EX-99.1 3 bajatechrep.htm UPDATED PRELIMINARY ECONOMIC ASSESSMENT ON THE BOLEO PROPERTY Exhibit 99.1 Updated Preliminary Economic Assessment on the Boleo Property



BAJA MINING CORP.



AN UPDATED PRELIMINARY ASSESSMENT OF EL BOLEO COPPER COBALT PROJECT



BAJA CALIFORNIA SOUTH,

MEXICO


Prepared for


Baja Mining Corp.


by


W.J.A. Yeo

MAusIMM

J. Wyche

MAusIMM, MMICA, CPEng

M. Holmes

MSAIMM, PEng

E. Norton

PEng

D. Hunter

FAusIMM, MIOM3, CPEng, CEng

T. Ross

CPEng

S.Britton

PEng




31 January 2007







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EL BOLEO PROJECT

MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT





TABLE OF CONTENTS


 

1     SUMMARY

1

1.1     GEOLOGY AND MINTERAL RESOURCE MODELLING

1

2     INTRODUCTION

7

3     RELIANCE ON OTHER EXPERTS

8

4     PROPERTY DESCRIPTION AND LOCATION

9

4.1     LOCATION

9

4.2     DESCRIPTION

10

4.3     OWNERSHIP

12

4.4     TAXES AND ASSESSMENT WORK REQUIREMENTS

12

4.4.1     TAXES

12

4.4.2     WORK REQUIREMENTS

12

4.4.3     OPTION PAYMENTS.

12

4.5     PERMITS AND LIABILITIES

12

4.5.1     PERMITS

12

4.5.2     LIABILITIES

13

5     ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

14

5.1     ACCESS

14

5.2     CLIMATE

14

5.3     LOCAL RESOURCES

14

5.4     INFRASTRUCTURE

15

5.5     PHYSIOGRAPHY

15

6     HISTORY

16

7     GEOLOGICAL SETTING

19

7.1     REGIONAL GEOLOGY

19

7.2     PROPERTY GEOLOGY

20

7.3     STRUCTURAL GEOLOGY

22

8     DEPOSIT TYPES

24

9     MINERALIZATION

25

9.1     BOLEO DISTRICT FALSE BOTTOMS

27

10   EXPLORATION

33

11   DRILLING

34

11.1   GENERAL

34

11.2   HISTORICAL

34

11.3   2004 – 2007

34

12   SAMPLING METHODS AND APPROACH

48

12.1   CORE RECOVERY

52

12.2   DENSITY DETERMINATIONS

53



 

 

 

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MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT




12.2.1     DIMENSIONAL (OR CALLIPER) METHOD

54

12.2.2     IMMERSION METHOD

54

12.2.3     RESULTS

54

13   SAMPLE PREPARATION, ANALYSIS AND SECURITY

56

13.1   HISTORICAL DRILL HOLES – PRE 2004

56

13.2   CURRENT DRILL HOLES

56

14   DATA VERIFICATION

58

14.1   HISTORICAL DATA – PRE 2004

58

14.1.1     PROBLEMS WITH THE SGS-XRAL ASSAY RESULTS

58

14.1.2     SUMMARY OF 1997 PROCESS REVIEW AND RE-ASSAYING PROGRAM

58

14.1.3     COMMENTS ON ASSAY QUALITY CONTROL RESULTS

63

14.1.4     DATABASE VERIFICATION

65

14.1.5     DRILL HOLE SURVERYING

66

14.2   CURRENT PROGRAM

66

14.2.1     ASSAY STANDARDS

66

14.2.2     CHECK ASSAYS

68

14.2.3     DUPLICATE ASSAYS

69

14.3   PROJECT DATABSE

70

14.3.1     DATA

70

14.3.2     DRILL HOLE SURVEYING

71

15   MINERAL PROCESSING AND METALLURGICAL TESTING

72

15.2   METALLURGY

72

15.1.1     BACKGROUND

72

15.1.2     PROOF OF CONCEPT PILOT CAMPAIGN

72

15.1.3     FULLY INTEGRATED PILOT CAMPAIGN

78

15.2   PROCESS PLANT DESIGN

99

15.2.1     INTRODUCTION

99

15.2.2     PROCESS PLANT DESCRIPTION

102

16   MINERAL RESOURCE AND MINTERAL RESERVE ESTIMATES

107

16.1   3D GEOLOGICAL INTERPRETATION

107

16.1.1     FAU;T SURFACES

107

16.1.2     MANTO FOOTWALL SURFACES

109

16.2   RESOURCE ESTIMATION TECHNIQUE

113

16.2.1     BLOCK MODEL VS. GRIDDED SEAM MODEL

113

16.2.2     DATA DOMAINS AND FLAT MODELS

114

16.3   3D BLOCK MODELS

116

16.3.1     COMPOSITE LENGTH AND BLOCK DIMENSIONS

116

16.3.2     UNIVERIATE STATISTICS OF DATA COMPOSITES

117

16.3.3     SPATIAL CONTINUITY OF GRADE

118

16.3.4     SEARCJ [ARA,ETERS AMD DATA CROTEROA

123

16.3.5     ,PDE; CPDOMG

124

16.3.6     RESOURCE CLASSIFICATION – 3D BLOCK MODELS

125

16.3.7     RESOURCE ESTIMATES – 3D BLOCK MODELS

127

16.3.8     MODEL VERIFICATION

129



 

 

 

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MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT




16.3.9     COMPARISON WITH 2005 3D BLOCK MODELS

135

16.4   SEAM MODELS

138

16.4.1     COMPOSITE LENGTH AND BLOCK DIMENSIONS

138

16.4.2     UNIVARIATE STATISTICS OF DATA COMPOSITES

140

16.4.3     SPATIAL CONTINUITY OF GRADE

141

16.4.4     SEARCH PARAMETERS AND DATA CRITERIA

143

16.4.5     RESOURCE CLASSIFICATION – SEAM MODELS

143

16.4.6     RESOURCE ESTIMATES – SEAM MODELS

144

16.4.7     MODEL VERIFICATION

145

16.5   TOTAL RESOURCE

151

16.6   MINERAL RESERVE ESTIMATES

153

16.6.1     REPORTING STATUS

153

17   OTHER RELEVANT DATA AND INFORMATION

154

17.1   MINING AND MINE DESIGN

154

17.1.1     OPEN CUT MINING

154

17.1.2     COPPER PITS

154

17.1.3     LIMESTONE QUARRY

156

17.1.4     OPEN CUT DEVELOPMENT SEQUENCE

156

17.1.5     OPEN CUT MINING COSTS

159

17.2   UNDERGROUND MINING

159

18   PRELIMINARY ECONOMIC ASSESSMENT

164

18.1.1     SUMMARY

164

18.1.2     PRODUCT MARKETING

166

18.1.3     CAPITAL COST ESTIMATE

167

18.1.4     OPERATING COSTS

168

18.1.5     PRELIMINARY ECONOMIC ASSESSMENT

169

19   CONCLUSIONS

182

19.1   GEOLOGT AND MINERAL RESOURCE MODELLING

182

19.2   METALLURGY AND PROCESS DESIGN

182

19.3   MINING

183

19.4   ENVIRONMENTAL

183

19.5   PRELIMINARY ECONOMIC ASSESSMENT

183

20   RECOMMENDATIONS

184

20.1   GEOLOGY AND MINERAL RESOURCE MODELLING

184

20.2   METALLURGY AND PROCESS DESIGN

185

20.3   MINING

185

20.4   UPDATED PRELIMINARY ECONOMIC ASSESSMENT

186

21   REFERENCES

187



 

 

 

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MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT




LIST OF FIGURES


 

Figure 1:

El Boléo Location Map

9

Figure 2:

El Boléo Property Map

11

Figure 3:

El Boléo Geological Setting

19

Figure 4:

Boléo Formation Stratigraphic Column

21

Figure 5:

Geological Cross-Sections

26

Figure 6:

Manto 1 Footwall Geology

28

Figure 7:

Manto 1 Hole DDH955 – False Bottom Geology

29

Figure 8:

Manto 3 Location of False Bottoms

31

Figure 9:

Drill Hole Locations –Historical & Infill

35

Figure 10:

Mechanical Core Splitter

49

Figure 11:

Prototype Mechanical Splitter

50

Figure 12:

Final Mechanical Splitter Design

51

Figure 13:

Core Recovery vs. Copper Grade- Current Drill Program

53

Figure 14:

Bulk Density vs. Sample Depth

55

Figure 15:

Boléo Cu, Round-Robin Cu Assays used by Peatfield & Smee

64

Figure 16:

Boléo Cu, Round-Robin Cu Assays used by H&S

64

Figure 17:

Boléo Assay Standards – Cu

67

Figure 18:

Boléo Assay Standards – Co

67

Figure 19:

Boléo Assay Standards – Zn

67

Figure 20:

Boléo Check Assay Results – Cu

68

Figure 21:

Boléo Check Assay Results – Co

69

Figure 22:

Boléo Duplicate Assay Results – Cu

69

Figure 23:

Boléo Duplicate Assay Results – Co

70

Figure 24:

Boléo Metal Extractions

91

Figure 25:

Boléo Pilot Plant CCD Thickener Mud Levels

93

Figure 26:

Manganese Loading On DSX Organic Phase

97

Figure 27:

Zinc in Solution vs. DSX Operating Time

97

Figure 28:

Cobalt in Solution vs. DSX Operating Time

98

Figure 29:

Schematic Flow Diagram Sheet 1

100

Figure 30:

Schematic Flow Diagram Sheet 2

101

Figure 31:

Schematic Flow Diagram Sheet 3

102

Figure 32:

Fault Framework in Section – Detail of Cross-Sections (900N) as Supplied by MMB

108

Figure 33:

Digitised Fault Traces – Faults Digitized (dashed red lines) Following Cross-Sectional Interpretation (900N)

108

Figure 34:

Correlation of Fault Traces

108

Figure 35:

Final 3D Fault Framework with Mintec Cross-Sections

109

Figure 36:

Triangulated Manto Surface – (Manto 3, oblique 60o)

110

Figure 37:

Triangulated Manto Surface Sliced Along Section Lines – (Manto 3, oblique 60o)

110

Figure 38:

Snapping Footwall Surface to Faults

111

Figure 39:

Re-Triangulation of Manto Surface Components

111

Figure 40:

Re-Triangulation of Manto Surface – Detail of Final Result

112

Figure 41:

Re-Triangulation of Manto Surface – Full View (Manto 3, Oblique 60o)

112

Figure 42:

Re-Triangulation of Manto Surface – Full View with Faults (Manto 3, Oblique 60o)

113

Figure 43:

Search Ellipses – Plan & Section Views

115

Figure 44:

Manto 3 Variogram Maps

119

Figure 45:

Variograms of Copper – Manto 3 (X, Y, Z directions)

120

Figure 46:

Variograms of Cobalt – Manto 3 (X, Y, Z directions)

121

Figure 47:

Variograms of Zinc – Manto 3 (X, Y, Z directions)

122

Figure 48:

Manto 3 Old Mine Areas – Location Drill Holes Which Encountered Voids

126

Figure 49:

Manto 3 Copper – Block Model with Assay Composites

129

Figure 50:

Detail of Manto 3 Copper – Block Model with Assay Composites

130




 

 

 

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MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT




Figure 51:

Manto 3 Cobalt – Block Model with Assay Composites

131

Figure 52:

Detail of Manto 3 Cobalt – Block Model with Assay Composites

132

Figure 53:

Manto 3 Zinc – Block Model with Assay Composites

133

Figure 54:

Detail of Manto 3 Zinc – Block Model with Assay Composites

134

Figure 55:

Manto 3 Resource Classification – Block Model with Drill Hole Locations

135

Figure 56:

Variograms of Copper – Manto 3 (X, Y, directions)

141

Figure 57:

Variograms of Cobalt – Manto 3 (X, Y, directions)

142

Figure 58:

Variograms of Zinc – Manto 3 (X, Y, directions)

142

Figure 59:

Manto 3 Copper – Block Model with Assay Composites

146

Figure 60:

Detail of Manto 3 Copper

147

Figure 61:

Manto 1 Copper – Block Model with Assay Composites

148

Figure 62:

Detail of Manto 3 Copper

149

Figure 63:

Manto 3 Resource Classification – Block Model with Drill Hole Locations

150

Figure 64:

Manto 3 Resource Classification – Block Model with Drill Hole Locations

151

Figure 65:

Open Cut Pit Locations

158

Figure 66:

Conceptual Mine Layout for a District

162

Figure 67:

Mining Area 3-10 Preliminary Layout

163



LIST OF TABLES


 

Table 1:

Boléo Property, Sole Concessions, January 2007

10

Table 2:

Boléo Project, Surface Property and Annual Taxes

11

Table 3:

Historical Mining Activities at Boléo 17

17

Table 4:

Manto 3 - Vertical Grade Variation Associated With False Bottoms

32

Table 5:

Diamond Drill Hole Locations – Current Programs (DDH928-1082)

36

Table 6:

Diamond Drill Hole Results – Current Programs (DDH928-1082)

37

Table 7:

Diamond Core Recovery – Historical

52

Table 8:

Diamond Core Recovery - Current Programs (DDH928-1082)

53

Table 9:

Bulk Density Summary by Manto

55

Table 10:

Boléo Assay Standards

59

Table 11:

Boléo Core Duplicate Results – 1st Program

62

Table 12:

Boléo Core Duplicate Results – 2nd Program

62

Table 13:

Boléo Formal Standards – Revised Values

64

Table 14:

Drill Holes Audited for Data Entry Errors

65

Table 15:

Comparison of As Received, 200 kg Composite and Scrubber Feed Assays

80

Table 16:

Leach Feed Slurry Assays

80

Table 17:

SGS Minerals Services Analytical Methods Employed

89

Table 18:

Boléo Pilot Plant Timeline

90

Table 19:

Leach and Partial Neutralisation Operating Result

90

Table 20:

Leach Efficiencies

90

Table 21:

CCD Operating Results

92

Table 22:

Vendor Testwork – CCD Paramet

94

Table 23:

Copper Solvent Extraction Solution Assays

94

Table 24:

Copper Solvent Extraction Organic Assays

94

Table 25:

Plating Cycle Times & Operating Conditions

95

Table 26:

Cathode Quality

95

Table 27:

DSX Metals Extractions

97

Table 28:

Manganese Carbonate Product Average Assay Values

98

Table 29:

Summary of Leach and CCD Reagent Consumptions

99

Table 30:

Univariate Statistics of Assay Composites – Copper

117

Table 31:

Univariate Statistics of Assay Composites – Cobalt

118

Table 32:

Univariate Statistics of Assay Composites – Zinc

118




 

 

 

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MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT




Table 33:

Variogram Models

123

Table 34:

Search Parameters – 2005 3D Block Models

123

Table 35:

Search Parameters – Current 3D Block Models

124

Table 36:

Measured and Indicated Resource at 0.5% CuEq. Cutoff

127

Table 37:

Measured and Indicated Resource at 1.0% CuEq. Cutoff

128

Table 38:

Inferred Resource

128

Table 39:

Manto 3 – Current vs. 2005 Resource Model, Measured and Indicated Resources

136

Table 40:

Manto 3 - Current vs. 2005 Resource Model, Inferred Resources

137

Table 41:

Comparison between Current and 2005 Resource Models, Manto 3a

137

Table 42:

Univariate Statistics of Assay Composites – Copper

140

Table 43:

Univariate Statistics of Assay Composites – Cobalt

140

Table 44:

Univariate Statistics of Assay Composites – Zinc

141

Table 45:

Variogram Models – Seam Composites

142

Table 46:

Search Parameters – Current 3D Block Models

143

Table 47:

Seam Models – Measured and Indicated Resource at 0.50% CuEq.

144

Table 48:

Seam Models – Measured and Indicated Resource at 1.0% CuEq.

144

Table 49:

Seam Models - Inferred Resource

145

Table 50:

Combined Resource at 0.50% CuEq.

152

Table 51:

Combined Resource at 1.0% CuEq.

152

Table 52:

Preliminary Economic Assessments – Base-Case Highlights

165

Table 53:

Sensitivity to Key Variables

166

Table 54:

Capital Cost Areas of Responsibility

167

Table 55:

Capital Cost Estimate Summary

167

Table 56:

Summary of Boléo Project Development Timing

170

Table 57:

Boléo Preliminary Economic Assessment – Base Case Summary

173

Table 58:

Boléo Preliminary Economic Assessment – 5 Year Price Case Summary

174

Table 59:

Boléo Preliminary Economic Assessment – Current Price Case Summary

175

Table 60:

Boléo PEA – Opportunity: Base Case with Manganese Production

176

Table 61:

20 Year Detailed Cash Flow – Base Case

177

Table 62:

20 Year Detailed Cash Flow – 5 Year Prices Case – (3 year trailing + 2 year leading)

179

Table 63:

20 Year Detailed Cash Flow – Current Prices Case

181







 

 

 

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MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT




1

SUMMARY

 

 


1.1

GEOLOGY AND MINERAL RESOURCE MODELLING


El Boléo Cu-Co-Zn-Mn deposit is located near the town of Santa Rosalia, Baja California Sur, Mexico.

The deposit, which occurs in the Boléo Formation, comprises seven mineralized units called “Mantos” (manto is a Spanish term used in mining parlance for a generally mineralized layer or stratum).  The mineralized mantos dip gently to the east but faulting, which is common throughout the project area, produces a step-like pattern in the present position of the mantos.

The deposit was actively mined from 1886 to 1972, during which time an estimated 18 Mt of ore were treated.  Modern exploration activities took place during the years 1993 to 1998 and again in 2004 to 2006, at which time the drill hole assay database used in the current Resource Study was acquired.

Independent geologic consultants, Hellman & Schofield (H&S) of Sydney, Australia were engaged in 2004 to oversee the geologic modelling of the Boléo deposit and to produce a resource estimate in accordance with NI 43-101 standards.

Analytical processes and quality control methods have been studied at length.  Matrix matched assay standards were developed for the project and round-robin assay trials used to establish the best analytical method.  The assays were deemed accurate and precision values were obtained from original samples versus diamond core duplicates indicating they were representative of the drill hole intervals.

A gridded block model for each Manto was developed for open pit mine design in preference to a gridded seam model because it allows greater versatility to evaluate mining methods at various cutoff grades.  Block dimensions used were 50 m EW x 100 m NS and 1 m vertical.

Gridded seam models have been developed for Mantos 1, 2, 3 and 4 for underground mine planning.  These models incorporate minimum and maximum mining heights.

A 3D digital geological interpretation was developed from data and information supplied by the company.  This incorporated both Manto surfaces and faults.

Statistical analysis of the assay data from each Manto showed that the histograms of Cu, Co, Zn and Mn are not highly skewed, indicating that ordinary kriging is an appropriate estimation method.

To account for the displacement of mantos by the numerous faults, block models and drill hole intersections for each Manto were re-aligned at the same height so the data was effectively re-positioned approximating pre-faulting locations.  This removed the need to have a complex series of fault bounded data domains and significantly streamlined the resource estimation



 

 

 

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UPDATED PRELIMINARY ECONOMIC ASSESSMENT



process.  These block models are referred to as ‘flat’ models.  Block centroid heights in real space, taken from gridded surfaces of the Manto footwalls, were used to translate flat model grade estimates into true 3D blocks.

Resource classification was determined by the number of composite data available for grade estimation, from increasingly localised data search regimes.

For 3D Block and Seam models:

·

Measured – search 200 m x 250 m x 2 m

·

Indicated – search 400 m x 500 m x 2 m

·

Inferred – search 400 m x 500 m x 4 m.


For 3D Block models:

·

Measured – minimum data 14 composites from 4 octants

·

Indicated – minimum data 14 composites from 4 octants

·

Inferred – minimum data 7 composites from 2 octants.


For Seam models:

·

Measured – minimum data 4 composites from 4 octants

·

Indicated – minimum data 4 composites from 4 octants

·

Inferred – minimum data 2 composites from 2 octants.


The total reported Measured and Indicated Resource, based on copper equivalent cutoff grades of 0.5 and 1.0% Cu are as follows:

CuEq. Cutoff Grade

0.5%

1.0%

Measured

Tonnes (106)

59.4

53.2

 

CuEq.%

2.15

2.30

 

Cu%

0.86

0.94

 

Co%

0.088

0.091

 

Zn%

0.46

0.48

 

Mn%

2.77

2.87

Indicated

Tonnes (106)

173.4

128.3

 

CuEq. %

1.72

2.05

 

Cu%

0.76

0.96

 

Co%

0.055

0.064

 

Zn%

0.54

0.60

 

Mn%

2.74

3.06

Total

Tonnes (106)

232.8

181.5

 

CuEq. %

1.83

2.13

 

Cu%

0.79

0.96

 

Co%

0.064

0.072

 

Zn%

0.52

0.56

 

Mn%

2.75

3.00



 

 

 

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MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT



The additional Inferred Resource is:

CuEq.. Cutoff Grade

0.5%

1.0%

Inferred

Tonnes (106)

202.6

114.3

 

CuEq. %

1.32

1.76

 

Cu%

0.46

0.66

 

Co%

0.043

0.055

 

Zn%

0.65

0.88

 

Mn%

2.67

3.38




A model was created to account for past work in the historic mining areas of Mantos 1 and 3.  Tonnes were factored down by 12% to account for material extracted and processed.

METALLURGY AND PROCESS DESIGN


A 'proof of concept' pilot plant campaign was conducted at the SGS-Lakefield Testing Laboratories, Lakefield, Ontario, Canada from November 16th to 28th, 2004 treating a bulk oxide sample of underground Boléo ore grading 1.6% Cu, 0.087 % Co, 0.58% Zn, 3.23% Mn and 8.71% Fe.  At this time various other bench scale test work was conducted including settling rate determination, slurry viscosity work, tailings characterisation test work (for environmental permitting purposes) and bench scale leach testing.  

The major focus of the proof of principle pilot plant was to confirm 1) the clay based Boléo ores could be thickened and washed in a conventional CCD train using high rate thickeners and 2) the CSIRO “DSX” solvent extraction system could be used to recover a cobalt and zinc product.  

Since the initial proof of concept pilot plant tests, further work was undertaken to 1) optimize the DSX solvent extraction reagent composition, 2) incorporate and test a manganese recovery process, 3) test the suitability of soda ash for pH adjustment in the circuit and 4) utilize local Boléo carbonate for bulk neutralization.    

The integrated demonstration pilot plant campaign to verify these changes was run from June 5th to 24th, 2006 on a bulk oxide sample of underground Boléo ore grading 2.18% Cu, 0.135 % Co, 0.491% Zn, 5.0% Mn and 8.26% Fe.

 The zinc solvent extraction and cobalt solvent extraction and electrowinning circuits were run in a separate campaign on accumulated solutions during the period of July 4th to 15th, 2006 thereby allowing the SGS, Bateman and client teams to initially focus on the processes that were either new to the flowsheet or modified since the original proof of concept pilot campaign.

Once again extensive bench scale test work was conducted including:

·

settling rate determination

·

agitator power estimation in the leach and iron removal circuits

·

tailings slurry viscosity and pumping test work



 

 

 

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UPDATED PRELIMINARY ECONOMIC ASSESSMENT



·

bulk materials handling test work

·

filtration testing of iron residues and manganese carbonate product

·

tailings characterisation test work for environmental permitting purposes.


The results of the test work campaign have been employed to generate plant design data for the purposes of advancing the definitive feasibility study for a planned greenfield development consisting of open pit and underground mines, a processing plant and all associated facilities and infrastructure necessary to produce copper and cobalt metal and a zinc sulphate monohydrate salt on site. Manganese carbonate production as an additional product was also considered.

The process plant is being designed to produce and treat 2.6 Mdmt/a of plant feed at an average head grade 2.2% Cu, 0.1% Co, 0.6% Zn and 2.2% Mn through an integrated hydrometallurgical facility to produce the following products and tonnages:

·

Up to 60,000 t/a of copper cathode

·

Up to 3,100 t/a of cobalt cathode

·

Up to 36,000 t/a of zinc sulphate monohydrate salt.


In addition, the potential to produce up to 100,000 t/a of a manganese carbonate product has been reviewed and is discussed in Section 17 of this report.

It is intended to “de-bottleneck” the plant operations at production Year 5 to ensure the product production levels remain at the target values as ROM head grade begins to decrease.  Debottlenecking of the production facility will require a modest capital injection at the appropriate time.  Capital costs have been allowed by Baja Mining Corp for this eventuality in the financial analysis of all of the base case scenarios.

MINING


The potential for surface and underground mining were assessed for the H&S resource model.  A series of underground mining operations, supported by small open-cut surface mines, delivering targeted high-grade copper-cobalt-zinc manganese ore (0.5% to 2.5% Cu) to the process facility was selected as the best alternative for the scale of operation envisaged by Baja Mining Corp.

The seam formation and low material strength of the mantos suggested conventional “soft rock” mining methods such as used in coal, potash, or salt mining would be successful.  Common methods were examined and considered for their application to the deposit.  Longwall mining was discounted due to the faulted and dipping manto structure and high initial capital cost.  Shortwall mining was also discounted when efficient mine layouts could not be readily designed due to the manto structure.  Room-and-pillar mining using continuous miners was chosen because of the methods’ flexibility for layout designs, its’ efficient recovery of resources and lower initial capital cost. The resource seam model was used to define manto areas that could be economically mined.  The basic criteria used were:



 

 

 

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·

Minimum mining height of 1.8 m to allow working room for the machines.  If the economic thickness of the manto was less than this it was diluted by the lower grade blocks above up to 1.8m height.

·

Maximum mining height of 4.2 m, matching the reach of the continuous miner.  Economic blocks above this height were ignored.

·

The composite copper equivalent grade of the manto over the mining height must exceed a cutoff grade of 0.5% Cu.  For mining purposes, the copper equivalent cutoff grade was calculated based upon base case metal prices.

·

An allowance for voids in old works and recovery of “retaque”, or previously mined material, was made in terms of both recovery and density.


Underground mining trials to test equipment, working methods and geotechnical ground responses to the chosen method were undertaken in two stages in the years 2005-2006 under the supervision of consultants, Australian Mine Design & Development (AMDAD), Sydney, Australia, and Agapito Associates, Inc. (AAI), Grand Junction, Colorado, USA.  These tests were very successful and confirmed the suitability of continuous mining methods to the Boléo deposit.

Initial mining plans for both underground and surface mining targets in all mantos have been advanced by AAI and AMDAD to provide ore feed at targeted production levels and head grades based on process plant schedules.  A limestone source, located on the Boléo property, has also been modelled as the source of calcium carbonate needed for process plant operations.

Preliminary Economic Assessment:

A financial model was created utilizing the current mine production schedule over an initial 20 years, the associated diluted metal grades based on the H&S geological resource and mining schedule, capital and operating costs as set out herein and base case metal prices of copper US$1.25/lb, cobalt US$12.00/lb and zinc sulphate US$950/mt.

In addition, sensitivity analysis was also conducted at various increased metal prices.  The project is sensitive to four key variables; copper price; cobalt price; capital costs and operating costs.  The sensitivity of the After-Tax IRR and NPV (at 8% discount rate) relative to the Base Case is shown in the table below to indicate the effect of plus or minus 10% changes in the key variables.  Note that the changes to the copper price apply to all of the annual prices, starting in 2009, and not just the long term price.

 

After Tax IRR

 

After Tax NPV at 8% ($ Millions)

 

-10%

Base Case

+10%

 

-10%

Base Case

+10%

Copper Price

16.0%

19.0%

21.7%

 

$236

$333

$420

Cobalt Price

18.1%

19.0%

19.9%

 

$298

$333

$367

Capital Cost

21.5%

19.0%

16.6%

 

$375

$333

$283

Operating Cost

20.7%

19.0%

17.1%

 

$396

$333

$266



 

 

 

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The modelling, based on the current un-optimised preliminary mine schedule, indicates that the project is financially attractive at base-case metal prices. Financial modelling, using the base case prices and only 20 years for the project life, shows that the project could generate a net after- tax Internal Rate of Return (IRR) of 19.0% with a discounted present value, at an 8% discount rate, of US$333.4 million.  Using a 6% discount rate generates an NPV, after tax, of $US445.9 million.



 

 

 

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2

INTRODUCTION

 

 


The Boléo deposit was actively mined from about 1888 to 1972.  During that time approximately 18 Mt of ore were mined and processed.  More recently, an intensive phase of exploration drilling was carried out between 1993 and 1998, by International Curator Resources Ltd (ICR), focused on defining a large resource amenable to open-cut mining.  The project was not developed at that time and in 2001 ownership of the project passed to Minera y Metalurgica del Boléo S.A de C.V. (MMB).  Various resource studies and a second intensive phase of exploration and infill drilling between 2004 and 2006 has been completed by MMB with the program focus being to define higher grade resources that could be exploited by underground mining methods.

The current study was commissioned to:

·

Produce a digital geological model, including a topographic surface that honours the faults, tops and bases of the seven mantos using a set of interpreted cross sections.

·

Produce a digital grade model, for all mantos, for Cu, Co, Zn, and Mn.

·

Produce an underground void model.

·

Provide a Confidence Classification, taking into account drill spacing, structural and other controls on mineralization, suitable for quotation to NI43-101 standards.

·

Review the QA/QC data as relevant to sampling, assaying precision and accuracy.

·

Assess the methodology for determining density and the adequacy of the density database, leading to the production of a density model for conversion of volumes to tonnages.


MMB supplied digital drill hole assay and geology data, geological interpretation, drill hole recovery data and rock density data.  Information and data regarding assay and sampling quality control was sourced from various reports by consultants G. Peatfield, B. Smee and D. Mehner (Peatfield and Smee 1997, Peatfield 1997, Peatfield 1998, Mehner 2003).

Drilling, sampling and assaying activities providing a significant part of the data used in the current resource study were completed between 1993 to 1998.  H&S, whose involvement with the project commenced in August 2004, were, therefore, unable to observe any drilling, sampling or assaying activities related to the drilling carried out during the 1990s. Three infill drill programs were carried out by Minera y Metalurgica del Boléo, S.A. de C.V., the first from December 6, 2004 to January 29 2005 (DDHs 928 to 941), the second program from May 11, 2005 to July 3, 2005 (DDHs 942 to 959), and the third program from February 1, 2006 to September 15, 2006 (DDHs 960 to 1082). These programs were observed and monitored by H&S's representative Dr. B Yeo.



 

 

 

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3

RELIANCE ON OTHER EXPERTS

 

 


The capital cost for development of the Boléo Project has been developed by a number of specialist organizations.  These organizations are listed below in a table that summarises areas of significant capital cost and the organizations responsible for development of capital costs for these respective areas.  The Capital Cost Estimate for the project development has been co-ordinated and integrated by Wardrop Engineering on behalf of Bateman Engineering Canada Corp.

Major Cost Area

Consultant

Location

Open Pit Mining

AMDAD Pty Ltd

Sydney, Australia

Underground Mining

Agapito Associates, Inc

Golden, Colorado

Mining Surface Infrastructure

Wardrop Engineering

Vancouver, Canada

Process Plant and General Infrastructure

Wardrop Engineering

Vancouver, Canada

Tailings Dam

Arcadis Geotechnica

Santiago, Chile

Co-Generation Plant

Fransen Engineering Ltd

Vancouver, Canada

Acid Plant

Fenco Pty Ltd

Toronto, Canada

SO2 Gas Production Facility

Noram Engineering & Constructors Ltd

Vancouver, Canada

Barging Facility

ATI

Vancouver, Canada

Liquid Sulphur Infrastructure

ICEC Canada Ltd

Calgary, Canada

Mexican Construction Labour Rates

UHDE Jacobs

Mexico City


With the exception of AMDAD Pty Ltd, Agapito Associates and Wardrop Engineering the consultants listed above cannot be classified as qualified persons for the purposes of this report.  

Bateman Engineering Canada Corp & Wardrop Engineering have relied on the consultants listed above for the generation of capital cost estimates in their particular areas of expertise.  Neither Bateman Engineering Canada Corp nor Wardrop Engineering have attempted to formally verify the accuracy or sufficiency of the cost estimates provided by these consultants.  

The portion of the report to which the above disclaimer applies is to Section 18.1.3, Capital Cost Estimate.



 

 

 

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4

PROPERTY DESCRIPTION AND LOCATION1

 

 


4.1

LOCATION


The Boléo project is located along the east coast of the Baja peninsula centred on the port town of Santa Rosalia in Baja California Sur, Mexico (Figure 1). The town is approximately 850 km. south of San Diego, California, USA.  Coordinates for the project are Latitude 27°14´ to 27°25´ N, Longitude 112°14´ to 112°22´ W.

Figure 1:

El Boléo Location Map

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1 Information based on Mehner 2003, reviewed and updated by T. Albinson (MMB) Nov 2004.  Maps supplied by MMB.




 

 

 

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4.2

DESCRIPTION


The Boléo property consists of 18 total mineral concessions covering 19,519.1872 ha, of which 17 concessions are contiguous. The “San Bruno” concession is not contiguous and is located 30 KM south of Sta. Rosalia in the San Bruno basin area. The titled concessions are listed in Table 1 and shown in Figure 2. One concession is in the process of becoming titled (“San Luciano 5”claim).  It should be noted as of January 1st, 2006, all claims in Mexico are “Concesiones Unicas” (Sole Concessions) and the older classification of  Exploration claims evolving after 6 years to Exploitation claims is no longer applicable or in use.

Table 1:

Boléo Property, Sole Concessions, January 2007

Claim

Title No.

Surface Area
(ha)

Date
Initiated

Expiry
Date

Annual
Taxes

El Boléo

218082

4,975.6132

Sept. 29-2000

Sept. 28-2050

285,004

El Boléo I

218092

72.4463

Aug. 31-2000

Aug. 30-2050

4,150

El Boléo19, 256.1872 II fracc. Uno

218179

1,296.6156

Sept. 29-2000

Sept. 28-2050

74,270

El Boléo II fracc. Uno A

218180

507.2841

Sept. 29-2000

Sept. 28-2050

29,058

Boléo III

212148

224.6410

Aug. 31-2000

Aug. 30-2050

12,868

Nuevo San Luciano

214189

150.0000

Aug. 10-2001

Aug. 9-2051

4,272

Boléo II fracc. IV

218975

267.1579

Jan. 28-2003

Jan. 27-2053

7,608

Boléo X fracc. 5

211055

1.3829

Mar. 24-2000

Mar. 23-2050

80

Boléo X fracc. 8

211058

3.9486

Mar. 24-2000

Mar. 23-2050

226

Boléo X fracc. 9

211059

9.9612

Mar. 24-2000

Mar. 23-2050

570

Boléo X  fracc 12

211062

3.1241

Mar. 24-2000

Mar. 23-2050

178

Boléo X  fracc 16

211066

0.0068

Mar. 24-2000

Mar. 23-2050

2

Biarritz B

219819

0.0055

April 16-2003

April 15-2053

2

San Luciano 2

220740

670.0000

Sept. 30-2003

Sept. 29-2053

9,220

San Luciano 3

221073

1,899.0000

Nov. 19-2003

Nov.18-2053

26,130

San Bruno

222772

8,783.0000

Aug. 27-2004

Aug. 26-2054

120,854

San Luciano 4

223358

392.0000

Dic. 3-2004

Dic.2-2054

5,394

San Luciano 5

E-429

263.0000

 

 

 

Total (Pesos)

 

19,519.1872

 

 

579,886

Total (US$)

 

 

 

 

53,635

Note:

the exchange rate used is Pesos 10.8116 =US $1

The project also includes three surface lots totalling 6,692.58 ha as shown in Table 2 and Figure 2.



 

 

 

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Table 2:

Boléo Project, Surface Property and Annual Taxes

Surface Lot

Size
(ha)

Annual
Taxes

El Boléo

6,553.55

21,604

Soledad Property

99.91

23,786

San Luciano Property

39.12

5,301

Total (Pesos)

6692.58

50,691

Total (US$)

 

4,688


Figure 2:

El Boléo Property Map

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4.3

OWNERSHIP


The mineral concessions covering El Boléo copper-cobalt-zinc manganese deposit are 100% owned by Minera y Metalurgica del Boléo S.A. de C.V. (MMB), a Mexican company involved in mineral exploration and development and a wholly owned subsidiary of Baja Mining Corp, who recently listed on the Toronto Stock Exchange (TSX).


4.4

TAXES AND ASSESSMENT WORK REQUIREMENTS


4.4.1

TAXES


Total annual fees payable in January 2007 as of this report are 579,886 pesos for mineral concessions and 50,679 pesos for surface leases, or using the exchange rate as of January 02, 2007 (10.8116 pesos/US$), US$58,323 (see Tables 1 and 2). The calculated annual fees are based on the latest published government tax guides.

4.4.2

WORK REQUIRMENTS


Work obligations on the property (known in Mexico as “Informes de Conprobaciones de Obras”) are in good standing. Based on past work expenditures of approximately US$22 million, enough credits have been accrued to keep the property in good standing until 2013.  

4.4.3

OPTION PAYMENTS


There are no royalties payables on the properties and there are no other agreements or encumbrances.

4.5

PERMITS AND LIABLITIES


4.5.1

PERMITS


During 2006, MMB successfully completed a full Environmental Impact Assessment that covers the construction, operation and closure phases of the Boléo project.  Given the complexities of the project itself and the environmental sensitivity surrounding the project location, the Mexican Federal environmental agency, Secretaria de Medio Ambiente y Recursos Naturales, (SEMARNAT) requested the submittal of an Environmental Impact Manifest with a regional scope.  The change of scope required additional field-work to fully characterize the regional area of influence of the project.  This, in turn, caused a delay in the EIM submission date from February to May 2006.  The evaluation process also included a request from SEMARNAT to submit additional information relating to the project to better clarify the identified environmental impacts.  This request was given to MMB on July 5th, 2006.  The information was formally filed on October 2nd, 2006.

Finally, after incorporating the observations and recommendations from the National Commission for Natural Protected Areas; the Secretariat for Urban Planning, Infrastructure and Ecology of the State Government of Baja California Sur and the Municipal Presidency of Mulegé



 

 

 

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at Baja California Sur, the environmental impact resolution was issued on November 27th, 2006 and delivered to MMB on December 7th, 2006.  This resolution authorizes the construction, operation and closure of El Boléo Mining Project.  The official document number containing this resolution is S.G.P.A.-DGIRA.-DDT.-2395.06 and is signed by the General Director for Environmental Impact and Risk (DGIRA).

This authorization allows MMB to initiate the procedures to obtain more specific permits.  In 2007, MMB will concentrate its efforts in securing these additional permits and in managing the terms and conditions that were established in the environmental impact authorization.

4.5.2

LIABILITIES


There are no outstanding liabilities associated with the property. The most recent disturbances were caused by the drilling and metallurgical sampling programs of International Curator in 1997-98 and MMB in 2004-2006, including the development and operation of the underground test mine. All of the disturbed areas from 1997-98 were remediated without any assessed environmental liabilities from that period.  The current MMB drilling program work is in the process of being remediated and no liabilities have been recorded at this time and no liabilities are expected to be incurred from this work going forward.  The test mine work remains and has been incorporated into the EIM permit as an identified disturbed area allowed under the permit.

The project is located within the “buffer zone” of the Vizcaino Biosphere Reserve which is centred on the Desierto de Vizcaino on the west central coast of the Baja.  The Biosphere extends south to encompass the historic Boléo Mining District and to protect certain environmental and cultural features in the town of Santa Rosalia.  It is believed the Biosphere intended to protect the historic buildings dating from the late 1800s associated with the early mining of the Boléo district.  It should be noted these buildings are outside the Boléo project and study area boundary and will not be directly affected by project development.  

Since the Boléo district has been mined for copper and cobalt since 1865, and with two large gypsum quarries currently operating in the region, the authorities have designated the local land suitable for mining and have established land management directives within the Biosphere for development.  The area within the reserve is therefore managed relative to a specific land usage description.

There are no tailings ponds on the lands owned by MMB or on the referenced concessions.  There are 88 small mine waste dumps2 located at the portals of historic mine workings.



 

 

 

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5

ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPY2

 

 


5.1

ACCESS


El Boléo property is located on tidewater on the east coast of the State of Baja California Sur, Mexico, adjacent to the town of Santa Rosalia. Current access for construction equipment would principally be via the Trans-peninsular highway, a trip roughly 850 km south of the US border.  This highway passes directly in front of the plant site location and through Santa Rosalia.  The highway carries heavy traffic volumes year round.  Heavy construction equipment and project supplies would also be brought in by barge to the port of Santa Rosalia and the Pacific Coast marine facilities at Guerro Negro. There are regularly scheduled air services from both the United States and mainland Mexico to the towns of Loreto, which is a two hour drive away to the south, and La Paz which is a six hour drive to the south.  The closest private airstrip is at Palo Verde a half hour drive south.  Port facilities, which originally serviced the copper processing plant and mines until 1985, are still being used on a regular schedule by a ferry service to the mainland at Guaymas.


5.2

CLIMATE


The project area is immediately adjacent to the Gulf of California, with a climate typical of the Sonoran desert region with warm to hot temperatures and minimal seasonal precipitation.  Rainfall is confined mainly to heavy cloud bursts at intervals of several years during tropical cyclones.  Mining operations can be scheduled 365 d/a heavy rainfall events not withstanding.


5.3

LOCAL RESOURCES


According to the most recent Mexican census, nearly 40,000 people live in Mulegé County, which is over 10% of the Baja California Sur population. Population concentrations within the county include Santa Rosalia, Santa Agueda, Mulegé (town), Guerrero Negro, San Ignacio, Bahía Tortugas, San Marcos Island, Gustavo Diaz Ordaz and Bahía Asunción.  Santa Rosalía is the largest town in the county.  Santa Rosalia has a population of 10,000 people and services a fishing fleet, fish processing facilities and two open pit gypsum mines.  Education levels are reasonably high but unemployment is also high.  An informal socio-economic study prepared in 1997 suggested almost all of the non-staff positions might be filled from the local population.  

Hotel accommodation, gasoline, groceries and various hardware goods can all be purchased in Santa Rosalia. Other items including machinery and trained personnel are readily available from

 

 

2 Information derived from Mehner, 2003




 

 

 

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mainland Mexico via the ferry or airplane.  Services and supplies are also obtainable from California, USA.

Fresh water for domestic and drilling purposes is scarce and the town currently obtains most of its supply from wells in Palo Verde, 30 km away. The planned process plant and mines are expected to use seawater for 100% of its requirements.  A desalination plant with a capacity of 200 m3/h will be required to supply process and potable water for use at the plant and mines, including sufficient fresh water for the final stages of metal production.  


