Tschudi Copper Deposit Geological Modelling and …weatherlyplc.com/.../03/Coffey_Tschudi_Resource... · Coffey Mining (SA) Pty Ltd DOCUMENT INFORMATION Tschudi Copper Deposit Geological

Coffey Mining (SA) Pty Ltd (2006/030152/079) VAT Number (415 023 9327)

Block D, Somerset Office Estate, 604 Kudu Street, Allen’s Nek 1737 Roodepoort, South Africa www.coffey.com/mining

Tschudi Copper Deposit Geological Modelling and Mineral Resource Estimate

Prepared by Coffey Mining (SA) (Pty) Ltd on behalf of:

Weatherly Mining Namibia Limited

Coffey Mining (SA) Pty Ltd

DOCUMENT INFORMATION

Tschudi Copper Deposit Geological Modelling and Mineral Resource Estimate 9 November 2009

Author(s): Mark McKinney Exploration Projects Manager (BSc Hons Geology, Pr.Sci.Nat)

Ian Brown Senior Field Geologist (BSc Geology)

Allan Goldschmidt Resource Consultant (BSc Hons Geology, Pr.Sci.Nat)

Ken Lomberg Regional Manager (BSc Hons Geology, BCom, Pr.Sci.Nat)

Date: 13 November 2009

Project Number: JTSC01

Version / Status: v.02 Final

Path & File Name: Document3

Print Date: Friday, 13 November 2009

Copies: Weatherly Mining Namibia Limited (2)

Coffey Mining – Johannesburg (1)

Document Review and Sign Off

Ian Brown Author

Mark McKinney Author

Alan Goldschmidt Author

Supervising Principal Reviewed By Ken Lomberg

This document has been prepared for the exclusive use of Weatherly Mining Namibia Limited (“Client”) on the basis of instructions, information and data supplied by them. No warranty or guarantee, whether express or implied, is made by Coffey Mining with respect to the completeness or accuracy of any aspect of this document and no party, other than the Client, is authorised to or should place any reliance whatsoever on the whole or any part or parts of the document. Coffey Mining does not undertake or accept any responsibility or liability in any way whatsoever to any person or entity in respect of the whole or any part or parts of this document, or any errors in or omissions from it, whether arising from negligence or any other basis in law whatsoever.



Table of Contents

Executive Summary ................................................................................................................................i

1 Introduction ................................................................................................................................. 1 1.1 Scope of Work .................................................................................................................. 1 1.2 Terms of Reference .......................................................................................................... 1 1.3 Site Visits .......................................................................................................................... 1 1.4 Participants ....................................................................................................................... 2 1.5 Technical Report .............................................................................................................. 2 1.6 Principal Sources of Information ...................................................................................... 2 1.7 Qualifications and Experience .......................................................................................... 3

2 Disclaimer .................................................................................................................................... 0

3 Property Description ................................................................................................................... 1 3.1 Project Location ................................................................................................................ 1 3.2 Physiography and Climate ............................................................................................... 2 3.3 Local Infrastructure and Services ..................................................................................... 3

4 History .......................................................................................................................................... 4 4.1 Ancient History ................................................................................................................. 4 4.2 Modern History ................................................................................................................. 4 4.3 Historical Mineral Resource Estimates ............................................................................. 5

5 Geological Setting ....................................................................................................................... 7 5.1 Regional Setting ............................................................................................................... 7 5.2 Deposit Types ................................................................................................................. 12 5.3 Local Structure ............................................................................................................... 14 5.4 Local Geology ................................................................................................................. 15 5.5 Geological Modelling Method ......................................................................................... 18

6 Mineralization .............................................................................................................................. 0 6.1 Introduction ....................................................................................................................... 0 6.2 Oxide Mineralization ......................................................................................................... 0 6.3 Transitional Mineralization ................................................................................................ 0 6.4 Sulphide Mineralization .................................................................................................... 1 6.5 Dolomite Mineralization .................................................................................................... 1 6.6 Ore Genesis ..................................................................................................................... 2 6.7 Zone Surfaces .................................................................................................................. 4 6.8 Distribution and Geometry ................................................................................................ 6

7 Exploration ................................................................................................................................. 14 7.1 Previous Exploration....................................................................................................... 14 7.2 Exploration Potential ....................................................................................................... 15



8 Drilling ........................................................................................................................................ 17 8.1 Drilling by Previous Owners ........................................................................................... 17 8.2 Drilling by Current Owners ............................................................................................. 19

8.2.1 Introduction ............................................................................................................ 19 8.2.2 Pre-Collaring ......................................................................................................... 23 8.2.3 Reverse Circulation Drilling ................................................................................... 23 8.2.4 Diamond Drilling .................................................................................................... 25 8.2.5 Down Hole Surveys ............................................................................................... 25 8.2.6 Geotechnical Core Logging ................................................................................... 26 8.2.7 Geological Core Logging ....................................................................................... 26 8.2.8 Geological Chip Logging ....................................................................................... 28 8.2.9 Drillhole Location and Topographical Survey ........................................................ 29 8.2.10 Data Storage and Management ............................................................................ 29

9 Sampling Method and Approach ............................................................................................. 30 9.1 Core Sampling ................................................................................................................ 30 9.2 RC Chip Sampling .......................................................................................................... 31 9.3 Relative Densities ........................................................................................................... 31 9.4 Sample Preparation and Analysis .................................................................................. 32

9.4.1 Sample Preparation ............................................................................................... 32 9.4.2 Analysis ................................................................................................................. 34

9.5 Analytical Quality Control Data ....................................................................................... 35 9.5.1 Introduction ............................................................................................................ 35 9.5.2 Definitions .............................................................................................................. 35 9.5.3 QA/QC Programme ............................................................................................... 35 9.5.4 Quality Control - Monitoring ................................................................................... 37 9.5.5 Methodology .......................................................................................................... 38 9.5.6 Blanks .................................................................................................................... 38 9.5.7 Standard GBM 306-16 .......................................................................................... 38 9.5.8 Standard GBM 996-3 ............................................................................................ 39 9.5.9 Standard GBM 905-11 .......................................................................................... 39 9.5.10 Standard GBM 303-2 ............................................................................................ 39 9.5.11 Field Duplicates ..................................................................................................... 39 9.5.12 Referee Analysis ................................................................................................... 39 9.5.13 Chain of Custody – Responsibility and Accountability .......................................... 40 9.5.14 Conclusions ........................................................................................................... 40

10 Historic Data Verification ......................................................................................................... 54 10.1 Twin Program ................................................................................................................. 54

10.1.1 Twin Drillholes ....................................................................................................... 54 10.1.2 Twin Results .......................................................................................................... 54 10.1.3 Conclusion ............................................................................................................. 58

10.2 Historic Database Reconstruction .................................................................................. 59



10.2.1 Geology ................................................................................................................. 59 10.2.2 Downhole Survey .................................................................................................. 59 10.2.3 Collars ................................................................................................................... 59 10.2.4 Assays ................................................................................................................... 59 10.2.5 Historic Core .......................................................................................................... 59 10.2.6 Comparison with Recent Drilling ........................................................................... 60

11 Mineral Resource Estimate ...................................................................................................... 61 11.1 Drillhole Database .......................................................................................................... 61 11.2 Resource Block Models .................................................................................................. 69 11.3 Grade Interpolation ......................................................................................................... 69

11.3.1 Lower Mineralized Zone (OB0) ............................................................................. 69 11.3.2 Upper Zones (OB1 – OB14) .................................................................................. 71

11.4 Resource Classification .................................................................................................. 72 11.5 Resource Statement ....................................................................................................... 75

12 Conclusions and Recommendations ...................................................................................... 82

13 References ................................................................................................................................. 83



List of Tables Table 1 – Mineral Resource Estimate and Classification iii

Table 4.3_1 – Mintek Mineral Resource Estimate 6

Table 5.1_1 – Stratigraphy of the Otavi Mountainland 10

Table 8.2.1_1 – Summary of 2007-2008 Drilling Campaign 22

Table 9.5.3_1 – Expected Values for Geostats Standards Used 36

Table 10.1.2_1 – Summary of Twin Drilling Results 55

Table 11.1_1 – Descriptive Statistics for Cu % for 1m Composites 62

Table 11.1_2 – Descriptive Statistics for Ag g/t for 1m Composites 63

Table 11.1_3 – Descriptive Statistics for Total Composite Length 64

Table 11.1_1 – Frequency Histograms 1m Composite Samples Cu 65

Table 11.1_2 – Frequency Histograms 1m Composite Samples Ag 67

Table 11.5_1 – Mineral Resource Estimate and Classification 75

Table 11.5_2 – Grade and Tonnage Estimates 76

Table 11.5_3 – Mineral Resource Estimate and Classification at Cu 0.3% Cut-off 77










List of Figures Figure 3.1_1 – Location of the Tschudi Project 1

Figure 3.2_1 – General View of the Tschudi Project 2

Figure 5.1_1 – Regional Geology and Prominent Mines of Namibia 9

Figure 5.1_2 – Regional Geology of the Otavi Mountain Land 11

Figure 5.4_1 – Local Tschudi Geology 16

Figure 6.7_1 – Longitudinal Projection, Showing Base of Weathering and Grade Shells 5

Figure 6.8_1 – Horizontal Section at the 1,270m Rl 7



Figure 6.8_4 – Section E250, Showing Simplified Geology and Grade Shells 10

Figure 6.8_5 – Section E1800, showing simplified geology and grade shells 11

Figure 6.8_6 – Isometric View Looking Southwest 13

Figure 7.2_1 – Regional Cu Geochemical Soil Sampling Map 16

Figure 8.1_1 – Historical Drilling Campaigns 18

Figure 8.2.1_1 – 2007 – 2008 Drillhole Collar Plan 20

Figure 8.2.2_1 – Pre-collar chips and RC Samples at Rig Awaiting Logging 23

Figure 8.2.3_1 – RC Drill Rigs 24

Figure 8.2.7_1 – Core Logging and Storage on Site 27

Figure 8.2.7_2 – Example of Core Marked for Sampling 27

Figure 8.2.8_1 – Drill Chip Storage on Site 28

Figure 9.4.1_1 – Tsumeb Laboratory Jaw Crusher 33

Figure 9.4.1_2 – Tsumeb Laboratory Ring Pulveriser 33

Figure 9.4.2_1 – Tsumeb Laboratory AAS Room 34

Figure 9.5.6_1 – QA/QC: Blank: Cu 41

Figure 9.5.6_2 – QA/QC: Blank: Ag 42

Figure 9.5.7_1 – QA/QC: GBM 306-16: Cu 43


Figure 9.5.8_2 – QA/QC: GBM 996-3: Ag 45





Figure 9.5.10_2 – QA/QC: GBM 303-2: Ag 48

Figure 9.5.11_1 – QA/QC: Field Duplicates: Cu 49

Figure 9.5.11_2 – QA/QC : Field Duplicates: Ag 50

Figure 9.5.13_1 – Referee Analysis Comparison: Cu 51

Figure 9.5.13_2 – Referee Analysis Comparison: Ag 52

Figure 9.5.13_3 – Referee Analysis Comparison: Ag >10 x Detection Limit 53

Figure 10.1.2_1 – Twin Drillhole Comparison Mineralized Length 56

Figure 10.1.2_2 – Twin Drillhole Comparison Average Cu % 56

Figure 10.1.2_3 – Twin Drillhole Comparison Average Ag g/t 57

Figure 10.1.2_4 – Twin Drillhole Comparison Cu % m 57

Figure 10.1.2_5 – Twin Drillhole Comparison Ag g/t m 58

Figure 10.2.5_1 – Photograph showing Condition of Historical Core 60

Figure 11.3_1 – Downhole Variogram Cu% 70

Figure 11.3.2_1 – Composite Variogram for Upper Mineralized Zones - Cu% 71

Figure 11.3.2_2 – Composite Variogram for Upper Mineralized Zones - Ag g/t 72

Figure 11.4_1 – Grade Shells and Drillhole Location Plan Showing Resource Classification Perimeters 73

Figure 11.5_1 – Mineral Resource Estimate Grade Tonnage Curves 81

List of Appendices Appendix A – Geological Logging Codes and Templates 1

Appendix B – Twin Hole Comparisons 1


Tschudi Copper Deposit Geological Modelling and Mineral Resource Estimate Page: i 13 November 2009

EXECUTIVE SUMMARY

Coffey Mining (SA) (Pty) Ltd (Coffey Mining) was engaged by Weatherly Mining Namibia Limited (Weatherly), to design and manage the exploration and delineation drilling of the Tschudi Copper Project in Namibia, and to estimate the mineral resources. This report details the geology of the deposit and the mineral resource estimate.

The Tschudi deposit is centred on coordinates of 19°15’55” South and 17°31’14” East, at a mean elevation of 1,298m amsl. The project area is located some 21km west of Tsumeb, northern Namibia. Access to the project area is by the sealed Tsumeb – Etosha road for 9km, and then 12km along a well maintained gravel road. Skilled labour and most services are available in Tsumeb, which has a long and mature history of mining, ensuring that most services and supplies are available.

The mineral rights for the Tschudi project are owned by Ongopolo Mining Limited, a wholly owned subsidiary of Weatherly Mining Namibia Limited. The area is covered by prospecting license EPL132A and Mining Licence ML125.

Coffey Mining personnel were on site on a continuous basis from January 2007 to September 2008. Mr Mckinney visited the site regularly in a supervisory role and Mr Lomberg visited the site in November 2007 for due diligence purposes.

There has been a long history of exploration on the deposit, as well as throughout the entire Otavi Mountainland region. The Tschudi deposit was first identified from a geochemical survey soil anomaly in 1968. Since then there have been various drilling campaigns, as well as a decline shaft sunk. For the purposes of this report, 218 historical drillholes have been used, as well as 220 drillholes from the recent 2007/8 campaign, drilled from surface and underground. The recent campaign was separated into various different stages, i.e. close spaced and diamond drilling adjacent to existing underground workings, as well as a combination for the infill drilling for delineation of the proposed open pit. Some 13 drillholes were also drilled underground.

