, with areas contributing the bulk of resource ounces drilled at the closer spacing. At

Anwia, the central part of the deposit is drilled to about 70 metres depth with holes spaced at 15m x

15m. Below this depth, and in peripheral areas, drill spacing averages 50m x 25m.

A further 26,712m of dri lling, all conducted by ARL, has been used in the resource estimations of the

Satellite Deposits. Drill spacings vary from 50 x 50m down to 20 x 20m.

The drilling and the associated surveying methods in the deposit areas are detailed in Section 10

below.

10.0 DRILLING

10.1 Anwia Deposit Drilling

Semafo drilled a total of 322 RC holes for c. 22,448m and 75 diamond core holes for 12,912m at the

Anwia deposit between 1995 and 1998. Most of the drilling was done by Canadian contractor, St

Lambert. All Semafo RC and diamond core holes were geologically logged, although much data was

subsequently lost when Samax later rationalized the lithological codes. Most holes were declined

between 50° and 70° and four major azimuths were drilled; UTM northeast, northwest, southeast, and

southwest using Semafo’s Anwia local grid. Semafo diamond drill core was not oriented for structural

measurements.

Two campaigns of RC drilling were undertaken by Samax at Anwia during the 1998-2000 period for a

combined 9,002m in 153 holes. The holes were on approximately 15 metre centres and all were

vertical. Two PQ diamond core tails were drilled from RC pre-collars and one PQ hole from surface,

for a total of 148 m. The PQ drill core was used for geotechnical and metallurgical work. Detailed

geological and geotechnical logs were compiled; a photographic record of the core (core blocks,

orientation marks and bulk density determination intervals) was prepared and magnetic susceptibility

measurements recorded.




46

Between 2003 and January 2006, ARL drilled 38 RC and diamond core holes at Anwia for a total of

4195m HQ and NQ diameter core and 2462m RC. Three holes were cored from surface, the majority

of diamond coring was from RC pre-collars. All drilling was undertaken by Minerex using a

multipurpose RC/DDC UDR650 drill rig with air capacity of 350psi/700cfm and a skid-mounted

diamond coring rig for some of the tails. A face-sampling RC hammer was employed with a bit

diameter of 5.25”. Air capacity of the rig was deemed inadequate for RC drilling much deeper than 80metres down hole, beyond which depth diamond core tails were generally used. RC drill holes were

typically collared with PVC pipe down to 6 metres, and recoveries from the first 3 metres were usually

quite poor. All ARL RC holes were logged on a one metre basis for lithology, weathering and

oxidation, and qualitative moisture content (dry, moist or wet) recorded. The one metre RC samples of

the cyclone were weighed using 60kg bench scales. Diamond drill core was typically extracted in 3

metre runs and fresh core was oriented using a spear or Craelius device at 6 to 12 metre intervals.

The core was placed in core trays by the drilling crew, with annotated core blocks inserted between

core runs. The core trays were then moved to the Adamus core yard in Nkroful for marking up,

geological logging, sampling and storage. The fresh core was marked with the intervals and a bottom-

of-hole line (based on spear orientation marks) prior to geological logging. Structure orientations were

then measured relative to the bottom-of-hole line. Logged diamond core recoveries within the primary

profile are typically 95-100%, in the oxide zone generally >70%.

The majority of the ARL drill holes were inclined toward the southwest, parallel to the Semafo grid, and

declinations range from -50º to -90º (dominantly -60º) designed to provide an optimal intersection

through mineralised zones. A few of the ARL holes were drilled to the northeast and southeast, again

parallel to the Semafo grid.

During 2006 and early 2007 ARL undertook further drilling:

• In January 2006, 1989m of RC drilling in 27 holes targeting extensions of mineralisation to the

NW and SE of Anwia, and mineralisation hosted by the granite dyke.

• In June and August 2006, 552m of RC drilling in 7 holes drilled to gain samples for

metallurgical test work.

• In June and August 2006, 364m of PQ diamond coring in 5 holes drilled to gain samples for

comminution and metallurgical test work.

• In August to October 2006, 902m (including 156m in 2 RC pre-collars) of HQ diamond coring

in 6 holes for geotechnical logging.

• In October 2006, 270m of RC and open hole drilling in 5 trial dewatering bores and

observation wells.

• In December 2006 to March 2007, 1784m in RC precollars and 2526m of HQ core in 20 holes

designed to infill the main Anwia deposit to approximately 25m x 25m spacing between 120m

and 180m depth below surface.

10.2 Salman Deposits Drilling

BHP (1994-1995) completed 71 RC drill holes for 6,965m and 4 NQ and HQ diamond core holes for

571m on 12 traverses spread over 4 kilometres strike of the Salman Trend. The diamond coring was

conducted by Stanley Drilling Services using a truck mounted LM55 electric hydraulic diamond drill rig.

Ausdrill carried out the RC drilling using UDR650 and Drilteck D25K rigs with 4.5” drill pipe and 5”

RC46 face sampling hammer with 5.25” bit. Most holes are oriented at -45º to 300º magnetic (Salman




47

local grid west). All holes were geologically logged on a one metre basis. Diamond drill core was not

oriented for the collection of structural information.

To the end of November 2005, the data cut-off date for the February 2006 resource estimate, Adamus

had completed 662 drill holes for 49,389m of RC drilling and 2,754m of HQ and NQ diamond coring.

The first 66 RC holes were drilled by Canadian contractor, St Lambert Drilling, using an MPD1500 RCdrill rig with an air capacity of 350psi/900cfm, a face-sampling hammer and button bits of 5.25” or

5.75” diameter. All subsequent Adamus RC drilling programs were undertaken by Minerex using

multipurpose UDR650 and KL900 drill rigs with name plate air capacities of 350psi/700cfm and

500psi/1150cfm respectively. Face-sampling hammers were used with bits of 5.25” or 5.75” diameter.

The RC drill holes were typically collared with PVC pipe down to 6 metres. All Adamus RC holes were

logged on a one metre basis for lithology, weathering and oxidation, and qualitative moisture content

(dry, moist or wet) recorded. RC samples were weighed on a campaign basis to quantitatively check

sample recoveries. Five RC holes at Salman South, Central and North were twinned with HQ3

diamond core holes to confirm the integrity of the RC drilling.

The Adamus diamond core holes were variously drilled from surface or from RC pre-collars up to a

nominal depth of 80 metres. Diamond drill core was typically extracted in 3 metre runs and fresh corewas oriented using a spear or Craelius device at 6 to 12 metre intervals. The core was placed in core

trays by the drilling crew, with annotated core blocks inserted between core runs. The core trays were

then moved to the Adamus core yard at the Salman exploration camp for marking up, geological

logging, sampling and storage. The fresh core was marked with the intervals and a bottom-of-hole

line (based on spear orientation marks) prior to geological logging. Structure orientations were then

measured relative to the bottom-of-hole line. Logged diamond core recoveries within the primary

profile are typically 95-100 per cent, in the oxide zone generally >70 per cent. Minor intervals within

the transition zone, especially within the highly deformed hanging wall phyllite sequence, have proved

less acceptable, with logged recoveries in the 30-50 per cent range.

During 2006, Adamus undertook additional drilling:

• In February and May to August 2006, 15,581m of RC drilling in 241 holes, mainly infilling

previous drill patterns to 25m x 25m spacing.

• In April 2006, 3370m of RC drilling in 50 holes to gain samples for metallurgical test work.

These holes also serve to infill portions of the resource.

• In July and August 2006, 572m of PQ diamond coring in 8 holes drilled to gain samples for

comminution and metallurgical test work.

• In September 2006, 635m of HQ diamond coring in 5 holes for pit-slope geotechnical

investigations.

• In August and September 2006, 800m of RC and open hole drilling in 17 trial dewatering bores

and observation wells.

Almost all of Adamus’ drilling within the Salman Trend has been conducted on UTM Zone 30N

WGS84 east-west lines which are, overall, approximately perpendicular to the strike of the gold

mineralisation. The exception is at Salman North where the Adamus drilling is parallel to that of BHP

(i.e. 120 - 300°). Most holes were drilled at -45° either east or west as appropriate for the dip of the

target lode (both east and west dipping at Salman). A few holes were drilled between -60° and -90°,

and a series of 6 RC holes were drilled approximately parallel to strike (c. north-south) at Salman

Central and Nugget Hill to test mineralisation continuity within lodes.




48

10.3 Satellite Deposits Drilling

All diamond and RC drill holes used in the resource estimates were conducted according to Adamus’

standard RC and diamond drill hole procedures as detailed above. Some historical drilling at Bokrobo

and Aliva which could not be verified was excluded from the resource calculations. At Avrebo, drilling

was undertaken on two different orientations. The original holes (50m spaced sections) were drilledeast-west while the later 20m spaced holes were drilled at 315º.

10.4 Surveying

10.4.1 Anwia Surface Surveying

There is little information about Semafo surveying methods at Anwia. Samax identified potentially

serious collar survey errors in the Semafo database and resurveyed the drill collars in 1999. The

Semafo drill collars were generally clearly labelled and Samax managed to pick-up most of the holesin the resurvey program. The survey was carried out with Sokkia total station equipment using new

local grid control stations which were linked to the Ghana National Grid.

Field checks by Adamus in 2004 indicated significant drill collar coordinate errors in the database

provided to Adamus upon acquisition of the Ebi – Teleku Bokazo concession. Most of the Semafo

and Samax drill collar pipes were still readily visible and clearly labelled, and a second resurveying

program was carried out by using both total station and Differential GPS equipment in UTM Zone 30N

WGS84 coordinates. Accuracy by both methods is nominally centimetre to decimetre in the easting

and northing, sub-decimetre for the RL by total station and sub-metre by Promark 2 DGPS.

As part of the same program, an area approximately 1 km2 centred on the Anwia Deposit was

topographically surveyed by combination of Promark 2 DGPS to establish base stations and Sokkiatotal station equipment along cut lines. Approximately 2000 stations were collected with a final grid

spot height spacing of approximately 25 m by 100 m. Those data have since been added to and

combined with surveyed collar positions to construct a digital terrain model suitable for resource

modelling based on approximately 5,200 spot heights.

10.4.2 Anwia Down-hole Surveys

The Semafo diamond core holes were sporadically surveyed during drilling by acid etch tube and the

results were of doubtful quality. Resource Services Group (RSG) was contracted by Semafo in 1997

to down-hole survey a selection of drill holes. Some 24 diamond core holes and 19 RC holes were

down-hole surveyed by RSG using a Reflex electronic survey instrument in the open hole. Theselected diamond drill holes were generally successfully surveyed, but many of the holes had

collapsed and could not be surveyed much beyond 20-40 metres down hole.

All Adamus diamond core holes and c. 40 per cent of Adamus RC holes at Anwia were down-hole

surveyed by the drilling contractor, Minerex, immediately at the completion of each hole, and/or by

Downhole Surveys using a Flexit Multismart electronic survey tool in the open hole. Minerex collected

both dip and azimuth for diamond core holes by hanging the Eastman single shot camera beyond the

end of the diamond core drill rods at approximately 30 metre intervals down the hole. For RC holes

Minerex surveyed inside the rods, hence only drill hole declination was reported.




49

To address concerns about significant deviation in the historic bore holes Adamus employed Minerex

drilling contractors to re-enter and down-hole survey 22 of the 75 diamond drill holes completed by

Semafo. Both azimuth and declination was collected as for Adamus’ standard diamond drill hole

surveying procedure. The holes selected for resurvey included those with relatively high metal content

and/or important geometrical implications. Digital Surveying was also contracted by Adamus to checksome of the historic drill holes using the Flexit Multismart electronic tool. Some 33 holes were

resurveyed representing all of the previous drilling programs. Results from the open hole resurveying

program were very consistent with the Minerex re-entry program but hampered by a high frequency of

collapsed holes.

10.4.3 Salman Deposi t Surface Surveying

All Adamus drill hole collars have been surveyed in UTM Zone 30N WGS84 coordinates with

centimetre accuracy using Sokkia total station equipment tied to a series of concrete control pillars

established by Promark2 DGPS. Any BHP drill collars which were still visible were also resurveyed

using the same technique.

A strip approximately 8 kilometres long and up to 1 kilometre across covering the Salman Trend has

been topographically surveyed using concrete pillars established by Promark 2 DGPS and Sokkia total

station equipment along cut lines. Over 17,000 spot heights on c. 10m x 50m centres were used to

construct a digital terrain model suitable for resource modelling.

Approximately 35 trenches (c. 2065 m) were at least partially surveyed with Sokkia total station

equipment, 50 trenches (c. 3604 m) were surveyed using a combination of Garmin GPS and chain

and compass then draped on the Adamus digital terrain model, and the remaining 22 trenches (c.

1371 m) were scaled from TEMCO local .grid plans and AGR drill plans (originally located using chain

and compass) then draped over the digital terrain model.

Locations of all pits manually dug in the 2006 campaign were surveyed by total station in the same

manner as drill hole collars.

10.4.4 Salman Deposi t Down-hole Surveys

None of the BHP drill holes were down-hole surveyed. At the time of the feasibility study,

approximately 83 per cent of Adamus diamond core holes and 49 per cent of the RC holes had been

down hole surveyed with a range of equipment including Tropari, single shot down hole cameras and

electronic multishot tools. The breakdown by equipment is: 13 RC holes surveyed to bottom of hole

for both dip and azimuth using Tropari inside stainless steel starter rods; 10 RC holes surveyed to

bottom of hole for both dip and azimuth using an electronic multishot tool inside stainless steel starterrods; 176 RC holes surveyed to bottom of hole for dip only by digital tool inside rods; 199 RC holes

surveyed to bottom of hole for dip only by single shot Eastman camera inside rods; 8 diamond core

holes surveyed to bottom of hole for dip and azimuth by hanging digital survey tool out end of drill

rods; 8 diamond core holes surveyed to bottom of hole for dip and azimuth by hanging single shot

Eastman camera out end of drill rods; 99 RC and diamond core drill holes partially to completely

surveyed with Flexit Multismart tool in the open hole after drilling. The Flexit Multismart tool was

operated by Digital Surveying; all other equipment by the drilling contractor during drilling. Open hole

surveys using the Flexit Multismart tool were hampered by widespread collapse of holes shortly after

drilling. The majority of RC holes at Salman are less than 100 metres long and the average length is




50

about 80 metres. The down hole surveying results show a mean deviation of 3° in dip (n=138) and

less than 2° in azimuth (n=27).

10.4.5 Satellite Deposi t Surveys

All collar and down hole surveys were conducted to Adamus’ standard RC and diamond drill hole

surveying procedures as detailed above.

11.0 SAMPLING METHODS

11.1 RC sampling

All RC samples were collected from the drill rig cyclone on a one metre basis into large plastic bags

then riffle split to collect sub-samples for assay. BHP, Semafo and Samax typically submitted onemetre sub-samples directly for assay, Adamus variably directly submitted one metre sub-samples or

composited the sub-samples for assay according to prospectivity and laboratory assay production

rates.

Semafo RC samples were passed through a 75:25 riffle splitter to produce a 3-4kg sub-sample from

each metre for submission to the assay laboratory. One sample in fifteen had a blind duplicate

prepared and submitted. Splitters were routinely cleaned with high pressure air. For the later part of

the Semafo drilling program a wet splitter was used as required. The wet sample was collected in a

30 litre tub and allowed to settle before the excess water was decanted off. The sample was then

passed through a wet splitter. Semafo had concerns that gold was lost during the decanting process

and a flocculent was added to prior to decanting. Blind field duplicate samples were prepared and

submitted to the laboratory along with the primary assay samples at a frequency of 1:15 samples.

The Samax RC drill samples were riffle split on a one metre basis to provide approximately 4kg from

each metre for assay. It is not known how any wet samples were dealt with. Blanks, standards and

field duplicates were inserted into the sampling sequence at a rates ranging from 1:10 to 1:50

samples.

The BHP RC samples were riffle split to 2kg on a one metre basis. It is not known how any wet

samples were dealt with. BHP used a computer generated random numbering system to label sub-

samples sent to the assay laboratory. Two standards were randomly included for 20 samples. The

standards, Sal1 and Sal2, were prepared by homogenizing crushed rock from a road cutting through

the main mineralised zone at Salman, and assayed values ranged 0.018 to 4.66 ppm and 0.01 to 0.29

ppm Au respectively. A blank was made from beach sand collected near Essiama and assayed less

than 0.02 ppm. Both of these materials are considered unsuitable for assay quality control purposes.




51

All Adamus RC cuttings were collected from the drill rig cyclone at one metre intervals into large

plastic bags before splitting. The sampling cyclone was cleaned at the end of each rod using high-

pressure air from the rig compressor, a scraper, and a rubber mallet to avoid progressive sample

accumulation. The bags were clearly marked with the appropriate hole number and interval using a

permanent marking pen. The plastic bags of RC drill cuttings were weighed on bench scales and laid

out sequentially in rows of 10 or 20 on the drill pad during drilling. Prior to sub-sampling for assay,each single metre field sample was rolled in the bag or tipped into a large clean plastic bucket and

mixed to reduce sample stratification. For the initial Adamus drill programs the one metre samples

were put through a three tier 87:13 riffle splitter and the small sub-samples combined into 4 metre

composite sub-samples weighing 2-3kg for assay. Any significantly mineralised intervals were

identified and single metre sub-samples collected from the original (bulk) samples using a single tier

50:50 riffle splitter then submitted for assay. For later programs the RC drill cuttings were sub-

sampled and submitted for assay as one metre intervals from the outset (i.e. no 4 metre composites).

Wet samples were not immediately split, but instead tube or grab sampled to produce the 4 metre

composite sample then the mineralised one metre bulk samples were air dried before being weighed

and split for one metre assays. In all cases the samples were placed into calico bags labelled with a

unique sample number. Quality control samples, including standards, blanks and field duplicates,

were inserted into the sampling sequence at a rate of 1:20 (5 per cent) in accordance with standardRSG Global quality control protocols (RSG Global, 2002). Laboratory submission sheets were then

completed and samples dispatched by the company or laboratory courtesy vehicle for assay. Upon

receipt of primary assays the pulps of selected mineralised samples were also periodically dispatched

to Genalysis Laboratories, Australia (ISO/IEC 17025: Accreditation No. 3244) for confirmatory

analysis.

11.2 Diamond Drill Core Sampling

The Semafo and BHP diamond drill core was mostly sampled in one metre intervals with the

remaining core (generally ½ core) returned to the trays for storage stored. The three Samax PQ

diamond drill core holes were sampled and assayed on a lithological basis.

Most of the Adamus oxide and transition zone core was wrapped with plastic film immediately after

drilling for bulk density determinations which were performed on site (see below) prior to sampling. All

Adamus diamond core sampling occurred at the Salman exploration camp. The drill core was

sampled in 1 or 2 metre intervals according to prospectivity: Single metre samples of ½ core were

collected through zones considered prospective by the supervising geologist, owing to them

containing visible gold or any mineral assemblage known from previous work to have potential to host

gold, quartz, quartz veining, fresh or oxidised sulphides, alteration and/or being within structurally

complex zones or packages. Two metre ¼ core samples were collected through zones considered to

be waste owing to the absence of any of the above markers of prospective mineralisation. Competent

(fresh and some transition) core was cut using core saw, soft oxide and transition material sampled

with a knife or spatula. The samples were placed into calico or plastic bags labelled with a unique

sampling number. The remaining core was returned to the trays for storage. Quality control reference

standards and blanks were inserted into the sampling sequence at a rate of 1:20 (5 per cent) in

accordance with RSG Global quality control protocols (RSG Global, 2002). Laboratory submission

sheets were then completed and samples dispatched by company or laboratory courtesy vehicle for

assay. Upon receipt of primary assays the pulps of selected mineralised samples were also

periodically dispatched to SGS Analabs (ISO/IEC 17025: Accreditation No. 1936) and Genalysis

Laboratories, Perth, Australia for confirmatory analysis.




52

11.3 Trench and Pit Sampling

All trenches (TEMCO, AGR and Adamus) were geologically logged and sampling intervals of 1, 2 or 3

metres as appropriate marked with coloured flagging tape and/or aluminium permatags along the wallusing tape. Samples were collected as a continuous channel running along the base of one wall or

floor of the trench or bulldozer line as appropriate, placed in calico bags, given a unique sample

number and submitted for assay. Quality control samples, including standards, blanks and field

duplicates, were inserted into the Adamus sampling sequence at a rate of 1:20.

Manually dug pits dug at Salman during 2006 were sampled in vertical channels with individual

sample intervals ranging from 0.5 metres to 1.5 metres length.

12.0 SAMPLE PREPARATION, ANALYSES AND SECURITY

12.1 Sample Preparation

Sample preparation is described in Section 11 of this Report. Beyond this point, no employee, officer,

director or associate of Adamus was involved in any aspect of the preparation or analysis of samples

from the Project.

12.2 Analyses

12.2.1 TEMCO

The TEMCO trench samples were sent to SGS Laboratory, Accra where they were pulverized and

assayed for Au by aqua regia digest presumably with AAS finish. Most samples were also analysed

for As, Sb, Cu and/or Mn. Assay certificates are not available but TEMCO exploration reports with a

complete printout of trench sample descriptions with assays are held by Adamus (Tropical Exploration

and Mining Co Ltd, 1992, Tropical Exploration and Mining Co Ltd, 1993).

12.2.2 Semafo

Semafo diamond core and RC samples were submitted to SGS Laboratories, Tarkwa for assay of gold

by 50g charge fire assay. Assay certificates are not available to Adamus.

12.2.3 Samax

Samax RC and diamond core drilling samples of approximately 4kg each were submitted to SGS

Laboratories, Tarkwa, pulped and assayed for gold by 2kg agitated cyanide leach (bottled roll) with

AAS finish. Assay certificates are held by Adamus.




53

12.2.4 BHP

BHP diamond core and RC samples were sent to SGS Laboratories in Accra where they were

assayed for gold by 50g charge fire assay with AAS finish. Lower limit of detection was 0.01 ppm. A

second 50g pulp sample was taken from each pulverized diamond core bulk residue at the laboratory

for aqua regia digest followed by assay of As, Cu and Sb by AAS. A batch of 149 duplicate samples

from RC holes with significant mineralisation was analyzed by CHEMEX laboratories in Canada; the

CHEMEX and SGS assays compared well (e.g. Bolton & Amegashi, 1996). A selection of assay

certificates and complete printout of digital BHP drill hole logs with assays is held by Adamus.

12.2.5 ARL

Adamus RC and diamond core samples were submitted to Transworld Laboratories and/or SGS

Laboratories both in Tarkwa, for industry standard 50g charge fire assay for gold with AAS finish and

0.01 ppm lower limit of detection. Approximately 1800 samples were pulped at SGS Tarkwa and 120

samples at Transworld then air-freighted to Genalysis, Perth for 50g charge fire assay for gold with

AAS finish (0.01 ppm lower limit of detection). Approximately 44,000 gold assays were successfully

conducted by the three laboratories, approx. 90 per cent by Transworld, 7 per cent by SGS Tarkwa,

and 3 per cent pulped by SGS Tarkwa and Transworld and assayed by Genalysis. During 2005, field

duplicates and pulps of selected mineralised samples from Transworld and SGS Tarkwa were

periodically air-freighted to SGS Analabs and Genalysis Laboratories, Perth for check assaying.

Approximately 5900 samples were also assayed for gold by 1kg agitated cyanide leach (bottle roll)

with Leachwell and tail fire assay at Transworld and SGS in Tarkwa. Approximately 200 samples

were assayed by 200 or 400g cyanide bottle roll with Leachwell and tail f ire assay at Genalysis, Perth.

Standard sample preparation and fire assay procedure at all laboratories involved oven drying ofsamples upon arrival, followed by jaw crushing to -2 mm, followed by complete pulverization to P90% -

75µm in LM2 or LM5 disk mills, followed by homogenisation and sub-sampling to obtain 150-200g

pulp, with 50g sub-sampled from the pulp for lead collection fire assay and AAS finish for Au.

Remaining pulps were returned to Adamus for storage or re-assay as appropriate. All assays were

supplied to Adamus in electronic form and as hardcopy assay certificates. From mid-2004 to mid-

2005, quality control data were periodically analysed by RSG Global and appropriate

recommendations made. Subsequent QCQA monitoring has been by Adamus.

12.3 Analyses - Bulk Densities

12.3.1 Anwia

About 1400 bulk density determinations have been undertaken by Adamus using air-dried HQ core

and the weight-in-air, weight-in-water method. Figure 12-1 to Figure 12-4 show summary statistics for

bulk densities in each of the weathering subdomains. In addition, Samax measured bulk densities on

27 samples taken from hand-dug pits at depths up to four metres below surface. Densities of those

samples ranged from 1.512 to 2.22 g/cc. Densities applied to the resource model are listed in Table

12-1.




54

Table 12-1 Bulk densities applied to the Anwia resource model

Weathering subdomain Bulk density g/cc

Very weathered, 1 1.8

Moderately weathered, 2 2.1

Weakly weathered, 3 2.5

Fresh rock, 4 2.8

Figure 12-1 Bulk densit ies of Anwia very weathered drill core

Univariate Statistics

variable:

weighted by:

mean:

varnc:

coefvrn:

min:

q1:

median:

q3:

max:

iqr:

no. of data:

density

--

1.878

0.030

0.092

1.590

1.750

1.850

1.950

2.540

0.200

89 / 1393(data is sub -setted)

1.0 1.5 2.0 2.5 3.00

0.05

0.10

0.15

0.20

0.25

grade class - density

P r o p o r t i o n o f s a m p l e s

Histogram of density

Figure 12-2 Bulk densi ties of Anwia moderately weathered drill core


variable:

weighted by:

mean:

varnc:

coefvrn:

min:

q1:

median:

q3:

max:

iqr:

no. of data:

density

--

2.155

0.036

0.088

1.710

2.000

2.140

2.280

2.850

0.280


1.0 1.5 2.0 2.5 3.00

0.05

0.10

0.15

0.20


P r o p o r t i o

n o f s a m p l e s





55

.

Figure 12-3 Bulk densit ies of Anwia weakly weathered drill core


variable:

weighted by:

mean:

varnc:

coefvrn:

min:

q1:

median:

q3:

max:

iqr:

no. of data:

density

--

2.446

0.028

0.068

1.990

2.300

2.430

2.600

2.770

0.300


1.0 1.5 2.0 2.5 3.00

0.05

0.10

0.15

0.20

0.25




Figure 12-4 Bulk densities of Anwia fresh rock drill core


variable:

weighted by:

mean:

varnc:

coefvrn:

min:

q1:

median:

q3:

max:

iqr:

no. of data:

density

--

2.774

0.016

0.046

1.500

2.750

2.780

2.810

4.750

0.060


1.5 2.0 2.5 3.0 3.50

0.1

0.2

0.3

0.4

0.5

0.6







56

12.3.2 Salman

Adamus has measured about 1480 bulk densities on HQ drill core from Salman. Measurements used

the weight-in-air, weight-in-water method on air-dried HQ core. Figure 12-5 to Figure 12-8 show

summary statistics for bulk densities in each of the weathering subdomains and densities applied to

the resource model are listed in Table 12-2

Table 12-2 Bulk densi ties applied to the Salman resource model





Fresh rock, 4 2.8

Figure 12-5 Bulk densit ies of Salman very weathered dril l core


variable:

weighted by:

mean:

varnc:

coefvrn:

min:

q1:

median:

q3:

max:

iqr:

no. of data:

density

--

1.815

0.046

0.118

1.220

1.720

1.810

1.900

2.670

0.180


1.0 1.5 2.0 2.5 3.00

0.05

0.10

0.15


P r o

p o r t i o n o f s a m p l e s





57

Figure 12-6 Bulk densi ties of Salman moderately weathered drill core


variable:

weighted by:

mean:

varnc:

coefvrn:

min:

q1:

median:

q3:

max:

iqr:

no. of data:

density

--

2.173

0.041

0.093

1.550

2.060

2.150

2.270

2.680

0.210


1.0 1.5 2.0 2.5 3.00

0.02

0.04

0.06

0.08

0.10

0.12

0.14




Figure 12-7 Bulk densi ties of Salman weakly weathered drill co re


variable:

weighted by:

mean:

varnc:

coefvrn:

min:

q1:

median:

q3:

max:

iqr:

no. of data:

density

--

2.560

0.010

0.039

2.200

2.490

2.580

2.630

2.760

0.140


1.0 1.5 2.0 2.5 3.00

0.05

0.10

0.15

0.20


P r o p o r t i o n o f s a m p l e

s





58

Figure 12-8 Bulk densit ies of Salman fresh rock drill core


variable:weighted by:

mean:

varnc:

coefvrn:

min:

q1:

median:

q3:

max:

iqr:

no. of data:

density --

2.763

0.013

0.041

2.140

2.740

2.760

2.790

5.740

0.050


1.5 2.0 2.5 3.0 3.50

0.1

0.2

0.3

0.4




12.3.3 Satelli te Deposi ts

No bulk density data were available for the Satellite Deposits. The bulk densities assigned to the four

weathering zones modelled in this study are shown in Table 12-3. They are the same as the densities

calculated at Aniwa and Salman.

Table 12-3: Bulk densi ties applied to the Satellite Deposits resource models





Fresh rock, 4 2.8

12.4 Sample Storage and Security

RC samples are prepared and collected from the drill rig on a daily basis and periodically delivered to

Transworld and SGS Laboratories in Tarkwa for analysis by company vehicle or laboratory courtesy

vehicle. Similarly, the diamond core is transferred to the core yard at the Salman exploration camp on

a daily basis for geological logging and sampling prior to submission. All Adamus drill core, BHP drill

core, BHP and Adamus assay pulps, BHP RC chip-boards, Adamus RC chip trays and significantly

mineralised RC bulk residues from Adamus programs are stored at the Salman exploration camp.

Semafo drill core from Anwia is stored under cover at the old Semafo core yard in Nkroful. Semafo

and Samax RC bulk residues and drill core from the three Samax diamond holes are no longer

available.




59

From mid-2004 to mid-2005, sample handling was monitored by RSG Global personnel to ensure

adequate sample quality and security. After establishment of standard operating procedures, all

subsequent sample handling has been by Adamus personnel.

12.5 Representivity of Samples

A discussion of the representivity of samples is set out in Section 13 of this Report.

12.6 Adequacy of Sample Preparation, Security and Analytical

Procedures

In the author’s opinion, the methods of sample preparation and analysis used conform to those

described elsewhere as “industry standards”.

The sample security procedures conform to those instituted and revised by RSG Global in 2004-5. All

QCQA monitoring since has been undertaken by Adamus. The results of subsequent QCQA checks

are such that, in the author’s opinion, sample security procedures have been adequate since 2004.




60

13.0 DATA VERIFICATION

13.1 Anwia Assay Accuracy

Adamus has no record of reference standards submitted with drilling at Anwia by Semafo or

Samax/AGC but records do exist for blank samples submitted by Samax/AGC (Figure 13-1). Samples

from core and RC drilling by Adamus in 2004 were submitted with 1:20 sample blanks and reference

standards. Figure 13-2 shows assays of blanks and Figure 13-3 to Figure 13-9 show run charts for

assays of reference standards. Blank samples indicate that with few exceptions, that may have

resulted from sample switches, there is no evidence of significant sample-to-sample contamination.

Reference standards do not indicate any consistent bias in assays by SGS or Transworld laboratories

although Transworld tend to report low.

Figure 13-1 Assays of Samax/AGC blanks submitted with Anwia drill samples

Sample Blank Contr ol Chart

-1

-0.5

0

0.5

1

0 20 40 60 80 100 120 140 160 180 200 220

Date sequ ence

g / t A u

SGS FA

Figure 13-2 Assays of Adamus blanks submitted with Anwia drill samples

Sample Blank Contr ol Chart

-1.00

-0.50

0.00

0.50

1.00

0 20 40 60 80 100 120 140 160 180 200 220 240 260

Date seq uence

g

/ t A u

SGS FA TWL FA




61

Figure 13-3 Assays of STD4B submitted with Adamus Anwia samples

Standard Control Chart

1

1.25

1.5

1.75

2

0 2 4 6 8 10 12 14 16 18 20 22 24 26 28 30 32

Date seq uence

g / t A u

STD 4B 1.48g/t +/- 10% SGS FA TWL FA



0.2

0.3

0.4

0.5

0.6

0.7

0.8

0 2 4 6 8 10 12 14 16 18 20 22 24 26 28 30 32

Date se quence

g / t A u




7

8

9

10

11

12

0 2 4 6 8 10 12 14 16

Date se quence

g / t A u





62


Standard Contr ol Chart

1.5

1.75

2

2.25

2.5

0 5 10 15 20 25 30 35 40 45 50 55

Date seq uence

g / t A u




2

2.25

2.5

2.75

3

0 2 4 6 8 10 12 14 16 18 20 22

Date se quence

g / t A u




0.75

1

1.25

1.5

1.75

2

0 2 4 6 8 10 12 14 16 18 20 22 24 26 28 30 32

Date s equence

g / t A u





63



10

12

14

16

18

0 2 4 6 8 10 12 14 16 18 20

Date seq uence

g / t A u


13.2 Comparison of Adamus to Semafo and Samax/AGC Sampling

Estimation of gold resources at Anwia relies significantly on assay data deriving from previous drilling

by Semafo and Samax/AGC. The only way to check how reliably these samples represent

mineralisation was to compare them to subsequent Adamus drilling. Using one-metre sample

composites, a search was undertaken to find pairs of samples from holes drilled by Semafo or Samax

and samples drilled by Adamus lying within 10mE x 10mN x 0.5mRL of each other. The resulting

pairs were then filtered to remove duplicate pairings, i.e., to retain the unique earlier sample lying

nearest to each Adamus drill sample within the search radii. The search located 1405 unique nearest

neighbour pairs with a mean separation distance of 7.1 metres. It was reasonable to assume thatthese pairs represent two independent samplings of the same region of mineralisation. Figure 13-10

shows a scatter plot of assays. Correlation between gold grades in individual pairs of samples is poor,

as is to be expected in a deposit such as Anwia. Figure 13-11 shows a quantile-quantile plot

comparing the marginal histograms of the two sample populations. There is no obvious bias to higher

gold grades in either sample population and the summary statistics of the two sample populations are

very similar. It may be concluded that the two samplings are equally representative of Anwia

mineralisation.




64

Figure 13-10 Scatter plot of nearest neighbour sample pairs

Data Statistics

variable:

weight by:

mean:

varnc:

coefvrn:

min:

q1:

median:

q3:

max:

iqr:

ADU_Au

--

0.527

2.730

3.134

0.005

0.030

0.070

0.320

17.200

0.290

AGC_Au

--

0.549

2.452

2.852

0.001

0.020

0.080

0.350

17.700

0.330

covarnc:

Pearson:

Spearman:

no. data:

0.423

0.163

0.454

1389 / 1405

(data is sub-setted)

0 5 10 15 200

5

10

15

20

ADU_Au

A G C

_ A u

Scatter pl ot

Figure 13-11 Q-Q plot of nearest neighbour sample pairs

Data Statistics

variable:

weight by:

mean:

varnc:

coefvrn:

min:

q1:

median:

q3:

max:

iqr:

ADU_Au

--

0.527

2.730

3.134

0.005

0.030

0.070

0.320

17.200

0.290

AGC_Au

--

0.549

2.452

2.852

0.001

0.020

0.080

0.350

17.700

0.330

covarnc:

Pearson:

Spearman:

no. data:

0.423

0.163

0.454

1389 / 1405


0 5 10 15 200

5

10

15

20

ADU_Au

A G C_

A u

Q-Q Plo t

13.3 Anwia Sampling and Assaying Precision

Field re-splits of RC drill samples measure the accumulated errors of the entire sampling and assaying

procedure. Available data include re-splits by Samax/AGC during their 1998-2000 drilling campaigns,

and 1:20 field re-splits of RC samples submitted by Adamus during 2004 drilling.