5.4

INFRASTRUCTURE


Aside from the many kilometres of drill roads built along each arroyo to access the property, the only other infrastructure on the plant site property is a warehouse and fenced yard. This was constructed in 2002 by MMB as a site improvement and will be used as the site office and base of operations for ongoing development.

A test underground mine and portal site was constructed in 2005 near the Texcoco area of the Boléo arroyo and is currently idle.  The site was used to conduct underground mining tests for geotechnical information and mine equipment evaluations.  The site includes:

·

an over-the-road container trailer modified to house a diesel electrical generator, caplamp light rack, mining supplies and some tools

·

several lean-to sheds for roof bolt supply storage and temporary office

·

a fuel tank (w/containment)

·

a small mine fan

·

steel tunnel cowlings

·

several pieces of underground mine equipment, including a continuous miner, electrical power center, diesel LHD, two diesel mine trucks, and a portable hydraulic roof drilling machine.  


At present, there are no plans to continue mining via these portals.   


5.5

PHYSIOGRAPHY


Property topography is best described as mesa-arroyo with relatively flat plateaus cut by deeply incised arroyo valleys resulting in rugged, steep sided valleys with arroyos that drain into the Gulf of California.  Project site elevations vary from 50 masl to 350 masl.

The project site is very arid with vegetation consisting of a wide variety of cactus.  Over most of the project area vegetation is quite sparse and only locally along the mesa tops, a few kilometres in from the coast, does it occur in significant amounts.



 

 

 

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6

HISTORY

 

 


The discovery of copper in the Boléo district is attributed to local rancher, Jose Rosa Villavicencio who found copper nodules, “boleos” in 1868 while traversing down an arroyo not far from present day Santa Rosalia (Wilson and Rocha, 1955).  The property was sold to two individuals from Guaymas, Sonora who in 1872 began mining and hand sorting high-grade oxidized copper ores from trenches and open cuts and shipping them to smelters in Europe and Guaymas.

Lower grade material was left on dumps or used as backfill in the stopes. This continued until 1884 when declining copper made operations difficult and the firm failed. Until then production was estimated to have been 60,000 short tons grading 24% Cu (Wilson and Rocha, 1955).  A further 120,000 tons averaging about 8% Cu is estimated to have been deposited on dumps or used as backfill in stopes.

In 1884, a number of French geologists and mining engineers including Messrs. Eduoard Cumenge and G. de la Bouglise visited Boléo and after recognizing the vast potential recommended a significant investment to develop the district. On May 16, 1885 the Compagnie du Boléo (later to be known in Mexico as the Compania del Boléo, S.A. - “the Boléo Company”) was formed in Paris, backed mainly by the banking interests of the French House of Rothschild.  

On July 7, 1885, the Boléo Company acquired, from the Mexican government, all mining claims in the region and a concession of about 20,655 ha.  Operations began in 1885 and early work involved a systematic organization of mining and construction of a smelter, port facility, town site and other infrastructure.  

Production started in 1886 and by 1894 had reached over 10,000 annual tons of copper contained in copper matte and “black copper”, which were transported to Europe for treatment.  In 1922, a new smelter was built to produce blister copper, which was shipped to Tacoma, Washington, for refining.  The Compagnie du Boléo was active from 1885 to 1938, when it went into liquidation.  It continued operations on a small scale until 1948, when it was reorganized as the Boléo Estudios e Inversiones Mineras, S.A.  From 1938 on, much of the smelter feed was supplied by small groups of independent miners called “poquiteros”, who re-worked backfilled stopes, robbed pillars and worked smaller, lower grade mines.  Their work is poorly documented.  Smelting operations were initially suspended in 1954.  Nearly all of the ore mined in this period was sourced from the numerous small underground mines throughout the district.

In 1954, operations were taken over by the Compañia Minera Santa Rosalia, S.A., jointly owned by Federal and State Governments and private Mexican interests and managed by the Comisión de Fomento Minero (Fomento Minero is the Mexican Bureau of Mines).  Fomento Minera attempted to sustain copper production by re-opening the smelter and building a leach-precipitation-flotation (LPF) plant to treat dump material, including small amounts of underground ore still being produced by the poquiteros, all to produce a concentrate for the smelter.  Recoveries in the LPF plant are reported to have been about 60% in the early years



 

 

 

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but diminished with time as the plant deteriorated. The smelter continued operation, treating material produced by poquiteros and concentrates from off-shore, until final closure in 1985.

During the latter years of operation at Boléo, there was some exploration in the form of diamond and churn drilling by both French and Mexican concerns.  Shafts were also sunk to intercept the high-grade mineralization.  This work was concentrated in a few relatively restricted areas of the district since the smelting operations needed a cutoff grade of > 4.5%. This exploration work showed that the required grades lay near the southeast corner of the present property and at depths deeper than 200 m.  It should be noted that these early operators assayed only for copper and only portions of the mineralized units were sampled.  The results were thus of little importance in the overall context of a modern exploration program.  

During the 1960s and early 1970s, the Compañia Minera Santa Rosalia S.A, in an effort to find more reserves for the LPF plant, commenced an underground program in which it blocked out a measured resource of backfill material in the Apollo mine area reported to be in the order of 660,000 tonnes grading about 1.60% Cu, with an unknown cobalt and zinc content.  This material was never mined due to lack of funding from the government.

Table 3:

Historical Mining Activities at Boléo

Period of Activity

Metric Tonnes Mined

Ave. Copper Grade

Tonnes Copper Produced

To 1884

~54,400

24.0%

~10,400

1888 – 1947

13,622,327

4.81%

540,3342

1948 – 1952

817,300

3.95%

~27,000

1953 – 1972

1,118,200

3.95%

~36,500

1973 – 1985

720,900

3.02%

~18,000

1964 – 1972

2,500,000

1.40%

n/a


After cessation of operations at Boléo in the 1980s, the bulk of the district was held in the Mexican Strategic National Mining Reserve until 1991. Some months after the release of the ground from the reserve, much of the district was acquired by Minera Terra Gaia, S.A. de C.V., a wholly-owned subsidiary of Terratech Environmental Corporation (Terratech), Barbados, which subsequently optioned the concessions to International Curator Resources Ltd. and its’ subsidiary, Mintec International Ltd., a Barbados company, now Mintec Processing Ltd. (Mintec) as a result of a continuation of Mintec International to British Columbia, Canada in 1993.

Over the period October 1993 to March 1997, Curator completed 68,685 m of HQ coring in 828 holes. In addition, there were 28 holes either re-drilled or twinned, and 58 large diameter holes drilled to recover metallurgical test samples. This was supplemented with 108 hand or excavator dug trenches put in to expose mineralized mantos for both assay data as well as to better define erosional limits of various mantos.  Ten larger trenches were dug, using bulldozers to expose the base of Mantos 2 and 3 and to provide sites for bulk sampling.  Only six of these trenches were successful in exposing the desired contact and, of these, 5 contained Manto 3.

By the end of 1997 a pre-feasibility study incorporating all work completed since 1993 was prepared by Fluor Daniel Wright and presented to the underlying owners.  



 

 

 

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In summary, Fluor Daniel Wright estimated proven reserves at Boléo to be 71.2 M tonnes grading 1.44% Cu, 0.092% Co and 0.55% Zn with further probable reserves of 13.1 M tonnes grading 1.57% Cu, 0.065% Co and 0.81% Zn.  These were deemed sufficient to support an 11,500 t/d operation for about 17 years with an estimated capital cost of about US$440.5 million.  The operation was envisioned to be an open-pit mine with on-site processing utilizing a hydrometallurgical plant producing copper, zinc and cobalt cathode with an option to produce a cobalt sulphide product instead of cathode.  Metal recovery would involve acid leaching with copper, cobalt, and zinc recovered from the leach slurry using a “novel”, in pulp method of recovery.

In 2001, following a significant down-turn in metal prices, International Curator withdrew from the project by handing back its interest in Mintec and the Boléo concessions to Terratech.

After re-gaining control of Boléo, Mintec engaged Bateman Engineering Pty. Ltd. of Australia to assess the pre-feasibility work conducted by Fluor Daniel back in 1997 and determine if significant improvements in mining, processing and capital costs could be achieved relative to the costs presented by Fluor Daniel.  

Most of Bateman’s work concentrated on the metal recovery part of the proposed flow sheet where they proposed using a conventional counter current decantation (CCD) solid-liquid separation circuit followed by base metal recovery from the CCD wash solution.  The Bateman flow sheet also involved acid leaching of the copper, zinc, cobalt, and manganese followed by rejection of the leach residue and separate recovery of copper, zinc sulphate, cobalt either as a metal product or a high value cobalt intermediate precipitate and manganese carbonate.

As a result of the Bateman work and the belief that the new process would greatly improve the economics of the project, Mintec embarked on a corporate re-organization. In April, 2002, Mintec International Corp. acquired all of the rights to the copper/cobalt/zinc/manganese concessions, as described above and registered them with Mintec’s wholly owned Mexican subsidiary Minera y Metalurgica del Boléo S.A. de C.V.  

In November 2003, the shareholders of Mintec entered into an agreement with First Goldwater Resources Inc., a TSX-Venture Exchange listed public company, to exchange all the issued shares of Mintec for shares in First Goldwater (now Baja Mining Corp) and effectively purchased the Boléo copper-cobalt-zinc asset by way of a reverse take-over. The TSX-V Exchange approved the transaction and, after an initial Cdn$10 million capital raising in April 2004 the name First Goldwater Resources Inc. was changed to Baja Mining Corp (Baja), trading on the TSX-V under the symbol BAJ until February 2007 when the shares of Baja Mining Corp were listed for trading on the TSX main board under the symbol BAJ-T.



 

 

 

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7

GEOLOGICAL SETTINGS3

 

 


7.1

REGIONAL GEOLOGY


The Boléo deposits occur within the Boléo sub-basin of the Santa Rosalia basin.  This basin formed as a result of Miocene rifting in the Gulf of California extensional province (Figure 3).  The northward extension of this province is the Basin and Range province of the southwest United States.

Figure 3:

El Boléo Geological Setting

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The timing of initial rifting varies from 13 Ma to 8 Ma. In the Boléo District, which is located near the western edge of the Gulf extensional province, rifting is believed to have started some time after 10 Ma (Sawlan and Smith, 1984).

The early rifting direction was east-northeast and produced north-northwest oriented basins and ranges in basement Miocene volcanic rocks flanking the rift axis. The latest movement, occurring after the Gulf transform and San Andreas wrench-fault systems were initiated, has been right-lateral oblique movement (Stock and Hodges, 1989). This has moved Baja California approximately 350 km. northwest relative to mainland Mexico and has created a number of deep pull-apart basins along the axis of the Gulf of California (Bailes, et al., 2001).

 

 

3 Information from Mehner, 2003




 

 

 

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Stratigraphically the Boléo copper-cobalt-zinc manganese deposits occur within the late Miocene age succession of fine to coarse clastic sedimentary rocks of the Boléo Formation, lying unconformably on andesitic rocks of early to middle Miocene age called the Comondú Volcanics. The Boléo Formation is characterized by a number of coarsening upward cycles of sediments that are believed to represent deltaic deposition in a shallow, near-shore marine basin. The upper part of the formation has been locally eroded and unconformably overlain by similar but barren and fossil-rich sedimentary successions of Pliocene and Pleistocene age delta and beach deposits, known as the Gloria, Infierno and Santa Rosalía Formations.

The Boléo and overlying formations collectively make up the so-called Boléo Basin.  Locally, the entire succession is capped by Pleistocene to recent flows and pyroclastic rocks of the Tres Virgenes Volcanics.  The geology of the district has been described in detail by Wilson and Veytia (1949) and by Wilson and Rocha (1955), in privately prepared reports for International Curator by Peatfield (1995) and Christoffersen (1997) and in numerous other published and unpublished papers and reports referred to in the above mentioned documentation.


7.2

PROPERTY GEOLOGY


The oldest rocks outcropping on the property are andesitic volcanics of the Comondú Formation. They include sub-aerially erupted flows and coarse explosive breccias which grade into coeval epiclastic sediments to the west. The volcanics have been dated from 24 to 11 Ma. They are underlain by Cretaceous granodiorite (Schmidt, 1975).

The overlying Boléo Formation consists of five coarsening upward cycles of sedimentation numbered “4” at the base and “0” at the top (Figure 4). This interpretation is based on work by Curator from 1993 to 1997 and is different from that published by Wilson and Rocha (1955) who interpreted conglomerates, the coarsest units in the stratigraphy, to be the basal unit in each cycle.

The basal unit in the Boléo Formation is a 1 to 5 meter thick limestone unit. It contains cherty lenses and non-diagnostic fossil fragments.  Its occurrence atop very steep paleo surfaces, combined with banding parallel to its base and the cherty horizons, suggests it is at least in part a chemical sediment.

Overlying the limestone or laying directly on Comondú over parts of the district, particularly over much of the coastal area, is an extensive gypsum deposit up to 80 m thick.  Although a few dome or mound structures have been noted, the gypsum unit is characteristically flat to shallow dipping exhibiting laminated to massive and even brecciated textures.  Intraformational carbonate beds are rare.



 

 

 

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Figure 4:

Boléo Formation Stratigraphic Column

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On top of the gypsum/limestone beds is the cyclic succession of clastic beds that average 150 m and range to 270 m thick. Individual cycles range from 20 to 140 m thick and consist of a basal mud and fine volcanic ash horizon (now altered to montmorillonite clay) that hosts the copper-cobalt-zinc manganese mineralization (the manto).

These are overlain by progressively coarser material of maroon coloured, tuffaceous claystone, siltstone, feldspathic sandstone, pebbly sandstone and eventually cobble to boulder orthoconglomerates.

Typically the earliest cycles (manto 4, then 3) are thickest with each successive cycle being thinner. The last cycle is thin and believed to be incomplete. All cycles thin over basement highs and wedge out toward the basin margins. The copper-cobalt-zinc manganese stratiform deposits only occur within Boléo formation rocks.

Unconformably overlying the Boléo clastics are fossiliferous marine sandstones and conglomerates of the lower Pliocene (about 5.3 Ma.). Gloria formation (Bailes, et al., 2001). These in turn are overlain by slight unconformity with a sequence of fossiliferous marine sandstones and conglomerates of the Infierno formation. Unconformably overlying these are fossiliferous sandstone and conglomerate of the Pleistocene, Santa Rosalia formation (Wilson and Rocha, 1955).


7.3

STRUCTURAL GEOLOGY


The Boléo Formation rests on an irregular volcanic basement, with several distinct basement highs and intervening troughs.  In places, these basement highs are so pronounced that they have influenced the deposition of the lower mantos, such that these pinch out against the volcanics and only the upper mantos are present.  There is also a tendency for the sediments of each cycle to thin towards the high ground, giving a stratigraphic compression and thus less vertical separation of mantos.

Faulting is common throughout the district.  The dominant faults are northwest to north-northwest striking and steeply dipping with normal movements.  These faults have downthrows to both east and west, with more of the major faults down dropping to the west.  This, coupled with the generally easterly dip of the mantos, yields a stepwise pattern of the present position of the mineralized beds (Figure 5).  Many of the faults appear to be long-lived, probably with their first movements influencing the initial basin formation and with continuing movements throughout time to the present day.  Vertical displacements can be as much as 50 m to 200 m maximum on the major faults, with much lesser movements toward the ends of these faults and on lesser structures throughout the district. Fault displacements will obviously be important in detailed mine planning; fortunately, in much of the district, the faults and their displacements are well documented in old mining records.

Major faults at Boléo are, in most cases, laterally separated by several hundred, to in some cases, over a thousand m.  Lesser faults are common and more closely spaced.  Faults displace mantos and as a consequence of their dip, may form “fault windows” in which the



 

 

 

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mantos are not present.  However an order of magnitude calculation suggests that the windows may represent less than 2% of the total area.

Many of the major faults have zones a few metres to tens of metres wide in which the rocks, including the mineralized mantos, are highly disrupted.

There has also been some oblique strike-slip movement on many of the faults.  The sense of this movement appears to be predominantly right lateral which would be expected given the spreading regime in the Gulf of California.  



 

 

 

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8

DEPOSIT TYPES4

 

 


The Boléo District hosts a number of mineral deposit types that have the potential to be of sufficient size and grade to be economically mined and processed. The most important of these and the subject of this report are the manto hosted copper-cobalt-zinc manganese deposits which occur in Boléo formation clastic sediments.

Another possible targets are the extensive gypsum beds which occur over portions of the property, particularly north and east of Arroyo Saturno.  It is believed they occur along the same stratigraphic position as those currently being mined immediately north of Boléo and 20 km southeast on San Marcos Island.  



 

 

4 Data from Mehner, 2003




 

 

 

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9

MINERALIZATION5

 

 


Deposits of copper-cobalt-zinc manganese mineralization in the Boléo District occur within widespread, stratiform clay-rich horizons or beds known as “mantos” (manto is a Spanish term used in mining parlance for a generally mineralized layer or stratum).  Within Boléo formation stratigraphy there are up to seven mantos, including two of very limited extent, that occur as relatively flat to generally shallow dipping, stratabound and stratiform beds. These include, with increasing depth, manto 0, 1, 2, 3AA, 3A, 3 and 4.  Historically the major producing manto has been number 3, which yielded approximately 83% of all production between 1886 and 1985 when the plant shut down.  Most of the remaining production has come from manto 1 in the southeast portion of the Boléo area where manto 3 is absent.  A small amount of production has come from the widespread but generally thin manto 2 while an even smaller level of production has come from the relatively restricted manto 3A.  Based on previous studies and exploration work, mantos 1, 2 and 3 still offer the most potential for hosting significant economic reserves.

The mantos themselves tend to be clay rich (ash altered to montmorillonite) with laminated basal zones generally less than 1 m thick overlain by intrabasin slump breccias up to 20 m thick.  Underlying lithologies vary from predominantly ortho-conglomerates in the heart of the Boléo basin to coarse sandstones typically containing pebbles of Comondú volcanics. The contact between the mantos and footwall rocks is sharp.

Overlying lithologies vary from fine to medium grained sandstones. The contact between them and the clay rich slump breccias is gradational.

In a general sense each manto has distinctive characteristics, especially with regard to copper-cobalt ratios and relative concentrations of zinc, manganese and carbonates.  

Metals of interest in the mantos include copper, cobalt, zinc and manganese.  Ore minerals include a fine grained, complex assemblage of primary sulphides including pyrite, chalcocite, chalcopyrite, bornite, carrolite, sphalerite and secondary minerals including malachite, azurite and the rare minerals of boleite, pseudoboleite and cumengite.  Mineralization is generally finely disseminated over intervals up to 20 m thick in the slump breccias.   The richest material typically occurs in the laminated basal section of the Manto, which was historically mined from 1886 to 1972 to an average of about 80 cm and graded 4% to 5% Cu.  

Recently, the presence of a thin layer of claystone within a manto creating a “false bottom” over large areas of mantos 1 and 3 has been recognized.  Within these areas, the rich material is often in the higher portions of the mantos above the false bottom layer.  For a detailed discussion on the concept of the "false bottom" please refer to Section 9.1 of this report.

 

 

5 Data from Mehner, 2003





 

 

 

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Figure 5:

Geological Cross-Sections

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Zoning of the principal economic metals occurs both vertically and laterally. Within individual mantos, copper is enriched at the base, zinc towards the top and cobalt is more or less evenly distributed.  Stratigraphically, vertical zoning shows a trend of zinc enrichment from the lowest manto (4) to the uppermost mantos.  Lateral variations indicate the central core of the Boléo sub-basin is copper rich flanked by a zinc rich marginal zone.  Cobalt is variable and shows no clear correlation with copper or zinc.

Individual mantos and their enclosing strata are “time transgressive”, in that they are progressively younger toward the present Gulf of California.  One very distinctive unit, the “Cinta Colorada” or “red ribbon” is a layer of reddish andesitic-basaltic tuff up to two m thick.  This is interpreted as the product of a single explosive volcanic event, which probably blanketed the entire region.  The Cinta Colorada represents a true “time horizon”, and can be seen to transgress stratigraphy, in some places lying within the unit 2 conglomerate (below Manto 2) and elsewhere in the underlying unit 3 clastic succession.  Thus it demonstrates the time transgressive nature of the enclosing stratigraphic units.

Individual mantos have great lateral continuity and relatively consistent thicknesses. In the principal areas of interest, the lowest Manto (4) lies at the base of the Boléo Formation, directly on the Comondú Formation.  Manto 3 is widespread and thick, and accounting for the largest proportion of the mineral resource.  Mantos 3A and 3AA are less continuous and thinner, lying higher in the succession.  In some places, especially in the Saturno-Jalisco area, 3A merges with 3.  Manto 2 is stratigraphically very continuous but because of its higher stratigraphic position, it is more commonly eroded.  

Manto 1 makes up the bulk of the mineral resource in the southeast portion of the Boléo property, where the lower mantos (3 and 4) were for the most part not deposited.  The beds dip to the southeast and as a consequence, Manto 1 lies deeply buried in this area.  To the northwest, Manto 1 overlies the well mineralized portion of Manto 3 (and in many places, 3A, 3AA and 2).  


9.1

Boléo District False Bottoms


Although in the greater part of the district the mineralized mantos sit directly and in sharp contact on conglomerate footwall, there is also a large portion of the district where the best mineralized horizon is detached from the conglomerate footwalls, or exhibits no conglomerate in the footwall. The “old timers” named these footwalls “false bottoms” which were usually described as sandstones (Wilson, 1955).  The extensive infill drill program carried out by Baja Mining during 2005 and 2006, in conjunction with the drill programs conducted by International Curator Resources during the 1990s have clarified that false bottoms consist not only of sandstone, but of limestones, claystones and polymictic breccias.  For the purpose of resource estimation, the presence of a false bottom is only defined if two or more adjacent diamond drill holes exhibit similar and correlate-able footwall stratigraphy.

The first new significant false bottom was identified in Manto 1 on the eastern basin-side of the manto where the footwall rocks are comprised mostly of finely stratified light coloured limestones (see Figures 6 and 7).



 

 

 

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Figure 6:

Manto 1 Footwall Geology

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Figure 7:

Manto 1 Hole DDH955 – False Bottom Geology

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In reference to Figure 6, Manto 1 has formed above the complex footwall stratigraphy reflecting a deepening basin on the eastern side of Cerro Montado island. The buried Cerro Montado island exhibits typical Boléo formation conglomerate directly overlying the island’s Comondú volcanics on the western side of the basin. The manto occurs directly above conglomerate footwall in a patchy pattern only in the western sector, although sandstones and claystone false bottoms are also present above the conglomerate in the western sector. The eastern limit with presence of conglomerate in the footwall is shown as an interrupted black line.  

East of this line there is no conglomerate present in the footwall rocks of the manto. The western limit of economic mineralization is shown as a thick red line in Figure 6 and mineralization tends to occur where “fingers” of Manto 1 overlie conglomerate with claystone false bottoms. This finger-like geometry seems to reflect the details of paleo-topography with well mineralized manto occurring along valley bottoms coming off Cerro Montado Island, where deeper and quieter submarine conditions led to deposition of claystone first and the entrapment of later richer brines forming Manto 1.



 

 

 

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Moving westwards, the conglomerate facies disappears altogether, and the footwall stratigraphy is dominated first by sandstones and claystones and further to the east by a base of well laminated hard and light coloured limestones which are clearly identifiable (Figure 7).

The thickness of this unit at the base of Manto 1 ranges from a few tens of centimetres to over 1 meter. Even further to the east, the manto itself exhibits an increase in calcareous content indicating contemporaneous deposition of limestone and mineralized manto. Some of the highest copper grades (3% ~4% Cu) are associated with the zone where Manto 1 directly overlies either claystone or limestone in what may have represented a feeder zone that was coincident with a ridge.

The location of old stopes also partly mimics the Manto 1 paleo-topography. Thus, the evidence provided by the stratigraphic changes in the footwall of the manto, as well as the pattern of the old stopes, and the paleo-topography of the base of Manto 1 all indicate that a hard finger-like geologic boundary for Manto 1 occurs to the west where the distribution of the brine was sharply limited by impoundment against the paleo-topographic barrier created by the buried Cerro Montado island. To the east, the manto shows a more gradational zoning pattern, where the best copper occurs in the central portion of the basin in the downthrown block of the Santa Agueda fault system, providing progressively better cobalt and lower copper grades further eastwards on the up thrown side of the Santa Agueda faults.

Several other areas with false bottoms have been identified in Manto 3 (see Figure 8).  The Rancheria false bottom is located in the southern extreme of Manto 3 where it ends against the northern flank of the buried Cerro Montado Island.  Where the manto under Providencia Creek sits directly above its footwall conglomerate, towards the south under the Rancheria ridge, the well mineralized horizon sits either in the middle (as in DDHs 06-463, 06-1072 and 06-1004) or towards the top (as in DDH06-970 and 06-971) of a wedge of low grade manto which occurs on the flanks of the buried island.  This stratigraphic change is demonstrated by the position of old stopes that followed the well mineralized horizon downwards to a position only metres away from the high grade DDH06-1072. The surface area influenced by this facies change is approximately 600 m long and 250 m wide.

The second and largest area of Manto 3 over which a false bottom occurs is in the central portion of the district in between and eastwards of the Boléo and Juanita islands (Purgatorio-Soledad wedge). The funnel-shaped wedge is approximately 2.5 km long, 1.5 km wide at its southern base and less than 200 m wide at its northern top. The wedge is situated where the structural pattern of the district is dominated by north-south trending faults, reflecting the position of an extensional rip-apart sub-basin.  This sub-basin is related to local normal and dextral fault displacements, which are responsible on a more regional scale for the formation of the Gulf of California.

What is common within the wedge is the presence of a low grade claystone or polymictic breccia false bottom above the footwall conglomerate, with the well mineralized horizon sitting at the top of Manto 3. The thickness of the false bottom ranges from less than a meter to several metres wide. In the central portion of the wedge the false bottom is absent and an “island” of footwall conglomerate directly underlies the mineralized manto.



 

 

 

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Figure 8:

Manto 3 Location of False Bottoms

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Both vertical grade distribution in the metals, as well as the position of old stopes in this area provide additional evidence the well mineralized horizon occurs at the top of the manto. The example from three diamond holes (Table 4) demonstrates vertical change in grade between the claystone false bottom and the well mineralized top of Manto 3 in this area.



 

 

 

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Table 4:

Manto 3 - Vertical Grade Variation Associated With False Bottoms

DDH

Interval (m)

Width (m)

Cu%

Co%

Zn%

Mn%

Lithology

95-167

117.68-118.41

0.73

1.33

0.112

0.67

2.15

Retaque

 

118.41-119.17

0.76

4.17

0.118

0.44

3.61

Retaque

 

119.17-120.09

0.92

0.52

0.013

0.11

0.29

Claystone

 

120.09-120.69

0.60

0.97

0.011

0.08

0.36

Claystone

 

120.69-121.72

1.03

0.50

0.010

0.10

0.11

Claystone

95-193

139.54-140.15

0.61

1.27

0.126

0.12

4.25

Breccia

 

140.15-141.14

0.99

2.75

0.161

0.16

4.20

Breccia

 

141.14-142.13

0.99

0.08

0.007

0.06

0.15

Claystone

 

142.13-142.50

0.37

0.02

0.010

0.13

0.25

Claystone

95-192

54.1-55.5

1.40

3.06

0.141

0.60

3.22

Polymictic Breccia

 

55.5-56.9

1.40

0.63

0.024

0.27

0.22

Claystone


DDH 95-167 shows that the better mineralized part of the manto is in fact “retaque” (stope fill) indicating that the old stopes were located at the top of the manto and that mining took place above a claystone false bottom.

The Boléo Creek false bottom is located where Manto 3 approaches and ends up against the southeastern side of Boléo Island.  The "2 de Abril" sub-basin, which is located in the northwestern sector of the district and at the northwestern side of Boléo Island, also exhibits the better mineralized horizon of Manto 3 at the top of the manto.

A false bottom has also been identified in parts of Manto 4 where the best mineralized area is also characterized by the higher grades being at the top of the manto.



 

 

 

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10

EXPLORATION

 

 


Since acquiring control of the Boléo project in 2001, and prior to December 2004, MMB has concentrated on carrying out a complete geological, mining and processing review of the Boléo Property in place of a straight field exploration work. This includes an independent review of the open pit copper-cobalt-zinc manganese reserves, an in-depth review of alternate flow sheets for the processing of ore from the Boléo Property and an investigation into alternative mining methods, in particular the potential to use underground mining techniques.

This review resulted in a new pre-feasibility study being issued by Bateman Engineering Pty Limited, of Perth, Western Australia in February 2002, which focused principally on a new metallurgical flow sheet for the processing of ore from the Boléo Property.

Drilling activities commenced on the property in December 2004 (see Section 10.3 Drilling, below).

All historical work, including that carried out by Curator in conjunction with Mintec prior to its re-organization, is documented in Section 10.2 under History.





 

 

 

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31 January 2007

Page 33 of 188




[bajatechrep001.jpg]

EL BOLEO PROJECT

MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT




11

DRILLING

 

 



11.1

GENERAL


The current resource study uses all the drill holes completed by previous owners of the Boléo property, plus the additional holes of the three infill drill programs carried out by MMB between December 2004 and September 15, 2006 (DDHs 04-928 to 06-982).

A small number of these holes can be classified as exploration holes and did not serve the purpose of infill drilling.  These holes are numbered 04-928, 04-929, 05-947 and 05-948.


11.2

HISTORICAL


The oldest recorded drilling program was carried out between 1927 and 1940 when 10,237 metres were drilled in 46 vertical churn holes (Wilson and Rocha, 1955). Most of these holes were drilled in the southeast portion of the property to explore for manto 1 in the Rancheria, San Luciano and Montado areas.  Further diamond drilling was carried during the latter years of mining operations when Fomento Minero was looking for high-grade reserves to exploit. Records of this drilling were either never kept or have been lost or destroyed.

By far the most extensive drilling program ever conducted in the Boléo District was that of International Curator Resources Ltd. between 1993 and 1997.  All exploration diamond drilling was completed using skid mounted Longyear 38 drill rigs moved with logging skidders.  Core size was HQ (63.5 mm diameter), reduced to NQ (47.6 mm diameter) when necessary because of drilling problems.  In areas where considerable thickness of overlying barren stratigraphy (Gloria, Infierno and Santa Rosalia formations) was expected, the upper portion of the hole was triconed before coring was begun near the target horizon.

The need to increase the Boléo resource to the measured and indicated classification led to the execution of three drill programs between December 2004 and January 2007.


11.3

2004 – 2007


Three infill drill programs were carried out by MMB on the property between December 2004 and January 2007. These drill programs were designed to infill the existing drill hole coverage with the aim of improving the confidence in the resource estimates.  The location of all the drill holes is shown in Figure 9.

The first program comprised 15 holes for 2,566 m, between December 6, 2004 to January 29 2005 (DDHs 928 to 941), the second program comprised 18 holes for 3,174 m from May 11, 2005 to July 3, 2005 (DDHs 942 to 959), and the third program which comprised 123 holes for a total of 20,609 m from February 1, 2006 to September 15, 2006 (DDHs 960 to 1082).  A total of 156 new drill holes were completed for a total of 26,349 m.



 

 

 

ZAB406-00233:EM15.02.NGI

31 January 2007

Page 34 of 188




[bajatechrep001.jpg]

EL BOLEO PROJECT

MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT




Figure 9:

Drill Hole Locations –Historical & Infill

[bajatechrep014.gif]


Table 5 contains the drill hole location information for the latest drill hole programs, covering diamond drill holes DDH 928 to 1082.

Table 6 contains the various results of the latest drilling programme, covering diamond drill holes DDH928 to 1082.  Results are reported for the logged manto intersections only from each hole and are quoted as single composite intervals for each manto.



 

 

 

ZAB406-00233:EM15.02.NGI

31 January 2007

Page 35 of 188




[bajatechrep001.jpg]

EL BOLEO PROJECT

MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT




Table 5:

Diamond Drill Hole Locations – Current Programs (DDH928-1082)

Hole ID

Easting

Northing

El.

Total
Depth

 

Hole ID

Easting

Northing

El.

Total
Depth

04-0928

-561.96

-1209.94

264.22

190.75

 

06-0987

425.02

730.93

154.5

166.55

04-0929

769.22

-3482.88

228.34

252.05

 

06-0988

497.43

811.61

152.72

167.85

05-0930

-2156.14

1806.45

152.55

67.05

 

06-0989

3121.87

-530.46

183.21

260.9

05-0931

-2328.86

1601.64

161.79

57.55

 

06-0990

-127.15

1060.82

227.46

210.4

05-0932

-2444.79

1421.11

174.14

132.5

 

06-0991

-245.12

1167.57

228.85

209.3

05-0933

-2535.9

1265.82

180.07

195.25

 

06-0992

1564.43

-251.47

64.8

45

05-0934

-1005.47

420.98

107.45

159.85

 

06-0993

1434.63

-395.35

70.18

41.5

05-0934a

-1010.42

417.43

107.62

192.15

 

06-0994

1483.44

-552.94

70.71

51.55

05-0935

-344.8

440.01

80.44

167.75

 

06-0995

1290.67

-544.41

73.36

48.3

05-0936

193.85

3.64

74.55

182.65

 

06-0996

2561.59

-881.73

202.26

243.58

05-0937

-913.46

516.64

105.29

124.6

 

06-0997

2791.285

-1400.491

202.72

177.92

05-0938

1068.7

-1038.58

89.46

188.5

 

06-0998

3723.1

-715.651

61.24

96.1

05-0939

978.24

-1216.07

93.01

142.85

 

06-0999

3481.557

-504.168

82.63

173.6

05-0940

708.32

-2033.87

119.99

186.05

 

06-1000

3324.58

-405.978

92.88

157.9

05-0941

-1279.98

954.24

267.34

326

 

06-1001

26.82

963.92

192.06

177.25

05-0942

-177.44

1094.05

232.21

208.15

 

06-1002

-14.16

1213.95

226.92

231.95

05-0943

-175.59

1103.88

231.82

344.65

 

06-1003

82.96

1401.86

226.73

276.2

05-0944

515.74

566.14

61.86

216.25

 

06-1004

1886.62

-829.83

222.07

295.63

05-0945

-1593.29

2478.35

117.04

208.9

 

06-1005

-272.81

1322.82

209.85

209.85

05-0946

-2362.16

2500.04

252.38

230.8

 

06-1006

-158.6

1469.56

224.01

248.7

05-0947

1212.12

-2708.22

243.84

117.9

 

06-1007

-8.238

1545.81

229.3

275.1

05-0948

989.89

-3186.24

238.08

238

 

06-1008

2567.85

-1485.04

202.7

183.63

05-0949

3119.52

-668.48

183.83

259.4

 

06-1009

2434

-347.39

201.74

225.93

05-0950

3422.26

-839.11

77.5

143.4

 

06-1010

3174.75

-307.66

104.06

222.2

05-0951

3079.72

-969.8

101.52

125.05

 

06-1011

2830.13

-478.42

135.66

201

05-0952

2982.47

-1083.21

110

173

 

06-1012

3034.68

-411.87

118.27

135.75

05-0953

3557.88

-782.85

67.35

112.5

 

06-1013

533.68

1289.7

78.32

107.35

05-0954

3751.26

-852.11

56.39

106.65

 

06-1014

373

1130

110

124.75

05-0955

3448.8

-1318.79

74.89

130.15

 

06-1015

223.55

1036.84

129.37

144.3

05-0956

317.8

418.34

69.97

56.9

 

06-1016

759.925

1159.34

78.75

128.7

05-0957

95.48

693.49

184.51

161.2

 

06-1017

671.336

764.103

55.47

73.25

05-0958

405.25

877.91

158.56

165.6

 

06-1018

697.127

653.225

57.18

63.65

05-0959

522.47

1083.51

150.71

175.8

 

06-1019

3894.46

-920.119

47.32

93.95

06-0960

3350.7

-1452.221

85.664

135.11

 

06-1020

2396.99

-1475.467

211.616

200.08

06-0961

3196.833

-1492.658

96.4

169.88

 

06-1021

446.387

1370.493

92.538

118.1

06-0962

2809.012

-800.552

199.362

205.72

 

06-1022

2787.684

-1215.667

126.783

120.85

06-0963

2636.241

-849.439

204.97

202.82

 

06-1023

2863.332

-1137.225

120.304

138.75

06-0964

2327.664

-821.916

211.163

252.09

 

06-1024

2499.2

-1285.33

148

142.1

06-0965A

2984.115

-779.126

187.485

208.54

 

06-1025

3544.254

-1499.62

163.457

232.35

06-0966

2883.369

-674.971

190.566

202.82

 

06-1026

3410.035

-1622.453

177.24

278.95

06-0967

3218.229

-673.91

193.04

250.3

 

06-1027

3202.275

-1674.148

178.337

262.2

06-0968

2655.783

-1034.359

207.774

224.17

 

06-1028

3023.133

-1702.861

190.199

180.4

06-0969

2642.35

-573.727

204.795

227.22

 

06-1029

2893.864

-1803.652

194.269

238.5

06-0970

2053.019

-811.571

214.976

230.27

 

06-1030

850.086

439.985

59.168

63.8

06-0971

2000.533

-657.837

218.508

245.52

 

06-1031

751.213

465.275

58.282

55.65

06-0972

2336.122

-496.693

205.894

254.2

 

06-1032

686.001

386.568

60.575

54.55

06-0973

-530.029

829.648

241.89

190.25

 

06-1033

462.774

102.808

64.628

48.75

06-0974

-395.135

848.398

240.541

190.2

 

06-1034

295.766

227.733

68.068

54.9

06-0975

-238

735.73

234.91

195

 

06-1035

-8.957

222.214

72.649

43.95

06-0976

-414.44

1041.63

235.267

195

 

06-1036

513.177

433.436

63.271

57.2

06-0977

3281.76

-1121.34

180.018

219.41

 

06-1037

2296.533

-1331.305

165

114

06-0978

2430.27

-701.08

206.865

261.31

 

06-1038

2192.228

-1281.247

175.499

128.15

06-0979

3182.74

-1287.86

184.307

193.85

 

06-1039

2973.117

-1087.651

110

118.3

06-0980

2944.6

-1305.54

194.1

245.07

 

06-1040

2921.715

-2095.619

101.316

177.7

06-0981

-66.61

577.75

197.843

170.8

 

06-1041

3350.474

-1826.613

80.091

184.95

06-0982

-0.45

585.01

195.053

172.1

 

06-1042

3243.21

-1894.093

84.925

170.7

06-0983

186.82

881.68

192.893

201.55

 

06-1043

2676.089

-2570.624

124.23

156.15

06-0984

145.34

556.07

161.717

144.7

 

06-1044

3848.69

-1133.866

45.992

175.55

06-0985

268.06

589.28

158.41

157.05

 

06-1045

3704.952

-1176.098

56.308

129.8

06-0986

348.96

636.18

156.03

161.25

 

06-1046

2888.242

-1589.836

117.881

128.15




 

 

 

ZAB406-00233:EM15.02.NGI

31 January 2007

Page 36 of 188




[bajatechrep001.jpg]

EL BOLEO PROJECT

MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT




Hole ID

Easting

Northing

El.

Total
Depth

 

Hole ID

Easting

Northing

El.