The Tschudi ore body is hosted within the basal arenites and conglomerates of the Mulden Group, which lies unconformably on the dolomitic carbonate sediments of the Otavi Group. The ore body is roughly planar, outcropping in a NE – SW direction, and dipping at approximately 30° to the NW. The strike length of the drilled mineralization is approximately 2,500m, although is open ended to depth and the southwest. The copper mineralization in the oxide zone is largely composed of chalcocite and malachite, and within the sulphide zone chalcopyrite and bornite dominate, while within the transitional zone there is a combination of these minerals. There are a host of secondary and supergene minerals, including azurite, chrysocolla, cuprite, digenite, dioptase, dufftite, mottramite and tennanite, amongst others. Silver is a common secondary economic metal that is usually contained within bornite, chalcocite, digenite and tennanite.

The deposit has been drilled out over some 2.5km. Oxide mineralization extends down to approximately 55m below surface. There is then a transitional zone of mixed sulphide-oxide mineralization to a vertical depth of approximately 75m, followed by a dominantly sulphide zone.


Tschudi Copper Deposit Geological Modelling and Mineral Resource Estimate Page: ii 13 November 2009

In order to model the mineralization sectional interpretation was completed on each of the 60 drill section lines and >0.15% Cu grade shells digitised, clipped to the drillhole traces. In order to gain continuity some <0.15% samples have been included within the grade shells where they are bound by >0.15% samples. Sectional interpretations were extended to some 30m beyond the deepest holes, 400 - 500m below surface. A base of oxide surface was created using a synthesis of >50% of Cu sulphuric acid soluble where data was available and base of weathering where it was not.

Silver broadly tracks copper (elevated silver values do not occur without elevated copper values) and so the grade shells are considered valid for both copper and silver. Sectional grade shell interpretations excluded all mineralization within the Otavi dolomite as this was considered too erratic to model with any confidence. In all cases the Lower Zone grade shells were modelled with the Otavi/Mulden contact as their lower edge. The sectional interpretations were then wireframed into solids in three dimensions. Shells were cropped to the surveyed surface topography.

The majority of the mineralization is confined to a robust zone directly above the contact that extends throughout the 2.5km strike length and 925m down dip length of the deposit modelled.

The true width of mineralization in this Lower Zone varies from 1 - 45m with large areas of 10 - 15m thick mineralization running roughly parallel to strike at 80 - 150m below surface. Mineralization generally narrows and decreases in tenor towards surface and down dip. It is open to depth and along strike although there is a general fall in tenor and width in both directions.

Varying from 0-15m above the Lower Zone, 14 discontinuous lenses forming the Upper Zone have been defined. These attain a maximum of approximately 25m thick, more commonly 2 - 5m, and plunge shallowly to the northwest along apparent dip of the strata at 20° -25°. Plan extent is typically 450m down plunge (varying from 250 - 700m) by 80m wide (varying from 35 - 200m).

As with the Lower Zone, mineralization is generally poorly developed and narrow near surface. The Upper Zone lenses all appear to terminate with depth, the deepest intersection being some 225m below surface. Upper Zone lenses may be locally stacked providing near continuous mineralization from the Otavi contact for significant distances. However, they separate from one another and the Lower Zone mineralization laterally.

The drillhole sample database contains samples that were usually 1m in length, hence all samples within the mineralized volumes were composited to 1m. Below detection limit values and values of 0 for Cu and Ag were replaced with half the assay trace value i.e. 0.005% for Cu and 0.5g/t for Ag. Very high Ag values in the sample database were removed by cutting the data at the 99.5 percentile value of 89.23g/t.

A Datamine block model of the mineralized zones was created. In order to model the thin mineralized volumes within the Upper Zone, the block model was created with parent cell measuring 10m x 10m in plan and 3m vertically. The block model was truncated at the surface using a topographic surveyed surface and model cells were classified as being oxide or sulphide based on a surface representing 50% acid soluble Cu. A volume enclosing the historical surface and underground workings was removed from block model.


Tschudi Copper Deposit Geological Modelling and Mineral Resource Estimate Page: iii 13 November 2009

Cu and Ag grade values were interpolated into the model of Lower Zone using the 1m composite samples and an inverse distance methodology. In order to preserve potential horizontal layering, the search radii used for interpolation was dynamic and changed orientation as the dip and strike of the mineralization changed. For the Upper Zone, full length composite samples were used for grade interpolation and the estimates were made using ordinary kriging.

The deposit has been subject to a number of surface drilling programmes. Areas representing three drillhole densities have been defined and have been used for the classification of the resource estimates. The area of underground mining has been excluded from the mineral resource estimate. The results of the exercise are given in Table 1.

Table 1Tschudi Project

Mineral Resource Estimate and Classification at Cu 0.3% Cut-off

Domain Resource Category Tonnage (kt)

Cu (%)

Ag (g/t)

Cu Metal (t)

Ag Metal (kg)

Oxide

Measured 81 1.11 10.71 896 865 Indicated 4,546 0.73 7.82 33,004 35,533 Measured and Indicated 4,627 0.73 7.87 33,900 36,398 Inferred

Sulphide

Measured 4,347 1.09 11.15 47,594 48,494 Indicated 19,869 0.94 11.82 185,990 234,886 Measured and Indicated 24,217 0.96 11.70 233,584 283,379 Inferred 18,874 0.74 9.85 140,482 185,966

Total

Measured 4,428 1.10 11.15 48,490 49,359 Indicated 24,416 0.90 11.08 218,994 270,419 Measured and Indicated 28,844 0.93 11.09 267,484 319,777 Inferred 18,874 0.74 9.85 140,482 185,966


Tschudi Copper Deposit Geological Modelling and Mineral Resource Estimate Page: 1 9 November 2009

1 INTRODUCTION

1.1 Scope of Work

Coffey Mining (SA) (Pty) Ltd (Coffey) was engaged by Weatherly Mining Namibia Limited (Weatherly) to design and manage the exploration and delineation drilling of the Tschudi Copper Project in northern Namibia and to estimate a mineral resource for the deposit. This report details the geology of the deposit and the mineral resource estimate. It is understood that this report is to be used for stock exchange reporting and due diligence purposes, and will form the basis of a feasibility study for an open pit development.

1.2 Terms of Reference

Collectively, Coffey Mining was commissioned to carry out the following activities in the resource estimation study of the Tschudi Project:-

Design and management of the drilling and sampling program.

Compilation, validation and migration of previous drilling data into a new database.

Three dimensional modelling of the geology of the deposit.

Quality Assurance and Quality Control (QA/QC) of assay laboratory.

Drilling of twin holes and validation of the previous drilling results.

Database generation, review and management.

Completion and validation of the mineral resource estimate based on historical data and additional data obtained during this phase of drilling, consistent with the JORC code.

Compilation of a report.

1.3 Site Visits

Mr Mckinney (Exploration Projects Manager – Southern Africa) visited the site on a regular basis during the on site project work conducted between January 2007 and September 2008. Mr Mckinney has assisted the project geologists with various aspects of the drilling programme including drill program design, database management, QA/QC requirements, geological interpretation and assay laboratory management. Site work was undertaken by a team of Coffey Mining geologists.

Mr Lomberg (Regional Manager – Southern Africa) visited the site in November 2007 as part of a due diligence review.



1.4 Participants

The Coffey Mining personnel involved in the technical review and mineral resource estimation of the Tschudi Project, including their principal areas of responsibility, are listed below:-

Mark Mckinney, Exploration Projects Manager – Southern Africa Management of exploration and drilling, QA/QC analysis, finalisation of the geological model, compilation, reporting.

Ian Brown, Senior Field Geologist Field work, modelling, reporting.

Allan Goldschmidt – Resource Consultant Mineral resource estimation, reporting

Ken Lomberg – Regional Manager Southern Africa Peer review

1.5 Technical Report

The report is consistent with the “Australasian Code for Reporting of Mineral Resources and Ore Reserves” of December 2004 (the JORC Code) as prepared by JORC of the Australasian Institute of Mining and Metallurgy, Australian Institute of Geoscientists and the Mineral Council of Australia.

The satisfaction of requirements under both the JORC and Valmin Codes is binding on the authors as Coffey Mining is a Corporate Member of the Australasian Institute of Mining and Metallurgy (AusIMM).

1.6 Principal Sources of Information

Data and information required to complete the exploration and mineral resource estimation were provided by Weatherly to Coffey Mining. These data and information are summarised as follows:-

Drillhole data from previous drilling campaigns in hardcopy format.

Previous mineral resource estimation parameters and reports.

Wireframe solids of existing underground development and stoping.

Excel tabulations of tonnages mined.



1.7 Qualifications and Experience

Coffey Mining is part of Coffey International Limited which is one of the top 300 companies on the Australian Stock Exchange. Coffey International Limited consists of a range of specialist companies working in social infrastructure and physical infrastructure and operates in more than 60 countries around the world.

Coffey Mining is an integrated Australian-based consulting firm, which has been providing services and advice to the international mineral industry and financial institutions since 1987. Coffey Mining, previously RSG Global, has maintained a fully operational office at Accra in Ghana since 1996, providing an operational base for consulting and contracting assignments throughout the West African region. An additional African office was established in Johannesburg, South Africa, in 1999 to support expanding activities within southern and eastern portions of the continent. In 2007 an additional office was established in Lusaka, Zambia to provide consulting services to the Zambian Copperbelt in particular.

Neither Coffey Mining, nor the key personnel nominated for the completed and reviewed work, have any material interest in Weatherly Mining Namibia Limited or its mineral properties. The proposed work, and any other work done by Coffey Mining for Weatherly, is strictly in return for professional fees. Payment for the work is not in any way dependent on the outcome of the work, nor on the success or otherwise of Weatherly’s own business dealings. As such there is no conflict of interest in Coffey Mining undertaking the Independent Qualified Person’s Report as contained in this document.



2 DISCLAIMER

Coffey Mining relied on data supplied by Weatherly for its estimate of the Tschudi Project. These data include historic drillhole data, third party technical reports prepared by previous company geologists, unpublished theses, and other relevant published and unpublished third party information. Coffey Mining has endeavoured, by making all reasonable enquiries, to confirm the authenticity and completeness of the third party technical data upon which this report is based. A final draft of this report was provided to Weatherly, along with a written request to identify any material errors or omissions.

Neither Coffey Mining nor the authors of this report are qualified to provide extensive comment on the legal aspects associated with ownership and other rights pertaining to Weatherly mineral tenements. Coffey Mining did not apply any legal due diligence to confirm such title. Similarly, neither Coffey Mining nor the authors of this report are qualified to provide comment on any environmental issues associated with Weatherly.

No warranty or guarantee, be it expressed or implied, is made by Coffey Mining with respect to the completeness or accuracy of the legal, environmental or metallurgical information referred to in this document.



3 PROPERTY DESCRIPTION

3.1 Project Location

The Tschudi Project is located some 21km west of Tsumeb in northern Namibia (Figure 3.1._1). Access to the project area is by a sealed road from Tsumeb for 9km and some 12km along a well maintained gravel road to the project site. The project is centred on coordinates 19°15’55” South and 17°31’14” East, at a mean elevation of 1,298m above mean sea level (amsl).

Figure 3.1_1 Location of the Tschudi Project



3.2 Physiography and Climate

The Tsumeb area is semi arid, with an average annual rainfall of 500mm, falling mainly from December to April. The climate is sub-tropical, with mean summer temperatures averaging 35°C, and mean winter temperatures average around 10°C. Thus the field conditions are such that it is possible to work all year round, poor weather conditions rarely disrupt exploration or mining operations.

The Tschudi Project area lies on a flat plain at an altitude of approximately 1,300m amsl, sandwiched between low lying dolomitic hills both to the southern and northern sides (Figure 3.2_1). The dolomite hills trend approximately parallel to the regional strike and are situated approximately 1km south of the deposit and 3.5km on the northern side.

Figure 3.2_1 General View of the Tschudi Project from a historic open pit across the sandstone plains to low

dolomite hills north of the deposit.



The soil type is closely related to the bedrock lithology. Over the sandstone plain is a thin red-brown sandy soil, overlying one to 10m of white calcrete. Below this calcrete are the arenaceous sandstones of the Mulden Group. The dolomite hills are covered by a thin layer of dark brown sandy soil, with abundant chert boulder rubble.

The vegetation in the plain is dominantly low bushland, comprised of Acacia and Dichrostachys species, interspersed with open grassland. On the hills there is a more diverse range of tree cover. The water table is at approximately 70 – 80m below surface.

The main land use in the area is cattle farming, as well as game farming in the reserve to the north of the property.

3.3 Local Infrastructure and Services

The project area lies almost adjacent to a major tarred road that links Tsumeb to Etosha and Angola to the north. There is an electrical power line located on the project site, utilised in the current underground operations. The railway linking Tsumeb to Swakopmund and Windhoek passes 5.5km to the east.

Skilled labour and most services are available in Tsumeb, and Namibia has a well established mining industry. Tsumeb serves as a base for providing a full range of urban amenities, including medical and educational facilities, financial, retail and commercial services. Tsumeb is a major tourist destination, thanks to the proximity of Etosha National Park. Thus modern hotels, lodges, shops and restaurants are able to provide most services. Telephone and mobile phone services are reliable, as are the high-speed internet facilities. Mobile phone reception is patchy, but present on site.



4 HISTORY

4.1 Ancient History

There is evidence of ancient peoples locally smelting copper in the Tsumeb area. It appears that the Bergdama tribe were smelting copper long before the Herero or Ovambo peoples. Primitive smelting works are common throughout the Otavi Mountainland, and evidence from localities such as Gross Otavi and Otjikoto suggests that the style and techniques used were similar to those used in Central Africa prior to 500 AD (Cairncross, 1997). The local Bushmen are said to have sold the copper ore to the Ovambos from the north, who were skilled metal workers (Emslie, 1979).

4.2 Modern History

Since the discovery of the Tsumeb deposit by Europeans in 1842, there has been a large amount of exploration throughout the entire Otavi Mountainland Province. Sir Francis Galton and a scientist named Charles Andersson reported the presence of copper smelting by the local population at Tsumeb – ‘The hill of the Frog’. Germany originally had possession of the territory, and granted a mining concession to the South West Africa Company (SWACO) in 1892. This concession included sole mineral rights for almost the entire northern half of what is now Namibia. In 1892 several of the known copper occurrences in the Otavi Valley, namely Tsumeb, Gross Otavi and Asis, were investigated. In 1900 SWACO granted the mineral rights for a 1,200km2 concession to the Otavi Minen Und Eisenbahn Gesellschaft (OMEG). SWACO and OMEG’s properties were lost at the end of the Fist World War, and subsequently sold to the newly formed Tsumeb Corporation Limited (TCL) (Misiewicz, 1988). Significant exploitation at Tsumeb began in 1906, and has continued intermittently since then (Lombaard et al, 1986). Since the early 1900’s there have been numerous exploration campaigns throughout the Otavi Mountainland by TCL, which have led to the discovery of many economic deposits of copper, silver, zinc, vanadium and lead, amongst others. Many of these deposits have been turned into operating mines, such as Kombat, Asis, Abenab, Berg Aukas, etc.