65

Figure 13-12 and Figure 13-13 show scatter and precision plots and summary statistics for field re-split

samples submitted by Samax/AGC. The correlation between first and second split samples is very

high because of good repeatability of high-grade samples. Those samples also influence the

precision statistics which, at 12 per cent, is high for RC drill samples from a gold deposit such as

Anwia. There is no obvious bias to higher grades in either sample split.

Figure 13-14 and Figure 13-15 show the same statistics for field re-split samples submitted by

Adamus. The correlation and precision are not as good as for the Samax/AGC pairs but there are far

fewer data to compare. At 34 per cent the precision is more typical of pairs of RC samples from lode

gold deposits.

Figure 13-12 Scatter plot: Samax/AGC Anwia field re-split s

Data Statistics

variable:

weight by:

mean:

varnc:

coefvrn:

min:

q1:

median:

q3:

max:

iqr:

Au_FDPP

--

0.700

5.253

3.272

0.005

0.020

0.090

0.430

25.600

0.410

AU_FDP1

--

0.712

5.579

3.316

0.001

0.020

0.090

0.400

28.800

0.380

covarnc:

Pearson:

Spearman:

no. data:

5.328

0.984

0.944

844 / 850

(data is sub -setted)

0 5 10 15 20 25 300

10

20

30

Au_FDPP

A U_

F D P 1

Scatter plot

Figure 13-13 Precision plo t: Samax/AGC field re-splits




66

Data Statistics

variable:

weight by:

mean:

varnc:

coefvrn:

min:

q1:

median:

q3:

max:

iqr:

Au_FDPP

--

0.700

5.253

3.272

0.005

0.020

0.090

0.430

25.600

0.410

AU_FDP1

--

0.712

5.579

3.316

0.001

0.020

0.090

0.400

28.800

0.380

covarnc:

Pearson:

Spearman:

precision:

no. data:

5.328

0.984

0.944

+/-12% @ 56%CI

844 / 850


0 10 20 300

1

2

3

4

5

6

Pair Average

P a i r A b s o l u t e D i f f e r e n c e

Paired Data Precision plot




67

Figure 13-14 Scatter plot: Adamus Anwia field re-splits

Data Statistics

variable:

weight by:

mean:

varnc:

coefvrn:

min:

q1:

median:

q3:

max:

iqr:

Au_FDPP

--

0.182

0.129

1.965

0.005

0.020

0.050

0.130

2.180

0.110

Au_FDP1

--

0.189

0.147

2.037

0.005

0.020

0.060

0.160

1.950

0.140

covarnc:

Pearson:

Spearman:

no. data:

0.120

0.872

0.897

95 / 98


0 0.5 1.0 1.5 2.0 2.50

0.5

1.0

1.5

2.0

2.5

Au_FDPP

A u_

F D P 1

Scatter plot

Figure 13-15 Precision plot : Adamus Anwia field re-splits

Data Statistics

variable:

weight by:

mean:

varnc:

coefvrn:

min:

q1:

median:

q3:

max:

iqr:

Au_FDPP

--

0.182

0.129

1.965

0.005

0.020

0.050

0.130

2.180

0.110

Au_FDP1

--

0.189

0.147

2.037

0.005

0.020

0.060

0.160

1.950

0.140

covarnc:

Pearson:

Spearman:

precision:

no. data:

0.120

0.872

0.897

+/-34% @ 61%CI

95 / 98


0 0.5 1.0 1.5 2.0 2.50

0.2

0.4

0.6

0.8

1.0

1.2

Pair Average






68

13.4 Salman RC Sample Recovery

During 2002-2003 RC drilling at Salman and Akanko RC samples were weighed on a campaign basis

to quantitatively check sample recoveries. A total of 5,144 samples in 75 holes were weighed as theywere recovered from the drill cyclone. Figure 13-16 to Figure 13-19 show histograms and summary

statistics for per cent sample recoveries, grouped by degree of weathering (and thus also

approximately by down-hole depth). Based on experience, sample recoveries from RC drilling

commonly range from about 55 to 70 per cent, with industry best practice achieving about 85 per cent.

Available data indicate that sample recoveries in strongly and moderately weathered materials in early

Salman drilling were quite poor, averaging about 40 per cent. Recoveries from shallow depths are of

course adversely affected by loss of sample in the first five or six metres of drilling when confining

pressure around the bit face is low. Sample recoveries in weakly weathered material and fresh rock

are more acceptable, averaging 50 and 60 per cent respectively.

A large number of RC samples from resource definition drilling in 2006 were weighed. Data areavailable for 13,016 samples from 212 holes. Figure 13-20 to Figure 13-23 show histograms and

summary statistics for sample recoveries, again grouped by degree of weathering. Sample recoveries

range from 57 per cent in near-surface, very weathered material though 69 per cent for moderately

weathered, 83 per cent for weakly weathered and 85 per cent for fresh rock. These are considered

better than normal industry practice.

Figure 13-16 Sample recoveries in 2002-2003 RC drilling, very weathered material


variable:

weighted by:

mean:

varnc:

coefvrn:

min:

q1:

median:

q3:

max:

iqr:

no. of data:

%recov

--

40.982

738.762

0.663

1.000

15.000

35.000

64.000

100.000

49.000

223 / 5144(data is sub-setted)

0 10 20 30 40 50 60 70 80 90 1000

0.02

0.04

0.06

0.08

0.10

0.12

0.14

grade class - %recov

P r o p o r t i o n o f s a m

p l e s

Histogram of %recov




69

Figure 13-17 Sample recoveries in 2002-2003 RC drilling , moderately weathered material


variable:

weighted by:

mean:

varnc:

coefvrn:

min:

q1:

median:

q3:

max:

iqr:

no. of data:

%recov

--

39.542

316.300

0.450

3.000

28.000

40.000

50.000

87.000

22.000


0 10 20 30 40 50 60 70 80 90 1000

0.02

0.04

0.06

0.08

0.10

0.12


P r o p o r t i o n

o f s a m

p l e s

Histogram of %recov

Figure 13-18 Sample recoveries in 2002-2003 RC drilling, weakly weathered material


variable:

weighted by:

mean:

varnc:

coefvrn:

min:

q1:

median:

q3:

max:

iqr:

no. of data:

%recov

--

51.505

278.469

0.324

3.000

40.000

52.000

63.000

100.000

23.000


0 10 20 30 40 50 60 70 80 90 1000

0.02

0.04

0.06

0.08

0.10

0.12


P r o p o r t i o n

o f s a m

p l e s

Histogram of %recov




70

Figure 13-19 Sample recoveries in 2002-2003 RC drilli ng, fresh rock


variable:

weighted by:

mean:

varnc:

coefvrn:

min:

q1:

median:

q3:

max:

iqr:

no. of data:

%recov

--

59.089

411.963

0.343

4.000

44.000

61.000

74.000

100.000

30.000


0 10 20 30 40 50 60 70 80 90 1000

0.02

0.04

0.06

0.08

0.10


P r o p o r t i o n

o f s a m

p l e s

Histogram of %recov

Figure 13-20 Sample recoveries in 2006 RC drilling , very weathered material


variable:

weighted by:

mean:

varnc:

coefvrn:

min:

q1:

median:

q3:

max:

iqr:

no. of data:

%recov

--

57.296

655.806

0.447

8.000

38.000

57.000

77.000

100.000

39.000


0 10 20 30 40 50 60 70 80 90 1000

0.02

0.04

0.06

0.08

0.10


P r o p o r t i o n o f s a m

p l e s

Histogram of %recov




71

Figure 13-21 Sample recoveries in 2006 RC drilling , moderately weathered material


variable:

weighted by:

mean:

varnc:

coefvrn:

min:

q1:

median:

q3:

max:

iqr:

no. of data:

%recov

--

68.999

471.846

0.315

6.000

54.000

70.000

86.000

100.000

32.000


0 10 20 30 40 50 60 70 80 90 1000

0.05

0.10

0.15


P r o p o r t i o n

o f s a m

p l e s

Histogram of %recov

Figure 13-22 Sample recoveries in 2006 RC dril ling, weakly weathered material


variable:

weighted by:

mean:

varnc:

coefvrn:

min:

q1:

median:

q3:

max:

iqr:

no. of data:

%recov

--

83.273

294.965

0.206

10.000

74.000

88.000

99.000

100.000

25.000

3273 / 13016


0 10 20 30 40 50 60 70 80 90 1000

0.1

0.2

0.3

0.4


P r o p o r t i o n

o f s a m

p l e s

Histogram of %recov




72

Figure 13-23 Sample recoveries in 2006 RC dril ling, fresh rock


variable:

weighted by:

mean:

varnc:

coefvrn:

min:

q1:

median:

q3:

max:

iqr:

no. of data:

%recov

--

84.676

212.565

0.172

7.000

79.000

89.000

96.000

100.000

17.000


0 10 20 30 40 50 60 70 80 90 1000

0.05

0.10

0.15

0.20

0.25


P r o p o r t i o n

o f s a m

p l e s

Histogram of %recov

13.5 Salman Assay Accuracy

There are no records available for any reference standards or other check samples that may have

been submitted with drill samples by BHP. However, BHP drill samples now form a very small

proportion of the data informing resource estimates. From the commencement of drilling at Salman by

Adamus, blanks, reference standards and field-re-splits of RC samples have been interleaved with

samples prior to submission for assay.

Figure 13-24 to Figure 13-32 show run charts for blank samples submitted to SGS, Transworld and

Genalysis laboratories. Blank samples have, at various times, comprised crushed fire assay pots,

rejects from previously assayed RC samples that returned very low gold grades and portions of

Voltaian sandstone and grit. Many of the blanks that stand out on the charts as having returned

elevated gold grades derive from RC sample rejects. Otherwise there is no marked evidence of cross-

sample contamination in assay laboratories.




73

Figure 13-24 Adamus blanks submitted for SGS fire assay with Salman drill samples

Sample Blank Control Chart

-1

-0.5

0

0.5

1

0 25 50 75 100 125 150 175 200 225

Date sequ ence

g / t A u

SGS FA

Figure 13-25 Adamus blanks submitted for Transworld CN leach assay with Salman drill

samples


-1

-0.5

0

0.5

1

0 10 20 30 40 50 60

Date sequ ence

g / t A u

TWL CN Leach

Figure 13-26 Adamus blanks submi tted for Transworld fire assay with 2002 dril l samples


-1

-0.5

0

0.5

1

0 10 20 30 40 50 60

Date sequ ence

g / t A u

TWL FA 2002




74

Figure 13-27 Adamus blanks submi tted for Transworld fi re assay with 2003 dril l samples


-1

-0.5

0

0.5

1

0 50 100 150 200 250 300 350

Date sequ ence

g / t A u

TWL FA 2003



-1

-0.5

0

0.5

1

0 25 50 75 100 125 150 175

Date sequ ence

g / t A u

TWL FA 2004

Figure 13-29 Adamus blanks submi tted for Transworld fi re assay with 2005 dril l samples


-1

-0.5

0

0.5

1

0 10 20 30 40 50 60 70 80 90

Date sequ ence

g / t A u

TWL FA 2005




75



-1

-0.5

0

0.5

1

0 50 100 150 200 250 300 350 400 450

Date sequ ence

g / t A u

TWL FA 2006



-1

-0.5

0

0.5

1

0 2 4 6 8 10 12 14 16

Date sequ ence

g / t A u

Genalysis FA

Figure 13-32: Adamus blanks submitted for Genalysis fire assay with Salman drill samples


-1

-0.5

0

0.5

1

0 2 4 6 8 10 12 14 16

Date sequ ence

g / t A u

Genalysis FA




76

Figure 13-33 through to Figure 13-42 show run charts for reference standards submitted with drill

samples fire assayed by SGS Tarkwa. Standards are sourced from Rocklabs New Zealand and from

Gannet Industries, Perth. The erratic results of some standards possibly relates to the reference

material itself. In the author’s experience, some reference materials produced by Gannet return

inconsistent gold grades. Bearing in mind that most sample batches submitted for analysis contain atleast four reference standards, there is no evidence for significant bias in the SGS assays.

Figure 13-43 to Figure 13-57 show run charts for standards submitted to Transworld Laboratories,

Tarkwa. Assays are mainly within 10 per cent error bounds and there is no consistent bias evident.

The erratic grades returned from high-grade standard 6B almost certainly reflect inhomogeneity in the

material.

Figure 13-33: Assays of STD1 submit ted to SGS with Salman samples


0.5

0.75

1

0 2 4 6 8 10 12 14 16 18 20

Date seq uence

g / t A u

STD1 0.802g/t +/- 10% SGS FA

Figure 13-34: Assays of STD4B submitted to SGS with Salman samples


1

1.25

1.5

1.75

2

0 2 4 6 8 10 12 14 16 18 20 22 24 26 28 30

Date sequ ence

g / t A u

STD4B 1.48g/t +/- 10% SGS FA




77



0.25

0.5

0.75

0 2 4 6 8 10 12 14 16 18 20 22 24 26 28 30

Date sequ ence

g / t A u

STD5B 0.50g/t +/- 10% SGS FA



8

9

10

11

12

0 5 10 15 20 25 30 35 40 45 50

Date seq uence

g / t A u

STD6B 9.70g/t +/- 10% SGS FA



1.25

1.75

2.25

2.75

0 1 2 3 4 5 6 7 8 9 10 11 12

Date sequ ence

g / t

A u

STD7B 2.06g/t +/- 10% SGS FA




78



1.75

2

2.25

2.5

2.75

0 10 20 30 40 50 60

Date sequ ence

g / t A u

STD8B 2.36g/t +/- 10% SGS FA



1

1.25

1.5

1.75

0 5 10 15 20 25 30 35 40 45 50

Date sequ ence

g / t A u

STD9B 1.33g/t +/- 10% SGS FA

Figure 13-40: Assays of STD10B submi tted to SGS with Salman samples


10

12

14

16

18

0 5 10 15 20 25 30 35 40 45

Date seq uence

g / t

A u

STD10B 13.9g/t +/- 10% SGS FA




79

Figure 13-41: Assays of STD11 submitted to SGS with Salman samples


1.25

1.5

1.75

2

2.25

2.5

0 2 4 6 8 10 12 14 16 18 20 22 24 26 28 30 32 34

Date seq uence

g / t A u

STD11 1.805g/t +/- 10% SGS FA

Figure 13-42: Assays of STD12 submitted to SGS with Salman samples


6

6.5

7

7.5

8

8.5

9

0 2 4 6 8 10 12 14 16 18 20

Date seq uence

g / t A u

STD12 7.615g/t +/- 10% SGS FA

Figure 13-43: Assays of STD1 submit ted to Transworld wi th Salman samples


0.5

0.75

1

0 10 20 30 40 50 60 70 80 90 100

Date s equence

g / t A u

STD1 0.802g/t +/- 10% TWL FA




80

Figure 13-44: Assays of STD02 submitted to Transwor ld with Salman samples


0.75

1

1.25

0 2 4 6 8 10 12 14 16 18 20 22 24 26 28

Date s equence

g / t A u

STD2 0.939g/t +/- 10% TWL FA



2

2.25

2.5

2.75

3

0 10 20 30 40 50 60 70 80 90 100 110 120

Date s equence

g / t A u

STD3 2.427g/t +/- 10% TWL FA



2

2.25

2.5

2.75

3

0 2 4 6 8 10 12 14 16 18 20 22 24 26

Date s equence

g / t A u

STD4 2.58g/t +/- 10% TWL FA




81

Figure 13-47: Assays of STD4B submi tted to Transworld w ith Salman samples


1

1.25

1.5

1.75

2

0 40 80 120 160 200 240 280 320

Date s equence

g / t A u

STD4B 1.48g/t +/- 10% TWL FA



9

10

11

12

0 2 4 6 8 10 12 14 16 18 20 22 24

Date sequence

g / t A u

STD5 10.47g/t +/- 10% TWL FA



0.35

0.45

0.55

0.65

0 20 40 60 80 100 120 140 160 180 200

Date sequence

g / t A u

STD5B 0.50g/t +/- 10% TWL FA




82



12

14

16

18

0 2 4 6 8 10 12 14 16 18 20 22 24

Date sequence

g / t A u

STD6 15.15g/t +/- 10% TWL FA



8

9

10

11

12

0 20 40 60 80 100 120 140 160 180 200

Date s equence

g / t A u

STD6B 9.70g/t +/- 10% TWL FA



1.25

1.75

2.25

2.75

0 20 40 60 80 100 120 140 160 180 200 220

Date sequence

g / t A u

STD7B 2.06g/t +/- 10% TWL FA




83



1.75

2

2.25

2.5

2.75

0 25 50 75 100 125 150 175 200 225 250 275 300 325 350 375

Date sequence

g / t A u

STD8B 2.36g/t +/- 10% TWL FA



1

1.25

1.5

1.75

0 15 30 45 60 75 90 105 120 135 150

Date sequence

g / t A u

STD9B 1.33g/t +/- 10% TWL FA

Figure 13-55: Assays of STD10B submit ted to Transwor ld with Salman samples


10

12

14

16

18

0 5 10 15 20 25 30 35 40 45 50 55 60 65 70 75 80

Date sequence

g / t A u

STD10B 13.9g/t +/- 10% TWL FA




84

Figure 13-56: Assays of STD11 submitted to Transwor ld with Salman samples


1.25

1.5

1.75

2

2.25

2.5

0 2 4 6 8 10 12 14 16 18

Date sequence

g / t A u

STD11 1.805g/t +/- 10% TWL FA

Figure 13-57: Assays of STD13B submit ted to Transwor ld with Salman samples


10

12

14

16

18

0 20 40 60 80 100 120 140 160 180 200 220 240Date sequence

g / t A u

STD13B 13.9g/t +/- 10% TWL FA

13.6 Salman Sampling and Assaying Precision

Almost from inception, Adamus has submitted field re-splits of Salman RC drill samples at a ratio of

1:20. Figure 13-58 and Figure 13-59 show scatter and precision plots and summary statistics for field

re-split samples submitted to SGS for fire assay and Figure 13-60 to Figure 13-63 show the same data

for samples fire assayed by Transworld and assayed by cyanide bottle roll by Transworld.Correlations between first and second spit samples are satisfactory and precisions, at 25-35 per cent,

are typical of sampling in lode gold deposits. Some outliers probably represent mismatches of pairs.

In each of the data sets there is no obvious bias to higher grades in either sample split.




85

Figure 13-58: Scatter p lot: SGS fire assays of Salman field re-splits

Data Statistics

variable:

weight by:

mean:

varnc:

coefvrn:

min:

q1:

median:

q3:

max:

iqr:

Au_FDPP

--

0.913

2.491

1.728

0.010

0.090

0.270

0.870

12.300

0.780

Au_FDP1

--

0.879

1.908

1.571

0.005

0.080

0.310

0.990

10.000

0.910

covarnc:

Pearson:

Spearman:

no. data:

1.856

0.852

0.906

369 / 1512


0 5 10 150

5

10

15

Au_FDPP

A u_

F D P 1

Scatter pl ot

Figure 13-59: Precision plot : SGS fire assays of Salman field re-spli ts

Data Statistics

variable:

weight by:

mean:

varnc:

coefvrn:

min:

q1:

median:

q3:

max:

iqr:

Au_FDPP

--

0.913

2.491

1.728

0.010

0.090

0.270

0.870

12.300

0.780

Au_FDP1

--

0.879

1.908

1.571

0.005

0.080

0.310

0.990

10.000

0.910

covarnc:

Pearson:

Spearman:

precision:

no. data:

1.856

0.852

0.906

+/-36% @ 63%CI

369 / 1512


0 2 4 6 8 100

1

2

3

4

5

6

7

8

9

Pair Average






86

Figure 13-60: Scatter plot: Transworld fire assays of Salman field re-splits

Data Statistics

variable:

weight by:

mean:

varnc:

coefvrn:

min:

q1:

median:

q3:

max:

iqr:

Au_FDPP

--

0.697

6.135

3.555

0.005

0.030

0.090

0.390

57.370

0.360

Au_FDP1

--

0.691

6.183

3.597

0.005

0.030

0.090

0.370

59.760

0.340

covarnc:

Pearson:

Spearman:

no. data:

5.539

0.899

0.943

2207

(data set at full limits)

0 3 6 9 12 150

3

6

9

12

15

Au_FDPP

A u_

F D P 1

Scatter plot

Figure 13-61: Precision plot: Transworld fire assays of Salman field re-splits

Data Statisti cs

variable:

weight by:

mean:

varnc:

coefvrn:

min:

q1:

median:

q3:

max:

iqr:

Au_FDPP

--

0.697

6.135

3.555

0.005

0.030

0.090

0.390

57.370

0.360

Au_FDP1

--

0.691

6.183

3.597

0.005

0.030

0.090

0.370

59.760

0.340

covarnc:

Pearson:

Spearman:

precision:

no. data:

5.539

0.899

0.943

+/-32% @ 63%CI

2207

(data set at full limits)

0 3 6 9 12 150

1

2

3

4

5

6

7

Pair Average






87

Figure 13-62: Scatter plot: Transworld CN leach assays of Salman field re-splits

Data Statistics

variable:

weight by:

mean:

varnc:

coefvrn:

min:

q1:

median:

q3:

max:

iqr:

Au_FDPP

--

1.925

5.028

1.165

0.030

0.470

0.950

2.580

11.300

2.110

Au_FDP1

--

1.937

5.541

1.215

0.020

0.500

1.080

2.180

13.220

1.680

covarnc:

Pearson:

Spearman:

no. data:

4.741

0.898

0.840

56 / 1512


0 5 10 150

5

10

15

Au_FDPP

A u_

F D P 1

Scatter pl ot

Figure 13-63: Precision plot : Transworld CN leach assays of Salman field re-splits

Data Statistics

variable:

weight by:

mean:

varnc:

coefvrn:

min:

q1:

median:

q3:

max:

iqr:

Au_FDPP

--

1.925

5.028

1.165

0.030

0.470

0.950

2.580

11.300

2.110

Au_FDP1

--

1.937

5.541

1.215

0.020

0.500

1.080

2.180

13.220

1.680

covarnc:

Pearson:

Spearman:

precision:

no. data:

4.741

0.898

0.840

+/-25% @ 61%CI

56 / 1512


0 2 4 6 8 10 12 140

1

2

3

4

5

Pair Average



13.7 Conclusion

The author considers sampling methods and assay accuracy and precision conform to industry

standards and that the sample data adequately reflect the tenor of mineralization in each of the

deposits.




88

14.0 ADJACENT PROPERTIES

14.1 Adjacent properties

Adjacent properties are not relevant to the updated mineral resource and ore reserve estimates.

15.0 METALLURGICAL TESTING

15.1 Introduction

The approach taken to metallurgical testwork was to compile master composite samples, representing

the majority of the known in pit resource established in the ARL scoping study of 2005. Variability

composite samples were also compiled from appropriate combinations of sample reserves used to

form the master composite samples.

Each of the master composite samples was subjected to gravity/leach testing to assess the potential

for gravity concentration and determine the optimum leach conditions. The variability composite

samples were then subjected to testwork using the optimum conditions established from the master

composite sample testwork. The rationale was that the results obtained on the master composite

samples, rather than any of the minor components of the total resource, should dictate the process

design parameters.

The complete testwork programme comprised the following:

• Unconfined compressive strength (UCS) determinations.

• SMC Drop-Weight Testwork.

• JK Drop Weight Testwork (for SAG mill amenability).

• Bond Abrasion Index (Ai) Determinations.

• Bond Rod Work Index (RWi) Determinations.

• Bond Ball Work Index (BWi) Determinations.

• Head Assay Analysis.

• Cyanidation Optimisation Testwork.

• Carbon Adsorption Testwork.

• Cyanide Destruction Testwork.

• Thickening and Viscosity Testwork on Slurries.

• Arsenic precipitation.

• Tailings Consolidation.

• Tailings Geochemistry.




89

15.2 Composite Samples and Sample Preparation

15.2.1 General

Three main types of composite samples were prepared:

• Comminution composite samples, used to determine crushing, grinding and abrasion

parameters.

• Leach master composite samples representing the main ore types and oxidation zones, used

to assess the potential for gravity concentration and determine the optimum leach conditions.

• Leach variability composite samples representing different pits, rock types and oxidation

zones, used to assess the effect of varying pits, rock types and oxidation levels on gravity and

leach recoveries.

15.2.2 Comminut ion Composites

Five master comminution composite samples and twenty six comminution variability samples were

selected and compiled from portions of whole PQ diamond drill core.

Advanced Media Competency (AMC) and JK Drop Weight tests were performed with the master

composite samples and all the comminution composite samples were subjected to SAG Mill

Comminution (SMC), Bond Rod and Ball Mill Work Index and Bond Abrasion Index testing.

Due to the specific requirements of the feed sizing for the AMC and JK Drop Weight tests, separate

composites were created for the AMC and JK Drop Weight testing of each ore type. As a

consequence of their low strength and the resulting lack of suitably sized “lump”, no samples were

produced for AMC and JK Drop Weight testing of most of the oxide ores and some of the transition

ores.

Details of the comminution composite samples are listed in Table 15-1Table 15-1 to Table 15-7.

Table 15-1 Anwia Comminut ion Transition Master Composite Sample

Drill Hole No From (m) To (m)

AWDD 079 27.2 55.2

AWDD 080 38.1 45

AWDD 081 40.7 46

AWDD 082 24.8 38.4




90

Table 15-2 Anwia Comminut ion Sulphide Master Composite Sample


AWDD 079 55.2 84.2

AWDD 081 51.9 55

AWDD 082 38.4 56

AWDD 085 57 94

Table 15-3 Anwia Comminut ion Variability Composi te Samples

Sample No. Descripti on Drill Hole From (m) To (m)

1 Transition, Weakly Weathered AWDD 079 27.2 45.1


3 Sulphide, Fresh AWDD 079 55.2 84.2

4 Oxide, Very-Moderately Weathered AWDD 079 4.5 27.6

5 Oxide, Weakly Weathered AWDD 079 27.6 38.1





10 Oxide, Very Weathered AWDD 082 0.0 10.0

11 Oxide, Moderately Weathered AWDD 082 10.0 24.8


13 Sulphide, Weakly Weathered AWDD 082 38.4 49.1




Table 15-4 Salman Comminut ion Oxide Master Composite Sample


SNDD 671 0 16

SNDD 672 0 36.2

SNDD 673 26 43

SNDD 674 0 23.9

SNDD 678 0 26.5

Table 15-5 Salman Comminut ion Transition Master Composite Sample


SNDD 671 16 36.8

SNDD 672 36.2 54

SNDD 673 43 47.7

SNDD 674 13 56.8

SNDD 675 8.2 11.2

SNDD 678 26.5 35.8




92

The leach master composite samples were produced to allow testwork to be conducted to establish

the optimum leach conditions for each ore type and oxidation zone.

All of the leach master composite samples were compiled from samples derived from multiple intervals

from multiple reverse circulation (RC) drill holes. The weight of each contributing sample was

apportioned to obtain a blended grade within ±10% of the mill feed grade determined in the preliminaryin pit resources and included approximately 10% waste dilution.

The leach master composite samples are listed in Table 15-8, together with the resource tonnes that

each composite represented. In total, they represented 94.2% of the in pit resource identified in the

scoping study.

Table 15-8 Leach Master Composite Samples

Leach Master Composite% of Resources Tonnes

Represented% of Resource Tonnes

Salman Oxide 32.2 36.2

Salman Transition 14.8 16.7

Anwia Oxide 13.9 13.9

Anwia Transition 9.8 9.8

Anwia Sulphide 23.4 23.4

15.2.4 Leach Variability Composites

The leach variability composite samples were selected so that they represented the main rock types

and oxidation zones in each mine area.

The weight of each contributing sample was apportioned to obtain a blended grade within ±10% of the

mill feed grade determined in the preliminary in pit resources and included approximately 10% waste

dilution.

The number of drill holes used to form each leach variability composite sample was varied according

to the proportion of the resource tonnage identified in the scoping study, ie, the number of drill holes

used for each composite increased with increasing resource tonnes.

The leach variability composite samples were tested using the optimum leach conditions determined

by testing of the corresponding leach master composite samples.

The leach variability composite samples are listed in Table 15-9 and Table 15-10.




93

Table 15-9 Salman Leach Variability Composite Samples

Leach VariabilityComposite ID

Composite Series Composite Details Drill Hole From (m) To (m)

1 Greywacke Oxide Akanko Central AKRC121 16 23

2 Greywacke Oxide Nugget Footwall SNRC562

SNRC564

9

18

14

233 Greywacke Oxide Salman Central SNRC568

SNRC570

SNRC571

SNRC572

SNRC576

SNRC583

16

0

1

58

3

10

26

53

33

64

27

13

4 Greywacke Oxide Salman North SNRC540

SNRC544

SNRC546

SNRC584

SNRC585

4

14

14

7

1

14

16

16

15

8

5 Greywacke Oxide Salman South SNRC573SNRC574

1451

3158

6 Greywacke Oxide Salman SW SNRC579

SNRC580

SNRC581

40

15

21

51

21

29

7 Greywacke Oxide Teberu Footwall SNRC551

SNRC552

SNRC553

SNRC586

16

7

19

20

20

17

26

22

8 GreywackeTransition

Salman Central tr-lo SNRC567

SNRC568

SNRC569

SNRC572SNRC576

SNRC582

37

34

55

9036

27

46

43

64

10140

35


Salman North tr-lo SNRC544 19 39


Salman South tr-lo SNRC573

SNRC574

63

80

78

90


Akanko Central tr-up AKRC100

AKRC106

AKRC121

18

15

23

20

20

27


Salman Central tr-up SNRC565

SNRC570

SNRC571SNRC572

SNRC576

SNRC583

17

17

3064

9

0

28

55

3569

36

10


Salman North tr-up SNRC543

SNRC544

SNRC545

SNRC585

7

12

22

8

17

19

24

10


Salman South tr-up SNRC574 58 66




94

Table 15.9 Salman Leach Variability Composite Samples (continued)


Composite Series Composite Details Drill Hole From (m) To (m)

15 Greywacke Transition Nugget Footwall SNRC562 7 28

16 Greywacke Transition Salman SW SNRC579SNRC580

SNRC581

3513

20

4015

21

17 Greywacke Transition Teberu Footwall tr SNRC551

SNRC586

20

22

23

27

18 Granite Oxide Akanko Central AKRC106

AKRC117

AKRC118

AKRC120

AKRC123

21

27

6

0

1

28

32

15

18

14

19 Granite Oxide North Hill SNRC548

SNRC549

37

7

38

14

20 Granite Oxide Salman North SNRC538

SNRC545

SNRC546

SNRC547

0

24

16

0

14

36

23

8

21 Granite Transition Salman North tr-lo SNRC540 57 68

22 Granite Transition Akanko Central tr-up AKRC100

AKRC106

AKRC117

AKRC120

AKRC123

11

16

32

18

9

18

26

34

22

11

23 Granite Transition Salman North tr-up SNRC538

SNRC543

SNRC545

14

17

28

17

26

38

Sulphide Comp #1 Sulphide Nugget Footwall SNRC561

SNRC563

SNRC564

36

41

27

44

55

31

Sulphide Comp #2 Sulphide Salman Central SNRC570 55 60

Sulphide Comp #3 Sulphide Salman Central SNRC575 74 88

Sulphide Comp #4 Sulphide Salman North SNRC537 60 76



Sulphide Comp #7 Sulphide Salman North, SNRC541 79 95


Sulphide Comp #9 Sulphide Salman South SNRC574 111 123Sulphide Comp #10 Sulphide Salman South SNRC577 82 120

Sulphide Comp #11 Sulphide Salman SW SNRC579 51 69

Sulphide Comp #12 Sulphide Teberu Footwall SNRC554 32 44

Sulphide Comp #13 Sulphide Teberu Footwall SNRC560 31 39

Sulphide Comp #14 Sulphide Akanko Central AKRC103 35 45



tr-lo: lower transition, tr-up: upper transition




95

Table 15-10 Anwia Leach Variabili ty Composite Samples


Composite Series Drill Hole From (m) To (m)

24 Oxide AWRC074 0 14

25 Oxide AWRC076 22 4426 Oxide AWRC078 1 22

28 Sulphide AWRC075 63 98




32 Transition AWRC074 41 48






15.2.5 Head Assays

Head assays for the master and variability leach composite samples are summarised in Table 15-11 to

Table 15-13 The master leach composite sample gold head grades were within ±10% of the scoping

study mill feed grades with the exception of the Anwia oxide master composite sample. Several

sample intervals used to create this sample were very high grade (>25g/t Au) and subject to significant

sampling and assay errors.

Table 15-11 Leach Master Composite Sample Head Assays

Composite Au*

(ppm)

Ag(ppm)

As(ppm)

Corg

(%)

Stot

(%)

Sulphide

(%)

Scoping Study

Mill Feed Grade

Au (ppm)

Salman Oxide 2.32 <0.3 1549 0.06 0.04 <0.02 2.52

Salman Transition 2.75 <0.3 2942 0.16 0.40 0.30 2.84

Anwia Oxide 3.26 0.40 387 0.05 0.02 <0.02 2.06

Anwia Transition 2.97 0.60 251 0.08 0.40 0.35 2.94

Anwia Sulphide 2.90 0.45 565 0.08 0.94 0.91 2.95

*Average of calculated head grades from leach optimisation testing




96

Table 15-12 Salman Leach Variability Composite Head Assays

Leach Variability Au Ag As Ctot Stot

Compos ite ID (ppm) (ppm) (ppm) (%) (%)

1 0.77 < 0.3 1250 0.09 <0.02

2 2.00 < 0.3 1965 0.05 0.06

3 2.77 < 0.3 1520 0.10 0.02

4 2.37 < 0.3 1706 0.07 0.05

5 1.51 < 0.3 627 0.03 <0.02

6 0.86 < 0.3 2382 0.06 0.21

7 2.21 < 0.3 2738 0.05 0.16

8 3.17 0.3 4805 0.35 0.74

9 2.62 < 0.3 3844 0.15 0.73

10 1.93 0.5 1576 0.19 0.05

11 1.56 < 0.3 784 0.17 <0.02

12 3.35 < 0.3 2159 0.15 <0.02

13 2.12 < 0.3 2662 0.15 0.5514 1.36 0.4 424 0.18 <0.02

15 1.94 < 0.3 1194 0.12 0.36

16 2.12 0.3 9296 0.09 1.25

17 3.06 < 0.3 2261 0.17 0.52

18 1.02 < 0.3 1045 0.04 <0.02

19 2.74 < 0.3 3212 0.04 0.02

20 2.59 < 0.3 3104 0.05 0.06

21 1.66 0.3 5813 0.13 0.36

22 6.63 0.4 1100 0.07 0.03

23 1.45 0.3 5293 0.05 0.35

Sulphide Comp #1 2.25 0.8 3029 0.19 0.70

Sulphide Comp #2 3.06 < 0.3 5324 0.97 0.71Sulphide Comp #3 4.12 0.3 6926 0.32 0.97

Sulphide Comp #4 6.59 < 0.3 8741 0.98 1.30

Sulphide Comp #5 1.88 < 0.3 2283 1.25 0.56

Sulphide Comp #6 1.91 < 0.3 2147 1.12 0.60

Sulphide Comp #7 7.74 < 0.3 7200 1.33 1.11

Sulphide Comp #8 2.06 < 0.3 2478 1.16 0.44

Sulphide Comp #9 5.98 < 0.3 361 1.23 0.50

Sulphide Comp #10 4.13 < 0.3 6251 1.58 1.17

Sulphide Comp #11 2.76 < 0.3 1350 2.06 0.79

Sulphide Comp #12 3.66 < 0.3 4033 1.10 0.90

Sulphide Comp #13 2.62 < 0.3 3760 1.48 0.60

Sulphide Comp #14 3.86 < 0.3 3620 0.39 0.45

Sulphide Comp #15 1.80 < 0.3 4050 0.24 0.70

Sulphide Comp #16 5.47 0.6 1591 0.96 0.16




97

Table 15-13 Anwia Leach Var iabi lit y Compos ite Head Ass ays

Leach Variability Au Ag As Ctot Stot

Compos ite ID (ppm) (ppm) (ppm) (%) (%)

24 2.03 < 0.3 429 0.09 <0.02

25 2.66 0.5 609 0.04 <0.02

26 4.63 0.7 279 0.06 0.02

28 2.08 0.5 720 0.42 0.83

29 1.97 0.5 439 0.59 1.11

30 4.74 1.0 876 1.45 1.65

31 1.97 0.3 148 0.94 0.71

32 2.76 0.6 2030 0.09 0.42

33 2.70 0.5 335 0.05 0.74

34 2.06 0.5 191 0.03 0.02

35 1.44 <0.3 429 0.07 0.02

36 4.30 0.7 323 0.05 0.67

37 1.06 <0.3 109 0.14 0.35

15.3 Comminution

15.3.1 General

Five master comminution composite samples and twenty six comminution variability samples were

selected and compiled from portions of whole PQ diamond drill core.