Total
Depth

06-1047

2764.409

-1660.283

126.838

151.45

 

06-1067

2593.19

-2860.708

140.921

113.7

06-1048

2536.325

-1667.435

141.525

149.26

 

06-1068

2406.14

-2733.43

141.621

68.5

06-1049

3534.311

-1840.815

71.921

190.55

 

06-1069

2448.127

-2594.405

142.433

81

06-1050

3651

-1776

66

178.2

 

06-1070

2765.874

-2181.574

109.119

152.8

06-1051

1730.157

-242.117

59.372

44.3

 

06-1071

2124.837

-998.746

213.4676

222.85

06-1052

1638.898

-405.448

65.267

47.88

 

06-1072

1754.549

-1052.788

223.001

280.95

06-1053

1713.964

-437.524

64.32

51.85

 

06-1073

2215.643

-350.787

206.548

216.25

06-1054

1383.348

-575.701

73.639

51.85

 

06-1074

2395.16

-164.752

196.685

218.2

06-1055

1365.22

-793.227

77.935

69.9

 

06-1075

2572.601

-410.941

207.362

179.5

06-1056

1925.381

-240.29

83.373

94.55

 

06-1076

1542.991

-709.676

139.376

137.7

06-1057

2191.465

-754.431

211.239

222.7

 

06-1077

1507.634

-847.329

147.345

153.15

06-1058

2256.946

-894.332

213.516

274.5

 

06-1078

3383.144

-186.706

185.193

206.4

06-1059

2399.632

-893.945

208.342

245.5

 

06-1079

2665.826

-278.444

193.13

224.55

06-1060

2714.409

-828.027

202.535

262.95

 

06-1080

2837.633

-267.714

186.03

253.15

06-1061

2749.111

-1875.109

198.11

213.65

 

06-1081

3895.591

-1609.201

50.764

193.35

06-1062

2584.629

-1943.91

205.96

174.95

 

06-1082

3998

-1653

45

196.75

06-1063

2422.407

-2010.307

210.66

118.85

 

 

 

 

 

 

06-1064

2519.457

-217.098

207.602

137.35

 

 

 

 

 

 

06-1065

2339.935

-1824.408

209.87

147.9

 

 

 

 

 

 

06-1066

2666.607

-2379.18

116.735

131.5

 

 

 

 

 

 


Table 6:

Diamond Drill Hole Results – Current Programs (DDH928-1082)

Hole ID

From

To

Thickness

Manto

Cu%

Co%

Zn%

Mn%

04-0928

167.04

170.36

3.32

4

0.13

0.009

0.32

2.47

04-0929

147.4

151.3

3.9

3

0.02

0.006

5.12

1.52

04-0929

174.72

177.4

2.68

4

0.00

0.003

1.75

4.74

04-0929

189.4

193.22

3.82

5

0.28

0.006

0.61

2.51

05-0930

3.5

6.1

2.6

3

2.01

0.109

0.72

5.25

05-0930

55.05

60.98

5.93

4

0.50

0.033

0.32

0.69

05-0931

5.35

9.1

3.75

3

1.66

0.039

0.41

2.16

05-0931

49.34

54.76

5.42

4

0.47

0.038

0.44

0.59

05-0932

111.65

113.27

1.62

4a

2.24

0.092

0.82

0.84

05-0932

119.6

126.83

7.23

4

1.01

0.075

0.39

2.88

05-0933

6.7

13.3

6.6

3

0.28

0.040

0.62

2.85

05-0933

172.8

179.27

6.47

4

0.29

0.022

0.14

0.18

05-0934

114.65

152.1

37.45

4

0.46

0.028

0.20

1.00

05-0934a

116.1

147.16

31.06

4

0.28

0.030

0.25

0.94

05-0935

19

21.35

2.35

3

1.10

0.076

0.56

7.41

05-0935

97.4

102.61

5.21

4a

0.18

0.022

0.33

0.08

05-0935

108.4

117.45

9.05

4

0.08

0.059

0.31

0.90

05-0936

26.3

27.7

1.4

3a

0.11

0.043

1.15

4.14

05-0936

33.35

36.14

2.79

3

1.04

0.066

0.44

2.54

05-0936

128.1

150.26

22.16

4

0.07

0.032

0.17

0.17

05-0937

78.1

86.12

8.02

4

0.14

0.016

0.15

0.96

05-0938

48.05

54.5

6.45

3

0.48

0.009

0.17

0.52

05-0938

110.43

122.31

11.88

4a

0.08

0.021

0.15

0.37

05-0938

128.67

131.5

2.83

4

0.32

0.041

0.17

3.12

05-0939

43.65

47.7

4.05

3a

0.16

0.040

0.30

2.04



 

 

 

ZAB406-00233:EM15.02.NGI

31 January 2007

Page 37 of 188




[bajatechrep001.jpg]

EL BOLEO PROJECT

MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT




Hole ID

From

To

Thickness

Manto

Cu%

Co%

Zn%

Mn%

05-0939

51.6

54.58

2.98

3

2.57

0.081

0.55

4.89

05-0939

109.7

123.25

13.55

4

0.65

0.038

1.08

0.57

05-0940

48.8

61

12.2

3a

0.01

0.004

0.07

0.06

05-0940

117.13

157.74

40.61

4

0.09

0.015

0.17

0.44

05-0941

73.7

78.77

5.07

2

0.09

0.009

0.37

2.09

05-0941

145.8

147.29

1.49

3

2.21

0.122

0.56

7.20

05-0941

296.1

309.8

13.7

4a

0.74

0.030

0.34

0.58

05-0942

143.15

148.25

5.1

2

1.28

0.061

0.26

3.78

05-0942

201.75

203.21

1.46

3a

0.33

0.052

0.63

5.42

05-0943

146.05

149.7

3.65

2

1.75

0.041

0.68

2.35

05-0943

200.35

203.4

3.05

3

0.12

0.040

0.73

7.21

05-0943

328.4

332.55

4.15

4

0.14

0.025

0.18

0.88

05-0944

52.1

53.1

1

3a

0.06

0.042

1.43

6.08

05-0944

55.28

57.57

2.29

3

1.87

0.189

0.44

3.73

05-0944

205.52

215.85

10.33

4

0.05

0.046

0.26

1.59

05-0945

34.91

36

1.09

2a

0.40

0.017

0.20

0.45

05-0945

39.9

44.35

4.45

2

0.02

0.040

0.66

4.56

05-0945

73.7

78.42

4.72

3

2.20

0.028

0.20

0.91

05-0945

197.15

206.88

9.73

4

0.03

0.019

0.16

0.21

05-0946

113.05

116.91

3.86

3

2.43

0.039

0.10

0.33

05-0946

187

188.3

1.3

4

0.01

0.013

0.13

0.34

05-0947

73.5

78.41

4.91

3

0.02

0.003

0.71

1.60

05-0947

88.34

94.19

5.85

4

0.06

0.004

0.35

1.16

05-0948

97.35

104.27

6.92

1

0.00

0.002

0.08

0.30

05-0948

108.9

113.75

4.85

2

0.05

0.004

1.19

4.10

05-0948

119.6

122.33

2.73

3

0.17

0.005

0.75

4.61

05-0948

157.4

166.19

8.79

4a

0.14

0.008

1.14

3.54

05-0948

174.9

179.3

4.4

4

0.31

0.018

0.29

5.38

05-0949

185.42

187.1

1.68

0a

0.01

0.008

0.93

3.30

05-0949

190.48

191.4

0.92

0

0.01

0.006

0.72

1.00

05-0949

225.25

227.27

2.02

1

2.55

0.109

0.56

4.24

05-0950

63.7

64.16

0.46

0a

0.03

0.012

0.96

0.60

05-0950

66.44

68.03

1.59

0

0.01

0.008

0.42

0.70

05-0950

97.17

100.03

2.86

1

4.37

0.083

0.40

3.25

05-0951

47.21

49.7

2.49

0

0.07

0.011

0.45

0.44

05-0951

90.11

91.82

1.71

1

6.15

0.172

0.23

1.15

05-0952

108.75

111.7

2.95

1

0.36

0.036

0.42

4.53

05-0952

152.02

154.17

2.15

2

0.42

0.103

0.61

6.46

05-0952

160.6

169.9

9.3

3

0.05

0.062

0.82

7.10

05-0953

74.15

78

3.85

0

0.01

0.006

0.25

0.83

05-0953

98.87

100.75

1.88

1

2.16

0.107

0.64

3.41

05-0954

55.03

56.83

1.8

0a

0.02

0.006

0.42

1.01

05-0954

58.89

60.77

1.88

0

0.02

0.027

0.59

1.46

05-0954

94.55

97.06

2.51

1

1.73

0.085

0.59

4.40



 

 

 

ZAB406-00233:EM15.02.NGI

31 January 2007

Page 38 of 188




[bajatechrep001.jpg]

EL BOLEO PROJECT

MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT




Hole ID

From

To

Thickness

Manto

Cu%

Co%

Zn%

Mn%

05-0955

70.27

71.41

1.14

0a

0.03

0.007

0.80

2.40

05-0955

73.92

76.78

2.86

0

0.02

0.008

0.80

0.75

05-0955

121.45

123.74

2.29

1

3.17

0.090

0.98

4.45

05-0956

46.45

53.98

7.53

3

1.00

0.041

0.36

1.88

05-0957

110.13

116.27

6.14

2

0.42

0.044

0.38

1.82

05-0957

152

157.88

5.88

3

1.36

0.071

0.41

2.65

05-0958

113.28

114.73

1.45

2

1.94

0.143

0.39

7.50

05-0958

157.48

163.56

6.08

3

0.75

0.033

0.59

3.42

05-0959

126.5

130.52

4.02

2

0.65

0.077

0.50

5.44

05-0959

166.6

174.25

7.65

3

0.76

0.018

0.26

0.80

06-0960

73.11

80

6.89

0

0.02

0.008

0.27

0.51

06-0960

128.55

132.24

3.69

1

0.64

0.070

0.55

2.22

06-0961

52.46

54.55

2.09

0

0.01

0.019

0.44

0.84

06-0961

100.87

104.16

3.29

1

0.20

0.024

0.26

1.69

06-0961

163.78

164.38

0.6

2

0.01

0.014

0.42

2.19

06-0962

159.94

160.85

0.91

0a

0.02

0.009

0.98

1.43

06-0962

166.37

168.56

2.19

0

0.02

0.007

0.42

0.67

06-0962

199

201.69

2.69

1

2.19

0.069

0.60

3.24

06-0963

134.96

138.2

3.24

0a

0.01

0.004

0.30

1.37

06-0963

139.47

143.51

4.04

0

0.01

0.011

0.57

0.65

06-0963

175.37

180.48

5.11

1

0.50

0.032

0.52

1.76

06-0964

103.37

105.8

2.43

0a

0.01

0.004

0.31

1.40

06-0964

108

112.62

4.62

0

0.01

0.010

0.66

0.98

06-0964

135.72

145.65

9.93

1

0.06

0.024

0.47

2.36

06-0964

204.85

207.03

2.18

2

6.31

1.435

4.67

10.64

06-0965A

161.65

166.15

4.5

0

0.04

0.014

0.53

0.72

06-0965A

198.15

202.64

4.49

1

1.86

0.075

0.58

4.73

06-0966

152.27

153.48

1.21

0a

0.03

0.008

0.53

2.59

06-0966

155.5

158.6

3.1

0

0.02

0.013

0.52

0.82

06-0966

191.92

194.62

2.7

1

1.41

0.090

0.80

3.01

06-0967

202.41

208.2

5.79

0

0.01

0.004

0.18

0.63

06-0967

242.98

244.73

1.75

1

2.89

0.065

0.27

1.46

06-0968

145.71

150.2

4.49

1

0.58

0.041

0.63

1.74

06-0968

214.26

217.3

3.04

2

2.47

0.252

2.19

7.78

06-0969

156.1

159.46

3.36

1

1.31

0.058

0.38

1.67

06-0969

218.07

225.82

7.75

2

0.12

0.047

0.46

2.75

06-0970

84.65

89.8

5.15

1

0.01

0.008

1.10

2.73

06-0970

120.44

123.87

3.43

2a

0.01

0.004

0.22

1.07

06-0970

128.22

130.32

2.1

2

0.16

0.014

0.60

3.43

06-0970

193.65

195.81

2.16

3

1.66

0.065

0.80

3.34

06-0970

199.26

209.94

10.68

4

0.35

0.036

0.17

3.90

06-0971

72.56

74.43

1.87

0a

0.01

0.005

0.17

2.23

06-0971

77.48

81.19

3.71

0

0.01

0.013

1.18

2.09

06-0971

108.27

120.47

12.2

1

0.05

0.006

0.37

0.96



 

 

 

ZAB406-00233:EM15.02.NGI

31 January 2007

Page 39 of 188




[bajatechrep001.jpg]

EL BOLEO PROJECT

MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT




Hole ID

From

To

Thickness

Manto

Cu%

Co%

Zn%

Mn%

06-0971

173.57

180.42

6.85

2

0.14

0.019

0.46

1.87

06-0971

218.07

222.02

3.95

3

2.44

0.039

0.47

2.67

06-0972

119.91

121.11

1.2

0a

0.01

0.006

0.34

1.53

06-0972

123.96

127.34

3.38

0

0.05

0.007

0.29

0.45

06-0972

157.89

158.66

0.77

1

1.86

0.041

0.29

2.38

06-0972

205.2

207.2

2

2

0.89

0.066

0.43

4.42

06-0972

234.42

236.26

1.84

3

0.17

0.030

0.62

1.72

06-0973

125

127.1

2.1

2

0.15

0.013

1.49

4.58

06-0973

183.95

184.9

0.95

3a

0.21

0.068

1.18

5.03

06-0973

188.45

189.07

0.62

3

0.63

0.027

0.28

0.79

06-0974

123.4

128.15

4.75

2

0.16

0.014

0.92

1.60

06-0974

182.7

184

1.3

3a

1.47

0.138

0.85

6.15

06-0974

187

188.35

1.35

3

1.79

0.121

0.49

2.56

06-0975

137.05

141.72

4.67

2

0.06

0.019

0.72

5.66

06-0975

184.35

185.7

1.35

3a

0.18

0.057

1.51

7.52

06-0975

189.07

191

1.93

3

1.99

0.065

0.46

2.18

06-0976

130.85

134.05

3.2

2

2.21

0.070

0.99

5.90

06-0976

186.25

187.8

1.55

3a

2.53

0.067

1.06

9.41

06-0976

188.7

191.45

2.75

3

7.07

0.076

0.35

1.16

06-0977

154.49

155.89

1.4

0a

0.05

0.008

0.20

1.05

06-0977

158.22

161.48

3.26

0

0.01

0.007

0.27

0.48

06-0977

205.24

207.53

2.29

1

3.68

0.135

0.57

2.98

06-0978

123.83

125.39

1.56

0a

0.02

0.007

0.12

1.36

06-0978

129.72

131.75

2.03

0

0.04

0.011

0.55

0.71

06-0978

157.34

159.71

2.37

1

0.59

0.026

0.28

2.33

06-0978

211.31

213.71

2.4

2

0.90

0.020

0.35

3.31

06-0978

241.72

243.74

2.02

3

0.24

0.016

0.12

2.74

06-0978

252

253.68

1.68

4

0.22

0.017

0.19

1.92

06-0979

136.5

138.4

1.9

0a

0.02

0.007

0.53

1.67

06-0979

140.34

142.94

2.6

0

0.01

0.010

0.61

0.59

06-0979

189.5

190.15

0.65

1

1.86

0.036

0.17

2.71

06-0980

164.66

169.46

4.8

1

1.35

0.063

0.42

2.74

06-0980

233.78

235.46

1.68

2

0.48

0.088

1.37

5.65

06-0980

238.37

245.07

6.7

3

0.02

0.042

0.47

4.26

06-0981

105.35

107.15

1.8

2

0.11

0.015

0.36

2.07

06-0981

161.8

162.64

0.84

3a

0.12

0.046

2.31

12.58

06-0981

167.1

168.73

1.63

3

3.68

0.041

0.33

0.42

06-0982

105.6

109.82

4.22

2

1.53

0.043

0.69

2.77

06-0982

161.2

162.45

1.25

3a

0.84

0.110

0.46

8.80

06-0982

167.7

169.1

1.4

3

2.38

0.086

0.28

2.36

06-0983

138.8

141.07

2.27

2

1.63

0.097

0.38

6.60

06-0983

192.9

195.2

2.3

3a

0.05

0.026

0.35

2.95

06-0983

199.2

199.6

0.4

3

3.72

0.080

0.34

2.47

06-0984

135.95

137.52

1.57

3a

0.38

0.040

0.36

3.68



 

 

 

ZAB406-00233:EM15.02.NGI

31 January 2007

Page 40 of 188




[bajatechrep001.jpg]

EL BOLEO PROJECT

MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT




Hole ID

From

To

Thickness

Manto

Cu%

Co%

Zn%

Mn%

06-0984

140.55

142.1

1.55

3

3.23

0.032

0.21

1.76

06-0985

92.6

96.65

4.05

2

0.66

0.052

0.63

4.78

06-0985

148.9

149.2

0.3

3a

0.02

0.015

0.17

0.28

06-0985

151.45

153.5

2.05

3

0.56

0.019

0.38

1.89

06-0986

99.25

104.47

5.22

2

0.13

0.045

0.31

2.29

06-0986

154.75

157.25

2.5

3a

0.28

0.077

0.74

5.93

06-0986

158.93

160.05

1.12

3

3.68

0.121

0.49

4.32

06-0987

107.9

110.52

2.62

2

0.09

0.039

0.47

2.32

06-0987

160.35

161.1

0.75

3a

0.04

0.010

0.16

0.70

06-0987

162.3

164.4

2.1

3

4.82

0.167

0.28

1.61

06-0988

110.02

112.64

2.62

2

0.16

0.067

0.61

5.58

06-0988

159.6

161.45

1.85

3a

0.11

0.035

0.58

3.43

06-0988

163.7

163.82

0.12

3

8.15

0.026

0.24

3.34

06-0989

182.65

187.25

4.6

0

0.06

0.016

0.50

0.83

06-0989

217.44

220.19

2.75

1

0.75

0.044

0.16

2.42

06-0990

151.5

153.75

2.25

2

1.57

0.106

1.62

8.32

06-0990

201.8

202.8

1

3a

0.20

0.066

0.59

6.22

06-0990

204.52

206.25

1.73

3

2.92

0.048

0.37

1.46

06-0991

141.6

144.4

2.8

2

2.22

0.135

0.75

5.33

06-0991

200

204.58

4.58

3a

0.09

0.018

0.28

1.88

06-0991

206.39

207.69

1.3

3

1.57

0.065

0.51

2.42

06-0992

32.8

37.4

4.6

3

2.09

0.038

0.53

3.84

06-0993

34.1

40.03

5.93

3

0.89

0.036

0.24

2.19

06-0994

45.34

46.02

0.68

3a

0.01

0.015

0.18

0.24

06-0994

46.77

49.41

2.64

3

2.59

0.107

0.43

5.02

06-0995

38

38.35

0.35

3a

0.08

0.017

0.36

3.49

06-0995

38.8

40.8

2

3

2.30

0.013

0.22

0.17

06-0996

110.41

112.67

2.26

0a

0.01

0.003

0.14

0.93

06-0996

116.6

119.04

2.44

0

0.04

0.014

0.85

0.97

06-0996

152

159.56

7.56

1

1.34

0.081

0.64

3.71

06-0996

230.85

233.95

3.1

2

0.58

0.103

0.94

8.75

06-0997

125.36

128.1

2.74

1

0.15

0.037

0.74

4.36

06-0997

166.58

172.2

5.62

2

0.03

0.018

0.27

1.30

06-0998

51.35

52.9

1.55

0a

0.01

0.007

0.64

1.07

06-0998

54.35

56.45

2.1

0

0.01

0.010

0.66

0.97

06-0998

89.5

91.15

1.65

1

1.01

0.119

0.64

2.88

06-0999

116

116.8

0.8

0a

0.02

0.013

0.83

1.93

06-0999

119.5

121.55

2.05

0

0.01

0.008

0.34

0.58

06-0999

168.9

170.7

1.8

1

0.13

0.093

0.73

7.20

06-1000

111.75

113.2

1.45

0a

0.01

0.007

0.28

1.03

06-1000

115.45

117.3

1.85

0

0.06

0.009

0.61

0.62

06-1000

149.3

151.16

1.86

1

2.03

0.067

0.47

3.76

06-1001

114.89

118.04

3.15

2

0.87

0.108

0.51

2.97

06-1001

172

175.01

3.01

3

0.54

0.084

0.75

5.22



 

 

 

ZAB406-00233:EM15.02.NGI

31 January 2007

Page 41 of 188




[bajatechrep001.jpg]

EL BOLEO PROJECT

MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT




Hole ID

From

To

Thickness

Manto

Cu%

Co%

Zn%

Mn%

06-1002

173.56

175.39

1.83

2

0.32

0.058

0.57

4.73

06-1002

226.6

227.97

1.37

3a

0.02

0.023

0.66

4.04

06-1002

228.71

230.1

1.39

3

2.49

0.046

0.43

1.53

06-1003

139.45

141.26

1.81

1

0.02

0.016

0.86

3.42

06-1003

218.9

220.31

1.41

2

0.23

0.033

0.77

4.58

06-1003

268.44

269.78

1.34

3a

0.02

0.021

0.55

4.74

06-1003

272.6

273.75

1.15

3

3.11

0.051

0.36

1.77

06-1004

119.19

121.53

2.34

1

0.27

0.018

0.59

5.08

06-1004

190.69

194.55

3.86

2

0.09

0.020

0.48

1.05

06-1004

237.72

244.4

6.68

3

0.14

0.016

0.30

0.20

06-1004

284.54

293.58

9.04

4

0.18

0.023

0.06

2.84

06-1005

137.64

147.2

9.56

2

0.18

0.022

0.34

1.32

06-1005

202.74

205.11

2.37

3a

0.03

0.027

0.57

3.76

06-1006

184.65

187.7

3.05

2

0.13

0.043

0.49

3.13

06-1006

239.55

241.35

1.8

3a

0.06

0.025

0.42

3.31

06-1007

119.5

120.37

0.87

0

0.01

0.009

3.32

5.24

06-1007

145.2

150.38

5.18

1

0.02

0.012

0.40

3.01

06-1007

222.69

227.65

4.96

2

0.10

0.023

0.76

3.24

06-1007

265.92

266.67

0.75

3a

0.20

0.033

0.60

4.91

06-1007

268.5

270.87

2.37

3

0.27

0.027

0.21

0.68

06-1008

96.25

100.45

4.2

0

0.01

0.007

0.90

1.93

06-1008

137.31

140.15

2.84

1

0.08

0.032

1.09

3.37

06-1008

155.38

157.91

2.53

2

0.31

0.047

0.81

5.23

06-1008

175.05

177.94

2.89

3

0.00

0.018

0.67

3.82

06-1009

117.69

118.67

0.98

0a

0.01

0.006

0.33

2.03

06-1009

121.14

121.71

0.57

0

0.02

0.006

0.82

0.97

06-1009

149.03

152.68

3.65

1

0.72

0.048

0.56

1.20

06-1009

194.7

198.23

3.53

2

0.58

0.126

0.43

5.80

06-1009

220.16

223.12

2.96

3

0.22

0.021

0.43

0.43

06-1010

109.7

113

3.3

0

0.08

0.009

0.67

0.96

06-1010

146.9

148.4

1.5

1

1.23

0.061

0.35

2.28

06-1010

215.7

217.85

2.15

2

0.24

0.085

0.66

7.31

06-1011

125.53

126.67

1.14

1

0.11

0.007

0.32

0.81

06-1011

178.42

180.25

1.83

2

0.10

0.067

1.09

15.99

06-1011

191.95

193.8

1.85

3

0.21

0.028

0.21

0.57

06-1012

92.86

96.35

3.49

0

0.04

0.006

0.41

0.97

06-1012

128.1

131.7

3.6

1

1.35

0.071

0.31

2.71

06-1013

36.44

37.53

1.09

1

0.01

0.004

0.03

0.10

06-1013

76.86

78.4

1.54

2

0.45

0.041

0.61

2.93

06-1013

90.85

96.47

5.62

3a

0.17

0.031

0.20

0.98

06-1013

96.47

99.95

3.48

3

0.50

0.052

0.22

1.53

06-1014

74.79

77.56

2.77

2

0.27

0.058

1.20

8.34

06-1014

116.11

118.5

2.39

3a

0.05

0.031

0.38

2.68

06-1014

119.1

122.08

2.98

3

0.16

0.023

0.19

0.88



 

 

 

ZAB406-00233:EM15.02.NGI

31 January 2007

Page 42 of 188




[bajatechrep001.jpg]

EL BOLEO PROJECT

MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT




Hole ID

From

To

Thickness

Manto

Cu%

Co%

Zn%

Mn%

06-1015

84.7

89.45

4.75

2

0.09

0.026

0.34

1.51

06-1015

137.29

140.93

3.64

3a

0.17

0.017

0.28

1.90

06-1015

140.93

142.66

1.73

3

5.91

0.068

0.51

5.35

06-1016

60.35

63.4

3.05

2

0.30

0.074

2.22

11.05

06-1016

87.65

89.87

2.22

3

0.47

0.062

0.14

0.62

06-1016

119.22

121.06

1.84

4

0.55

0.021

0.14

0.43

06-1017

24.45

25.93

1.48

2

0.11

0.006

0.08

0.73

06-1017

60.4

61.74

1.34

3a

0.23

0.050

0.52

4.65

06-1017

62.32

65.68

3.36

3

1.27

0.095

0.26

2.17

06-1018

54.75

60.36

5.61

3

2.23

0.054

0.33

3.54

06-1019

66

70.05

4.05

2

0.10

0.080

0.68

9.07

06-1019

83.64

85.45

1.81

3

0.24

0.027

0.27

0.62

06-1019

87.3

87.5

0.2

4

3.25

0.352

0.01

0.89

06-1020

96.15

97.18

1.03

0

0.08

0.007

1.12

3.68

06-1020

111.67

112.68

1.01

1

0.23

0.008

0.54

4.21

06-1020

146.52

150.17

3.65

2

0.16

0.038

0.69

2.39

06-1020

175.6

178.75

3.15

3

0.01

0.016

0.78

5.99

06-1021

71.88

74.47

2.59

2

0.25

0.062

0.88

3.84

06-1021

94.35

95.35

1

3a

0.46

0.025

0.43

2.26

06-1021

96

99.8

3.8

3

1.06

0.088

0.32

1.81

06-1022

61.89

64.77

2.88

1

0.73

0.069

1.03

2.26

06-1022

111.44

116.18

4.74

2

1.08

0.140

1.65

5.29

06-1023

66.45

68.71

2.26

1

2.52

0.089

0.86

2.57

06-1023

132.96

134.19

1.23

2

0.13

0.080

1.24

8.93

06-1024

39.25

43.24

3.99

0

0.01

0.007

1.05

2.53

06-1024

88.8

92.14

3.34

1

2.15

0.063

0.73

1.79

06-1024

135.32

137.48

2.16

2

0.12

0.043

2.96

10.74

06-1025

176.65

178.56

1.91

0a

0.01

0.005

0.22

1.23

06-1025

180.98

182.3

1.32

0

0.08

0.019

0.69

1.47

06-1025

226.92

228.52

1.6

1

0.99

0.112

1.63

5.20

06-1026

169.65

170.77

1.12

0a

0.02

0.008

0.86

2.29

06-1026

171.65

176.46

4.81

0

0.01

0.007

0.24

0.44

06-1026

221.42

223.53

2.11

1

0.86

0.147

1.25

4.62

06-1027

129.73

131.25

1.52

0a

0.03

0.005

0.47

1.43

06-1027

133.73

136.44

2.71

0

0.01

0.008

0.63

0.54

06-1027

182.4

186

3.6

1

1.29

0.051

0.67

1.18

06-1027

253.1

254.85

1.75

2

0.11

0.070

0.58

5.74

06-1028

105.75

108.85

3.1

0

0.04

0.004

0.79

0.57

06-1028

158.08

159.8

1.72

1

1.93

0.153

0.52

0.37

06-1029

125.2

127.05

1.85

0a

0.02

0.004

0.09

0.86

06-1029

130.65

132

1.35

0

0.03

0.011

0.60

0.93

06-1029

184.75

186.06

1.31

1

1.44

0.141

0.42

0.84

06-1029

233.85

235

1.15

2

0.16

0.086

1.06

10.73

06-1030

15.3

17.76

2.46

2

0.34

0.052

0.72

6.11



 

 

 

ZAB406-00233:EM15.02.NGI

31 January 2007

Page 43 of 188




[bajatechrep001.jpg]

EL BOLEO PROJECT

MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT




Hole ID

From

To

Thickness

Manto

Cu%

Co%

Zn%

Mn%

06-1030

54.56

55.35

0.79

3a

0.08

0.067

0.76

5.40

06-1030

58.57

60.37

1.8

3

1.05

0.052

0.31

2.09

06-1031

45.83

46.63

0.8

3a

0.78

0.077

0.61

7.82

06-1031

49.26

52.08

2.82

3

2.65

0.069

0.20

1.59

06-1032

43.57

45.06

1.49

3a

0.05

0.014

0.30

2.02

06-1032

47.45

49.7

2.25

3

3.54

0.075

0.43

4.26

06-1033

34.29

41.4

7.11

3

0.32

0.020

0.25

0.64

06-1034

42.44

47.82

5.38

3

0.56

0.030

0.34

1.86

06-1035

30.5

31.83

1.33

3a

0.68

0.087

0.68

6.97

06-1035

35.23

41.1

5.87

3

1.73

0.073

0.17

0.04

06-1036

48.11

49.68

1.57

3a

0.08

0.038

2.40

6.93

06-1036

50.59

52.33

1.74

3

2.36

0.178

0.39

4.42

06-1037

35.9

40.15

4.25

0

0.05

0.007

0.91

3.07

06-1037

65.61

68.44

2.83

1

0.10

0.014

0.95

6.39

06-1037

99

101.25

2.25

2

0.35

0.059

1.16

5.75

06-1038

75.77

78.33

2.56

1

0.12

0.018

0.77

7.89

06-1038

89.1

92.14

3.04

2

0.78

0.070

1.34

6.97

06-1038

121.35

124.3

2.95

3

0.01

0.019

1.20

10.55

06-1039

108.5

110.64

2.14

1

1.11

0.062

0.57

3.69

06-1040

40.93

43.15

2.22

0a

0.01

0.005

0.25

1.58

06-1040

44.45

48.25

3.8

0

0.01

0.005

0.30

1.04

06-1040

100.45

103.3

2.85

1

0.77

0.025

0.60

1.61

06-1040

145.5

148.49

2.99

2

0.42

0.070

0.41

4.61

06-1041

61.87

63.98

2.11

0a

0.04

0.007

0.27

1.10

06-1041

66.85

70.1

3.25

0

0.03

0.008

0.26

0.91

06-1041

119.01

120.85

1.84

1

1.97

0.107

1.12

3.15

06-1041

177.53

178.95

1.42

2

0.10

0.083

0.87

6.30

06-1042

45.4

47.7

2.3

0a

0.01

0.004

0.23

0.57

06-1042

51.05

53.2

2.15

0

0.03

0.016

0.47

0.76

06-1042

103.23

105.4

2.17

1

2.00

0.160

0.72

3.33

06-1042

157.2

157.8

0.6

2

0.02

0.114

0.78

9.40

06-1043

54.64

55.99

1.35

0a

0.02

0.006

0.07

0.89

06-1043

57.15

61.08

3.93

0

0.01

0.005

0.52

0.47

06-1043

97.16

100.14

2.98

1

1.52

0.128

1.13

1.61

06-1043

114.15

117.57

3.42

2

1.02

0.099

2.88

8.52

06-1044

68.63

71.25

2.62

0a

0.01

0.007

0.53

1.21

06-1044

74.4

75.65

1.25

0

0.01

0.012

0.28

0.90

06-1044

107.7

109.15

1.45

1

2.48

0.106

0.65

5.89

06-1044

169.05

171.45

2.4

2

0.38

0.092

0.82

5.72

06-1045

78.7

80.05

1.35

0a

0.02

0.006

0.30

1.04

06-1045

81.65

84.53

2.88

0

0.02

0.006

0.41

0.88

06-1045

123.75

124.85

1.1

1

1.91

0.122

1.06

6.96

06-1046

39.88

42.1

2.22

1

0.08

0.043

1.24

7.84

06-1046

63.7

65.65

1.95

2

0.29

0.100

1.06

6.61



 

 

 

ZAB406-00233:EM15.02.NGI

31 January 2007

Page 44 of 188




[bajatechrep001.jpg]

EL BOLEO PROJECT

MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT




Hole ID

From

To

Thickness

Manto

Cu%

Co%

Zn%

Mn%

06-1046

93.1

94.5

1.4

3

0.01

0.024

0.47

0.16

06-1046

100

103.9

3.9

4

0.42

0.072

0.40

4.89

06-1047

35.89

37.53

1.64

0a

0.01

0.004

0.09

0.94

06-1047

40.65

42.9

2.25

0

0.02

0.010

0.77

1.43

06-1047

82.85

86

3.15

1

1.08

0.086

1.77

2.18

06-1048

36.3

47.55

11.25

0

0.02

0.006

0.41

1.42

06-1048

86.05

87.76

1.71

1

0.03

0.044

2.52

7.59

06-1048

129.32

131.25

1.93

2

0.60

0.055

0.99

5.35

06-1049

73.35

74.4

1.05

0a

0.01

0.004

0.18

0.46

06-1049

75.4

78.7

3.3

0

0.03

0.007

0.35

0.85

06-1049

122.7

125.55

2.85

1

0.10

0.037

0.59

1.33

06-1049

178.28

181.25

2.97

2

0.03

0.029

0.32

2.84

06-1050

70.1

71.2

1.1

0a

0.02

0.006

0.08

0.76

06-1050

72.35

74.85

2.5

0

0.03

0.009

0.58

0.84

06-1050

113.8

116.4

2.6

1

0.91

0.082

0.72

3.12

06-1050

169.65

172.9

3.25

2

0.42

0.091

0.50

7.09

06-1051

37

37.37

0.37

3a

0.01

0.018

0.49

0.10

06-1051

38.23

41.52

3.29

3

0.82

0.026

0.45

2.14

06-1052

40.55

41.3

0.75

3a

0.04

0.017

0.21

0.16

06-1052

42.23

42.7

0.47

3

0.18

0.018

0.28

0.53

06-1053

44.52

45.03

0.51

3a

0.01

0.016

0.29

0.16

06-1053

46.63

49

2.37

3

0.94

0.051

0.62

5.21

06-1054

45.4

46.71

1.31

3a

0.08

0.021

0.47

3.91

06-1054

47.8

50.15

2.35

3

3.50

0.064

0.21

1.33

06-1055

3.25

8.2

4.95

2

0.81

0.039

0.37

3.50

06-1055

64.78

66.55

1.77

3

0.37

0.023

0.26

0.16

06-1056

49.61

53.95

4.34

2

0.16

0.051

1.45

6.31

06-1056

85.4

90.28

4.88

3

0.09

0.024

0.44

1.32

06-1057

77.25

80.2

2.95

0

0.03

0.018

0.81

1.68

06-1057

109.66

113.66

4

1

0.20

0.019

0.76

1.73

06-1057

162.88

165.78

2.9

2

0.31

0.059

1.02

9.05

06-1057

198.77

200.6

1.83

3

0.44

0.042

0.56

6.36

06-1057

206.97

209.94

2.97

4

0.04

0.033

0.39

2.11

06-1058

91.5

92.95

1.45

0a

0.01

0.007

0.25

3.06

06-1058

97.1

98.9

1.8

0

0.01

0.014

1.33

2.77

06-1058

133.15

135.4

2.25

1

0.04

0.032

1.73

8.64

06-1058

199.88

203.19

3.31

2

0.22

0.042

3.05

7.00

06-1058

244.61

251.65

7.04

3

0.07

0.014

0.31

1.10

06-1058

259.65

263.2

3.55

4

0.06

0.018

0.15

2.80

06-1059

107.65

108.51

0.86

0a

0.01

0.008

0.38

2.27

06-1059

112.01

113.92

1.91

0

0.01

0.023

0.90

1.36

06-1059

145.06

151.81

6.75

1

0.04

0.040

1.22

4.24

06-1059

219.55

221.47

1.92

2

0.42

0.130

0.38

2.10

06-1059

240.81

243.3

2.49

3

0.10

0.017

0.23

0.52



 

 

 

ZAB406-00233:EM15.02.NGI

31 January 2007

Page 45 of 188




[bajatechrep001.jpg]

EL BOLEO PROJECT

MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT




Hole ID

From

To

Thickness

Manto

Cu%

Co%

Zn%

Mn%

06-1060

150.21

152.27

2.06

0a

0.01

0.003

0.54

0.33

06-1060

154.97

156.73

1.76

0

0.01

0.009

0.46

0.93

06-1060

190.33

193.17

2.84

1

0.53

0.026

0.20

1.16

06-1060

256.39

257.7

1.31

2

0.19

0.065

0.87

8.80

06-1061

160.53

164.06

3.53

1

0.41

0.033

0.79

1.65

06-1061

197.92

199.72

1.8

2

0.67

0.079

0.65

6.35

06-1062

107.05

109.31

2.26

0

0.09

0.016

1.38

1.98

06-1062

127.77

132.01

4.24

1

0.05

0.012

0.57

3.53

06-1062

166.95

169.24

2.29

2

0.31

0.060

1.50

10.46

06-1063

86.7

88.99

2.29

0

0.01

0.003

0.04

0.89

06-1063

93.95

96.66

2.71

1

0.05

0.008

1.09

4.24

06-1063

102.4

108.7

6.3

2

0.16

0.020

0.39

3.48

06-1064

100.29

101.42

1.13

0a

0.01

0.005

0.17

3.59

06-1064

103.7

105.45

1.75

0

0.02

0.008

1.81

3.13

06-1064

125.27

126.78

1.51

1

0.14

0.022

0.34

1.63

06-1065

85.63

86.97

1.34

0a

0.01

0.005

0.06

2.59

06-1065

93.5

96.3

2.8

0

0.01

0.007

1.20

3.45

06-1065

115.48

117.15

1.67

1

0.25

0.012

0.95

5.63

06-1065

142.21

145.58

3.37

2

3.24

0.045

0.65

5.23

06-1066

42.7

47.64

4.94

0

0.01

0.005

0.34

1.78

06-1066

80.67

85.28

4.61

1

2.07

0.057

0.23

0.83

06-1066

89.42

92.8

3.38

2

0.12

0.021

0.42

2.84

06-1066

117.67

121.84

4.17

3

0.32

0.031

0.48

2.85

06-1067

53.3

54.45

1.15

0a

0.01

0.002

0.19

0.85

06-1067

60.1

61.45

1.35

0

0.33

0.023

0.37

1.82

06-1067

66.7

69

2.3

1

0.00

0.001

0.06

2.01

06-1067

87

91

4

2

0.49

0.024

2.21

5.12

06-1067

102.9

109.7

6.8

3

0.22

0.040

0.70

5.99

06-1068

43

45.5

2.5

0

0.02

0.005

0.15

0.79

06-1068

46.62

50.8

4.18

1

0.01

0.011

0.59

0.72

06-1068

54.3

58.75

4.45

2

0.43

0.016

0.63

5.91

06-1069

42.3

45.4

3.1

0

0.01

0.003

0.07

0.60

06-1069

46.6

52.3

5.7

1

0.01

0.006

0.29

0.59

06-1069

71.7

73.75

2.05

2

0.56

0.027

1.16

2.16

06-1070

44.6

47.16

2.56

0a

0.01

0.007

0.32

2.14

06-1070

50.55

52.55

2

0

0.01

0.008

0.75

1.61

06-1070

98.5

102.57

4.07

1

0.15

0.034

0.44

1.42

06-1070

141.5

147.5

6

2

0.57

0.038

0.19

0.99

06-1071

109

116.2

7.2

1

0.03

0.021

0.97

2.79

06-1071

154.2

159.1

4.9

2

0.13

0.040

0.51

2.50

06-1071

195

202.45

7.45

3

0.03

0.027

0.41

2.31

06-1071

206.2

210.83

4.63

4

0.02

0.046

0.48

4.67

06-1072

115.9

127.7

11.8

1

0.02

0.008

0.29

1.34

06-1072

187.25

189.87

2.62

2

2.86

0.075

0.74

1.41



 

 

 

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Hole ID

From

To

Thickness

Manto

Cu%

Co%

Zn%

Mn%

06-1072

248.9

255.9

7

3

1.81

0.078

0.57

1.95

06-1072

258.95

272.95

14

4

0.06

0.051

0.41

5.78

06-1073

121.85

122.93

1.08

0a

0.01

0.006

0.63

1.96

06-1073

128.14

129.39

1.25

0

0.01

0.009

0.88

0.73

06-1073

161.59

163.45

1.86

1

5.89

0.053

0.19

1.18

06-1073

184.15

184.35

0.2

2

0.12

0.032

0.53

6.32

06-1073

204.35

204.82

0.47

3a

0.01

0.025

0.36

0.50

06-1073

205.7

208.65

2.95

3

0.24

0.026

0.32

1.53

06-1073

210.3

214.8

4.5

4

0.03

0.042

0.30

3.66

06-1074

142.03

145

2.97

0a

0.00

0.003

0.03

0.19

06-1074

145.1

146.7

1.6

0

0.05

0.006

0.08

0.41

06-1074

161.55

163.7

2.15

1

0.01

0.004

0.04

0.28

06-1074

187.9

192.07

4.17

2

0.13

0.043

0.39

5.65

06-1074

210.9

214.57

3.67

3

1.73

0.157

0.29

4.64

06-1075

123.4

126.23

2.83

0

0.01

0.008

0.57

0.81

06-1075

156.5

159.4

2.9

1

0.18

0.020

0.27

1.60

06-1076

13.85

15.88

2.03

1

0.11

0.021

0.83

2.21

06-1076

80.4

83.05

2.65

2

0.11

0.032

0.69

2.01

06-1076

131.57

136.1

4.53

3

0.19

0.018

0.24

0.26

06-1077

22

30.75

8.75

1

0.04

0.012

0.92

3.41

06-1077

88.95

92.7

3.75

2

0.15

0.015

0.28

0.91

06-1077

144.78

148.5

3.72

3

0.55

0.010

0.20

0.31

06-1078

177.5

179.26

1.76

1a

0.01

0.008

0.45

0.80

06-1078

193.87

195.45

1.58

1

3.70

0.165

0.85

3.50

06-1079

133.25

136.1

2.85

1

0.83

0.071

0.37

1.26

06-1079

202.35

204.47

2.12

2

0.28

0.092

0.59

10.21

06-1079

218.45

220.8

2.35

3

1.74

0.042

0.21

1.94

06-1080

146.9

148.35

1.45

0a

0.02

0.006

0.25

0.95

06-1080

181.9

185.5

3.6

1

0.14

0.020

0.37

1.38

06-1080

243.7

247.2

3.5

3

0.63

0.020

0.21

0.77

06-1081

82.93

87.2

4.27

0

0.02

0.007

0.23

0.88

06-1081

122.17

124.64

2.47

1

1.22

0.087

0.60

3.22

06-1081

187.25

189.63

2.38

2

0.31

0.088

0.59

9.57

06-1082

83.36

87.25

3.89

0

0.01

0.006

0.35

0.69

06-1082

120.73

123.3

2.57

1

1.22

0.099

1.01

4.33

06-1082

185.6

191.15

5.55

2

0.02

0.021

0.36

3.06



 

 

 

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12

SAMPLING METHODS AND APPROACH

 

 



All samples used in the resource estimation are from diamond drill holes.  Down hole sample intervals vary as intervals were selected on the basis of geology.  A total of 12,675 samples have been collected from the mineralized manto units and the mean sample interval was 0.95m.

Historical drilling activities prior to 2004 have not been observed.  Sampling procedures (adopted by Curator) comprised:

·

Core was transported from the drill site by either helpers or geologists to company warehouses in Santa Rosalia, where the boxes were labelled and core recoveries calculated.

·

The core was then logged by a company geologist who simultaneously marked out all sample intervals.

·

Core was split with a mechanical splitter (or a knife in poorly consolidated material) by a trained local helper.

·

Core logging was based on geological intervals with detailed written descriptions for each interval.  Mineralogical, structural and textural information was not recorded in dedicated fields, making it difficult to extract anything other than summary data from the logs.  Logged geology intervals did not always correspond to assay intervals and in many cases overlaps exist between geology and assay intervals.


Drilling rigs were inspected during the recent drilling program in May 2006.

Procedures for drill site supervision and handling of core for this program were as follows:

·

Drilling Supervision.  At the time of manto intersection geologists were on-site to take notes on core recovery and the conditions of the core.

·

Transport of the core.  On completion of a hole the core boxes with lids are loaded into a pick up and secured with a rope to avoid core losses during transport.

·

Layout of core for logging.  In the core shack the core is placed in sequential order on benches.  The benches are designed for ease of use and are well lit.

·

Calculation of core recovery and RQD.  Measurement of the length of recovered core between each drilling interval is used to calculate the percent recovery of the core.  The total length of core that occurs in individual core pieces greater than 10 centimetres in length is measured in order to calculate the RQD.

·

Logging procedures. The core is logged based on geological criteria such as lithology, formation number and members, paying special attention to subdivisions in the mineralized zones.  Geologists complete a fully descriptive written log of each interval.  This descriptive log is then computer coded for entry into a geological database.  Each sample interval corresponds exactly to an assay interval.



 

 

 

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·

Sample selection.  Sample selection is carried out in the mineralized manto intervals with a black marker pen indicating where each sample initiates and ends and showing arrows that indicate the interval to be sampled. The length of the samples varies depending on the type of material being sampled. The maximum length for individual samples does not exceed 1.5 m.

·

Numbering of samples.  The numbering sequence used for sampling is continuous and ascending with respect to depth in a specific drill hole. Numbering control is recorded in a sampling book, sample numbers are also recorded on the full log and the computer coded log sheets.  Assay Quality Control items such as standards, blanks, and pulp duplicates are inserted at this stage.