The first record of prospecting in the Tschudi area was in 1913, when a Mr. Hoepker pegged the land on farm Uris. An OMEG report dated 25 October 1920 reported copper mineralization in limestone and dolomite beds near the Otavi dolomite – Mulden sandstone contact some 200m east of the Post II water hole on Uris. This was noted as two ‘aplite’ (feldspathic sandstone) lenses, both well mineralized with malachite. From 1916 to 1920 three other small scale mines were operating in the vicinity, namely Alt Bobos (Cu-V), Karavatu (Cu-Pb-V), and Uris (Cu-Pb-V) (Murphy, 1980).

The main outcrop area was trenched and sampled by TCL geologist A. T. Griffis at the beginning of 1948. Since then there have been further more detailed exploration work carried out, described in Chapter 7.1.



In 1991 a decline shaft was sunk and limited development put in including five cross-cuts through the oxide ore at the 1,184m elevation. The material mined was used as a bulk sample in a trial run through the Tsumeb plant. In addition two small open pits on the oxide material were excavated.

In 2007 Weatherly re-opened and extended the trial mining development and mined approximately 209,000 tonnes of oxide and supergene material. This was treated at the Tsumeb plant. Mining operations ceased in December 2008 due to the fall in the copper price.

4.3 Historical Mineral Resource Estimates

Mineral resource estimation and pre-feasibility studies were completed by Geotech Africa cc, on behalf of TCL, in August 1997. This study estimated an open pittable Indicated Resource of 15.15 million tonnes grading 0.93% Cu, with negligible Pb and Ag.

This mineral resource estimate was updated by James Lonergan of Mintec Inc, and Abraham Saayman of Geotech Africa cc, in August 2002. The geological modelling and mineral resource estimate was compiled using historic geologic logging information and assay laboratory test results supplied by Ongopolo Mining and Processing Limited. The nature of the data allowed for an accuracy range of +/- 30%, placing the report in the pre-feasibility category.

The orebody model was based on geologic logging information from 190 drillholes, and was separated into Oxide, Transitional and Sulphide mineralized zones. A three dimensional block model with a block size of 10m by 10m by 5m vertical was constructed for the resource classification. These estimated resources are presented in Table 4.3_1.



Table 4.3_1 Tschudi Project

Mintek Mineral Resource Estimate

Ore Type Category Tonnage (kt)

Grades

Cu (%) Ag (g/t)

Oxide

Measured 3,444 0.661 6.1197

Indicated 3,457 0.574 5.3963

Inferred 63 0.771 14.0254

TOTAL: 6,964 0.619 5.8322

Transitional

Measured 1,574 0.724 10.5434

Indicated 1,542 0.633 9.4825

Inferred 3 0.171 1.1179

TOTAL: 3,118 0.679 10.0103

Sulphide

Measured 8,555 0.896 11.052

Indicated 2,1314 0.884 11.8316

Inferred 3,465 0.841 10.9628

TOTAL: 33,334 0.882 11.5412

TOTAL

Measured 13,573 0.816 9.7414

Indicated 26,312 0.828 10.8485

Inferred 3,531 0.839 11.0096

TOTAL: 43,416 0.825 10.5155



5 GEOLOGICAL SETTING

5.1 Regional Setting

The Tschudi Project area is located within the Otavi Mountainland of northern Namibia, which forms part of the Northern Carbonate Platform of the Pan African Damaran orogen. The Damara Supergroup is an orogenic belt that was deposited on a pre-1.0 Ga granitoid basement, the Grootfontein Basement Complex. It is composed of a 400km wide north-east trending arm, as well as two coastal arms, that all join in the region of Swakopmund, on the western coast of Namibia. This entire sequence was formed by the deposition of a geosyclinal sequence approximately 900 to 650Ma, caused by the separation of the Kalahari, Congo and proto-South American cratons. This rifting allowed the deposition that formed the Damara, which was then followed by a period of compression. The northern arm is labelled as the Kaoko Belt, and the southern arm is the Gariep Belt, which overlies the Namaqua Metamorphic Complex in southern Namibia. The Otavi Mountainland is located within the central intra-continental arm (Figure 5.1_1) (Misiewicz, 1988).

The sediments of the Damara Supergroup were unconformably deposited on the folded and peneplaned Grootfontein Basement, composed of granite, gneiss and poorly exposed mafic complex, characterised by common rift grabens. The oldest Damara sediments comprise mafic lavas, mica schists, conglomerates and arenites that form the Nosib Group. This is a discontinuous succession up to 750m thick, that was deposited in five NE trending, fault bounded, grabens, and is unconformably overlain by the Otavi Group. The Otavi sediments are divided into two groups by a regional disconformity:-

The lower Abenab Sub-group (dolomite and limestone, with minor shale and arenaceous units).

The upper Tsumeb Sub-group (dolomite with subordinate chert and limestone bands).



The Otavi Group reaches a maximum thickness of approximately 7,000m and is interpreted as being deposited on a stable marine shelf (Murphy, 1980). Sedimentation of the 3,000m thick Tsumeb Group started with the deposition of fluvoglacial diamictites (Chuos Formation), followed by shallow shelf marine sedimentation of laminated limestones and marls with abundant slump breccias and argillite bands (Maieberg Formation). Conformably overlying these are the dolomites and cherts of the Elandshoek Formation that have been exposed to secondary silicification, forming the rugged terrain around Tsumeb and the exposed hills of the Otavi Valley Syncline. The uppermost portion of the Tsumeb Subgroup, the Huttenberg Formation, is a light grey dolomite sequence characterised by algal chert lenses and oolitic and pisolitic concentrations towards the top of the sequence (Misiewicz, 1988). There are eight distinguishable informal lithozones within the Tsumeb Subgroup (Table 7.1_1). Sedimentological and carbon-isotope studies have shown the older zones to have been deposited in relatively deeper cold water during the collision with the Kaoko Belt to the west, during and after a glacial event, observed in the lowermost zone. The deposition of the later zones are characteristic of extentional tectonics, and a progressively warmer climate, as evidenced by the presence of chert, oolites, pisolites, stromatolites and algal mats within the Huttenberg Formation (Melcher, 2003).



Figure 5.1_1

Regional Geology and Prominent Mines of Namibia




Stratigraphy of the Otavi Mountainland

GROUP SUBGROUP FORMATION LITHOLOGY MULDEN Owambo Pelite, marl, carbonate

Kombat Shale Tschudi Arenite

OTAVI

Tsumeb

Huttenberg T8 Dolomite T7 Pelite T6 Dolomite

Elandshoek T5 Chert T4

Dolomite Maieberg

T3 T2 Limestone

Ghaub T1 Diamictite

Abenab

Auros Carbonate, pelite Gauss

Dolomite Berg Aukas Chuos Diamictite

NOSIB Askevold Volcanics Nabis Clastics

Grootfontein Basement Complex (adapted from Melcher, 2003)

Above the Otavi dolomites are the Mulden Group sandstones which attain a maximum thickness of about 700m. The Mulden Group is a clastic molasse sequence which was deposited during the early stages of the Damara Orogeny. It marks a drastic change in the depositional environment, from the stable marine shelf environment of the Otavi Group sediments, to one of fluvial and deltaic deposition in an intermontane setting (Misiewicz, 1988). The Mulden Group can be divided into three formations by variations in sedimentary facies and geographical accumulation; the Tschudi, Kombat and Owambo Formations (Table 5.1_1). These lithologies are located within the synclines formed by the Otavi dolomites.

The basal Mulden facies has no official status, but consists of a sporadic dirty subgreywacke with common localised chert pebble conglomerates, unconformably lying on the Huttenberg dolomites. These conglomerates were deposited as playa mud flats or shallow lacustrine sediments on the karstic dolomite basement, and become progressively thicker towards the west. Argillite and shale lenses are commonly intermixed within this basal unit. Above this basal unit is the Tschudi Formation, dominated by clean, feldspathic and quartz-arenites, greywackes and silty mudstone lenses, with very poorly developed bedding.



This clastic sequence was probably developed as a delta fan above the basal conglomerate unit, with the clastic sediments being locally derived from the denudation of up-domed basement complex. Above these clean arenites the Mulden Group becomes more strongly bedded, with an increase in pyrite and secondary silicification, as well as occasional hydrothermal quartz-veinlets. Above this facies change is the start of the Kombat Formation, composed of a shale unit that has been metamorphosed to phyllites and slates, deposited by deltaic processes in a deep water environment. These shales gradually grade into the carbonate pelites of the Owambo Formation. The regional Otavi-Mulden contact is mostly unconformable, with sedimentation having occurred on an erosional karst surface (Misiewicz, 1988). These sediments were subjected to a period of orogenic folding during the Palaeozoic, which was followed by a prolonged period of erosion and denudation. It is thought probable that the Otavi Mountainland was at one stage covered by Karoo sediments, although nothing remains of post-Mulden deposition today (Murphy, 1980).

Figure 5.1_2 Regional Geology and of the Otavi Mountain Land



The regional structure of the Otavi Mountainland is characterised by east-west trending, open to isoclinals folds, and north-vergent thrust faults, which were a result of continent-continent collision between the Congo and Kalahari Cratons, approximately 545Ma. These regional structures have been overprinted by a second, northwards compressive, folding phase creating NE-trending open, upright folds. During this folding, the sediments of the Otavi Mountainland were subjected to lower greenschist to prehnite-pumpellyite facies metamorphism (Cairncross, 1997, Melcher, 2003).

5.2 Deposit Types

The carbonate-hosted mineral deposits of the Otavi Mountain Land have traditionally been classified into two main categories of mineralization, separated by stratigraphic, isotopic and mineralogical factors:-

The Tsumeb-type (e.g., Hughes, 1987).

The Berg Aukas-type (e.g., Misiewicz, 1988).

In general, both the Tsumeb-type and Berg-Aukas-type mineralization display features which broadly affiliate them to Mississippi Valley-type (MVT) deposits.

The Tsumeb-type ore deposits are characterized by polymetallic phases of sulfide minerals containing copper, lead, zinc, silver and a host of other secondary metals. The ore minerals are usually disseminated in a variety of structures such as solution breccias and pipes, shear zones and fractures. The orebodies usually occur in the upper portions of the Tsumeb Subgroup, but are not stratabound. The most significant deposits are located close to north-east trending fractures and faults regarded as reactivated basement structures, as well as having an intimate relationship with feldspathic sandstone of the Mulden which has intruded into paleo-casts.

In contrast, the Berg Aukas-type ores are composed of sulfides containing lead, zinc and vanadium, resembling ore from the Zn-Pb-rich Mississippi Valley-type deposits. Copper is rare to absent and Ag, Ge, Ga and Cd, although present, are much less abundant than in the Tsumeb-type ores. Furthermore, the Berg Aukas-type deposits are generally confined to the Abenab Subgroup and middle-lower portions of the Tsumeb Subgroup. Mineralization occurs as breccia bodies, but may be both stratabound and discordant. There is a close relationship with the Grootfontein Basement high.



On the basis of various data, including isotopic and fluid inclusion analyses, the two different types of deposits are considered to have formed during two different mineralizing events, both in time and space. An earlier event produced the Berg Aukas-type deposits and involved fluids derived from compaction and dewatering of sediments. Lead and zinc were scavenged by the fluids which were then deposited as sulfides over structural "highs" in karst breccias and along faults in the Abenab Subgroup carbonates. The vanadium mineralization that accompanies the Berg Aukas deposits resulted from supergene processes, during which vanadium was carried in solution as calcium metavanadate under oxidizing conditions. When these fluids encountered sulphides, and hence a reducing environment, the solution reacted with the metal ions liberated from the sulfides, thereby forming vanadates such as descloizite. This vanadium event postdates the second or Tsumeb-type deposits which were emplaced after the deposition of the overlying Tsumeb Subgroup carbonates. The vanadium event involved copper-rich, higher temperature/lower salinity fluids (Cairncross, 1997).

On the basis of these two broad deposit types, a number of subtypes can be described throughout the OML, with various different ore genesis models.

The Tschudi Deposit is the only significant economic deposit that has been discovered to date that is hosted by the Mulden sandstone, and is closely related to the MVT deposit types (Emslie, 1979) which are characterised by:-

An absence of any apparent igneous activity or igneous rocks which could be potential sources of ore solutions.

Occur in limestone and/or dolomite.

Consist mainly of bedded replacements and vein deposits, with evidence of solution activity, brecciation, slumping and collapse structures.

Form at relatively low temperatures.

The deposits are most common in passive structural regions.

The mineralogy is simple and the metal content is usually low.

The ore bodies occur at shallow depths relative to the present day surface.

The only igneous rocks in the OML region are the granitic suite of the Grootfontein Basement, which is considerably older than both the deposition of the sandstones and the mineralization. The granites may be a secondary source of the mineralizing fluids, but are clearly not a primary source of the mineralization.

The majority of MVT deposits lie in limestone or dolomite and range from Late Cambrian to Jurassic age. The dolomitic rocks of the Otavi Group are also of Late Cambrian age. Although the Tschudi deposit is hosted by sandstone, it is closely related to the underlying dolomite.



The Tschudi orebody is a roughly planar stratiform deposit with the mineralization confined to a distinct stratigraphic zone, bounded by bedding planes. The mineralization is evenly, finely disseminated throughout the deposit, with minor associated solution activity and slumping.

Although the Mulden sediments are rich in pyrite, the copper mineralized areas are low in pyrite. The very low iron content of the copper mineralization is evidence of the low temperature of formation of the deposit.

Tschudi, along with most of the ore deposits of the OML, occurs along the flanks of a fold. The uniformly bedded to massive nature of the Tschudi arenites indicates a very passive structural depositional environment.