Advanced Media Competency and JK Drop Weight tests were performed with the master comminution

composite samples and all the comminution composite samples were subjected to SAG Mill

Comminution (SMC), Bond Rod and Ball Mill Work Index and Bond Abrasion Index testing.

15.3.2 Crushing Work Index

Bond impact crushing work index tests were performed for the Anwia and Salman comminution

transition and sulphide master composite samples. Due to the low competency exhibited by the oxide

ores no crushing work index tests were conducted with the oxide ores.

Results from the crushing work index testing are shown in Table 15-14




98

Table 15-14 Crushing Work Index Test Results

Sample Description Crushing Work Index (kWh/t)

Average Maximum Minimum Std. Dev.

Anwia Transition Master Composite 7.9 18.3 4.5 3.2

Anwia Sulphide Master Composite 20.1 32.4 7.6 6.8

Salman Transition Master Composite 6.3 10.2 3.7 2.0

Salman Sulphide Master Composite 11.3 18.1 5.1 3.9

The crushing work indices for both ores increased with reduced weathering. The Anwia ore samples

were harder than the corresponding Salman ore samples and also showed greater variability in the

range of crushing work indices.

15.3.3 Unconfined Compressive Strength Tests

Unconfined compressive strength (UCS) tests were performed for the Anwia and Salman comminution

variability composite samples. Results from the UCS testing of the comminution variability composite

samples are shown in Table 15-15 and Table 15-16. No competent core was available for some

samples and as a consequence no UCS test results are available for these samples.

Table 15-15 Anwia Variability Composite Sample Unconfined Compressive Strength

Sample No. Descrip tion Drill Hole Interval (m) UCS (MPa)



3 Sulphide, Weakly Weathered AWDD 079 60.9 60.34 Oxide, Moderately Weathered AWDD 080 17.4 4.7

5 Oxide, Moderately Weathered AWDD 080 28.8 195.2





11A Oxide, Moderately Weathered AWDD 082 15.3 10.2

11B Oxide, Weakly Weathered AWDD 082 23.8 1.4


13 Sulphide, Weakly Weathered AWDD 082 43.8 43.9


15 Oxide, Weakly Weathered AWDD 085 5.5 0.816 Sulphide, Fresh AWDD 085 78.5 134.2




99

Table 15-16 Salman Variability Composite Sample Unconfined Compressive Strength

SampleNo.

Pit Description Drill Hole Interval(m)

UCS(MPa)

1 Salman Central Oxide Greywacke, Weakly Weathered SNDD 672 28.6 120.6

2 Salman Central Transition Greywacke, Weakly Weathered SNDD 671 36.7 12.5

3 Salman Central Sulphide Greywacke, Fresh SNDD 671 56.1 17.3

5 Salman South Transition Greywacke, Weakly Weathered SNDD 674 42.0 3.2

6 Salman South Sulphide Greywacke, Fresh SNDD 673 79.0 63.5

8 Salman North Transition Greywacke, Weakly Weathered SNDD 678 35.8 13.7

9 Salman North Sulphide Greywacke, Fresh SNDD 675 34.0 *

9A Salman North Sulphide Greywacke, Fresh SNDD 678 43.0 33.3

11 Salman North Transition Granite, Weakly Weathered SNDD 676 20.5 9.4

12A Salman North Sulphide Granite, Fresh SNDD 676 49.7 87.3

12B Salman North Sulphide Granite, Fresh SNDD 677 63.0 100.5*Sample failed before testing

The Anwia samples showed a general increase in UCS with decreasing oxidation, as shown in Figure

15-1. The main exception to this trend was Anwia variability sample 5 (massive grey vein quartz)

which gave the highest UCS for all samples tested. This was a marked contrast to the other Anwia

oxide samples which were very weak and lacked any significant degree of competence.

Figure 15-1 Unconfined Compressive Strength of Anwia Ore Variability Samples

Anwia Variability Sample Unconfined Compressive Strength

0

50

100

150

200

250

1 5 - O x i d e

1 1 B - O x i d e

4 - O x i d e

7 - O x i d e

6 - O x i d e

1 - T r a n s i t i o n

1 1 A - O x i d e


9 - S u l p h i d e

1 3 - S u l p h i d e

3 - S u l p h i d e

1 4 - S u l p h i d e


1 6 - S u l p h i d e

1 2 - T r a n s i t i o n

5 - O x i d e

Composite Number and Oxidation

U C S ( M P a )




100

The Salman samples also showed a general increase in UCS with decreasing oxidation, as shown in

Figure 15-2. The main exceptions to this trend were Variability Sample 1 which gave the highest UCS

for all Salman samples tested and Variability Sample 9, which failed prior to testing.

The compressive strength of the Anwia and Salman ores is moderate to low. The compressive

strength increased with the proportion of quartz veins in the ore,

Figure 15-2 Unconfined Compressive Strength of Salman Ore Variability Samples

Salman Variability Sample Unconfined Compressive Strength

0

20

40

60

80

100

120

140

9 - S u l p h i d e


1 1 - T r a n s i t i o n



3 - S u l p h i d e

9 A - S u l p h i d e

6 - S u l p h i d e

1 2 A - S u l p h i d e

1 2 B - S u l p h i d e

1 - O x i d e

Composite Number and Oxid ation

U C S ( M P

a )

The crushing work index and UCS results indicate that jaw crushers are suitable for crushing the

Anwia and Salman ores.

15.3.4 Advanced Media Competency Tests

Only the Anwia and Salman sulphide core samples had sufficient quantities of suitably sized pieces of

core to make up the 180kg sample necessary for Advanced Media Competency (AMC) testing. The

product size distributions obtained from the AMC tests with each sample are shown in Figure 15-3 andare compared with the product size distribution obtained from the AMC testing of an extremely hard

ore.




101

Figure 15-3 Media Competency Test Product Sizing

MEDIA COMPETENCY TEST - PRODUCT SIZING

0

10

20

30

40

50

60

70

80

90

100

100 1000 10000 100000

SIZE (µm)

C U M U L A T I V E W E I G H T

R E T A I N E D ( %

Anwia Sulphide

Salman Sulphide

Extremely Hard Ore

Results from impact work index testing of the “survivor” particles from the AMC tests are shown in

Figure 15-4 and are compared with the impact work index results obtained from testing of ore from the

Three Mile Hill mine in Western Australia.

Figure 15-4 Impact Work Index Testing of AMC Test Survivors

Impact Work Index Testing of AMC Test Survivors

0

10

20

30

40

50

60

70

80

22 32 45 64 89

Mean Partic le Size (mm)

I m p a c t W

o r k I n d e x ( k W h / t )

Anwia Sulphide

Salman Sulphide

Three Mile Hill




102

The Salman and Anwia sulphide comminution master composite samples showed similar responses to

the AMC testing, with both samples exhibiting moderate to low resistance to tumbling and impact

breakage. Both ore types are considered amenable to SAG-Ball milling, with a very low potential for

build-up of critical size particles within the SAG mill.

15.3.5 JK Drop Weight Tests

JK Drop weight tests were performed for the Anwia transition and sulphide master composites and the

Salman oxide, transition and sulphide master composite samples.

No JK drop weight test was conducted with the Anwia oxide ore due to the lack of sufficient suitably

sized material in the available core.

Results from the JK drop weight testing are shown in Table 15-17.

Table 15-17 Master Compos ite JK Drop Weight Test Parameters

Sample A b A*b t10 @ 1kWh/t ta

Anwia Transition Master Composite 53.70 2.45 131.5 49.1 1.50

Anwia Sulphide Master Composite 59.16 1.33 78.5 43.5 0.65

Salman Oxide Master Composite 69.75 3.75 261.4 68.1 2.63

Salman Transition Master Composite 72.95 4.09 298.7 71.7 4.33

Salman Sulphide Master Composite 60.09 1.41 84.9 45.5 0.90

Anwia Transition Master Composite

The Anwia Transition Master Composite has an A*b value of 131.5, which puts this material in the

very soft range of resistance to impact breakage. 90.7% of the 2,140 ore types contained In the

JKTech database have lower A*b values (are harder than the test sample).

With a ta of 1.50, the Anwia Transition Master Composite falls into the very soft abrasion range with

92.7% of the 2,255 ore types contained in the JKTech database having lower ta values (have greater

abrasion resistance than the test sample).

At low input energy levels (0.25 and 1.0 the Anwia Transition Master Composite displays decreasing

resistance to breakage with increasing particle size, as shown in Figure 15-5, with t10 values (t10 is

defined as the percentage of material passing 1/10th of the initial particle size) increasing withincreasing particle size. This indicates that particles in the 100 to 200mm size range (normal media

size) may not be strong enough to survive SAG milling.

The crusher model parameters for Anwia Transition Master Composite also show a decrease in

impact resistance with increasing particle size, as shown in Figure 15-6




103

Figure 15-5 Variation of Impact Resistance with Particle Size – Anwia Transit ion

Anw ia Transition Master Compos ite

0

10

20

30

40

50

60

70

80

90

100

10 15 20 25 30 35 40 45 50 55 60

Particle Size (mm)

t 1 0 ( % )

0.25 kWh/t

1.0 kWh/t

2.5 kWh/t

Figure 15-6 Variation of Crushing Energy with Partic le Size – Anwia Transiti on

Anw ia Transition Master Composite

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

0.50

10 15 20 25 30 35 40 45 50 55 60

Particle Size (mm)

E c s ( k W h / t )

10% minus t10

20% minus t10

30% minus t10




104

Anwia Sulphide Master Composite

The Anwia Sulphide Master Composite has an A*b value of 78.5, which puts this material in the soft

range of resistance to impact breakage. 77.4% of the 2,140 ore types contained In the JKTechdatabase have lower A*b values (are harder than the test sample).

With a ta of 0.65, the Anwia Sulphide Master Composite falls into the moderately soft abrasion range

with 69.3% of the 2,255 ore types contained in the JKTech database having lower ta values (have

greater abrasion resistance than the test sample).

The Anwia Sulphide Master Composite also shows decreasing resistance to breakage with increasing

particle size, as shown in Figure 15-7 and Figure 15-8

Salman Oxide Master Composite

The Salman Oxide Master Composite has an A*b value of 261.8, which puts this material in the very

soft range of resistance to impact breakage. 98.3% of the 2,140 ore types contained In the JKTech

database have lower A*b values (are harder than the test sample).

With a ta of 2.63, the Salman Oxide Master Composite falls into the very soft abrasion range with

98.5% of the 2,255 ore types contained in the JKTech database having lower ta values (have greater


The data for Salman Oxide Master Composite follows a trend of decreasing slope with decreasing

energy (Ecs values), as shown in Figure 15-9.

The crusher model parameters for Salman Oxide show no variation in impact resistance with particle

size, as shown in Figure 15-10.




105

Figure 15-7 Variation of Impact Resistance with Particle Size – Anwia Sulphide

Anwia Sulph ide Master Compos ite

0

10

20

30

40

50

60

70

80

90

100

10 15 20 25 30 35 40 45 50 55 60

Particle Size (mm)

t 1 0 ( % )

0.25 kWh/t

1.0 kWh/t

2.5 kWh/t

Figure 15-8 Variation of Crushing Energy with Partic le Size – Anwia Sulphide

Anwia Sulphide Master Composite

0.0

0.1

0.2

0.3

0.4

0.5

0.6

0.7

0.8

10 15 20 25 30 35 40 45 50 55 60

Particle Size (mm)

E c s

( k W h / t )

10% minus t10

20% minus t10

30% minus t10




106

Figure 15-9 Variation of Impact Resistance with Particle Size – Salman Oxide


0

10

20

30

40

50

60

70

80

90

100

10 15 20 25 30 35 40 45 50 55 60

Particle Size (mm)

t 1 0 ( % )

0.25 kWh/t

1.0 kWh/t

2.5 kWh/t

Figure 15-10 Variation of Crush ing Energy wi th Particle Size – Salman Oxide


0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

0.50

10 15 20 25 30 35 40 45 50 55 60

Particle Size (mm)

E c s

( k W h / t )

10% minus t10

20% minus t10

30% minus t10




107

Salman Transition Master Composite

The Salman Transition Master Composite has an A*b value of 298.7, which puts this material in the

very soft range of resistance to impact breakage. 98.8% of the 2,140 ore types contained In theJKTech database have lower A*b values (are harder than the test sample).

With a ta of 4.33, the Salman Transition Master Composite falls into the very soft abrasion range and

has the highest ta (lowest abrasion resistance) of all ore types in the JKTech database.

The data for Salman Transition Master Composite follows a trend of decreasing slope with decreasing

energy (Ecs values), as shown in Figure 15-11. The crusher model parameters for Salman Transition

Master Composite show a slight decrease in impact resistance with increasing particle size, as shown

in Figure 15-12.

Salman Sulphide Master Composite

The Salman Sulphide Master Composite has an A*b value of 84.9, which puts this material in the soft

range of resistance to impact breakage. 80.1% of the 2,140 ore types contained In the JKTech

database have lower A*b values (are harder than the test sample).

With a ta of 0.90, the Salman Sulphide Master Composite falls into the soft abrasion range with 81.6%

of the 2,255 ore types contained in the JKTech database having lower ta values (have greater


The data for Salman Transition Master Composite follows a trend of decreasing slope with decreasing

energy (Ecs values), as shown in Figure 15-13.

The crusher model parameters for Salman Sulphide Master Composite show a marked increase in

impact resistance with decreasing particle size, as shown in Figure 15-14.




108

Figure 15-11 Variation of Impact Resistance with Particle Size – Salman Transition


0

10

20

30

40

50

60

70

80

90

100

10 15 20 25 30 35 40 45 50 55 60

Particle Size (mm)

t 1 0 ( % )

0.25 kWh/t

1.0 kWh/t

2.5 kWh/t

Figure 15-12 Variation of Crushing Energy with Particl e Size – Salman Transiti on


0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

0.50

10 15 20 25 30 35 40 45 50 55 60

Particle Size (mm)

E c s

( k W h / t )

10% minus t10

20% minus t10

30% minus t10




109

Figure 15-13 Variation of Impact Resistance with Particle Size – Salman Sulphide

Salman Sulph ide Master Composite

0

10

20

30

40

50

60

70

80

90

100

10 15 20 25 30 35 40 45 50 55 60

Particle Size (mm)

t 1 0 ( % )

0.25 kWh/t

1.0 kWh/t

2.5 kWh/t

Figure 15-14 Variation of Crush ing Energy w ith Particle Size – Salman Sulphide

Salman Sulph ide Master Composite

0.00

0.10

0.20

0.30

0.40

0.50

0.60

0.70

0.80

10 15 20 25 30 35 40 45 50 55 60

Particle Size (mm)

E c s

( k W h / t )

10% minus t10

20% minus t10

30% minus t10




110

15.3.6 SMC Testing

SAG Mill Comminution (SMC) tests were performed for the Anwia and Salman variability samples.

Results from the JK drop-weight tests conducted with the master comminution composites were used

to calibrate the DWi versus A and b correlations. Database values were used where no results were

available for a comparable sample. Results from the SMC testing are shown in Table 15-18.

Table 15-18 SMC Test Results

Test Calibrati on SMC SMC Test Derived Values

Sample Sample SG DWi A B

Anwia Transition Master Anwia Transition Master 2.58 2.2 69.2 2.02

Anwia Sulphide Master Anwia Sulphide Master 2.71 3.5 61.2 1.27 Anwia Variability 1 Anwia Transition Master 2.47 2.0 71.6 2.08

Anwia Variability 2 Anwia Transition Master 2.58 2.6 72.0 1.66 Anwia Variability 3 Anwia Sulphide Master 2.66 2.8 62.6 1.50 Anwia Variability 4 SMC Database 2.57 1.1 73.7 3.21

Anwia Variability 5 SMC Database 2.28 2.1 73.4 1.50 Anwia Variability 6 SMC Database 2.58 1.5 68.5 2.48

Anwia Variability 7 SMC Database 2.49 2.0 70.6 1.81 Anwia Variability 8 Anwia Transition Master 2.70 3.4 73.5 1.27

Anwia Variability 9 Anwia Sulphide Master 2.69 3.8 64.4 1.09

Anwia Variability 11 SMC Database 2.34 1.0 75.8 3.15 Anwia Variability 12 Anwia Transition Master 2.56 2.4 73.9 1.73 Anwia Variability 13 Anwia Sulphide Master 2.64 2.8 67.7 1.41

Anwia Variability 14 Anwia Sulphide Master 2.65 4.1 63.7 1.01

Anwia Variability 16 Anwia Sulphide Master 2.71 3.5 63.0 1.23Salman Oxide Master Salman Oxide Master 2.31 0.7 77.9 4.17Salman Transition Master Salman Transition Master 2.39 0.8 78.4 3.92

Salman Sulphide Master Salman Sulphide Master 2.77 3.5 66.4 1.21

Salman Variability 1 Salman Oxide Master 2.46 2.0 70.6 1.76Salman Variability 2 Salman Transition Master 2.11 0.7 63.7 4.53Salman Variability 3 Salman Sulphide Master 2.62 3.4 66.1 1.18

Salman Variability 4 Salman Oxide Master 2.33 1.2 73.0 2.64Salman Variability 5 Salman Transition Master 2.43 0.6 71.0 5.53

Salman Variability 6 Salman Sulphide Master 2.58 2.5 64.8 1.61

Salman Variability 7 Salman Oxide Master 2.15 0.5 75.4 5.43Salman Variability 8 Salman Transition Master 2.30 1.6 79.8 1.86Salman Variability 9 Salman Sulphide Master 2.59 3.0 68.5 1.24

Salman Variability 10 SMC Database 2.14 0.9 79.0 3.09Salman Variability 11 SMC Database 2.59 2.4 77.1 1.38

Salman Variability 12 SMC Database 2.68 5.0 77.3 0.69

The Salman and Anwia ores exhibit moderate to very low resistance to tumbling and impact breakage.

None of the JK Dropweight, SMC or AMCT tests suites showed any indication of the potential for

build-up of critical size particles within the SAG mill.

Due to the low resistance to impact breakage, the Anwia and Salman ores are not considered to be

suitable for fully autogenous (FAG) milling.




111

The wide range of variability in comminution characteristics is likely to make the control of a single

stage SAG mill difficult. Blending of feed material will be necessary to minimise fluctuations in the

feed and appropriate instrumentation for SAG mill load control and measurement will be necessary to

ensure stable operation of the milling circuit.

A SAG-Ball mill circuit can be expected to provide a more consistent product and throughput and isbetter able to accommodate variations in feed size and ore hardness.

15.3.7 Bond Comminut ion Tests

Bond abrasion, rod and ball mill index tests were performed for the Anwia and Salman master and

variability comminution composite samples. Results from the testing are shown in Table 15-19.

Table 15-19 Bond Test Results

Test Ai Rod Wi Ball Wi (kWh/t)

Sample (g) (kWh/t) 106µm* 63µm*

Anwia Transition Master 0.2742 12.4 11.0

Anwia Sulphide Master 0.2003 13.1 12.5

Anwia Variability 1 0.1876 12.8 9.1


Anwia Variability 3 0.2316 13.0 13.0 15.0







Anwia Variability 10 0.0212 7.7 Levin Test Anwia Variability 11 0.2183 9.6 8.6




Anwia Variability 15 0.0234 5.7 Levin Test


Salman Oxide Master 0.1503 9.5 Levin Test

Salman Transition Master 0.1187 10.0 Levin Test

Salman Sulphide Master 0.1173 13.2 7.8

Salman Variability 1 0.1698 7.3 Levin Test

Salman Variability 2 0.0928 8.2 6.1


Salman Variability 4 0.1701 5.6 Levin TestSalman Variability 5 0.3016 7.1 Levin Test







Salman Variability 12 0.2385 14.3 12.1*Closing screen aperture




112

Some samples were too fine to perform Bond ball mill work index tests and Levin tests were

performed with these samples. Results from the testing are shown in Table 15-20.

Table 15-20 Levin Test Results

Sample P80 (µm)

Feed 2.5 kWh/t 5 kWh/t 10 kWh/t

Anwia Variability 10 1466 273 135 90*

Anwia Variability 15 1473 92 72* 49*

Salman Oxide Master 1228 172 96 57

Salman Transition Master 1575 222 116 56

Salman Variability 1 1657 231 111 51




Salman Variability 10 1407 149 85 64*Estimated By Extrapolation

Results from the Bond Comminution testing are consistent with the results from the AMCT and JK

Drop Weight tests, ie, the ores exhibit moderate to low resistance to breakage and abrasion. Both the

Salman and Anwia ores are considered amenable to SAG-Ball milling.

15.4 Mineralogy

15.4.1 General

Five master leach composite samples and selected Salman transition leach variability samples were

submitted to Roger Townend and Associates for mineralogical examination. The method used for the

examinations was:

• TBE separation to produce sinks (SG>~3) and floats (SG<~3) fractions.

• Optical and SEM examination of "sinks" for Au and major minerals.

• XRD of floats.

Results from the examinations are summarised in the following sections.

15.4.2 Anwia Oxide Master

The sinks fractions were comprised mainly of goethite with some titanium oxides and trace amounts of

pyrite and arsenopyrite.

Four gold occurrences were detected; all were associated with goethite and were between 3µm and

6µm in size.




113

15.4.3 Anwia Transition Master

The sinks fractions were comprised mainly of goethite with pyrite and trace amounts of arsenopyrite,

chalcopyrite, covellite, bornite, pyrrhotite, sphalerite, galena and titanium oxides.

Twelve gold occurrences were detected; 8 were associated with pyrite, 3 were fully liberated and 1

was in goethite. The gold particles ranged in size from 1µm as inclusions in pyrite, up to 40µm free

gold particles and 50µm by 5µm rims on coarse pyrite.

15.4.4 Anwia Sulphide Master

The sinks fractions were comprised mainly of pyrite with arsenopyrite and trace amounts of

chalcopyrite, covellite, bornite, pyrrhotite, sphalerite, galena and titanium oxides.

Nine gold occurrences were detected; 7 were associated with pyrite and 2 were associated with

arsenopyrite. The gold particles ranged in size from 1µm as inclusions in pyrite, up to 50µm by 3µm

rims on coarse pyrite.

15.4.5 Salman Oxide Master

The sinks fractions were comprised mainly of goethite with some titanium and manganese oxides and

trace amounts of pyrite, arsenopyrite, chalcopyrite, pyrrhotite, sphalerite and galena.

Four gold occurrences were detected; all were associated with goethite and were between 3µm and

6µm in size.

15.4.6 Salman Transi tion Master

The sinks fractions were comprised mainly of goethite, pyrite and arsenopyrite with some marcasite

and titanium and manganese oxides and trace amounts of, chalcopyrite and sphalerite.

One 1.5µm gold particle was detected in goethite.

The amount of gold observed was much lower than that expected from the assayed gold content.

This may indicate the possibility of some of the gold being present in solid solution in the pyrite or

arsenopyrite.

15.4.7 Salman Central Upper Transition (Variabilit y Composite 12)

The sinks fractions were comprised mainly of goethite with minor rutile and trace arsenopyrite, pyrite

and graphite.

Five types of gold particles were detected; 3 within goethite, 1 in quartz and 1 liberated. The gold

particles ranged in size from 1µm inclusions in coarse goethite up to one 150µm free particle.




114

15.4.8 Salman Central Lower Transition (Variability Composite 8)

The sinks fractions were comprised mainly of goethite, arsenopyrite and pyrite with rutile and graphite

and trace chalcopyrite, covellite and sphalerite.

Four types of gold particles were detected; 3 within arsenopyrite and 1 in pyrite. The gold particles

ranged in size from 0.5µm to 2µm inclusions in arsenopyrite up to one 7µm particle in pyrite.

15.4.9 Salman North Upper Transition (Variabil ity Composite 13)

The sinks fractions were comprised mainly of goethite, arsenopyrite and pyrite with rutile and graphite

and trace chalcopyrite covellite, galena and sphalerite.

Three types of gold particles were detected, all within arsenopyrite, and ranged in size from 0.5µm to

17µm.

15.4.10 Salman North Lower Transit ion (Variability Composite 9)

The sinks fractions were comprised mainly of pyrite, arsenopyrite and rutile with goethite, chalcopyrite

and marcasite and galena and sphalerite.

Three types of gold particles were detected, all within arsenopyrite, and ranged in size from 0.5µm to

10µm.

15.4.11 Salman North Upper Transit ion (Variabilit y Composite 23) – Granite

The sinks fractions were comprised mainly of pyrite, arsenopyrite and rutile with marcasite, magnetite

and goethite and trace pyrrhotite, chalcopyrite and galena.

Three types of gold particles were detected, 2 were within arsenopyrite and 1 within pyrite. The

particles ranged in size from 1µm to 6µm.

15.4.12 Salman North Lower Transition (Variability Composite 21) – Granite

The sinks fractions were comprised mainly of arsenopyrite with pyrite, marcasite and rutile and trace

chalcopyrite, pyrrhotite and galena.

Four types of gold particles were detected; all were within arsenopyrite. The particles ranged in size

from 2µm to 16µm.

Bismuth and bismuth tellurides were also detected as 1µm to 2µm inclusions in arsenopyrite.




115

15.4.13 Salman Sulphide AMC Comminut ion Composite

The sinks fractions were comprised mainly of arsenopyrite, pyrite and goethite with marcasite, rutile

and graphite and trace pyrrhotite, chalcopyrite, covellite, bornite and galena.

Three types of gold particles were detected; all were within arsenopyrite. The particles ranged in size

from 0.5µm to 16µm.

A large proportion of the gold occurrences observed in the Anwia ore samples are associated with

pyrite, however the ore can be considered to be “free-milling”, with no evidence of any “refractory” gold

occurring in the Anwia ores.

The Salman ores contain higher levels of graphitic carbon and arsenopyrite than the Anwia ores.

Some of the gold in the transition and sulphide ores is likely to be present in solid solution in the pyrite

or arsenopyrite. The presence of organic carbon supports the use of carbon-in-leach processing to

minimise soluble gold losses to “pre-robbing” carbon.

15.5 Thickening

15.5.1 General

Samples of master leach composite slurries were prepared by Ammtec and submitted to Outokumpu

Technology for flocculant screening and dynamic bench scale thickening testwork.

15.5.2 Flocculant Screening Tests

The flocculant screening tests showed a low charge anionic polyacrylamide (SNF AN910VHM)

flocculant achieved the most rapid settling rate and a good clarity at a reasonable flocculant dosage

for all ore types, except for the oxide ores which showed a better response to a medium charge

anionic polyacrylamide (SNF AN923VHM).

15.5.3 Dynamic Thickening Tests

All dynamic thickening tests were performed at pH 10.5 with AN923VHM flocculant for the oxide

samples and AN910VHM flocculant for the transition and sulphide samples.

Anwia Oxide Master

Results from the dynamic thickening testing of the Anwia oxide master sample are shown in Table

15-21.




116

Table 15-21 Anwia Oxide Master Dynamic Thickening Tests

Feed Floc Underflow Overflow

Run

No.

Solids

(t/m².h) (%w/w)

Liquor Rise Rate

(m/h)

Dose

(g/t) (%w/w)

Shear

(Pa)

Clarity

(ppm)1 1.03 12.1 7.8 15 54.7 50 80

2 1.01 12.1 7.7 10 51.7 29 170

3 1.20 11.8 9.4 15 47.8 * 440

4 1.15 11.6 9.2 43 49.2 * 85

5 1.14 11.7 9.0 32 47.5 * 100

6 0.96 12.0 7.4 56 50.7 * 50*No measurement

Anwia Transition Master

Results from the dynamic thickening testing of the Anwia transition master sample are shown in Table15-22.

Table 15-22 Anwia Transition Master Dynamic Thickening Tests


RunNo.

Solids

(t/m².h) (%w/w)

Liquor Rise Rate

(m/h)

Dose

(g/t)

(%w/w) Shear

(Pa)

Clarity

(ppm)

1 0.97 11.8 7.6 21 58.1 34 40

2 1.18 12.0 9.0 22 54.8 * 70*No measurement

Anwia Sulphide Master

Results from the dynamic thickening testing of the Anwia sulphide master sample are shown in Table

15-23.

Table 15-23 Anwia Sulphide Master Dynamic Thickening Tests


Run

No.

Solids

(t/m².h)

(%w/w) Liquor Rise Rate

(m/h)

Dose

(g/t)

(%w/w) Shear

(Pa)

Clarity

(ppm)

1 1.05 12.3 7.8 10 56.4 111 30

2 1.15 11.7 9.1 21 51.3 * 70*No measurement

Salman Oxide Master

Results from the dynamic thickening testing of the Salman oxide master sample are shown in Table

15-24.




117

Table 15-24 Salman Oxide Master Dynamic Thickening Tests


Run

No.

Solids

(t/m².h)


(m/h)

Dose

(g/t)

(%w/w) Shear

(Pa)

Clarity

(ppm)

1 0.65 12.6 4.8 19 53.9 59 20

2 0.85 12.6 6.2 19 53.4 43 20

3 1.11 12.6 8.1 18 50.4 32 50

4 1.06 12.6 7.7 14 49.5 28 110

Salman Transition Master

Results from the dynamic thickening testing of the Salman transition master sample are shown in

Table 15-25.

Table 15-25 Salman Transit ion Master Dynamic Thickening Tests


Run

No.

Solids

(t/m².h)


(m/h)

Dose

(g/t)

(%w/w) Shear

(Pa)

Clarity

(ppm)

1 0.89 11.7 7.0 23 53.0 31 10

2 0.96 11.8 7.5 19 48.6 540

3 1.22 12.1 9.3 20 47.4 10

All of the ore types tested can be flocculated and thickened to produce underflow densities between

50% and 55% solids w/w with an overflow of suitable clarity for recycling.

The Salman transition ore is the most difficult to flocculate and settle and the Anwia sulphide ore the

easiest, with the former requiring a higher flocculant addition.

A single high rate thickener would suffice for the intended thickening application, with a flocculant

addition of 20 to 25g/t and a design solids settling rate of 0.9 t/m2.h.

15.6 Viscosity

15.6.1 General

Viscosity tests were conducted using slurry samples prepared from master leach composites for

gravity and leach testing. Results from the viscosity testing of slurries are summarised in the following

sub-sections.

15.6.2 Anwia Oxide Master

Results from viscosity testing of Anwia oxide master slurries at varying pulp density are summarised in

Table 15-26.




118

Table 15-26 Anwia Oxide Master Viscosity Test Summary

Sample SHEAR 40%w/w 45%w/w 50%w/w 55%w/w 60%w/w

ID RATE

(sec-1)

Viscosity

(cP)

Viscosity

(cP)

Viscosity

(cP)

Viscosity

(cP)

Viscosity

(cP)

Pre-CIL 4.1 n/a n/a 500 1120 3288

No Gravity 7.4 n/a n/a 298 668 2020

13.1 n/a n/a 188 430 1292

21.8 n/a n/a 124 282 853

38.9 n/a 59 86 185 543

67.4 37 52 76 130 350

119.2 49 59 80 127 239

209.4 70 81 96 140 239

Pre-CIL 4.1 n/a n/a n/a 644 991

Gravity 7.4 n/a n/a n/a 280 673

Tailings 13.1 n/a n/a n/a 187 453

21.8 n/a n/a n/a 129 312

38.9 n/a n/a 67 89 20067.4 35 43 62 82 135

119.2 42 52 68 84 135

209.4 58 73 89 98 146

CIL 4.1 n/a n/a n/a n/a 594

Tailings 7.4 n/a n/a n/a n/a 341

13.1 n/a n/a n/a n/a 242

21.8 n/a n/a n/a 86 187

38.9 n/a n/a 49 81 148

67.4 35 43 52 80 148

119.2 46 51 68 90 158

209.4 60 71 98 121 186

Anwia oxide ore displays moderate viscosity at high pulp density and exhibits shear thinning, with the

viscosity decreasing with increasing shear rates. Viscosity also reduced following gravity separation

and leach processing. The effect of varying pulp density and shear rate for Anwia oxide ore is shown

in Figure 15-15. Anwia oxide ore could be leached at up to 50% solids w/w without the pulp viscosity

adversely affecting mixing/interstage screening.




119

Figure 15-15 Variation in Viscosity with Shear Rate and Pulp Density – Anwia Oxide

Anw ia Oxide - Pre -CIL, No Gravi ty

10

100

1000

10000

1 10 100 1000Shear Rate (sec-1)

V i s c o s

i t y ( c P )

40% SOLIDS (w /w )

45% SOLIDS (w /w )

50% SOLIDS (w /w )

55% SOLIDS (w /w )

60% SOLIDS (w /w )

15.6.3 Anwia Transition Master

Results from viscosity testing of Anwia transition master slurries at varying pulp density are

summarised in Table 15-27. Anwia transition ore displays low viscosity at high pulp density and could

be leached at up to 50% solids w/w without the pulp viscosity adversely affecting mixing/interstage

screening.

Table 15-27 Anwia Transit ion Master Viscosity Test Summary


ID RATE

(sec-1)

Viscosity

(cP)

Viscosity

(cP)

Viscosity

(cP)

Viscosity

(cP)

Viscosity

(cP)

Pre-CIL 4.1 n/a n/a n/a n/a n/a

Gravity 7.4 n/a n/a n/a n/a n/a

Tailings 13.1 n/a n/a n/a n/a n/a

21.8 n/a n/a n/a n/a 101

38.9 n/a n/a n/a 58 92

67.4 31 39 47 55 91

119.2 42 51 57 64 107

209.4 64 64 80 100 121

CIL 4.1 n/a n/a n/a n/a 1279


13.1 n/a n/a n/a n/a 253

21.8 n/a n/a n/a n/a 185

38.9 n/a n/a 61 76 176

67.4 34 44 63 74 196

119.2 46 52 73 100 216

209.4 63 71 94 120 253




120

15.6.4 Anwia Sulphide Master

Results from viscosity testing of Anwia sulphide master slurries at varying pulp density are

summarised in Table 15-28. Anwia sulphide ore displays very low viscosity at high pulp density and

could be leached at up to 50% solids w/w without the pulp viscosity adversely affecting

mixing/interstage screening.