·

Sampling of the Core.  Splitting of the core is completed using a mechanical splitter.  One half of each sample interval is sent to the laboratory for assaying the other half is retained in the original core box.


The use of a saw is generally better than mechanical splitting in that it cuts through grains and particles, so that some material is collected in the sample whilst the rest is retained in the reject half.  Where a splitter is used the split surface is typically uneven, with breaks that do not exactly honour the desired half core proportion.  Particles along a break can be left intact and are either taken 100% into the sample or left behind in the reject.  This can lead to increased sampling error and variability in the resulting assay data.  In extreme circumstances this may result in sampling induced bias in the assay results.

During two site visits in 2006 Dr B Yeo was able to observe the core splitting process using the mechanical splitter as well as similar material being ‘cut’ using a core saw as described above.

It was concluded that the use of the core saw was inappropriate because typical manto ores are soft and friable.  The cut core disintegrated immediately and created greater problems with regard to collection of representative material than caused by the mechanical splitter.

The mechanical splitter was quick and easy to use but comprised a very short cutting blade that required breaking of the core prior to splitting, and during use the material often crumbled making it difficult to take an accurate half core sample (see Figure 10).  

Figure 10:

Mechanical Core Splitter

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The process of allocating material to the assay sample or to the retained portion at times appeared confusing and a little arbitrary.  As a result a new mechanical splitter was designed by H&S and MMB personnel.  

This initially comprised a wooden holder, fixed to a work bench, with a closing lid.  An entire core box length of core was then placed in the holder, lid closed and a metal blade was guided through a slot in the lid.  This blade was then hammered through the entire core interval.  The blasé was left in place whilst all material from one side of the blade was removed to a sample bag, either by hand picking large fragments or brushing fines.  This produced a very clean split, with all material from one half of core being easily collected.  The concept can be appreciated by reference to Figure 11.

Figure 11:

Prototype Mechanical Splitter

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[bajatechrep018.jpg]

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A more robust metal version of the device was subsequently built (see Figure 12) that was similar in concept but had the blade attached to a screw similar to the original splitter, rather than requiring hammering to break the core.


Figure 12:

Final Mechanical Splitter Design

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12.1

CORE RECOVERY


Core recovery from the historical drill holes was measured for each retrieved core run.  Measured recoveries range from 82-90%, with a mean of 83% (Table 7).  Where recovery was less than 100% no attempt was made to specifically identify location of lost core within the recovered interval.  

Table 7:

Diamond Core Recovery – Historical

Manto

Intervals

Recovery %

0

49

90

1

128

87

2

381

83

3

772

82

3a

564

84

3aa

50

88

4

141

84

Total

2,085

83%


Core Recovery for the current drill program (DDH928-1082) shows better overall values with a mean recovery of 94%.  See Table 8 below.




 

 

 

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Table 8:

Diamond Core Recovery – Current Programs (DDH928-1082)

Manto

Intervals

Recovery %

0

270

94

1

356

95

2

388

96

3

346

93

3a

75

95

3aa

 

 

4

313

95

Total

1,748

94%


For the historical holes, the core run intervals over which recoveries were determined do not correspond to specific assay intervals so it is not possible to evaluate whether a relationship, of any sort, exists between recovery and Cu or Co grade.  

For the current drill holes recoveries have been determined for each sample interval.  Figure 13 below shows core recovery against copper grade.  There is no strong correlation between recovery and grade, although lower recoveries tend to be lower grade suggesting possible loss of friable mineralization.


Figure 13:

Core Recovery vs. Copper Grade- Current Drill Program

[bajatechrep028.gif]



12.2

DENSITY DETERMINATIONS


Data and methodology used to determine density were taken from an internal company report (Felix 1996).  Measurements of specific gravity were taken during the exploration drilling



 

 

 

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campaigns of 1995 and 1996.  A total of 2,112 measurements were obtained of which 418 volumes were determined by a dimensional method (discussed below) and 1,694 by water immersion.  

The dimensional method was used to obtain volume determinations in 1995 and the first weeks of 1996, after which only the water immersion method was used.  Approximately a hundred samples were measured for specific gravity using both methods.  The difference between the results averaged 0.4%.  The samples were taken from both mantos (995 determinations) and non mineralized rocks (1,117 determinations).  

Densities were determined on wet or undried samples.  Dry density for mineralized Manto samples was calculated from the wet density by factoring the density value by the proportion of water content lost during sample drying.  To determine water content, samples were dried at 110º C for a minimum of 6 hours.  Water content was calculated, as the weight loss on drying, as a percentage of the original sample weight, by SGS-XRAL Laboratories in Hermosillo.


12.2.1

DIMENSIONAL (OR CALLIPER) METHOD


The core sample was cut with a saw perpendicular to its axis; samples were nominally 30cm in length.  A tape was used to measure the diameter of the two ends, with two measurements taken at each end with the average of the four taken to establish the diameter of the sample.  In the same way, the length of the core was measured with the same tape in two directions and the average taken as the length.  The volume is calculated using the formula for the volume of a cylinder and bulk density determined dividing the weight by the volume after adjusting for core recovery.


12.2.2

IMMERSION METHOD



The immersion is based on the Archimedes Principle, which states that the density of an object is equal to the weight divided by the difference between the weight in air and the weight in water.

In the laboratory, the samples were sprayed with lacquer to form a thin and uniform coat.  The lacquer is used to prevent water being absorbed by the sample.  The sprayed sample was allowed to dry a few minutes and was weighed in normal (in air) conditions and then weighed submerged in water.


12.2.3

 RESULTS


For the purposes of resource calculations, a global in situ dry bulk density of 1.41 tonnes per cubic meter has been used for manto material, based on an average calculated from 995 wet density measurements and the corresponding analysed water contents of the samples (Table 9).

Although there is some indication of increasing bulk density with sample depth, a graph of bulk density vs. depth (Figure 14) shows only a weak correlation.



 

 

 

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Table 9:

Bulk Density Summary by Manto

Manto

Wet (Natural)

Dry

Bulk Density

Data

H2O%

Bulk Density

Data

Tbcu0

1.84

53

26.75

1.36

53

Tbcu1

1.88

86

26.15

1.40

86

Tbcu1A

1.90

2

25.53

1.41

2

Tbcu2

1.87

136

26.93

1.38

136

Tbcu2A

1.86

11

26.86

1.36

11

Tbcu3

1.89

543

25.71

1.41

489

Tbcu3A

1.90

133

26.15

1.41

128

Tbcu3AA

1.88

9

26.89

1.38

9

Tbcu4

1.91

78

24.19

1.46

78

Tbcu4A

1.93

3

25.48

1.44

3

All Mantos

1.89

1,054

25.93

1.41

995


Figure 14:

Bulk Density vs. Sample Depth

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13

SAMPLE PREPARATION, ANALYSIS AND SECURITY

 

 



13.1

HISTORICAL DRILL HOLES – PRE 2004


Prior to 1997, assay samples were sent to the SGS-XRAL laboratory in Hermosillo where they were dried, crushed and pulverised.  Analysis for Cu, Co, Zn, Fe and Mn was carried out at the same laboratory, with a perchloric acid digest and AAS finish.  Samples that reported grades greater than 1% of Cu, Co or Zn were re-assayed using a method more appropriate for higher grade samples.

From 1997 onwards sample preparation was carried out at Chemex facilities in Hermosillo with pulps being forwarded to the Chemex laboratory in Vancouver, Canada for analysis.  A four acid digest with AAS finish was used for Cu, Co, Zn, Fe and Mn.

The reject and remaining pulp material were returned to Santa Rosalia where they are securely and systematically stored in warehouses.  

The reason for the change in assaying laboratory and technique was due to the identification of a systematic under-reporting of, primarily, Co by SGS-XRAL.  The extent of the analytical problems and re medial measures taken are discussed in more detail in section 14.1 below.

All samples were stored, prior to shipping, in one of three locked warehouses in Santa Rosalia.  When shipped, the samples were taken to the ferry by company personnel where they were put on the ferry and shipped across the Gulf of California to Guaymas. The representatives of the laboratory picked the samples up and delivered them directly to the laboratory in Hermosillo.


13.2

CURRENT DRILL HOLES


The samples for assaying are placed into 30 kilogram sacks and kept in locked premises until such time as they are transported to the laboratory.  A company vehicle is used to transport the samples from Santa Rosalia to Guaymas via a ferry and then driven directly to Hermosillo, Sonora.  The samples are driven and accompanied by a company employee.

Remaining core is stored in boxes in an underground storage area that is secured by a locked metal door.

The sample shipment is delivered to ALS Chemex located at:

Ignacio Salazar 688 Local 5, Fracc. Los Viñedos
C.P. 83147
Hermosillo, Sonora, México.

The sample preparation protocol used is as follows:



 

 

 

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·

Samples are oven dried at 110 degrees until dry

·

Samples are fine crushed to >70% <2 mm  (TM Rhino or Terminator crusher)

·

Crushed sample is split using a riffle (Jones) splitter to 250 g

·

250 g of crushed material is pulverised (Labtech LM2)>85% <75 µm.


After sample preparation pulps are sent from Hermosillo to ALS Chemex in North Vancouver for assaying.

The assay method currently used for Copper, Cobalt, Zinc and Manganese is ME-ICP61a (four acid “near-total” digestion).  Samples reporting high grades for Cu, Co, Zn or Mn are re-assayed for those elements by method AA62.

Total sulphur has been assayed for using ME-ICP61a with sulphide sulphur determined by either S-IR08, or S-IR07.  Moisture content was determined by OA-GRA05s.



 

 

 

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14

DATA VERIFICATION

 

 


Data verification can be considered as having four separate components:

·

Are the assay results reported for each sample accurate?

·

Are the samples, used in assaying process, representative of the sample interval?

·

Are the reported assay values, which are identified by a unique sample number, assigned to the correct down hole sample intervals.

·

Are the drill hole locations known accurately, and locations correctly entered in the database.


To evaluate accuracy and representativity, a number of quality control samples have to be inserted into the sample stream.  These samples include Standard Reference Materials (SRMs or “standards”) of known grade, blank samples with no grade, and possibly field duplicate samples.  Field duplicates of diamond core samples should be the un-sampled half of core retained after the original round of sampling.  To be most effective, the Quality Control samples need to be anonymous to the assaying laboratory.


14.1

HISTORICAL DATA – PRE 2004


14.1.1

PROBLEMS WITH THE SGS-XRAL ASSAY RESULTS


During the original round of assaying by SGS-XRAL, no SRMs or field duplicate samples were used.  Some internal repeat assays were reported by the laboratory and on quick visual examination these appeared to be adequate, but no statistical analysis was completed to support this.

At a late stage in the process, a group of samples were retrieved and sent to an independent laboratory (Bondar-Clegg, Vancouver) for check analysis.  The results of this check program revealed some considerable departures from the original values, though no standards were included with these samples, so the same samples were forwarded to Chemex Laboratories, also in Vancouver.  The results of these second repeat check assays were again considerably different to both the original SGS-XRAL and Bondar-Clegg check assays, such that sufficient doubt was generated as to the validity of the entire assay database at Boléo.  

Consequently a decision was taken by the company (International Curator Resources Ltd, “Curator”) to review the entire assaying process used for the Boléo samples.

14.1.2

SUMMARY OF 1997 PROCESS REVIEW AND RE-ASSAYING PROGRAM


The review was carried out on behalf of Curator by consultants G. Peatfield and B. Smee.  The process and subsequent re-assay program is reported in detail (Peatfield & Smee 1997,



 

 

 

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Peatfield 1997, Peatfield 1998).  These reports are available and a limited synopsis is included in Sections 14.2.1 to 14.2.6.

Several problems were identified with respect to the original SGS database.  These were:

·

Despite repeated assurances from SGS that the perchloric acid digest would give a total metal extraction, data from check assaying indicated otherwise.

·

Above 4%, Cu SGS assay values showed very poor correspondence with check assays from other laboratories.  It was thought that this was to do with a dilution step in the process but SGS failed to provide a rational explanation.

·

SGS reporting left much to be desired, with issues such as changes in reporting format and numerous data errors.

·

Check assaying strongly suggested that Co values from SGS were reported systematically low, possibly by as much as 15%.


PREPARATION OF ASSAY STANDARDS


The use of standards to monitor assay accuracy was decided on immediately.  It was also deemed essential that these standards should be prepared from Boléo material so that the matrix of the standards matched that of the samples.

Initially two standards (Interim Standards) were prepared from Boléo material (Boléo I, II).  These were prepared at the SGS laboratory.  Sub-samples were dispatched to 7 laboratories for round-robin analysis.  The material was subjected to a 4 acid digest (nitric, hydrochloric, perchloric and hydrochloric), with an AAS finish.  From the data received from each laboratory the mean and relative standard deviations were determined.  The interim standards were used only as a stop-gap measure whilst the full review of the assay process and the preparation of formal standards were completed.

The formal standards (Boléo III, IV, V) were prepared at the Colorado Minerals Research Institute and CDN Resource Laboratories, Burnaby B.C. Canada.  Composites of varying grade were combined to form three single 25kg samples.  Samples were again dispatched to several laboratories in the USA, Canada and to SGS in Mexico for round-robin analysis.  The same 4 acid digest method was used.  The acceptable grade and limits for each standard (Peatfield 1997) are shown in Table 10.

Table 10:

Boléo Assay Standards

 

Copper

Cobalt

Zinc

Standard

mean

+2sd

-2sd

mean

+2sd

-2sd

mean

+2sd

-2sd

Interim Standards

 

 

 

 

 

 

 

 

 

Boléo I

1.546

1.682

1.410

0.0595

0.0640

0.0550

0.34

0.329

0.278

Boléo II

3.813

4.159

3.466

0.0527

0.5677

0.4377

0.355

0.389

0.321

Formal Standards

 

 

 

 

 

 

 

 

 

Boléo III

0.514

0.555

0.472

0.0520

0.0630

0.0490

0.322

0.338

0.306

Boléo IV

1.124

1.202

1.047

0.0934

0.1082

0.0786

0.387

0.415

0.358

Boléo V

7.405

7.900

6.910

0.0986

0.1146

0.0827

0.459

0.459

0.434




 

 

 

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ANALYTICAL METHOD AND ASSAYING LABORATORY


To determine the most appropriate analytical method to use, 12 samples were selected, prepared in replicate pulps and dispatched to several laboratories where they were assayed using four different techniques:

·

aqua-regia digest

·

sodium peroxide fusion

·

perchloric acid/aqua-regia digest

·

four acid digest (nitric, hydrochloric, perchloric, hydrofluoric).


The sodium fusion method tended to report substantially lower grades indicating less complete dissolution.  The four acid method gave the most consistent results although slightly lower than the aqua-regia or perchloric/aqua-regia digests.

The four acid digest was chosen as the analytical technique and Chemex was chosen as the assaying laboratory.  A second laboratory, Mineral Environments Laboratories, was chosen for check assaying.

RE-ASSAY PROGRAM


All pulp and reject material held by SGS in Hermosillo were returned to Santa Rosalia.  Samples were selected for re-assay on the premise that any sample obviously or likely to be used in resource estimations would be re-assayed.  When in doubt samples were included rather than omitted.

Thirty gram sub-samples of existing pulps were prepared and a completely new sample number sequence applied.  Standards, blanks and duplicate pulps were inserted into the sample stream in sequence so that for every 40 samples sent to Chemex would include, two standards, two blanks and two duplicate pulps.  In addition Chemex inserted two of their own standards and a sample blank.  In all about 6,800 samples were re-assayed.  To monitor quality 1,200 internal quality control samples (standards, blanks and duplicates) were inserted by Curator and an additional 800 laboratory samples were inserted by Chemex.

The results of the internal standards and blanks were monitored as data was received by means of control charts, which show the acceptable value and upper/lower limits (± 2s.d.).  No concerning results or trends were identified.  Blanks sample results were monitored by eye and no obvious contamination was seen.  Chemex standards were monitored in a similar way.

Sample duplicate data was analysed using a mathematical technique (Thompson and Howarth 1978) to determine analytical precision.  The results indicated that at grades likely to be mined, levels of precision were very good (± <5%) for both Cu and Co.



 

 

 

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SECOND LABORATORY CHECK ASSAYING


A total of 441 duplicate pulps were sent to a second laboratory (Mineral Environments Laboratories, “Min-En”) for check assays.  Typically results were similar, there was however a clear laboratory bias, with Min-En results reporting lower than Chemex.  

Similar results were seen in the data reported for the standards returned from Min-En.  It was concluded that the check laboratory did not match their performance in the original round-robin tests.

CORE DUPLICATE ASSAYING


As part of the re-assaying program, pulp duplicates were routinely inserted into the sample stream.  However, because they are from the same original sample, they do not provide a means for determining overall precision resulting from all steps in the process, from sampling, sample preparation and assaying.  To do this the remaining half-core has to be sampled and subjected to the same sample preparation and assaying regime.

One hundred samples were selected for core duplicate assaying.  For each of these samples the remaining half-core was collected from the core tray and sent for assay.  Half core was crushed and split into two equal samples.  From each of these a 250 g sub-sample was taken and fine pulverized and a final 30 g sub-sample taken from each and sent to Chemex for assaying.  As a result it was possible to compare the original assay against two identical assays of the remaining core, referred to as ‘A’ and ‘B’ core duplicates.

In addition, duplicate samples were also taken from the coarse residue of the original assay samples retained by the laboratory.

Standards were inserted into the sample stream at a rate of 1 in 20, as done previously.

The results showed, as expected, lower precision between the core duplicates and original samples than for the coarse residue duplicates and the original samples.  Surprisingly though, there were also noticeable differences between the calculated precision from the ‘A’ and ‘B’ core duplicates.  This lead to a decision to include core duplicate sampling programs as part of all future drilling and assaying programs.

Average grades from both the ‘A’ and ‘B’ core duplicates were lower than the original assays by ~10% for Cu, 5% for Co and 2% for Zn.  No explanation was given for this difference.

Table 11 summarises the results of the ‘A’ and ‘B’ core duplicate sampling program.

Subsequently a further 55 core duplicate samples were assayed (Table 12).  Levels of precision clearly improve from core duplicates to coarse residue duplicates to pulp duplicates and, as in the first program, the core duplicate assays show lower average grades than the original assays, although in the second program this difference is less.  



 

 

 

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Table 11:

Boléo Core Duplicate Results – 1st Program

Core
Duplicates

Range
(%)

Ave.
Precision

Original Assay
Ave.

Dupl. Core Assay
Ave.

%
Diff.

Copper

 

 

 

 

 

‘A’

0.10 – 4.78

25.9

1.001

0.895.

-10.6

‘B’

0.10 – 4.78

26.8

1.001

0.903

-9.8

Cobalt

 

 

 

 

 

‘A’

0.010 – 0.0525

20.2

0.063

0.060

-4.8

‘B’

0.010 – 0.0525

20.1

0.063

0.060

-4.8

Zinc

 

 

 

 

 

‘A’

0.2 – 6.69

23.9

0.751

0.736

-2.0

‘B’

0.2 – 6.69

14.2

0.751

0.736

-1.5


One reason for the improved result in the second program may be due to the fact that the duplicates and original samples were assayed in the same batch, whilst in the first program duplicates were submitted in two batches, one at a later date than the other.  

Therefore any problem unique to a batch or prevalent at a laboratory for a limited time can impact the entire program.  The average grades of both the coarse residue and pulp duplicates are almost identical to the original samples.

Table 12:

Boléo Core Duplicate Results – 2nd Program

Duplicate
Type

Range
(%)

Ave.
Precision

Original Assay
Ave.

Dupl. Assay
Ave.

%
Diff.

Copper

 

 

 

 

 

Core

0.10 – 4.38

14.4

1.020

0.957.

-6.2

Coarse Residue

0.10 – 3.34

5.7

0.816

0.811

-0.6

Pulp

0.10 – 3.72

1.7

0.595

0.587

-1.3

Cobalt

 

 

 

 

 

Core

0.010 – 0.206

22.3

0.045

0.044

-2.2

Coarse Residue

0.010 – 0.335

8.8

0.051

0.051

0.0

Pulp

0.010 – 0.187

3.7

0.053

0.052

-1.9

Zinc

 

 

 

 

 

Core

0.20 – 1.30

12.9

0.398

0.386

-3.0

Coarse Residue

0.20 – 0.65

1.5

0.326

0.325

-0.3

Pulp

0.20 – 1.48

1.8

0.441

0.440

-0.2



CONCLUSIONS DERIVED FROM THE RE-ASSAY PROGRAM


The conclusions arrived at by Peatfield & Smee were:

·

The initial assay method used by SGS was unreliable and uncontrolled, and therefore unacceptable.



 

 

 

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·

Tests carried out showed that a multi acid digestion provided the most consistent assay results.

·

Round-Robin assaying showed the original laboratory had difficulty in generating statistically acceptable results.

·

The re-assay program was completed under controlled QC conditions, and provided revised database assay information with acceptable accuracy.

·

Subsequent core duplicate sampling produced relatively high but acceptable levels of precision.  

·

Precision based on reject duplicates showed little difference to pulp duplicate precision, indicating error attributable to sample preparation was minimal.

·

The systematic decrease in metal in the core duplicates is not readily explicable.  The improvement between the 1st and 2nd core duplicate sampling program may be due to the fact that in the 2nd program duplicates were assayed in the same batch as the original samples.

·

Routine monitoring of internal and Chemex standards indicates that there are no recognizable problems with the accuracy of Cu, Co, Zn and Mn assays in the Boléo database.


14.1.3

COMMENTS ON ASSAY QUALITY CONTROL RESULTS


As a result of the problems identified in the original Boléo assay database, the results from the re-assay program have been subjected to a high level of scrutiny.

The preparation of matrix matched standards is commendable and is the best means available to monitor and ensure assay accuracy.

The control charts for each standard (Peatfield & Smee 1997) show that although the Cu results do fall within the acceptable limits, they are mostly higher than the accepted value, rather then being scattered above and below.  Conversely, the standards reported by the second laboratory were also within the acceptable limits but systematically low.  These patterns are consistent, in all cases, with the Chemex value for each standard reported in the round-robin assaying relative to the accepted value (the mean of 7 participating laboratories).  The implication of these results is that the Boléo assay data may be reported marginally high, in the order of 2% to 4%.  

However, H&S consider that a more rigorous evaluation of the round-robin results used in determining the standard values standards should have been applied.  Of the seven participating laboratories, results from six were used, in most cases, to determine the true values (Peatfield & Smee 1997).  H&S believe that using results from fewer laboratories (3 or 4) of superior quality and consistency is best in determining the recommended values.  

Figure 15 shows graphically the Cu data reported by each laboratory used for Boléo V, whilst Figure 16 shows the data used by H&S.  A similar exercise was completed copper and cobalt data for Boléo III and IV the results are compared in Table 13.  




 

 

 

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Figure 15:

Boléo Cu, Round-Robin Cu Assays used by Peatfield & Smee

[bajatechrep032.gif]


Figure 16:

Boléo Cu, Round-Robin Cu Assays used by H&S

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Table 13:

Boléo Formal Standards – Revised Values

 

Copper

Cobalt

Standard

mean

+2sd

-2sd

mean

+2sd

-2sd

Formal Standards – Peatfield & Smee (1997)

Boléo III

0.514

0.555

0.472

0.0520

0.0630

0.0490

Boléo IV

1.124

1.202

1.047

0.0934

0.1082

0.0786

Boléo V

7.405

7.900

6.910

0.0986

0.1146

0.0827

Revised Formal Standards – H&S

Boléo III

0.525

0.530

0.519

0.0511

0.0567

0.0455

Boléo IV

1.135

1.159

1.111

0.0929

0.0988

0.0871

Boléo V

7.579

7.709

7.450

0.0986

0.1045

0.0927


Copper values determined by H&S are slightly higher than Peatfield & Smee, whilst cobalt values are slightly lower.  If revised copper standard values were used, the potential assay bias of 2% to 4% would be reduced to less than 2%.



 

 

 

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The core duplicate sampling programs show a consistent difference in the means of the original assays and the duplicate samples, by as much as 10% for copper.  This may be in part attributable to batch issues but this has not been confirmed.  

Another aspect that can cause differences between reported grades is the method and completeness of the collection of material from the core tray.  It is unlikely that the fines in the bottom of a core tray are identical in make up as the larger segments of core.  If, for example, fine grained chalcocite is dislodged, through handing of core, and collects in the tray it needs to be carefully collected, failure to do this may result in understating copper grade.  

The use of a soft brush is often needed to ensure that all the fines are collected from the bottom of the tray.  Unfortunately, sample collection was not observed.  The 55 duplicate core samples from the second program were dispatched to the laboratory at the same time as the original samples, suggesting that they were bagged immediately on splitting and not returned to the core tray to be re-sampled at a later date.  Yet results for these duplicate samples were still lower than for the original samples.

14.1.4

DATABASE VERIFICATION


To ensure that the database is accurate, in other words, that assays are assigned to the correct sampled interval, an audit of original assay certificates against the database files was carried out.  

The assay data is kept in 10 Microsoft Excel spread sheets (a separate file for each hundred holes).  Two holes were checked from each file (Table 14), with Cu, Co, and Zn assays checked.  Geological logs and summary logs contained in the same Excel files were not audited.

Table 14:

Drill Holes Audited for Data Entry Errors

Hole ID

Assay Job Num.

Hole ID

Assay Job Num.

22

A 9631180

491

A 9627709

57

A 9629734

526

A 9627711

115

A 9634051

592

A 9628562

173

A 9626536

656

A 9626533

226

A 9626538

695

A 9629737

287

A 9626541

746

A 9717119

324

A 9626542

765

A 9718436

361

A 9626544

818

A 9720883

410C

A 9626545

888

A 9737165


No data entry errors were detected.



 

 

 

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14.1.5

DRILL HOLE SURVEYING


Drill hole collars at Boléo have not been surveyed.  In place of conventional surveying, high resolution Orthophotos have been acquired from which drill hole coordinates have been calculated.  To assist in identifying drill holes in the photos each collar was marked with a large white cross, with the hole at the centre.  It is estimated that the accuracy using this method is to sub meter.  

The drill hole spacing, at its closest, is in the order of 150 m x 150 m so sub meter accuracy for the easting and northing coordinates is adequate.  Mantos are generally only a few metres thick so accuracy of the elevation coordinate is more critical.  A tolerance of ±0.65 m has been quoted for this dimension (Albinson pers. comm. 2004).  This has not been verified.  The potential implications of an error in this regard are not considered serious.  

In an open-pit environment, mining is carried out with strict grade control practices that are used to identify ore and waste prior to mining.  In an underground situation geological control, particularly the different lithologies that define the base of each Manto should be adequate to avoid losses due to uncertainty in the collar surveys.

All the holes at Boléo are vertical, which is appropriate.  No down hole surveys were conducted.

No drill holes were intersected during the mining trial.


14.2

CURRENT PROGRAM


Assay standards, duplicate sampling, blank samples and check assaying have been used during the 2006 drilling programs to validate the assay results reported by ALS.

14.2.1

ASSAY STANDARDS


The three Boléo standards (Boléo III, IV, V) discussed above have been used again during the current assaying program.  Results are shown in the figures below.  The reported results are combined and presented as a percentage difference from the accepted value of each standard.



 

 

 

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Figure 17:

Boléo Assay Standards – Cu

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Figure 18:

Boléo Assay Standards – Co

[bajatechrep038.gif]


Figure 19:

Boléo Assay Standards – Zn

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There is a persistent under-stating of standard grades by between 5% and 10% on the 2006 drill program.  The reason for this is not at present well understood but the issue is currently under investigation and review.   

14.2.2

CHECK ASSAYS


Check assays from ACTLabs in Phoenix however show very good correlation to the ALS-Chemex results (Figure 20).  This suggests that the problem may be that the accepted values are not correct.  Additional check assaying of the three Boléo standards is required to confirm this.

Figure 20:

Boléo Check Assay Results – Cu

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Figure 21:

Boléo Check Assay Results – Co

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14.2.3

DUPLICATE ASSAYS


Duplicate assays have been determined from crusher split reject material.  Although the total number of duplicates is low (<50) the comparison is very good (Figures 22 and 23).

Figure 22:

Boléo Duplicate Assay Results – Cu

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Figure 23:

Boléo Duplicate Assay Results – Co

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14.3

PROJECT DATABASE


14.3.1

DATA


A new Microsoft Access database was customised for the Boléo project.  This database was set up to accommodate the historical data and the new data from the current programs.

Historical data exists in the form of 10 Excel spreadsheets, a single spread sheet for each sequential hundred holes.  Each file contained several sheets including a drill hole coordinate sheet, assay sheet and summary geological logging sheet. The data is stored in a number of non-active tables, i.e., no new data is to be added to the tables.  The data was taken as supplied and no modifications were made.

Data from the current drill programs is imported into a number of active tables to which the data is progressively added as it becomes available.  These tables are:

·

Assay Header Table – Dispatch number, sample number sequence, elements assayed, method of assay, detection limits.

·

Assay Table – Sample number, reported results for each sample.

·

Header Table – Drill hole coordinates, hole depth, hole azimuth, dip, drilling dates, drilling type, bit diameter etc.

·

Geology Table – Logging and sampling intervals, sample number, dispatch number, geology logging data.



 

 

 

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Assay files are loaded directly from digital files supplied by the laboratory without any modification or manipulation, therefore avoiding traditional errors from manual data entry.

Original geology logs are hand written in full descriptive text.  These logs are then translated by the relevant geologist into computer coding on paper data entry forms.  The data on these forms is then manually entered into Excel spreadsheets with identical column names and formats.  Finally these Excel forms automatically uploaded into the database.

Data is verified by means of a number of database queries run that detect such errors as mismatched ‘From’ and ‘ To’ intervals, drill hole depth greater than last ‘To’ depth, missing sample numbers, duplicated sample numbers, duplicated sample intervals, missing key geology fields such as formation and lithology.  A list of acceptable codes for each geological field is used to prevent incorrect and inappropriate codes being used.

On completion of all data entry the assay data from the assay table is merged with the geology data from the geology table.  Sample number and dispatch number are used to ensure correct data merging.  Historical data and new data are also combined together at this time to produce two tables that contain data for all holes, which are:

·

Collar and Survey Table – Easting, Northing, Elevation, Total depth, Azimuth and Dip

·

Geology and Assay Table – H ole ID, Interval ‘From and To’, sample number, preferred assays fields, geology descriptors for each sample interval.


14.3.2

DRILL HOLE SURVEYING


Drill hole locations were surveyed using a Total Station instrument (model:  Topcon – Hiper GD/Legacy H L1/L2 RTK).  Accuracy is reported as ±5 mm in the horizontal and 10 mm vertically.

No down hole surveys were carried out.



 

 

 

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15

MINERAL PROCESSING AND METALLURGICAL TESTING

 

 


15.1

METALLURGY


15.1.1

BACKGROUND


Treatment strategies for the Boléo polymetallic mixed oxide/sulphide ore were studied by Fluor in the mid 1990s during the Curator PFS development that resulted in a complex, high capacity flowsheet matching the requirements of a low-grade, ‘super-pit’ design. The flowsheet featured a combination of roasting, leaching, precipitation and metal refining.

Recently, in parallel with the adoption of selective mining of higher grades (and significantly reduced waste mining), Bateman sought to simplify the flowsheet via a more direct approach incorporating leaching, solid-liquid separation, solvent extraction and electrowinning.

Key to the revised process was the successful demonstration of the solid-liquid separation characteristics of the leached clay ore followed by an effective process for dealing with the manganese in the pregnant leach solution (PLS).  These changes in processing strategy result in a more robust, operable flowsheet with reduced operating and capital cost.  

Significantly, the flowsheet has been successfully tested in two separate pilot plant campaigns held at the SGS Lakefield facility in Ontario, Canada.

The metal recovery circuits are typical of those deployed in numerous operations world-wide.  In developing the Boléo flowsheet, Bateman was able to incorporate the results of earlier testwork; supplementing these with recent bench scale and pilot plant testwork results, as well as information from recent Bateman projects featuring similar processes and unit operations.

The following summarises the metallurgical testwork history and goes on to describe the proposed Bateman process flowsheet.

15.1.2

PROOF OF CONCEPT PILOT CAMPAIGN

 

SOLID LIQUID SEPARATION TESTWORK


As a precursor to conducting a proof of concept pilot campaign a bench scale test work campaign was conducted at SGS Lakefield's facility in Ontario.  Batch leach testing was carried out on six different ore samples from the Boléo Reserve.  The samples varied widely in their copper and cobalt grades and were sourced from widely geographically dispersed outcrops of Manto 3.  

The principal objective of the work was to generate leached pulp samples for subsequent settling test work by specialist vendors (Outokumpu) and consultant representatives (Pocock



 

 

 

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Industrial).  The samples were tested for their amenability to High Rate thickening in a series of tests, the success of which was key to the flowsheet development.  

In essence the solid-liquid separation tests conducted at SGS Lakefield Research have demonstrated that Boléo ores can be settled and washed in a conventional CCD circuit utilizing high rate thickeners.  Washing of settled solids to recover metal values from the PLS solution is fundamental to the economic success of the project.

The samples were found to settle best when diluted to between 2.0% and 3.5% solids.  The optimum flocculant dosage for the leach varied from 3 ppm to 6 ppm.  Underflow densities of 20% to 22% were achieved in test work with clear overflows being produced.  

This test work was sufficiently successful to warrant taking the next step in the process development, namely the operation of a 'proof of concept' pilot plant campaign.  

SOLVENT EXTRACTION TESTWORK


The SX Group of the Parker Centre, CSIRO Minerals in Perth, Australia have developed a process that allows for the separation of copper, cobalt and zinc from calcium, magnesium and manganese by solvent extraction with synergistic mixtures of commercially available organic extractants.

The combination of organic extractants results in a large 'synergistic' shift in the pH 50 for the system under consideration, allowing efficient separations of metals in solution to take place.  This system of reagents allows for the separation of copper, cobalt and zinc from manganese, magnesium and calcium.  This technology provides a means of dealing with the high levels of manganese in solution and is both fundamental to and appropriate for metals recovery from Boléo solutions.    

In preliminary test work at pH 4.5, 100% of the copper was extracted, cobalt extraction was in the range of 94% to 98%; zinc extraction was in the range of 64% to 80% and manganese extraction in the range of 0.76% to 1.55%.  In almost all cases, no calcium and magnesium were extracted at pH 4.5.  The separation factors of copper and cobalt over manganese ranged in the ten thousands and thousands, respectively, suggesting an easy and complete separation of copper and cobalt from manganese for all the systems tested.  The separation factors of zinc over manganese ranged in the hundreds, suggesting that a good separation of zinc from manganese can be obtained.

The success of this test work promoted further investigation and after some additional bench scale testing the concept was included in the flowsheet for 'proof of concept' piloting.   

PROOF OF CONCEPT PILOT CAMPAIGN TESTWORK


A 'proof of concept' pilot plant campaign was conducted at the SGS Lakefield facility from November 16 to 28th , 2004 treating a bulk sample of Boléo ore grading 1.6% Cu, 0.087% Co, 0.58% Zn, 3.23% Mn and 8.71% Fe.  The pilot plant flowsheet comprised:



 

 

 

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·

attritioning of the ore with grinding of coarse ore particles to form an ore slurry

·

acid oxidation leaching of the ore with sulphuric acid in seawater

·

acid reduction leaching of the ore with sulphur dioxide and sulphuric acid in seawater

·

partial neutralization with limestone


The picture below shows the oxidative leach, the reductive leach and the partial neutralization steps undertaken at the SGS Lakefield facility in Ontario.

[bajatechrep050.gif]


·

Counter current decantation (CCD) washing of the leach residue in thickeners, to separate the metal rich aqueous solution from the clay waste


The picture below shows 4 of the 6 stage CCD thickeners.  



 

 

 

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[bajatechrep052.gif]

·

Copper solvent extraction/electrowinning


The picture below shows the copper solvent extraction mixers/settlers to the left and the electrowinning process to the right, under a fume extraction hood.  The cathodes can be clearly seen suspended in the electrowinning cell.

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·

Iron removal by pH adjustment and oxidation with air (and polish with hydrogen peroxide)



 

 

 

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·

Thickening of the iron residue prior to disposal;

[bajatechrep056.gif]


·

Direct Solvent Extraction technology ('DSX ®') for selective recovery of cobalt and zinc and small amounts of residual copper.  As mentioned above DSX ® technology is the property of Commonwealth Scientific Industrial Research Organization (CSIRO), Perth, Australia.


The picture below shows the setup of the DSX circuit.

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SGS Lakefield and Bateman Engineering have previously jointly reported the summary findings from the pilot plant.  Highlights of the campaign included the following:

·

A total of nearly 2 metric tonnes of ore were treated through the pilot plant.

·

The pilot plant operated continuously for a total of 12 days in leaching, 11.5 days in CCD, 9.5 days in Copper SX/EW and 9 days in Cobalt and Zinc SX using the DSX ® technology.

·

The oxidation, reduction leaching circuit gave excellent extractions of copper, cobalt and zinc. Copper extraction exceeded 90% during pilot operation.  Cobalt extraction varied from 80% to as high as 90%.  Zinc extraction was generally above 70%.  These numbers are indicative of the potential of the Boléo process to extract the three pay metals copper, cobalt, and zinc.  

·

The CCD circuit was set up to simulate the use of the “high rate” type of thickeners with recirculation of overflow solution to dilute the feed slurry prior to flocculation.  This method of settling and washing was based on recommendations from bench-scale testing by Outokumpu and Pocock Industrial and proved to be highly effective.  The leach residue settled quickly, producing clear overflow solutions to advance to copper, cobalt and zinc recovery.

·

15.5 kg of copper metal were electrowon from the solvent extraction strip solutions at high efficiency.  A picture of the first copper cathode produced during this campaign appears below.  These cathode samples assayed at better than LME grade.

[bajatechrep060.gif]


·

The iron removal circuit was designed to remove iron, aluminium and other impurities from the solution prior to recovery of cobalt and zinc using DSX ® technology. The iron removal circuit consistently produced very low concentrations of key impurities in solution with negligible losses of cobalt and zinc.  

·

The DSX ® circuit for cobalt and zinc recovery performed very well.  The advantage of the DSX ® circuit for Boléo plant design is that cobalt and zinc can be separated from manganese, magnesium and calcium.  In the Lakefield pilot plant, cobalt and zinc were recovered with high overall extraction efficiency (+95%) to produce a concentrated zinc sulphate solution (for subsequent production of zinc sulphate monohydrate crystals) and a concentrated cobalt solution (for subsequent production of cobalt metal cathode).


Two further metallurgical tests were performed on the product streams from the pilot plant.



 

 

 

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·

Production of zinc sulphate monohydrate crystals.  Zinc sulphate monohydrate was produced by evaporative crystallization of the zinc strip solution from the DSX ® circuit.

·

Production of cobalt cathode.  The cobalt strip solution from the DSX ® circuit containing cobalt along with small amounts of zinc and nickel was treated in a zinc solvent extraction circuit followed by a cobalt SX/EW circuit to purify the cobalt solution for electrolysis as high grade (+99.9%) cobalt cathode.  Conventional solvent extraction reagents and process steps were utilized for this purpose.


In addition certain bench scale tests were conducted to obtain additional data for definitive feasibility study purposes including:

·

Environmental testing of residues and solutions produced in the pilot plant program

·

Characterization of High Acid Consuming (HAC) material from the Boléo site containing limestone and other alkali minerals.  HAC material is intended to be used as a low cost neutralizing agent in the commercial Boléo plant

·

Ore scrubbing and grinding testwork for developing final design for ore preparation circuit, including Bond Work Index determination for both ROM ore and HAC material

·

Testwork on oxidation and precipitation of iron from the DSX ® circuit feed solution to ensure maximum removal of iron with minimum treatment time and reagent consumption

·

Leach testing on 24 individual samples of ore that were composited to form the pilot plant feed.  These tests were used to assess leach variability and acid consumption variability of the ore, and the production of an acid consumption model.


The successful completion of this pilot program was an important milestone in moving the Boléo project forward.  For the first time Boléo ores had been treated in a continuous pilot plant program to leach, separate and recover pay metals in final commercial form.

15.1.3

FULLY INTEGRATED PILOT CAMPAIGN


Refinements to the process flowsheet followed in an attempt to optimise the capital and operating costs for the proposed operation as did targeted testwork initiatives designed to support the proposed flowsheet variations.  

Targeted DSX system testwork conducted by Bateman's Solvent Extraction Group in late 2005 produced an optimum set of conditions for the operation of the DSX circuit, relating in particular to the relative proportions of the extractants and the operating pH.  

Boléo limestone, available on the Boléo Reserve, was found to be effective in bulk neutralization duties and iron precipitation in the Iron Removal Stage during bench scale testing at SGS Lakefield.  Soda ash was found to be effective at pH control.  Milling in acidic raffinate was found to bolster circuit tenors and improve the overall 'water balance' for the circuit.

These various initiatives gave rise to the need to conduct a second pilot campaign with certain targeted objectives.  The campaign was duly scheduled for July 2006 and the pilot plant constructed and commissioned to meet these dates.    



 

 

 

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Objectives for the 19 day pilot campaign included:

·

Building on and verifying the “Proof of Concept” pilot plant results of 2004

·

Demonstrating the flowsheet feasibility on a larger scale and to a greater extent of integration than in the previous campaign

·

Assessing larger scale CCD settling behaviour and performance

·

Testing the suitability of the DSX technology for the selective recovery of zinc and cobalt from a Mn matrix using a modified (optimised) mixture of the synergistic reagents

·

Confirming key reagent consumptions, including the use of Boléo Limestone (HAC) for pH adjustment

·

Demonstrating the use of soda ash in pH control in the various Solvent Extraction circuits as a lower cost replacement for sodium hydroxide

·

Demonstrating the Cadmium Cementation step

·

Demonstrating Manganese Carbonate production

·

Confirming product quality

·

Verifying existing design criteria and confirming current design assumptions

·

Extracting engineering design data

·

Testing the proposed plant control philosophy.


In addition to continuous piloting of the Boléo flowsheet, on-site bench-scale testwork was conducted by the following vendors and industry specialists:

·

Outokumpu Technology – CCD & high rate thickening

·

Pocock Industrial Inc – CCD & high rate thickening

·

RPA Process – Filtration of iron residues, manganese carbonate product

·

Mixtec – Agitation testing in oxidative and reductive leach, partial neutralization and tailings neutralization

·

SGS Lakefield – Environmental characterisation of Boléo Pilot tailings

·

SGS Lakefield – Production of purified cobalt carbonate

·

SGS Lakefield – Cobalt removal from DSX zinc solution (precipitation by Caro’s acid).