Galena and sphalerite are the major ore minerals in classic MVT deposits. Although very rare at Tschudi, these minerals are found throughout the deposits of the OML. The major primary sulphide mineralogy at Tschudi is very simple, consisting of bornite and chalcopyrite.

The mineralization at Tschudi is at the present-day surface and extends down to approximately 400m vertical depth. Many of the deposits in the OML are similar, which is comparable to many MVT deposits around the world (Emslie, 1979).

The combination of these factors all indicate that the Tschudi orebody can be classified as an MVT deposit. This indicates that the mode of genesis of the sulphides is probably similar to that proposed for other low temperature lead-zinc deposits.

5.3 Local Structure

The deposit is located on the southern limb of the Tschudi-Uris Syncline and dips between 25° and 35° to the NNW. The northern limb has a steeper dip of about 50° and the syncline plunges at between 5° to 9° to the west. The axis of the syncline trends between 070° to 085°. No major faults have been identified, but a set of NW trending open fractures intersect the area, observed in underground exposure and drill core.

Three major fracture trends are present in the chert and dolomite horizons. The major set as observed from underground trends at 045° to 070° and is sub-parallel to bedding and dips steeply towards the south. This set has been observed to be associated with some calcite filling and in places has been striated. Other sets trend at 100° to 120° and 140° to 160° and are both steeply dipping. Regionally a major set of fractures trending at 050°, as well as the Heidelberg and Tsumeb Dyke Fault Zones, occur in Otavi Group sediments immediately southwest of the Tschudi/Uris Syncline. Prominent bedding plane shear zones occur on or near the Otavi/Mulden contact and this has been observed in most of the drilled core. Drillhole sections and some underground observations indicate that this contact undulates gently. Some micro-folding has been observed in the dolomite.



5.4 Local Geology

The Tschudi orebody is hosted within the basal arenite of the Mulden Group, the Tschudi Formation, unconformably overlying the Otavi dolomites. The orebody transgresses from the Huttenberg dolomites, through the basal conglomerate, and up into the clean arenites for approximately 15 - 20m.

The Huttenberg dolomites of the Otavi Group form the basement rocks within the Tschudi area. These folded dolomites form the hills exposed to the north and south of the orebody, in what is known as the Tschudi-Uris Syncline. The Tschudi deposit is located on the southern limb of this syncline (Figure 5.4_1), as a planar stratiform deposit, dipping at approximately 25°- 30° to the NNW. In the Tsumeb District the top of the formation is represented by a 20 - 40m thick succession of grey calcitized dolomite, but this is largely absent at Tschudi.

In the few present exposures of this dolomite, it is highly recrystallised and calcitized, and commonly has a high Mn concentration. Either it was largely eroded before Mulden sedimentation, or was not deposited to any great degree. Thus the Mulden sediments dominantly rest on the oolitic-pisolitic chert horizon. Cross bedding is occasionally visible in the cherts, as well as elongation of the oolites, probably due to regional deformation.

Below the chert horizon the dolomite is commonly brecciated, probably due to small isolated solution collapse zones. In some cases the more extensive breccias appear to be caused by slumping in partly consolidated sediments. Throughout the area the dolomite and chert horizons commonly display extensive karst features. In the drill core and underground exposures, these features can be seen as cracks, fissures and caves that have been filled with later Mulden arenitic material, calc-arenites and the basal conglomerate.



Figure 5.4_1 Local Tschudi Geology



In the Tschudi project area the basal conglomerate is very sporadically developed, thought to be confined to karstic and erosional depressions in the dolomite. From the drillhole core and exposures underground the conglomerate can be described as a 5 - 20cm thick, rarely reaching 2m, dark brown polymictic unit, with varying amounts of angular to sub-rounded quartz, chert, dolomite, argillite, and sandstone pebbles, very poorly sorted. The origin of the clasts is thought to be localised, most likely eroded chert clasts from the Huttenberg Formation, as well as quartz pebbles from the exposed Grootfontein Basement complex. Pebble size ranges from granules up to cobbles but small pebbles, 1 - 2 cm wide, predominate. The basal conglomerate and first few metres of arenite are usually oxidised, even when below the average weathering profile. The thickness of this oxidised layer varies, but decreases with depth. In the deeper drillholes this oxidation is usually restricted to the immediate contact zone. Conformably above the conglomerate, and often directly (unconformably) on the dolomite, are the clean arenites of the Tschudi Formation. These arenites continue to the current surface erosion profile, and the upper portions of the Mulden Group are not represented in the project area.

The lower Mulden sediments are composed of quartz-arenite, sub-greywacke, arkose, siltstone and mudstone. The lowermost arenite is a grey, fine to medium grained feldspathic quartzite to sub-greywacke, with very poorly developed bedding. The unoxidised arenites are moderately pyritic, with minor amounts of graphite, creating a black coloured arenite. In the mineralized zones the pyrite content decreases. These arenites are dominantly composed of detrital, sub-rounded quartz grains, with clay minerals (kaolin and montmorillonite), calcite and feldspars (orthoclase and microcline) as the main inter-granular components. There is common sericitic alteration, especially in the oxidised zone, above 70 - 80m below surface. Minor amounts of muscovite, biotite and chlorite are also present, along with reported trace amounts of rutile, zircon and sphene (Murphy, 1980). Interspersed amongst the arenites are occasional siltstone and mudstone lenses, varying from a few centimetres to a few metres thick. These units are confined to the lower 40m of the Mulden arenites. Additionally, there are a few minor thin, coarse and gritty conglomerate beds within the arenites, similar in composition to the basal conglomerate. The clean, poorly bedded, arenites extend for 20 - 60m above the Otavi-Mulden contact, after which the arenites become more strongly bedded, with a slight foliation, and increase in pyrite content. These upper arenites are commonly very hard and glassy (below the oxidised zone) with minor hydrothermal quartz veining, indicating secondary silicification (Murphy, 1980).

Above the arenites, near surface, there is layer of very hard, silicified white calcrete that ranges from one to 15m thick. Above this is a thin sandy soil covering that ranges from 0 - 2m thick.



5.5 Geological Modelling Method

Some 60 historical drill section lines that were used in previous sectional models were digitized, and provided the basis for sectional interpretation. This interpretation was completed with sections centred on drill lines (bearing 331°). In the area surrounding the underground workings the sectional spacing was 25m, which increased to roughly 50m over the remainder of the area. The actual sectional separation varied depending on historical and recent drillhole location. The half distance between two adjacent section lines was used as the clipping window to accommodate line spacing. Strings/polylines were then digitized onto the drillholes, clipped to the actual hole traces in three dimensions, to represent each of the major rock types. This generated a series of three dimensional polylines centred on the drill line but true to the actual downhole lithological contacts.

A series of three dimensional wireframes were then created by triangulation between adjacent sectional polylines. The terminations of these were closed to a point halfway between section lines and closed off vertically at end of drilled strike, at a distance equivalent to the previous section line. These solids were then cut horizontally at 50m elevation separations and the plan interpretation checked against drill results. The sectional polylines were then adjusted for any changes in interpretation arising from this and a final series of solid models generated. These incorporated more detailed terminations generated from plan interpretation to increase accuracy.



6 MINERALIZATION

6.1 Introduction

The Tschudi Deposit hosts various different mineralization facies, separated out into oxide, transitional and sulphide zones. The copper mineralization is preferentially developed in the base of the arenite sequence on the southern limb of the syncline, as a disseminated, continuously distributed roughly planar sheet, varying from two metres to at least 40m thick. There is a continuous basal mineralized zone termed the Lower Zone lying on the dolomite surface, with occasional lenses/pods of mineralization occurring several metres above constituting the Upper Zone. These lenses plunge down-dip, towards the base of the syncline. The mineralization is best developed within the medium to fine grained feldspathic arenites and sub-greywackes. Pyrite often occurs in the fine grained argillites near the base of the arenites, but these seldom contain copper mineralization. At surface outcrop the oxide mineralization occurs over a strike length of approximately 2.5km, and continues down to approximately 55m vertical depth. There is then a transitional zone of mixed sulphide-oxide mineralization to a vertical depth of approximately 75m, followed by a dominantly sulphide zone. The mineralization does locally transgress into the dolomites, as void fillings, in joints, fractures and shear zones. The mineralization is open ended to depth, but is not present to any great degree in the opposite limb of the syncline. There are sporadic soil anomaly indications of copper further along the dolomite-sandstone contact, as well as in the nose of the fold.

Two preliminary petrological studies have been carried out, with thin and polished sections being taken from drillhole core (Murphy, 1980). Copper is the main economic metal targeted, with secondary silver. Lead and zinc have been recorded, with values rarely exceeding 0.1%. No significant concentrations of gold, uranium, molybdenum, arsenic, barium, bismuth, cobalt or vanadium have been detected (Viviers, 1992).

6.2 Oxide Mineralization

In the oxide zone the dominant copper mineral is malachite, accompanied by cuprite, azurite and minor amounts of chalcocite, digenite and covellite. Chalcocite is often seen as rimming pyrite grains, and this appears to be a primary phase of chalcocite.

In the basal conglomerate and lower oxidised arenites the mineralization is dominantly malachite and chalcocite, with lesser amounts of covellite and cuprite. The cuprite is disseminated in clays and gossany limonitic material. Although very poorly represented in the drill core, chrysocolla is commonly present in the surface pits, as infill in veins and fractures.

6.3 Transitional Mineralization

The highest copper grades in the deposit occur in the supergene zone. The transitional zone occupies a poorly defined zone from approximately 55 – 75m vertical depth.



6.4 Sulphide Mineralization

In the unoxidised arenites, hypogene sulphide mineralization occurs as very fine intergranular disseminated grains, often difficult to determine with the naked eye. Pyrite is abundant throughout the Mulden clastics, although it does decrease in the copper mineralized zones.

Bornite and chalcopyrite are the dominant copper sulphide minerals, with some minor chalcocite and digenite. No discrete silver mineral phases have been identified, and the silver is thought to be contained within the lattices of bornite, chalcocite and digenite. One observation of native silver has been recorded (Viviers, 1992). Isolated occurrences of disseminated galena and sphalerite have been noted, marginal to and overlying the copper mineralization.

A vertical zoning of the sulphides is apparent in several of the intersections; the general association varying from iron-rich sulphides at the top, through iron-copper-rich then copper-iron-rich sulphides, to dominantly copper-rich sulphides at the base. A typical sequence of sulphides from top to bottom is:-

Pyrite only.

Pyrite-chalcopyrite.

Pyrite-bornite-chalcopyrite.

Bornite-pyrite-chalcopyrite-chalcocite.

Chalcocite-bornite-pyrite.

Chalcocite-covellite.

In several cases there is a general trend of the copper sulphides to become increasingly iron deficient with depth (Murphy, 1980). Textural evidence indicates that the bulk of the pyrite formed as an early diagenteic phase, and the chalcopyrite, bornite and chalcocite selectively replaced the pyrite.

The copper sulphides occur as discrete interclastic grains which range in size from 1 to 330µm, with the majority being 25 to 50µm in diameter. Minor quantities of bornite are found as inclusions in pyrite (Viviers, 1992).

6.5 Dolomite Mineralization

No primary mineralization has been identified from the dolomite, but chalcocite, covellite, malachite and azurite are sporadically present in the upper 10m of the dolomite/chert sequence. In some cases the mineralization continues down to 34.5m below the contact. Some early drill intersections showed significant mineralized lengths in the dolomite in isolated instances. It was postulated that these may be feeder zones of the Tsumeb style to the diffuse mineralization in the porous overlying sandstone. Where these have been re-drilled with inclined holes they have proved to constitute narrow (<2m wide) zones of spaced



sub-vertical <5cm thick chalcocite and malachite mineralized fractures. Apparent widths were due to vertical drillholes superimposed on sub-vertical stringers.

Where sandstone has infilled karst cavities within the dolomite, these are commonly oxide mineralized. Where mineralization occurs in the dolomite it is commonly associated with strong calcitisation, manganese alteration and graphitic material.

6.6 Ore Genesis

From the observed characteristics of the deposit, the genesis of the Tschudi orebody could be described by several processes, namely a sedimentary-diagenetic or an epigenetic model of origin. In the general diagenetic model, ascribed to the majority of MVT deposits worldwide as well as the OML, basin-derived fluids acquire heat, metals and other solutes during transportation, and then deposit the sulphides in the pore spaces as they emerge from adjacent parts of the basin. The fluids are either driven upwards by sediment compaction (low porosity at depth, higher porosity at surface), or are derived from the basin stream systems. The fluids collect the metals by brine leaching, carrying them as chloride or organic complexes. The metal sulphides are then precipitated when they come into contact with H2S, presumably from nearby evaporates (Anderson & Macqueen, 2003).

In the epigenetic red-bed model, a chemically reducing, carbonaceous and pyritic grey-bed sediment is originally enriched in sulphur (iron sulphide or anhydrites) by primary, syndiagenetic processes. Copper and associated metals are then zonally overprinted on the sulphur rich host during a post-sedimentary influx of dissolved base metals from adjacent, coarse grained, highly porous and permeable red-bed sediments. The deposition of metals is a low temperature chemical reaction between the reducing sulphur (pyrite) in the host rock, and base metals added to these sediments (Brown, 2003).

From these general models, two different scenarios can be postulated for the Tschudi sulphide ore genesis:-

Copperbelt Model (Epigenetic)

Pyrite formed in an early stage of the diagenesis, soon after deposition of the host rock. Metals leached from basement sources were transported to the depository by stream systems. These low temperature, metal bearing, waters encountered the reducing, pyritic environment. Selective replacement of pyrite by chalcocite, bornite and chalcopyrite occurred. Much of the chalcocite, digenite and covellite now present would have been derived by supergene alteration (Viviers, unknown).

Hydrothermal Model (Diagenetic)

The fact that no mineralization similar to Tschudi has been found within the OML, combined with the presence of many showings of hydrothermally derived, carbonate hosted mineralization localities within 6km of Tschudi, could be used as an argument for a localized hydrothermal origin of the deposit. Mineral fluids permeated a structurally prepared locus in



the dolomite lying directly below what is now the main Tschudi orebody and eventually reached the Mulden contact. The contact itself could also have been the conduit for migration of the fluids. The hydrothermal fluids then spread laterally along the basal units of the Mulden Group and precipitated copper sulphides in the arenites, after coming into contact with hydrogen sulphide from the bacterial action evidenced by the abundance of stromatolites and algal beds (Viviers, unknown).