Table 15-28 Anwia Sulphide Master Viscosity Test Summary


ID RATE

(sec-1)

Viscosity

(cP)

Viscosity

(cP)

Viscosity

(cP)

Viscosity

(cP)

Viscosity

(cP)

Pre-CIL 4.1 n/a n/a n/a n/a n/a

Gravity 7.4 n/a n/a n/a n/a n/a


21.8 n/a n/a n/a n/a 131

38.9 n/a n/a 48 68 138

67.4 38 39 53 65 146

119.2 45 52 66 85 153

209.4 63 70 87 122 174

CIL 4.1 n/a n/a n/a n/a 696


13.1 n/a n/a n/a n/a 306

21.8 n/a n/a n/a 78 216

38.9 n/a n/a 53 80 150

67.4 35 41 54 80 137

119.2 45 46 66 86 140

209.4 61 69 95 105 166

15.6.5 Salman Oxide Master

Results from viscosity testing of Salman oxide master slurries at varying pulp density are summarised

in Table 15-29.

Salman oxide ore displays high viscosity at high pulp density and exhibits shear thinning, with the

viscosity decreasing with increasing shear rates. The effect of varying pulp density and shear rate for

Salman oxide ore is shown in Figure 15-16.

Salman oxide ore had the highest viscosity of all ores tested and the pulp density for leaching will be

limited to less than 45% w/w when processing this ore without the inclusion of other ores in the feed

blend.




121

Table 15-29 Salman Oxide Viscosity Test Summary


ID RATE

(sec-1)

Viscosity

(cP)

Viscosity

(cP)

Viscosity

(cP)

Viscosity

(cP)

Viscosity

(cP)

Pre-CIL 4.1 608 1055 2074 5405 17870

No Gravity 7.4 352 607 1233 3204 11420

13.1 222 407 788 2089 6965

21.8 144 253 498 1321 4379

38.9 95 167 322 831 2601

67.4 73 106 202 520 1559

119.2 69 94 147 314 985

209.4 80 103 137 242 602

Pre-CIL 4.1 N/A 536 940 1514 2230

Gravity 7.4 N/A 335 573 702 1442

Tailings 13.1 N/A 205 343 464 958

21.8 83 133 220 305 631

38.9 59 85 138 192 39467.4 49 68 97 124 248

119.2 52 65 85 108 167

209.4 66 75 93 110 157

CIL 4.1 530 738 1713 2302 3663

Tailings 7.4 274 476 978 1397 2476

13.1 170 297 587 819 1599

21.8 112 191 367 561 1143

38.9 76 121 226 349 686

67.4 61 88 145 232 444

119.2 61 80 116 165 277

209.4 76 88 119 156 217

15.6.6 Salman Transi tion Master

Results from viscosity testing of Salman transition master slurries at varying pulp density are

summarised in Table 15-30.

Salman transition ore displays high viscosity at high pulp density and exhibits shear thinning, with the

viscosity decreasing with increasing shear rates. Viscosity also reduced following gravity separation

and leach processing. The pulp density for leaching of Salman transition ores will be limited to less

than 45% w/w when processing this ore without the inclusion of other ores in the feed blend.




122

Figure 15-16 Variation in Viscosity w ith Shear Rate and Pulp Density – Salman Oxide

Salman Oxide - Pre-CIL, No Gravi ty

10

100

1000

10000

100000

1 10 100 1000Shear Rate (sec-1)

V i s c o s

i t y ( c P )

40% SOLIDS (w /w )

45% SOLIDS (w /w )

50% SOLIDS (w /w )

55% SOLIDS (w /w )

60% SOLIDS (w /w )

Table 15-30 Salman Transition Viscosity Test Summary


ID RATE

(sec-1)

Viscosity

(cP)

Viscosity

(cP)

Viscosity

(cP)

Viscosity

(cP)

Viscosity

(cP)

Pre-CIL 4.2 N/A N/A 711 1728 2496

Gravity 7.4 N/A N/A 411 853 1960Tailings 13.1 N/A 144 267 552 1295

21.9 N/A 94 177 374 852

38.9 48 69 113 235 554

67.4 42 58 89 155 364

119.2 49 61 87 137 240

209.6 69 80 101 143 225

CIL 4.2 N/A N/A 518 721 2113

Tailings 7.4 N/A N/A 311 597 1481

13.1 N/A N/A 197 374 927

21.9 N/A N/A 130 259 647

38.9 N/A 56 88 170 451

67.4 40 49 78 125 303

119.2 47 55 78 119 232

209.6 67 80 96 130 226




123

15.7 Gravity Recovery

15.7.1 General

Gravity recovery tests were conducted prior to leach optimisation and variability testing using 3kg

samples. Bulk gravity recovery tests were also performed on slurry samples prior to carbon

adsorption and cyanide detoxification testing.

Results from the gravity testing are summarised in the following sections.

15.7.2 3kg Batch Gravity Tests

Gravity recovery tests were performed with 3kg feed samples to provide slurry for leach optimisation

testing. The samples were ground to the required size and treated with a 3” Knelson concentrator per

the following conditions:

• Feed rate ~300g "dry solids"/minute.

• Initial fluidisation water pressure: 17kPa.

• Concentrate Treatment: Amalgamation.

The Knelson tails and amalgamation residue were then combined for subsequent leach tests.

Average results from gravity testing of 3kg samples of Anwia ore are summarised in Table 15-31.

Anwia ores show a weak relationship between gravity recovery, head grade and grind size, with

gravity recovery increasing with decreasing grind size and increasing head grade.

Table 15-31 Average 3kg Batch Gravity Test Results – Anwia

Grind Oxide Transition Sulphide

P80

(µm)

Head Grade(g/t)

Gravity

Rec (%)

Head Grade(g/t)

Gravity

Rec (%)

Head Grade(g/t)

Gravity

Rec (%)

75 3.31 31.83 2.81 27.91 3.12 49.23

106 3.35 28.58 3.09 34.36 2.87 47.61

150 3.05 28.92 2.92 37.39 2.87 42.34

Results from gravity testing of 3kg samples of Anwia ore variability samples ground to P80 75µm are

summarised in Table 15-32, 15-33, and 15-34.




124

Table 15-32 3kg Batch Gravity Test Results – Anwia Oxide Variability Samples

Sample

ID

Drill

Hole

Head Grade(g/t)

Gravity

Rec (%)

Oxide Master Composite - 2.78 33.17

Variability Composite 24 AWRC074 2.34 19.36



Table 15-33 3kg Batch Gravity Test Results – Anwia Transit ion Variability Samples

Sample

ID

Drill

Hole

HeadGrade (g/t)

Gravity

Rec (%)

Transition Master Composite - 2.64 36.99



Variability Composite 34 AWRC078 1.37 51.54Variability Composite 35 AWRC083, 27m-36m 1.09 25.95

Variability Composite 36 AWRC083, 36m-45m 4.30 51.90


Table 15-34 3kg Batch Gravity Test Results – Anwia Sulph ide Variabili ty Samples

Sample

ID

Drill

Hole

HeadGrade (g/t)

Gravity

Rec (%)

Sulphide Master Composite - 2.64 56.42


Variability Composite 29 AWRC076 1.68 54.75Variability Composite 30 AWRC077 4.64 54.28


Results from gravity testing of 3kg samples of Salman ore are summarised in Table 15-35.

Results from gravity testing of 3kg samples of Salman ore variability samples ground to P80 75µm are

summarised in Table 15-36, 15-37 and 15-38.

There is a very weak relationship between gravity recoveries, head grade and grind size, with gravity

recovery increasing with decreasing grind size and increasing head grade. Gravity gold recovery from

transition ore was substantially lower than that obtained from the oxide ore.

Gravity recoveries for Salman ores are much lower than those obtained for the Anwia ores.




125

Table 15-35 Average 3kg Batch Gravity Test Results – Salman

Grind Oxide Transition

P80

(µm)

Head Grade(g/t)

Gravity

Rec (%)

Head Grade(g/t)

Gravity

Rec (%)

38 - - 3.23 9.38

53 - - 3.28 11.08

75 2.22 12.07 2.72 4.51

90 2.52 13.31 2.56 3.98

106 2.34 11.12 2.66 6.31

150 2.37 10.92 3.23 9.38

Table 15-36 3kg Batch Gravity Test Results – Salman Oxide Variability Samples

Comp.

No.

Rock

Type

Pit HeadGrade (g/t)

Gravity

Rec (%)

1 Greywacke Akanko Central 1.55 44.12

2 Greywacke Nugget Footwall 2.59 2.60

3 Greywacke Salman Central 3.06 9.83

4 Greywacke Salman North 2.44 8.11

5 Greywacke Salman South 1.96 50.54

6 Greywacke Salman SW 2.61 64.33

7 Greywacke Teberu Footwall 2.32 11.13

18 Granite Akanko Central 1.68 37.83

19 Granite North Hill 3.70 35.99

20 Granite Salman North 2.20 8.18

Table 15-37 3kg Batch Gravity Test Results – Salman Transiti on Variability Samples

Comp.

No.

Rock

Type

Pit/

Oxidation Zone

HeadGrade (g/t)

Gravity

Rec (%)

8 Greywacke Salman Central tr-lo 3.62 5.98

9 Greywacke Salman North tr-lo 2.92 7.36

10 Greywacke Salman South tr-lo 2.45 6.79

11 Greywacke Akanko Central tr-up 1.59 29.60

12 Greywacke Salman Central tr-up 3.37 8.67

13 Greywacke Salman North tr-up 2.30 6.52

14 Greywacke Salman South tr-up 1.57 38.92

15 Greywacke Nugget Footwall 2.32 15.33

16 Greywacke Salman SW 2.23 18.94

17 Greywacke Teberu Footwall tr 3.21 23.98

21 Granite Salman North tr-lo 2.19 55.51

22 Granite Akanko Central tr-up 2.73 45.69

23 Granite Salman North tr-up 1.54 18.38




126

Table 15-38 3kg Batch Gravity Test Results – Salman Sulphide Variabili ty Samples

SulphideComp. No.

Rock

Type

Pit/Dril l Hole HeadGrade (g/t)

Gravity

Rec (%)

1 Greywacke Nugget Footwall, SNRC561 2.19 12.63

2 Greywacke Salman Central, SNRC570 2.57 4.32

3 Greywacke Salman Central, SNRC575 4.47 9.71

4 Greywacke Salman North, SNRC537 6.52 8.43


6 Greywacke Salman North, SNRC541 44m-79m 1.83 10.90

7 Greywacke Salman North, SNRC541 79m-95m 7.36 3.39


9 Greywacke Salman South, SNRC574 1.45 70.60

10 Greywacke Salman South, SNRC577 3.88 11.72

11 Greywacke Salman SW, SNRC579 1.53 67.29

12 Greywacke Teberu Footwall, SNRC554 3.57 15.14

13 Greywacke Teberu Footwall, SNRC560 2.42 4.6814 Granite Akanko Central, AKRC103 1.89 49.25

15 Granite Salman North, SNRC541 2.12 9.36

16 Granite Salman North, SNRC542 2.41 69.14

15.7.3 Bulk Gravity Tests

Gravity recovery tests were performed with 32 to 45kg feed samples to provide slurry for carbon

adsorption and cyanide detoxification testing. The samples were ground to the required size and

treated with a 3” Knelson concentrator per the following conditions:

• Grind P80: 75 µm.

• Feed rate ~300g "dry solids"/minute.

• Initial fluidisation water pressure: 17kPa.

• Concentrate Treatment: Intensive cyanidation, with diagnostic leaching of intensive

cyanidation residue.

Results from the bulk gravity tests are summarised in Table 15-39 and Table 15-40.

Table 15-39 Anwia Bulk Gravity Test Results

Sample

ID

Test

No.s

Feed

Mass

(kg)

Head Grade

(g/t)

Concentrate

Leach Rec(%)

Gravity

Rec (%)

Anwia Oxide SN1625,1619-1621 32 2.82 99.48 39.68

Anwia Transition SN1624, 1616-1618 45 2.27 99.32 41.70

Anwia Sulphide SN1567, 1568-1570 45 2.59 98.91 36.15




127

Table 15-40 Salman Bulk Gravity Test Results

Sample

ID

Test

No.s

Feed

Mass (kg)

Head Grade(g/t)

Concentrate

Leach Rec(%)

Gravity

Rec (%)

Salman Oxide SN1614, 1608-1610 45 3.07 98.01 27.53Salman Transition SN1613, 1605-1607 36 2.97 69.93 12.02

High intensity cyanidation gave excellent gold recovery from the bulk gravity concentrates, except for

the Salman transition sample. Diagnostic leaching of the leach residue obtained following high

intensity cyanidation of gravity concentrates from the Salman transition ore showed 92.8% of the

remaining gold was locked in sulphides.

Gravity gold recovery from all Anwia ore samples was significant and the installation of a well

designed gravity gold recovery circuit, with intensive cyanidation of the gravity concentrates, can be

expected to achieve 25 to 30% gold recovery via gravity and minimise gold losses to tailings due to

incomplete leaching of coarse free gold.

Gravity gold recovery from the Salman ores will be significantly lower than that obtained from the

Anwia ores (5 to 20%), but removal of gravity gold will help to ensure that the loss of free gold to the

leach tailings is minimised.

15.8 Leaching

15.8.1 General

Leach optimisation tests were conducted for each leach master composite sample. Variability tests

were then performed using the optimum conditions determined by results from the leach optimisation

tests and preliminary grinding and leaching capital and operating costs.

The leach optimisation testwork involved investigating the response of the master composite samples

to variations in grind size, pH, cyanide concentration and other leach conditions.

15.8.2 Leach Optimisation Testing

Samples of the master leach composites were ground in 3kg batches to the desired P80 and

processed by 3” Knelson concentrator. Gravity gold was removed from the concentrates by mercury

amalgamation. The amalgamation residue was then combined for subsequent leach testing.

Results from the leach optimisation tests are summarised in Table 15-41 to Table 15-45 and

discussed in the following sections.



Adamus Resources LimitedSouthern Ashanti Gold Project, Ghana, Western Africa Technical Report

Table 15-41 Anwia Oxide Leach Optimisation Test Results

Test No. P80 NaCN pH Head Gravity Overall Recovery NaCN Lim

(µm) (%) (g/t) (%) 2h 4h 8h 24h 48h (kg/t) (kg

SN1450 106 0.025 10.5 3.22 28.75 72.27 79.36 87.27 95.37 95.95 1.03 1.0

SN1451 75 0.050 10.5 3.03 30.21 80.03 86.70 92.38 95.79 97.28 1.47 0.9

SN1452 150 0.050 10.5 3.02 27.71 77.32 83.54 90.75 93.81 94.72 1.50 1.0

SN1453 106 0.100 10 3.32 26.02 81.42 87.88 93.83 96.61 97.17 2.76 0.0

SN1454 75 0.025 10 3.54 25.93 74.64 87.28 94.44 97.09 97.63 1.23 0.9

SN1455 150 0.025 10 3.04 30.32 79.78 87.88 93.62 96.07 96.68 1.43 0.8

SN1456 106 0.050 10 3.06 33.07 87.06 93.69 97.00 97.59 98.21 2.04 0.6

SN1457 75 0.100 10 3.07 34.97 87.96 93.32 96.09 96.54 96.96 2.87 0.6

SN1458 150 0.100 10 3.21 32.35 84.04 89.56 92.99 95.25 97.27 2.95 0.5

SN1459 106 0.025 9.5 3.25 31.96 83.01 88.64 92.03 94.22 96.21 1.31 0.6

SN1460 75 0.050 9.5 3.59 36.19 85.47 89.51 92.68 93.20 93.71 1.62 0.3

SN1461 150 0.050 9.5 2.94 25.31 74.05 84.43 92.50 92.93 93.56 1.67 0.3

SN1462 106 0.100 9.5 3.26 18.98 73.77 85.76 89.98 91.30 91.88 2.65 0.3

SN1463 106 0.050 9.5 3.83 21.94 81.07 89.91 94.70 96.40 96.88 2.07 0.3

SN1464 106 0.050 10.5 3.51 27.17 83.64 91.37 94.05 94.58 95.11 1.53 0.9

SN1465 106 0.050 10.5 3.33 35.90 78.92 84.63 90.52 95.64 97.57 0.63 0.7

SN1466 106 0.050 10.5 3.03 34.00 82.10 89.42 93.93 96.30 97.67 0.67 1.0

SN1467 106 0.050 10.5 3.68 28.01 78.79 89.36 94.70 96.74 97.17 1.28 1.2

SN1468 106 0.050 10.5 3.45 27.81 74.67 83.23 89.60 93.38 97.43 1.08 1.1

SN1469 106 0.050 10.5 3.57 0.00 51.58 68.95 83.69 92.88 96.88 1.43 1.3

SN1470 106 0.050 10.5 3.57 0.00 51.23 70.35 85.05 93.44 97.43 1.42 1.2




Table 15-42 Anwia Transition Leach Optimisation Test Results


(µm) (%) (g/t) (%) 2h 4h 8h 24h 48h (kg/t) (kg

SN1499 106 0.025 10.5 2.66 36.47 70.24 85.47 89.55 91.43 93.53 0.62 0.6

SN1500 75 0.050 10.5 2.60 19.04 84.90 88.48 92.56 94.69 96.84 1.09 0.6

SN1501 150 0.050 10.5 2.85 33.90 81.75 85.97 89.32 91.04 93.01 1.25 0.5

SN1502 106 0.100 10 3.43 36.01 87.84 91.03 94.12 95.74 97.37 2.02 0.4

SN1503 75 0.025 10 2.74 17.20 63.12 79.82 88.83 92.65 94.68 0.67 0.3

SN1504 150 0.025 10 3.20 40.50 79.77 84.79 89.30 92.55 94.28 0.79 0.3

SN1505 106 0.050 10 2.98 36.86 85.62 89.36 92.87 94.74 96.60 1.59 0.2

SN1506 75 0.100 10 2.93 36.44 86.24 90.00 93.60 95.48 97.39 2.64 0.2

SN1507 150 0.100 10 2.76 40.88 82.62 86.60 90.42 92.43 94.44 2.61 0.1

SN1508 106 0.025 9.5 2.94 38.86 81.12 86.30 89.89 91.76 93.65 0.95 0.1

SN1509 75 0.050 9.5 2.98 38.95 86.55 90.24 93.76 95.61 97.48 1.43 0.1

SN1510 150 0.050 9.5 2.88 34.27 83.08 87.74 91.38 93.31 95.24 1.56 0.1

SN1511 106 0.100 9.5 3.24 28.52 82.17 87.59 90.86 92.57 94.28 2.40 0.1

SN1512 106 0.050 9.5 3.26 30.84 78.15 85.12 91.33 94.71 96.41 1.62 0.1

SN1513 106 0.050 10.5 3.27 38.82 84.48 89.86 93.07 94.76 96.46 1.45 0.4

SN1514 106 0.050 10.5 3.19 37.15 83.03 89.22 90.73 94.63 96.18 0.70 0.3

SN1515 106 0.050 10.5 3.15 30.99 78.71 85.60 86.94 92.70 92.73 0.64 0.3

SN1516 106 0.050 10.5 2.76 29.03 81.43 88.78 92.07 93.79 95.52 0.95 0.3

SN1517 106 0.050 10.5 3.00 35.54 85.94 89.97 92.59 93.96 95.33 0.95 0.3

SN1518 106 0.050 10.5 2.66 0.00 69.36 83.64 89.39 93.33 95.42 1.24 0.4

SN1519 106 0.050 10.5 2.97 0.00 73.43 83.92 90.97 94.47 96.34 1.27 0.4




Table 15-43 Anwia Sulphide Leach Optimisation Test Results


(µm) (%) (g/t) (%) 2h 4h 8h 24h 48h (kg/t) (kg

SN1471 106 0.025 10.5 3.61 41.17 71.73 77.94 85.42 90.55 95.85 0.88 0.9

SN1472 75 0.050 10.5 3.41 45.93 83.84 88.41 92.60 95.85 97.48 1.46 0.7

SN1473 150 0.050 10.5 3.08 42.11 78.27 85.21 90.03 93.62 95.39 1.40 0.6

SN1474 106 0.100 10 3.09 43.25 81.92 85.62 89.21 93.79 97.34 2.45 0.5

SN1475 75 0.025 10 3.00 46.83 83.38 88.13 91.79 95.48 97.34 1.18 0.4

SN1476 150 0.025 10 2.86 38.43 73.41 82.93 88.24 92.15 94.10 0.86 0.0

SN1477 106 0.050 10 2.92 41.85 77.16 84.92 88.72 92.57 96.58 1.47 0.3

SN1478 75 0.100 10 3.00 49.03 86.44 90.11 93.80 95.62 97.48 2.27 0.3

SN1479 150 0.100 10 2.73 44.77 82.31 86.29 90.35 92.35 94.39 2.50 0.2

SN1480 106 0.025 9.5 2.44 48.30 80.49 84.92 89.54 91.76 94.06 0.98 0.2

SN1481 75 0.050 9.5 3.09 55.11 85.24 88.83 92.46 94.26 96.09 1.55 0.1

SN1482 150 0.050 9.5 2.80 44.03 78.58 82.59 86.55 88.54 90.51 1.64 0.1

SN1483 106 0.100 9.5 2.79 51.50 82.90 86.80 90.77 92.79 94.80 2.60 0.1

SN1484 106 0.050 9.5 2.82 52.64 82.78 86.65 90.66 92.64 94.59 1.50 0.1

SN1485 106 0.050 10.5 2.93 46.49 83.64 87.41 91.20 93.06 94.96 1.27 0.5

SN1486 106 0.050 10.5 2.73 45.73 87.30 91.21 91.93 93.85 96.06 0.61 0.3

SN1487 106 0.050 10.5 2.79 49.94 87.42 91.16 92.72 94.56 96.13 0.62 0.3

SN1488 106 0.050 10.5 2.59 55.25 86.12 89.48 92.95 94.81 96.65 1.20 0.4

SN1489 106 0.050 10.5 2.93 53.49 87.73 90.43 93.24 94.66 96.07 0.99 0.4

SN1490 106 0.050 10.5 2.52 0.00 40.49 67.24 82.39 91.64 96.08 1.33 0.6

SN1491 106 0.050 10.5 2.82 0.00 38.53 66.43 83.78 91.94 95.85 1.35 0.7




Table 15-44 Salman Oxide Leach Optimisation Test Results


(µm) (%) (g/t) (%) 2h 4h 8h 24h 48h (kg/t) (kg

SN1230 106 0.025 10.5 2.12 12.85 82.77 88.98 92.10 92.54 92.97 0.82 1.4

SN1231 75 0.050 10.5 2.39 12.96 81.18 87.99 91.66 93.48 94.54 1.29 1.9

SN1232 150 0.050 10.5 2.48 8.28 73.90 82.71 88.22 91.38 93.69 1.36 1.7

SN1233 106 0.100 10.5 2.49 11.95 77.92 84.53 88.68 91.75 94.51 2.01 1.8

SN1234 75 0.025 10 2.13 13.04 82.27 89.56 92.68 93.12 93.55 0.91 1.8

SN1235 150 0.025 10 1.99 12.04 81.97 88.32 91.53 92.00 92.46 0.96 1.5

SN1236 106 0.050 10 2.31 15.19 85.65 90.70 93.85 94.25 94.65 1.22 1.4

SN1237 75 0.100 10 2.09 10.80 86.43 91.26 93.39 93.83 94.27 2.32 1.4

SN1238 150 0.100 10 2.49 15.18 82.15 88.08 89.73 91.43 91.81 2.19 1.3

SN1239 106 0.025 9.5 2.41 14.20 76.21 83.77 86.13 88.48 88.86 0.93 1.2

SN1240 75 0.050 9.5 2.29 11.48 85.68 88.89 90.80 91.20 91.60 1.62 1.0

SN1241 150 0.050 9.5 2.53 8.19 73.18 80.13 85.68 88.16 88.53 1.62 1.0

SN1242 106 0.100 9.5 2.36 7.76 79.35 85.92 88.21 88.61 89.00 2.39 0.8

SN1243 106 0.050 9.5 2.41 6.91 76.41 84.67 87.31 89.29 90.67 1.54 1.0

SN1244 106 0.050 10.5 2.37 4.78 78.36 84.00 87.67 91.42 92.56 1.54 1.3

SN1444 90 0.025 10.5 2.25 11.77 80.37 86.89 87.38 89.48 91.24 0.51 2.0

SN1445 90 0.025 10.5 2.74 18.46 86.48 91.92 91.59 92.58 92.84 0.45 2.1

SN1446 90 0.025 10.5 2.50 10.83 81.61 85.98 87.31 88.61 91.85 0.39 2.0

SN1447 90 0.025 10.5 2.60 12.17 81.61 85.98 87.31 88.61 91.85 0.40 2.0

SN1448 90 0.025 10.5 2.35 0.00 57.77 73.65 86.57 90.87 91.27 0.70 2.3

SN1449 90 0.025 10.5 2.44 0.00 55.66 73.08 85.12 88.43 90.67 0.67 2.3




Table 15-45 Salman Transition Leach Optimisation Test Results


(µm) (%) (g/t) (%) 2h 4h 8h 24h 48h (kg/t) (kg

SN1245 106 0.025 10.5 2.76 3.18 43.34 51.55 56.31 59.64 62.11 0.70 1.0

SN1246 75 0.050 10.5 2.58 4.98 55.02 58.95 59.95 60.31 60.66 1.34 1.3

SN1247 150 0.050 10.5 2.79 7.90 45.12 50.13 51.91 54.29 54.62 1.28 1.2

SN1248 106 0.100 10.5 2.35 6.21 46.31 50.64 52.47 53.54 53.93 2.00 1.1

SN1249 75 0.025 10 2.98 3.35 50.74 60.00 63.68 65.24 65.55 0.87 1.0

SN1250 150 0.025 10 2.60 6.86 37.40 53.75 58.69 61.45 62.45 0.94 0.8

SN1251 106 0.050 10 2.80 17.85 60.45 66.37 68.52 71.00 72.61 1.39 0.7

SN1252 75 0.100 10 2.54 3.47 56.44 58.80 60.79 62.81 63.83 2.05 0.7

SN1253 150 0.100 10 2.58 5.01 51.28 55.52 58.83 61.21 62.65 2.30 0.7

SN1254 106 0.025 9.5 2.68 4.28 40.29 54.40 59.01 61.85 63.10 1.03 0.6

SN1255 75 0.050 9.5 2.60 5.20 49.45 56.49 58.78 61.19 63.08 1.47 0.4

SN1256 150 0.050 9.5 2.61 4.17 47.91 55.58 58.67 60.72 62.58 1.41 0.5

SN1257 106 0.100 9.5 2.43 4.29 58.81 61.89 64.31 66.10 68.31 2.31 0.4

SN1258 106 0.050 9.5 2.53 3.49 57.05 62.07 65.06 67.62 69.41 1.58 0.4

SN1259 106 0.050 10.5 2.40 5.28 57.99 63.20 65.19 69.46 71.28 1.51 0.6

SN1434 75 0.100 10.5 2.78 10.15 59.35 60.46 61.94 63.44 64.12 2.13 1.1

SN1495 53 0.025 10.5 3.24 9.11 40.70 55.99 60.47 64.21 67.77 0.81 1.2

SN1496 53 0.025 10.5 3.33 13.05 44.20 57.60 64.51 67.82 69.49 0.88 1.3

SN1497 38 0.025 10.5 3.23 10.28 45.16 56.70 62.71 67.80 69.52 0.88 1.1

SN1498 38 0.025 10.5 3.22 8.48 43.60 56.89 61.47 65.20 68.64 0.88 1.1



Adamus Resources LimitedSouthern Ashanti Gold Project, Ghana, Western Africa October 2007Technical Report Page 133

133

Effect of Grind Size

Anwia oxide ore displays litt le variation in recovery when varying P80 between 75µm and 150µm, with

similar recoveries obtained at 75µm and 106µm (~95%), and decreasing slightly to ~94% when thegrind is coarsened to 150µm.

Anwia transition ore displays a slight variation in recovery when varying P80 between 75µm and

150µm, with recoveries increasing from ~94% at 150µm to ~96% at 75µm.

Anwia sulphide ore displays a pronounced variation in recovery when varying P80 between 75µm and

150µm, with recoveries increasing from ~93% at 150µm to ~97% at 75µm.

Salman oxide ore gave similar recoveries at 75µm and 106µm (~92.5%) decreasing ~91% when the

grind is coarsened to 150µm.

Salman transition ore shows little variation in recovery when varying P80 between 38µm and 150µm,with ~1g/t residues obtained at all grind sizes.

Effect of pH and Cyanide Concentration

The effect of varying pH and cyanide concentration are relatively unimportant from a design

perspective, since both variables can be easily changed in operation. However, an understanding of

the interaction of pH and cyanide concentration on recovery and reagent consumptions is necessary

for the estimation of operating costs and recovery for the selected operating conditions.

Comparison of tests conducted at 0.05% NaCN and pH 9.5 and pH 10.5 shows leaching at pH 10.5

gives higher recoveries for the majority of the ores and grind sizes tested.

Comparison of tests conducted at varying cyanide concentration shows that leaching at higher

cyanide concentration gave increased leaching rates. The effect of increased cyanide concentration

diminished with decreasing grind size and increasing leach time.

Cyanide consumption increased two to three times when increasing the cyanide concentration from

0.025% NaCN to 0.10% NaCN. The variation of cyanide consumption with cyanide concentration for

Salman and Anwia ores is shown in Figure 15-17 and Figure 15-18




134

Figure 15-17 Effect of Cyanide Concentration on Cyanide Consumption – Salman

0.0

0.5

1.0

1.5

2.0

2.5

3.0

0.00 0.02 0.04 0.06 0.08 0.10 0.12

Cyanide Concentration (% NaCN)

C y a n i d e C o n s u m p t i o n ( k g / t )

Salman OxideSalman Transition

Figure 15-18 Effect of Cyanide Concentration on Cyanide Consumption – Anwia

0.0

0.5

1.0

1.5

2.0

2.5

3.0

3.5

0.00 0.02 0.04 0.06 0.08 0.10 0.12

Cyanide Concentrati on (% NaCN)

C y a n i d e C o n s u m p t i o n ( k g / t )

Anwia Oxide

Anwia Transition

Anwia Sulphide




135

Effect of Gravity

Results from tests performed with and without gravity recovery were compared to assess the effect of

gravity gold recovery on overall recovery.

The effect of gravity gold recovery for Salman and Anwia ores is shown in Figure 15-19 to Figure

15-22.

The tests with Salman oxide ore with gravity recovery prior to leaching at 106µm provide 2% higher

recovery after 48 hours leaching, when compared with leaching without gravity recovery at 90µm.

Gravity recovery prior to leaching gave higher recovery for leach times less than ~38 hours for Anwia

oxide ore.

Gravity recovery prior to leaching of Anwia transition ore gave 0.6% higher recovery after 48 hours

leaching, when compared with leaching for the same time without gravity recovery.

Gravity recovery prior to leaching gave higher recovery for leach times less than ~42 hours for Anwia

sulphide ore.

Figure 15-19 Effect of Gravity Gold Recovery – Salman Oxide

Salman Oxide

50

60

70

80

90

100

0 10 20 30 40 50 60

Leach Time (h)

R e c o v e r y ( % )

No Gravity (90µm)

Gravity (106µm)




136

Figure 15-20 Effect of Gravity Gold Recovery – Anwia Oxide

Anw ia Oxide

40

50

60

70

80

90

100

0 10 20 30 40 50 60

Leach Time (h)

R e c o v e r y ( % )

No Gravity

Gravity

Figure 15-21 Effect of Gravity Gold Recovery – Anwia Transition

Anwia Tran sition

60

70

80

90

100

0 10 20 30 40 50 60

Leach Time (h)

R e c o v e r y ( % )

No Gravity

Gravity




137

Figure 15-22 Effect of Gravity Gold Recovery – Anwia Sulphide

Anwia Sulphide

30

40

50

60

70

80

90

100

0 10 20 30 40 50 60

Leach Time (h)

R e c o v e r y ( % ) No Gravity

Gravity

A preliminary financial analysis indicates that an improvement in recovered grade of 0.008g/t (30”

Concentrator) to 0.014g/t (48” Concentrator) is necessary to justify the installation of a gravity recovery

circuit. For a mill feed grade of 2.70 g/t this equates to an improvement of 0.3% to 0.5% recovery.

Inclusion of a gravity circuit when processing Salman oxide ore provides a payback period of 5 months(30” Concentrator) to 9 months (48” Concentrator) and the installation of a gravity circuit with a 40”

concentrator is recommended.

The main parameters used in the economic analysis are summarised in Table 15-46.

Table 15-46 Gravity Circuit Economic Analys is Parameters

Parameter Unit Value

Plant Throughput Mt/a 1.32

Plant Availability % 91.32

Interest rate % 6

Loan period years 3

Gold Price US$/oz 600

Capital Cost - 48” Concentrator US$M 1.00






138

Effect of Carbon Addition

Previous testwork conducted for the Scoping Study indicated that the Salman ore had a slight preg

robbing capacity and the majority of tests conducted during this programme were performed with

carbon present through out leaching. Some tests were carried out to without carbon present toconfirm the requirement for the presence of carbon during leaching.

Salman oxide and Anwia transition ore gave improved recovery with carbon present during leaching,

where as higher recovery was obtained for Anwia oxide and sulphide ore without carbon present

during leaching.

To provide suitable flexibility within the circuit it is recommended that provision be made for carbon

addition to all tanks, with recovery of loaded carbon possible from either of the first two tanks.

Effect of Pulp Density

The majority of tests conducted during the leach optimisation programme were performed at 40% w/w

slurry density. Tests were also carried out at 45% w/w and 50% w/w to determine if there were any

beneficial, or adverse, effects from increasing the pulp density.

The tests with Salman oxide ore at increased pulp density showed reduced reagent consumption with

increasing pulp density. Residue grades increased with increasing pulp density; however the feed

grade for the tests varied widely, with the test performed at 50% w/w having the highest head grade,

making any comparison of recoveries difficult.

The tests with Anwia oxide ore at increased pulp density showed reduced reagent consumption and

increased recovery with increasing pulp density.

The tests with Anwia transition ore at increased pulp density showed reduced reagent consumption

and a slight decrease in recovery with increasing pulp density.

The tests with Anwia sulphide ore at increased pulp density showed reduced reagent consumption

and increasing recoveries with increasing pulp density.

Leaching should be performed at the highest possible slurry density, within the constraints of cyclone

performance, thickening capacity and slurry viscosity.




139

Optimum Grind Size and Leach Residence Time Determination

A simple economic analysis was performed to establish the optimum grind size and leach residence

time for plant design. The economic analysis considered the following criteria, for P80 from 75µm to

150µm and up 48 hours leach residence time:

• Mill operating power (calculated by Orway Mineral Consultants).

• Mill liners and grinding media consumption (calculated by Orway Mineral consultants).

• Cyanide and lime consumption.

• Grinding and Leaching Capital.Costs (estimated to ±30% by Lycopodium).

• Gold Recovery.

The main parameters used in the economic analysis are summarised in Table 15-47.

Table 15-47 Economic Analysis Parameters

Parameter Unit Value

Plant Throughput Mt/a 1.32

Plant Availability % 91.32

Fixed Annual Operating Cost US$M/year 10.13

Power Cost US$/kWh 0.053

Cyanide US$/t 2035

Lime US$/t 246

SAG Mill Balls US$/t 1181

SAG Mill Liners US$/t 2700

Ball Mill Balls US$/t 1163

Ball Mill Liners US$/t 2700

Interest rate % 6Loan period years 3

Gold Price US$/oz 600

The effect of varying grind size and leach residence time for the two major ore types, Anwia sulphide

and Salman oxide, are shown in Figure 15-23 and Figure 15-24. (Note: Revenue only considers the

capital and operating costs associated with grinding and leaching.)