Samples were also sent off-site for treatment by the following:

·

GLV Pty Ltd – Paste Thickening options for the CCD Circuit

·

Saskatchewan Research Council – Tailings pump loop testing

·

Jenike & Johanson Ltd – Flowability testing of crushed ore.


PILOT CAMPAIGN ORE FEED  


Approximately 10 tonnes of Boléo ore, composited from 143 channel samples of underground oxide Manto 3 ore from the test mining campaign, were homogenised, and crushed to produce



 

 

 

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the feed to the pilot plant.  A 200 kg composite sample was prepared and used for preliminary grinding testwork.

Of the initial 10 tonnes prepared approximately 8 tonnes were milled during the campaign, and approximately 5 tonnes were consumed in the pilot campaign.  The average grade of the elements of interest for the duration of the campaign is shown in Table 15.

Table15:

Comparison of As Received, 200 kg Composite and Scrubber Feed Assays

 

H2O %

Cu %

Co g/t

Zn g/t

Mn %

Fe g/t

Baja Assays – By Chemex (As Received from Site)

 

 

 

 

 

 

Average

27.39

2.0078

1252

4753

3.9601

8.2383

Length Weighted Average

1.9928

1235

4757

3.9464

8.1712

Boléo Bench Testwork - 200 kg SGS Composite Sample

 

Calculated

27.81

2.1388

1177

4820

3.6855

8.3434

Assayed

 

1.8200

1130

4020

4.8200

8.0000

Composite Samples Taken off the Feed Belt to Scrubber During Milling Campaigns

Weighted Average

 

2.1700

1320

4800

5.0400

8.1400


Grade variation of the prepared feed slurry over the 19-day campaign is shown in Table 16:

Table 16:

Leach Feed Slurry Assays

Sample

Cu %

Co g/t

Zn g/t

Mn %

Si %

Fe %

Feed Ore 1A

2.16

1020

4780

4.14

18.7

8.04

Feed Ore 1B

2.25

1230

4650

5.58

19.4

8.15

Feed Ore 2A

2.26

1410

4810

5.99

18.5

8.22

Feed Ore 2B

2.26

1240

5270

4.27

20.4

8.90

Feed Ore 2C

2.27

1570

4990

5.21

19.0

7.90

Feed Ore 3A

2.18

1370

5130

5.25

20.1

8.59

Feed Ore 3B

2.04

1550

4640

5.45

20.6

7.71

Feed Ore 3C

2.04

1370

4990

4.16

21.2

8.58

Minimum Feed Ore

2.04

1020

4640

4.14

18.5

7.71

Maximum Feed Ore

2.27

1570

5270

5.99

21.2

8.59

Average Feed Ore

2.18

1350

4910

5.01

19.7

8.26



OXIDATIVE, REDUCTIVE AND PARTIAL NEUTRALISATION LEACH STAGES


Recovery of Cu, Co and Zn from Boléo ore requires three atmospheric leach stages.

The first stage is an oxidative leach with sulphuric acid (pH 1.2) at 70°C to 80°C for three hours.  No reagent addition was required to maintain a redox potential of 600 mV to 950 mV due to the natural occurrence of MnO2 within the Boléo ore body.  



 

 

 

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A second stage reductive leach with sulphuric acid (pH 1.5) at 70°C to 80°C for three hours, ensures dissolution of the manganese minerals. A redox potential of 400 mV is maintained by addition of SO2 gas. This reductive step releases additional copper, cobalt, and zinc locked within the manganese mineralization.

A partial neutralization stage, conducted at pH 2.0 at 70°C 80°C for one hour follows prior to copper solvent extraction to ensure high extraction efficiencies in copper SX by subtle pH adjustment. The pH modification is also carried out to minimize gel formation, especially silica gel.

Impurities leached include calcium, iron, magnesium, and aluminum, contributing to the overall acid consumption of the ore.

 [bajatechrep062.gif]

Photo

Overview of the Leach Circuit.



COUNTER CURRENT DECANTATION (CCD)


Due to the fine nature of the feed slurry, counter current decantation was selected as the residue wash step in order to reduce soluble losses. The CCD circuit setup included six stages of thickeners, feed dilution to 2% to 3% solids, underflow washing with brine and manganese carbonate precipitation thickener overflow at a wash ratio of 1.75 m3/t and flocculation with Hychem 301.



 

 

 

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Photo

Overview of the piloting Counter Current Decantation (CCD) equipment setup.

The CCD circuit was operated at 40°C as per the requirements of the downstream solvent extraction circuit. The final tailings were drummed, with a portion blended with iron removal circuit residue, neutralised with HAC and lime to pH 7, and sent for pump loop testing by the Saskatchewan Research Council and environmental characterisation testwork by SGS Lakefields' Environmental Group.

COPPER SOLVENT EXTRACTION (CUSX)


The Cu SX process utilized 20%v/v LIX 664N extractant and Orfom SX 80 CT diluent. The circuit consisted of two extraction, one scrub and two strip stages operated at 40°C.



 

 

 

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Photo

Overview of the piloting Copper Solvent Extraction equipment setup.

Pregnant Leach Solution (PLS) from CCD 1 overflow was collected and filtered to remove suspended solids using a 1 µm in-line cartridge filter prior to being pumped to a storage container ahead of the CuSX circuit.  The solution was filtered with a 0.5 µm filter prior to entering the first extraction stage. No pH adjustment was done in the scrub and strip stages.  Organic entrainment removal was via a plastic chip coalescer for the raffinate stream and a combination of a plastic chip coalescer and a multimedia activated carbon filter for the loaded electrolyte stream.

COPPER ELECTROWINNING (CUEW)


The CuEW process utilized a conventional circuit with lead-calcium-tin anodes and stainless steel starter cathode sheets, operated at 40°C and a target current density of 250 A/m2 over a nominal four day strip cycle.  The lead anodes were conditioned prior to the start of the campaign to reduce the potential for lead contamination of the copper product.



 

 

 

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Figure 20:

Cu Solvent Extraction and Electrowinning Flowsheet, and Operating Parameters


[bajatechrep068.gif]


IRON REMOVAL


The objective of this step is to precipitate the iron, aluminium and residual copper as hydroxides by adjusting the pH of the process feed stream (CuSX raffinate) with HAC to pH4, effectively oxidizing Fe+2 to Fe+3.

Initially, the iron was removed in two stages with HAC and air being fed to the first stage, and lime and oxygen fed to the second stage. The circuit was modified during the campaign by removing one of the original thickeners giving a circuit configuration consisting of a feed tank and four reaction tanks followed by one thickener.



 

 

 

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The iron precipitate was thickened using Hychem 301. The thickener overflow proceeded to the Co Zn DSX circuit and the underflow was filtered.  The filter cake was bagged and drummed. The filtrate was returned to the second reaction tank. The circuit was operated at 50°C during the campaign to be representative of the operating temperature of the industrial plant.

[bajatechrep070.gif]

[bajatechrep072.gif]

Photos:

Pilot Plant Iron Removal Reactors and Iron Removal Thickener


CSIRO Direct Solvent Extraction (DSX)

The DSX process utilized a 6.25% Versatic 10 (0.33M) and 13.2% LIX 63 (0.3M) organic mix in Orfom 80 SX CT diluent. This organic mixture was developed by the Commonwealth Scientific & Industrial Research Organisation (CSIRO) in Australia for the selective recovery of Co and Zn in a matrix containing high manganese concentrations (typically 20 g/L to 50 g/L Mn).

Extensive bench-scale testing by CSIRO indicated selective recovery of Co and Zn from Boléo solutions and this was successfully demonstrated in the previous pilot plant campaign. At that time, however, the process suffered the disadvantage of organic degeneration, thought to be caused by manganese oxidation, acerbated by elevated temperature (>30°C).

Subsequent testwork by CSIRO and Bateman Advanced Technologies (BAT, Israel) confirmed the effect of manganese on the organic stability and lead to a series of tests by CSIRO to optimise both the circuit conditions and the relative proportion of the synergistic reagents.  Testwork successfully demonstrated that degradation can be all but eliminated by ensuring that manganese does not load onto the Lix 63 extractant.  



 

 

 

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The key process parameter in this regard is to ensure that the pH is maintained below 4.5 pH units.   

The circuit consisted of three extraction, two scrub, two Zn strip and two bulk (Co, Zn) strip stages operated at 30°C.

 [bajatechrep074.gif]


Photo

Overview of the Pilot Plant Direct Solvent Extraction (DSX) Circuit


Iron free solution was collected and filtered to remove suspended solids with an in-line cartridge filter prior to being fed to the DSX circuit.  pH in extraction was controlled by dosing a 100 g/L sodium carbonate solution, pH in Zn strip and bulk strip stages was controlled by dosing 200 g/L sulphuric acid. Zn and Co strip solutions were collected for further processing to zinc sulphate and cobalt metal respectively in the second phase of the testwork campaign.

The DSX operating parameters were as follows:



 

 

 

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Figure 21:

Direct Solvent Extraction (DSX) Flowsheet & Operating Parameters

[bajatechrep076.gif]

MANGANESE CARBONATE


The objective of this circuit is to precipitate manganese from the DSX raffinate stream. The raffinate from the Co Zn DSX circuit is fed to the manganese precipitation tank, which is maintained under an inert atmosphere with nitrogen gas.  



 

 

 

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Sodium carbonate (Na2CO3) is used to precipitate manganese as its carbonate. The treated stream is processed in a thickener; the overflow is recycled to the CCD circuit to be used as wash water and the underflow is filtered.

[bajatechrep078.gif]


Photo

Overview of The Manganese Carbonate Precipitation Circuit (Reaction Tanks and Thickener)

[bajatechrep080.gif]

[bajatechrep082.gif]


Photos:

Manganese Carbonate Thickener Underflow And Filter Cake


PILOT PLANT CONTROLS


Pilot plant controls included regular flowrate/mass measurements, on-line process variable trending (temperature, redox potential and pH), analytical laboratory assay trending, bench titrations, reagent consumption monitoring (flowmetres, load cell measurements, containers)



 

 

 

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and log sheets.  Actual flowrates were cross-checked with the load cell data to produce a full mass balance with the following accuracies:

·

solids within 5 % deviation

·

solution within 1 % deviation.


A high level process review was performed daily by the management team on site.

Analytical methods utilized by SGS Minerals Services Analytical Group included:

Table 17:

SGS Minerals Services Analytical Methods Employed

Area

Sample Type

Method

Batch Grinding

Pulp

SG

 

Solids

Malvern

 

Liquids

SG, ICPLa, Cl, Fe2+, FAT

Leach and Partial Neutralization

Pulp

SG

 

Solids

ICPS

 

Liquids

SG, ICPLa, Cl, Fe2+, FAT

CCD

Pulp

SG, TSS

 

Solids

ICPS

 

Liquids

SG, ICPLb

Iron Removal

Pulp

SG, TSS

 

Solids

ICPS

 

Liquids

SG, ICPLa, Cl, Fe

Cu SXEW

Aqueous

SG, ICPLa, Cu, Ge/In/Ga, Cl, FAT

 

Organic

SG, ICPO

 

Cathode

ICPC

DSX

Aqueous

SG, ICPLa, Fe, Ge/In/Ga, Cl

 

Organic

SG, ICPO


Special sample preparation for both analytical and SG determination samples included the multiple washing (two-stage repulp) of solids to minimize errors from soluble metal contributions.

COMMISSIONING AND CAMPAIGN DURATION


Pilot plant commissioning commenced on 5th June 2006.  The integrated pilot plant campaign ran until 24th June. The zinc solvent extraction and cobalt solvent extraction and electrowinning circuits were run during the period of 4th to 15th July in a separate campaign. The start-up times for the various sections are shown in Table 18.



 

 

 

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Table 18:

Boléo Pilot Plant Timeline

Section

Time

Date

Milling

Batch

5 June

Leach

23h00

6 June

CCD1 (Solids feed)

10h30

7 June

CuSX

10h45

8 June

Fe Removal

19h30

8 June

CuEW

09h00

10 June

DSX

01h00

10 June

MnCO3

10h30

11 June

First Cu Cathode

15h00

13 June

End of Campaign

 

 

Leach

06h00

24 June

CuSX

23h00

25 June

Fe Removal

07h00

27 June

DSX

07h00

27 June

MnCO3

07h00

27 June



SUMMARY OF PILOTING RESULTS


LEACH AND PARTIAL NEUTRALISATION


The Leach and Partial Neutralization operational KPIs were as follows:

Table 19:

Leach and Partial Neutralisation Operating Result

Parameter

Unit

Oxidative Leach
Tank 1 Tank 2

Reductive Leach
Tank 1 Tank 2

Partial
Neutralization

Temperature

ºC

80

80

80

80

80

pH (Ag/AgCl – sat’d KCL)

 

1.5

1.4-1.7

1.2-1.5

1.2 -1.5

2.06

Redox Potential

mV

>800

>800

400

400

~400


The average leach extractions for the various pH conditions of the campaign in Oxidative Leach Tank 2 are shown below. The figures were calculated using a silica tie method to compare the metals values in the feed with the values in the washed residue from CCD.

Table 20:

Leach Efficiencies

Condition

OL2
pH

% Extraction
Cu

% Extraction
Co

% Extraction
Zn

% Extraction
Mn

OL2 pH = 1.7

1.7

90.5

80.8

52.9

96.6

OL2 pH = 1.5

1.5

90.5

79.4

54.4

95.8

OL2 pH = 1.4

1.4

92.4

82.1

60.1

97.7

Start-up

1.2

94.1

89.3

71.6

94.9

Overall average

1.4

91.8

82.4

59.1

96.5




 

 

 

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Graphs of the metals extractions for the duration of the pilot plant campaign are shown below.

Figure 24:

Boléo Metal Extractions

[bajatechrep084.gif]



 

 

 

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[bajatechrep086.gif]

Note:

    The following abbreviations apply - Leach Feed (LFS), Oxidative Leach (OL2), Reductive Leach (RL2), Partial Neutralisation (Neut 1) and CCD6 Underflow (CCD 6 U/F).


COUNTER CURRENT DECANTATION (CCD)


The CCD circuit average operational KPIs were as follows:

Table 21:

CCD Operating Results

Parameter

Unit

CCD 1

CCD 2

CCD 3

CCD 4

CCD 5

CCD 6

 

 

 

 

 

 

 

 

U/F Density

% solids

10

12

14

15

16

15

Wash Ratio

m3/t

1.75

Wash Efficiency

%

91

Flocculant Addition – Total

g/t

470




 

 

 

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The graph below shows the improved operational performance of the CCD circuit as the campaign progressed.

Figure 25:

Boléo Pilot Plant CCD Thickener Mud Levels

[bajatechrep088.gif]


Observations:

·

The CCD circuit produced underflow densities ranging from 10% to 24 % solids as reported on a shift basis. However, some of the results, particularly the higher ones, could be suspect as the calculation depended on a fixed value for the SG of the dry solids. The underflow densities obtained during vendor testwork (shown below) give a more accurate indication of the values attainable.

·

Wash efficiencies as high as 99% were achieved but only at wash ratios in excess of 4 m3/t.

·

CCD 1 overflow clarities were generally good with clarity wedge readings of 47 on a 0 to 47 scale.

·

There is potential for two-stage flocculation to reduce flocculant consumption.

·

Analysis of the CCD6 underflow assays revealed that additional leaching was taking place in the CCD circuit.


VENDOR TESTWORK


Thickening testwork was carried out by Outokumpu Technology and Pocock Industrial Inc during the campaign. In addition, samples were sent to GL&V for paste thickening testwork. The table below summarises the vendor results.



 

 

 

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Table 22:

Vendor Testwork – CCD Parameters

Parameter

Outokumpu

Pocock

GLV

Diluted feed, % solids

2-3

2-3

1.5

Max underflow, % solids

16.1

16.1

21-22

Overflow clarity – TSS, ppm

60-85

80-130

-

Optimum rise rate, m/h

1.64

0.87

-

Typical bed height, mm

180-230

1000

300

Flocculant

Hychem NF301

Hychem NF301

Hychem 302


It was noted that underflow densities were as much as 6% (absolute) lower at similar rise rates achieved in comparative tests conducted by Outokumpu in 2004.  The reasons for the difference in performance are not well understood but it is postulated that they result from the combined effects of milling in acidic raffinate and the use of Boléo limestone for neutralization purposes.

COPPER SOLVENT EXTRACTION (CUSX)


 The CuSX circuit extracted an average of 98.6% Cu. The average pregnant leach solution and raffinate assay values are shown below.


Table 23:

Copper Solvent Extraction Solution Assays

Element

Unit

PLS

Raffinate

Cu

mg/L

2515

34.5

Co

mg/L

265

257

Zn

mg/L

1023

976

Mn

mg/L

11710

11601

Fe Total

mg/L

5916

5837

Ca

mg/L

601

589

Mg

mg/L

7573

7442

Al

mg/L

3369

2901

Ni

mg/L

28

28

Si

mg/L

50-170

<140

Cl

mg/L

14535

13570


Only minor amounts of manganese and iron were co-extracted with the copper.  The average organic assays are listed in Table 24.

Table 24:

Copper Solvent Extraction Organic Assays

 

 

Organic Assays

Element

Unit

Loaded

Scrubbed

Stripped

Cu

mg/L

5620

5417

3200

Mn

mg/L

0.10

0.07

<0.05

Fe Total

mg/L

36

8

0.9

Ni

mg/L

< 2

< 2

< 2




 

 

 

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COPPER ELECTROWINNING (CUEW)


Four plating cycles were carried out during the campaign.  The conditions are shown in Table 25.

Table 25:

Plating Cycle Times & Operating Conditions

Cycle Number

1

2

3

4

Start Time

June 9 17:50

June 13 15:33

June 16 18:43

June 21 15:18

End Time

June 13 15:06

June 16 11:51

June 21 15:18

June 25 23:18

Cathode Mass (kg)

A

B

A

B

A

B

A

B

8.78

8.88

4.41

4.52

5.74

5.88

4.52

4.61

Av. Current Density,A/m2

183

162

155

167

Current Efficiency, %

94.4

97.3

99.5

99.6


In spite of the difficulties encountered in copper solvent extraction all four cycles produced copper of the following quality. All exceeded the LME grade-A specification.

Table 26:

Cathode Quality

Sample Number

Cathode, Cu%

1

>99.996

>99.996

2

>99.993

>99.995

3

>99.995

>99.995

4

>99.996

>99.996


[bajatechrep090.gif]

Photo

Cathode From Cycle No.1 – 13 June 2006




 

 

 

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IRON REMOVAL


The occasionally erratic performance of the iron removal circuit did not adversely affect the downstream DSX operation. Many of the operational problems resulted from the relatively small size of the pilot plant equipment. The blockage of spargers and the poor oxygen utilization are not expected to occur in the industrial plant with the frequency that they did in the pilot plant.

The average concentration of iron in the feedstream was 5454 mg/L and in the thickener overflow (feed to DSX) it was 17 mg/L giving an average iron rejection for the campaign of 99.7%.

It must be noted that the iron levels in the thickener overflow were generally less than 1 mg/L and the average quoted above is a result of excursions resulting from blockage of oxygen spargers and / or limestone (HAC) addition.

The iron removal thickener in the pilot plant gave an average percent solids in the underflow of 49%. This was supported by Outokumpu testwork results which reported a maximum underflow density of 46% solids, at a rise rate 3.81 m/h and an overflow clarity of 72 ppm solids.

Observations

·

In general the iron precipitate settled and filtered easily and gave a clearer overflow at higher temperatures. The initial operating temperature of 60°C was dropped incrementally to 45°C over a period of two days from 15th June.  However, this resulted in poorer settling performance, reduced overflow clarity and poorer filtration characteristics and a decision was taken to run the circuit at 50°C for the remainder of the campaign.

·

The occasional use of hydrogen peroxide to “trim” the remaining ferrous iron did not cause problems with the DSX extractant during the campaign.

·

A combination of coagulant (Magnafloc 368) and flocculant (Magnafloc 155) was found to best flocculate the iron removal thickener feed stream.


CSIRO DIRECT SOLVENT EXTRACTION (DSX)


The DSX circuit ran trouble-free for the 425 hours of the campaign. Crud was formed in the extraction stages but did not adversely affect the physical or chemical behaviour of the circuit. The crud is likely to have been caused by particulates or precipitates in the feed stream and was not formed by products of organic degradation as was the case in the first pilot plant campaign in November 2004.

None of the black “manganese crud” of 2004 was evident and this is attributed to the excellent control of pH that ensured low manganese loading on the organic at approximately 100 mg/L throughout the campaign.



 

 

 

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Figure 26:

Manganese Loading On DSX Organic Phase

[bajatechrep092.gif]


METAL EXTRACTION


The DSX extractions for copper, cobalt, and zinc are shown in Table 27.

Table 27:

DSX Metals Extractions

Description

Unit

Cu

Co

Zn

Extraction

%

99.57

99.48

99.03


ZINC SELECTIVE STRIP


The graph below shows that a zinc tenor in the zinc strip liquor of 40 g/L to 45 g/L can be maintained while keeping a Zn:Co ratio of approximately 50:1.

Figure 27:

Zinc in Solution vs. DSX Operating Time

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BULK STRIP


The graph below shows the cobalt and zinc tenors in the bulk strip liquor.  Cobalt concentrations in excess of 10 g/L were attained while maintaining a Zn:Co ratio of 1.6:1.

Figure 28:

Cobalt in Solution vs. DSX Operating Time

[bajatechrep096.gif]


Observations

·

Excellent pH control in extraction was achieved with 100 g/L sodium carbonate.


MANGANESE CARBONATE


Manganese carbonate was precipitated from the DSX raffinate by the addition of 150 g/L sodium carbonate slurry. The sodium carbonate was added at 80% of stoichiometric requirements to minimize the co-precipitation of calcium and magnesium. The table below shows the average results of 19 of the 21 batches produced. Intermittent overdosing of the sodium carbonate resulted in some low manganese values but in general values of 45% or greater were achieved.

Table 28:

Manganese Carbonate Product Average Assay Values

Element

Value

Units

Mn

44.2

%

Ca

1.1

%

Mg

0.2

%

Al

0.4

%

Fe

173

g/t

Co

29

g/t

Zn

201

g/t

Al

0.4

g/t

Ni

365

g/t

Si

443

g/t

Cu

< 5

g/t

Cd

< 5

g/t




 

 

 

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Observations:

·

The pale pink manganese carbonate precipitate settled, filtered and washed well.

·

Seeding of the feed from 1.4% to 5.8% solids with recycled underflow led to better flocculation, improved underflow percent solids (62% to 67%) and clearer overflows (47 on the clarity wedge).


Reagent Consumption

Consumption (kg of reagent per tonne of dry feed) of the key reagents in the leach and CCD circuits for various pH conditions is shown below.

Table 29:

Summary of Leach and CCD Reagent Consumptions

Condition

RL2
ORP

H2SO4 Rate
(kg/t)

SO2 Rate
(kg/t)

HAC Solids Rate
(kg/t)

Flocculant Rate
(kg/t)

OL2 pH = 1.7

399

228

81

69

0.53

OL2 pH = 1.5

397

228

71

73

0.40

OL2 pH = 1.4

399

330

129

116

0.46

Start up

427

496

144

362

0.55

Overall average

404

311

106

138

0.47



15.2

PROCESS PLANT DESIGN


This section remains largely unchanged from the previous document but does contain some fundamental changes to the flowsheet and the process design as a result of the knowledge and experience gained from the Fully Integrated Pilot Campaign.

15.2.1

INTRODUCTION


The proposed treatment route for the Boléo ore consists of a two-stage, whole of ore, sulphuric acid and sulphur dioxide leach followed by solid-liquid separation in a counter current decantation circuit prior to solution purification and metal concentration using solvent extraction technologies.  Cobalt and copper will be electrowon to produce high purity products for export to global metal markets.  Zinc will be recovered in a granulator as a zinc sulphate crystal suitable for incorporation in animal feed.  

Sulphuric acid is manufactured on the site in a stand-alone acid producing facility, employing sulphur as the main feedstock.  Power is produced by harnessing the energy in the steam produced during the acid production process to generate power in a so-called 'co-generation' facility consisting of a steam turbine and generator.  Limestone, which is available on the mine site, is crushed in the field and milled on site to provide for plant bulk neutralization duties.     

Revision I of the schematic flow diagrams summarizing the proposed plant processing route is included below.  



 

 

 

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Figure 29:

Schematic Flow Diagram Sheet 1

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Figure 30:

Schematic Flow Diagram Sheet 2

[bajatechrep100.gif]



 

 

 

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Figure 31:

Schematic Flow Diagram Sheet 3

[bajatechrep102.gif]


15.2.2

PROCESS PLANT DESCRIPTION


The proposed process plant consists of the following main areas:

Primary Crushing – ROM ore is crushed through a single-stage, toothed roll crusher selected specifically to maintain high availabilities on the high clay containing Boléo ore types.

Scrubbing, Screening, Secondary Crushing and Milling – The crushed ROM is treated in two scrubbers designed to slurry the high clay component of the ore.  The slurry exiting the scrubber is screened.  The remaining coarse ore fraction is crushed in a secondary crushing



 

 

 

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operation and reports via a pump box to a cyclone cluster for classification.  Screen underflow is also cycloned to remove the fine sized material.  The coarse fraction from the cyclone is returned to the ball mill for further size reduction.  Slurrying and milling take place in acidic raffinate returned from copper solvent extraction.  Ceramic balls are used as grinding media.

Oxidative Leach – The ground ore is leached in sulphuric acid and recycled copper SX raffinate to leach (dissolve) the copper, manganese and zinc containing minerals and produce a solution rich in copper and zinc sulphate solution.  A small portion of the cobalt is leached at this stage too.

Reductive Leach – The slurry is subjected to a reductive leach where SO2 gas is bubbled through the slurry to facilitate the leaching of the remaining cobalt minerals.  At the end of the reductive leach typically 90% of the copper and 80% of the cobalt has been leached into solution.

Partial Neutralisation The slurry overflowing the reductive leaching circuit will contain residual acid which will be partially neutralised to ensure optimal copper recovery in the downstream copper solvent extraction circuit.  Partial neutralization is achieved via the oxidation of ferrous to ferric ions using blower air and the addition of a relatively small quantity of Boléo Carbonate.  

Slurry Cooling – Partially neutralised slurry cooling takes place in a heat exchanger step followed by further cooling in a forced draft cooling tower.  Cooling of the slurry is required to facilitate the performance of downstream operations.  Heat exchangers are used to recover heat into raffinate for subsequent use elsewhere in the circuit.

Counter-Current Decantation (CCD) Circuit – Cooled leach residue slurry discharged from the cooling tower is washed with sea water in a six stage CCD circuit to recover soluble copper, zinc and cobalt.  Deep bed, high rate thickeners are utilized.  Thickener underflow is pumped from the last CCD stage to a tailings neutralization facility.  Thickener overflow from the first CCD stage contains the bulk of the leached copper, zinc and cobalt and is termed pregnant leach solution or PLS.  The PLS is transferred to the copper SX circuit for copper extraction and concentration prior to electrowinning.  

Copper Solvent Extraction (SX) – The copper solvent extraction circuit consists of a 2-stage extract, 1-stage wash, 1-stage parallel extract and 2-stage strip operation.  Copper is recovered from the PLS into an organic liquor known as an extractant. The PLS aqueous solution, now stripped of copper and known as raffinate (containing residual copper, zinc, cobalt, iron and manganese, all as sulphate species) is split between the scrubbing and milling circuit, (where it is used as dilution and make-up solution in the comminution circuit) and the iron removal process where levels of certain metallic species, principally iron and aluminium, are reduced prior to further concentration of the zinc and the cobalt in the DSX ® operation.

The loaded organic stream from the copper extraction circuit is ‘stripped’ with spent copper electrolyte from the copper electrowinning operation, producing loaded copper electrolyte and stripped organic, the latter being recycled to copper SX.  Loaded copper electrolyte is pumped,



 

 

 

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via multi media filters to remove carry-over organic and entrained solids, to Copper Electrowinning where copper is recovered as LME grade cathode.  

Electrowinning (EW) – Copper metal is electro-deposited from filtered loaded electrolyte, onto stainless steel blanks, known as cathodes, over a nominal 7-day cycle.  Cathodes are harvested via an automated stripping machine on a semi-continuous basis.  

An overhead crane will lift out sets of cathodes from the EW cells.  The cathodes are then washed, and fed on a chain conveyor type system to the automated stripping machine.  Copper cathode is  automatically 'stripped' from the stainless steel blanks, sampled, weighed and packaged for sale.  Spent copper electrolyte is eventually returned to the copper SX circuit.

Iron Removal – The copper raffinate from the copper SX will contain residual copper, zinc, cobalt, iron and manganese.  This step has been designed to remove residual acid and iron from solution.  Air is sparged through the solution to convert the ferrous ion to ferric ion and, by increasing the pH using limestone, encourage the formation of iron hydroxide precipitates.  The products of the iron removal process, primarily goethite, will be thickened and washed in a small CCD circuit in order to recover as much process liquor and dissolved metal values as possible before being diverted to tailings neutralization.  

Thickener overflow reports to a pinned bed clarifier for final polishing ahead of the DSX ® circuit.  This solution is at a pH of 4.5 in preparation for the DSX ® extraction step.    

Tailings Neutralization and Disposal – Thickened tailings from the main CCD circuit and thickened underflow from the iron removal CCD circuit are treated in a neutralization operation where limestone is added to neutralise excess acid precipitate metal salts contained in the tailings solution.  Subsequently lime will be added to the slurry in a second tank to raise the pH to that required for disposal.  

The slurry discharged from this process is pumped to a tailings dam situated to the west of the plant site known as the Curuglu area.

DSX ® – Copper, zinc and cobalt, contained in solution from the iron removal clarifier overflow, are separated from manganese, magnesium and calcium into an organic extractant phase via the DSX® process.  The DSX process employs two extractants known as LIX63i and Versatic 10 to effect this separation.  The zinc solvent extraction circuit consists of a 2-stage extract, 2-stage scrub and 2-stage strip operation.  The organic phase is scrubbed with aqueous solutions to remove small quantities of impurity from the loaded organic.  

The loaded organic – containing the cobalt and the zinc - is then stripped and concentrated into an acidic stream for further treatment.  This stream reports to the zinc solvent extraction circuit after a cadmium reduction step.  

Cadmium Removal – the DSX strip solution is diverted to a cadmium removal step where zinc dust is used to purify the solution of excess cadmium at a pH of 3.3, enabling the product to meet the quality specifications required for use as an additive to animal feed.  The solution is



 

 

 

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filtered with the filter cake being diverted to tailings and the filtrate being passed to a zinc SX step.  

Zinc Solvent Extraction – The purified bulk strip solution from DSX ® becomes the feed to a zinc removal solvent extraction operation, prior to cobalt SX and cobalt EW.  The zinc solvent extraction circuit consists of a conventional 4-stage extract, 2-stage scrub and 2-stage strip operation.  Strip liquor, rich in zinc, reports to two places in the flowsheet.  A small proportion is returned to the DSX ® as scrub liquor and the remainder the feed to a facility producing zinc sulphate monohydrate crystals in a fluidized bed granulator.  

Cobalt Solvent Extraction, Ion-Exchange and Electrowinning – Raffinate from the zinc secondary SX circuit, now containing only cobalt and very small quantities of zinc, nickel, and iron, will report to the cobalt SX operation.  The circuit consists of a 4-stage extract, 2-stage scrub and 3-stage strip operation. Strip liquor from the cobalt SX operation is further purified through ion exchange where highly selective resins remove extraneous metal ions prior to electrowinning to produce cobalt metal. Cobalt electrowinning requires high purity feed solutions; the IX circuit provides a final ‘catch-all’ scavenging process. Electrowon cobalt is crushed, screened and dispatched in drums as a premium metal product.

Zinc Sulphate Monohydrate Production – Zinc strip solution reports to a fluidised bed granulator where heat is applied to the solution via the use of hot air to affect the crystallisation process.   Particles are 'grown' to the appropriate size and shape, cooled, screened and packaged in a stand-alone production facility.  The air used in the process is heated by combustion of diesel.   

Sulphuric Acid Generation and SO2 Production – Sulphuric acid and sulphur dioxide are produced in a 'stand alone' facility which consists of a number of unit operations including:  

·

liquid sulphur storage

·

air drying

·

sulphur burning

·

gas conversion (employing catalysts)

·

economizers and heat exchangers

·

gas scrubbing and acid storage

·

liquid SO2 production and storage.


The SO2 is sparged directly into the slurry in the reductive leach operation.

Power Production – Steam exported from the sulphuric acid and SO2 gas plant is used to generate electricity via the use of a gas driven turbine and power generation system.  The power generated via this co-generation system exceeds the overall mining and plant operational requirements at design throughputs, making the Boléo Operation self sufficient in electricity.  Any power shortfall experienced at elevated throughputs can be made up by the use of standby diesel powered generating sets.  



 

 

 

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Limestone Milling – A limestone milling circuit is required to produce ground limestone slurry for various bulk neutralization duties throughout the circuit namely Partial Neutralisation, Tailings Neutralisation and Iron Removal.  Limestone is mined on the Boléo Reserve, crushed in the field, stockpiled, milled and then circulated through the plant on a ring main system.

Process Water – Seawater will be employed throughout the circuit as process water.  

Cooling Water – Seawater will be employed throughout the circuit as cooling water.  

Desalinated Water – Desalinated water, required for particular duties throughout the plant, will be produced in a Multiple Effect Distillation type desalination facility.  A small quantity of this water production will be diverted for use throughout the plant and mining operations as potable water.

Tailings Dam – The Boléo Creek tailings dam, located in the Curuglu area, is designed to accommodate approximately 20 years worth of plant tailings.  The dam wall will be raised in 3 stages over the life of the mine.  The first dam wall will accommodate approximately 5 to 7 years worth of plant tailings production at design feedrates.  




 

 

 

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16

MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES

 

 


16.1

3D GEOLOGICAL INTERPRETATIONS


The geology of the Boléo deposit is well known (see sections 7, 8, and 9 above) and has been extensively studied.  Mineralization is effectively confined to seven shallowly dipping layers, referred to as 'mantos'.  A large number of faults have been identified from either surface mapping or mapping historical workings.  These faults play an important role with respect to local continuity of both geology and grade.

To enable the resource estimates to be accommodated accurately within the geological framework the existing geological interpretation was used to construct a 3D framework comprising faults and Manto footwall surfaces.

Digital geological data supplied by MMB comprises:

·

Manto footwall contours (for manto 1, 2, 3aa, and 3), in plan as *.dxf files

·

Traces of faults in plan, effectively as lines of intersection on Manto 1 and 3 footwall surfaces, as *.dxf files

·

Erosional windows of mantos, as *.dxf files

·

Approximately 230 cross-sections (each 2km in width), amounting to 460 line kilometres.


The geological interpretation was built in the software package MineSight 3D.

16.1.1

FAULT SURFACES


To build the fault framework, faults were firstly digitised as ‘polylines’ from each cross-section.  No nomenclature or labelling exists to correlate faults from section to section, so this was achieved by using the plans of fault line of intersections with Mantos 1 and 3 as guide to connect the digitised polylines correctly.  The process and results are depicted below in
Figures 32 to 35.






 

 

 

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Figure 32:

Fault Framework in Section – Detail of Cross-Sections (900N) as Supplied by MMB

[bajatechrep103.jpg]

Figure 33:

Digitised Fault Traces – Faults Digitized (dashed red lines) Following Cross-Sectional Interpretation (900N)

[bajatechrep104.jpg]

Figure 34:

Correlation of Fault Traces

[bajatechrep105.jpg]

Note:

    Digitized Polylines of Faults (red) Joined from Section to Section using Fault Lines of Intersection on Manto 3 (yellow) as a guide



 

 

 

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Figure 35:

Final 3D Fault Framework with Mintec Cross-Sections

[bajatechrep106.jpg]

16.1.2

MANTO FOOTWALL SURFACES


The Manto footwall surfaces were more complicated to define as they had to fit with the fault framework described above.  

To create these footwall surfaces, each drill hole was tagged with a set of points corresponding to the lowest logged occurrence of each manto.  For each Manto, these points were triangulated to form a continuous 3D surface (Figure 36).

The resulting 3D surface was sliced along the same 100 section lines as the Mintec geological interpretation (Figure 37), these slice lines were then superimposed over the Mintec sections and ‘dragged’ and ‘snapped’ to the digitised fault framework (Figure 38).






 

 

 

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Figure 36:

Triangulated Manto Surface – (Manto 3, oblique 60o)

[bajatechrep107.jpg]

Figure 37:

Triangulated Manto Surface Sliced Along Section Lines – (Manto 3, oblique 60o)

[bajatechrep108.jpg]



 

 

 

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Figure 38:

Snapping Footwall Surface to Faults

[bajatechrep109.jpg]

Note:

    Triangulated Surface Slices (yellow), New Surface Slices after Snapping to Faults (dashed Red)

A new manto surface was then created from the resulting profiles that honour the faults, the original drill hole tag points and tie lines along the upper and lower intersections of each fault with the manto surface.  The latter are required to force the triangulation to precisely honour the fault offsets (Figures 39 and 40).

Figure 39:

Re-Triangulation of Manto Surface Components

[bajatechrep110.jpg]

Note:

    New Fault Snapped Section Lines (yellow), Fault Tie Lines (blue) and DH tag Points (red)



 

 

 

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Figure 40:

Re-Triangulation of Manto Surface – Detail of Final Result

[bajatechrep111.jpg]

The resulting triangulation is a new manto footwall surface (Figure 41) that honours the base of the manto as defined by each drill hole intersection but also honours the faults as defined on the geological cross-sections (Figure 42).

Figure 41:

Re-Triangulation of Manto Surface – Full View (Manto 3, Oblique 60o)

[bajatechrep112.jpg]



 

 

 

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Figure 42:

Re-Triangulation of Manto Surface – Full View with Faults (Manto 3, Oblique 60o)

[bajatechrep113.jpg]

The same process was followed for all seven Manto footwall surfaces.


16.2

RESOURCE ESTIMATION TECHNIQUE


16.2.1

BLOCK MODEL VS. GRIDDED SEAM MODEL


The Boléo copper-cobalt resource will be mined by a combination of open cut and underground methods.  Consequently, different resource models need to be built that reflect these different requirements.

In 2005 3D block models were produced for the purposes of open cut optimisation.  This approach was appropriate and has been used once again in the latest studies to produce models suitable for open cut optimisation.

As predicted the 3D block models proved to be less useful for underground mine design.  The planned underground mining method is constrained by minimum and maximum height parameters and it is not possible to exercise the same degree of mining selectivity using a 3D block model.  Consequently a seam model approach to resource modelling has been used to produce models for underground mine design and scheduling.

The predominant mining method used for each manto will vary.  Mantos 1, 2 and 4 will be mined by underground methods only.  Manto 0, 3aa, 3a will be mined by open cut methods and Manto 3 will be mined by a combination of both open cut and underground methods.



 

 

 

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16.2.2

DATA DOMAINS AND FLAT MODELS


The definition of data domains is required to limit data used for grade estimation to the data within the area being considered.  The most obvious control that needs to be applied at Boléo is to keep separate the data and estimation process for each Manto.  Data for each manto was loaded into different data files and individual models were created and the estimations process was carried out separately for each manto.

The second controlling influence on the data that needs to be considered is the effect of the faulting.  Faulting is predominantly post mineralization and as such has resulted in fragmentation of once continuous Manto units into numerous blocks, each bounded by faults.  In plan, the continuity of the mantos remains.  In section, continuity before a fault termination is significantly reduced to distances as short as 100 m.

Previous resource estimates have used faults as domain boundaries with ore blocks and data restricted to specific fault blocks.  The effect of this, where faulting is dense, is to severely limit data available for use in the grade estimation process.  To estimate grade of a particular block, Kriging techniques capture data by means of a search ellipse (with dimensions defined by the user) with its centre at the block mid-point.  All data that falls with this search ellipse are used in the estimation process.  

In Figure 43, a search ellipse is shown in plan and in section.  In plan the ellipse, centred over a highlighted block, appears to capture three drill holes that fall within it (yellow points).  However, when viewed in section, the block, which is adjacent to but on the up-throw side of a fault, has its search ellipse well above the data on the down-throw side and consequently no data will be captured from that side of the fault.

Because the faulting at Boléo is predominantly post mineralization, the two segments of Manto now off-set by faulting would originally have been juxtaposed.  This being the case, the inclusion of the holes on the down-throw side of the fault would be justified.  It is not practical to capture data significantly off-set by faults simply by means of modifying the search ellipse parameters (i.e., increasing the vertical search distance) as this would require very large vertical searches that may also result in too much data from higher levels in the manto being captured as well.  

To achieve the desired effect the base of each manto intersection in each drill hole was set at the same (fixed) elevation.  The base of the model was also set at the same elevation.  This step removes the effect of the faulting and returns drill hole intersections to their pre-faulting position, relative to adjacent holes.  It is then possible to consider all the data for each Manto as a single mineralized domain.

The effect of this is to create a resource model that has the same EW and NS lateral extents but exists in a reduced or flat vertical space.  Aligning each hole at the same level not only removes the faulting off-sets but also the easterly dip of the mantos.  This has an additional advantage in that each block model has a significantly smaller vertical extent as each model only has to be slightly thicker than the maximum Manto thickness.  These models are referred to as flat models.



 

 

 

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Figure 43:

Search Ellipses – Plan & Section Views

[bajatechrep115.gif]

Plan View

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Section View



 

 

 

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All grade modelling steps for both 3D block models and seam models was carried out using the flat models.  A final step was required to convert the flat model block elevations back into true space.

This process was achieved by:

·

Generating a 2D gridded surface from the already created Manto footwall surfaces.  These surfaces were gridded on 25 m x 25 m centres such that a sub-set of grid points were coincident with block mid-points.

·

Deleting all points from the gridded surface file that were not coincident with block centroid points.

·

Further editing the reduced gridded surface file, so that the elevation was set as the mid-point elevation of the lowest blocks in the flat model.  The base of the flat block models is set at zero so the mid-point elevation is 0.5 m.

·

Retain the true footwall elevation values to be simply imported into the flat model as a separate model item.

·

Exporting (after grade estimation was completed) block coordinates and the item containing the true footwall elevation, from all blocks with estimates, into a spreadsheet where the true mid-point elevation of each block stacked above a footwall block was calculated by adding 1m progressively to the footwall elevation for each higher block.

·

Importing true model elevations for each block as a new item in the flat models.


All blocks now have both flat and true mid-point elevations.


16.3

3D BLOCK MODELS


16.3.1

COMPOSITE LENGTH AND BLOCK DIMENSIONS


In the resource estimation process when original assay samples are of varying lengths it is necessary to composite these into equal lengths, because each original assay value is representative of different proportions of a mineralized interval.  Compositing results in an assay population where each value or composite has equal weight.