Although both models are plausible, the diagenetic model is considered more likely. Emslie (1979) concludes that the base metals were associated with brines, originated in sedimentary basins. The marbles and schists of the Swakop Group, which lie to the south of the OML, are considered to be the source of the base metals. These metals were released during diagenesis, were leached into brines which migrated northwards during the higher grade metamorphism of the Swakop Group. Upon reaching the dolomitic rocks of the OML, these brines then reacted with the hydrogen sulphide released by faulting and brecciation, which cased the deposition of the metal sulphides into the dolomitic rocks (Emslie, 1979).

After this structural event, when the dolomite surface had been weathered and karsting took place, the deposition of the Mulden sediments started. Thus the copper already in place in the dolomites, similar to the nearby Alt-Bobos or Karavatu deposits, could act as a source for the Tschudi deposit. A deep seated hydrothermal origin for the mineralization has been suggested, but is now widely dismissed. It is possible there is a main conduit through the dolomite, similar to the Tsumeb ore body, but all evidence goes against this, as all of the dolomite mineralization has been shown to weaken or disappear with depth.



6.7 Zone Surfaces

Due to the poorly developed nature of the supergene mineralization and extremely fine nature of the sulphide grains a single mineralization facies surface was modelled for the base of oxides/top of hypogene. Previous workers had modelled oxide, supergene and sulphide surfaces for the deposit, but these proved highly unreliable in the underground mining and were discarded.

Surface definition was based on mineralogical observations in drill core where holes intersected the surface within mineralization and by sulphuric acid soluble copper data. The historical data was not analysed for this, but all samples from the recent campaign were. A value of “>50% of total copper sulphuric acid soluble” was taken to represent oxide mineral dominance. Where holes penetrated this surface away from mineralization the base of weathering was used.

This process yielded a fairly uniform surface with a mean elevation of approximately 1,230m amsl, roughly 65m below surface (Figure 6.7_1). Localised irregularities do occur, presumably indicating areas of fracturing/faulting with deeper groundwater penetration. There is a zone of slight weathering along the contact well below this surface, where the contact has acted as an aquifer for groundwater, but this has not been considered in the weathering surface.

While the 50% of Cu sulphuric acid soluble figure has proven a reliable guide on other southern African deposits in Coffey Mining’s experience, a more metallurgically effective surface should be derived in future in conjunction with metallurgical and mineralogical test work.



Figure 6.7_1

Longitudinal Projection, Showing Base of Weathering and Grade Shells



6.8 Distribution and Geometry

In order to model the mineralization sectional interpretation was completed on each of the 60 drill section lines and >0.15% Cu grade shells digitised, clipped to the drillhole traces. A Cu grade of 0.15% was used as a natural cut-off after reviewing numerous drillholes and the sample assay value distribution. In many cases the fall off in mineralization is fairly sharp and not much difference would be attained between using 0.25% and 0.15% Cu cuts. In order to gain continuity some <0.15% samples have been included within the shells where they are bound by >0.15% samples. Sectional interpretations were extended to some 25m beyond the deepest holes, 400 - 500m below surface.

Silver broadly tracks copper (elevated silver values do not occur without elevated copper values) and so the grade shells are considered valid for both copper and silver. Sectional grade shell interpretations excluded all mineralization within the Otavi dolomite as this was considered too erratic to model with any confidence. In all cases the basal grade shells were modelled with the Otavi/Mulden contact as their lower edge, even if (unusually) this initial portion ran <0.15% Cu. This was done as it was felt likely that >0.15% Cu mineralisation could occur within this zone immediately adjacent to the drill hole and the contact forms the obvious lower contact to underground stopes and open pits.

The sectional interpretations were then wireframed into solids in three dimensions. The resulting solids were then cut in long section and plan section and reviewed for continuity and crossovers. Corrections were made and the grade shells re-wireframed. The ends of the wireframes were then terminated using the guidelines of half the distance between holes along strike where they terminated or projected one drill line spacing along strike where they were open. Finally shells were cropped to the surveyed surface topography.

Significant mineralization is confined to the zone directly above the Otavi-Mulden contact. As discussed previously the mineralization is not completely confined to the arenite and does occur in the adjacent dolomite, but these occurrences are minor and highly erratic and have not been considered in the grade shell modelling. The majority of the mineralization is confined to a robust zone directly above the contact that extends throughout the 2.5km strike length and 925m down dip length of the deposit modelled (Figures 6.7_1 and 6.8_1 to 6.8_5).

The true width of mineralization in this Lower Zone varies from 1 - 45m with large areas of 10 - 15m thick mineralization running roughly parallel to strike at 80 - 150m below surface. Mineralization generally narrows and decreases in tenor towards surface and down dip. It is open to depth and along strike although there is a general fall in tenor and width in both directions.



Figure 6.8_1

Horizontal Section at the 1,270m Rl, 28m below Surface, Showing Simplified Geology and Grade Shells



Figure 6.8_2

Horizontal Section at the 1,200m Rl, 98m below Surface, Showing Simplified Geology and Grade Shells



Figure 6.8_3 Horizontal Section at the 1,090m Rl, 208m below Surface, Showing Simplified Geology and Grade Shells



Figure 6.8_4 Section E250, Showing Simplified Geology and Grade Shells



Figure 6.8_5

Section E1800, Showing Simplified Geology and Grade Shells



Varying from 0-15m above the Lower Zone mineralization, 14 discontinuous lenses have been defined constituting the Upper Zone (Figures 6.7_1 and 6.8_1 to 6.8_6). These attain a maximum of approximately 25m thick, more commonly 2 - 5m, and plunge shallowly to the northwest along apparent dip of the strata at 20° -25°. Plan extent is typically 450m down plunge (varying from 250 - 700m) by 80m wide (varying from 35 - 200m).

As with the Lower Zone, mineralization is generally poorly developed and narrow near surface. The Upper Zone lenses all appear to terminate with depth, the deepest intersection being some 225m below surface. The lenses may be locally stacked providing near continuous mineralisation from the Otavi contact for significant distances. However, they separate from one another and the Lower Zone mineralization laterally.

Given the geometries evident in the Upper Zone grade shells some of the thicker areas in the Lower Zone may well have a similar plunge. The current density of information is too low to model this reliably, however.



Figure 6.8_6

Isometric View Looking Southwest, Showing Topographic Surface and Grade Shells



7 EXPLORATION

7.1 Previous Exploration

The Tschudi area has undergone several phases of exploration since its discovery in 1913. A 1920 report noted the presence of malachite mineralization in the sandstone, but the first definite exploration was in 1948, when the main outcrop area was trenched and sampled by TCL geologist A. T. Griffis. Griffis recommended diamond drilling, but none was carried out; instead regional mapping of the area was carried out in 1950. Nothing then happened until 1968 when TCL carried out a regional soil geochemistry sampling program, which extended over the entire Tschudi area (Murphy, 1980). The results from this survey prompted further close spaced geochemical sampling, and ultimately drilling programmes, listed below:-

1978: 85 wagon and 31 diamond drillholes proved copper mineralization at the base of Mulden sandstones.

1978-1982: 75 diamond drillholes were drilled, and a “geological resource” of 20 Mt was declared. Metallurgical testwork was carried out on drill core to investigate mineralogical characteristics and recoverability of the sulphide ore.

1979 - A 26m deep exploration shaft was sunk to obtain bulk samples of oxide ore.

1980 - Decline shaft from surface with crosscuts into oxide ore at 40 and 50m levels.

1984 - RC drilling over selected oxide areas.

1990 - 12 diamond drillholes in oxide zone.

1991-1992 - Detailed diamond drill program on a 25m grid in the area where underground mining was planned, to develop a proven resource.

1991 - Exploration development – oxide ore from five cross cuts used as a trial run in the Tsumeb smelter.

1992-1994 - Diamond drilling to prove oxide mineralization for a planned open pit.

1996-1997 - 12 infill drillholes for resource estimate.

1997 - Mineral resource estimate and geological model by Geotech Africa, updated 2002.



7.2 Exploration Potential

The bulk of the Otavi-Mulden contact throughout the Otavi Mountain Land has been geochemically sampled by previous workers. To date Tschudi is the only significant mineralization associated with this contact that has been discovered. While Coffey Mining has not reviewed all historic data on exploration throughout the region it appears that there are several geochemical anomalies along the southern edge of the Tschudi syncline that have not been tested by drilling (Figure 7.2_1).

The mineralization intersected in drilling to date at Tschudi remains open to depth and to the southwest and northeast. However mineralization to the northeast of the deposit becomes increasingly low grade and narrow. Increased drill density at depth may well elucidate additional Upper Zone mineralized lenses.

A recurring theme throughout historic reports on the project is the potential for a Tsumeb style mineralized feeder pipe to the deposit. Such a zone has not been intersected to date, where significant lengths of footwall mineralization has been intersected in vertical holes, follow up inclined holes have shown it to be hosted by narrow sub-vertical stringers (see Section 6.5 for more detail). Intersecting such a zone when drilling to define Mulden hosted mineralization would be a matter of chance. A concerted program to search for such a body would require geophysical techniques including gravity.



Figure 7.2_1

Regional Cu Geochemical Soil Sampling Map



8 DRILLING

The drillholes used to define the Tschudi orebody have been completed in four main exploration phases, by TCL and Weatherly.

8.1 Drilling by Previous Owners

Drilling of the Tschudi Project area was conducted by TCL during three periods, from 1978-1982, 1990-1993, and 1996-1997. This included some 30,000m of diamond and percussion drilling. The first phase drilling was wide spaced (approximately 100m grid) to test the original soil anomaly, as well as a few drillholes to determine the extent of the mineralization to depth. This stage of drilling included both the sulphide and oxide zones, and the majority of the drillholes were vertical. Once the orientation of mineralization was better understood the holes were angled at -70° to better intersect the orebody. The second phase of drilling was mainly focused on the oxide zone, with the grid spacing dropping to 25m over the area where the decline shaft was sunk and initial mining commenced. The western extension of the orebody was drilled on a 50 x 100m grid to prove up oxide mineralization for a possible open pit. The majority of these drillholes were drilled vertically, barring nine drillholes in the shallow south-west corner of the drilling block. The third phase of diamond drilling was undertaken in 1996 in order to increase confidence in the mineral resource estimate. This comprised some infill drilling in the oxide block, as well as drillholes along the fringe of the deposit to better understand the extents of the orebody (Figure 8.1_1).



Figure 8.1_1

Historical Drilling Campaigns



8.2 Drilling by Current Owners

8.2.1 Introduction

Weatherly began an extensive drilling program in April 2007, totalling 26,056m from 200 completed drillholes. The majority of the drilling was as infill to the historical information, in order to increase confidence in the mineral resource estimate, as well as further define the orebody in the planned underground mining areas. The drilling programme was designed by Coffey Mining, based on the historical drilling and other information, using Micromine. Collar positions were placed in the field by the mine surveyors. The drilling was undertaken in four main phases, focusing on separate target areas within the orebody, displayed in Figure 8.2.1_1. These targets were:-

Oxide Block: 4,101m from 48 drillholes were drilled by Reverse Circulation (RC) percussion drill, for definition of shallow oxide/transitional mineralization near the current underground workings. These drillholes were all drilled at -60° to the south-east (151°), to intersect the orebody perpendicularly.

Sulphide Block: 73 drillholes were drilled, totalling 11,257m, for resource definition of the deeper sulphide mineralization for the proposed extension of underground mining. These drillholes were all drilled at -65° to the south-east (151°), to intersect the orebody perpendicularly. The majority of these drillholes (54) were drilled entirely by RC. Only one drillhole was drilled entirely by Diamond Drill (DD). Due to the increased depth and thickness of the arenite hangingwall, the remainder of these holes were pre-collared by RC through the unmineralized areas, and completed by DD through the orezone (18 drillholes).

Underground: The underground campaign was drilled from an access drive off the current underground workings. These holes were designed to provide further confidence in the geometry of the target mineralization, as well as some cover drilling for water ahead of the mining face. Only five holes were drilled to the planned depth. Eight other drillholes were started, but due to bad ground conditions and extremely poor contractor performance were abandoned. Due to this poor performance, the underground campaign was cancelled.



Figure 8.2.1_1

2007 – 2008 Drillhole Collar Plan



Open Pit: 71 drillholes, totalling 10,124m, were drilled as infill to the historical drilling, along the western strike extension, to increase confidence in the mineral resource estimate for the design of the proposed open pit. All of these holes were drilled vertically, by RC. It was planned to drill additional drillholes (41 holes, 6,231m) as pre-collared diamond holes to provide detailed mineralogy down dip and for geotechnical logging and testwork, but the programme was cancelled before this could be achieved. To this end, and additional 450m was drilled by RC, but these holes did not reach the orezone, and were abandoned before they could be completed with DD.

Twin Program: A twinning program was planned in order to validate the historical information. The programme was cancelled before many of these could be drilled, but 3 drillholes were completed, totalling 278m. These drillholes were planned to be at the same azimuth and dip as the historical holes.

The breakdown of metres for each phase of drilling is represented in Table 8.2.1_1.

All drilling was undertaken by drilling contractors (Murray and Roberts Cementation, Kalahari and Hard Rock), to industry standard. For the RC drilling, recoveries were estimated to be above 95%. In some cases the drillholes were stopped due to poor recovery or hole blockage, resulting in the re-drilling of five holes in the sulphide block, and five in the open pit block. For the diamond core, the recoveries were poor, especially in the weathered oxide zone, with an overall average of 88.4%. Problems were often encountered at the sandstone-dolomite contact, where the karst features created irregularities in the lithological succession. When drilling in the deeper, unweathered sulphide zones, the recoveries increased to approximately 94%.

The underground drilling was ineffective, with core recoveries commonly below 80%. The underground drillholes were used for geological modelling, but were removed from the database before resource modelling, due to the low core recovery.

Downhole surveys were completed for 120 of the 195 drillholes drilled from surface. The pre-collared drillholes were surveyed before and after the diamond drilling. In some cases there was extreme deviation of the pre-collared portions, causing these holes to be abandoned and re-drilled. The underground drillholes were not surveyed.