A 75µm P80 provides the maximum revenue for both the Anwia sulphide and Salman oxide ores at all

leach times considered. The optimum leach time was 24 hours for Salman oxide ore, compared with

36 hours for the Anwia sulphide ore.

Based on this analysis a 106µm P80 and 30 hour leach time were considered to provide the optimum

economic benefit.




140

Figure 15-23 Effect of Varying Grind Size and Leach Time – Anwia Sulphide

Sensitivity to Grind Size and Residence Time

- Anw ia Sulph ide ($600/oz & 5.3¢/kWh)

38.5

39.0

39.5

40.0

60 75 90 105 120 135 150 165

Grind P80 (µm)

R e v e n u e ( U S $ / t )

48 h

36 h

30 h

24 h

Figure 15-24 Effect of Varying Grind Size and Leach Time – Salman Oxide

Sensitivity to Residence Time and Grind Size

- Salman Oxide ($600/oz & 5.3¢/kWh)

36.5

37.0

37.5

38.0

38.5

39.0

60 75 90 105 120 135 150 165

Grind P80 (µm)

R e v e n u e ( U S $ / t )

48 h

36 h

30 h

24 h




141

15.8.3 Leach Variability Testing

Conditions for the variability tests were selected using results from the leach optimisation tests in

conjunction with preliminary grinding and leaching capital and operating costs to determine the

grind/leach circuit configuration and operating conditions providing the optimum financial benefit for

the respective ores.

Samples of the variability leach composite samples were ground in 3kg batches to the desired P80

and processed by 3” Knelson concentrator. Gravity gold was removed from the concentrates by

mercury amalgamation. The amalgamation residue was then combined for subsequent leach testing.

Results from the leach variability tests are summarised in Table 15-48 to Table 15-50 All leach

variability tests were performed using the same conditions, except the Salman oxide samples which

were leached at 45% w/w due to the higher viscosity exhibited by this ore:

•

P80 75 µm

• 0.025% NaCN

• pH 10.5.

• 50% w/w

15.8.4 Effect of Oxygen Addit ion

A series of leach tests were performed with samples of Anwia sulphide AMC composite to determine

the effect of varying oxygen/aeration conditions on leach recoveries. Results from these tests are


Tests SN1690 to SN1697 were carried out as CIL tests, with carbon addition after 5 hours of leaching.

The feed for each test batch in this series was treated in the Knelson concentrator separately. The

calculated recoveries from the tests display a significant decrease following the addition of activated

carbon. This most likely due to the sampling and assay errors associated with the carbon (carbon was

sampled by “dips”, rather than total replacement).

The uncertainty of the results obtained, in particular the intermediate recoveries (between 10 and 30

hours leaching), and the differences in feed grades for the tests the makes it difficult to determine the

effect of the varying aeration conditions with a high level of confidence. For leach residence times

less than 24 hours, it appears that oxygen may improve gold recovery by ~1% (~0.03g/t). There

appeared to be no benefit in more than 5 hours of oxygen addition.

A further series of tests was performed with a modified procedure to eliminate variability in feed grades

and reduce sampling/assay errors. The feed samples for this series of tests (SN1701 to SN1709)

were split from the tailings obtained by processing a single 24kg batch treated in the Knelson

concentrator. No carbon was added during the leach to remove carbon sampling and assaying errors.




Table 15-48 Anwia Leach Variability Test Results

Test No. Sample Head Gravity Overall Recovery

ID (g/t) (%) 2h 4h 8h 24h 30h

SN1623 Oxide Master Composite 1 and 2 Blend 2.44 34.52 73.23 87.97 91.42 91.52 94.87

SN1553 Oxide Master Composite 1 2.78 33.17 74.56 92.86 94.00 95.65 95.70

SN1554 Variability Composite 24 2.34 19.36 80.03 90.45 89.40 90.68 90.79



SN1622 Transition Master Composite 1 and 2 Blend 2.28 37.60 78.14 85.59 91.42 93.67 93.87

SN1557 Transition Composite 1 2.64 36.99 69.17 87.16 93.47 93.99 95.62







SN1548 Sulphide Master Composite 2.64 56.42 90.24 94.32 93.49 93.54 93.86








Table 15-49 Salman Leach Variability Test Results – Oxide and Transition

Test No. Comp Deposit Rock Ox. Head Gravity Overall Recovery

No. Type* Class (g/t) (%) 2h 4h 8h 24h 30h

SN1593 18 Akanko Central IGRA Ox 1.68 37.83 89.22 91.02 92.42 92.99 96.15

SN1594 19 North Hill IGRA Ox 3.70 35.99 92.66 92.95 93.16 94.11 95.00

SN1595 20 Salman North IGRA Ox 2.20 8.18 39.56 54.99 71.08 83.68 83.82

SN1576 1 Akanko Central SW Ox 1.55 44.12 70.32 78.29 94.38 94.87 99.60

SN1577 2 Nugget Footwall SW Ox 2.59 2.60 79.68 80.27 80.87 81.13 81.18

SN1578 3 Salman Central SW Ox 3.06 9.83 76.20 83.38 86.70 87.72 89.34

SN1579 4 Salman North SW Ox 2.44 8.11 80.65 82.56 82.79 83.54 83.65

SN1580 5 Salman South SW Ox 1.96 50.54 94.32 95.74 95.74 96.33 96.61

SN1581 6 Salman SW SW Ox 2.61 64.33 78.55 87.08 88.97 92.16 93.94

SN1582 7 Teberu Footwall SW Ox 2.32 11.13 74.43 74.13 74.77 75.89 76.75

SN1590 15 Nugget Footwall SW Tr 2.32 15.33 66.07 72.68 78.45 78.81 78.97

SN1591 16 Salman SW SW Tr 2.23 18.94 47.60 71.07 80.24 83.70 85.69

SN1592 17 Teberu Footwall SW Tr 3.21 23.98 64.61 68.22 69.23 69.81 70.13

SN1597 22 Akanko Central IGRA Up-Tr 2.73 45.69 89.12 94.41 94.81 95.68 96.21

SN1598 23 Salman North IGRA Up-Tr 1.54 18.38 37.64 38.07 58.41 68.59 69.24

SN1586 11 Akanko Central SW Up-Tr 1.59 29.60 81.77 90.02 92.71 94.28 96.48

SN1587 12 Salman Central SW Up-Tr 3.37 8.67 80.26 84.24 84.40 84.44 85.37

SN1588 13 Salman North SW Up-Tr 2.30 6.52 35.83 36.87 37.14 40.00 40.47

SN1589 14 Salman South SW Up-Tr 1.57 38.92 82.90 86.08 87.30 88.13 88.59

SN1596 21 Salman North IGRA Lo-Tr 2.19 55.51 80.32 85.73 91.79 92.16 94.27

SN1583 8 Salman Central SW Lo-Tr 3.62 5.98 31.10 32.82 36.26 36.29 36.35

SN1584 9 Salman North SW Lo-Tr 2.92 7.36 24.17 26.01 28.29 31.59 33.26

SN1585 10 Salman South SW Lo-Tr 2.45 6.79 69.51 72.35 73.44 73.46 73.93




Table 15-50 Salman Leach Variability Test Results – Sulphide

Test No. Comp Deposit Rock Ox. Head Gravity Overall Recovery

No. Type* Class (g/t) (%) 2h 4h 8h 24h 30h

SN1648 1 Nugget Footwall SW Su 2.19 12.63 27.92 33.10 33.78 34.54 34.75

SN1649 2 Salman Central SW Su 2.57 4.32 38.60 40.57 43.62 44.26 45.55

SN1650 3 Salman Central SW Su 4.47 9.71 31.13 40.10 41.80 43.63 44.13

SN1651 4 Salman North SW Su 6.52 8.43 14.79 15.54 15.99 16.30 16.99





SN1656 9 Salman South SW Su 1.45 70.60 92.01 92.11 93.63 94.12 94.28

SN1657 10 Salman South SW Su 3.88 11.72 28.57 28.87 29.57 30.29 30.35

SN1658 11 Salman SW SW Su 1.53 67.29 85.68 91.53 91.54 94.03 96.86

SN1659 12 Teberu Footwall SW Su 3.57 15.14 35.91 36.54 38.02 38.10 38.29

SN1660 13 Teberu Footwall SW Su 2.42 4.68 12.46 14.72 15.85 16.39 16.63

SN1661 14 Akanko Central IGRA Su 1.89 49.25 75.88 83.47 88.76 93.77 94.59

SN1662 15 Salman North IGRA Su 2.12 9.36 17.74 21.09 21.63 22.74 22.97

SN1663 16 Salman North IGRA Su 2.41 69.14 91.49 91.65 95.19 96.14 97.74




Table 15-51 Effect of Varying Oxygen/Aeration

Test No. Aeration Head Gravity Overall Recovery Final Residu

Condi tions (g/t) (%) 2h 5h 10h 24h 30h 36h (g/t)

SN1692 36 Hrs O2 1.27 52.96 91.06 96.83 90.00 94.09 95.98 92.43 0.10

SN1693 36 Hrs O2 1.39 51.74 84.77 88.28 85.55 87.44 91.13 92.91 0.10

SN1696 10 Hrs O2/26 Hrs Air 1.46 52.18 89.77 91.88 86.53 88.31 88.33 89.98 0.15

SN1697 10 Hrs O2/26 Hrs Air 1.43 45.98 83.51 88.22 84.70 84.71 86.44 89.80 0.15

SN1694 5 Hrs O2/31 Hrs Air 1.47 48.32 85.34 89.13 82.36 85.84 89.18 90.77 0.14

SN1695 5 Hrs O2/31 Hrs Air 1.44 50.75 83.15 89.17 83.43 86.93 90.37 90.38 0.14

SN1690 36 Hrs Air 1.49 47.23 82.23 88.23 80.89 82.64 85.90 92.16 0.12

SN1691 36 Hrs Air 1.67 61.68 85.32 89.83 88.08 88.03 90.96 92.37 0.13

SN1703 36 Hrs O2 2.03 32.68 75.96 84.89 87.06 89.73 90.47 92.19 0.14

SN1704 36 Hrs O2 2.01 33.01 76.23 85.25 87.92 90.13 91.86 93.10 0.12

SN1707 10 Hrs O2/26 Hrs Air 1.98 33.56 73.97 82.09 86.77 90.49 91.23 93.98 0.10

SN1708 10 Hrs O2/26 Hrs Air 2.02 32.84 76.33 84.82 86.99 90.16 90.90 92.63 0.13

SN1705 5 Hrs O2/31 Hrs Air 1.98 33.54 75.42 85.57 88.78 90.00 92.25 93.50 0.11

SN1706 5 Hrs O2/31 Hrs Air 1.99 33.38 73.07 83.65 86.83 90.03 91.77 93.01 0.12

SN1701 36 Hrs Air 2.00 33.20 64.18 75.58 80.14 86.25 87.93 89.13 0.20

SN1702 36 Hrs Air 2.02 32.86 63.52 74.32 79.81 85.86 86.55 88.72 0.21




147

Table 15-52 Summary of Oxygen Uptake Test Results

Sample/Oxidant Oxygen Demand (mg/L/min) Av

0h 1h 2h 3h 4h 5h 6h 24h dO2

Anwia Oxide (O2) 0.023 0.013 0.113 0.116 0.118 0.100 0.153 0.020 19.5

Anwia Transition (O2) 0.002 0.130 0.130 0.136 0.147 0.145 0.117 0.165 17.5

Anwia Sulphide (air) 0.015 0.008 0.013 0.030 0.024 0.021 0.020 0.007 7.9

Anwia Sulphide (O2) 0.030 0.108 0.079 0.108 0.084 0.101 0.084 0.100 32.3

Salman Oxide (O2) -0.030 0.114 0.093 0.063 0.079 0.077 0.035 0.038 20.3

Salman Transition (O2) 0.091 0.108 0.100 0.087 0.061 0.064 0.095 0.074 23.1

The Anwia transition ore sample has the highest measured oxygen demand, as illustrated in Figure

15-26. The oxygen demand for this ore also increased significantly between 6 and 24 hours. This

may be due to the oxidation of sulphide minerals; however the observed behaviour doesn’t match the

typical response for ores containing reactive sulphides, which often display very high oxygen

consumption at the commencement of the test followed by a gradual decline in demand. The results

may have been affected by the loss of oxygen due to poor sealing of the test vessel lid.

Figure 15-26 Oxygen Demand - Anwia and Salman Ores

Oxygen Demand

-0.05

0.00

0.05

0.10

0.15

0.20

0 4 8 12 16 20 24

Time (h)

O 2

D e m a n d ( m g / L /

m i n )

Anwia Oxide Anwia Transi tion

Anwia Sulphide (air) Anwia Sulphide (O2)

Salman Oxide Salman Transition

A 1.5 t/day oxygen plant would be suitable to satisfy the calculated oxygen demand for the first leach

tank, with oxygen levels in the remaining tanks maintained by sparging up to 225Nm³/h of air.




148

15.10 Carbon Adsorption

15.10.1 General

Equilibrium loading and kinetic tests were conducted with slurry samples produced from bulk gravityrecovery and leach testing of the leach master composite samples. Results from these tests are


Table 15-53 Summary of Carbon Adsorption Test Results

Sample Equilibrium Carbon Loading (g/t) Kinetic Constants

0.2 ppm Au 0.5 ppm Au 1 ppm Au k (h-1) n

Anwia Oxide 3947 5806 7774 142 0.73

Anwia Transition 2479 3559 4679 192 0.67

Anwia Sulphide 2100 3149 4277 200 0.64

Salman Oxide 3721 4902 6037 160 0.77

Salman Transition 3357 4945 6630 202 0.67

The measured equilibrium loading and kinetic constants fall within the typical “normal” range of values.

Analysis of loaded carbon from these tests showed low to moderate levels (100 to 500 ppm) of copper

and nickel. The presence of copper and nickel at these levels is not expected to have any adverse

affect on gold recovery.

15.11 Cyanide Detoxification and Arsenic Precipitation

15.11.1 General

Cyanide detoxification (SO2/air) and arsenic precipitation (ferric salt addition) tests were conducted

with tailings slurry samples produced from bulk gravity recovery, leach and carbon adsorption testing

of the leach master composite samples. Results from these tests are summarised in Table 15-54 to

Table 15-57.

15.11.2 Cyanide Detoxi fication

All samples tested responded well to cyanide detoxification using the SO2/air process, with WAD

cyanide levels less than 0.5ppm (by picric acid assay method) achieved following detoxification.

The total cyanide concentration for some samples exceeded 1ppm due to the presence of residualferricyanide following detoxification. If there is a requirement to achieve total cyanide levels less than

1ppm then increased copper addition will be required to precipitate ferricyanide as Cu2Fe(CN)6.




149

15.11.3 Arsenic Precipitation

Arsenic precipitation tests were conducted for slurries that contained more than 1ppm arsenic

following cyanide detoxification, per the World Bank Environment, Health and Safety Guidelines for

“effluent discharged to receiving waters from tailings impoundments, mine drainage, sedimentation

basins, sewage systems, and stormwater drainage”.

The conditions used for the initial arsenic precipitation tests were:

• Retention time: 15 minutes

• pH: 8.5

• Target Fe:As mol Ratio: 4:1

With the exception of one test with Anwia Sulphide sample (D26) the arsenic concentration in the final

solution exceeded the 1ppm target after ferric precipitation. Additional tests were performed to

determine the effect of the following criteria:

• Increased Fe:As mol ratio.

• Air/Na 2S2O5 oxidation prior to ferric addition.

• Oxidation by Caro’s acid salt (KHSO 5) prior to ferric addition.

Sighter tests were also performed to assess the suitability of phosphate as an alternative precipitant

with prior oxidation by Na2S2O5 and KHSO 5. Results from these tests are summarised in Table 15-56.

Although the tests conducted using the Anwia sulphide sample with the increased Fe:As mol ratio

achieved arsenic levels less than 1ppm, the arsenic levels in tests with the Salman transition sample

increased significantly. The poor performance achieved with the Salman transition sample may have

been due to the higher pH at which these tests were conducted. The tests were conducted at pH 8.6

to 8.8, compared with the Anwia sulphide tests which were conducted at pH 8.0 to 8.1.

The arsenic levels increased significantly in all tests conducted with phosphate addition and no further

tests were conducted with phosphate.

Following review of the results from the preliminary and sighter arsenic precipitation tests a series of

optimisation tests were carried out to determine the effect of increased residence time, varying ferric

addition and reduced pH. Conditions used for the optimisation tests were:

• Retention time: 60 minutes

• pH: 7 and 8

• Target Fe:As mol Ratio: 3.7, 7 and 10

Results from these tests are summarised in Table 15-57. Figure 15-27 shows the arsenic precipitation

optimisations results and the response surface for a fitted equation.




Table 15-54 Summary of Cyanide Detoxif ication Test Results - Anwia

Sample ID SO2 CuSO4.5H2O Lime

Cu Ni Fe Zn As CNfree CNWAD CNtotal Addition Addition Addition (ppm) (ppm) (ppm) (ppm) (ppm) (ppm) (ppm) (ppm) (g SO2/g CNWAD) (ppm) (g Ca(OH)2/g S

Oxide Feed 2.40 <0.05 0.35 0.28 0.25 52 55 56 - - -

D17 0.07 0.04 0.18 0.15 <0.1 - 0.17 0.39 5.03 77 0.19

D18 0.21 0.05 0.19 2.59 0.13 - 0.18 0.57 3.77 74 0

D19 0.07 0.05 0.34 0.09 0.20 - <0.1 0.94 2.52 74 0

D24 0.20 0.34 0.58 0.05 0.15 - 0.27 1.87 2.39 74 0.10

D25 0.25 0.29 1.48 0.05 0.20 - 0.35 4.43 1.98 33 0

Transition Feed 13.20 0.07 1.32 0.51 9.00 49 65 69 - - -

D14 0.17 0.03 0.76 0.03 9.25 - 0.29 2.28 5.18 41 0.99 D15 0.09 0.03 1.04 <0.02 9.99 - 0.29 3.05 3.77 38 0.38

D16 0.22 0.06 1.23 6.75 7.88 - 0.29 2.44 2.51 41 0

D22 2.15 0.26 0.97 0 2.16 - 2.92 5.59 2.0 83 0.38

D23 0.19 0.17 2.40 0 3.78 - 0.48 7.09 3.0 40 0

Sulphide Feed 4.20 0.08 2.52 0.40 16.60 52 57 64 - - -

D1 0.21 0.09 1.06 0.04 7.85 - 0.38 3.30 7.71 0 0.83

D2 0.08 0.06 1.42 <0.02 6.93 - 0.34 4.26 6.42 0 0.71

D3 0.10 <0.05 0.54 <0.02 6.17 - 0.13 1.61 6.42 39 0.40

D4 0.10 <0.02 1.26 <0.02 6.66 - 0.53 4.00 5.18 82 0.64

D5 0.15 0.05 0.09 <0.02 2.97 - 0.30 0.53 6.00 79 0.41

D26 1.22 1.08 0.36 <0.02 0.77 - 0.28 1.27 2.55 84 0.41




Table 15-55 Summary of Cyanide Detoxi fication Test Results – Salman

Sampl e ID SO2 CuSO4.5H2O Lime

Cu Ni Fe Zn As CNfree CNWAD CNtotal Addition Addition Addition(ppm) (ppm ) (ppm) (ppm) (ppm) (ppm) (ppm) (ppm) (g SO2/g CNWAD) (ppm) (g Ca(OH)2/g

Oxide Feed <0.02 <0.05 0.15 0.16 0.40 53 53 54 - - -

D11 0.21 0.02 0.15 0.05 0.19 - 0.46 0.88 5.03 86 0.19

D12 0.17 <0.02 0.11 0.03 0.20 - 0.29 0.59 4.02 86 0

D13 0.12 <0.02 0.06 0.02 0.24 - <0.1 0.17 3.11 88 0

D20 1.71 <0.05 0.11 <0.02 0.15 - 1.86 2.16 3.06 82 0.32

D21 0.06 <0.05 0.20 <0.02 0.24 - 0.46 1.02 3.59 85 0.62

Transition Feed 8.82 <0.05 0.32 0.45 34.00 50 61 62 - - -

D6 0.16 <0.02 0.28 0.04 28.83 - 0.19 0.96 5.99 44 0.63 D7 0.15 <0.02 0.32 0.05 35.27 - 0.34 1.23 4.53 46 0.49

D8 0.31 0.04 0.27 0.03 38.83 - 0.28 1.03 3.02 44 0.79

D9 0.33 0.11 0.22 0.09 3.68 - 0.58 1.01 3.06 46 0.13

D10 0.44 <0.02 0.31 <0.02 4.96 - 0.97 1.81 2.51 46 0



Adamus Resources Limited August 2008Southern Ashanti Gold Project Feasibility Study – Open Pit Mining

Page 152

Table 15-56 Arsenic Precipitation Sighter Tests

Sample Test Na2S2O5 KHSO5 P:As

mol

Ratio

Fe:As

Molar

Ratio

pH As

(ppm)

Anwia Sulphide Feed - - - - - 1.50

MH7030 - - - 7 8.08 0.70

MH7031 - - 7 7.98 0.53

MH7032 - - 7 7.97 0.47

MH7033 - 7 - 9.05 2.90

MH7034 - 7 - 9.06 3.00

Salman Transition Feed - - - - - 3.68

MH7035 - - - 7 8.78 10.30

MH7036 - - 7 8.59 9.52

MH7037 - - 7 8.57 8.29

MH7038 - 7 - 9.06 22.10

MH7039 - 7 - 9.05 19.30

Table 15-57 Summary of Arsenic Precipitation Optimisation Test Results

Sample ID Precip. Acid Fe:As

pH Addition Molar As

(kg H2SO4/t) Ratio (ppm)

Salman Feed - - - 3.87

Transition D27 8.01 1.17 3.39 1.71

JK Comp D28 8.01 0.84 4.12 1.39

D29 8.02 0.67 10.16 1.13

D30 7.28 3.76 2.94 0.78

D31 6.99 5.37 7.72 0.47

D32 7.04 4.24 11.32 0.54

Increasing the Fe:As mol ratio increased arsenic precipitation, however reducing the pH gave a much

greater reduction in arsenic levels.

Additional precipitation tests are recommended for the Anwia transition and sulphide ores to confirm

the optimum pH, Fe:As mol ratio and corresponding acid consumption.




Page 153

Figure 15-27 Final Arsenic Concentration Variation with pH and Fe:As Molar Ratio

Recommended conditions for cyanide detoxification and arsenic precipitation are summarised in

Table 15-58.

Table 15-58 Recommended Cyanide Detoxi fication and Arsenic Precip itation Parameters

Sample g SO2/g CNWAD g Na2S2O5/g CNWAD ppm CuSO4.5H2O Precip.pH

Fe:As

Ratio

Anwia Oxide 2.5 3.7 75 7.5 5

Anwia Transition 3.0 4.5 80 7.5 5

Anwia Sulphide 2.5 3.7 80 7.5 5

Salman Oxide 3.6 5.3 90 7.5 5

Salman Transition 3.0 4.5 50 7.5 5




Page 154

15.12 Flowsheet Selection

Testwork has been completed to a level that is sufficient to provide process plant design criteriarequired for the detailed design. General conclusions drawn from the testwork performed on the

samples examined are:

• The preferred Base Case milling circuit comprises a SAG-Ball mill comprising a 1500kW SAG

mill and 2000kW ball mill. The selected circuit will generate a product having a P80 of 75µm

when treating 162.5t/h of the hardest ore type ((Anwia sulphide).

• The ores show a wide range of variability in comminution characteristics which could make

the control of a single stage SAG mill difficult. Single stage SAG milling is not recommended

where a grind size less than 80% passing 106µm is required.

• Irrespective of the selected milling circuit configuration, blending of the mill feed will be

necessary to minimise fluctuations in throughput and grind size and appropriateinstrumentation for SAG mill load control and measurement will be necessary to ensure stable

operation of the milling circuit.

• Inclusion of a gravity circuit when processing Salman oxide ore provides a payback period of

less than 9 months and the installation of a gravity circuit with 40” concentrator with intensive

leaching of gravity concentrates is recommended.

• Some Salman ores exhibit low-level preg-robbing and CIL processing is recommended when

processing Salman ores.

• A leach residence time of 30 hours is recommended.

• Oxygen addition over the first 5 hours of leaching is beneficial and a 1.5t/d PSA oxygen plant

is considered sufficient to meet this demand.

• The measured carbon adsorption parameters are within the normal range of values and

carbon adsorption can be expected to provide satisfactory adsorption recovery of gold.

• Thickening of the leach feed reduces leach reagent consumptions and the capital cost of the

leach circuit. Leach feed thickening is recommended

• All ore samples tested responded well to cyanide detoxification using the SO 2/air process with

moderate reagent additions able to produce slurries meeting the required discharge levels for

cyanide compounds.

• Ferric precipitation is the recommended process for arsenic stabilisation, with 1 hour

residence time, pH 7.5 and a 5:1 Fe:As molar ratio required to obtain arsenic levels less than

1 ppm.

•

Additional ferric precipitation tests are required for the Anwia transition and sulphide ores toconfirm the operating conditions for these ores.




Page 155

15.13 Recovery Forecasts

Data from the leach optimisation tests conducted at 0.025% - 0.50% NaCN and pH 9.5 – 10.5 wasanalysed using regression analysis to obtain expressions for the prediction of recovery with varying

grind size and leach times.

A typical response curve for the fitted data is shown in Figure 15-28.

Figure 15-28 Recovery Variation with P80 and Leach Time

Tailings grades were calculated using for each ore type over the project life, based on the following

criteria:

• The average P80 was calculated by Orway Mineral Consultants using the average feed blend

for each year from the proposed mine schedule.

• The leach residence times were calculated using the weighted average of the proportion of

each ore in the feed blend and operating leach densities for each ore type advised byLycopodium for a throughput of 1.32 Mtpa and a total leach capacity of 8,100m³.

• Where the forecast P80 was coarser than 75 µm, the predicted recoveries for Salman oxide

and transition variability samples were adjusted on a pro-rata basis using the grind-leach time

relationships determined from the leach optimisation testing of the associated leach

composite sample.

• Recoveries for Anwia oxide, transition and sulphide ores were forecast from regression

analysis expressions derived from associated the leach optimisation tests.




Page 157

Table 15-60 Predicted Tailings Grades

Ore Source, Oxidation, Rock type Year

1 2 3 4 5 6

Akanko Central, Oxide, Granite 0.08 0.08 0.08 0.07 0.07 0.08

North Hill, Oxide, Granite 0.21 0.21 0.21 0.21 0.20 0.20

Salman North, Oxide, Granite 0.36 0.37 0.38 0.37 0.36 0.36

Akanko Central, Oxide, Greywacke 0.03 0.02 0.02 0.01 0.02 0.02

Nugget Footwall, Oxide, Greywacke 0.50 0.50 0.51 0.51 0.50 0.50

Salman Central, Oxide, Greywacke 0.35 0.35 0.34 0.34 0.34 0.35

Salman North, Oxide, Greywacke 0.41 0.41 0.42 0.42 0.41 0.41

Salman South, Oxide, Greywacke 0.07 0.08 0.08 0.08 0.07 0.07

Salman SW, Oxide, Greywacke 0.18 0.17 0.18 0.17 0.17 0.17

Teberu Footwall, Oxide, Greywacke 0.55 0.55 0.56 0.55 0.55 0.55

Nugget Footwall, Transition, Greywacke 0.50 0.50 0.54 0.51 0.49 0.50

Salman SW, Transition, Greywacke 0.34 0.33 0.38 0.34 0.33 0.33

Teberu Footwall, Transition, Greywacke 0.97 0.97 1.04 0.99 0.97 0.97

Akanko Central, Upper Transition, Granite 0.12 0.11 0.19 0.13 0.11 0.12

Salman North, Upper Transition, Granite 0.48 0.48 0.50 0.48 0.48 0.48

Akanko Central, Upper Transition, Greywacke 0.07 0.06 0.10 0.07 0.06 0.07

Salman Central, Upper Transition, Greywacke 0.52 0.51 0.57 0.52 0.51 0.52

Salman North, Upper Transition, Greywacke 1.38 1.38 1.40 1.38 1.38 1.38

Salman South, Upper Transition, Greywacke 0.19 0.19 0.22 0.19 0.18 0.19

Salman North, Lower Transition, Granite 0.40 0.40 0.40 0.40 0.40 0.40

Salman Central, Lower Transition, Greywacke 2.32 2.31 2.34 2.32 2.31 2.32

Salman North, Lower Transition, Greywacke 1.97 1.96 1.97 1.96 1.96 1.96

Salman South, Lower Transition, Greywacke 0.65 0.65 0.69 0.66 0.64 0.65

Nugget Footwall, Sulphide, Greywacke 1.44 1.44 1.43 1.43 1.44 1.44

Salman Central, Sulphide, Greywacke 1.97 1.96 1.94 1.95 1.96 1.96

Salman North, Sulphide, Greywacke 1.44 1.43 1.43 1.43 1.44 1.44

Salman South, Sulphide, Greywacke 1.40 1.40 1.39 1.40 1.40 1.40

Salman SW, Sulphide, Greywacke 0.06 0.05 0.04 0.05 0.06 0.06

Teberu Footwall, Sulphide, Greywacke 2.12 2.12 2.11 2.11 2.12 2.12

Akanko Central, Sulphide, Granite 0.11 0.11 0.10 0.10 0.11 0.11

Salman North, Sulphide, Granite 0.85 0.85 0.84 0.85 0.85 0.85

Anwia Oxide 0.07 0.09 0.08 0.06 0.06 -

Anwia Transition - 0.18 0.20 0.20 0.18 -

Anwia Sulphide - 0.23 0.21 0.21 0.18 -




Page 158

16.0 MINERAL PROCESSING

16.1 Engineering Design and Control Philosophy

The design of the treatment plant will reflect:

• A simple and robust process flowsheet based on the testwork completed by ARL.

• Sturdy, well proven equipment.

• A control philosophy for a plant with an appropriate level of automation and remote control

facilities, supplemented by sufficient alarming and diagnostics to facilitate troubleshooting.

The proposed flowsheet has been selected to suit the various orebodies associated with the Project.

The major characteristics of the plant design are:

• Inclusion of 2 stage milling to achieve 80% passing 75 micron in leach feeds for Anwia

Sulphide ores.

• Inclusion of a gravity circuit based on testwork results indicating high gravity gold recoveries

for Anwia ores.

• Inclusion of a cyanide detoxification circuit to meet International Cyanide Code standards.

• Inclusion of an arsenic precipitation stage due to elevated arsenic levels in Salman

transitional ores.

The general control philosophy is for a plant with minimal automation. The plant will be provided with

a crusher control panel and a central mill control room from which the status of the major electrical

equipment can be monitored, and from which some of the regulatory control loops can be monitored

and adjusted. The starting and stopping of most electrical drives will be performed at the stop/start

control stations located adjacent to each drive or in the case of major equipment, started locally or by

remote from the control room.

16.2 Plant Configuration Options

On completion of the Base Case plant design, a review of several options was completed for changes

to the processing plant configuration to reduce the initial capital cost. For each of the options, the

operating costs and also impact on gold recovery due to the plant changes were considered. The five

options reviewed are as follows:

Option 1: Deletion of the two stage grinding circuit and insertion of a 5.5m dia. x 6.0m EGL

SAG mill (based on Golden Pride mill size such that design time can be minimised), deletion of the

Intensive Cyanidation Reactor and insertion of a Gemini Table to treat gravity circuit concentrates,

deletion of leach feed thickening, deletion of one CIL tank, deletion of the PSA plant, treatment of tails

slurry via decant return dilution to meet <50ppm CN WAD target and Arsenic precipitation only on

supernatant solutions prior to discharge.


crusher to the SAG mill deletion of the two stage grinding circuit and insertion of a 5.5m dia x 7.32m




Page 159

EGL SAG mill ( correct size to meet required 1.3 Mtpa throughput), deletion of the Intensive

Cyanidation Reactor and insertion of a Gemini Table to treat gravity circuit concentrates, deletion of

leach feed thickening, deletion of one CIL tank, deletion of the PSA plant, treatment of tails slurry via

thickening and decant return dilution to meet <50ppm CN WAD target and Arsenic precipitation only

on supernatant solutions prior to discharge.


crusher to the SAG mill deletion of the two stage grinding circuit and insertion of a 5.5m DIA * 6.0m

EGL SAG mill, deletion of the Intensive Cyanidation Reactor and insertion of a Gemini Table to treat

gravity circuit concentrates, deletion of leach feed thickening, deletion of one CIL tank, deletion of the

PSA plant, treatment of tails slurry via thickening and decant return dilution to meet <50ppm CN WAD

target and Arsenic precipitation only on supernatant solutions prior to discharge and use of a single

column for both acid wash and elution cycles.

Option 4: Deletion of the Intensive Cyanidation Reactor and insertion of a Gemini Table to treat


Option 5: Deletion of the two stage grinding circuit and insertion of a 5.5m DIA * 6.0m EGL

SAG mill, deletion of the Intensive Cyanidation Reactor and insertion of a Gemini Table to treat gravity

circuit concentrates, and deletion of one CIL tank.

Following a review of all options, ARL concluded that Option 5 provided the best Project return in

terms of capital and operating costs and gold recovery. The process flowsheet for Option 5 plan

configuration can be seen in Figure 16-1.

The treatment plant flowsheet is thus based on single stage crushing, single stage SAG milling,

gravity recovery of free gold from a portion of cyclone feed, pre-leach thickening, a single stage of

leaching and a five stage CIL circuit. Gold will be recovered by a 5 tonne Zadra elution circuit with

electrowinning of the gold onto stainless steel cathodes. The electro-deposited gold will be removed

with high pressure water sprays and smelted to a final bullion product.

16.3 Run of Mine (ROM) Pad

ROM ore will be delivered by haul trucks and dumped on the ROM pad.

There will be limited stockpiles maintained at the various pit locations and provision has been made

for a ROM stockpile to allow blending to optimise plant performance. The ROM pad will be

constructed with material found near the plantsite or mine waste as required.




Page 160

Tail

Conc

Underflow

Overflow

ROM Ore Stockpile

Primary Crusher

Surge BinCrushed OreStockpile

SAG Mill

Cyclones

Gravity Concentrator

Cyanide Leach

Leach/Carbon Adsorption (CIL)

Carbon Stripping

Carbon Regeneration

Thickener

Cyanide Destruction

Arsenic Precipitation

Tailings StorageFacility

Gold Room

Gold Bullion

Figure 16-1 Summary Process Flowsheet Option 5




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16.4 Crushing

ROM ore will be loaded into the crusher ROM bin either by a CAT 988 or equivalent front end loader

(FEL). The loader for ROM pad reclaim will be provided by the mining contractor. The crusher will beoperated on a 24 hour basis.

The ROM bin will have a live capacity of 200 tonnes and ore will be extracted from the bin by a

primary apron feeder and fed directly into a 42” x 55” (1,070 mm x 1,400 mm) single toggle jaw

crusher. The crusher will produce a product with a P80 of 110 mm, at the rate of 220 wet tph to allow

for a 24 hour per day crusher operation. The jaw crusher will handle ROM ore with a maximum lump

size of 800 mm x 800 mm. It is likely that slabby material will be encountered and a static grizzly

screen will be installed above the ROM bin to ensure that the crusher feed chamber does not become

jammed with oversize material. Handling of oversize from the static grizzly will be via the mine

contractor’s mobile excavator using a rock breaker tool. Provision will be made for the installation of a

suitable pedestal at the jaw crusher level should operational experience indicate the requirement for a

rock breaker at this level.