At Boléo original sample lengths vary considerably, from as little as 10 cm to >1.5 m.  The average sample length for all mantos is about 0.95 m.  A composite of 1m was considered appropriate.

After determining a 1 m composite length, a similar 1m block height for the 3D model is sensible.  Ideally, model blocks should be greater than the composite interval, but thicker blocks would not be suitable as the manto thickness is generally <5 m.

The drill hole spacing and, therefore, data density, in plan, is used to decide the block dimensions.  The closest drill hole spacing at Boléo, over a meaningful area, is approximately 140 m x 140 m in the Saturno – Arroyo Boléo area.  Blocks 100 m x 100 m, or larger would be



 

 

 

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best from a modelling perspective, but large blocks would not fit readily into the framework of faults.  Consequently blocks 50 m EW by 100 m NS were chosen.

16.3.2

UNIVARIATE STATISTICS OF DATA COMPOSITES


Univariate statistics and histograms of grade for Cu, Co, and Zn for each Manto are shown in Tables 30, 31, and 32.  

Although new 3D block models have not been produced for Manto 1 and 2 the 1m composite statistics are shown for completeness and comparison.  Resource models for manto 0, 3aa and 4 have not been updated at this stage, so composite statistics are the same as reported previously.

Coefficients of Variations (CV) are generally less than 2, which indicate the data is not highly skewed and that non-linear estimation techniques, such as Multiple Indicator Kriging, are not warranted.  Exceptions to this are the Copper distributions for Mantos 1 and 2, which have CV values of 2.29 and 2.55 respectively, whilst Manto 3a and 4 have CV values close to 2.  Cobalt and Zinc CV values are uniformly low.

Ordinary Kriging is considered an appropriate method to use for Cu, Co and Zn, for Manto 0, 3a, 3a, 3 and 4 because:

·

Manto 1 and 2 are not reported as 3D block models based on the 1 m composites

·

The copper resource is dominated by Manto 3 which has a CV well below 2

·

Co and Zn CV values are uniformly low.


Table 30:

Univariate Statistics of Assay Composites – Copper

 

Manto

Copper

0

1

2

3aa

3a

3

4

Mean

0.02

0.55

0.27

0.53

0.35

.84

0.32

CV

1.9

2.29

2.55

1.3

2.04

1.46

2.0

Min

0.002

0.001

0.001

0.001

0.001

0.001

0.001

Q1

0.01

0.01

0.02

0.09

0.02

0.09

0.05

Median

0.01

0.04

0.08

0.20

0.07

0.342

0.14

Q3

0.02

0.43

0.25

0.66

0.34

1.14

0.32

Max

0.29

15.35

13.45

3.96

7.45

14.60

8.53

IQR

0.01

0.42

0.23

0.58

0.32

1.05

0.28

Data

237

775

1,859

127

1,781

4,283

1,312




 

 

 

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Table 31:

Univariate Statistics of Assay Composites – Cobalt

 

Manto

Cobalt

0

1

2

3aa

3a

3

4

Mean

0.008

0.0375

0.043

0.082

0.068

0.069

0.028

CV

1.0

1.29

1.82

0.8

1.04

0.97

1.0

min

0.001

0.001

0.001

0.007

0.001

0.002

0.001

Q1

0.004

0.007

0.013

0.028

0.023

0.027

0.013

Median

0.006

0.017

0.026

0.064

0.045

0.049

0.020

Q3

0.009

0.047

0.058

0.11

0.083

0.089

0.034

Max

0.079

0.475

2.82

0.281

0.630

1.510

0.464

IQR

0.005

0.040

0.045

0.082

0.060

0.062

0.464

Data

237

775

1,859

127

1,781

4,283

1,312


Table 32:

Univariate Statistics of Assay Composites – Zinc

 

Manto

Zinc

0

1

2

3aa

3a

3

4

Mean

0.42

0.81

0.87

0.73

0.56

0.35

0.22

CV

0.9

1.71

1.2

0.9

0.8

1.26

0.8

min

0.01

0.003

0.025

0.06

0.00

0.00

0.00

Q1

0.12

0.27

0.3

0.30

0.26

0.18

0.11

Median

0.29

0.49

0.57

0.61

0.43

0.27

0.16

Q3

0.60

0.86

1.01

0.92

0.71

0.41

0.27

Max

1.73

20

9.28

4.99

5.11

13.76

1.56

IQR

0.49

0.59

0.71

0.62

0.45

0.23

0.17

Data

237

775

1,859

127

1,781

4,283

1,312



16.3.3

SPATIAL CONTINUITY OF GRADE


Variograms of Copper, Cobalt and Zinc were created and modelled using Hellman & Schofield proprietary software “GS3”.  Variogram analysis utilized data re-aligned for flat models.

Variogram maps for each metal showed very poor structure with no strong directional controls (Figure 44).  A weak anisotropy in a NE-SW direction is apparent, particularly for Copper.  The range, or distances over which there is a spatial relationship between the grades at two points appear to be relatively short, a few hundred metres, compared to the full extent of the mantos (i.e., several kilometres).

Directional variograms (using trigonometric rather than grid coordinate conventions, i.e., 000 = East, 090 = North) were generated for all data.  Modelled variograms for Manto 3 are shown below (Figures 45 to 48).

Variograms were modelled with a relatively low nugget as determined from the well structure down hole (Z direction) variograms.  Variogram models for Mantos 3a, 3 and 4 are shown in Table 33.  Note that no variograms were possible for Manto 0 and 3aa due to lack of data.  Variograms for Manto 1 were used when estimating grade in Manto 0 and similarly variograms for Manto3a used for estimating grade in Manto 3aa.



 

 

 

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Figure 44:

Manto 3 Variogram Maps

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Figure 45:

Variograms of Copper – Manto 3 (X, Y, Z directions)

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Figure 46:

Variograms of Cobalt – Manto 3 (X, Y, Z directions)

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Figure 47:

Variograms of Zinc – Manto 3 (X, Y, Z directions)

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Table 33:

Variogram Models

 

 

Manto 3a

Manto 3

Manto 4

Metal

Structure

Nugget

c1

c2

Nugget

c1

c2

Nugget

c1

c2

Copper

type

 

exp

sph

 

exp

exp

 

exp

sph

 

variance

0.02

0.68

0.3

0.02

0.75

0.23

0.05

0.2

0.75

 

range - X

-

100

1600

-

90

1820

-

155

180

 

range - Y

-

170

1080

-

95

2290

-

180

240

 

range - Z

-

5

6

-

5.5

8

-

5.5

8.5

 

azimuth

292

 

 

292

 

 

330

 

 

Cobalt

type

 

exp

sph

 

exp

exp

 

exp

sph

 

variance

0.02

0.73

0.25

0.02

0.74

0.24

0.1

0.46

0.44

 

range - X

-

100

4000

-

105

640

-

90

350

 

range - Y

-

110

4000

-

80

855

-

175

570

 

range - Z

-

4

6

-

6

9

-

15

7.5

 

azimuth

315

 

 

292

 

 

360

 

 

Zinc

type

 

exp

sph

 

exp

sph

 

exp

sph

 

variance

0.02

0.7

0.28

0.02

0.65

0.33

0.05

0.1

0.85

 

range - X

-

200

300

-

100

435

-

85

105

 

range - Y

-

80

600

-

100

1205

-

90

110

 

range - Z

-

5

2.5

-

8

9

-

7

15

 

azimuth

360

 

 

337

 

 

360

 

 



16.3.4

SEARCH PARAMETERS AND DATA CRITERIA


For each metal, three estimation passes were completed.  Each pass progressively reduced the extent of the search radii or increased the number of data required before a grade estimate was calculated for each block.  Therefore each pass progressively improves the accuracy or confidence of the block estimates.

In the previous resource estimates reported in 2005, search and data parameters shown in Table 34, were used.

Table 34:

Search Parameters – 2005 3D Block Models

Parameter

Manto 2, 3aa,3a,3, 4

Manto 0, 1

Pass

1

2

3

1

2

3

Search Radii (m)

 

 

X – direction

200

280

400

500

750

1000

Y – direction

250

350

500

500

750

1000

Z – direction

2

2

4

2

2

4

Data Criteria

 

 

Min Data

18

8

6

18

8

6

Max Data

32

32

32

32

32

32




 

 

 

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For the current models these search parameters were modified in an attempt to improve the confidence in the grade estimates (Table 35).  

The first pass remains very similar in terms of both radii and data.  There is a small reduction in data required but tighter controls on the configuration of data around the block.  This is achieved but specifying the number of search octants that must have at least one data composite used in the grade estimation.  

For the second pass the search radii have been expanded by a factor of 2, but much the same data and configuration as used in pass one has been applied.  The third pass uses the same search radii but slightly stricter data configuration.  Only the pass 2 parameters are significantly different.  The purpose of making these changes was to increase the data number used and to improve the configuration of these data around a block.  This could only be achieved by expanding the search radii.

Table 35:

Search Parameters – Current 3D Block Models

Parameter

Manto 3a,3

Pass

1

2

3

Search Radii (m)

 

X – direction

200

400

400

Y – direction

250

500

500

Z – direction

2

2

4

Data Criteria

 

Min Data

14

14

7

Octants

4

4

2

Max Data

32

32

32


Resource estimates for Manto 0, 3aa and 4 have not been updated at this stage so the original search and data parameters still apply to these models.

Block discretization has been changed from 5 x 5 x 1 used in 2005 to 3 x 8 x 1.  This simply results in a more even distribution of estimation nodes throughout a block.

16.3.5

MODEL CODING


The block models have been coded as follows:

·

The proportion of a block below the topographic surface – Topo item

·

The proportion of a block below the upper surface of the Manto – ore percent item

·

A flag to specify whether a block is oxide or sulphide

·

A flag to specify whether a block is within the land held by the company – claim item

·

A flag to specify whether a block is within the areas affected by historic workings – mined item.



 

 

 

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Only those blocks that are below both the topographic surface, the upper manto surface and within the land boundary held by the company have a metal grade estimate.

The oxide – sulphide distinction was based on polygons provided by MMB.

16.3.6

RESOURCE CLASSIFICATION – 3D BLOCK MODELS


The resources have been classified on the basis of the 3 estimation passes as Measured, Indicated and Inferred.  

A block can only be classified as:

·

Measured if 14 individual assay composites from at least 4 search octants are located within a search ellipse with radii of 200 m x 250 m x 2 m.

·

Indicated if 14 individual assay composites from at least 4 search octants are located within a search ellipse with radii of 400 m x 500 m x 2 m.

·

Inferred if 7 individual assay composites from at least 2 search octants are located within a search ellipse with radii of 400 m x 500 m x 4 m.


Manto 0 is classified only as inferred because of the significantly wider drill hole spacing, approximately 500 m x 450 m.

Mantos 1 and 3 have been extensively mined in the past.  The location of the old mines are well known (Wilson 1955), however due to the method of mining adopted the precise location and extent of voids, pillars and back-filled excavations are not known.  

Ore was mined using a shortwall mining method, however, only high grade ore (>3.5% Cu) was taken to the surface for processing.  It is estimated that the ore processed amounted to only 40% of mined material, the remaining 60%, referred to as ‘Retaque’, was side-cast into old excavations as mining progressed.  Exploration drill holes have intersected voids in about 50 instances and a similar number of Retaque intervals have been identified.  The estimation process has been carried out effectively ignoring this issue other than treating voids as ‘missing’ or ‘no data'.

The resource, however, should be modified to account for the previous mining activities.  This could be achieved by simply removing an amount of ore equal to the known treated ore tonnage, 13.6Mt.  However, this may be overly conservative since many of the workings have collapsed and the void volume is now lower than was originally the case.  

The lower grade clayey breccias, that typically overlie the higher grade laminated manto material that was mined, have collapsed or sagged into the original voids and areas of back-fill, prior to the resource drilling programs of the 1990s.  The higher grade ore mined and processed has effectively been ‘replaced’ by lower grade material, firstly due to back-filling as mining progressed and later by the collapse of the workings where voids were left.  This replaced material has been sampled and can be considered in situ.  Voids still remain though and these have been accounted for by:




 

 

 

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·

digitising a simplified footprint shape around the area of historic workings (Figure 48)

·

the footprint area was extended up 2 m to form a 3D object that enveloped the workings

·

all model blocks that fell within this ‘mined envelope’ are flagged as ‘mined’

·

determining the proportion of sample intervals within the area affected by past mining that intersected voids

·

factoring the tonnes by the proportion of sample interval that were voids.


For manto 3 a total of 335 m of drill hole intersection occur within the mined volume of which 12% are voids (see Figure 48).  It is appropriate therefore, to factor the tonnes down by 12%.

Figure 48:

Manto 3 Old Mine Areas – Location Drill Holes Which Encountered Voids

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16.3.7

RESOURCE ESTIMATES – 3D BLOCK MODELS


Grade estimation in 2005 was completed using a specialist mining software package known as “Minesight-3D” (v 3.2), available from Metech Pty Ltd.  

For the current resource models a new Ordinary Kriging routine added to the GS3 software used for data analysis and variogram estimation has been used.  The benefit of using this new software option is that now the entire data analysis and resource modelling process can be completed within one software package.

The Boléo Resource includes Measured, Indicated and Inferred Categories.

Resources are quoted at copper equivalent cutoff grades of 0.5% and 1.0%.  Copper Equivalent is a means of describing the effective grade of a polymetallic resource by bringing the value of all the pay metals (copper, cobalt and zinc in this instance) to bear.   The proportion of each metal in the copper equivalent is weighted in the formula according to its 'value' i.e., its price.  The prices used in the formula below for each of the metals are based on long term average prices as follows:

·

copper – US$0.95/lb

·

cobalt – US$12/lb

·

zinc – US$0.45/lb.


The copper equivalent formula used is therefore:

·

 CuEq% = Cu% + 12Co%/0.95 + 0.45Zn%/0.95.


The resource models for mantos 0, 3aa and 4 are unchanged from the 2005 resource estimates.  Manto 3 and 3a are new resource estimates, incorporating the latest drilling results.  Manto 1 and 2 are not reported as 3D block models and have been removed from the following tables.  These models are discussed and reported as seam models in a following section.

MEASURE AND INDICATED RESOURCE


Table 36:

Measured and Indicated Resource at 0.5% CuEq. Cutoff

0.5% CuEq Cutoff

Manto

0

3aa

3a

3

4

Measured

Tonnes (106)

 

1.6

10.4

39.4

2.4

 

CuEq.%

 

2.289

2.03

2.24

1.82

 

Cu%

 

0.647

0.43

1.02

0.96

 

Co%

 

0.096

0.101

0.089

0.056

 

Zn%

 

0.91

0.67

0.30

0.33

 

 

 

 

 

 

 

Indicated

Tonnes (106)

 

1.8

35.7

75.5

27.6

 

CuEq.%

 

2.039

1.36

2.03

1.10

 

Cu%

 

0.55

0.36

1.09

0.50

 

Co%

 

0.087

0.058

0.059

0.037

 

Zn%

 

0.816

0.57

0.42

0.29



 

 

 

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0.5% CuEq Cutoff

Manto

0

3aa

3a

3

4

Total

Tonnes (106)

 

3.4

46.1

114.9

30.0

 

CuEq.%

 

2.16

1.51

2.10

1.16

 

Cu%

 

0.60

0.38

1.07

0.53

 

Co%

 

0.091

0.068

0.069

0.039

 

Zn%

 

0.86

0.59

0.38

0.29


Table 37:

Measured and Indicated Resource at 1.0% CuEq. Cutoff

0.5% CuEq Cutoff

Manto

0

3aa

3a

3

4

Measured

Tonnes (106)

 

1.6

9.0

36.5

1.4

 

CuEq.%

 

2.318

2.20

2.35

2.57

 

Cu%

 

0.656

0.49

1.09

1.42

 

Co%

 

0.097

0.110

0.089

0.075

 

Zn%

 

0.92

0.70

0.31

0.44

 

 

 

 

 

 

 

Indicated

Tonnes (106)

 

1.7

24.2

65.8

10.6

 

CuEq.%

 

2.093

1.63

2.21

1.77

 

Cu%

 

0.565

0.48

1.22

0.89

 

Co%

 

0.090

0.067

0.062

0.054

 

Zn%

 

0.839

0.63

0.43

0.41

 

 

 

 

 

 

 

Total

Tonnes (106)

 

3.3

33.2

102.3

12.0

 

CuEq.%

 

2.20

1.78

2.26

1.86

 

Cu%

 

0.61

0.48

1.17

0.95

 

Co%

 

0.093

0.079

0.072

0.057

 

Zn%

 

0.88

0.65

0.39

0.41



INFERRED RESOURCE


Table 38:

Inferred Resource

Cutoff

Manto

0

3aa

3a

3

4

0.5% CuEq.

Tonnes (106)

9.7

0.5

25.2

58.5

56.9

CuEq.%

0.56

1.92

1.19

1.51

0.87

Cu%

0.04

0.58

0.34

0.63

0.38

Co%

0.012

0.083

0.046

0.048

0.030

Zn%

0.77

0.64

0.58

0.58

0.23

1.0% CuEq.

Tonnes (106)

0.0

0.4

14.3

40.8

12.7

CuEq.%

1.11

2.03

1.50

1.81

1.47

Cu%

0.02

0.63

0.49

0.83

0.79

Co%

0.039

0.086

0.056

0.054

0.041

Zn%

1.26

0.66

0.64

0.65

0.32





 

 

 

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16.3.8

MODEL VERIFICATION


The model grade estimates have been verified by plotting block grades against assay composite grades, in plan, for different levels of the model.  Plans of the basal level of Manto 3 are shown for copper, cobalt, and zinc (Figures 49 to 56).

Figure 49:

Manto 3 Copper – Block Model with Assay Composites

[bajatechrep129.gif]

Notes:

    Total Resource (basal layer 0 m to 1 m). Box highlights area shown in detail in Figure 50.






 

 

 

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Figure 50:

Detail of Manto 3 Copper – Block Model with Assay Composites

[bajatechrep131.gif]

Note:

    the legend for Fig 50 is as per Figure 49.





 

 

 

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Figure 51:

Manto 3 Cobalt – Block Model with Assay Composites

[bajatechrep133.gif]

Notes:

    Total Resource (basal layer 0-1 m).  Box highlights area shown in detail in Figure 52.





 

 

 

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Figure 52:

Detail of Manto 3 Cobalt – Block Model with Assay Composites

[bajatechrep135.gif]

Note:

    the legend for Fig 52 is as per Figure 51.





 

 

 

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Figure 53:

Manto 3 Zinc – Block Model with Assay Composites

[bajatechrep137.gif]

Notes:

    Total Resource (basal layer 0-1 m).  Box highlights area shown in detail in Figure 54.





 

 

 

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Figure 54:

Detail of Manto 3 Zinc – Block Model with Assay Composites

[bajatechrep139.gif]

Note:

    the legend for Fig 54 is as per Figure 53.

The appropriateness of the resource classification has been assessed by plotting the classification against the drill hole collar locations (Figure 55).  The pattern shows a tight concentric zonation around areas of closer spaced drilling.  Note that the occurrence of a drill hole either in or very close to a block does not guarantee a high level of resource classification.






 

 

 

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Figure 55:

Manto 3 Resource Classification – Block Model with Drill Hole Locations

[bajatechrep141.gif]

Note:

    Measured – Red, Indicated – Yellow, Inferred – Blue (basal layer 0 to 1 m).

16.3.9

COMPARISON WITH 2005 3D BLOCK MODELS


The reported resource estimates for Manto 3a and 3 are compared against the 2005 resource estimates.  

For Manto 3 the differences between the reported resource from 2005 and the new resource are attributable to three causes:

·

addition of new drill hole data and re-interpretation of the manto footwall

·

changes to the kriging search parameters



 

 

 

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·

re-classification of the retaque proportion of the 2005 resource as Measured, Indicated or Inferred.


In the table below the new Manto 3 resource is reported in four columns with a different label in the row headed “Year of Estimate”.  The data reported in each of these columns is:

·

2005 – As estimated and reported in the year 2005

·

2007 A – New estimate, new drill hole data, Retaque tonnes segregated out

·

2007 B – New estimate, new drill hole data. Retaque tonnes allocated to Measured, Indicated or Inferred resource based on kriging search and data parameters.


The differences between 2005 data and 2007A are due to the new data, the changes made to the kriging search parameters and the identification of “false bottom” manto footwalls.  The false bottom footwalls have allowed low grade or barren basal intervals to be removed from the grade estimation.  When the effect of the false bottom is included in the resource calculations it allows for increased mining heights at similar grades. This results in a small net increase in tonnage.

The differences between 2007A and 2007B are only due to the removal of the Retaque category and re-allocation of this material to either Measured, Indicated, or Inferred resource according to kriging search and data parameters.  This step adds approximate 10 Mt to the combined measured and indicated resource and elevates the grade from 1.97% CuEq. to 2.10%.  This happens because the Retaque material represents the old mined areas which are generally the higher grade areas.

Table 39:

Manto 3 – Current vs. 2005 Resource Model, Measured and Indicated Resources

 

 

Manto 3

Cutoff Grade

0.5% CuEq.

1.0% CuEq.

Year of Estimate

2005

2007 A

2007 B

2005

2007 A

2007 B

Measured

Tonnes (106)

34.4

37.5

39.4

31.5

34.6

36.5

 

CuEq%

2.14

2.21

2.24

2.27

2.33

2.35

 

Cu%

0.88

0.98

1.02

0.95

1.05

1.09

 

Co%

0.089

0.086

0.089

0.094

0.090

0.089

 

Zn%

0.29

0.30

0.30

0.30

0.30

0.31

 

 

 

 

 

 

 

 

Indicated

Tonnes (106)

62.5

63.0

75.5

51.7

53.3

65.8

 

CuEq%

1.87

1.84

2.03

2.10

2.02

2.21

 

Cu%

0.93

0.92

1.09

1.08

1.05

1.22

 

Co%

0.059

0.058

0.059

0.064

0.061

0.062

 

Zn%

0.41

0.40

0.42

0.43

0.41

0.43

 

 

 

 

 

 

 

 

Total

Tonnes (106)

96.9

100.5

114.9

83.2

87.9

102.3

 

CuEq%

1.97

1.98

2.10

2.16

2.14

2.26

 

Cu%

0.91

0.94

1.07

1.03

1.05

1.17

 

Co%

0.070

0.068

0.069

0.075

0.072

0.072

 

Zn%

0.37

0.36

0.38

0.38

0.37

0.39



 

 

 

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Table 40:

Manto 3 - Current vs. 2005 Resource Model, Inferred Resources

 

 

Manto 3

Cutoff Grade

0.5% CuEq.

1.0% CuEq.

Year of Estimate

2005

2007 A

2007 B

2005

2007 A

2007 B

Inferred

Tonnes (106)

54.9

55.5

58.5

39.7

37.8

40.8

 

CuEq%

1.51

1.46

1.51

1.79

1.77

1.81

 

Cu%

0.65

0.59

0.63

0.83

0.78

0.83

 

Co%

0.049

0.047

0.048

0.054

0.054

0.054

 

Zn%

0.52

0.58

0.58

0.58

0.65

0.65

 

 

 

 

 

 

 

 

“Retaque”

Tonnes (106)

18.2

17.4

 

17.8

17.4

 

 

CuEq%

2.57

2.86

 

2.60

2.87

 

 

Cu%

1.60

1.83

 

1.62

1.84

 

 

Co%

0.058

0.062

 

0.058

0.062

 

 

Zn%

0.52

0.52

 

0.52

0.52

 

 

 

 

 

 

 

 

 

Total

Tonnes (106)

73.1

72.9

58.5

57.5

55.2

40.8

 

CuEq%

1.77

1.79

1.51

2.04

2.12

1.81

 

Cu%

0.88

0.88

0.63

1.07

1.11

0.83

 

Co%

0.051

0.051

0.048

0.056

0.057

0.054

 

Zn%

0.52

0.57

0.58

0.56

0.61

0.65



Manto 3a shows an overall drop in tonnes at each cutoff grade although the amount and grade of Measured and Indicated material remains very similar.  Infill drill holes for Manto 3a generally reported low grades and where located outside the main area of mineralization for this Manto.  The differences between the two models are mostly attributable to the change in kriging parameters.  

Table 41:

Comparison between Current and 2005 Resource Models, Manto 3a

 

 

Manto 3a

CuEq. Cutoff %

0.5

1.0

Year of Estimate

2005

2007

2005

2007

Measured

Tonnes (106)

9.7

10.4

8.4

9.0

 

CuEq%

2.12

2.03

2.31

2.20

 

Cu%

0.45

0.43

0.51

0.49

 

Co%

0.106

0.101

0.115

0.110

 

Zn%

0.71

0.67

0.74

0.70

Indicated

Tonnes (106)

38.1

35.7

23.8

24.2

 

CuEq%

1.32

1.36

1.63

1.63

 

Cu%

0.35

0.36

0.47

0.48

 

Co%

0.055

0.058

0.067

0.067

 

Zn%

0.59

0.57

0.66

0.63



 

 

 

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Manto 3a

CuEq. Cutoff %

0.5

1.0

Year of Estimate

2005

2007

2005

2007

Measured & Indicated

Tonnes (106)

47.8

46.1

32.2

33.2

 

CuEq%

1.48

1.51

1.81

1.78

 

Cu%

0.37

0.38

0.48

0.48

 

Co%

0.065

0.068

0.080

0.079

 

Zn%

0.61

0.59

0.68

0.65

Inferred

Tonnes (106)

29.6

25.2

17.8

14.3

 

CuEq%

1.25

1.19

1.55

1.50

 

Cu%

0.38

0.34

0.54

0.49

 

Co%

0.047

0.046

0.057

0.056

 

Zn%

0.58

0.58

0.63

0.64



16.4

SEAM MODELS


Seam models have been produced for Manto 1, 2 and 3.  For Mantos 1 and 2 these are the only models reported as these mantos will only be mined from underground.  Manto 3 however will be mined from both open cuts and under ground developments.  Therefore Manto 3 seam models reported here are not in addition to, but are a sub-set of and are included in the reported open cut 3D block models.

16.4.1

COMPOSITE LENGTH AND BLOCK DIMENSIONS


The type of the seam composite used from each drill hole has a big impact on the resource estimate.  Three alternative strategies for seam definition are:

·

A geological seam.  The full logged manto interval is composited into a single seam, from the footwall to the top of the last identified manto interval.

·

A grade seam.  A grade cutoff is used to define the top and bottom of the seam.

·

A mining seam.  The base of the seam is defined by the manto footwall whilst the top is defined by mining constraints.


The grade of copper mineralization generally decreases from higher grade towards the base to lower grades at the top of each manto.  If a geological seam is used significant lengths of material below economic cutoff grades can be included in the composite.  If the thickness of the manto is greater than the minimum mining height then the inclusion of sub-economic low grade material will result in an unrealistic lower grade resource than bears no relation to the anticipated mining method.  Therefore geological composites have not been used.

Definition composites by grade alone can be problematical.  There is not obvious ‘natural’ cutoff in the data that can be consistently applied throughout.  The main problem with grade defined seams is the confidence in the lateral continuity of the seam.  Seams defined this way typically show quite variable heights above the known manto floor and many intervals would include a



 

 

 

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certain amount of internal waste.  Both of these issues show that the lateral continuity of grade seams cannot be demonstrated.  Variogram analysis which shows relatively short ranges also supports this contention.  The grade seam method can only be used if the seam can be referenced against a defined geological horizon that has demonstrated continuity.

The manto footwall is a well defined geological entity with well established and demonstrated continuity.  Over much of Manto 3 for example a boulder conglomerate defines the floor.  Elsewhere though, other consistent lithologies are present, such as the calcareous sandstone unit that occurs beneath parts of Manto 1 and the barren clayey unit that overlies the conglomerate in certain parts of Manto3.  

To ensure continuity of the seams the manto footwall or floor has to be used to define the base of each composite.  The thickness of the seam can then be based on any other criteria.

The most sensible method to determine the seam thickness is to use a combination of grade and mining constraints.  Mining studies at Boléo have determined a minimum height of 1.8 m and a maximum height of 4.2 m.  Consequently it was decided to produce seam composites that reflect this.  For each hole two different seam composites would be produced, both with a base defined by the manto footwall:

·

A 1.8 m seam.  Where the full natural manto interval in a hole is actually less than 1.8 m a certain amount of barren material is added to bring the thickness up to 1.8 m.  Material greater than 1.8 m above the Manto floor is ignored and not used in the resource estimates.

·

A variable thickness seam from 1.8 m to 4.2 m.  As for the 1.8 m seam where the full natural manto interval in a hole is less than 1.8 m a certain amount of barren material is added to bring the thickness up to 1.8 m.  Where the full natural thickness of the seam is greater than 1.8 m then additional increments of 0.6 m, from 1.8 m to 4.2 m were added to the seam.  If the grade of these increments were greater than a specified grade the seam thickness would be increased to include these increments.  If the grade of additional increments was not above the specified grade then the minimum 1.8 m composite would be used.  Material greater than 4.2 m above the Manto floor is ignored and not used in the resource estimates.


The grade threshold used justify inclusion of additional increments was 1% Cu.  Increments were determined at 2.4 m, 3.0 m, 3.6 m, and 4.2 m.

Model block dimensions were the same in terms of X (50m) and Y (100m) as used in the 3D block models.  The minimum thickness models would have a fixed vertical height of 1.8 m, whilst models produced with seam composites up to 4.2 m would have variable height as determined by kriging seam thickness.

After producing the seam composites using the method described above it was clear that for Mantos 1 and 2 that only the minimum 1.8 m seams were appropriate.  Consequently only minimum thickness resource estimates would be produced.



 

 

 

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16.4.2

UNIVARIATE STATISTICS OF DATA COMPOSITES


Univariate statistics of Copper, Cobalt and Zinc for Manto 1, 2 and 3 are shown below (Tables 42 to 44).  Manto 1 has been separated into two geological domains (D1, D2) based on interpretation of palaeo geography at the time of deposition.  Domain 2 is a waste domain overlying a rise in the submerged basement floor, while D1 is the mineralized domain distal to the basement rise.

Generally the means of the seam composites are higher than the individual 1m composites and the coefficients of variation are lower.

Table 42:

Univariate Statistics of Assay Composites – Copper

 

Manto

Copper

1

1

1

2

3

3

Seam/Domain

1.8

1.8/D1

1.8/D2

1.8

1.8

4.2m

Mean

0.872

2.132

0.309

0.421

1.406

1.388

CV

1.612

0.79

2.428

1.976

0.936

0.894

Min

0.003

0.13

0.003

0.001

0.004

0.004

Q1

0.016

0.927

0.012

0.056

0.473

0.506

Median

0.133

1.757

0.03

0.164

1.114

1.141

Q3

1.155

2.538

0.19

0.403

1.893

1.91

Max

8.744

8.744

4.928

7.585

11.241

11.241

IQR

1.139

1.611

0.178

0.347

1.42

1.404

Data

214

66

148

477

873

873


Table 43:

Univariate Statistics of Assay Composites – Cobalt

 

Manto

Copper

1

1

1

2

3

3

Seam/Domain

1.8

1.8/D1

1.8/D2

1.8

1.8

4.2 m

Mean

0.049

0.089

0.031

0.056

0.088

0.088

CV

1.097

0.515

1.51

1.571

0.895

0.839

min

0

0.013

0

0.001

0.002

0.002

Q1

0.01

0.054

0.008

0.018

0.034

0.034

Median

0.025

0.085

0.014

0.041

0.069

0.07

Q3

0.075

0.113

0.032

0.074

0.119

0.123

Max

0.372

0.201

0.372

1.683

1.215

0.95

IQR

0.065

0.059

0.024

0.056

0.085

0.089

Data

214

66

148

477

873

873




 

 

 

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Table 44:

Univariate Statistics of Assay Composites – Zinc

 

Manto

Copper

1

1

1

2

3

3

Seam/Domain

1.8

1.8/D1

1.8/D2

1.8

1.8

4.2m

Mean

0.995

0.638

1.154

1.029

0.405

0.405

CV

1.666

0.635

1.692

0.979

1.372

1.34

Min

0.016

0.062

0.016

0.018

0.008

0.008

Q1

0.353

0.322

0.387

0.463

0.192

0.197

Median

0.637

0.568

0.691

0.716

0.29

0.29

Q3

0.895

0.794

0.945

1.203

0.46

0.46

Max

14.265

2.446

14.265

7.628

9.094

9.094

IQR

0.542

0.472

0.558

0.74

0.268

0.263

Data

214

66

148

477

873

873



16.4.3

SPATIAL CONTINUITY OF GRADE


Despite the reduction in sample variance the continuity determined from the experimental variograms is not significantly improved compared to the variograms determined from the 1 m composite data.

Variograms of seam composites for Manto 3 are shown below for copper, cobalt and zinc.

Figure 56:

Variograms of Copper – Manto 3 (X, Y, directions)

[bajatechrep143.gif]



 

 

 

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Figure 57:

Variograms of Cobalt – Manto 3 (X, Y, directions)

[bajatechrep145.gif]

Figure 58:

Variograms of Zinc – Manto 3 (X, Y, directions)

[bajatechrep147.gif]

Full variogram models are shown in Table 45.

Table 45:

Variogram Models – Seam Composites

 

 

Manto 1

Manto 2

Manto 3

Metal

Structure

Nugget

c1

c2

Nugget

c1

c2

Nugget

c1

c2

Copper

type

 

Exp

sph

 

exp

sph

 

exp

sph

 

variance

0.02

0.7

0.28

0.02

0.58

0.4

0.02

0.85

0.13

 

range – X

-

200

130

-

110

130

-

155

600

 

range – Y

-

360

1150

-

130

280

-

130

1800

 

azimuth

330

 

 

315

 

 

330

 

 

Cobalt

type

 

Exp

sph

 

exp

exp

 

exp

sph

 

variance

0.02

0.69

0.29

0.02

0.52

0.46

0.02

0.83

0.15

 

range – X

-

100

500

-

120

210

-

170

1665

 

range – Y

-

175

2805

-

120

350

-

160

1300

 

azimuth

330

 

 

315

 

 

360

 

 



 

 

 

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Manto 1

Manto 2

Manto 3

Metal

Structure

Nugget

c1

c2

Nugget

c1

c2

Nugget

c1

c2

Zinc

type

 

Exp

sph

 

exp

sph

 

exp

sph

 

variance

0.02

0.66

0.32

0.02

0.65

0.33

0.02

0.4

0.58

 

range – X

-

185

200

-

245

270

-

220

730

 

range – Z

-

460

940

-

240

340

-

200

1200

 

azimuth

330

 

 

315

 

 

360

 

 



16.4.4

SEARCH PARAMETERS AND DATA CRITERIA


The same search strategy with 3 consecutive estimation passes has been used for the seam models as used for the 3D block models.  The only difference being that the data requirements are modified to take into account the single seam composites rather than several individual 1m composites for each hole.

Table 46:

Search Parameters – Current 3D Block Models

Parameter

Manto 1, 2, 3

Pass

1

2

3

Search Radii (m)

 

 

 

X – direction

200

400

400

Y – direction

250

500

500

Data Criteria

 

 

 

Min Data

4

4

2

Octants

4

4

2

Max Data

8

8

8



16.4.5

RESOURCE CLASSIFICATION – SEAM MODELS


The resources have been classified on the basis of the three estimation passes as Measured, Indicated and Inferred.  A block can only be classified as:

·

Measured if 4 seam composites from at least 4 search octants are located within a search ellipse with radii of 200 m x 250 m.

·

Indicated if 4 seam composites from at least 4 search octants are located within a search ellipse with radii of 400 m x 500 m.

·

Inferred if 2 seam composites from at least 2 search octants are located within a search ellipse with radii of 400 m x 500 m.


No measured resource is reported for Manto 1.  Blocks that fell in this category based on the criteria defined above did not form coherent domains.  Additional drilling currently being completed will help resolve this.



 

 

 

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16.4.6

RESOURCE ESTIMATES – SEAM MODELS


The Boléo seam models include Measured, Indicated and Inferred Categories.

Resources are quoted at the 0.5% and 1.0% CuEq. cutoff grades.

·

CuEq.%= Cu % + 12Co%/0.95 + 0.45Zn%/0.95.

For Manto 3 both 3D block model and seam models are presented.  Manto 3 is generally thicker than the minimum mining height of 1.8 m so both a 1.8 m and variable (1.8 - 4.2 m) seam model is reported.  The seam estimates for Manto 3 are not in addition to the 3D block model report previously in this document.  The seam models are a sub-set of the 3D block model.

Table 47:

Seam Models – Measured and Indicated Resource at 0.50% CuEq.

0.5% CuEq

Manto

1

2

3

3

Seam Height

1.8

1.8

1.8

1.8 - 4.2

Measured

Tonnes (106)

 

5.6

17.4

20.0

 

CuEq%

 

1.88

2.93

2.93

 

Cu%

 

0.51

1.56

1.56

 

Co%

 

0.067

0.096

0.096

 

Zn%

 

1.12

0.34

0.34

Indicated

Tonnes (106)

8.3

24.5

25.0

27.7

 

CuEq%

2.61

1.64

2.62

2.63

 

Cu%

1.34

0.46

1.54

1.54

 

Co%

0.067

0.056

0.067

0.067

 

Zn%

0.90

1.01

0.50

0.50

Total

Tonnes (106)

8.3

30.1

42.4

47.7

 

CuEq%

2.61

1.68

2.75

2.76

 

Cu%

1.34

0.47

1.55

1.55

 

Co%

0.067

0.058

0.079

0.079

 

Zn%

0.90

1.03

0.43

0.43


Table 48:

Seam Models – Measured and Indicated Resource at 1.0% CuEq.

1.0% CuEq

Manto

1

2

3

3

Seam Height

1.8

1.8

1.8

1.8 - 4.2

Measured

Tonnes (106)

 

4.7

17.1

19.8

 

CuEq%

 

2.07

2.97

2.96

 

Cu%

 

0.57

1.58

1.58

 

Co%

 

0.074

0.097

0.096

 

Zn%

 

1.21

0.34

0.34

Indicated

Tonnes (106)

7.1

18.9

23.6

26.3

 

CuEq%

2.92

1.89

2.72

2.73

 

Cu%

1.54

0.55

1.62

1.61



 

 

 

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1.0% CuEq

Manto

1

2

3

3

 

Co%

0.075

0.065

0.069

0.069

 

Zn%

0.91

1.10

0.50

0.51

Total

Tonnes (106)

7.1

23.6

40.7

46.1

 

CuEq%

2.92

1.93

2.83

2.83

 

Cu%

1.54

0.55

1.60

1.60

 

Co%

0.075

0.067

0.081

0.081

 

Zn%

0.91

1.12

0.43

0.44

Note:

    Manto 3 resource quoted above is a sub-set of NOT additional to the 3D block model.

Table 49:

Seam Models - Inferred Resource

Cutoff

Manto

1

2

3

3

Seam Height

1.8

1.8

1.8

1.8 - 4.2

0.5% CuEq.

Tonnes (106)

10.2

51.3

26.4

28.4

CuEq%

2.19

1.48

1.95

2.01

Cu%

0.87

0.34

0.93

0.97

Co%

0.063

0.048

0.053

0.055

Zn%

1.13

1.13

0.73

0.73

1.0% CuEq.

Tonnes (106)

8.6

37.5

22.7

24.8

CuEq%

2.47

1.75

2.14

2.19

Cu%

0.99

0.41

1.06

1.09

Co%

0.071

0.056

0.057

0.059

Zn%

1.22

1.32

0.78

0.77

Note:

     Manto 3 resource quoted above is a sub-set of, NOT additional to, the 3D block model.

16.4.7

MODEL VERIFICATION


The model grade estimates have been verified by plotting block grades against assay composite grades, in plan, for different levels of the model.  Plans of Manto 1 and 3 are shown for copper (Figures 59 to 64).





 

 

 

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Figure 59:

Manto 3 Copper – Block Model with Assay Composites

[bajatechrep149.gif]

Note:

     Box highlights area shown in detail in Figure 60.





 

 

 

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Figure 60:

Detail of Manto 3 Copper

[bajatechrep151.gif]

Note

   : The legend for Fig. 60 is as per Fig. 59.





 

 

 

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Figure 61:

Manto 1 Copper – Block Model with Assay Composites

[bajatechrep153.gif]

Note:

     Box highlights area shown in detail in Figure 62





 

 

 

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Figure 62:

Detail of Manto 3 Copper

[bajatechrep155.gif]

Note:

    The legend for Fig. 60 is as per Fig. 59.





 

 

 

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Figure 63:

Manto 3 Resource Classification – Block Model with Drill Hole Locations

[bajatechrep157.gif]

Note:

    Measured – Red, Indicated – Yellow, Inferred – Blue (basal layer 0-1 m)





 

 

 

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Figure 64:

Manto 3 Resource Classification – Block Model with Drill Hole Locations

[bajatechrep159.gif]

Note:

      Measured – Red, Indicated – Yellow, Inferred – Blue.  (basal layer 0-1 m)

16.5

TOTAL RESOURCE


The tables below combine the 3D block models and the seam models in a single table for each cutoff grade.

·

Manto 1 and 2 are seam models

·

Manto 3aa, 3a, 3 and 4 are the 3D block models.





 

 

 

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Table 50:

Combined Resource at 0.50% CuEq.

Cutoff
CuEq

0.50%

Manto

0

1

2

3aa

3a

3

4

Model Type 

3D blk

Seam

Seam

3D blk

3D blk

3D blk

3D blk

Measured

Tonnes (106)

 

 

5.6

1.6

10.4

39.4

2.4

 

CuEq%

 

 

1.88

2.29

2.03

2.24

1.82

 

Cu%

 

 

0.51

0.65

0.43

1.02

0.96

 

Co%

 

 

0.067

0.096

0.101

0.089

0.056

 

Zn%

 

 

1.12

0.91

0.67

0.30

0.33

Indicated

Tonnes (106)

 

8.3

24.5

1.8

35.7

75.5

27.6

 

CuEq%

 

2.61

1.64

2.04

1.36

2.03

1.10

 

Cu%

 

1.34

0.46

0.55

0.36

1.09

0.50

 

Co%

 

0.067

0.056

0.087

0.058

0.059

0.037

 

Zn%

 

0.90

1.01

0.82

0.57

0.42

0.29

Measured & Indicated

Tonnes (106)

 

8.3

30.1

3.4

46.1

114.9

30.0

 

CuEq%

 

2.61

1.68

2.16

1.51

2.10

1.16

 

Cu%

 

1.34

0.47

0.60

0.38

1.07

0.53

 

Co%

 

0.067

0.058

0.091

0.068

0.069

0.039

 

Zn%

 

0.90

1.03

0.86

0.59

0.38

0.29

Inferred

Tonnes (106)

9.7

10.2

51.3

0.5

25.2

58.5

56.9

 

CuEq%

0.56

2.20

1.48

1.92

1.19

1.51

0.87

 

Cu%

0.04

0.87

0.34

0.58

0.34

0.63

0.38

 

Co%

0.012

0.063

0.048

0.083

0.046

0.048

0.030

 

Zn%

0.77

1.13

1.13

0.64

0.58

0.58

0.23


Table 51:

Combined Resource at 1.0% CuEq.