Table 8.2.1_1Tschudi Project

Summary of 2007-2008 Drilling Campaign

Phase

Type of Drilling

No of Holes

Completed

Depth

Dip

RC PC DD

Total m Completed Suffix Mean Min Max

m Completed

m Completed

m Completed

Oxide block TSO RC 48 85 40 132 -60 4,101 4,101

Sulphide block TS PD/RC 73 154 120 215 -65 8,183 1,723 1,351 11257

Underground TSU DD 5 59 33 141 297 297

Open Pit RC RC 71 143 72 228 -90 10,124 450 10,124

Twinning TAP PD 3 93 80.3 102 -60 to -90 90 193 278

Total 200 107 33 228 22,408 2,263 1,841 26,057

RC = Reverse Circulation

DD = Diamond Drilling

PD = Percussion Pre-Collars and Diamond Drill Tails


Tschudi Copper Deposit Geological Modelling and Mineral Resource Estimate Page: 23 9 November

8.2.2 Pre-Collaring

With the exception of drillholes drilled from underground, all drilling was completed from the hangingwall. Commonly there was an arenite hangingwall portion ranging from 10 - 180m that had to be drilled prior to intersecting the mineralized zone. For expediency and cost saving pre-collar drilling was carried out on the planned diamond drill holes.

The pre-collar holes were drilled using open hole RC percussion drilling. Initially chips representing the whole of each 1m interval were collected in a large plastic bag from the cyclone. Once the project was well underway, and confidence was attained in modelling mineralization, it was decided that to cut costs, only a portion of the sample would be collected. Thus, samples representing each metre were collected from the outside return on a shovel, and placed neatly in rows (Figure 8.2.2_1). A representative portion of each metre was then wet sieved by a geologist to collect the coarse chips, which were then logged and stored in chip trays.

Figure 8.2.2_1 Pre-collar chips (piles) and RC Samples (bags) at Rig Awaiting Logging

8.2.3 Reverse Circulation Drilling

The RC program was drilled using 4.5” diameter equipment (Schramm truck and track mounted) (Figure 8.2.3_1). The entire sample was bagged at the rig from the cyclone, no in built splitters on the cyclones were utilised. The drillers were supplied with pre-determined sample depths, and sample bags with sample numbers allocated to them. These sample bags were stored at the rig for logging (Figure 8.2.2_1).



Figure 8.2.3_1 RC Drill Rigs Utilised during 2007-2008 Drilling



8.2.4 Diamond Drilling

The diamond drilling program was drilled using NQ diameter (47.6mm) core, after one to two metres of HQ diameter (63.5mm) was used at the start of the hole either at surface or at the end of the pre-collar. For diamond tails of pre-collars plastic casing was inserted for 10 - 12m in the weathered surface zone. Core was returned by standard practice using a core tube retrieved by a steel wire line attached to a winch. The majority of the programme was carried out with single tube, but in some difficult ground triple tube drilling was used to try and increase core recovery.

Each 3m core run when retrieved was cleaned and carefully inserted into a metal core box with the end of run depth recorded on plastic blocks with permanent UV resistant markers. These core bocks were accurately placed between the bottom of one core run and the top of the subsequent run. A core block was also placed after every failed run. The drillhole number and box number were recorded on the end of the box, and the core was then transported to the logging shed.

For the underground programme, NQ diameter was also used, after a TBW core diameter reaming hole was started. This reaming hole was grouted with cement before insertion of the NQ drill string. Core was returned manually by removing the entire drill string and then the core barrel. Metre marks were written on the core with permanent marker every metre, as well as placing a yellow core block after every run.

8.2.5 Down Hole Surveys

Initial rig setups were completed using a sighting compass for azimuth, inclinometer on the drill head for inclination and a spirit level to ensure that the rig bed was horizontal. Down hole surveys were undertaken on most complete holes to record the azimuth and inclination at depth, every 6 to 30m, depending on the instrument used (Reflex Multi-shot or the Reflex EZ Single-shot respectively).

Magnetic azimuth, inclination, magnetic field strength and down hole temperature were recorded by the driller at each survey station electronically. These were submitted to the geologist at the completion of every survey. Once downloaded from the survey tool, the magnetic azimuth was corrected to a true azimuth using a 10°W declination and entered into the database. Once in the database an accurate spatial plot of each hole in three dimensions was attained using Micromine software.

Underground drillholes were not surveyed.



8.2.6 Geotechnical Core Logging

Upon arrival at the core shed the core was matched up, oriented and a median line marked along the core. An attempt was made to keep the median line along the ‘true’ top of the core. Recoveries and Rock Quality Designation (RQD) were then measured and calculated prior to lithological work. These helped to ensure that all driller’s depth marking discrepancies and inverted core runs were corrected prior to logging and sampling.

Core recovery percentage and RQD was completed on all diamond core using standard industry practice. Values were captured onto handwritten sheets, which were then entered into a formatted spreadsheet for upload into the database. All handwritten logs were filed and retained, along with a printed copy from the database.

Recoveries were continually monitored. Contractually the drillers were obliged to redrill intersections at their expense where mineralized zone recovery fell below 95%. In practice recoveries of less than 95% were accepted at times where extensive karst features were developed, or ‘soft’ zones were present due to weathering or faulting.

8.2.7 Geological Core Logging

Core was logged wet by experienced geologists using a simple and consistent lithological code system. Logging was completed using handwritten sheets with each geological unit delineated and labelled with a four letter geological code and then a short description of the rock, followed by quantification of various parameters such as alteration, vein intensity, and mineralization. An example of the logging template and the lithological coding parameters can be found in Appendix 1. The handwritten logs were then entered into a formatted spreadsheet and uploaded to Micromine. All handwritten logs were transcribed electronically, filed and retained. Once logging was complete the geologist marked the core for sampling and then the core was photographed wet (Figure 8.2.7_1 and 2), two trays at a time, with all pictures entered into the database.



Figure 8.2.7_1

Core Logging and Storage on Site

Figure 8.2.7_2 Example of Core Marked for Sampling, Photographed for the Record Prior to Cutting



8.2.8 Geological Chip Logging

A representative sample of chips from each metre were removed from the large chip bag (or the pile on the floor for pre-collars) and were then sieved and washed. A geologist logged the chips in detail, using an identical coding system to that used for the core, on a similar logsheet. This logged portion was placed in a chip tray to be stored for future reference (Figure 8.2.8_1).

Figure 8.2.8_1 Drill Chip Storage on Site



8.2.9 Drillhole Location and Topographical Survey

Initial drillhole collar positions were planned in Micromine, and placed in the field by the mine surveyors. A hand held Geographic Positioning System (GPS) was used to initially identify the area for clearing. Once the drillholes were complete, the collar position was then surveyed again, by the mine survey team (Total Station) and an accredited contract surveyor (E. Schwarting) using a differential GPS.

All of the historical collars that could be located were also surveyed by Mr Schwarting. These points in conjunction with the new surveyed collars and toe and crest surveys of the old pits using differential GPS were used to create a surface DTM.

All survey data uses the Namibian Lo17 system.

8.2.10 Data Storage and Management

All geological, geotechnical and sampling data was recorded on handwritten sheets initially, and then entered into Excel. Survey data was collected electronically, and then transferred into Excel format. These data were then imported into Micromine on site. Only senior expatriate geologists, and the database manager, were responsible for making changes to the data in Micromine, and validation checks were run regularly. All changes to the data were then made on both Micromine and the relevant Excel spreadsheets. Backups were routinely conducted, using an external hard drive. In addition, backups were made onto DVD and transferred to the Coffey Mining server in Johannesburg. Hardcopy data and original handwritten logs and mapping information were filed and retained on site.



9 SAMPLING METHOD AND APPROACH

9.1 Core Sampling

Industry standard core sampling methods were employed and each geologist was responsible for sampling the drillholes they logged, ensuring that all sampling procedures from start to finish were consistently being adhered to.

Once the geologist completed logging a drillhole, the sample intervals over mineralized zones were marked out, using the median line marked on the core. The median line extended through locked core as far as possible to ensure that the same half of core was sampled consistently.

Where possible lithological intervals were not transgressed by samples, but in the case of low angle contacts to mineralized zones some waste material was included in order to include all the mineralized material. Samples were taken continuously from three metres above the start of visible mineralization, all the way to the end of hole. Sample widths varied from 5cm to 8.4m in extreme cases, with a median at 0.79m. In areas of high core loss, especially in the underground drillholes, very long samples were taken in order to attain minimum sample weight. In areas of extremely high mineralization, or thin lithologies, very short samples were taken to give definition. Outside of these unusual cases sample lengths varied between 0.5m and 1.5m.

Cross cuts were marked on the core where natural breaks were not utilised, and the sample start and end depths were written on the core, on both sides of the median line. Sample numbers were written, with permanent paint markers, on both sides of numerous pieces of core within each sample (Figure 8.2.7_2). Sample tickets were obtained from Weatherly, and these dictated the numbering system used. Runs of sample numbers were entered into Excel and Quality Control and Quality Assurance (QA/QC) standards, blanks and duplicate positions assigned for the entire run.

The core was then cut in half using a core-cutter, supplied by Corstor. Once cut, the appropriate side of each sample was then placed into plastic sample bags with a sample ticket and the sample number written on the outside with permanent paint marker. The bags were laid out in order on the floor of the coreshed, to allow the insertion of standards, blanks and empty bags for the re-splits. Once QA/QC samples had been inserted the sample bags were closed using a stapler, and transported to the assay laboratory.

The sample intervals and numbers were then marked on the flat face of the half core remaining in the tray so that the assay results could be easily referred back to the core.



9.2 RC Chip Sampling

Industry standard chip sampling methods were employed throughout the programme. Chip samples were taken by the driller at the rig. The drillers were supplied with numbered sample bags, the numbering system used identical to that of the core sampling. The samples were taken directly from the cyclone.

The driller was supplied with sampling depths before the drillhole was started. These depths were determined using Micromine, where a point 15m above the top of the expected mineralized zone was chosen as the start of the sampling. QA/QC sample positions were determined, the sample numbers written on the bags, and the QA/QC sample bags removed. The sample bags were then supplied to the driller, along with the initial sampling depth. Every metre was then sampled to the end of hole. Sample bags were stored in rows at the rig, before being transported to the core shed. Sample masses were spot checked regularly and several holes were re-drilled where recovery was poor.

The large bags of drill chips were re-split manually to yield an approximately 2kg sub sample for analysis. This was accomplished using a single-tier riffle splitter which was cleaned thoroughly between samples. The samples were split approximately 4 - 5 times, until a small enough representative sample was obtained. An identical sample numbering procedure as that used for diamond drill core samples was applied to these samples.

The sample bags were then laid out in order, to allow the insertion of the previously removed QA/QC sample bags. Once these had been re-inserted the sample bags were closed with a stapler, and transported to the assay laboratory.

9.3 Relative Densities

Relative density measurements were taken on whole pieces of diamond core for each assay interval within a drillhole where possible. Measurements were taken on ±15cm long clean, solid, core billets using an electronic scale accurate to 0.1g. The mass of the dry core billet was measured, then the mass of the core billet suspended in water in the cage below the scale was measured.

The relative density (RD) was calculated by the formula:-

RD = Md / (Md – Mw),

Where Md = weight in air and Mw = weight in water.

All readings were recorded on paper by the geotechnician and then entered digitally into an Excel spreadsheet. Handwritten records were filed and retained. These data were then loaded into Micromine and compared against rock type, weathering and sulphide content and any anomalous determinations re-done.



In order to estimate in-situ bulk density using relative density measurements the material to be measured must be non-porous. Samples with a high porosity due to weathering or abundant vugs were sealed with a clear lacquer prior to immersion. The bulk density of broken core fragments cannot be accurately measured by this technique and where samples were fragmented this was recorded and the results excluded from the database.

9.4 Sample Preparation and Analysis

All sample preparation and analysis for all drilling completed at Tschudi was undertaken by Weatherly’s production laboratory in Tsumeb. According to long serving laboratory staff members the sample preparation and analysis methodology has remained little changed since the mid 1970’s. As the primary assay laboratory was not an external independent contractor Coffey Mining visited the laboratory to review procedures and standards on a regular basis. The following summary of procedures is that recorded by Coffey Mining staff.

9.4.1 Sample Preparation

Sample Tracking

Upon receipt of the samples from the field, the laboratory staff compares the received samples to the dispatch form that accompanies them. The dispatch form contains information on the number of samples, unique sample numbers and the analysis requirements. If there are any discrepancies picked up the geologist is notified. The sample ticket placed in each sample bag is removed from the bag and transported with the sample material throughout the entire process of preparation.

Crushing and Pulverising

Once the samples have been checked and admitted to the laboratory the samples that require crushing are dried for at least 6 hours in a drying oven at 110° - 120°C. The coarser rock samples are put through an oscillating jaw crusher (Figure 9.4.1_1), and then a finer rock crusher. The resultant material is typically 95% passing -1mm. Some 300g of this material is separated off and sent for pulverisation. The coarse reject is stored separately in the original plastic sample bags. The 300g of crushed material is transferred to a brown paper bag, along with the original sample ticket and a laboratory unique sample number that is related to the original sample number. The jaw crusher is cleaned with barren quartz better than every 10th sample, blown down with compressed air after each sample.

The sample is sent for pulverisation, using a ring pulveriser (Figure 9.4.1_2). During the pulverization of a batch of samples, a sample of clean silica sand is put through the pulveriser every ten samples, to ensure minimum contamination. If the material being sampled is a dark colour (signifying finely ground sulphides), this is done more regularly. Once pulverised the material is returned to the brown paper bag, and sent for analysis. Pulveriser plots are blown down and dusted in a dust hood between samples.



Figure 9.4.1_1 Tsumeb Laboratory, Jaw Crusher

Figure 9.4.1_2 Tsumeb Laboratory, Ring Pulveriser



9.4.2 Analysis

For the analysis 0.1g of pulverised material is taken from the approximately 300g of pulverised material using a spatula. This is mixed into a solution of 7:3 concentrated, hydrochloric/concentrated nitric acid, which is then heated for 3 hours at 180°C. Then 15ml of 40% hydrochloric acid is added and the material shaken in an agitator. The solution is left to settle overnight, and then analysed for Cu and Ag with an Atomic Absorption Spectrometer (AAS), a PerkinElmer Analyst 200. A similar solution process using only sulphuric acid was utilised separately on each sample and analyses by AAS to yield acid soluble Cu.