Primary crushed material will discharge onto a 1,200 mm wide conveyor, CV-01, for feeding direct to

a 90 tonne surge bin. From the surge bin, the primary crushed material will be fed at a controlled rate

to a 1,200 mm wide conveyor belt, CV-02, via a 900 mm wide electro-mechanically driven, variable

speed apron feeder. Conveyor CV-02 will transfer the crushed ore directly to the SAG mill feed chute.

Excess ore, not withdrawn to feed the SAG mill, will overflow from the surge bin and discharge onto a

750 mm wide stockpile feed conveyor, CV-03. The stockpile will have an 8,000 tonne capacity equal

to 48 hours of SAG Mill feed and will provide a reserve stockpile of crushed ore to ensure crusher

maintenance can be scheduled without interruptions to the mill operations. Crushed ore will be

reclaimed from the stockpile by FEL and transferred to the surge bin for feeding to the SAG mill.

Quicklime will be added directly from a lime silo onto the mill feed conveyor CV-02 for CIL circuit pH

control.

16.5 Grinding and Classification

The grinding circuit will consist of a Semi Autogenous Grinding (SAG) mill in closed circuit with

hydrocyclones. The 5.50 m diameter x 6.0 m EGL SAG mill with a 3,000 kW drive will operate at up

to 15% volumetric ball loading. Variable speed control of the mill, accomplished through a liquid

resistance starter and heat exchanger system, will provide additional flexibility for processing of

various ore types, ranging from 65 to 80% critical speed.

The discharge from the mill will feed the mill discharge hopper and will be diluted with process water

to approximately 60% solids prior to classification. The mill discharge slurry will be pumped to a

cluster of 6 duty and 1 standby classifying hydrocyclones.




Page 162

Cyclone overflow at 40% w/w solids will gravitate to a trash screen ahead of the leach feed thickener.

A portion of the cyclone feed will also be fed from a spigot off take to the gravity circuit. The trash

screen will be a horizontal vibrating unit installed with 0.63 by 18 mm slotted aperture crossflow

polyurethane panels. Trash screen oversize will gravitate to the tails hopper and underflow will

gravitate to the leach feed thickener.

The cyclone underflow will report to the SAG mill feed chute.

The cyclone feed offtake to the gravity circuit will feed a single-deck degritting screen located above

and feeding a single 30" gravity concentrator. The concentrator tails stream will gravitate to the mill

discharge hopper.

The mill floor slab will be sloped to a sump where a vertical spindle pump will return spillage to the mill

discharge hopper.

Grinding balls will be delivered to site and will be unloaded into concrete storage bunkers. Balls will

be loaded into the SAG mill via the Surge Bin Secondary Feeder and conveyor CV-02.

16.6 Gravity Concentration

The gravity concentrates recovered from the gravity concentrator will report to the gravity concentrate

hopper in the goldroom. The concentrates will then be fed over a Gemini Table with table tails

pumped to the mill discharge hopper and table concentrates collected by hand and dried in the drying

oven prior to being smelted.

16.7 Leach Feed Thickening

Trash screen underflow will launder to a 15 m diameter high rate thickener. Thickener overflow will

gravitate to the process water tank. Thickener underflow will be pumped to the CIL feed distributor

with the discharge stream being directed to either the leach tank, CIL tank 1 or the first adsorption

tank, CIL tank 2.

16.8 Leach and Adsorption Circuit

The CIL circuit will consist of one leach and five adsorption tanks each with a live capacity of 1,350

m3. The tanks will be interconnected with launders and slurry will flow by gravity through the tank

train. Each tank will be fitted with a dual stage mechanical agitator to ensure uniform mixing.

The five adsorption tanks will each be fitted with a mechanically swept wedge wire screen to retain the

carbon. A travelling gantry hoist will facilitate the removal of the screens for maintenance and routine

cleaning and will allow maintenance of all tank top equipment including agitators. All tanks will be

fitted with bypass facilities to allow any tank to be removed from service for agitator or screen

maintenance.

Provision will be made for the addition of sodium cyanide solution to the CIL feed distributor from a

main header pipe fed from the cyanide recirculation pumps.




Page 163

Carbon will enter the circuit at CIL tank 6 and will be advanced counter-current to the slurry flow by

pumping slurry, with a recessed impeller pump, from CIL tank 6 to CIL tank 5. The carbon will be

retained by the intertank screens in tank CIL 5 and the slurry will flow by gravity back to tank CIL 6.

This counter-current process will be repeated until the carbon eventually reaches CIL tank 2, the first

adsorption tank. A recessed impeller pump will be used to transfer slurry and loaded carbon to the

loaded carbon recovery screen mounted above the carbon elution column in the stripping plant. Theloaded carbon, reporting as screen oversize, will gravitate to the elution column and the screen

undersize slurry will return to CIL tank 1.

The tanks will be constructed on concrete ring beams in a bunded area with a sloping concrete floor.

Any spillage from the circuit will report to one of two sumps located on the periphery of the bunded

area. The spillage will either be pumped back to the circuit via the trash screen or the CIL feed

distributor or to tailings, either directly to the tails pumps or via the carbon safety screen.

Discharge from the last tank (CIL tank 6) will gravitate to the tailings hopper via a vibrating carbon

safety screen designed to recover any carbon leaking from a holed screen in the last tank. The

carbon safety screen will be a horizontal vibrating unit installed with slotted aperture crossflow

polyurethane panels. Carbon recovered on the carbon safety screen will be manually returned to the

circuit via the carbon sizing screen.

Barren carbon returning to the adsorption circuit from the carbon regeneration kiln will report to the

carbon sizing screen above CIL tank 6.

16.9 Elution and Gold Room Operations

The following operations will be carried out in the elution and gold room areas:

•

Acid Washing of Carbon.

• Stripping of gold from loaded carbon using the Zadra method.

• Electrowinning of gold from pregnant solution.

• Smelting of electrowinning and gravity concentration products.

The stripping and gold room areas will operate 5 days per week, with the majority of loaded carbon

preparation and stripping occurring during day shift. The Zadra stripping circuit will be manually

operated and will contain a separate rubber lined mild steel acid wash column and a stainless steel

elution column.




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16.9.1 Acid Wash

Loaded carbon will be received into the 5 tonne capacity acid wash column. Transfer, fill and acid

wash operations will be controlled manually.

During acid washing the dilute solution of hydrochloric acid pumped into the bottom of the column will

remove contaminants, predominantly carbonates, from the carbon. This process improves the elution

efficiency and has the beneficial effect of reducing the risk of calcium magnesium slagging within the

carbon during the regeneration process.

A metered 0.6 bed volumes of dilute acid solution will be pumped into the acid wash column and after

the predetermined soaking period the loaded carbon will be rinsed with water. Water rinsing will

consist of pumping 4 bed volumes of raw water through the column in order to displace any residual

acid from the loaded carbon. Dilute acid and rinse water will be disposed of to the tailings hopper.

Acid washed loaded carbon will be hydraulically transferred from the acid wash column to the elutioncolumn.

16.9.2 Zadra Elution Circuit

Fresh strip solution will be prepared prior to stripping each new batch of carbon. Sodium hydroxide

and sodium cyanide will be pumped from the respective storage and mixing tanks into the strip

solution tank and mixed with raw water to the required concentrations of cyanide (0.2 w/v %) and

caustic soda (2.0 w/v %).

The strip solution will be pumped from the strip solution tank through the reclaim heat exchanger and

then the inline strip solution heater and injected into the base of the elution column at a temperature of

125 C.

The strip solution will be pumped through the carbon in the stripping column and then pass through

eluate filters and the reclaim heat exchanger before entering the flash vessel.

The pregnant strip solution will then gravitate from the flash vessel to the electrowinning cells.

16.9.3 Electrowinning and Gold Room

In the Zadra system the elution and electrowinning are integral and continuous operations. When the

stripping solution exits the electrowinning cells it will gravitate into the strip solution tank and then be

recirculated to the elution column. The strip solution will be recirculated continuously for a designatedtime of 10 hours or until the gold level of the strip solution exiting the elution column reaches a desired

level. At that stage the barren carbon level should be less than 70 g Au/t carbon. Once elution and

electrowinning are completed the strip solution will be directed to the CIL feed distributor, allowing any

residual gold to be recovered in the CIL circuit.




Page 165

The electrowinning cells will be of polypropylene construction with stainless steel lockable lids and

sloping floors and will be located within the security area of the gold room. Rectifiers, one per cell, will

be located in a non-secure room below the cells allowing maintenance access without going through

security. Rectifier remote ammeters and manual controls will be located external to the gold room

security area. The cells will be arranged in parallel.

The two electrowinning cells will be fitted with stainless steel anodes and stainless steel wool

cathodes. A direct current will be passed through the cells between the electrodes and the electrolytic

action will cause the gold in solution to plate out on the cathodes. The gold will be removed from the

stainless steel wool cathodes by high pressure water blasting and then filtered in a pressure fil ter prior

to drying.

An overhead electric chain hoist will be provided to assist with handling of cathodes as necessary.

16.9.4 Gold Barring

The filtered dried gold sludge recovered from the cathodes of the carbon elution/electrowinning circuitand the Gemini Table concentrates will then be direct smelted with fluxes in a diesel-fired furnace to

produce doré bars.

Fume extraction equipment will be provided to remove noxious and explosive gases from the

electrowinning cells and barring furnace.

16.9.5 Gold Room Security

The gold room design is based on full security surveillance by a security guard and a second level of

surveillance by remote control CCTV cameras with viewing facilities in the Process Manager and

Security Foreman offices. Clean and dirty change rooms are provided adjacent to the security office.Toilet and crib rooms will be provided within the secure area to minimise entries. Access to the gold

room will be via proximity card and turnstile.

16.9.6 Carbon Regeneration

After completion of the elution process, the barren carbon will be transferred from the elution column

to a dewatering screen prior to entering the feed hopper of the carbon regeneration kiln. In the kiln

feed hopper any residual and interstitial water will be drained from the carbon before it enters the kiln.

Kiln off-gases will also be used to dry the carbon prior to entering the kiln.

The carbon will be heated to 650 - 750°C and held at this temperature for 15 minutes to allowregeneration to occur. Regenerated carbon from the kiln will discharge to a single deck carbon sizing

screen located above CIL tank 6. The carbon sizing screen will be fitted with a 0.6 mm square

aperture woven wire screen cloth.

Carbon reporting to screen oversize will be returned to CIL tank 6, and the fine carbon reporting to

screen undersize will discharge directly to the tailings hopper.




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16.10 Cyanide Destruction and Tailings Disposal

Tailings will gravitate from the final adsorption tank (CIL tank 6) via the carbon safety screen to the

Cyanide Detoxification Tanks. Two stages of detoxification each having one hour’s residence time

followed by a third stage for arsenic precipitation shall be provided. Detoxification shall be via the

Air/SO2 method.

Post cyanide destruction tails will be pumped with variable speed pumps to the tailings storage facility

(TSF). The tailings facility will comprise a cross-valley storage located to the north of the plant site.

Tailings will be discharged into the facility using sub-aerial deposition methods, through a combination

of banks of conventional spigot type discharges at regular intervals, and also single point discharges

from the heads of valleys. This will generate and maintain a supernatant pond at the southern end of

the facility where a decant tower based return water pumping system will be located. The supernatant

water will be pumped back to the process water pond.

Due to the site having a positive water balance there will be the need to periodically discharge water

to the environment. Where the cyanide and arsenic levels do not meet Ghana discharge

requirements the excess water to be discharged shall be treated via a cyanide detoxification step

followed by arsenic precipitation to meet the required discharge levels. It is expected that this second

stage of detoxification and arsenic precipitation may not be required given the first stage prior to

discharge to the Dam will meet Ghana requirements. Detoxification shall be via the Air/SO2 method.

The TSF design will incorporate an underdrainage system to minimise seepage to the local

groundwater. The basin underdrainage collection system will take advantage of the natural drainage

pattern existing in the tailings storage area, and will consist of two drainage units:

• A main collector drain; and

• finger drains.

The collector drains would be placed along the spine of the major drainage systems. The collector

drains consist of lengths of 160 mm diameter perforated drainage pipe, surrounded by drainage sand,

running either side for the existing stream beds. The drainage pipe is installed in a 600 mm deep vee-

shaped ditch, which is backfilled with sand drainage material to within approximately 200 mm of the

stripped ground surface. The rest of the excavation is backfilled to just above ground level with spoil

from the excavation.

The remaining areas of the under-drainage system are covered by finger drains at approximately 25

m spacings. The finger drains comprise sand surrounded, 63 mm diameter perforated drainage pipe

installed in a 400 mm deep, grader-cut, vee-ditch. The spoil from the vee ditch is used to construct a

compacted earth, erosion protection bund immediately upstream of the drain.

The placement of the drainage pipelines is arranged to take advantage of the existing ground fall and

minimal reshaping is therefore required for the drainage system.

The main collector drains will terminate in a sump constructed adjacent to and upstream of the main

embankment. From the sump seepage collected would be pumped back to the facility for collection in

the supernatant pond.




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16.11 Reagents

16.11.1 Lime

Provision has been made for quicklime to be delivered in bulk 20 tonne road tankers. The roadtankers will be pneumatically unloaded directly to the 80 tonne silo. Quicklime will be metered via a

rotary valve directly onto the SAG mill feed conveyor CV-02 for circuit pH control. Quicklime required

for the detoxification circuit shall be loaded into a hopper via a screw feeder and transported to the

detoxification circuit.

16.11.2 Cyanide

Cyanide will be delivered in 1.0 tonne bulk bags. The bulk bags will be lifted by monorail hoist to an

enclosed bag breaker above a cyanide mixing tank. Cyanide will be mixed to a 20% w/v solution with

process water and then transferred to a cyanide solution storage tank.

Cyanide will be added to the leach feed distributor and the Intensive Cyanidation Reactor from a

single ring main system fed by duty and standby fixed speed, centrifugal distribution pumps. Cyanide

will be dosed into the strip solution tank during preparation and mixing of the strip solution via a

dedicated variable speed, positive displacement pump.

16.11.3 Caustic

Caustic will be delivered in 25 kg bags and will be manually added to the caustic mixing tank. Caustic

will be mixed to a 20% w/v solution with raw water. Caustic will be dosed into the strip solution tank

during preparation and mixing of the strip solution via a dedicated variable speed, positive

displacement pump.

16.11.4 Hydroch lori c Acid

Hydrochloric acid will be delivered in bulk in an isotainer and will be transferred into a storage tank as

required. The dilute acid will be delivered to the acid wash column by the dedicated centrifugal

magnetic drive pump.

16.11.5 Activated Carbon

Fresh activated carbon will be delivered in 600 kg bulk bags. The bulk bags will be lifted by the CIL

gantry crane to the chute above the carbon sizing screen. The carbon sizing screen will remove

carbon fines from the fresh material and feed new coarse carbon into CIL tank 7.




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16.11.6 Sodium Metabisulphite

Sodium Metabisulphite (SMBS) will be delivered in 1 t bulk bags. The bulk bags will be lifted by

monorail hoist to an enclosed bag breaker above a mixing tank. SMBS will be mixed to a 20% w/v

solution with raw water before being pumped to a storage tank. It will then be dosed to the cyanidedestruction circuit.

16.11.7 Copper Sulphate

Copper Sulphate will be delivered in 1 t bulk bags. The bulk bags will be lifted by monorail hoist to an

enclosed bag breaker above a mixing tank. Copper Sulphate will be mixed to a 15% w/v solution with

raw water before being pumped to a storage tank. It will then be dosed to the flotation plant and

cyanide destruction circuit.

16.11.8 SAG Mill Balls

The 125 mm SAG mill grinding media will be delivered in bulk and stored in the SAG ball bunker.

SAG balls will be loaded into the Surge Bin and deposited into the mill via conveyor CV-02 as

required.

16.11.9 Flocculant

Flocculant will be received in bulk bags and lifted by chain hoist to the bag breaker/surge hopper of a

package mixing plant. It will be blown into a cyclone mixer and mixed to a solution strength of 0.25%

w/v with raw water and then aged in the mixing tank. It will then be transferred to a storage tank from

which it will be distributed to the thickener feed well by metering pump. The solution will be diluted

with raw water to a solution strength of 0.025% w/v prior to the thickener.

16.11.10 Ferric Sulphate

Ferric Sulphate will be delivered in 1 t bulk bags. The bulk bags will be lifted by monorail hoist to an

enclosed bag breaker above a mixing tank. Ferric Sulphate will be mixed to a 15% w/v solution with

raw water before being pumped to a storage tank. It will then be dosed to the arsenic precipitation

tank.

16.11.11 Sulphur ic Acid

Sulphuric acid will be delivered in 1,000 litre bulki-boxes and will be delivered to the arsenic

precipitation tank by a dedicated dosing pump.




Page 169

16.12 Services and Water

16.12.1 Raw Water Suppl y

A raw water storage tank of 500 m3

nominal capacity will be in the reagents mixing/storage area. Thevolume includes 200 m3 reserved for fire fighting purposes. This will receive water from:

• Pit dewatering Bores.

Overflow from the raw tank will be diverted to the process water pond.

16.12.2 Process Water

A process water pond of 2,000 m3 capacity will receive process water from the thickener overflow,

decanted supernatant from the TSF and make-up water from the overflow of the raw water tank.

Duty and standby low pressure pumps will be provided for plant process water with an additional

dedicated duty and standby set of low pressure pumps for the gravity concentrators.

16.12.3 Potable Water

Potable water will be supplied from the Raw Water Tank. This water will be filtered through a sand

filter and chemically treated prior to storage in a 60 m3 potable water tank. Duty and standby

pressure system pumps will be connected to this tank and distribute potable water around the plant, to

the mine workshop, the safety showers in the plant, the laboratory, and the administration and plant

buildings.

16.12.4 Instrument Air

The plant instrument air system will be provided via take-off from the plant air system, two stage

coalescing air filter system, refrigerated air driers and a dedicated air receiver. The system will feed

the carbon regeneration kiln, the CV-01 and lime silo dust collectors and mill girth gear lubrication

systems.

16.12.5 Plant Air

The plant air system will be supplied with clean, dry air from high pressure screw compressorscomplete with an air filter system. A discharge manifold will then distribute the air around the plant

from the air receiver.

16.12.6 Low Pressure Air

Low pressure air for the leach circuit and cyanide destruction circuits shall be supplied from multi-

stage blowers complete with air filtration system. Discharge manifolds will distribute the air to the

usage points.




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16.12.7 Oxygen

Oxygen will be generated on site via a pressure swing absorption (PSA) plant of 1.5 tonne per day

capacity. The system will be fed from the plant/instrument air system and will consist of PSA airreceiver, PSA oxygen plant and an oxygen receiver.

Oxygen will be distributed to the leach tank and the first two adsorption tanks, CIL tanks 2 and 3, and

injected through the shaft of the agitators.

16.12.8 Diesel Fuel

Diesel fuel will be received into a 10 m3 storage facility. A diesel distribution pump will deliver fuel to

the smelting furnace in the gold room and the regeneration kiln in the CIL plant.

16.12.9 Emergency Power Supply

An emergency power generating set will be provided, sufficient to provide lighting and sufficient power

to drive agitator motors, thickener drives, mill lubrication systems, heaters etc. In the event of a power

outage all drives will have to be started manually following bringing on line of the emergency power

supply.




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17.0 MINERAL RESOURCE ESTIMATE

17.1 Data Preparation and Treatment

17.1.1 Anwia Modelling Domains

Although structural controls on gold mineralisation at Anwia are not well understood, drilling in the

central part of the deposit is sufficiently close spaced that the orientations of mineralised zones are

beyond dispute. The present interpretation favours mineralization occurring in stacked primary lode

structures and this is supported by exposures presently being exploited by small-scale miners.

However, there are significant drill intercepts, particularly at depth, that fall outside of interpreted

mineralised domains. In the previous modelling, mineralised lenses were delimited using an

approximate mineralisation indicator grade of 0.2g/t Au on cross-sections looking both N40W and

N50E. Digitised outlines were then joined to form three-dimensional wireframes. In the author’s

opinion, the resulting interpretations were too constrained and resulted in an overly complicatedarrangement of domain boundaries inappropriate to the modelling method employed.

An alternative set of domains used in the current study simply groups the resource data into to

primary mineralised domains and are shown in Figure 17-1 and described in Table 17-2.

Note that the grid coordinates shown in the figures refer to a rotated local grid; grid transformation

parameters are listed in Table 17-1

Table 17-1: Anwia grid transformation parameters

Origin point (UTM) Rotation

574630.83E 40 degrees clockwise

550671.00N

The current study has also re-interpreted the weathering and oxidation surfaces using the appropriate

logging code provided in the geological data set. These logs were used to interpret weathering

profiles on cross-sections (looking toward 320 degrees on UTM grid) that were then joined to form

surfaces describing the bases of very weathered and moderately weathered rock and the top of fresh

rock for the purpose of assigning bulk densities in resource estimation. Depth to fresh rock at Anwia

varies between 25 and 50 metres over the resource area. Figure 17-2 shows the interpreted geology,

weathering profiles and mineralisation of a representative cross-section.

Table 17-2: Anwia resource modelling domains

Domain

number

Description

0 Peripheral, essentially barren, undefined mineralisation

1 Main Zone – series of stacked primary lode structures

2 Granite hosted gold mineralisation




Page 172

Figure 17-1 Plan view of Anwia model domain wireframes

Figure 17-2 Interpreted geology, weathering and mineralisation, Anwia section 550350N




Page 173

17.1.2 Salman Modell ing Domains

A series of mineralised lenses was interpreted over the eight and a half kilometres established strike

length of gold mineralisation, delimiting mineralised zones of similar tenor and directional trends at a

cut-off grade of about 0.2g/t Au. The approach is essentially identical to that applied by SRK in the

February 2005 resource estimate (Kentwell et al, 2005) and the Adamus January 2007 estimate, andthe wireframes are very similar. Outlines were digitized on E-W cross-sections, with points snapped

to drill traces in three-dimensions and those outlines then joined to form three-dimensional

wireframes. May 2005 logging data was used to interpret weathering profiles on cross-sections that

were then joined to form surfaces describing the bases of very weathered and moderately weathered

rock and the top of fresh rock to assign bulk densities in resource estimation. The elevation of the

fresh rock interface is fairly consistent across the resource area but because recent erosion is cutting

into the previously developed weathering profile, topography results in depth to fresh rock varying

from 80 metres beneath hills to less than 10 metres in low-lying areas.

Mineralisation in near-surface remnant lateritic scree and soil was delimited by the base of very

weathered (vox) material, equivalent to the top of recognizable weathered bedrock or saprolite.

Figure 17-3 shows a plan view of mineralisation wireframes and Figure 17-4 to Figure 17-7 show

interpreted geology, weathering profiles and mineralisation on several cross-sections. Table 17-3 lists

domain names and extents.

Table 17-3: Salman-Akanko resource modelling domains

Domain

Number

Description Extents

0 Waste domain: all samples outside of mineralisation wireframes and

below vox surface

1 All surficial material above vox surface

2 Salman South mineralisation associated with main Salman Shear 550837.5N – 551412.5N3 Salman South: west-dipping mineralisation in footwall of main shear 551437.5N – 551925N

4 Northern part of Salman South mineralisation on main shear 551687.5N – 551937.5N

5 Western phyllite-hosted mineralisation, west of Salman South 551312.5N – 551487.5N

6 Salman Central main lens 551937.5N – 552662.5N

7 West-dipping footwall lens east of Salman Shear 551787.5N – 552087.5N

8 Nugget Hill 552762.5N – 552962.5N

9 West-dipping footwall lenses east of Nugget Hill 552987.5N – 553112.5N

10 Nugget North main shear 553137.5N – 553412.5N

11 Nugget North west-dipping footwall lens 1 553187.5N – 553362.5N

11 Nugget North west-dipping footwall lens 2 553400N – 553562.5N

13 Teberu footwall lens 553587.5N – 554037.5N

14 Salman North granite and phyllite-hosted mineralisation associated with

Salman Shear

554125N – 555105N

15 Salman North east-dipping mineralisation hosted by footwall greywacke 554125N – 554680N

16 North Hill granite-hosted mineralisation 555262.5N – 555512.5N

17 Akanko South granite-hosted mineralisation 556237.5N – 556462.5N

18 Akanko Central granite-hosted mineralisation associated with Salman

Shear

557062.5N – 557887.5N

19 East-dipping Akanko footwall reef 557137.5N – 557462.5N




Page 174

Figure 17-3: Plan view of the Salman-Akanko mineralisation wi reframes

North

Central

South Model




Page 175

Figure 17-4: Interpreted geology and mineralisation on Salman South section 551850N

Figure 17-5: Interpreted geology and mineralisation on Salman Central section 552200N




Page 176

Figure 17-6: Interpreted geology and mineralisation on Teberu section 553800N

Figure 17-7: Interpreted geology and mineralisation on Salman North skewed section 554300N




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17.1.3 Satellite Deposi t Modelling Domains

Mineralised zones of similar tenor and directional trends were delimited at a cut-off grade of

approximately 0.2g/t Au for each Satellite Deposit. The selection approach and the procedure used indeveloping the wire-frame models is similar to that used by Hellman and Schofield in their previous

SAGP resource estimates (Hellman and Schofield; 2008a). Table 17-4 lists the domains used for

each satellite deposit and their summary descriptions

Table 17-4: Satellite Deposi ts resource modelling domains

The main mineralised zone at Bokrobo dips steeply to the west and strikes in a northerly direction. An

east dipping barren dyke stopes out the main zone on some levels. The gold mineralisation at

Bokrobo has been defined over a strike of some 350 metres and up to approximately 200 metres

vertically. The bulk of the available sample information is located on approximately 50 metre spaced

sections between 548,050 and 548,400mE. Figure 17-8 shows the mineralisation and dyke

wireframes used in the MIK estimate.

Deposit Domain

Number

Description

Bokrobo 0

1 Main Mineralised Zone-Strikes N and dips steeply to W

2 Dyke-Barren

Aliva 0 Peripheral (Essentially Barren)

1 Stacked, moderately E-dipping Mineralised Zones

2 Central Area- Narrow Steeper E-dipping Mineralised Zones

3 North Area- Narrow Steeper E-dipping Mineralised Zones

Avrebo 0 Peripheral (Essentially Barren)

1 Main Mineralised Zone (Weakly Kinked NNW-SSW striking lode)

Nfutu 0 No geometry of mineralization able to be interpreted.




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Figure 17-8 Pseudo 3D view of Bokrobo model domain wireframes

The gold mineralisation at Aliva has been defined over a strike of some 2,000 metres and up to 90

metres vertically. The bulk of the available sampling from drill holes is located on approximately 50

metre spaced sections between 548,500 and 550,500mE. The trenches sampled are spacedirregularly along the deposit, at approximately 100m intervals.

The Aliva drill holes and trench coordinates were rotated before modeling. Table 17-5 details the

rotation used to transform UTM coordinates into the local modeling grid.

Table 17-5: Aliva grid transformation parameters

Figure 17-9 shows the mineralisation and dyke wireframes used in the MIK estimate at Aliva.

UTM Coordinate Equivalent Local Grid

Coordinate

East 574,630.83 574,000

North 550,671.00 550,700

Rotation -40




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Figure 17-9: Pseudo 3D view of Ali va model domain wireframes in rotated grid

The gold mineralisation at Avrebo has been defined over a strike of some 670 metres and up to

approximately 170 metres vertically. The bulk of the available sampling from drill holes is located on

approximately 50 and 20 metre spaced sections between 544,350E and 545,050mE. The trench

sampling is spaced irregularly along the deposit. Two orientations have been used during drilling.The 50 metre-spaced sections are drilled along roughly east-west sections while the 20 metre spaced

sections are oriented at 315º.

Figure 17-10 shows the mineralised envelope used in the MIK estimate at Avrebo. This domain was

interpreted from logged geology based on alteration.




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Figure 17-10: Pseudo 3D view of Avrebo model domain w ireframe

The gold mineralisation at Nfutu has been defined over a strike of some 400 metres and up to 90

metres vertically. The bulk of the available sampling from drill holes is located on approximately 50

metre spaced sections and 2 surface trenches are present in the resource data set. The Nfutu drill

holes and trench coordinates were rotated before modelling. Table 17-6 details the rotation used to

transform UTM coordinates into the local modelling grid.

Table 17-6: Nfutu grid transformation parameters

UTM Coordinate Equivalent Local GridCoordinate

East 577,985 577,985

North 551,565 551,565

Rotation 45

Figure 17-11 shows the drill holes, trenches used in the MIK estimate at Nfutu. The Nfutu data set

was insufficient to formalise an interpretation of the mineralised geometry, therefore all data was

grouped into a single modelling domain.




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Figure 17-11: Pseudo 3D view of Nfutu dr ill holes and trenches

17.1.4 Mine Voids

Areas known to be affected by voids from previous mining are the Akanko footwall lode mined in the

late 1800s and early 1900s and the shallow pits and shafts recently mined by small-scale miners at

Anwia. The impact of these voids on resource estimates is considered negligible and no attempt has

been made to deplete resources for material previously mined.

At the Bokrobo Satellite Deposit an allowance has been made for resources which have been

depleted by artisanal mining activities. ARL produced a wireframe solid that defined the area/volume

affected by mining and the resources estimated within the mined wireframe have been removed prior

to reporting the resources.

17.1.5 Derivation of Preferred Assay Values

At both Anwia and Salman, 50g fire assays have been used in all cases where they are available.

Where repeat assays are available the results have not been averaged; initial assays have been

used. Table 17-7 lists the origins of assays used to inform the Anwia resource estimate. Samples

from RC holes drilled by Samax/Ashanti Goldfields at Anwia were routinely assayed by cyanide leach.

Of the 9039 samples assayed by that method, 710 have fire assays available for the leach residues

and total gold could be calculated.




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The samples with residue assays indicate that cyanide leach recoveries were generally greater than

90 per cent and normally in the range 92-96 per cent. The use of cyanide leach assays for samples

for which total gold cannot be calculated is therefore not considered unduly conservative. One

hundred and thirty-six samples from several RC holes drilled by Adamus soon after acquisition of

Anwia were also assayed by cyanide leach. Fire assays of leach residues are available for all of

those samples and total gold was calculated and used to inform the resource estimate. A number ofpulp residues from core samples were shipped to Australia for check assays by Genalysis in Perth.

The Genalysis assays have been used in preference to previous assays.

Table 17-8 lists the sources of assays used to inform resource estimates at Salman. A number of

core and RC samples have been re-assayed using cyanide leach but in all cases initial fire assays

have been used. Pulp residues from 346 core samples and 3363 RC samples were shipped to

Genalysis Laboratories in Perth for check fire assaying. The Genalysis assays have been used in

preference to the original assays by SGS or Transworld in Ghana.

Table 17-7: Anwia drill holes preferred gold assay sources

Company Drill holes SampleType

MetresSampled

Laboratory Assay Method

Semafo TB0001-TB0075 core 12912 SGS Accra 50g fire assay

Semafo RC0001-RC0412 RC 22448 SGS Accra 50g fire assay

Samax/AGC RC0450-RC0599 RC 8104 SGS Tarkwa 2kg CN leach

Samax/AGC RC0450-RC0599 RC 710 SGS Tarkwa 2kg CN leach with 50gfire assay of residues

Samax/AGC RCD600-602 core 225 SGS Tarkwa 2kg CN leach

Adamus AWRCD027-032 core 243 Genalysis Perth 50g fire assay

Adamus AWDD006, 015, 016, AWRCD009, 010,012-014, 027-032

core 2086 SGS Tarkwa 50g fire assay

Adamus AWDD006, AWRCD005, 007, 008,033-038

core 1850 TransworldTarkwa

50g fire assay

Adamus AWRC017-019, 022-026

RC 136 SGS Tarkwa 1kg CN leach with 50gfire assay of residues

Adamus AWRCD005, 008-010,012, 013, 027, 031

RC 69 SGS Tarkwa 50g fire assay

Adamus AWRC001-004, AWRCD005, 007-010,012-014, 017-038

RC 2254 TransworldTarkwa

50g fire assay




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Table 17-8: Salman dril l holes preferred gold assay sources

Company Drill holes Sample Type MetresSampled

Laboratory Assay Method

BHP SDDH001-004 core 571 SGS Accra 50g fire assay

BHP SRCH001-096 RC 6929 SGS Accra 50g fire assay

Adamus SNDD088, 136, 236,SNRCD246-247, 390, 395

core 496 SGS Tarkwa 50g fire assay

Adamus SNDD088, 136,SNRCD087, 110-112, 137-138, 405, 407, 422, 430,434-435, 445

core 2107 TransworldTarkwa

50g fire assay

Adamus SNDD088, 136, 210, 213,217, 225, 236, SNRCD246

core 346 Genalysis Perth 50g fire assay

Adamus AKRC013-033, SNRC192-401, SNRCD395

RC 6673 SGS Tarkwa 50g fire assay

Adamus AKRC001-181, SNRC001-797, SNRCD087-445

RC 58972 TransworldTarkwa

50g fire assay

Adamus AKRC014-015, 017,SNRC148-282

RC 1287 Genalysis Perth 50g fire assay

17.1.6 Compositing

Prior to compositing, all intervals for which assays were not available were allocated gold grades of-999 g/t.

Of the 64,209.6 metres of drilling available at Anwia, 4,204 metres has been assayed in four metre

intervals, 1,288 metres in intervals between 1 and 4 metres in length, and the remainder in intervals of

one metre or less. The longer sample intervals are almost exclusively in barren or very low grade

material so weighted average grades were calculated for one metre sample composites. Composites

with negative gold grades, indicating they were affected by intervals for which assays are not

available, were discarded. Residual intervals less than 0.5m length were also discarded. The

resulting located sample composites total 29,855 samples. The few available trench samples at

Anwia were not used to inform the resource estimate.

A number of trenches in the Salman database have been excluded from data informing the resource

estimate because of uncertainty concerning sample locations or because they lie outside of the

resource area. Names of excluded trenches are: AKTROADCUT, EBT001, EY2T1, EY3T1, EYR01,

SET01 to SET11, SNT0E, SNT0FE, SNT0FW, SNT26TEM, SNT30TEM, SNT51WTEM, all TUT prefix

trenches, SNT04CTEM, SNT05ATEM, SNT34TEM, SNT41ETEM, SNTRTTEM, SNT07ATEM and

SNT09ITEM. In remaining trenches over the Salman resource area there are 8169 metres of

horizontal channel sampling available from 137 trenches and bulldozer cuts. About 150 metres of

trenching has been sampled in three metre intervals, 1870 metres in two metre intervals and the

remainder in one metre intervals. Considering that the longer sample intervals are generally in barren




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or very low grade material, and desiring to render the trench sample data compatible with drill hole

sampling, weighted average grades were calculated for 1.5 metre composites of channel samples.

Intervals with negative gold grades (1266) and residuals less than 0.75 metres length were discarded,

leaving 5225 located sample composites.

Gold assays are available for 1316 metres of vertical channel sampling in 357 manually dug pits fromthe 2006 sampling campaign. One thousand and sixty four channel samples are in intervals of one

metre or less; 255 samples are in 1-2 metre intervals. Pit channel samples were composited to

uniform one metre length and residuals less than 0.5 metres length were discarded, leaving 1327

sample composites to inform estimates.

The Salman-Akanko resource estimate utilises samples from drill holes up to AKRC263 and

SNRC826, available at 31 October 2007. Drill holes SNRC064, 065, 369, 370, 371 and 372 lie

outside of the resource area and were excluded from the resource drill data. Gold assays are

available for 85,824.8 metres of drilling in 1,141 holes, including 4740.7 metres of HQ and PQ

diamond core in 49 holes. About 34,340 metres of RC drilling in material grading less than 0.2g/t Au

has been sampled in four metre intervals, 507 metres in three metre intervals and 936 metres in two

metre intervals; all remaining RC drilling has assays available for one metre intervals. In diamond drill

core, 2,698.5 metres has been sampled and assayed in lengths of one metre or less; 1,242.2 metres

sampled in intervals longer than 1.5 metres is almost exclusively in waste material. Weighted

average grades were calculated for two metre down-hole composites and intervals with negative gold

grades (affected by intervals for which assays are not available) and residuals less than 2.0 metres

length were discarded. The remaining data set comprised 45,711 composites (inc. channel samples).