Cutoff
CuEq

1.0%

Manto

0

1

2

3aa

3a

3

4

Model Type 

3D blk

Seam

Seam

3D blk

3D blk

3D blk

3D blk

Measured

Tonnes (106)

 

 

4.7

1.6

9.0

36.5

1.4

 

CuEq%

 

 

2.07

2.32

2.20

2.35

2.57

 

Cu%

 

 

0.57

0.66

0.49

1.09

1.42

 

Co%

 

 

0.074

0.097

0.110

0.089

0.075

 

Zn%

 

 

1.21

0.92

0.70

0.31

0.44

Indicated

Tonnes (106)

 

7.1

18.9

1.7

24.2

65.8

10.6

 

CuEq%

 

2.92

1.89

2.09

1.63

2.21

1.77

 

Cu%

 

1.54

0.55

0.57

0.48

1.22

0.89

 

Co%

 

0.075

0.065

0.090

0.067

0.062

0.054

 

Zn%

 

0.91

1.10

0.84

0.63

0.43

0.41

Measured & Indicated

Tonnes (106)

 

7.1

23.6

3.3

33.2

102.3

12.0

 

CuEq%

 

2.92

1.93

2.20

1.78

2.26

1.86

 

Cu%

 

1.54

0.55

0.61

0.48

1.17

0.95



 

 

 

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Cutoff
CuEq

1.0%

Manto

0

1

2

3aa

3a

3

4

Model Type 

3D blk

Seam

Seam

3D blk

3D blk

3D blk

3D blk

 

Co%

 

0.075

0.067

0.093

0.079

0.072

0.057

 

Zn%

 

0.91

1.12

0.88

0.65

0.39

0.41

Inferred

Tonnes (106)

0.03

8.6

37.5

0.4

14.3

40.8

12.7

 

CuEq%

1.11

2.47

1.75

2.03

1.50

1.81

1.47

 

Cu%

0.02

0.99

0.41

0.63

0.49

0.83

0.79

 

Co%

0.039

0.071

0.056

0.086

0.056

0.054

0.041

 

Zn%

1.26

1.22

1.32

0.66

0.64

0.65

0.32



16.6

MINERAL RESERVE ESTIMATES


16.6.1

REPORTING STATUS


Mineral Resource and Reserve estimates for the Boléo Copper Project are being prepared to comply with the requirements of the Canadian National Instrument 43-101 (NI 43-101).  Mine planning is well advanced on the open cut and underground sections of the orebody but the resource drilling conducted in late 2006 is yet to be incorporated in the resource model.  

Mineral Reserve Estimates will not be stated until the final resource models, including the mineral resource categories, are available for inclusion in the mine plan.  This is planned to occur in March 2007.

Any mining quantities or grades expressed in this Updated Preliminary Economic Assessment are subject to verification and amendment against the final resource model and as such they do not constitute a Mineral Reserve Estimate.




 

 

 

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17

OTHER RELEVANT DATA AND INFORMATION

 

 


17.1

MINING AND MINE DESIGN


17.1.1

OPEN CUT MINING


Open cut operations at Boléo will consist of ore mining from the shallow sections of Manto 3 and overlying mantos in the northern half of the deposit and quarrying of limestone from the Dos de Abril area in the north western corner of the mining area.

17.1.2

COPPER PITS


Curator undertook extensive close spaced drilling over the northern section of the deposit during the 1990s to define an Open cut resource, mainly on Mantos 3 and 3A.  Although the current mine plan focuses on extraction of the deeper, higher grade resources in Mantos 1, 2 and 3 the close drilling means that the areas most suitable for open cut mining are at a higher level of resource confidence and it is anticipated that they will be classed as either Indicated or Measured Mineral Resources and thus be available for definition of Mineral Reserves under NI 43-101.

RESOURCE MODEL


Since open cut operations allow selective mining based on close spaced grade control sampling in the cuts, a block model was used as opposed to the single layer seam models used for underground mining.  Block grades were estimated as whole block values using ordinary kriging with a block size of:

·

100 m north south

·

50 m east west

·

1 m vertically.


This allows for selecting a particular mining block vertically as well as laterally.

MINING LOSS AND DILUTION


Grade control drilling is allowed for on a 20 m x 20 m pattern with samples every 0.5 m vertically through the mantos.  The open cut mine plan in this report uses whole block grade estimates.  The increased definition provided by the grade control should allow exclusion of lower grade portions of the large resource blocks.  For this reason it is assumed that the whole block estimates represent a diluted resource and no further dilution adjustments are made.  This assumption will be reviewed further prior to issue of the Mineral Reserves estimate.



 

 

 

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PIT WALL SLOPES


A series of geotechnical core holes were drilled during the mid 1990s.  Geotechnical analysis of the data at that time suggested an average pit wall slope of 45°.  This slope was adopted for current open cut mine planning.

MINING METHODS


Four mining methods are proposed for the ore cuts:

·

Conventional mining by 110 tonne hydraulic excavator and 50 tonne trucks will be used to open the initial voids in most of the cuts.  Waste from this opening will be placed in dumps outside the pit, although in some cases it can placed adjacent to the pit to be pushed back into the pit void at a later date.  In the case of Pit 4 in the Texcoco area, there is an opportunity for the waste to be hauled a short distance for use as bulk fill on the downstream side of the tailings dam wall.

·

Once a void of 100 m to 200 m width is formed, bulldozers will be used to push overburden from above the manto into the mined out areas.  Experience in many coal mines across the world has shown dozer pushing to be a cheap and effective stripping method.  The Boléo overburden is well suited to dozer pushing because is relatively weak and does not contain large boulders.  In most cases the push directions will be planned to keep the waste within the pit limits thus minimizing the total area of disturbed land.  In some areas, such as Pits 3 and 4 at Texcoco and Pit 2 in Boléo Arroyo, Manto 3 outcrops high on the valley wall creating excellent conditions for dozer push to form out of pit waste dumps off the valley side.  Any such areas will require careful planning to minimize the disturbed area and ensure long term stability of the pushed out of pit dumps.

·

Overburden stripping will be stopped 5 m above the estimated position of the target mantos.  Grade control drilling will then be undertaken to define the target mining surfaces and ore grades.  The final 5 m will then be mined by the excavator and trucks.

·

Ore zones from each manto will be mined by the 110 tonne excavator and the 50 tonne trucks will haul the ore over purpose built haul roads to the run of mine (ROM) stockpile area at the plant site at the mouth of the Soledad arroyo.


Light blasting will be used in the manto overburden to promote productivity of the excavator and dozer fleets.

PIT SELECTION


Whittle and Minex pit optimization software tools were applied to the Boléo deposit to guide pit selection.  Both systems showed that very large pits can be economically mined.  However, when the cost of overburden removal is included, open cut ore costs 3 to 4 times more to deliver to the ROM stockpile than the estimated cost of underground mine production.  

The best copper grades are in the deeper sections of the mantos so the project is committed to underground mining.  Given these facts, the open cuts were defined by selecting the sections of the optimized pit shells which are not well suited to underground mining because they are too



 

 

 

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shallow, too steep or too faulted or in locations where the open cut wall provides a good entry point for underground mining.

The defined pits contain just over 10 Mt of ore grade manto at an average ratio of 12 bcm waste per tonne of ore.  The current mine plan averages 590,000 tonnes of potential ore per year, although this schedule requires further smoothing and matching against the underground mining schedule.

17.1.3

LIMESTONE QUARRY


The ore treatment process requires an average of 718,000 tonnes of limestone grading 65% CaCO3 per year which will be sourced from a section of the Coquina fossiliferous limestone bed in the north western corner of the mining area.  The Coquina bed is part of the Gloria formation which occurs high in the Boléo sequence and can be seen outcropping along many ridge lines.  This section of the Coquina bed averages just over 64% CaCO3.

Conventional mining with a 180 tonne hydraulic excavator matched with 90 tonne trucks is planned for both overburden and limestone.  All material is blasted.  The trucks haul waste to an out of pit dump in the valley immediately east of the quarry and limestone to a stockpile at the southern end of the quarry.

The 6.5 km haul from the quarry to the plant site is too long for the 90 tonne off-highway dump trucks so a wheeled loader will rehandle limestone from the quarry stockpile into purpose built 50 tonne road haulers for transport to the plant site.  

Haul roads have been designed to take advantage of the natural topography wherever possible and their costing is included as part of the surface infrastructure capital cost estimate. In the case of an occasional heavy rainfall event brought about by cyclonic activity the roads will be repaired and re-built on an as needed basis.  In addition, there are allowances in the capital cost estimate for pit dewatering equipment.   

Part of the limestone deposit overlies potential ore grade Manto 3.  This area was considered too deep to mine initially but after the southern portion of the quarry is removed, the reduced overburden depth makes it possible to extend copper Pit 1 further north.

17.1.4

OPEN CUT DEVELOPMENT SEQUENCE


Excavator mining of copper Pits 3, 4 and 5 starts in Year 1 to provide early highwall access to underground mining blocks on the down throw side of the San Antonio fault south west of Pits 6 and 7 and in the Texcoco area.

The excavator moves into the higher sections of Pit 2 in Year 2 and the dozers commence stripping in Pits 3 and 4 before moving into Pit 2 in Year 4.  Pit 1 is commenced in Year 7 when the limestone quarry has mined out the sections overlying the copper pit.  Pit 8 in Purgatorio Arroyo is commenced Year 11 once underground mining is finished in that area.  Finally, Pits 6 and 7 are commenced in Year 14 and 15.  Even though they are closest to the plant site their average grade is lower than the other pits.



 

 

 

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Although the schedule will require smoothing it achieves its basic aims of supplementing high grade underground feed over the first five to eight years then ramping up the production rate to make up for decreasing underground tonnes and grades in the latter years of the mine life.

17.1.5

OPEN CUT MINING COST


Mine operating costs were estimated from first principles by sizing the mining fleet, workforce and consumables (mainly explosives) to match the production schedule.

Operating hours were estimated for each machine each year and hourly operating costs were applied for overhauls, parts, fuel, lube, tyres and wear parts.

Lease costs were estimated for all mining equipment assuming a 10% lease rate on rolling 5 year terms.  This was done in preference to estimating an equipment capital schedule because it spreads the capital costs over time and it makes the overall mining costs closer to a contract mining cost which is one of the options under consideration.

Explosives consumption was estimated against the mining volumes and estimated powder factors.  Costs per tonne were applied against ANFO and emulsion usage and a percentage allowance was made for blasting accessories such as detonators, primers and detonating cord.

Total annual costs were estimated for each employee classification and applied against the workforce schedule.

Total grade control drilling metres and numbers of samples were estimated against the ore mining schedule.

A contingency was not applied to the total estimated cost to cover items, but this will be examined in further detail as the final feasibility study is prepared during the first quarter of 2007.

Total open cut mining costs averaged:

·

Copper pits ore and waste combined

US$2.15/bcm or US$1.14/t (wet)

·

Limestone quarry ore and waste combined

US$3.55/bcm or US$1.64/t

·

Copper pits, total cost ore to ROM

US$19.79/t (dry)

·

Quarry, total cost limestone to plant

US$10.65/t.


The open pit locations are shown below in Figure 65.



 

 

 

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Figure 65:

Open Cut Pit Locations

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17.2

UNDERGROUND MINGING


Underground mining operations at Boléo will consist of room-and-pillar mining with pillar extraction.  The initial mining approach is to target and sequence mining extraction to take advantage of the higher copper (Cu) grade “bull’s eyes” that exist in each of the mantos during the early production years.  The project life is a minimum of 20 years.  Preliminary mine planning has been based on a geologic model provided by Dr W Yeo of Hellman & Schofield Pty Ltd.

Underground mining will take place in Mantos 1, 2, and 3 where (1) there are reasonably sized, contiguous mineralized zones expected to be classified as either measured, indicated or inferred in the current resource model, (2) the mantos have a Cu content greater than 0.2% at a minimum mining height of 1.8 m, and (3) the mantos have an overburden thickness of at least 20 m.  Approximately 70% of the underground mining is in Manto 3, 16% in Manto 2, and 14% in Manto 1.  

Of the approximately 48.8 million underground dry tonnes initially modelled for the Preliminary Economic Assessment (PEA), none are measured, approximately 38.5 Mt are indicated, and approximately 10.3 Mt are inferred.  An additional 10 Mt of manto ore are required to be mined during the last 4 years of the initial mine plan and were not modelled as of this publication date.   

The underground mining resource was divided into independent mines, each of which can be accessed from logical portal locations (manto outcrops, box cuts and/or shallow slopes).  To the extent practical, the mines segregate the resource by copper grade.  The primary objectives for sequencing the mines were winning the higher copper grade ore first, while optimizing the initial plant feed grade at 2.3% Cu and then maintaining a higher feed grade for as many years as practical.  The copper grade gradually decreases by year for the life of the project.  Production and mined ore grades for the first 16 years were modelled using SurvCADD™ software, with production and grades for the remaining 4 years of the project life estimated by percent aerial recovery.

Underground mine production starts with one mining unit approximately 9 months prior to plant commissioning.  Production is systematically ramped up over the next 21 months to five mining units operating with three production crews each and a targeted annual production rate of approximately 2.6 million dry tonnes per year, coinciding with the start of the second year of plant operation.  Underground production is maintained at 2.6 Mt through Year 5, when a sixth underground mining unit is added to compensate for declining ore grade, increasing raw ore production by 20% to approximately 3.1 Mdmt/a.  

Beginning in Year 15, underground production is gradually decreased to approximately 1.5 Mmt/a and maintained at that rate through Year 19.  In Year 20, the production is reduced to approximately 0.75 Mt.  The reduction in underground production between years 15 through 20 is expected to be offset by equivalent production from open-cut mining.



 

 

 

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UNDERGROUND MINING METHOD


As a consequence of the Boléo mantos depositional and geotechnical similarities with coal seam deposits (bedded and relatively soft ground), the recommended underground mining method is similar to those successfully used for coal, trona, potash, and salt seam room-and-pillar mining with pillar removal in North America, Australia, and South Africa.

Underground mining will utilize (1) a combination of hydraulic excavators, continuous miners, rubber-tired batch haulage, and mobile roof-bolting equipment for access drift and pillaring crosscut development and (2) remote-controlled continuous miners, continuous haulage, and mobile roof supports (MRS) for pillar extraction.  Development mining and pillar extraction may be independent activities with different equipment and crews; however, mine scheduling and synchronizing mine production with plant feed requirements dictates some units conducting both development and pillar mining operations.  For both development and pillar removal, ore haulage from the producing units to the portal stockpile will be by belt conveyor.

Ore from the underground mines will be transported to the plant either by haulage truck from the surface stockpiles at the mine portal or via an overland conveying system. Both systems are expected to work on a 24/7 basis.

Depending on mine size, three or four parallel drifts (access mains), connected at regular intervals by crosscuts, will be excavated from the mine portal through the manto to access mining districts.  All mine portals will be established above the typical food levels and where this is not possible flood protection measures will be taken to minimize the possibility of flooding the workings.  Emergency drainage equipment is included in the capital cost estimate for the underground mine.

Where mine conditions and production scheduling allow, development work for an entire district will be completed prior to initiating pillar removal.  Multiple, independent districts will be simultaneously mined in various stages of development and pillar removal.  Equipment will be moved between districts as needed.  A conceptual mine plan for a mining district is shown in Figure 66 and a larger area map is shown in Figure 67.

UNDERGROUND RESOURCE RECOVERY


Underground resource recovery will be impacted by extensive faulting, steep manto dip, mining method/equipment limitations, and historical mine workings.  Except where access drifts cross faults, a 10 m buffer is left unmined on both sides of faults.  Areas where the manto dip exceeds 25% are not mined.  To account for mining method/equipment limitations and historical mine workings, it is assumed that only 95% of the development excavations and 75% of pillared areas will be recovered.  These assumptions will be reviewed prior to issuing the mineral reserves estimate.

UNDERGROUND RESOURCE GRADE DILUTION


The underground resource model for each manto initially calculates ore grade for a 1.8 m interval.  Where total manto thickness is less than 1.8 m, the grade is diluted to account for a



 

 

 

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minimum 1.8 m mining height.  Where the manto thickness is greater than 1.8 m, the grade of each adjacent 0.6 m intervals are sequentially checked.  If the immediately adjacent 0.6 m interval exceeds 1% Cu, 0.6 m is added to the resource height and average grade is adjusted to include that additional increment, and only then is the next 0.6 m interval evaluated for inclusion in the resource model.

The mining equipment selected for Boléo can adapt to infinitely variable mining heights between 1.8 and 4.2 m.  Within those mining limitations, the excavation height will approximate the grade interval, not the 0.6 m increments of the resource model.  Consequently, the resource model already accounts for reasonable grade dilution and no additional dilution adjustments are appropriate.  This mechanism for addressing grade dilution during the mining operation will be reviewed prior to issuing the final mining plan.

UNDERGROUND GEOTECHNICAL MINE DESIGN


All underground pillar and excavated opening dimensions used for mine layout and design are based on (1) observations made and documented6 during test mining at Boleo’s Texcoco mine during the last quarter of 2005 and 2006, and (2) a geotechnical study conducted by AAI.7

UNDERGROUND PRODUCTIVITY


Productivity at the underground mines is based on observations made during test mining at the Texcoco Mine including a roof-bolting productivity test conducted in November of 2006 and AAI’s extensive experience with similar production equipment in coal, potash, and trona mines.  It is assumed that developing mining units will produce 490 dt per unit shift and pillaring units will produce 2,536 dt per unit shift.

UNDERGROUND MINING COSTS


Underground mine costs were based on the following:

·

Capital/leasing costs were estimated on an annual basis for the equipment and infrastructure required to produce the required ore tonnage and grade from the proposed underground mine sequencing plan.

·

Labour costs were estimated for each employee classification and applied against a detailed workforce schedule developed to produce the required ore tonnage and grade from the proposed underground mine sequencing plan.

·

Maintenance cost per dry tonne of ore produced was estimated by averaging actual maintenance costs from five US underground coal mines utilizing similar equipment as proposed for Boléo.

·

Supply cost per dry tonne of ore produced was estimated by calculating the anticipated installed roof support cost per ton of ore produced and adding an estimated cost per ton of


 

6Agapito Associates, Inc. (2006), Preliminary Geotechnical Performance Study for Underground Mining of El Boléo Copper Cobalt Project, Texcoco Test Mine Including Operations Observations and Recommendations, draft report to Baja Mining Corp, July.

7Agapito Associates, Inc. (2007), “Geotechnical Evaluation for Underground Mine Design,” draft report to Baja Mining Corp, February



 

 

 

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other supplies based on experience from US underground coal mines utilizing similar mining methods as proposed for Boléo.


Figure 66:

Conceptual Mine Layout for a District

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Figure 67:

Mining Area 3-10 Preliminary Layout

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18

PRELIMINARY ECONOMIC ASSESSMENT

 

 


18.1.1

SUMMARY


This Preliminary Economic Assessment (“PEA”) of El Boléo project is based upon:

·

The mineral resource estimate for copper, cobalt and zinc prepared by Hellman and Schofield discussed in Section 15

·

The process flowsheet developed by Bateman and recoveries of copper, cobalt and zinc achieved during the Fully Integrated Pilot Plant testing program at SGS Lakefield Research Ltd., Lakefield, Ontario, conducted under the joint guidance of Baja Mining Corp. & Bateman, as discussed in Section 14

·

The open cut mine design developed by Australian Mine Development and Design, as discussed in Section 16.1

·

The underground mine design developed by Agapito Associates, Inc., as discussed in Section 16.2.


Reference is made to sections 14, 15 and 16.1 and 16.2 for a detailed discussion of the Process Flowsheet, Mineral Resource Estimate and Mine Design.

US$ and Metric measures are used throughout this section.

The current base-case is for annual mine production to deliver 3.5 Mwt (2.6 Mdt) per year to the process facility; with maximum annual metal production of 50,000 tonnes of copper, 2,600 tonnes of cobalt and 30,000 tonnes of zinc sulphate monohydrate.  

Capital cost of the construction of the mine and process plant complex is currently estimated at US$540 million (including Owner’s Costs and Contingencies).  Total Operating costs are estimated at US$31.62/dt of ore treated, averaged over the 20 year project life being modelled.  During the initial 20 years 58.5 Mt of ore will be processed.  The Total Measured and Indicated Resource is 232.8 dmt, potentially leaving a large amount of ore to be processed after the initial 20 years.  

Preliminary mine scheduling was based upon the scheduling of resource blocks which include Inferred material, using copper equivalent grades and process recoveries.  A copper equivalent grade was determined using the following base-case metal prices:

·

Copper – US$0.95/lb

·

Cobalt – US$12.00/lb

·

Zinc – US$0.45/lb

·

Copper recovery – 91.2%

·

Cobalt recovery – 78.2%



 

 

 

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Base Case financial modelling of the project was done using conservative long term metal prices. The assumed Base Case long term copper price is $1.25, starting in 2013, with step down pricing between current levels and long term according to the LME 5 year forward price curve published as of January 25, 2007.  Those prices are  $2.20 in 2009, $1.95 in 2010, $1.75 in 2011, and $1.50 in 2012.  

Expressions of interest have been received from potential offtake partners indicating that net-back price of copper would include a slight premium (assumed to be $0.02/lb) above the LME, or COMEX, price after adjusting for freight. The Base Case price for cobalt is $12.00/lb flat over the life of the project.  The Base Case price for zinc sulphate is $950/mt flat over the life of the project.  Provision is made for delivery of the products to the end markets, including packaging, freight, and freight insurance.  

The modelling, based on the current un-optimised preliminary mine schedule, indicates that the project is financially attractive at base-case metal prices. Financial modelling, using the base case prices and only 20 years for the project life, shows that the project could generate a net after- tax Internal Rate of Return (IRR) of 19.0% with a discounted present value, at an 8% discount rate, of US$333.4 million.  

Using a 6% discount rate generates an NPV, after tax, of $US445.9 million.

Cash flow analysis was also conducted using the following metal price assumptions to test the project's economic robustness and sensitivity to changes in those prices, as shown in Table 51:

·

The SEC approved 5 year average prices, comprised of the weighted average of the 3-year trailing price and 2 year leading price.  This calculates to be $2.20/lb for copper and $16.00/lb for cobalt.

·

The current prices which, as of the end of January 2007 are $2.50/lb of copper and $26.00/lb of cobalt and $1,500/ts of zinc sulphate monohydrate.

·

An “Opportunity Case” is also shown to demonstrate the potential impact of adding the recovery of Manganese (as Manganese Carbonate) to the basic project.


Table 52:

Preliminary Economic Assessments – Base-Case Highlights

Description

Quantity or Value

Preliminary Mine Production Schedule

2,600,000 dmt/a, increasing to 3,100,000 dmt/a in Year 6

Metal Production

Up to 60,000 t/a Cu cathode

Up to 3,100 t/a Co cathode

Up to 36,000 t/a ZnSO4 salt

Capital Cost

US$540 million

19 year average Operating Cost, excluding start-up year

US$31.62/t of ore

Long term metal prices

Copper – US$1.25/lb

Cobalt  – US$12.00/lb

ZSM – US$950/t

(After tax) Internal rate of return (IRR)

19.0 %



 

 

 

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The project is sensitive to 4 key variables:  Copper price; cobalt price; the capital cost and to operating costs.  The sensitivity of the After-Tax IRR and NPV (at 8% Discount Rate) relative to the Base Case is shown in the table below to indicate the effect of plus or minus 10% changes in the key variables.  Note that the changes to the Copper price apply to all of the annual prices, starting in 2009, and not just the long-term price.

Table 53:

Sensitivity to Key Variables

 

After Tax IRR

 

After Tax NPV at 8% ($Millions)

 

-10%

Base Case

+10%

-10%

Base Case

+10%

Copper Price

16.0%

19.0%

21.7%

$236

$333

$420

Cobalt Price

18.1%

19.0%

19.9%

$298

$333

$367

Capital Cost

21.5%

19.0%

16.6%

$375

$333

$283

Operating Cost

20.7%

19.0%

17.1%

$396

$333

$266



18.1.2

PRODUCT MARKETING


The processing facility at El Boléo property will produce London Metal Exchange (LME) grade copper cathode (metal), and high purity cobalt metal. It is not proposed to produce zinc metal but rather to evaporate the zinc sulphate stripped in the zinc solvent extraction circuit to produce zinc sulphate monohydrate for subsequent sale into the soil micro-nutrient market, animal feed market, or delivery to a zinc refinery.  

Off-take agreements with respect to the above products have not yet been negotiated, but “Expressions of Interest” have been received from potential offtake candidates.  

The copper cathode produced on site is expected to exceed LME purity specifications for sale and may command a premium on LME pricing. Indicative terms from offtake parties indicate that a premium of $0.02/lb above LME could be expected and this assumption has been built into the pricing assumptions. If an off-take agreement is not negotiated for the copper it could be offered for sale on the London or Comex metal markets.

The current process flowsheet provides for the production of cobalt metal on-site.  For purposes of this study it is presumed cobalt metal will be marketed through an off-take agreement that will be negotiated prior to the commencement of production.  Costs for packaging (drums), freight insurance, and freight to market will be deducted from the indicated selling prices.

The zinc sulphate monohydrate is considered to be a “value added” product and should command a price equivalent to the value of contained zinc metal, or higher.  It is assumed that the material would be sold through an agent and that the capital cost of the production facility would be paid for by the agent’s company and recovered through a processing fee.  The net price received by Baja is also adjusted for costs of packaging (jumbo bags), freight to market, and the agent’s selling commission.



 

 

 

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18.1.3

CAPITAL COST ESTIMATE


The capital cost for development of the Boléo Project has been developed by a number of specialist organizations.  These organizations are listed below in a table that summarises major areas of significant capital cost and the organizations responsible for development of capital costs for these respective areas.  The Capital Cost Estimate for the project development has been co-ordinated and integrated by Wardrop Engineering on behalf of Bateman Engineering Canada Corp.  

Table 54:

Capital Cost Areas of Responsibility

Major Cost Area

Consultant

Location

Open Pit Mining

AMDAD

Sydney, Australia

Underground Mining

Agapito Associates, Inc

Golden, Colorado

Mining Surface Infrastructure

Wardrop Engineering

Vancouver, Canada

Process Plant and General Infrastructure

Wardrop Engineering

Vancouver, Canada

Tailings Dam

Arcadis Geotechnica

Santiago, Chile

Co-Generation Plant

Fransen Engineering Ltd

Vancouver, Canada

Acid Plant

Fenco Pty Ltd

Toronto, Canada

SO2 Gas Production Facility

Noram Engineering & Constructors Ltd

Vancouver, Canada

Barging Facility

ATI

Vancouver, Canada

Liquid Sulphur Infrastructure

ICEC Canada Ltd

Calgary, Canada

Mexican Construction Labour Rates

UHDE Jacobs

Mexico City


The overall capital cost is estimated at US$540 million.  

This number includes the capital cost components of all infrastructure, the process plant, and the mining operation, including the mining fleet, mining infrastructure, tailings dam, haul road construction, electrical power and water reticulation to the various mine sites, waste disposal, construction camp and various community initiatives.  

Note that although these capital cost estimate numbers are still under review by both Bateman Engineering Canada Corp and Baja Mining Corp. and have not yet been signed off internally by either organisation it is believed that there are no significant omissions in the analysis.

Table 55 provides a summary breakdown of the estimated total project capital costs.  The base date of the cost estimate is July 31, 2006.  All costs are listed in US$.  The numbers are rounded to reflect that they are approximations.

Table 55:

Capital Cost Estimate Summary

Project Area

Estimate Capital Cost
US$

Mining & Tailings

45,563,000

Process Plant

181,950,000

Services & Infrastructure

116,493,000

Buildings (incl. Construction Camp)

15,316,000

Construction Indirects & Freight

38,171,000

Sub-total – Direct Field Costs

397,493,000



 

 

 

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Project Area

Estimate Capital Cost
US$

EPCM

45,433,000

Owners Costs

35,681,000

Contingency

62,041,000

Total – Project Capital Cost

540,648,000


The Capital Costs set out in the above table exclude:

·

mine rehabilitation costs

·

mine closure and environmental costs

·

working capital

·

capital spares

·

first fills of reagents

·

sustaining capital.


Provision for the above costs was made in the economic model as individual line items.  Attempts will be made to minimize Working Capital requirements through the involvement of off-take agreements offering prompt payment for metals following transfer of title as the products leave the plant gate.

The project evaluation has been presented for a 20 year life, but it is expected that the life of the project will be considerably longer.  There is no credit given for recapture of the salvage value of the equipment after the initial 20 years of operation.  It is assumed that any costs for closure and reclamation of the process plant would be adequately covered by the salvage value of the plant.  Reclamation of the open cut mines is expensed as it is incurred during the initial 20 year life on an ongoing basis, so that there will not be a significant closure and reclamation cost associated with mining.

18.1.4

OPERATING COSTS


Operating costs were developed by Bateman and Baja from the following sources:

·

Extensive bench scale metallurgical and pilot plant test work data

·

Quoted budget prices for reagents and consumables, typically from North American suppliers

·

Appropriate labour costs for Expatriates and Mexican nationals for project development in the Baja California area of Mexico, drawing on remuneration experience from the local Gypsum Project, adjacent to the Boléo Resource on the Baja Peninsula and other Mexican Operations

·

Maintenance costs based on other plant operations of a similar nature

·

Estimates of open cut mining costs from AMDAD Pty Ltd



 

 

 

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·

Estimates of underground mining costs from Agapito Associates, Inc based on the Boléo test mine experience as well their operating costing data base of similar operations in North America in coal, potash and trona

·

A factored approach to product marketing and product freight costs based on discussions with freight forwarders, shipping agents and interested off take parties.

Note that although these operating cost estimate numbers are still under review by both Bateman Engineering Canada Corp and Baja Mining Corp. and have not yet been signed off internally by either organisation it is believed that there are no significant omissions in the analysis.


18.1.5

PRELIMINARY ECONOMIC ASSESSMENT


SUMMARY


The financial model utilizes the current preliminary mine production scheduled over a 20 year mine life, the associated diluted metal grades based on the H&S geological resource and AMDAD  and Agapito mine plans, the capital and operating costs as submitted by Bateman Engineering, and projected metal prices.  The Economic Analysis of the Boléo project has been done on an all equity basis (with no financial leveraging) using the Discounted Cash Flows, after taxes, for the construction period and first 20 years of the project life.  The project is not limited to 20 years by the orebody, and it can be expected that the project will continue beyond that length of time, adding to the 20 year value of the project.  The projected cash flows allow for all capital expenditures, including construction, working capital and sustaining capital.  Since this is an unleveraged model no financing costs or interest were accounted for in the Project’s costs.

This Preliminary Economic Assessment includes the use of some inferred resources to enable a 20 year project life to be modelled.  Approximately 81% of the ore mined in this model is quantified as Measured or Indicated.  Some Inferred resources were scheduled, primarily in the last 5 years of the life of the project to complete the desired 20 year schedule but conservative estimates have been used for the grades mined in those years.  It is expected that the incorporation of the results from the 18,000 m of drilling that was completed between October and December, 2006 will upgrade these resources into the Indicated or Measured categories, enabling a full 20 year project to be modelled.  

On an all-equity basis the PEA indicates that El Boléo project has a payback period for invested capital within 42 months of start-up of commercial production of copper (on an after-tax basis).  The potential net present value of the project utilizing base-case prices at a 6% discount rate is US$445.9 million, or $333.3 million at 8%, and the internal rate of return on investment is 19.0% on an after tax basis.

Using the assumed “Base Case” metal prices the annual revenue from cobalt and zinc sulphate sales is potentially adequate to cover almost all of the operating costs.  This will result in a cash cost of copper production, net of by-product credits just slightly above zero cents per pound of LME grade copper production.  Using higher by-product prices, but still below current levels, will drop the net cash cost of production of copper below zero.  The addition of a Manganese product could also drop the net cash cost of copper to less than zero.



 

 

 

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MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT




BOLÉO PROJECT ECONOMIC ANALYSIS ASSUMPTIONS AND DISCUSSION


Capital Cost and Expenditure Timing


The total project capital costs for the construction of the mine and process plant are detailed in Sections 17.1.3 and 17.1.4 of this report.  The timing of expenditures is based on an accelerated project schedule that enables the start-up of the copper production circuit in 2 years, with construction continuing after that time on the DSX circuit for the recovery of cobalt and zinc.  At this time there is no allowance for the capital cost for the construction of the manganese carbonate circuit, and it remains an opportunity for future production.  

It is expected that there will be an offset of approximately 8 months between the start-up of copper production, and that of cobalt and zinc.  This philosophy was adopted to allow the operating team to focus on the start-up of the main revenue generator (copper) without the distraction of simultaneously starting up the more complex DSX circuit.   This should result in a faster ramp up rate for copper which will more that offset the loss of revenue from the other two products, and make the most efficient use of the technical resources.  This schedule will also enable more efficient use of a smaller construction crew and minimize the size of construction camp required, along with potential impact on the infrastructure of the neighbouring town of Santa Rosalia and its residents.  

Although the projected construction schedule indicates that the DSX circuit can potentially be built in 11 months, the economic model has allowed for a full year of stagger between the two start-ups.  

A summary of the Boléo Project development timing is shown in Table 56.  The timing of these activities forms the basis of timing of the cash flows in the economic model.

Table 56:

Summary of Boléo Project Development Timing

Year

Activity

Percentage Expenditure

-2

Mine pre-development

Site preparation for process plant

Commencement of construction of acid plant

10%

-1

Construction of Site and Process Plant

Underground development

Construction of permanent wharf

70%

1

Start-up of copper production

Construction of DSX circuit, cobalt EW, ZSM granulation

20%

2

Start-up of cobalt and zinc production

 


Sustaining Capital


In the economic analysis provision is made each year for sustaining capital.  This allowance covers cost associated with capitalized rebuilds, refurbishment rebuilds, and replacement of equipment and major spares to maintain the operation at the design capacity.



 

 

 

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UPDATED PRELIMINARY ECONOMIC ASSESSMENT




The provision for sustaining capital includes:

·

General allowance based on 3% of the direct capital costs of the process plant and infrastructure

·

Mine mobile equipment  rebuilds and replacements from Year 6 onwards

·

Development of underground mine portals

·

Lifts of the tailings dam in years 4 and 9.


Working Capital


In the cash flow analysis an estimate of annual working capital requirements is made, and includes:

·

The net balance of accounts payable and receivable

·

Inventories, including first fills of reagents and capital spares

·

In process inventories

·

Finished product inventories up to the point of transfer to the purchaser.  This is expected to be a warehouse at the port of Guaymas as has been discussed with interested offtake parties.


Revenue


·

Net Sales Revenue from the sale of copper is calculated on the basis of the LME (or COMEX) price, plus a premium, and minus the total cost of delivery to the buyer including freight and insurance.  It is assumed that copper will be delivered FOB to a warehouse at the port of Guaymas.  The premium received depends on quality and market conditions, which may vary.  It is assumed that the quality will be LME Grade A.  An average premium of $0.02/lb is assumed.  A long term LME price of $1.25/lb is assumed for copper.  As the project will be starting up in 2009 the published forward pricing curve is used for the first four years of production (2009 to 2012), and the long term price is used thereafter.

·

A long term price of $12.00/lb is assumed for cobalt, relative to the current level of $26, or the 5 year average of $16.  Deductions are made for packaging (drums) and freight.  The quality of cobalt cathode produced during the pilot plant campaign was very high and would indicate that a premium may be available, but this is not assumed in this analysis.

·

Zinc Sulphate Monohydrate is assumed to be sold into the US fertilizer and animal feed markets, using agents.  Provision is made for freight into the core area of use and for a selling commission to the agents.  A long term selling price of $950/t of ZSM is assumed, before the deductions for freight and commissions.


Depreciation and Taxation


·

Mexican tax laws provide for 2 methods of calculation of depreciation.  Assets can either be depreciated at a rate of 12% for 8 years, with the remainder taken in the ninth year or at the rate of 87% in the first year of operation with the remaining 13% taken at the end of the project life.  For the purposes of evaluation of this project, under the all equity funding assumption, it has been determined that it is tax effective to claim the 87% depreciation in



 

 

 

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the first year, and the remainder in the final year.  This assumption could change when the project financing is put into place.

·

Mexican federal income tax rates are currently 28%. There is no state income tax.  Tax rates have been decreasing at a rate of 1% per year and it has been stated that the intent is to continue decreasing the rate until it reaches 25%.  However, the legislation enabling this has not been passed by the current congress and until it is in place the project has assumed that the rate remains at 28%.

·

There is a Value Added Tax in Mexico and the applicable rate is 10% for Baja California.  The tax is paid on purchases and recovered on sales, so that the full amount is normally recoverable by mining projects, which has been assumed for this project.


Manganese Opportunity Case


·

The Boléo orebody contains a substantial amount of Manganese (Mn) which could easily be recovered, with a minor capital cost increment.  The technology was tested and proven during the fully integrated pilot plant campaign conducted at the SGS Lakefield  facility in June, 2006.  

·

There is a substantial operating cost increase, however, due to the consumption of soda ash to produce the manganese as a carbonate.  

·

The quality and purity of the manganese carbonate is high, however market studies have indicated that there is a limited market for this product per se, and it would likely be used as an intermediate product for the production of higher value forms of manganese compounds.  

·

Manganese sulphate could easily be produced at site, or at other locations closer to the markets, for use in fertilizer and animal feeds.  This would be synergistic with the planned production of zinc sulphate.  

·

This market could absorb approximately one third of the anticipated production of manganese from Boléo.  

·

Another potential end product for the conversion of manganese carbonate is electrolytic manganese dioxide.  The current estimated price of manganese carbonate is $500/t ($1,100/t of contained Mn) and that of manganese sulphate is $550/t ($1,212/t of contained Mn).  

·

The addition of a large quantity of manganese carbonate into the market could however negatively affect pricing.  An opportunity case is included in the scenarios below, which assumes a selling price of $400/t for manganese carbonate.


Tables 57 to 64 contain outputs from detailed calculations of the project economics at various metal prices.



 

 

 

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MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT




Table 57:

Boléo Preliminary Economic Assessment – Base Case Summary

 

 

Pre-Start-up

Yr 1

Yr 2 - 5

Yr 5-10

Yr 10-15

Yr 15-20

Yr 1 - 20

Capital (total)

 

($432,000)

($108,000)

Average

Average

Average

Average

Average

Ore grades:

Cu

%

2.10

2.19

1.54

1.33

0.83

1.47

 

Co

%

0.08

0.12

0.09

0.09

0.07

0.09

 

Zn

%

0.62

0.74

0.63

0.58

0.33

0.57

 

Mn

%

3.14

2.31

2.50

2.88

1.86

2.43

Ore treated

 

t/a

1,170

2,633

3,120

3,120

3,120

2,925

Production:

Cu

t/a

22,408

52,682

43,934

37,901

23,560

38,006

 

Co

 

0

2,251

2,298

2,074

1,610

2,048

 

ZSM

 

0

33,607

37,075

34,151

19,415

30,928

 

MnCO3

 

0

0

0

0

0

0

Revenue:

Cu

$000/a

$109,664

$188,418

$123,002

$106,113

$65,962

$116,936

 

Co

 

$0

$59,550

$60,800

$54,862

$42,599

$51,475

 

ZSM

 

$0

$31,926

$35,221

$32,443

$18,444

$27,912

 

MnCO3

 

$0

$0

$0

$0

$0

$0

 

Total

 

$109,664

$279,894

$219,023

$193,418

$127,005

$196,323

Op. Costs:

Mining

$000/a

$26,635

$28,520

$26,687

$26,213

$25,470

$26,628

 

Process

 

$22,186

$59,949

$67,492

$67,446

$56,153

$60,872

 

G & A

 

$1,876

$2,396

$1,916

$1,926

$1,911

$1,915

 

Sales, Dist'n

 

$199

$4,422

$4,718

$4,324

$2,515

$3,784

 

Total

 

$50,896

$95,287

$100,813

$99,909

$86,049

$84,622

 

 

 

Yr 1

Yr 2 - 5

Yr 5-10

Yr 11-15

Yr 15-20

19 Yr Avg.

Unit Op Costs:

Mining

$/t ore

$22.77

$10.83

$8.55

$8.40

$8.16

$8.82

(wtd averages)

Process

 

$18.96

$22.77

$21.63

$21.62

$18.00

$20.85

 

G & A

 

$1.60

$0.72

$0.61

$0.62

$0.61

$0.63

 

Sales, Dist'n

$0.17

$1.68

$1.51

$1.39

$0.81

$1.32

 

Total

 

$43.50

$36.01

$32.31

$32.02

$27.58

$31.62

Before Tax Cash Flow

$000/a

($73,754)

$176,337

$109,717

$87,760

$41,213

$103,757

After Tax Cash Flow

$000/a

($73,754)

$156,299

$78,645

$62,986

$34,900

$83,208

Earnings

 

$000/a

($422,791)

$159,097

$79,900

$63,704

$16,233

$79,733

Cash Cost/lb Cu:

 

 

 

 

 

 

 

 

Gross

$/lb CuEq.