Detection limits were 10ppm for Cu and 1ppm for Ag.

The assay techniques utilised are considered appropriate to the deposit and meet industry standards.

Figure 9.4.2_1 Tsumeb Laboratory, AAS Room



9.5 Analytical Quality Control Data

9.5.1 Introduction

The application of a QA/QC procedure is used to ensure that the assay results of an exploration programme can be confidently relied upon. The procedure requires the introduction of sufficient blanks, standards and duplicates with the samples to ensure that the procedures of the laboratory are not introducing a bias to the results.

The following aspects of the procedure may be checked:-

Sample management – transposition of samples, reported missing or added samples.

Sample preparation – contamination of one sample to another, sample mix ups, sample homogeneity.

Calibration of instrumentation – accuracy.

Repeatability of analysis – precision of analysis.

Of the 14,087 samples submitted to the Tsumeb laboratory during the program, 2,885 of these were QA/QC samples, representing 20.4% of the total.

9.5.2 Definitions

A standard is a reference sample with a known (statistically) element abundance and standard deviation. Reference standards are used to gauge the accuracy of analytical reporting by comparing the pre-determined values to those reported by the laboratory during an exploration project.

A blank is a standard with abundance of the element of interest below the level of detection of the analytical technique.

A duplicate is a second nominally identical sample taken at a particular stage of the sampling process; e.g. Field Duplicate.

9.5.3 QA/QC Programme

Blanks

Blanks (washed silica sand) were introduced with each batch submitted to the laboratory to monitor contamination in the crushing and pulverising stages. Some 100g of blank material was supplied for each blank sample included in the sample batch.

The blanks were introduced at a frequency of 1 in 20 (5%).



Standards

The accuracy of laboratory results during the drilling/sampling programme was monitored with the use of four commercial standards from Geostats (Pty) Ltd, Perth, Australia (Table 9.5.3_1). Some 10g of standard material was supplied for each standard sample included in the sample batch. The standards were not crushed or pulverised as they were sufficiently fine grained. In addition the laboratory introduced their own standards for internal QA/QC monitoring.

Standard material samples were introduced at a frequency of 1 in 10 (10%) or greater, with alternating standards used.

Table 9.5.3_1 Tschudi Project

Expected Values for Geostats Standards Used

Standard Element Grade (ppm) Standard Deviation

No of Analyses

Confidence Interval

GBM306-16 Copper 13,409 355 45 ± 104

Silver Not certified

GBM996-3 Copper 22,599 1,436 60 ± 363.4

Silver 44.2 4.0 52 ± 1.09

GBM905-11 Copper 31,758 1,017 47 ± 291

Silver Not certified

GBM303-2 Copper 72,921 2,847 65 ± 692

Silver 26.1 2.8 59 ± 0.72

Duplicates

Duplicates of the diamond core samples were generated from the coarse rejects by the laboratory. A designated sample was crushed and riffle split to provide a field duplicate rather than resubmitting duplicates from previous sample batches. This was deemed to be the most practical method of providing duplicates due to the volume of samples being submitted.

Duplicates of the RC chip samples were generated in the field by producing a second riffle split sample from the large sample bag.

Duplicates were introduced at a frequency of 1 in 20 (5%) or greater.



Inter-Laboratory Analyses (Referee Checks)

Some 6.5% of the drillhole pulps (including QA/QC samples) were submitted to an independent accredited laboratory (ALS Chemex, Johannesburg, South Africa) for referee analysis. Referee analysis was undertaken by four acid digestion with Inductively Coupled Plasma Atomic Emission Spectrometry (ICP-AES) finish for 33 elements (ALS method code ME-ICP61a). Detection limits were 1ppm for Ag and 10ppm for Cu.

9.5.4 Quality Control - Monitoring

The analytical QA/QC data was analysed on an on-going basis and several queries raised with the laboratory. Where significant deviation from the standards expected value was encountered, re-assaying of at least 5 samples on either side of the standard was undertaken to ensure that there was no bias in the surrounding samples.

All assay results were checked against the core and stored chips to ensure that the assay values represented observed mineralization. Where serious discrepancy was noted between mineralization and assay values, samples were quarter split and re-assayed and coarse rejects of the original sample re-assayed in order to track problems.



9.5.5 Methodology

The precision and accuracy will be discussed in terms of the following statistical measures routinely applied by Coffey Mining:-

Thompson and Howarth Plot showing the mean relative percentage error of grouped assay pairs across the entire grade range, used to visualise precision levels by comparing against given control lines.

Rank HARD Plot, which ranks all assay pairs in terms of precision levels measured as half of the absolute relative difference from the mean of the assay pairs (HARD), used to visualise relative precision levels and to determine the percentage of the assay pairs population occurring at a certain precision level.

Mean vs HARD Plot, used as another way of illustrating relative precision levels by showing the range of HARD over the grade range.

Mean vs HRD Plot is similar to the above, but the sign is retained, thus allowing negative or positive differences to be computed. This plot gives an overall impression of precision and also shows whether or not there is significant bias between the assay pairs by illustrating the mean percent half relative difference between the assay pairs (mean HRD).

Correlation Plot is a simple plot of the value of assay 1 against assay 2. This plot allows an overall visualisation of precision and bias over selected grade ranges. Correlation coefficients are also used.

Quantile-Quantile (Q-Q) Plot is a means where the marginal distributions of two datasets can be compared. Similar distributions should be noted if the data is unbiased.

9.5.6 Blanks

Minor anomalous Cu grades up to 0.05% (Figure 9.5.6_1) and Ag grades up to 2g/t (Figure 9.5.6_2) occur sporadically. At the start of the project a contaminated bag of blank material was used, reflecting Cu grades up to 0.5% and Ag grades up to 2 g/t. Once this was highlighted, commercial quality silica sand was obtained for the blank material. A single anomalous Ag value was reported after this, which can be ascribed to contamination in sample prep, as the previous sample recorded extremely high Ag values.

9.5.7 Standard GBM 306-16

The standard shows good precision and accuracy for Cu with a minor positive bias of 2.7% (Figure 9.5.7_1). The standard is not certified for Ag.




The standard shows good precision and accuracy for Cu with a minor negative bias of 0.85% (Figure 9.5.8_1). The standard shows good precision and accuracy for Ag with a minor negative bias of 1.1% (Figure 9.5.8_2). A slight correction in AAS machine calibration after the analysis of initial results is evident.


The standard shows good precision and accuracy for Cu with a minor negative bias of 0.29%. The standard is not certified for Ag (Figure 9.5.9_1).


The standard shows good precision and accuracy for Cu with a minor positive bias of 0.3% (Figure 9.5.10_1). The standard shows good precision and accuracy for Ag with a minor positive bias of 2.8% (Figure 9.5.10_2).

9.5.11 Field Duplicates

The analyses of field duplicates demonstrates precision in the assay technique for both Cu (Figure 9.5.11_1) and Ag (Figure 9.5.11_2). Both data sets show 91% of the data pairs being within 20% (HARD) precision. This includes data assaying below ten times detection limit for both elements.

9.5.12 Referee Analysis

A total of 727 sample pulps (6.5% of total drillhole samples) were submitted to ALS Chemex for analysis.

The comparative results for Cu show very good correlation with a -4% bias on the Tsumeb results relative to ALS Chemex (Figure 9.5.12_1).

Comparison of all Ag data (Figure 9.5.12_2) shows reasonable correlation, but the results are skewed by the preponderance of data near detection limit. Comparison of data for samples grading better than ten times detection limit (10ppm) shows good correlation with 7% bias in the Tsumeb data (Figure 9.5.12_3).



9.5.13 Chain of Custody – Responsibility and Accountability

A full chain of custody was implemented for the sample submission by the geologist to the analytical laboratory.

The details of the samples to be submitted were recorded on standard documentation on site. The samples were checked by sampling personnel and the supervising geologist prior to dispatch. The samples were transported by the geologist from the site to the laboratory directly.

The results were emailed to the senior geologist on site, in the form of Excel spreadsheets, as well as a hardcopy page of the results being delivered to the geologist. Cross checking of the assay certificates with the results was possible as these included details of each batch.

9.5.14 Conclusions

The analysis of analytical QA/QC data shows the data to be of sufficient quality to use in a mineral resource estimate.



Figure 9.5.6_1

Analysis of QA/QC: Blank: Cu

Standard: BLANK No of Analyses: 670Element: Cu Minimum: 0.01Units: % Maximum: 0.51Detection Limit: Mean: 0.01Expected Value (EV): 0.03 Std Deviation: 0.03E.V. Range: 0.03 to 0.03 % in Tolerance 0.15 %

% Bias -67.04 %% RSD 302.23 %

0.0

0.1

0.2

0.3

0.4

0.5

0.6

04

-Au

g-2

00

7

26

-No

v-2

00

7

06

-Feb

-20

08

28

-Mar-2

00

8

12

-Jun

-20

08

19

-Au

g-2

00

8C

u %

(%

)

DATE

Standard Control Plot(Standard: BLANK)

Cu % Expected Value = 0.03 EV Range (0.03 to 0.03) Mean of Cu % = 0.01

-1.0

0.0

1.0

2.0

3.0

04

-Au

g-2

00

7

26

-No

v-2

00

7

06

-Feb

-200

8

28

-Ma

r-20

08

12

-Jun

-200

8

19

-Au

g-2

00

8C

um

ula

tiv

e S

um

of

Cu

% -

Mea

n (

%)

DATE

Cumulative Deviation from Assay Mean(Standard: BLANK)

Cu % Mean of Cumulative Sum of Cu % - Mean (%) = 1.45

-15

-10

-5

0

5

04-A

ug

-20

07

26-N

ov-2

007

06-F

eb

-20

08

28-M

ar-2

008

12-Ju

n-2

00

8

19-A

ug

-20

08C

um

ula

tive S

um

of

Cu

% -

Ex

pect

ed V

alu

e (

%)

DATE

Cumulative Deviation from Expected Value(Standard: BLANK)

Cu % Mean of Cumulative Sum of Cu % - Expected Value (%) = -5.30



Figure 9.5.6_2

Analysis of QA/QC: Blank: Ag

Standard: BLANK No of Analyses: 670Element: Ag Minimum: 0.50Units: g/t Maximum: 2.00Detection Limit: Mean: 0.52Expected Value (EV): 3.00 Std Deviation: 0.13E.V. Range: 2.70 to 3.30 % in Tolerance 0.00 %

% Bias -82.69 %% RSD 24.38 %

0.0

1.0

2.0

3.0

4.0

04

-Au

g-2

00

7

26

-No

v-2

00

7

06

-Fe

b-2

00

8

28

-Ma

r-20

08

12

-Jun

-20

08

19

-Au

g-2

00

8A

g g

/t

(g/

t)

DATE

Standard Control Plot(Standard: BLANK)

Ag g/t Expected Value = 3.00 EV Range (2.70 to 3.30) Mean of Ag g/t = 0.52

-202468

1012

04

-Au

g-2

00

7

26

-No

v-2

00

7

06

-Fe

b-2

00

8

28

-Ma

r-20

08

12

-Jun

-20

08

19

-Au

g-2

00

8C

um

ula

tiv

e S

um

of

Ag

g/

t -

Me

an

(g

/t)

DATE

Cumulative Deviation from Assay Mean(Standard: BLANK)

Ag g/t Mean of Cumulative Sum of Ag g/t - Mean (g/t) = 4.93

-2000

-1500

-1000

-500

0

04

-Au

g-2

00

7

26

-No

v-2

00

7

06

-Fe

b-2

00

8

28

-Ma

r-20

08

12

-Jun

-20

08

19

-Au

g-2

00

8Cu

mu

lati

ve S

um

of

Ag

g/

t -

Ex

pecte

d V

alu

e (

g/

t)

DATE

Cumulative Deviation from Expected Value(Standard: BLANK)

Ag g/t Mean of Cumulative Sum of Ag g/t - Expected Value (g/t) = -827.31



Figure 9.5.7_1

Analysis of QA/QC: Standard GBM 306-16: Cu Standard: GBM306-16 No of Analyses: 545Element: Cu % Minimum: 1.310Units: Maximum: 1.430Detection Limit: - Mean: 1.377Expected Value (EV): 1.341 Std Deviation: 0.016E.V. Range: 1.207 to 1.475 % in Tolerance 100.000 %

% Bias 2.699 %% RSD 1.178 %

1.10

1.20

1.30

1.40

TD

01

2

TD

05

6

TD

08

5

TD

11

5

TD

14

0C

u %

(g

/t)

CERT #

Standard Control Plot(Standard: GBM306-16)


-0.2

0.0

0.2

0.4

0.6

0.8

1.0

1.2

TD

01

2

TD

05

6

TD

08

5

TD

11

5

TD

14

0C

um

ula

tiv

e S

um

of

Cu

% -

Me

an

(g

/t)

CERT #

Cumulative Deviation from Assay Mean(Standard: GBM306-16)

Cu % Mean of Cumulative Sum of Cu % - Mean (g/t) = 0.580

0

5

10

15

20

TD

01

2

TD

05

6

TD

08

5

TD

11

5

TD

14

0C

um

ula

tiv

e S

um

of

Cu

% -

Exp

ect

ed

Va

lue (

g/

t)

CERT #

Cumulative Deviation from Expected Value(Standard: GBM306-16)

Cu % Mean of Cumulative Sum of Cu % - Expected Value (g/t) = 10.460



Figure 9.5.8_1


% Bias -0.849 %% RSD 1.380 %

2.0

2.1

2.2

2.3

2.4

TD

01

0

TD

03

6

TD

06

2

TD

09

0

Cu

% (

g/t)

CERT #



-6

-5

-4

-3

-2

-1

0

1

TD

01

0

TD

03

6

TD

06

2

TD

09

0

Cu

mula

tiv

e S

um

of

Cu %

- M

ean

(g

/t)

CERT #


Cu % Mean of Cumulative Sum of Cu % - Mean (g/t) = -2.616

-10

-8

-6

-4

-2

0

TD

01

0

TD

03

6

TD

06

2

TD

09

0

Cu

mu

lati

ve S

um

of

Cu

% -

Exp

ect

ed

Valu

e (

g/t)