Similar assay data compositing was undertaken for the Satellite Deposits resource modelling. At

Bokrobo, assays were available for 68 drillholes totalling some 8300 metres. The final resource data

set comprised 3,637 composites. At Aliva, assays were available for 36 surface trenches (composited

to 1.5m intervals) and 88 drill holes totalling 4778m. The final resource data set comprised 3,036

composites. At Avrebo, assays were available for 10 surface trenches totalling 1,139m (composited to

1.5m intervals) and 96 drill holes totalling 8579m. The final resource data set comprised 4,417

composites. At Nfutu, assays were available for 2 surface trenches totalling 1,139m (composited to

1.5m intervals) and 63 drill holes totalling 5,055m. The final resource data set comprised 2,505

composites.

The volume differences between one metre RC samples, one metre diamond core samples and 1.5

metre trench channel samples are considered insignificant and the compositing down of 4 metre

sample intervals in waste material is considered unlikely to significantly impact models of spatial

continuity of gold grades.

At Anwia, Salman and the Satellite Deposits, sample composites were flagged with primary

(mineralisation) domain codes using the wireframes of mineralised domains. Sub-domain(weathering) codes 1 to 4 (very weathered, moderately weathered, weakly weathered and fresh rock)

were allocated using the interpreted weathering surfaces. Salman pit and trench sample codes were

forced to Dom1sub1. The numbers of sample composites reporting to each of the modelling domains

at Anwia, Salman and the Satellite Deposits are listed in Table 17-9,,Table 17-10 and Table 17-11




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Table 17-9: Numbers of sample composites contained in Anwia modelling domains

Domain Subdomain 1

very

weathered

Subdomain 2

moderately weathered

Subdomain 3

weakly weathered

Subdomain 4

fresh rock

0 1349 2418 1565 5623

1 1921 3225 2967 8296

2 378 593 384 1136

Table 17-10: Numbers of sample composites contained in Salman modelling domains

Table 17-11: Numbers of sample composites contained in Satellite Deposits modelling

domains

Model Domain Subdomain 1

very

weathered

Subdomain 2

moderately

weathered

Subdomain 3

weakly weathered

Subdomain 4

fresh rock

0 2471 3574 5236 7232

1 248 466 561 510

2 25 35 73 36

3 278 677 913 585

4 20 25 58 116

5 545 599 1325 525

6 37 22 18 38

7 44 87 153 165

8 11 48 188 228

South

9 14 53 74 143

0 640 669 1059 1613

1 184 299 424 878

Central

2 186 387 921 22340 983 2346 1093 1807

1 28 103 95 469

2 19 68 34 42

3 144 386 121 270

4 22 170 87 179

5 9 96 6 -

North

6 22 149 13 -

Model Domain 0 Domain 1 Domain 2 Domain 3

Bokrobo 3287 203 147 -

Aliva 2194 460 278 104

Avrebo 1989 2428 - -

Nfutu 2505 - - -




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17.2 Exploratory Data Analysis

17.2.1 Anwia

In some of the Anwia domains only low numbers of samples lie within particular weathering horizons,

rendering it necessary to combine the samples prior to calculation of conditional statistics for resource

estimation. Considering that there are no pronounced differences in the tenor of gold mineralisation

across weathering boundaries, such as may be caused by supergene gold enrichment, the

approximations are considered reasonable.

The mineralised domains, with mean sample grades ranging from 1.5 to 4.5g/t Au, show coefficients

of variation (CV) from about 3.0 to 5.0, which are high, and are typical for gold deposits with gold

mineralisation similar to that seen at Anwia. CV at these levels indicate that reliable estimation of

recoverable gold grades using a linear estimator would be difficult. The grade populations show the

positive skew typical of gold deposits with high maximum grades seen in the main domain (Domain1).

17.2.2 Salman

As at Anwia, some of the domain/subdomain subsets contain too few samples to provide reliable

conditional statistics for input to resource estimation and have been combined as indicated by the

histograms in the appendix. In contrast to Anwia mineralisation, most Salman mineralised domains

show coefficients of variation less than 3.0 with many less than 2.0 and 70 per cent of the domains

contain maximum sample grades less than 30g/t Au.

17.2.3 Satelli te Deposits

At Bokrobo, Domains 0 and 2 contain very few composites with elevated gold values . Statistics show

the Domain 1 composites have an average grade of 2.69 g/t Au and a CV of 2.04. The maximum

sample grade is 35.9 g/t. The high CV suggests a modelling method such as MIK would be needed to

effectively model the Bokrobo Deposit gold resources

At Aliva, statistics for the mineralised domain composites show their CV’s lie between 1.1 and 1.5.

Maximum sample grades in the 3 mineralised domains are less than 9 g/t.

At Avrebo, the CV of the mineralised domain composites is 2.64 and at Nfutu, where all thecomposites were lumped into one domain, the CV of 5.23 reflects this.




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17.3 Spatial Continuity Analysis

17.3.1 Measures of Spatial Continu ity

Most resource estimation methods use a measure of spatial continuity to estimate the grade of blocksin a resource model. In some methods the measure is implicit; for example, a polygonal method

assumes that the grade is perfectly continuous from the sample to its surrounding polygon boundary.

Geostatistical methods like Ordinary Kriging and Indicator Kriging are among those methods for which

the continuity measure is explicit and is customised to the data set being studied. This measure in its

many forms is usually called the variogram.

Geostatistics provides several measures for describing spatial continuity: the variogram, the

covariance, the correlogram and many others. All are valid descriptions but not all provide a basis for

constructing kriging models of mineralisation. Whatever the method of description used, it is common

to use the term variogram in a generic sense to describe contour plots and directional plots of spatial

continuity measures. Throughout the present work, the maps and directional variograms used are all

based on the correlogram measure. Directional correlograms are displayed inverted so as toresemble familiar variogram plots. The use of the correlogram as a robust and reliable measure of

spatial continuity is proposed by Srivastava & Parker (1988) and Isaaks & Srivastava (1989). The

correlogram measure has the advantages of being standardised to a sill of 1 and being robust with

respect to clustering in the sample data. Models of the sample correlogram can be used directly in

Ordinary Kriging and Indicator Kriging.

The various parameters of the variogram model, such as the nugget effect and ranges in different

directions, describe properties of the statistical continuity of metal grades. For example, a variogram

with high nugget may indicate that there is a high level of error in the sample grades being used to

construct the variograms or that there is a high degree of variability in the grade over very short

distances in the mineralisation. A different range in one direction compared to another is likely to be

indicating that grade is more continuous in one direction than another.

17.3.2 Directional Control s on Gold Mineralisation

Gold and indicator variograms were calculated and modelled for each of the mineralised domains and

for the waste domains at each of Anwia, Salman and the Satellite Deposits of Bokrobo, Aliva and

Avrebo. Indicator transforms were undertaken with probability thresholds 0.1, 0.2, 0.3, 0.4, 0.5, 0.6,

0.7, 0.75, 0.8, 0.85, 0.9, 0.95, 0.97 and 0.99 for data in each domain subset. In all cases the data from

moderately weathered, weakly weathered and primary sub domains were combined for variogram

modeling. The modelled variograms as inputted into the resource model are complicated and are

difficult to visualise. Figure 17-12 to Figure 17-26 have been included to aid the interpretation of

these variogram models. The plots show the 3D-variogram surface maps for the median indicator

variogram for each of the mineralised domains modelled. The viewing angle is generally looking north

and down (arrow indicates north on the plots). No indicator variograms were able to be interpreted for

Nfutu owing to the limited data. Indicator variograms from Anwia were used as the Nfutu

mineralization is believed to be similar in nature to that seen at Anwia.

The spatial continuity of gold grades in Domains 1 and 2 at Anwia show contrasting orientations.

Figure 17-12 shows the median variogram for the main domain (Domain 1) and Figure 17-13 shows

the same plot for the granite domain (Domain 2). Whereas the dominant orientation to gold

mineralised structures in Domain 1 tends to strike NE and dip steeply towards north (modelling grid)




Page 188

the mineralisation in Domain 2 strikes in a northerly direction and dips steeply towards east. The

indicator and gold variograms for Domain 1 have been used to model the gold in Domain 0 at Anwia.

Figure 17-12: 3D variogram map,Indicator Threshold P 0.5, Anwia, Domain 1

Figure 17-13: 3D variogram map, Indicator Thresho ld P 0.5, Anwia, Domain 2




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At Salman, the spatial continuity of gold mineralization below the surficial cover, changes from

moderate to steep west dipping in the southern areas (Salman South domains) to being vertical or

east dipping at Salman Central , Salman North and Akanko areas. The local strike and dip changes

to the gold mineralisation are reflected in the variograms and these orientations are confirmed by

observations on the plots of gold grades in section and plan.

Figure 17-14, Figure 17-15 and Figure 17-16 show the 3D variogram map for the domains modeled

for indicator and gold variograms for use in the Salman South model. The variogram modeling was

restricted to Domains 1, 3 and 5 for this area owing to there being too few data in the remaining

domains to obtain useful variograms. The variograms modelled based on Domain 1 data were used

in the kriging of gold in Domain 2, Domain 5 variograms were used in Domains 0, 6, 7 and 9 and

Domain 3 variograms were used for modeling the gold grades in Domains 4 and 8.

Figure 17-14 3D variogram map, Indicator Threshold P 0.5, Salman South Model, Domain 1




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Figure 17-15 3D variogram map, Indicator Thresold P 0.5, Salman South Model, Domain 3

Figure 17-16 3D variogram map, Indicator Thresold P 0.5, Salman South Model, Domain 5




Page 191

Figure 17-17 and Figure 17-18 show the 3D variogram map for the domains modelled for indicator

and gold variograms for use in the Salman Central model. Domain 1 indicator and gold variograms

were used for modelling the gold in Domain 0.

Figure 17-17 3D variogram map, Indicator Threshold P 0.5, Salman Central Model, Domain 1

Figure 17-18 3D variogram map, Indicator Threshold P 0.5, Salman Central Model, Domain 2




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Figure 17-19, Figure 17-20 and Figure 17-21show the 3D variogram map for the domains modeled for

indicator and gold variograms for use in the Salman North model. The gold and indicator variograms

for Domain 2, 3 and a combined data set for Domains 4, 5 and 6 were modelled. The modelled

variograms of Domain 3 were used in the kriging of gold in Domains 0 and 1.

Figure 17-19 3D variogram map, Indicator Threshold P 0.5, Salman North Model, Domain 2

Figure 17-20: 3D variogram map, Indicator Threshold P 0.5, Salman North Model, Domain 3




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Figure 17-21: 3D variogram map, Indicator Threshold P 0.5, Salman North Model, Domain

4, 5 and 6

The gold and indicator variograms used in the kriging of gold in the surficial zones at Salman (Sub

Domain 1 for all primary domains) were sourced from previous ARL work. For completeness, the 3D

variogram map at the median indicator threshold is shown in Figure 17-22

Figure 17-22: 3D variogram map, Indicator Thresho ld P 0.5, Salman all models and all Primary

Domains, Sub Domain 1




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Figure 17-23 to Figure 17-26 show the 3D variogram maps for the domains modelled for indicator and

gold variograms for use in the models for individual Satellite Deposits..

Figure 17-23 3D variogram surface for the median indicator variogram model, Bokrobo Main

Zone (used for all domains)

Figure 17-24: 3D variogram surface for the median indicator variogram model, Aliva Domain 1




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Figure 17-25: 3D variogram surface for the median indicator variogram model, Aliva Domain 2

and 3 (also Domain 0)

Figure 17-26: 3D variogram surface for the median indicator variogram model,

Avrebo-All Domains




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17.4 Indicator Kriging

17.4.1 Indicator Krig ing for Recoverable Resource Estimation

Data viewing, compositing and wireframing at Anwia and Salman have been performed usingMicromine software. Exploratory data analysis, variogram calculation and modelling, and resource

estimation have been performed using Hellman & Schofield’s GS3 software. GS3 is designed

specifically for estimation of recoverable resources using multiple indicator kriging (MIK).

The MIK method was developed in the early 1980’s to address some of the problems associated with

estimation of resources in mineral deposits. These problems arise where sample grades show the

property of extreme variation and consequently where estimates of grade show extreme sensitivity to

a small number of very high grades. These characteristics are typical of many lode gold deposits,

where the coefficient of variation in samples commonly exceeds 2. MIK is one of a number of

methods that can be used to provide better estimates than the more traditional methods such as

ordinary kriging and inverse distance weighting.

It is fundamental to the estimation of resources that the estimation error is inversely related to the size

of the volume being estimated. To take the extreme case, the estimate of the average grade of a

deposit generated from a weighted average grade of the entire sample data set is much more reliable

than the estimate of the average grade of a small block of material within the deposit generated from

a local neighbourhood of data. Small blocks cannot provide the basis for reliable estimates of

recoverable resources.

Another fundamental notion relevant to the optimisation of resources to develop an open pit mine and

schedule is that the optimisation algorithm does not require the resource be defined on extremely

small blocks relative to data spacing.

The basic unit of an MIK block model is a panel that normally has the dimensions of the average drill

hole spacing in the horizontal plane. The panel should be large enough to contain a reasonable

number of mining blocks, or Selective Mining Units (SMUs; about 15). The SMU is the smallest

volume of rock that can be mined separately as ore or waste and is usually defined by a minimum

mining width. Based on experience at a number of open pit mining operations in hydrothermal gold

deposits, the dimensions of SMUs at Anwia and Salman are assumed to be in the order of 5mE x

8mN x 3mRL. The Satellite Deposits have slightly smaller SMUs in the order of 4mE x 8mN x 3mRL

excepting Nfutu which has SMU dimensions of 5mE x 8mN x 2mRL

The goal of MIK is to estimate the tonnage and grade of ore that would be recovered from each panel

if the panel were mined using the SMU as the minimum selection criteria to distinguish between ore

and waste. To achieve this goal, the following steps are performed:

• Estimate the proportion of each domain within each panel. This estimation can be achieved

by kriging of indicators of domain classifications of sample data points or by intersecting a

“template” model with the domain wireframes. At both Anwia and Salman panel domain

proportions were calculated by passing the panels through the domain wireframes.

• Estimate the histogram of grades of sample-sized units within each domain within each panel

using MIK. MIK actually estimates the probability of the grade within each panel being less

than a series of indicator threshold grades. These probabilities are interpreted as panel

proportions.




Page 197

• For each domain, and for each panel that receives an estimated grade greater than 0.0g/t Au,

implement a block support correction (variance adjustment) on the estimated histogram of

sample grades in order to achieve a histogram of grades for SMU-sized blocks. This step

incorporates an explicit adjustment for Information Effect.

• Calculate the proportion of each panel estimated to exceed a set of selected cut-off grades,

and the grades of those proportions.

• Apply to each panel, or portion of a panel below surface, a bulk density to achieve estimates

of recoverable tonnages and grades for each panel.

Apart from considerations of resource confidence classification, step 5 completes construction of the

resource model. The estimates of recoverable resources for each panel may be combined to provide

an estimate of global recoverable resources for the deposit.

17.4.2 Indicator Krig ing Parameters

The input parameters to Indicator Kriging of the Anwia, Salman and Satellite Deposits mineralisation

include:

• Indicator variogram models describing the spatial continuity of indicator variables within each

domain at each indicator threshold.

• Variogram models describing the spatial continuity of gold grades within each domain.

• Mean gold grades of each of the indicator classes within each domain.

Variogram models were rotated, where appropriate, to conform to the dip and plunge of mineralisation

as indicated by raw data plots and by variogram maps. The rotations specified conform to the

Cartesian convention wherein a positive rotation is clockwise when looking toward the positive end of

the rotation axis. Rotations are performed in the order in which they are listed. At Salman there are

insufficient regularly spaced samples in Domain 5 to allow calculation of reliable variograms;variogram models for nearby Domain 2 were applied to estimation. Similarly, resources in Domain 18

were estimated using variogram models of the east-dipping, greywacke-hosted mineralisation at

Salman North Domain 14 and variogram models from granite-hosted mineralisation in Domains 15, 16

and 17 were applied to estimation in Domain 19.

No cuts were applied to high-grade assays at either Anwia or Salman. The reduced ranges of

variogram models at high indicator thresholds effectively reduce the spatial influence of extreme

grades.

Table 17-12 and 17.13 show the grid framework and kriging parameters used in the indicator kriging

models at Aniwa and Salman. Table 17-14 to 17.17 show the grid framework and kriging parameters

used in the indicator kriging models at the Satellite Deposits. Panels were placed so that, as near aspossible, panel centroids lie between drill sections. No search rotations were imposed. The

allocation of resource confidence categories is described in Section 16.7 below.

At Anwia, all domain boundaries were treated as soft boundaries in the kriging process, i.e. estimation

in any one domain was permitted to “see” samples in neighbouring domains. This minimises

boundary problems and permits consistency of estimates in each panel. At Salman, all domain

boundaries were treated as soft boundaries with the exception of the interface between Domain 1, the

surficial material, and all underlying material. A hard boundary was imposed at that interface to




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prevent trench and pit samples influencing estimates in underlying saprolite material. At the Satellite

Deposits all domain boundaries were treated as soft boundaries.

Table 17-12: Anwia Model Framework & Krig ing Search Parameters (Rotated Space)

Panel Model Extents

East North Elevation

Panel origin (centroid) 574850 550045.5 -298.5

Panel Dimensions 20 25 3

No. of panels 44 39 119

Panel Discretisation 5 5 2

Kriging Parameters (all domains)

Criteria Measured Indicated Inferred

Min no. of data 16 16 8

Max no. of data per octant 4 4 4

Min no. of octants with data 4 4 2

X (east) search radius (metres) 25 32.5 32.5

Y (north) search radius (metres) 30 39 39Z (rl) search radius (metres) 10 13 13

Table 17-13: Salman Model Framework and Kr iging Search Parameters

Panel Model Extents, South Model


Panel origin (centroid) 553850 550812.5 836.5




Panel Model Extents, Central ModelEast North Elevation





Panel Model Extents, North Model






Kriging Parameters (all model areas)

Criteria Measured Indicated Inferred

Min no. of data 16 16 8

Max no. of data per octant 6 6 6

Min no. of octants with data 4 4 2

X (east) search radius (metres) 25 37.5 37.5

Y (north) search radius (metres) 30 45 45

Z (rl) search radius (metres) 10 15 15




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Table 17-14: Bokrobo Model Framework and Kriging Search Parameters

Panel Model Extents


Model origin (centroid) 575,130 548,062.5 712.5

Block Dimensions 20 25 3

Block Discretisation 5 5 2

GC SMU size 4 8 3

Kriging Parameters

Criteria Category 1 Category 2 Category 3

Min no. of data

Max no. of data per octant

Min no. of octants with data

X (east) search radius (metres)

Y (north) search radius (metres)

Z (rl) search radius (metres)

16

4

48

25

40

10

16

4

48

37.5

60

15

8

2

48

37.5

60

15

Table 17-15: Aliva Model Framework and Kr iging Search Parameters

Panel Model Extents


Model origin (centroid) 577,510 548,587.5 -61.5



GC SMU size 4 8 3

Kriging Parameters


Min no. of data






16

4

48

30

40

10

16

4

48

45

60

15

8

2

48

45

60

15

Table 17-16: Avrebo Model Framework and Kriging Search Parameters

Panel Model Extents





GC SMU size 4 8 3

Kriging Parameters


Min no. of data






16

4

32

20

25

10

16

4

32

30

37.5

15

8

2

32

30

37.5

15




Page 200

Table 17-17: Nfutu Model Framework and Kriging Search Parameters

Panel Model Extents





GC SMU size 4 8 3

Kriging Parameters


Min no. of data






16

4

32

20

25

10

16

4

32

30

37.5

15

8

2

32

30

37.5

15

17.5 Block Support Adjustment (Variance Adjustment)

17.5.1 General

The block support adjustment is one of the most important properties of a recoverable resource model

based on non-linear estimation methods like MIK. It is an essential part of the model and involves

important assumptions about the nature of the block grade distribution within each panel of the model.

Indicator Kriging provides a direct and reliable estimate of the histogram of grades of sample-sized

units within each panel of the model provided the panel dimensions are of an appropriate size.

However, ore is not selected on sample-sized units during mining; it is selected by shovels that have a

minimum mining width and loaded into trucks that are despatched to either ore or waste. The

flexibility of digging equipment and the size of the trucking equipment provide an indication of the size

of the smallest block of rock that will be mined as ore or waste. To estimate with some accuracy the

resources in a deposit that will be recovered with a certain set of mining equipment, the histogram of

grades of sample-sized units in a panel provided by MIK must be adjusted to account for the size of

the mining block.

There are a number of adjustment methods that can be used and most of these are described well in

Journel & Huijbregts (1978) or Isaaks & Srivastava (1989). These methods make three reasonable

assumptions:

• The average grade of sample-sized units and blocks within the panel is the same and is equal

to the estimated average grade of the panel.

• The variance, or spread, of the block grades within the panel is less than the variance of

grades of sample-sized units within the panel and the change of variance from sample-sized

units to blocks can be calculated from the variogram of gold grades.

• The approximate shape of the histogram of block grades can be reasonably predicted by

some appropriate assumptions.




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17.5.2 The Variance Adjustment

The size of the variance adjustment needed to obtain the variance of the block grade distribution

within the panel can be calculated using the rule of additivity of variances, which in the case of blocksupport adjustment is often called Krige’s Relationship:

Var (samples in a panel) = Var (samples in a block) + Var (blocks in a panel)

The variance of sample grades in a panel and the variance of samples within a block can be directly

calculated from the variogram of gold grades for the particular domain. The ratio of Var (blocks in

panel) to Var (samples in panel) is that required to implement the block support adjustment.

17.5.3 Shape of the Block Grade Distribution

There are a number of rules of thumb that are useful when making judgements about the shape of the

block grade distribution within each panel and they relate to the size of the variance adjustment ratio:

• If the variance adjustment ratio is greater than 0.7, it may be useful to assume that the shape

of the histogram of block grades is similar to that of the histogram of grades of sample-sized

units. This is known as the Affine Correction method. Its application to gold deposits is

usually inappropriate.

• If the variance adjustment ratio is between 0.3 and 0.7 and the information adjustment is

negligible, then the Indirect Lognormal Correction method of Isaaks & Srivastava (1989) can

be useful.

• If the variance adjustment ratio is less than 0.3, it is reasonable to assume there is a high

degree of symmetrization in the block grade histogram. If the histogram of sample grades in

a panel is positively skewed, the histogram of block grades is assumed to be lognormal in

shape. If the histogram of sample grades in a panel is approximately symmetrical or

negatively skewed, the block grade histogram is assumed to be normal in shape. The

theoretical support for these assumptions comes from the Central Limit Theorem of

probability. The theory supports the interpretation that as the variance adjustment ratio

becomes very small, the shape of the block grade distribution must approach that of a normal

distribution. In the GS3 software the shape of the histogram of sample-sized units is

assessed on a panel-by-panel basis and this approach to variance adjustment is called the

Lognormal-Normal Correction method. This model is well supported by reconciliation studies

of resource and grade control models.

17.5.4 The Information Effect

The variance adjustment described above is only part of the adjustment required in many gold

deposits because the short scale variation in gold grades is extreme, as is the case at Anwia and in

some of the Salman domains. This variance adjustment provides an estimate of the variance of true

block grades under the assumption that grade control selection will operate with knowledge of the true




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block grades. While this assumption is never absolutely true, it can be a reasonable assumption in

some deposits where the short scale variability is small and the grade control sampling density is

high. In many deposits, however, an additional variance adjustment must be undertaken to account

for the “Information Effect”.

In the absence of production information or grade control sampling, the Information Effect ratio is

based on the variograms of gold grade and on the grade control sample spacing expected to be usedduring mining.

17.5.5 Variance Adjus tments Applied to the Resource Models

Variance adjustment ratios applied in estimating Anwia and Salman recoverable gold resources are

listed in Table 17-18 and 17.19 respectively. These ratios have been applied using the Lognormal-

Normal Correction method (i.e., incorporating symmetrization of block grade distributions). Selective

mining (SMU) dimensions of 5mE x 8mN x 3mRL and grade control sample spacing of 8mE x 5mN x

1.5mRL have been assumed. Variance adjustment ratios applied in estimating Satellite Deposit

recoverable gold resources are listed in Table 17-20. These ratios have been applied using the

Lognormal-Normal Correction method (i.e., incorporating symmetrization of block grade distributions).

With the exception of Nfutu, selective mining (SMU) dimensions and grade control sample spacing of4mE x 8mN x 3mRL have been assumed. At Nfutu, a GC SMU size of 4mE x 8mN x 2mRL is

assumed. The variance adjustments applied to the models represent large reductions of variance,

typical of hydrothermal gold deposits.

Table 17-18: Variance adjustments appl ied to the Anwia resource model

Domain Panel to block

adjustment Information effect Total ratio

Domain 0 0.087 0.326 0.029

Domain 1 0.087 0.326 0.029Domain 2 0.183 0.482 0.088




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Table 17-19: Variance adjustments applied to the Salman resource model

Model Area Domain Panel to block

adjustment

Information effect Total ratio

South Domain 0 0.354 0.805 0.285

Domain 1 0.118 0.549 0.065

Domain 2 0.118 0.549 0.065

Domain 3 0.149 0.376 0.056

Domain 4 0.149 0.376 0.056

Domain5 0.354 0.805 0.285

Domain 6 0.354 0.805 0.285

Domain 7 0.354 0.805 0.285

Domain 8 0.149 0.376 0.023

Domain 9 0.354 0.805 0.285

Central Domain 0 0.176 0.321 0.057

Domain 1 0.176 0.321 0.057

Domain 2 0.108 0.572 0.062

North Domain 0 0.160 0.382 0.061

Domain 1 0.160 0.382 0.061

Domain 2 0.203 0.774 0.157

Domain 3 0.160 0.382 0.061

Domain 4 0.151 0.413 0.062

Domain 5 0.151 0.413 0.062

Domain 6 0.151 0.413 0.062

Table 17-20: Variance adjustments appli ed to the Satellite Deposits resource models

Deposit Domain Total Block Variance Correction

Bokrobo Domain 0, 1 and 2 0.172

Aliva Domain 0, 2 and 3 0.216

Domain 1 0.215

Avrebo Domain 0 and 1 0.082

Nfutu Domain 1 0.034




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17.6 Resource Classification

Panels in the resource models have been allocated confidence categories based on the number and

location of samples used to estimate proportions and grade of each panel. The approach is based onthe principle that larger numbers of samples, which are more evenly distributed throughout the search

neighbourhood, will provide a more reliable estimate. The number of samples and the particular

geographic configurations that may qualify the panel as Measured rather than Indicated or Inferred

are essentially the domain of the Qualified Person. The search parameters used to decide the

classification of a panel resource in this study are:

• Minimum number of samples found in the search neighbourhood.

For Indicated resources, this parameter is set to sixteen. For Inferred category, a minimum of eight

samples is required. This parameter ensures that the panel estimate is generated from a reasonable

number of sample data.

• Minimum number of spatial octants informed.

The space around the centre of a panel being estimated is divided into eight octants by the axial

planes of the data search ellipsoid. This parameter ensures that the samples informing an estimate

are relatively evenly spread around the panel and do not all come from one drill hole. For Indicated

resources, at least four octants must contain at least one sample. For Inferred panels, at least two

octants must contain data.

• The distance to informing data.

The search radii define how far the kriging program may look in any direction to find samples to

include in the estimation of resources in a panel. Panel dimensions and the sampling density in

various directions usually influence the length of these radii. It is essential that the search radii be

kept as short as possible while still achieving the degree of resolution required in the model. For

Measured resources at Anwia, the easting, northing and elevation search radii were set to 25, 30 and

10 metres respectively. For Indicated and Inferred resources the radii were expanded by 30 per cent

to 32.5mE x 39mN x 13mRL. Measured resources at Salman were estimated using search radii of

20mE x 30mN x10mRL; Indicated and Inferred resources were estimated with radii expanded by 50

per cent to 30mE x 45mN x 15mRL. For resources within the Satellite Deposits, the easting, northing

and elevation search radii were set according to the sample densities which varied for each deposit.

Search radii for resource confidence categories are detailed in Table 17-12 through to Table 17-17

At Anwia, Salman and the Satellite Deposits, the majority of panels in areas drilled at 25m x 25m

spacing or closer report to measured category, most panels in areas consistently drilled at 50m x 50m

spacing or less report to indicated category and panels in peripheral areas and at depth with less

consistent drill coverage report to inferred category.

17.7 Anwia Resource Model

Figure 17-27 to Figure 17-29 show an example cross-section through the Anwia model. The plots

show estimated mean panel grades (e-type estimates), recoverable proportions above 1.0g/t cut-off

and panel confidence categories. Figure 17-30, Figure 17-31 and Figure 17-32 show mean panel




Page 205

grades, recoverable proportions and confidence categories on an example plan view slice through the

model at about 40 metres below surface.

Figure 17-27: Anwia Panel Mean Grade Estimates, Section 550400N

Figure 17-28: Anwia Panel Recoverable Proporti ons at 1 g/t Cut-off, Section 550400N




Page 206

Figure 17-29: Anwia Panel Confidence Categories, Section 550400N

Figure 17-30: Anwia Panel Mean Grade Estimates, Plan at 1.5RL




Page 207

Figure 17-31: Anwia Panel Recoverable Proport ions at 1 g/t Cut-off , Plan at 1.5RL

Figure 17-32: Anwia Panel Conf idence Categories, Plan at 1.5RL




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17.8 Salman Resource Model

Figure 17-33 to Figure 17-35 show estimated mean panel grades, recoverable proportions above 1g/t

cut-off and panel confidence categories on an example cross-section through Salman Central, an

area that contributes a large proportion of resource tonnes and ounces. Figure 17-36, Figure 17-37

and Figure 17-38 show the same on an example plan view slice through the model at about 45 metresbelow surface.

Figure 17-33: Salman Panel Mean Grade Estimates, Section 552137.5N




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Figure 17-34: Salman Panel Recoverable Propor tions at 1 g/t Cut-off, Section 552137.5N

Figure 17-35: Salman Panel Confidence Categories, Section 552137.5N




Page 210

Figure 17-36: Salman Central Panel Mean Grade Estimates, Plan at 1001.5RL

Figure 17-37: Salman Central Panel Recoverable Proportions at 1g/t Cut-of f, Plan at 1001.5RL




Page 211

Figure 17-38: Salman Central Panel Confidence Categories, Plan at 1001.5RL

17.9 Satellite Deposits Resource Models

The resource estimates for each of the SAGP Satellite Deposits have been calculated at cut-off

grades which span the range appropriate for open pit mining: in the case of Bokrobo, between natural

surface and a maximum depth of about 200 metres; at Avrebo to a maximum depth of about 170

metres and at Aliva and Nfutu to a maximum depth of about 90 metres

The estimates have been truncated to the current land surface (interpreted from drill hole collars). At

Bokrobo, an allowance has also been made for resources which have depleted by artisanal mining

activities. The resources that were estimated within the “mined” wireframe were removed prior to

reporting the resources.

Figure 17-39 through to Figure 17-42 show 3D views of the resource composites and MIK models of

the various Satellite Deposits. The MIK model panels on these plots have been scaled to show theproportion of recoverable resource above 1.0 g/t cut-off and coloured by the average block grade.




Page 212

Figure 17-39: Bokrobo MIK Model showing 1.0g/t cut-off resource (model panels scaled in the

east dimension by the proportion of contained resource)




Page 213

Figure 17-40: Aliva MIK Model showing 1.0g/t cut -off resource (model panels scaled in the east

dimension by the proporti on of contained resource)




Page 214

Figure 17-41: Avrebo MIK Model showing 1.0g/t cut-off resource (model panels scaled in the

east dimension by the proportion of contained resource)




Page 215

Figure 17-42: Nfutu MIK Model show ing 1.0g/t cut-off resource (model panels scaled in the east

dimension by the proporti on of contained resource)

NB: Southern Zone

not modeled




Page 216

17.10 Mineral Resource Statement

Considering the precision inherent in the resource estimates and their informing data, the resource

estimate for the Project should be reported as listed in Table 17-21. The figures in the table have

been rounded to reflect the level of confidence in the resource. Figures may not sum owing to the

effects of rounding.

Table 17-21: Summary of Southern Ashanti Gold Project Resources at 0.8g/t cut-off

Category Measured Indicated Inferred

Deposit Cut off

grade

(g/t)

Mtonnes g/t

Au

k oz

Au

Mtonne

s

g/t

Au

k oz

Au

Mtonnes g/t

Au

k oz

Au

Anwia 0.8 6.2 2.01 400 2.8 2.00 180 2.6 1.7 140

Salman 0.8 11.4 1.73 630 5.6 1.54 280 2.5 1.5 125

Satellite

Deposits0.8 1.0 2.10 70 1.5 1.57 70 1.3 1.8 75

Total 18.6 1.84 1,100 9.8 1.67 530 6.4 1.6 340

17.11 Other

Information regarding legal titles, environmental, permitting, taxation, socio-eceonomic and political

issues that may impact upon the mineral resources are described elsewhere in this report and have

not been independently verified by the author. Nevertheless the author believes that there are noissues arising from such considerations that would materially impact the quantum of the mineral

resources or Adamus’ ability to access them.




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18.0 ORE RESERVE ESTIMATE

18.1 Introduction

The scoping study completed in March 2006 was based on a processing throughput rate of 1.3 million

tonnes of gold ore per annum, requiring the movement of 6.0-8.0 million tonnes of total material per

annum from the open pits. The study indicated a mine life of more than 6 years.

Following the completion of the scoping study, Mining Solutions Consultancy Pty Ltd (MSC) prepared

the open pit mining section of the feasibility study for further development of the Project. In line with

the scoping study, the feasibility study was based on an ore throughput rate of 1.3 million tonnes per

annum. The finalisation of the mining study followed the completion of the geotechnical, metallurgical

and resource studies based on the in-fill resource and test-work drilling results.

In April 2008 Mining Solutions Consultancy Pty Ltd (MSC) prepared an update of the open pit mining

section of the feasibility study incorporating changes to the mining study parameters brought about by

an increase in the estimated mineral resources, an increase in the gold price to $800/oz, a

government royalty of 3% and an Anwia royalty of 1% and cost escalations in mining and processing

since December 2006.

The selected mining method for the Project is conventional open pit mining including drilling, blasting,

loading and hauling operations carried out by a mining contractor with experience in Ghana. The

construction of the processing plant is currently planned close to the Salman deposit. The mining

contractor will also be responsible for the haulage of the ore from the Anwia pit to the plant site via a

public road.

The topography in the Project area consists of a series of 20-50m high hills with the peaks separatedby horizontal distances of 200-400m. The average slope of the hills is approximately 20 degrees,

varying generally between 15 and 30 degrees. The low lying areas between the wider spaced hills

contain standing surface water in the form of swamps. The low lying area to the east of the northern

areas of the Salman deposit has been identified as a major swamp. Generally, the mine haul roads

and waste dumps will be located at the relatively higher elevations avoiding the swamp areas.

The Salman and Anwia deposits, located 9km apart to the east and west respectively in the Project

area are the subject of the open pit mining study. The depths of the proposed pits at Salman will

generally vary between 30 and 70m depending on the variable topography over a strike length of

7km. The final open pit at Anwia was to be developed in two major stages to a depth of 150m.