$1.19

$0.53

$0.59

$0.65

$0.91

$0.67

 

Net of by-product credits

$1.19

$0.06

$0.05

$0.15

$0.48

$0.15

Note:

     Long Term Prices: Copper $1.25/lb, Cobalt $12.00/lb, Zinc Sulphate $950/t

After Tax 20 year IRR:

19.04%

NPV at

0%

$1,002,100

 

6%

$445,883

 

8%

$333,377



 

 

 

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MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT



Table 58:

Boléo Preliminary Economic Assessment – 5 Year Price Case Summary

Long Term Metal Prices: Co: $2.20, Co: $16.00, ZnSO4: $950

 

 

Pre-Start-up

Yr 1

Yr 2 - 5

Yr 5 - 10

Yr 10-15

Yr 15 - 20

Yr 1 - 20

Capital (total)

 

($432,000)

($108,000)

Average

Average

Average

Average

Average

Ore grades:

Cu

%

2.10

2.19

1.54

1.33

0.83

1.47

 

Co

%

0.08

0.12

0.09

0.09

0.07

0.09

 

Zn

%

0.62

0.74

0.63

0.58

0.33

0.57

 

Mn

%

3.14

2.31

2.50

2.88

1.86

2.43

Ore treated

 

t/a

1,170

2,633

3,120

3,120

3,120

2,925

Production:

Cu

t/a

22,408

52,682

43,934

37,901

23,560

38,006

 

Co

 

0

2,251

2,298

2,074

1,610

2,048

 

ZSM

 

0

33,607

37,075

34,151

19,415

30,928

 

MnCO3

 

0

0

0

0

0

0

Revenue:

Cu

$000/a

$109,664

$257,826

$215,011

$185,488

$115,304

$185,999

 

Co

 

$0

$79,400

$81,067

$73,149

$56,798

$68,634

 

ZSM

 

$0

$31,926

$35,221

$32,443

$18,444

$27,912

 

MnCO3

 

$0

$0

$0

$0

$0

$0

 

Total

 

$109,664

$369,152

$331,298

$291,081

$190,546

$282,545

Op. Costs:

Mining

$000/a

$26,635

$28,520

$26,687

$26,213

$25,470

$26,628

 

Process

 

$22,186

$59,949

$67,492

$67,446

$56,153

$60,872

 

G&A

 

$1,876

$2,396

$1,916

$1,926

$1,911

$1,915

 

Sales, Dist'n

 

$199

$4,491

$4,810

$4,404

$2,564

$3,853

 

Total

 

$50,896

$95,357

$100,905

$99,988

$86,098

$84,665

 

 

 

Yr 1

Yr 2 - 5

Yr 5-10

Yr 11-15

Yr 15 - 20

19 Yr Avg.

Unit Op Costs

Mining

$/t ore

$22.77

$10.83

$8.55

$8.40

$8.16

$8.82

Weighted Averages

Process

 

$18.96

$22.77

$21.63

$21.62

$18.00

$20.85

G&A

 

$1.60

$0.72

$0.61

$0.62

$0.61

$0.63

 

Sales, Dist'n

$0.17

$1.71

$1.54

$1.41

$0.82

$1.34

 

Total

 

$43.50

$36.04

$32.34

$32.05

$27.60

$31.65

Before Tax Cash Flow

$000/a

($73,754)

$265,525

$221,901

$185,344

$104,704

$194,369

After Tax Cash Flow

$000/a

($73,754)

$220,515

$159,417

$133,247

$80,614

$148,448

 

 

 

 

 

 

 

 

 

Earnings

 

$000/a

($422,791)

$223,313

$160,672

$133,965

$61,947

$144,974

 

 

 

 

 

 

 

 

 

Cash Cost/lb Cu:

Gross

$/lb CuEq.

$1.19

$0.60

$0.68

$0.76

$1.07

$0.78

 

Net of by-product credits

$1.19

($0.11)

($0.16)

($0.07)

$0.21

($0.06)


After Tax 20 year IRR:

30.22%

NPV at

0%

$2,242,695

 

6%

$1,114,821

 

8%

$890,983




 

 

 

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MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT




Table 59 :

Boléo Preliminary Economic Assessment – Current Price Case Summary

Current Metal Prices Cu: $2.50, Co: $26.00, ZnSO4: $1,500

 

 

Pre-Start-up

Yr 1

Yr 2 - 5

Yr 5 - 10

Yr 10 - 15

Yr 15 - 20

Yr 1 - 20

Capital (total)

 

($432,000)

($108,000)

Average

Average

Average

Average

Average

 

 

 

 

 

 

 

 

 

Ore grades:

Cu

%

2.10

2.19

1.54

1.33

0.83

1.47

 

Co

%

0.08

0.12

0.09

0.09

0.07

0.09

 

Zn

%

0.62

0.74

0.63

0.58

0.33

0.57

 

Mn

%

3.14

2.31

2.50

2.88

1.86

2.43

 

 

 

 

 

 

 

 

 

Ore treated

 

t/a

1,170

2,633

3,120

3,120

3,120

2,925

 

 

 

 

 

 

 

 

 

Production:

Cu

t/a

22,408

52,682

43,934

37,901

23,560

38,006

 

Co

 

0

2,251

2,298

2,074

1,610

2,048

 

ZSM

 

0

33,607

37,075

34,151

19,415

30,928

 

MnCO3

 

0

0

0

0

0

0

 

 

 

 

 

 

 

 

 

Revenue:

Cu

$000/a

$124,483

$292,667

$244,066

$210,554

$130,885

$211,134

 

Co

 

$0

$129,025

$131,733

$118,868

$92,297

$111,530

 

ZSM

 

$0

$50,410

$55,612

$51,226

$29,122

$44,072

 

MnCO3

 

$0

$0

$0

$0

$0

$0

 

Total

 

$124,483

$472,102

$431,412

$380,649

$252,304

$366,736

 

 

 

 

 

 

 

 

 

Op. Costs:

Mining

$000/a

$26,635

$28,520

$26,687

$26,213

$25,470

$26,628

 

Process

 

$22,186

$59,949

$67,492

$67,446

$56,153

$60,872

 

G & A

 

$1,876

$2,396

$1,916

$1,926

$1,911

$1,915

 

Sales, Dist'n

 

$214

$4,526

$4,839

$4,429

$2,580

$3,878

 

Total

 

$50,911

$95,391

$100,934

$100,013

$86,114

$84,679

 

 

 

Yr 1

Yr 2 - 5

Yr 5-10

Yr 11-15

Yr 15-20

19 Yr  Avg.

 

 

 

 

 

 

 

 

 

Unit Op Costs:

Mining

$/t ore

$22.77

$10.83

$8.55

$8.40

$8.16

$8.82

(weighted  averages)

Process

 

$18.96

$22.77

$21.63

$21.62

$18.00

$20.85

G & A

 

$1.60

$0.72

$0.61

$0.62

$0.61

$0.63

 

Sales, Dist'n

$0.18

$1.72

$1.55

$1.42

$0.83

$1.35

 

Total

 

$43.51

$36.05

$32.35

$32.06

$27.60

$31.66

Before Tax Cash Flow

$000/a

($58,950)

$368,440

$321,985

$274,886

$166,447

$282,940

After Tax Cash Flow

$000/a

($58,950)

$293,577

$231,478

$197,717

$125,077

$211,962

Earnings

 

$000/a

($407,986)

$296,375

$232,733

$198,435

$106,410

$208,488

Cash Cost/lb Cu:

Gross

$/lb CuEq.

$1.19

$0.53

$0.59

$0.66

$0.91

$0.68

 

Net of by-product credits

$1.19

($0.70)

($0.89)

($0.84)

($0.68)

($0.79)


After Tax 20 year IRR:

40.63%

NPV at

0%

$3,454,721

 

6%

$1,786,587

 

8%

$1,457,416




 

 

 

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EL BOLEO PROJECT

MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT



Table 60:

Boléo PEA – Opportunity:  Base Case with Manganese Production

Long Term Metal Prices:  Cu: $1.25, Co: $12.00, MnCO3: $400, ZnSO4: $950

 

 

Pre-Start-up

Yr 1

Yr 2 - 5

Yr 5 - 10

Yr 10 -15

Yr 15 - 20

Yr 1 - 20

Capital (total):

 

($448,000)

($112,000)

Average

Average

Average

Average

Average

Ore grades:

Cu

%

2.10

2.19

1.54

1.33

0.83

1.47

 

Co

%

0.08

0.12

0.09

0.09

0.07

0.09

 

Zn

%

0.62

0.74

0.63

0.58

0.33

0.57

 

Mn

%

3.14

2.31

2.50

2.88

1.86

2.43

Ore treated

 

t/y

1,170

2,633

3,120

3,120

3,120

2,925

Production:

Cu

t/a

22,408

52,682

43,934

37,901

23,560

38,006

 

Co

 

0

2,251

2,298

2,074

1,610

2,048

 

ZSM

 

0

33,607

37,075

34,151

19,415

30,928

 

MnCO3

 

0

102,351

134,533

155,073

100,013

117,875

Revenue:

Cu

$000/a

$109,664

$188,418

$123,002

$106,113

$65,962

$116,936

 

Co

 

$0

$59,550

$60,800

$54,862

$42,599

$51,475

 

ZSM

 

$0

$31,926

$35,221

$32,443

$18,444

$27,912

 

MnCO3

 

$0

$40,940

$53,813

$62,029

$40,005

$47,150

 

Total

 

$109,664

$320,835

$272,836

$255,447

$167,010

$243,473

Op. Costs:

Mining

$000/a

$26,635

$28,520

$26,687

$26,213

$25,470

$26,628

 

Process

 

$22,186

$78,217

$91,504

$95,123

$74,003

$81,910

 

G&A

 

$1,876

$2,396

$1,916

$1,926

$1,911

$1,915

 

Sales, Dist'n

 

$199

$9,539

$11,445

$12,078

$7,516

$9,677

 

Total

 

$50,896

$118,672

$131,551

$135,340

$108,900

$103,050

 

 

 

Yr 1

Yr 2 - 5

Yr 5 - 10

Yr 11-15

Yr 15-20

19 Yr Avg.

Unit Op Costs:

Mining

$/t ore

$22.77

$10.83

$8.55

$8.40

$8.16

$8.82

(wtd averages)

Process

 

$18.96

$29.71

$29.33

$30.49

$23.72

$28.19

 

G & A

 

$1.60

$0.72

$0.61

$0.62

$0.61

$0.63

 

Sales, Dist'n

$0.17

$3.62

$3.67

$3.87

$2.41

$3.37

 

Total

 

$43.50

$44.89

$42.16

$43.38

$34.90

$41.02

Before Tax Cash Flow

$000/a

($77,914)

$193,348

$132,638

$114,123

$58,880

$124,747

After tax Cash Flow

$000/a

($77,914)

$169,643

$95,129

$81,926

$47,943

$98,660

Earnings

 

$000/a

($440,330)

$172,914

$96,451

$82,792

$28,124

$95,070

Cash Cost/lb Cu:

Gross

$/lb CuEq.

$1.19

$0.65

$0.77

$0.88

$1.14

$0.86

 

Net of by-product credits

$1.19

($0.09)

($0.19)

($0.17)

$0.15

($0.10)


After Tax 20 year IRR:

20.93%

NPV at

0%

$1,277,645

 

6%

$580,218

 

8%

$441,228




 

 

 

ZAB406-00233:EM15.02.NGI

31 January 2007

Page 176 of 188




[bajatechrep001.jpg]

EL BOLEO PROJECT

MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT



Table 61:

20 Year Detailed Cash Flow – Base Case

 

 

2007

2008

2009

2010

2011

2012

2013

2014

Throughput:

dmt/a

 

 

1,170

2,470

2,600

2,600

2,860

3,120

Grades:

%Cu

 

 

2.10

2.10

2.49

1.96

2.22

1.62

 

%Co

 

 

0.08

0.15

0.15

0.12

0.06

0.06

 

%Zn

 

 

0.62

0.81

1.00

0.71

0.42

0.84

Recoveries:

Cu

 

 

91.2%

91.2%

91.2%

91.2%

91.2%

91.2%

 

Co

 

 

0.0%

60.0%

78.2%

78.2%

78.2%

78.2%

 

Zn

 

 

0.0%

50.0%

65.6%

65.6%

65.6%

65.6%

Production:

t/a Cu

 

 

22,408

47,305

59,043

46,476

57,905

46,096

 

t/a Co

 

 

0

2,238

2,989

2,481

1,297

1,440

 

t/a ZSM

 

 

0

28,581

48,731

34,599

22,514

49,121

Prices:

Cu

 

 

$2.20

$1.95

$1.75

$1.50

$1.25

$1.25

 

Co

 

 

$12.00

$12.00

$12.00

$12.00

$12.00

$12.00

 

ZSM

 

 

$950

$950

$950

$950

$950

$950

Revenue:

Cu

 

 

$109,664

$205,441

$230,383

$155,732

$162,117

$129,056

($000's)

Co

 

 

$0

$59,199

$79,066

$65,619

$34,316

$38,081

 

Zn

 

 

$0

$27,152

$46,295

$32,869

$21,388

$46,665

 

Total

 

 

$109,664

$291,793

$355,744

$254,221

$217,821

$213,802

Total Op Cost $000's

 

 

$58,896

$106,402

$104,388

$95,192

$88,193

$106,736

Initial Capital

$54,000

$378,000

$108,000

 

 

 

 

 

Sustaining Capital

 

 

$0

$1,350

$2,430

$3,980

$2,430

$18,915

Working Capital at YE

 

 

$8,522

$11,591

$12,471

$11,853

$11,388

$12,688

Working Capital Change

 

$8,522

$3,069

$880

-$619

-$465

$1,300

Income Taxes

 

 

$0

$0

$887

$43,559

$35,704

$25,371

Net Cash Flow

($54,000)

($378,000)

($73,754)

$173,971

$247,159

$112,109

$91,958

$61,480

 

 

2015

2016

2017

2018

2019

2020

2021

2022

Throughput:

dmt/a

3,120

3,120

3,120

3,120

3,120

3,120

3,120

3,120

Grades:

%Cu

1.53

1.64

1.60

1.33

1.69

1.33

1.29

1.31

 

%Co

0.06

0.13

0.12

0.09

0.08

0.09

0.08

0.09

 

%Zn

0.55

0.63

0.56

0.59

0.38

0.47

0.51

0.79

Recoveries:

Cu

91.2%

91.2%

91.2%

91.2%

91.2%

91.2%

91.2%

91.2%

 

Co

78.2%

78.2%

78.2%

78.2%

78.2%

78.2%

78.2%

78.2%

 

Zn

65.6%

65.6%

65.6%

65.6%

65.6%

65.6%

65.6%

65.6%

Production:

t/a Cu

43,535

46,665

45,527

37,844

48,088

37,844

36,706

37,275

 

t/a Co

1,537

3,196

3,025

2,293

1,952

2,171

1,830

2,245

 

t/a ZSM

32,163

36,841

32,748

34,502

22,222

27,485

29,824

46,197

Prices:

Cu

$1.25

$1.25

$1.25

$1.25

$1.25

$1.25

$1.25

$1.25

 

Co

$12.00

$12.00

$12.00

$12.00

$12.00

$12.00

$12.00

$12.00

 

ZSM

$950

$950

$950

$950

$950

$950

$950

$950

Revenue:

Cu

$121,886

$130,649

$127,463

$105,953

$134,633

$105,953

$102,767

$104,360

($000's)

Co

$40,662

$84,552

$80,034

$60,671

$51,635

$57,444

$48,408

$59,380

 

Zn

$30,555

$34,999

$31,110

$32,777

$21,110

$26,110

$28,332

$43,888

 

Total

$193,103

$250,200

$238,607

$199,401

$207,378

$189,507

$179,507

$207,628

Total Op Cost $000's

$98,638

$98,449

$100,324

$99,919

$95,322

$101,245

$96,418

$107,545

Initial Capital

 

 

 

 

 

 

 

 

Sustaining Capital

$2,700

$14,580

$2,700

$2,700

$6,382

$14,432

$2,700

$2,700

Working Capital at YE

$12,173

$12,098

$12,294

$12,254

$11,984

$12,432

$12,011

$12,780

Working Capital Change

-$515

-$76

$196

-$40

-$270

$448

-$421

$769

Income Taxes

$25,793

$38,939

$38,061

$27,197

$29,821

$21,198

$22,607

$27,366

Net Cash Flow

$66,487

$98,308

$97,325

$69,625

$76,124

$52,185

$58,203

$69,248



 

 

 

ZAB406-00233:EM15.02.NGI

31 January 2007

Page 177 of 188




[bajatechrep001.jpg]

EL BOLEO PROJECT

MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT




 

 

 

2023

2024

2025

2026

2027

2028

Throughput:

dmt/a

 

3,120

3,120

3,120

3,120

3,120

3,120

Grades:

%Cu

 

1.04

1.23

0.94

0.74

0.71

0.52

 

%Co

 

0.09

0.10

0.06

0.06

0.07

0.05

 

%Zn

 

0.77

0.62

0.38

0.26

0.23

0.17

 

 

 

 

 

 

 

 

 

Recoveries:

Cu

 

91.2%

91.2%

91.2%

91.2%

91.2%

91.2%

 

Co

 

78.2%

78.2%

78.2%

78.2%

78.2%

78.2%

 

Zn

 

65.6%

65.6%

65.6%

65.6%

65.6%

65.6%

 

 

 

 

 

 

 

 

 

Production:

t/a Cu

 

29,593

34,999

26,747

21,056

20,203

14,796

 

t/a Co

 

2,171

2,342

1,561

1,440

1,610

1,098

 

t/a ZSM

 

45,028

36,256

22,222

15,204

13,450

9,941

 

 

 

 

 

 

 

 

 

Prices:

Cu

 

$1.25

$1.25

$1.25

$1.25

$1.25

$1.25

 

Co

 

$12.00

$12.00

$12.00

$12.00

$12.00

$12.00

 

ZSM

 

$950

$950

$950

$950

$950

$950

 

 

 

 

 

 

 

 

 

Revenue:

Cu

 

$82,851

$97,987

$74,884

$58,952

$56,562

$41,425

($000's)

Co

 

$57,444

$61,962

$41,308

$38,081

$42,599

$29,045

 

Zn

 

$42,776

$34,443

$21,110

$14,444

$12,777

$9,444

 

Total

 

$183,071

$194,392

$137,303

$111,476

$111,938

$79,914

Total Op Cost $000's

 

 

$99,016

$91,756

$89,385

$85,937

$85,970

$77,196

Initial Capital

 

 

 

 

 

 

 

 

Sustaining Capital

 

 

$2,700

$2,700

$2,700

$2,700

$2,700

$0

Working Capital at YE

 

 

$12,085

$11,559

$11,505

$11,289

$11,307

$0

Working Capital Change

 

 

-$695

-$526

-$54

-$216

$19

-$11,307

Income Taxes

 

 

$22,878

$28,080

$12,759

$6,493

$6,613

-$22,383

Net Cash Flow

 

 

$59,173

$72,381

$32,512

$16,562

$16,636

$36,408



 

 

 

ZAB406-00233:EM15.02.NGI

31 January 2007

Page 178 of 188




[bajatechrep001.jpg]

EL BOLEO PROJECT

MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT



Table 62:

20 Year Detailed Cash Flow – 5 Year Prices Case – (3 year trailing + 2 year leading)

 

 

2007

2008

2009

2010

2011

2012

2013

2014

Throughput:

t/a

 

 

1,170

2,470

2,600

2,600

2,860

3,120

Grades:

%Cu

 

 

2.10

2.10

2.49

1.96

2.22

1.62

 

%Co

 

 

0.08

0.15

0.15

0.12

0.06

0.06

 

%Zn

 

 

0.62

0.81

1.00

0.71

0.42

0.84

Recoveries:

Cu

 

 

91.2%

91.2%

91.2%

91.2%

91.2%

91.2%

 

Co

 

 

0.0%

60.0%

78.2%

78.2%

78.2%

78.2%

 

Zn

 

 

0.0%

50.0%

65.6%

65.6%

65.6%

65.6%

Production:

t/a Cu

 

 

22,408

47,305

59,043

46,476

57,905

46,096

 

t/a Co

 

 

0

2,238

2,989

2,481

1,297

1,440

 

t/a ZSM

 

 

0

28,581

48,731

34,599

22,514

49,121

Prices:

Cu

 

 

$2.20

$2.20

$2.20

$2.20

$2.20

$2.20

 

Co

 

 

$16.00

$16.00

$16.00

$16.00

$16.00

$16.00

 

ZSM

 

 

$950

$950

$950

$950

$950

$950

Revenue:

Cu

 

 

$109,664

$231,512

$288,955

$227,451

$283,385

$225,594

($000's)

Co

 

 

$0

$78,932

$105,421

$87,492

$45,754

$50,774

 

Zn

 

 

$0

$27,152

$46,295

$32,869

$21,388

$46,665

 

Total

 

 

$109,664

$337,597

$440,671

$347,812

$350,527

$323,033

Total Op Cost $000's

 

 

$58,896

$106,428

$104,447

$95,263

$88,314

$106,833

Initial Capital

$54,000

$378,000

$108,000

 

 

 

 

 

Sustaining Capital

 

 

$0

$1,350

$2,430

$3,980

$2,430

$18,915

Working Capital at YE

 

 

$8,522

$11,591

$12,471

$11,853

$11,388

$12,688

Working Capital Change

 

 

$8,522

$3,069

$880

-$619

-$465

$1,300

Income Taxes

 

 

$0

$0

$37,468

$69,744

$72,828

$55,929

Net Cash Flow

($54,000)

($378,000)

($73,754)

$219,749

$295,446

$179,443

$187,420

$140,057

 

 

2015

2016

2017

2018

2019

2020

2021

2022

Throughput:

dmt/a

3,120

3,120

3,120

3,120

3,120

3,120

3,120

3,120

Grades:

%Cu

1.53

1.64

1.60

1.33

1.69

1.33

1.29

1.31

 

%Co

0.06

0.13

0.12

0.09

0.08

0.09

0.08

0.09

 

%Zn

0.55

0.63

0.56

0.59

0.38

0.47

0.51

0.79

Recoveries:

Cu

91.2%

91.2%

91.2%

91.2%

91.2%

91.2%

91.2%

91.2%

 

Co

78.2%

78.2%

78.2%

78.2%

78.2%

78.2%

78.2%

78.2%

 

Zn

65.6%

65.6%

65.6%

65.6%

65.6%

65.6%

65.6%

65.6%

Production:

t/a Cu

43,535

46,665

45,527

37,844

48,088

37,844

36,706

37,275

 

t/a Co

1,537

3,196

3,025

2,293

1,952

2,171

1,830

2,245

 

t/a ZSM

32,163

36,841

32,748

34,502

22,222

27,485

29,824

46,197

Prices:

Cu

$2.20

$2.20

$2.20

$2.20

$2.20

$2.20

$2.20

$2.20

 

Co

$16.00

$16.00

$16.00

$16.00

$16.00

$16.00

$16.00

$16.00

 

ZSM

$950

$950

$950

$950

$950

$950

$950

$950

Revenue:

Cu

$213,061

$228,379

$222,809

$185,210

$235,342

$185,210

$179,640

$182,425

($000's)

Co

$54,217

$112,736

$106,712

$80,895

$68,846

$76,592

$64,544

$79,173

 

Zn

$30,555

$34,999

$31,110

$32,777

$21,110

$26,110

$28,332

$43,888

 

Total

$297,832

$376,114

$360,631

$298,881

$325,299

$287,912

$272,516

$305,486

Total Op Cost $000's

$98,729

$98,547

$100,420

$99,998

$95,422

$101,325

$96,494

$107,623

Initial Capital

 

 

 

 

 

 

 

 

Sustaining Capital

$2,700

$14,580

$2,700

$2,700

$6,382

$14,432

$2,700

$2,700

Working Capital at YE

$12,173

$12,098

$12,294

$12,254

$11,984

$12,432

$12,011

$12,780

Working Capital Change

-$515

-$76

$196

-$40

-$270

$448

-$421

$769

Income Taxes

$55,091

$74,167

$72,201

$55,030

$62,811

$48,729

$48,628

$54,744

Net Cash Flow

$141,826

$188,896

$185,114

$141,194

$160,954

$122,979

$125,114

$139,650



 

 

 

ZAB406-00233:EM15.02.NGI

31 January 2007

Page 179 of 188




[bajatechrep001.jpg]

EL BOLEO PROJECT

MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT




 

 

2023

2024

2025

2026

2027

2028

 

 

Throughput:

dmt/a

3,120

3,120

3,120

3,120

3,120

3,120

 

 

Grades:

%Cu

1.04

1.23

0.94

0.74

0.71

0.52

 

 

 

%Co

0.09

0.10

0.06

0.06

0.07

0.05

 

 

 

%Zn

0.77

0.62

0.38

0.26

0.23

0.17

 

 

Recoveries:

Cu

91.2%

91.2%

91.2%

91.2%

91.2%

91.2%

 

 

 

Co

78.2%

78.2%

78.2%

78.2%

78.2%

78.2%

 

 

 

Zn

65.6%

65.6%

65.6%

65.6%

65.6%

65.6%

 

 

Production:

t/a Cu

29,593

34,999

26,747

21,056

20,203

14,796

 

 

 

t/a Co

2,171

2,342

1,561

1,440

1,610

1,098

 

 

 

t/a ZSM

45,028

36,256

22,222

15,204

13,450

9,941

 

 

Prices:

Cu

$2.20

$2.20

$2.20

$2.20

$2.20

$2.20

 

 

 

Co

$16.00

$16.00

$16.00

$16.00

$16.00

$16.00

 

 

 

ZSM

$950

$950

$950

$950

$950

$950

 

 

Revenue:

Cu

$144,826

$171,284

$130,900

$103,049

$98,871

$72,413

 

 

($000's)

Co

$76,592

$82,616

$55,077

$50,774

$56,798

$38,726

 

 

 

Zn

$42,776

$34,443

$21,110

$14,444

$12,777

$9,444

 

 

 

Total

$264,194

$288,343

$207,088

$168,267

$168,447

$120,583

 

 

Total Op Cost $000's

$99,078

$91,830

$89,441

$85,981

$86,012

$77,227

 

 

 

Initial Capital

 

 

 

 

 

 

 

 

 

Sustaining Capital

$2,700

$2,700

$2,700

$2,700

$2,700

$0

 

 

 

Working Capital at YE

$12,085

$11,559

$11,505

$11,289

$11,307

$0

 

 

 

Working Capital Change

-$695

-$526

-$54

-$216

$19

-$11,307

 

 

 

Income Taxes

$45,575

$54,366

$32,283

$22,382

$22,424

-$11,004

 

 

 

Net Cash Flow

$117,537

$139,974

$82,717

$57,420

$57,292

$65,667

 

 

 



 

 

 

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MINERA Y METALURGICA DEL BOLEO, SA DE CV

UPDATED PRELIMINARY ECONOMIC ASSESSMENT



Table 63:

20 Year Detailed Cash Flow – Current Prices Case

 

 

2007

2008

2009

2010

2011

2012

2013

2014

Throughput:

dmt/a

 

 

1,170

2,470

2,600

2,600

2,860

3,120

Grades:

%Cu

 

 

2.10

2.10

2.49

1.96

2.22

1.62

 

%Co

 

 

0.08

0.15

0.15

0.12

0.06

0.06

 

%Zn

 

 

0.62

0.81

1.00

0.71

0.42

0.84

Recoveries:

Cu

 

 

91.2%

91.2%

91.2%

91.2%

91.2%

91.2%

 

Co

 

 

0.0%

60.0%

78.2%

78.2%

78.2%

78.2%

 

Zn

 

 

0.0%

50.0%

65.6%

65.6%

65.6%

65.6%

Production:

t/a Cu

 

 

22,408

47,305

59,043

46,476

57,905

46,096

 

t/a Co

 

 

0

2,238

2,989

2,481

1,297

1,440

 

t/a ZSM

 

 

0

28,581

48,731

34,599

22,514

49,121

Prices:

Cu

 

 

$2.50

$2.50

$2.50

$2.50

$2.50

$2.50

 

Co

 

 

$26.00

$26.00

$26.00

$26.00

$26.00

$26.00

 

ZSM

 

 

$1,500

$1,500

$1,500

$1,500

$1,500

$1,500

Revenue:

Cu

 

 

$124,483

$262,798

$328,003

$258,187

$321,680

$256,080

($000's)

Co

 

 

$0

$128,265

$171,309

$142,175

$74,351

$82,508

 

Zn

 

 

$0

$42,872

$73,097

$51,899

$33,771

$73,682

 

Total

 

 

$124,483

$433,935

$572,410

$452,261

$429,802

$412,270

Total Op Cost $000's

 

 

$58,911

$106,460

$104,486

$95,294

$88,353

$106,863

Initial Capital

$54,000

$378,000

$108,000

 

 

 

 

 

Sustaining Capital

 

 

$0

$1,350

$2,430

$3,980

$2,430

$18,915

Working Capital at YE

 

 

$8,522

$11,591

$12,471

$11,853

$11,388

$12,688

Working Capital Change

 

$8,522

$3,069

$880

-$619

-$465

$1,300

Income Taxes

 

 

$0

$0

$105,455

$98,981

$95,014

$80,906

Net Cash Flow

($54,000)

($378,000)

($58,950)

$316,056

$359,159

$254,625

$244,470

$204,285

 

 

2015

2016

2017

2018

2019

2020

2021

2022

Throughput:

dmt/a

3,120

3,120

3,120

3,120

3,120

3,120

3,120

3,120

Grades:

%Cu

1.53

1.64

1.60

1.33

1.69

1.33

1.29

1.31

 

%Co

0.06

0.13

0.12

0.09

0.08

0.09

0.08

0.09

 

%Zn

0.55

0.63

0.56

0.59

0.38

0.47

0.51

0.79

Recoveries:

Cu

91.2%

91.2%

91.2%

91.2%

91.2%

91.2%

91.2%

91.2%

 

Co

78.2%

78.2%

78.2%

78.2%

78.2%

78.2%

78.2%

78.2%

 

Zn

65.6%

65.6%

65.6%

65.6%

65.6%

65.6%

65.6%

65.6%

Production:

t/a Cu

43,535

46,665

45,527

37,844

48,088

37,844

36,706

37,275

 

t/a Co

1,537

3,196

3,025

2,293

1,952

2,171

1,830

2,245

 

t/a ZSM

32,163

36,841

32,748

34,502

22,222

27,485

29,824

46,197

Prices:

Cu

$2.50

$2.50

$2.50

$2.50

$2.50

$2.50

$2.50

$2.50

 

Co

$26.00

$26.00

$26.00

$26.00

$26.00

$26.00

$26.00

$26.00

 

ZSM

$1,500

$1,500

$1,500

$1,500

$1,500

$1,500

$1,500

$1,500

Revenue:

Cu

$241,853

$259,241

$252,918

$210,238

$267,145

$210,238

$203,915

$207,077

($000's)

Co

$88,102

$183,196

$173,407

$131,454

$111,875

$124,461

$104,883

$128,657

 

Zn

$48,244

$55,261

$49,121

$51,753

$33,332

$41,227

$44,735

$69,296

 

Total

$378,199

$497,699

$475,446

$393,445

$412,353

$375,926

$353,534

$405,030

Total Op Cost $000's

$98,758

$98,578

$100,450

$100,023

$95,454

$101,350

$96,519

$107,647

Initial Capital

 

 

 

 

 

 

 

 

Sustaining Capital

$2,700

$14,580

$2,700

$2,700

$6,382

$14,432

$2,700

$2,700

Working Capital at YE

$12,173

$12,098

$12,294

$12,254

$11,984

$12,432

$12,011

$12,780

Working Capital Change

-$515

-$76

$196

-$40

-$270

$448

-$421

$769

Income Taxes

$77,586

$108,202

$104,341

$81,500

$87,177

$73,366

$71,307

$82,609

Net Cash Flow

$199,670

$276,414

$267,759

$209,261

$223,610

$186,331

$183,430

$211,304

 

 

2023

2024

2025

2026

2027

2028

 

Throughput:

t/a

3,120

3,120

3,120

3,120

3,120

3,120

 



 

 

 

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MINERA Y METALURGICA DEL BOLEO, SA DE CV

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19

CONCLUSIONS

 

 



19.1

GEOLOGY AND MINERAL RESOURCE MODELLING


A Measured and Indicated Resource has been defined, amounting to:

·

233 Mt at 1.83% CuEq. at a cutoff grade of 0.5% CuEq.

·

182 Mt at 2.13% CuEq. at a cutoff grade of 1.0% CuEq.


In addition an Inferred Resource has been defined, amounting to:

·

203 Mt at 1.32% CuEq. at a cutoff grade of 0.5% CuEq.

·

114 Mt at 1.76% CuEq. at a cutoff grade of 1.0% CuEq.



19.2

METALLURGY AND PROCESS DESIGN


A 'Fully Integrated' Pilot Plant Campaign was conducted at the SGS Lakefield facility from June 23 rd to July 20th, 2006 treating a bulk sample of Manto 3 oxide ore from the Boléo Property.  

Highlights of the campaign include the following:

·

Pilot plant commissioning commenced on 5th June 2006.  The integrated pilot plant campaign ran until 24th June

·

A total of nearly 5 mt of ore were treated through the pilot plant

·

The zinc solvent extraction and cobalt solvent extraction and electrowinning circuits were run during the period of 4th to 15th July in a separate campaign using the solutions collected in the earlier phase.  

·

The oxidation and reductive leaching circuit once more gave excellent extractions of copper, cobalt and zinc. Copper extraction exceeded 90% during pilot operation.  Cobalt extraction varied from 80% to as high as 90%.  Zinc extraction was generally above 70%.  These numbers are confirmation of the earlier proof of concept flowsheet test performance.   

·

The iron removal circuit consistently produced very low concentrations of key impurities in solution with negligible losses of cobalt and zinc.  

·

The CSIRO DSX ® circuit for cobalt and zinc recovery performed extremely well.  In the Lakefield pilot plant, cobalt and zinc were recovered with high overall extraction efficiency (+95%) to produce a concentrated zinc sulphate solution (for production of zinc sulphate monohydrate crystals for sale) and a concentrated cobalt solution (for production of cobalt metal cathode).

·

The successful completion of this pilot program is an important milestone in advancing the design of the hydrometallurgical facility and moving the Boléo project forward.  Once again Boléo ores have been treated in a continuous pilot plant program to leach, separate



 

 

 

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and recover pay metals in final commercial form.  Importantly the use of soda ash for pH control and the use of Boléo carbonate for neutralization duties was successfully demonstrated, as was the production of manganese carbonate and zinc sulphate monohydrate.

·

Design criteria have been confirmed and data for the purposes of ultimately formulating process guarantees for the Boléo flowsheet have been gathered.



19.3

MINING


·

The potential for surface and underground mining were assessed for the H&S resource model.  A series of underground mining operations, supported by small open-cut surface mines, delivering targeted high-grade copper-cobalt-zinc manganese ore (0.5% to 2.5% Cu) to the process facility was selected as the best alternative for the scale of operation envisaged by Baja Mining Corp.  

·

The mine modelling process has not been optimized in this preliminary economic assessment.  Additional modelling and cost estimating work is required to fully develop the Boléo property.  It is expected this effort will produce a more efficient mine plan in the future.

·

A suitable limestone source was located on the property and will be mined by surface methods to provide calcium carbonate for plant process needs.

·

A tailings dam and associated tailings disposal facility has been designed with the capacity to support the life-of-mine projections.



19.4

ENVIRONMENTAL


An environmental plan was submitted to Mexican Federal authorities and the basic Environmental Impact Manifest (EIM) permit was approved and issued to the Company in December, 2006.  This allows the company to begin submitting additional specific permitting requests for construction and operational activities in accordance with the EIM provisions.


19.5

PRELIMINARY ECONOMIC ASSESSMENT


·

Financial modelling based on the current, un-optimized preliminary mine schedule indicates that the project is attractive using conservative (base-case) metal prices while recovering copper, cobalt and zinc.

·

Modelling of these base-case metal prices over a projected 20 year mine life shows that the project could generate a cumulative after tax cash flow of US$1,434,100 million with a discounted present value of US$445.9 million at a 6% discount rate, or US$333.4 million using a 8% discount rate.

·

The cash cost of production of copper, net of by-product prices, is close to zero using conservative pricing and becomes significantly less than zero if by-product prices approach current, or recent average, price levels.

·

The addition of manganese production could add extra value to the project.



 

 

 

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20

RECOMMENDATIONS

 

 



20.1

GEOLOGY AND MINERAL RESOURCE MODELLING


H&S made the following recommendations in the 2005 PEA, all of which have been adopted.  For ease of reference these are re-quoted below.

"…The drilling data is currently kept in a series of Excel spread sheets.  These spread sheets need to be replaced by a formal database.  H&S suggest the use of Microsoft Access.  The advantages a database has over spread sheets include:

·

improved data security

·

automated data entry

·

automated data checking

·

data easily exported in correct format use in other software packages

·

data is easily interrogated

·

activity reporting

·

intercept reporting

·

automated monitoring of QC data

·

checking of QC data during data entry and automated alerts for ‘out of spec’ values.


Drill hole logging has not been carried out in a way that is amenable for use in any mining or geological software processing package.  Descriptions of geological intervals are completed as full text rather than coded to allow for computer processing.  H&S recommend that new logging sheets be drafted to facilitate recording the data in a way that allows all the pertinent geological characteristics to be recorded, whilst in a format that can be uploaded directly into a database, without manual data entry.  This typically results in a logging form with dedicated data fields that allows the geologist to record quantitative as well as qualitative observations.  This can add a lot of value to the data and can be extremely useful during resource modelling.  It is also important that future assaying and logging is completed over the same down hole intervals.  Matching of grade and geological items in this way facilitates statistical analysis of grade according the relevant geological criteria.  It would also enable accurate modelling of ore type characteristics.

Sampling practices were not observed during the present study.  It is recommended that these be reviewed by H&S at an early opportunity during the next drilling program…"

There are no further recommendations.



 

 

 

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20.2

METALLURGY AND PROCESS DESIGN


Bateman made a small number of recommendations in the 2005 PEA, most of which related to the upcoming 2006 pilot campaign, all of which were implemented at that time.  For ease of reference these are re-quoted below.

·

…"That every effort be made to ensure that the sample(s) utililized in the proposed confirmatory pilot plant campaign represent the diversity of ore type anticipated in the first 5 years of the mine life as per the preliminary mining schedule

·

That Boléo HAC be employed in all neutralization operations during the pilot campaign

·

That the pilot campaign be fully integrated – including zinc salt production, cobalt metal production and all recycle streams – to verify the overall recovery of copper, cobalt & zinc from the circuit

·

That the pilot campaign be undertaken at a sufficiently large scale that regular sampling does not unduly influence the overall process performance

·

That appropriate vendors be contracted to conduct bench scale testwork during the campaign in the areas of agitation and solid-liquid separation with the ultimate objective of obtaining process guarantees from the successful vendors for these process functions..."


Further recommendations include:

·

Getting better definition of the acid consumption variability of the ore for the first five years of operation via testwork.  This acid consumption distribution will serve to provide confidence in the ability of the processing plant to operate at design values for a given mining plan and blending strategy.

·

Getting improved definition of the proposed process flowsheets ability to successfully process small quantities of sulphide ore as part of the ore feed mix.  


Work in both these important areas has commenced in late January 2007 and results will be incorporated in the plant design as appropriate.


20.3

MINING


AMDAD and H&S made the following recommendations in the 2005 PEA, all of which have been adopted.  For ease of reference these are re-quoted below.

·

…“The trial mining operation should be kept as close as possible to anticipated operating conditions so that it can be reliably calibrated against the design and production parameters to be used in the feasibility study.

·

An infill drilling program should be conducted to bring the first two to five years of production up to at least Indicated Resource status.  A range of drill hole spacings should be planned in virgin and previously mined areas to assess the continuity of grade over the anticipated mining heights under both conditions.  This may allow a review of the overall deposit to re-assess the resource categories.



 

 

 

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·

A “mining” resource model should be constructed using economic mining height thickness calculated at each drill hole and including roof and floor data for this height.  This could be used to prepare hazard plans for roof and floor conditions covering information such as the thickness of clay breccia below the overlying sandstone or the presence of false sandstone floors above the conglomerate.  It may also assist in defining the extent of voids or stope fill.

·

Preliminary development layouts should be designed to assess the mining recovery and to define/assist definition of ore haulage and ventilation options.

·

More detailed analysis of mining methods can be deferred until the data from the mining trial is available…"


At this stage of the project development there are no further recommendations with respect to Mining.


20.4

UPDATED PRELIMINARY ECONIMIC ASSESSMENT


There are no recommendations pertinent to the economic assessment:


 

 

 

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21

REFERENCES

 

 


Bailes, R.J., Christoffersen, J.E., Escandon, F., and Peatfield, G.R. 2001.  Sediment–Hosted Deposits Of the Boléo Copper-Cobalt-Zinc District, Baja California Sur, Mexico. In Society of Economic Geologists, SP8, 2001, p. 291-306

Felix, F.C., 1996.  Specific Gravity, Nov 1996.  Unpublished Internal Company Activity Report.

Peatfield, G.R., & Smee, B. 1997, 1997.  Assay Quality Control Report for the Boléo Copper-Cobalt Project, Baja California Sur, Mexico. Jan 1997, Unpublished Internal Company Report.

Peatfield, G.R. 1997, Update to the Assay Quality Control Report for the Boléo Copper-Cobalt Project, Baja California Sur, Mexico.  Feb 1997, Unpublished Internal Company Report.

Peatfield, G.R. 1998. Analytical Quality Control at the Boléo Copper-Cobalt-Zinc Project, Baja California Sur, Mexico.  Pathways 98, Cordilleran Roundup and Exploration Methods 98, Pathways to Discovery.  Short Course #4 Analysis and Quality Control in Mineral Exploration. Jan 25-26 1998.

Sawlan, M.G. and Smith, J.G., 1984.  Petrologic Characteristics, Age and Tectonic Setting of Neogene Volcanic Rocks in Northern Baja California Sur, Mexico, In Frizzel, V.A., ed., Geology of the Baja California Peninsula:  Society of Economic Paleontologists and Mineralogists, Pacific Section (Los Angeles), San Diego, California, April 18-21, 1984, Symposium Proceedings, p. 237-251.

Schmidt, E.K., 1975.  Plate Tectonics, Volcanic Petrology and Ore Formation of the Santa Rosalia  Area, Baja California, Mexico:  Unpublished M.Sc. thesis, Tucson, University of Arizona,  191 p.

Thompson, M. and Howarth, R.J., 1978.  A New Approach to the Estimation of Analytical Precision.  Elsevier Scientific Publishing Company, Journal of Geochemical Exploration, Vol. 9, pp 23-30.

Wilson, I.F., and Rocha, V.S., 1955.  Geology and Mineral Deposits of the Boléo Copper District. Baja California, Mexico:  U.S. Geological Survey Professional Paper 273, 134p.

Wright, F.D. 1997, Pre-feasibility Study Final Report:

Volume 1, Final Report; Volume 2, Cost Report; Boléo Project, Sept 1997.  Unpublished report prepared for Minera Curator, S.A. de C.V.

Agapito Associates, Inc. (2006), Preliminary Geotechnical Performance Study for Underground Mining of El Boléo Copper Cobalt Project, Texcoco Test Mine Including Operations Observations and Recommendations, draft report to Baja Mining Corp, July 2006.



 

 

 

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Agapito Associates, Inc. (2007), “Geotechnical Evaluation for Underground Mine Design,” draft report to Baja Mining Corp, February 2006.




























 

 

 

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