Page 218

18.2 Mining Study Scope

The tasks completed in parallel for the open pit planning of the Salman and Anwia deposits can be

summarised as follows:

• mining contract budget pricing tender preparation and evaluation of the received submissions

• Data transfer and preparation of the mining resource, metallurgical recovery and operating

costs models for pit optimisation and reserve estimation purposes

• Pit limit optimisations and sensitivity analysis to determine the ultimate pit limits and stages for

incremental pit development

• Pit designs, analysis of mining quantities and preliminary schedules to determine the open pit

development strategy

• Site layout and waste dump designs including waste backfill options

•

Various iterations of the interdependent planning tasks above as the confidence in theparameters increased during the study

• Calculation of open pit mining inventories and reserves statement

• Preparation of detailed production schedules for mining, milling and stockpiling operations

based on multiple pits and ore types

• Mining cost model and schedule based on final production schedule and budget prices

received from an experienced Ghana based mining contractor

• Mining equipment and manpower estimates based on final production schedule

• Shadow mining cost estimate from first principles to confirm the budget prices received from

mining contractors

• Reconciliation of the study results and preparation of final study report

• Site visit before the completion of the study report to identify any issues that may affect the

development of the open pits and operating cost estimates

In terms of the mine planning software, Whittle Four-X optimisation software was used in the

generation and analysis of the optimal pit shells. MineSight was used as the main software to store

the resource models, create the pit optimisation models; design the open pits and waste dumps,

calculate the mining inventories, visualise and plot the resource and design data. The resultant

production schedules and mining cost models were prepared using an advanced spreadsheet

capable of evaluating and balancing ore and waste production from multiple open pits.

18.3 Parameters

This section of the Report discusses only the changed parameters in the updated Mining Study. Refer

to the previous Technical Report for all other details used previously for pit optimisations, designs,

production scheduling and operating cash flow estimates.

The main changes in the mining study parameters can be summarised as follows:

• Gold price increase of approximately 35%

• Contract mining cost increase of approximately 15%




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• Processing cost increase of approximately 15%

• Update of the Salman and Anwia resource models by Hellmann & Schofield (H&S,

March 2008) as a result of further drilling

18.3.1 Parameters Summary

The parameters used in this ore reserve update study have been summarised in Table 18-1 and

Table 18-2 for Salman and Anwia deposits respectively.

18.3.2 Gold Price and Royalty

The most significant change in the reserves update is the change in the gold price from $575/oz to

$800/oz. The government royalty of 3.0% and other royalty on the Anwia ore of 1.0% have been

deducted from the gold price in the calculation of the net revenues in pit limit optimisations.

18.3.3 Throughput Costs

Lycopodium provided the update for the general, administration and processing costs for the project

based on the changes in the consumable, power and labour rates. The mining supervision, grade

control, ore haulage and crusher loading costs have been increased depending on the increases in

the contract mining rates and labour costs. The throughput costs assigned to the ore tonnes

processed have increased by approximately 15% since the feasibility study.

18.3.4 Contract Mining Costs

The mining contractors were requested to confirm / revise their budget prices for the feasibility study.

The data sent to the contractors were based on the feasibility study final results reported in May 2007.

Although the requests were sent to PMC, AMS and PWL mining contractors, the responses were

received only from AMS and PWL as can be seen in Appendix C.

The mining costs used in the study have been revised on the basis of the indications and revised

rates received from the mining contractors. The overall mining costs have increased by

approximately 15% based on the 20% increase in load and haul rates, 10% increase in drilling and

blasting rates and 5% increase in the supervision and overheads. The most significant increase in the

load haul rates is due to the increase in fuel price from US$0.72/litre to US$1.16/litre.




Page 220

Throughput Rate Oxide Trans. Primary

Mill Feed Mtpa 1.30 1.30 1.30

Throughput Costs Oxide Trans. Primary

Metallurgy Labour $/t ore 1.35 1.35 1.35 Based on Lycopodium Memo 7 Dec 07

General & Admin Labour $/t ore 1.54 1.53 1.53 Based on Lycopodium Memo 7 Dec 07

Operating Consumables $/t ore 3.45 5.73 5.73 Based on Lycopodium Memo 7 Dec 07

Power $/t ore 2.59 2.78 2.78 Based on Lycopodium Memo 7 Dec 07

Maintenance $/t ore 0.92 0.92 0.92 Based on Lycopodium Memo 7 Dec 07

Laboratory $/t ore 0.44 0.44 0.44 Based on Lycopodium Memo 7 Dec 07

General & Admin $/t ore 1.56 1.55 1.55 Based on Lycopodium Memo 7 Dec 07

Total Processing and G & A 11.85 14.30 14.30

Mine Supervision $/t ore 1.12 1.12 1.12 Escalated by 5% since Dec'06

Grade Control $/t ore 1.18 1.18 1.18 Escalated by 10%, Since Dec'06

Ore Haulage $/t ore 0.54 0.54 0.54 Escalated by 21%, since Dec'06Crusher Loading $/t ore 0.59 0.59 0.59

Rehandling Cost (25% of ore) $/t ore 0.32 0.32 0.32 Escalated by 21%, Since Dec'06

Rehabilitation Cost $/t ore 0.15 0.15 0.15

Total Mining $/t ore 3.91 3.91 3.91

Total Throughput Cost $/t ore 15.76 18.21 18.21

Processing Parameters Oxide Trans. Primary

Recovered Au Cut-off at US$800/oz Price g/t 0.64 0.73 0.73

Residual grades and recoveries vary in the recovered grade model by rock types, oxidisation degree and deposit areas.

Gold Price and Royalty Min Base Max

Gold Price US$/oz 600 800 900 from min to max in $25/oz increments

US$/gm 19.29 25.72 28.94

Royalty 3.6% 3.6% 3.6%

Gold Price after Royalty US$/oz 578.4 771.2 867.6

US$/gm 18.60 24.79 27.89

Discount Rate %pa 10%

Overall Slope Angles Oxide Fresh

East Wall deg 41.0 48.0

Other Walls deg 44.5 44.5

Resource Model

Waste dilution & Recovery Included in the recoverable resource model

Classif ication used: Measured and Indicated only

Table 18-1 Salman Pit Optimisation Parameters Summary

Table 18-2 Anwia Pit Optimisation Parameters Summary




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Throughput Rate Oxide Trans. Primary

Mill Feed Mtpa 1.30 1.21 1.30

Throughput Costs Oxide Trans. Primary

Metallurgy Labour $/t ore 1.36 1.46 1.36 Based on Lycopodium Memo 7 Dec 07

General & Admin Labour $/t ore 1.54 1.65 1.54 Based on Lycopodium Memo 7 Dec 07

Operating Consumables $/t ore 4.20 5.09 5.23 Based on Lycopodium Memo 7 Dec 07

Power $/t ore 2.94 3.08 3.16 Based on Lycopodium Memo 7 Dec 07

Maintenance $/t ore 0.93 0.99 0.93 Based on Lycopodium Memo 7 Dec 07

Laboratory $/t ore 0.44 0.47 0.44 Based on Lycopodium Memo 7 Dec 07

General & Admin $/t ore 1.55 1.67 1.55 Based on Lycopodium Memo 7 Dec 07

Total Processing and G & A 12.96 14.40 14.21

Mine Supervision $/t ore 1.12 1.21 1.12 Escalated by 5% since Dec'06

Grade Control $/t ore 1.18 1.26 1.18 Escalated by 10%, Since Dec'06

Ore Haulage $/t ore 2.29 2.29 2.29 Escalated by 21%, since Dec'06

Crusher Loading $/t ore 0.59 0.59 0.59

Rehandling Cost (25% of ore) $/t ore Salman Only

Rehabilitation Cost $/t ore 0.15 0.16 0.15

Total Mining $/t ore 5.33 5.51 5.33

Total Throughput Cost $/t ore 18.29 19.91 19.54

Processing Parameters Oxide Trans. Primary Oxide LG Trans LG Prim LG

Leach Residue Grades g/t 0.086 0.202 0.242 0.086 0.202 0.242

Average Grade g/t 0.88 1.06 1.09 0.76 0.90 0.95

Process Recovery at Cut-off % 90.2% 80.9% 77.8% 88.7% 77.6% 74.5%

Cut-off Grade at US$800/oz Price g/t 0.82 0.99 1.01 0.64 0.79 0.82

Gold Price and Royalty Min Base Max

Gold Price US$/oz 600 800 900 from min to max in $25/oz increments

US$/gm 19.29 25.72 28.94

Royalty No royalty i 3.6% 3.6% 3.6%

Gold Price after Royalty US$/oz 578.4 771.2 867.6

US$/gm 18.60 24.79 27.89

Discount Rate %pa 10%

Overall Slope Angles

South & North Walls (Incl. ramp) deg 39.5

East and West Walls (Incl. ramp) deg 43.5

Resource Model

Waste dilution & Recovery Included in the recoverable resource model

Classification used: Measured and Indicated only

`




Page 222

Resource Area Oxide Transitional Sulphide Total

Salman Deposit 85.2% 72.9% 58.0% 79.3%

Anwia Deposi t 95.2% 90.7% 90.6% 90.9%

Project Total 86.4% 83.8% 87.2% 86.2%

18.3.5 Resource Model and Surfaces

The revised Salman and Anwia resource models were received from Hellmann and Schofield (H&S)

in the same format as received in the feasibility study. The surfaces defining the oxide, transitional

and sulphide resource zones were also updated by H&S. The resource summaries for Salman and Anwia deposits received from H&S can be seen in Appendix D.

18.3.6 Other Parameters

Other parameters used in the reserve update study are the same as used in the feasibility study in

2007.

The resultant process recoveries summarised in Table 18-3 are slightly (-1%) lower due to the lower

cut-off grades and slightly lower resource model grades within the pit limits.

Table 18-3 Resultant Average Process Recoveries for Resource Areas

18.4 Pit Limit Optimisations

The pit optimisations have been carried out for a range of gold prices from $600/oz to $900/oz in

$25/oz increments. Table 18-4 to Table 18-6 summarise the pit optimisation results for Salman,

Anwia deposits and project totals. The details of the pit optimisation input parameters and results

have been provided in Appendix E.

Although in general the global resources have increased significantly, especially in the Salman group

of deposits, the review of the pit optimisation results indicated that most of the resource increase isperipheral to the optimal pit limits. The results also indicated a slight reduction in the resource grade

within the optimal pit limits.

Compared to the feasibility study results reported in May 2007, the operating cost per ounce is

significantly higher for the optimal pit shells. The review of the optimisation results and the resource

models indicated to the following breakdown of factors causing an increase of $130/oz in the

operating costs:

• $50/oz cost increase due to the approximately 15% increase in the operating costs




Page 223

• $50/oz cost increase due to the higher gold price, larger pit shells and lower cut-off

grade (approximately $700/oz cost for the incremental ounces between $575/oz and

$800/oz gold price)

• $30/oz cost increase due to the approximately 10% reduction in resource grade within

the feasibility study pit designs



Adamus Resources Limited Southern Ashanti Gold Project Feasibility Study – Open Pit MiningDoc.Ref.: Draft Technical Report FINAL 21 Aug 2008.doc

Physicals Economics

Gold Ore Waste Total W/O Contn'd Recv'd Recv. Mining Th'put Reven. Undisc'd Mining Th'put Unit

Price Pit Total Au Ratio Ounces Ounces Cost Cost C'Flow Cost Cost Cost

US$/oz No Tonnes g/t Tonnes Tonnes t/t koz koz % $'000 $'000 $'000 $'000 $/t $/t $/oz

600 14 2,855,966 2.57 6,232,469 9,088,435 2.18 236,367 198,354 83.9% 16,718 46,289 114,729 51,722 1.84 16.21 318

625 15 3,025,535 2.53 6,759,780 9,785,315 2.23 245,877 205,477 83.6% 18,008 49,094 123,801 56,699 1.84 16.23 327

650 16 3,242,899 2.48 7,582,572 10,825,471 2.34 258,068 214,712 83.2% 19,924 52,735 134,541 61,882 1.84 16.26 338

675 17 3,437,366 2.43 8,110,957 11,548,323 2.36 268,051 222,018 82.8% 21,268 55,988 144,464 67,208 1.84 16.29 348

700 17 3,565,795 2.38 7,982,528 11,548,323 2.24 272,621 225,158 82.6% 21,268 58,054 151,935 72,613 1.84 16.28 352

725 18 3,797,482 2.33 8,807,522 12,605,004 2.32 284,913 233,917 82.1% 23,240 61,921 163,483 78,322 1.84 16.31 364

750 19 4,034,449 2.29 9,789,615 13,824,064 2.43 296,648 242,831 81.9% 25,522 65,877 175,566 84,167 1.85 16.33 376

775 20 4,160,614 2.27 10,418,076 14,578,690 2.50 303,945 247,702 81.5% 26,924 68,010 185,059 90,125 1.85 16.35 383

800 21 4,632,581 2.24 12,492,747 17,125,328 2.70 333,881 265,557 79.5% 32,090 76,182 204,801 96,529 1.87 16.44 408

825 21 4,705,919 2.22 12,419,409 17,125,328 2.64 336,519 267,070 79.4% 32,090 77,374 212,397 102,933 1.87 16.44 410

850 22 4,955,407 2.21 13,700,316 18,655,723 2.76 351,874 276,410 78.6% 35,122 81,710 226,486 109,654 1.88 16.49 423

875 22 4,971,013 2.21 13,684,710 18,655,723 2.75 352,519 276,742 78.5% 35,122 81,987 233,430 116,321 1.88 16.49 423

900 22 5,096,427 2.17 13,559,296 18,655,723 2.66 355,923 279,099 78.4% 35,122 83,984 242,146 123,040 1.88 16.48 427

Table 18-4: Salman Pit Optimisation Results




Physicals Economics


Price Pit Total Au Ratio Ounces Ounces Cost Cost C'Flow Cost Cost Cost

US$/oz No Tonnes g/t Tonnes Tonnes t/t koz koz % $'000 $'000 $'000 $'000 $/t $/t $/oz

600 12 3,251,187 2.79 17,613,134 20,864,321 5.42 291,696 269,757 92.5% 45,737 64,343 156,028 45,948 2.19 19.79 408

625 13 3,753,971 2.69 19,987,959 23,741,930 5.32 324,435 299,122 92.2% 51,986 74,302 180,222 53,934 2.19 19.79 422

650 15 4,116,663 2.62 21,439,243 25,555,906 5.21 346,105 318,078 91.9% 56,034 81,498 199,311 61,779 2.19 19.80 432

675 15 4,118,504 2.61 21,437,402 25,555,906 5.21 346,180 318,135 91.9% 56,034 81,535 207,006 69,437 2.19 19.80 432

700 16 4,404,756 2.54 22,073,699 26,478,455 5.01 359,903 330,144 91.7% 58,169 87,200 222,778 77,409 2.20 19.80 440

725 16 4,609,490 2.48 21,868,965 26,478,455 4.74 367,443 336,109 91.5% 58,169 91,254 234,906 85,483 2.20 19.80 445

750 17 5,005,786 2.49 26,860,902 31,866,688 5.37 400,530 366,189 91.4% 70,485 99,124 264,754 95,145 2.21 19.80 463

775 18 5,488,372 2.46 31,017,390 36,505,762 5.65 433,286 395,431 91.3% 81,673 108,756 295,429 105,000 2.24 19.82 482

800 18 5,828,717 2.37 30,677,045 36,505,762 5.26 444,338 404,182 91.0% 81,673 115,423 311,709 114,613 2.24 19.80 488

825 18 5,847,483 2.37 30,658,279 36,505,762 5.24 444,960 404,663 90.9% 81,673 115,796 321,823 124,354 2.24 19.80 488

850 19 5,879,490 2.37 30,930,709 36,810,199 5.26 447,075 406,529 90.9% 82,412 116,429 333,107 134,266 2.24 19.80 489

875 20 6,108,978 2.32 31,283,487 37,392,465 5.12 455,981 413,922 90.8% 83,787 121,051 349,142 144,304 2.24 19.82 495

900 21 6,155,353 2.31 31,465,297 37,620,650 5.11 458,077 415,721 90.8% 84,316 121,968 360,680 154,396 2.24 19.81 496

Table 18-5: Anwia pit Optimisation Results




Physicals Economics


Price Total Au Ratio Ounces Ounces Cost Cost C'Flow Cost Cost Cost

US$/oz Tonnes g/t Tonnes Tonnes t/t koz koz % $'000 $'000 $'000 $'000 $/t $/t $/oz $

600 6,107,153 2.69 23,845,603 29,952,756 3.90 528,062 468,110 88.6% 62,455 110,632 270,757 97,670 2.09 18.12 370 625 6,779,506 2.62 26,747,739 33,527,245 3.95 570,312 504,599 88.5% 69,994 123,396 304,023 110,633 4.03 18.20 383

650 7,359,562 2.55 29,021,815 36,381,377 3.94 604,173 532,791 88.2% 75,958 134,233 333,852 123,661 4.03 18.24 395

675 7,555,870 2.53 29,548,359 37,104,229 3.91 614,231 540,154 87.9% 77,302 137,523 351,470 136,645 4.03 18.20 398

700 7,970,551 2.47 30,056,227 38,026,778 3.77 632,524 555,302 87.8% 79,437 145,254 374,713 150,022 4.04 18.22 405

725 8,406,972 2.41 30,676,487 39,083,459 3.65 652,357 570,026 87.4% 81,409 153,175 398,389 163,805 4.04 18.22 412

750 9,040,235 2.40 36,650,517 45,690,752 4.05 697,178 609,020 87.4% 96,007 165,001 440,320 179,312 4.06 18.25 429

775 9,648,986 2.38 41,435,466 51,084,452 4.29 737,230 643,132 87.2% 108,597 176,766 480,488 195,125 4.08 18.32 444

800 10,461,298 2.31 43,169,792 53,631,090 4.13 778,219 669,739 86.1% 113,763 191,605 516,510 211,142 4.11 18.32 456

825 10,553,402 2.30 43,077,688 53,631,090 4.08 781,479 671,733 86.0% 113,763 193,170 534,220 227,287 4.11 18.30 457

850 10,834,897 2.29 44,631,025 55,465,922 4.12 798,948 682,939 85.5% 117,534 198,139 559,593 243,920 4.12 18.29 462

875 11,079,991 2.27 44,968,197 56,048,188 4.06 808,500 690,664 85.4% 118,909 203,038 582,572 260,625 4.12 18.32 466

900 11,251,780 2.25 45,024,593 56,276,373 4.00 814,000 694,820 85.4% 119,438 205,952 602,826 277,436 4.12 18.30 468

Table 18-6: Project Total Pit Optimisation Results



Adamus Resources Limited April 2008Southern Ashanti Gold Project Feasibility Study – Open Pit MiningDoc.Ref.: Draft Technical Report FINAL 21 Aug 2008.doc Page 227

18.5 Mine Design

The pit optimisation results update provided the basis for the new pit designs. The criteria for the pitdesigns were maintained the same as reported in the feasibility study (May 2007).

The size of the Salman pit designs shown in Figure 18-1 to 18-3 are similar to the previous feasibility

study designs. Similarly, the South and Central areas have been planned to mine in two pit stages

each along strike. The designs extend further north with the addition of a small pit (Akanko North 2)

500m north of the Akanko North Pit.

The new Anwia pit design is 30m deeper than the feasibility design and larger enough to mine the

main pit in three cutbacks. The Anwia ultimate pit design and the pit stage designs have been shown

in Figure 18-4 and Figure 18-5.




Figure 18-1: Salman Central and South Pit Designs




Figure 18-2: Salman North Pit Designs




Figure 18-3: Akanko Pit Designs




Figure 18-4: Anwia Ultimate Pit Design (Local Grid)




Figure 18-5: Anwia Pit Stage Designs (Local Grid)




Mill Feed Ore Cut-off Grades

Salman Granite Formation Greywacke Formation

Deposit Oxide Transition Sulphide Oxide Transition Sulphide

Area Upper Lower Upper Lower Au g/t Au g/t Au g/t Au g/t Au g/t Au g/t Au g/t Au g/t

Akanko 0.8 0.8 0.9 1.0 0.8 0.9 1.2 1.7

Akanko South 0.8 0.9 1.1 1.2 1.0 1.3 1.7 2.2

North Hill 0.9 1.1 1.2 1.5 1.0 1.6 2.1 2.2

Salman North 0.9 1.1 1.2 1.5 1.0 1.6 2.1 2.2

Teberu Footwall 1.1 1.3 1.7 2.7

Teberu 0.9 1.1 1.2 1.5 1.1 1.3 1.7 2.2

Nugget Hill 0.9 1.1 1.2 1.5 1.1 1.2 1.7 2.2

Salman Central 0.9 1.1 1.7 2.7

Salman South 0.8 1.0 1.4 2.1

Salman SW 0.8 1.0 1.4 1.7

Oxide Transition Sulphide

Au g/t Au g/t Au g/t

Anw ia Depo si t 0.8 1.0 1.0

Low Grade Ore Cut-off Grades

Salman Granite Formation Greywacke Formation

Deposit Oxide Transition Sulphide Oxide Transition Sulphide

Area Upper Lower Upper Lower

Au g/t Au g/t Au g/t Au g/t Au g/t Au g/t Au g/t Au g/t

Akanko 0.6 0.6 0.7 0.8 0.6 0.7 1.1 1.5

Akanko South 0.7 0.8 1.0 1.0 0.9 1.2 1.6 2.0

North Hill 0.7 0.9 1.1 1.3 0.9 1.5 2.0 2.0

Salman North 0.7 0.9 1.1 1.3 0.9 1.5 2.0 2.0

Teberu Footwall 1.0 1.2 1.6 2.5

Teberu 0.7 0.9 1.1 1.3 1.0 1.2 1.6 2.0

Nugget Hill 0.7 0.9 1.1 1.3 1.0 1.1 1.6 2.0

Salman Central 0.8 0.9 1.6 2.5

Salman South 0.6 0.8 1.2 1.9Salman SW 0.6 0.8 1.2 1.5

Oxide Transition Sulphide

Au g/t Au g/t Au g/t

Anw ia Depo si t 0.7 0.8 0.9

Note:

In the mining inventory, the Salman ore quantities are based on the operating value calculation for each model block.

18.6 Mining Quantities and Reserves

18.6.1 Cut-of f Grades

The revised cut-off grades for the calculation of the ore reserves have been listed in Table 18-7 for the

deposit areas and ore types. The cut-off grades are slightly lower (~0.1g/t) as a results of the changes

in the gold price, operating costs and process recoveries.

Table 18-7: Cut-off Grades for Ore Reserves




18.6.2 Open Pit Quanti ties

The quantities reported within the pit designs have been summarised in Table 18-8 classified by the

ore types and deposit areas. The indicative economic values and the operating costs in the table

provided guidance to determine the mining order of the pit areas in the production schedules as

follows:

1. Salman Central pits would be mined first at the lowest operating cost per ounce.

2. The mining of the larger Anwia pit would be completed second at a reasonable cost and

providing mill feed for 4 years from a single source.

3. Salman South and Southwest pits would be mined third and waste would be backfilled

into the Central Stage 1 and 2 pits.

4. Salman North pits would be mined next with partial backfilling of the mined pits.

5. The Akanko pits would be mined last, with the cost effective backfilling of the smaller pits

to the South and North.

Note that the mining transition between the areas 3, 4 and 5 above would not be a clear cut since

these areas need to be mined simultaneously due to the smaller size of the open pits.

The details of the pit quantities by mining levels have been reported in Appendix G.

18.6.3 Ore Reserve Estimate

Based on the revised feasibility study parameters and in compliance with the JORC guidelines

(AusIMM 2004), the ore reserves estimate for the Southern Ashanti Gold Project is provided in Table

18-9. All the estimated ore reserves are included within the mineral resource as defined in Table

17-21

This report summarises the changes in the mining study parameters and the effect on the ore reserve

estimates following the completion of the feasibility study in May 2007. The reader should refer to the

feasibility study report for the complete details of the project

All the work providing the basis for the ore reserves statement has been carried out by Mr Tamer

Dincer, BSc and MSc degrees in mining engineering, a fulltime employee of Mining Solutions

Consultancy Pty Ltd, who is a Member of Australasian Institute of Mining & Metallurgy, a Member of

the Mineral Industry Consultants Association and who is a Competent Person under the JORC

definition of competent person for the estimation of ore reserves.




Proved Reserve Probable Reserve T

Tonnesx1000 Au g/t Ouncesx1000 Tonnesx1000 Au g/t Ouncesx1000 Tonnesx1000

ANWIA TOTAL 5,497 2.08 367 1,579 2.31 117 7,076

SALMAN TOTAL 4,622 2.09 310 320 1.91 20 4,942

PROJECT TOTAL 10,119 2.08 677 1,899 2.24 137 12,018

Anwia 1 1,381 2.01 89 35 1.60 2 1,416

Anwia 2 1,645 2.15 114 267 2.44 21 1,912

Anwia 3 2,327 2.10 157 1,238 2.32 92 3,565

Anwia Nor th 144 1.48 7 39 1.56 2 183

ANWIA TOTAL 5,497 2.08 367 1,579 2.31 117 7,076

Akanko N2 112 1.41 5 7 1.69 0 119

Akanko North 92 2.09 6 1 1.17 0 93

Akanko 539 1.54 27 74 1.70 4 613

Akanko South 181 1.66 10 23 1.77 1 204

Salman North 719 2.20 51 58 2.20 4 777

Teberu Footwall 138 2.20 10 28 2.23 2 166

Nugget Hill 193 2.45 15 69 2.11 5 262

Salman Central 1 973 2.73 85 10 2.06 1 983Salman Central 2 681 2.66 58 8 1.70 0 689

Salman South 1 664 1.30 28 7 0.96 0 671

Salman South 2 197 1.12 7 24 0.98 1 221

Salman SW 133 2.07 9 11 2.83 1 144

SALMAN TOTAL 4,622 2.09 311 320 1.91 20 4,942

Table 18-9: Ore Reserves by Mining Areas and Ore Types




19.0 OTHER RELEVANT DATA AND INFORMATION

19.1 Production Schedules

As the reserve base for the project increased with the increasing gold price, the larger throughput

rates have become an option. In addition to the 1.3Mtpa throughput rate schedule prepared in the

original feasibility study, 1.5Mtpa and 1.8Mtpa cases have been included in the update study. Table

19-1 below summarises and compares the schedules while the detailed schedules can be seen in

Appendix H.

Table 19-1 Product ion Schedule Results Summary

Schedule Version 7a 7b 7c

Processing Rate Mt pa 1.3 1.5 1.8

Processing life Years 9.25 8.00 6.75

Average mill feed ore grade (First 5 years) g/t Au 2.46 2.44 2.38

First 5 years’ average gold production Oz pa 91,000 104,000 120,000

Maximum ROM stockpile balance Tonnes 310,000 525,000 510,000

Low grade stockpile feed grade g/t Au 0.91 0.91 0.91

Low grade gold production (last year) oz pa 27,500 32,000 38,500

Open pit mine life Years 8.5 7.5 6.5

Pre-stripping requirement Tonnes 260,000 260,000 260,000

Pre-stripping duration (incl. ramp up) Months 1-2 1-2 1-2

Total material movement for 1st year Mt pa 4.1 5.1 6.7

Total movement for 2nd year Mt pa 6.9 7.8 11.8

Total movement for 3rd and 4 th years Mt pa 11.6 13.2 16.8 , 9.7

Total movement for 5th year Mt pa 7.3 7.0 8.0

Average movement for rest of mine life Mt pa 6.3 7.0 8.0




20.0 INTERPRETATION AND CONCLUSIONS

This updated Technical Report incorporates new data into the mining and resource sections of the

previously lodged Technical Report of December 2007. The principal components of the Update are a

Project Resource Update completed in January 2008 (Hellman and Schofield; 2008a) and the

Summary Resource Report for the Satellite Deposits (Hellman and Schofield; 2008b) and an UpdateStudy of the open pit mining components of the feasibility study completed in April 2008 (Dincer;

2008).

Since the compilation of the Technical Report, Adamus’ exploration in the SAGP has focussed on two

objectives:

• Converting inferred resources to measured and indicated resources by infill drilling (and

thereby allow their incorporation into the mining reserve); and

• Proving up measured and indicated resources on smaller prospects proximal to the major

deposits to provide more flexibility in the mining and processing of the SAGP resources.

The January 2008 resource estimate at Anwia incorporated data from the 2006 to 2007 diamond drill

program undertaken to infill drill coverage between about 130 and 175 metres below surface. The

updated resource model converted much of the previous inferred resource to Measured and Indicated

categories. The updated resource model resulted in a deeper optimum pit. Further drilling northwest

of the main mineralisation in this area may delineate resources amenable to underground mining.

The areas that contribute the most resource ounces to the Salman deposit have been drilled at a

spacing that allows confident estimation of recoverable resources. Infill drilling in the Salman north

area to convert resources to Measured and Indicated status has resulted in extensions of optimum pit

shells and additional ore reserves. Additional resource and reserve ounces could be found with

shallow drilling in the Teberu and Akanko North areas. Deeper drilling at many of the deposits on theSalman trend will result in substantial increases to the Measured and Indicated resources in the

sulphide zones of these deposits.

Continued exploration of the Satellite Deposits has resulted in the addition of 2.5 Million tonnes @

1.78 g/t to the SAGP Measured and Indicated resources. Newly identified mineralisation in granite at

Bokrobo is likely to substantially increase the resource there.

Although in general the global mineral resources have increased significantly, especially in the Salman

group of deposits, the pit optimisation results indicate that most of the mineral resource increase is

peripheral to the optimal pit limits. The Proven and Probable Mineral Ore Reserves have increased to

12.02 Million tonnes. with a slight reduction in the resource grade within the optimal pit limits.

Compared to the feasibility study results, the operating cost per ounce is $130/oz higher for the

optimal pit shells.




21.0 RECOMMENDATIONS

Exploration and development activities in 2008 should focus on further enhancing the economics of

the Project by endeavouring to add additional ore reserves and reducing, where possible, the

estimated capital costs, by further focused exploration and the examination of alternative plant optionswhere possible

In the event of a decision to mine, the recommended development methodology for the design and

construction management of the Project is the EPCM approach, thus allowing ARL to maintain control

of the budget, schedule and quality of the end product through all stages of project development. The

project capital cost estimate has been developed on the basis that a single organisation (the Engineer)

will provide the EPCM services with the assistance of specialist sub-consultants as required.

Detailed planning of the roads and other site layout will be required before the implementation stage of

the Project.




22.0 REFERENCES AND BIBLIOGRAPHY

• ACA Howe International Limited, 1996. Review of 1995 – 1996 Exploration Programme,

Teleku Bokazo Property Ghana. Unpublished report for Semafo Ghana, held by AdamusResources Ltd.

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unpublished report held by Adamus Resources Ltd.

• Adamus Resource Limited, 2007. Southern Ashanti Gold Project. Feasibility Study Report,

unpublished report held by Adamus Resources Ltd.

• African Gold Properties Ltd, 2003. Ankobra Prospecting Licence. Annual Report.

Unpublished report for the Minerals Commission.

• African Gold Properties Ltd, 2003. Salman Prospecting Licence. Exploration Report for the

2001 – 2003 Period. Unpublished report for the Minerals Commission.

•

African Gold Properties Ltd, 2003. Salman East Prospecting Licence. Annual Report.Unpublished report for the Minerals Commission.

• African Gold Properties Ltd, 2003. Tumentu Prospecting Licence. Annual Report.


• AGR Limited, 2000. Information Memorandum on the Salman Gold Project, Ghana.

Unpublished internal AGR report held by Adamus Resources Ltd.

• Allibone, A., McCuaig, T. C., Harris, D., Etheridge, M., Munroe, S., Byrne, D., Amanor, J.,

Gyapong, W., 2002a. Structural Controls on Gold Mineralisation at the Ashanti Deposit,

Obuasi, Ghana. Society of Economic Geologists, Special Publication 9, p. 65-93.

• Allibone, A., Teasdale, J., Cameron, G., Etheridge, M., Uttley, P., Soboh, A., Apiah-Kubi, J.,

Adanu, A., Arthur, R., Mamphey, J., Odoom, B., Zuta, J., Tsikata, A., Pataye, F., and Famiyeh,

S., Lamb., E., 2002b. Timing and Structural Controls on Gold Mineralisation at the BogosoGold Mine, Ghana, West Africa. Economic Geology, 97, p. 949-969.

• Amegashie, A. Agbeli, 1995. Salman Phase I Drill Report. Unpublished internal BHP

memorandum held by Adamus Resources Ltd.

• Amegashie, A. A., 1998. Terminal Report on Ankobra Gold Prospect, South Ashanti Project,

Ghana. Unpublished report for the Minerals Commission.

• Amegashie, A. A., 1998. Terminal Report on Tumentu Gold Prospect, South Ashanti Project,

Ghana. Unpublished report for the Minerals Commission.

• Amegashie, Augustus, 1997. Report on Fiscal Years 1996-97 Exploration Activities on South

Ashanti Properties, Western Region, Ghana. Unpublished report for the Minerals

Commission.

• Ammtec Limited, 2003. Optimal grind size and gravity-leach evaluation of 8 gold ore

composites from the Salman Gold Project for Adamus Resources Ltd. Report No.A8729, July

2003. Unpublished report held by Adamus Resources Ltd.

• Ashanti Goldfields Ltd, 2000. MSD 1088, Metallurgical evaluation of Anwia Prospect samples.

Unpublished internal report held by Adamus Resources Ltd.

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report for the Minerals Commission.

• Bertram, Dean A., 2001. Salman East Gold Project, Terminal Report, AGR Limited.





• Bloomer, Tony, 1997. Teleku Bokazo Project - Specific Gravity determinations on diamond

drill cores. Unpublished report for Semafo held by Adamus Resources Ltd.

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in West Africa at 2.1 Ga. Journal Geophysical Research, 97, p. 345-369.

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the Salman, Fidelity, and parts of the Tumentu and Ankobra Prospecting Licences, Western

Region, Ghana. Unpublished internal BHP report held by Adamus Resources Ltd.

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Prospecting Licence, Western Region, Ghana (December 1994-July 1996). Unpublished BHP


• Brabham, 2006. Southern Ashanti Gold Project, West Africa – February 2006 Resource

Report. Company report filed to SEDAR in March 2006.

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– February 2003 RC Drilling Program. Unpublished internal report held by Adamus

Resources Ltd.

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Africa. Precambrian Research, 67, p. 89-107.

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Corporation. Unpublished internal GNMC report held by Adamus Resources Ltd.

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Birimian and Tarkwaian rocks of southwest Ghana, West Africa. Journal of African Earth

Science, 14/3, p. 313-325.

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Feasibility Purposes

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mid 1950s. Available in the library at the Geological Survey, Accra, Ghana.

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Prepared on behalf of the Minerals Commission, c/o P.O. Box M248, Cantonments, Accra,

Ghana, 431p.

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January 2008 Resource Update. Unpublished report held by Adamus Resources Ltd.

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Avrebo and Nfutu Gold Deposits, June 2008. Unpublished report held by Adamus Resources

Ltd.

• Hirdes, W., Davis, D. W., and B. N. Eisenlohr, 1992. Reassessment of Proterozoic granitoid

ages in Ghana on the basis of U/Pb zircon and monazite dating. Precambian Research, 56, p.

89-96.

• Isaaks, E.H. & Srivastava, R.M., 1989. An Introduction to Applied Geostatistics. Oxford

University Press, New York.

• John, T., Klemd, R., Hirdes, W., and Loh, G., 1999. The metamorphic evolution of the

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southern ashanti gold project ghana, west africa (21 august 2008)

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