CERT #


Cu % Mean of Cumulative Sum of Cu % - Expected Value (g/t) = -7.176



Figure 9.5.8_2

Analysis of QA/QC: Standard GBM 996-3: Ag Standard: GBM996-3 No of Analyses: 474Element: Ag g/t Minimum: 39.0Units: Maximum: 46.0Detection Limit: - Mean: 43.7Expected Value (EV): 44.2 Std Deviation: 1.0E.V. Range: 39.8 to 48.6 % in Tolerance 99.2 %

% Bias -1.1 %% RSD 2.3 %

38

40

42

44

46

48

50

TD

01

0

TD

03

6

TD

06

2

TD

09

0

Ag

g/t

(g/

t)

CERT #



-30

-20

-10

0

10

20

TD

01

0

TD

03

6

TD

06

2

TD

09

0

Cu

mu

lati

ve S

um

of

Ag

g/

t -

Me

an

(g

/t)

CERT #


Ag g/t Mean of Cumulative Sum of Ag g/t - Mean (g/t) = -7.9

-250

-200

-150

-100

-50

0

TD

01

0

TD

03

6

TD

06

2

TD

09

0

Cu

mu

lati

ve S

um

of

Ag

g/

t -

Ex

pecte

d V

alu

e (

g/

t)

CERT #


Ag g/t Mean of Cumulative Sum of Ag g/t - Expected Value (g/t) = -120.0



Figure 9.5.9_1


% Bias -0.294 %% RSD 0.675 %

2.7

2.8

2.9

3.0

3.1

3.2

3.3

3.4

TD

02

7

TD

02

8

TD

02

8

TD

03

0

TD

03

0

TD

03

2

TD

03

4

TD

03

5

TD

03

6

Cu

% (

g/

t)

CERT #



-0.4

-0.3

-0.2

-0.1

0.0

0.1

TD

02

7

TD

02

8

TD

02

8

TD

03

0

TD

03

0

TD

03

2

TD

03

4

TD

03

5

TD

03

6

Cu

mu

lati

ve S

um

of

Cu

% -

Me

an

(g

/t)

CERT #



-1.0

-0.8

-0.6

-0.4

-0.2

0.0

TD

02

7

TD

02

8

TD

02

8

TD

03

0

TD

03

0

TD

03

2

TD

03

4

TD

03

5

TD

03

6

Cu

mu

lati

ve S

um

of

Cu

% -

Exp

ect

ed

Va

lue (

g/

t)

CERT #


Cu % Mean of Cumulative Sum of Cu % - Expected Value (g/t) = -0.482



Figure 9.5.10_1


% Bias 0.296 %% RSD 0.934 %

6.5

7.0

7.5

8.0

8.5

TD

11

3

TD

11

6

TD

11

8

TD

11

9

TD

12

0

TD

12

2

TD

12

5

TD

13

0

TD

13

2

TD

13

5

TD

13

9

TD

14

0

TD

14

3

TD

14

8

TD

14

9

TD

15

2

Cu

% (

g/

t)

CERT #



-2.0

-1.5

-1.0

-0.5

0.0

0.5

TD

11

3

TD

11

6

TD

11

8

TD

11

9

TD

12

0

TD

12

2

TD

12

5

TD

13

0

TD

13

2

TD

13

5

TD

13

9

TD

14

0

TD

14

3

TD

14

8

TD

14

9

TD

15

2

Cu

mu

lati

ve S

um

of

Cu

% -

Me

an

(g

/t)

CERT #



-1

0

1

2

3

4

TD

11

3

TD

11

6

TD

11

8

TD

11

9

TD

12

0

TD

12

2

TD

12

5

TD

13

0

TD

13

2

TD

13

5

TD

13

9

TD

14

0

TD

14

3

TD

14

8

TD

14

9

TD

15

2

Cu

mu

lati

ve S

um

of

Cu

% -

Exp

ect

ed

Va

lue (

g/

t)

CERT #


Cu % Mean of Cumulative Sum of Cu % - Expected Value (g/t) = 0.976



Figure 9.5.10_2

Analysis of QA/QC: Standard GBM 303-2: Ag

Standard: GBM303-2 No of Analyses: 162Element: Ag g/t Minimum: 25.0Units: Maximum: 28.0Detection Limit: - Mean: 26.8Expected Value (EV): 26.1 Std Deviation: 0.7E.V. Range: 23.5 to 28.7 % in Tolerance 100.0 %

% Bias 2.8 %% RSD 2.7 %

23

24

25

26

27

28

29

TD

11

3

TD

11

6

TD

11

8

TD

11

9

TD

12

0

TD

12

2

TD

12

5

TD

13

0

TD

13

2

TD

13

5

TD

13

9

TD

14

0

TD

14

3

TD

14

8

TD

14

9

TD

15

2

Ag

g/t

(g/

t)

CERT #



-10

0

10

20

30T

D1

13

TD

11

6

TD

11

8

TD

11

9

TD

12

0

TD

12

2

TD

12

5

TD

13

0

TD

13

2

TD

13

5

TD

13

9

TD

14

0

TD

14

3

TD

14

8

TD

14

9

TD

15

2

Cu

mu

lati

ve S

um

of

Ag

g/

t -

Me

an

(g

/t)

CERT #


Ag g/t Mean of Cumulative Sum of Ag g/t - Mean (g/t) = 11.3

-20

0

20

40

60

80

100

120

TD

11

3

TD

11

6

TD

11

8

TD

11

9

TD

12

0

TD

12

2

TD

12

5

TD

13

0

TD

13

2

TD

13

5

TD

13

9

TD

14

0

TD

14

3

TD

14

8

TD

14

9

TD

15

2

Cum

ula

tiv

e S

um

of

Ag

g/

t -

Ex

pecte

d V

alu

e (

g/

t)

CERT #


Ag g/t Mean of Cumulative Sum of Ag g/t - Expected Value (g/t) = 70.1



Figure 9.5.11_1

Analysis of QA/QC: Field Duplicates: Cu Cu % RCu % Units Result

No. Pairs: 612 612 Pearson CC: 0.98Minimum: 0.01 0.01 % Spearman CC: 0.99Maximum: 8.20 8.40 % Mean HARD: 6.62Mean: 0.58 0.59 % Median HARD: 1.82Median 0.25 0.24 %Std. Deviation: 0.87 0.89 % Mean HRD: 0.30Coefficient of Variation: 1.49 1.51 Median HRD 0.00

0

50

100

0.001 0.01 0.1 1 10

HA

RD

(%

)

Mean of Data Pair (g/t)

Mean vs. HARD Plot(All Data)

Mean HARD: 6.62 Median HARD: 1.82Precision: 20%

0

20

40

60

80

100

0 10 20 30 40 50 60 70 80 90 100

HA

RD

(%

)Rank (%)

Rank HARD Plot(All Data)

Precision: 20%

90.85% of data are withinPrecision limits

0

20

40

60

-1.0 0.0 1.0

Fre

qu

en

cy (

%)

HRD (/100)

HRD Histogram(All Data)

Mean HRD: 0.30 Median HRD: 0.00Precision: +/-20%

-100

-50

0

50

100

0.001 0.01 0.1 1 10

HR

D (

%)


Mean vs. HRD Plot(All Data)


0.0001

0.001

0.01

0.1

1

10

0.001 0.01 0.1 1 10

Ab

solu

te D

iffe

rence (

%)

Mean of Data Pair (%)

T & H Precision Plot (Assay Pairs)(All Data)

10% 20% 30% 50%

0.0001

0.001

0.01

0.1

1

10

0.001 0.01 0.1 1 10

Me

dia

n A

D (

%)

Grouped Mean of Pair (%)

T & H Precision Plot (Grouped Pairs)(All Data)

10% 20% 30% 50%

0

5

10

0 1 2 3 4 5 6

RC

u %

(%

)

Cu % (g/t)

Correlation Plot(All Data)

P.CC= 0.98 S.CC= 0.99 Ref. Liney = 1.01x + 0.00

-5

0

5

10

0 1 2 3 4 5 6 7 8 9 10

RC

u %

(%

)

Cu % (%)

QQ Plot(All Data)

Ref. Line y = 1.02x -0.01



Figure 9.5.11_2

Analysis of QA/QC: Field Duplicates: Ag Ag g/t RAg g/t Units Result

No. Pairs: 612 612 Pearson CC: 0.98Minimum: 0.01 0.01 g/t Spearman CC: 0.99Maximum: 8.20 8.40 g/t Mean HARD: 6.62Mean: 0.58 0.59 g/t Median HARD: 1.82Median 0.25 0.24 g/tStd. Deviation: 0.87 0.89 g/t Mean HRD: 0.30Coefficient of Variation: 1.49 1.51 Median HRD 0.00

0

50

100

0.001 0.01 0.1 1 10

HA

RD

(%

)




0

20

40

60

80

100

0 10 20 30 40 50 60 70 80 90 100

HA

RD

(%

)

Rank (%)


Precision: 20%


0

20

40

60

-1.0 0.0 1.0

Fre

qu

en

cy (

%)

HRD (/100)



-100

-50

0

50

100

0.001 0.01 0.1 1 10

HR

D (

%)




0.0001

0.001

0.01

0.1

1

10

0.001 0.01 0.1 1 10

Ab

solu

te D

iffe

rence (

g/t)



10% 20% 30% 50%

0.0001

0.001

0.01

0.1

1

10

0.001 0.01 0.1 1 10

Me

dia

n A

D (

g/

t)

Grouped Mean of Pair (g/t)


10% 20% 30% 50%

0

5

10

0 1 2 3 4 5 6

RA

g g

/t

(g/

t)

Cu % (g/t)



-5

0

5

10

0 1 2 3 4 5 6 7 8 9 10

RA

g g

/t

(g/t)

Ag g/t (g/t)

QQ Plot(All Data)

Ref. Line y = 1.02x -0.01



Figure 9.5.12_1

Comparison of Tsumeb Laboratory and ALS Chemex: Cu

T Cu % ALS Cu % Units ResultNo. Pairs: 728 728 Pearson CC: 0.98Minimum: 0.00 0.00 % Spearman CC: 0.96Maximum: 9.40 9.87 % Mean HARD: 5.63Mean: 1.14 1.10 % Median HARD: 2.59Median 0.81 0.80 %Std. Deviation: 1.14 1.12 % Mean HRD: 2.23Coefficient of Variation: 1.00 1.02 Median HRD 1.16

0

50

100

0.001 0.01 0.1 1 10

HA

RD

(%

)




0

20

40

60

80

100

0 10 20 30 40 50 60 70 80 90 100

HA

RD

(%

)Rank (%)


Precision: 10%


0

20

40

60

-1.0 0.0 1.0

Fre

qu

en

cy (

%)

HRD (/100)



-100

-50

0

50

100

0.001 0.01 0.1 1 10

HR

D (

%)




0.0001

0.001

0.01

0.1

1

10

0.001 0.01 0.1 1 10

Ab

solu

te D

iffe

ren

ce

(%

)



10% 20% 30% 50%

0.0001

0.001

0.01

0.1

1

10

0.001 0.01 0.1 1 10

Me

dia

n A

D (

%)

Grouped Mean of Pair (%)


10% 20% 30% 50%

0

5

10

0 1 2 3 4 5 6 7 8 9 10

ALS

Cu %

(%

)

T Cu % (%)



-5

0

5

10

0 1 2 3 4 5 6 7 8 9 10

ALS

Cu %

(%

)

T Cu % (%)

QQ Plot(All Data)

Ref. Line y = 0.98x -0.02



Figure 9.5.12_2

Comparison of Tsumeb Laboratory and ALS Chemex: Ag T Ag g/t ALS Ag g/t Units Result


0

50

100

0.1 1 10 100 1000

HA

RD

(%

)




0

20

40

60

80

100

0 10 20 30 40 50 60 70 80 90 100

HA

RD

(%

)

Rank (%)


Precision: 10%


0

10

20

30

40

-1.0 0.0 1.0

Fre

qu

en

cy (

%)

HRD (/100)



-100

-50

0

50

100

0.1 1 10 100 1000

HR

D (

%)




0.01

0.1

1

10

100

0.1 1 10 100 1000

Ab

solu

te D

iffe

ren

ce (

g/

t)



10% 20% 30% 50%

0.01

0.1

1

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Figure 9.5.12_3

Comparison of Tsumeb Lab and ALS Chemex: Ag >10 x Detection Limit T Ag g/t ALS Ag g/t Units Result


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10 HISTORIC DATA VERIFICATION

10.1 Twin Program

10.1.1 Twin Drillholes

In order to develop a mineral resource model for the Tschudi Project it was necessary to integrate historic drillhole data. These historic data were conducted in several campaigns by TCL, and are described in Section 7.1. In all cases twin drillholes were drilled within 10m of the historic drillholes, and only drillholes who’s collars could be located in the field were twinned. The twin drillholes were drilled at the same azimunth and inclination as the historic holes.

10.1.2 Twin Results

As a basis for acceptance of the historical data, each of the four twin pairs was compared for lithology/geology, style of mineralization and tenor of mineralization. These comparisons are summarised in Table 10.1.2_1, and downhole logs are shown in Appendix B.

Lithology/Geology

AP55 and TAP55: The Otavi/Mulden contact is defined by a basal conglomerate that is in direct contact with a chert band, followed by and arenite filled cavity, evidenced in both drillholes.

AP95 and TAP95: Siltstone /argillite beds can be correlated in both the original and twin drillholes, although two conglomerate beds are shown in the historic that are not represented in the twin drillhole.

AP115 and TAP115: In the upper horizons there are isolated siltstone and conglomerate bands that correlate between drillholes. At the Otavi/Mulden contact in the original hole the arenite lies on a basal conglomerate followed by oolitic chert then brecciated dolomite. In the twin drillhole there is first a thin chert layer above the conglomerate, as well as a cavity with conglomerate infill not represented in the original drillhole.

AP241 and TAP241: AP241 has a silty/clayey layer above the orezone, which is composed of arenitic sandstone, that is in direct contact with an oolitic chert layer, followed by brecciated/altered dolomite. In the twin drillhole, TAP241, the silty bands are represented, and the ore zone is composed of black arenite, directly in contact with the oolitic chert layer, but this chert contains a conglomerate layer that is not represented in the original. Similarly, there are arenitic cavities in the dolomite of the twin drillhole that are not described in the original.



Mineralization Style

AP55 and TAP55: The mineralization in both drillholes is confined to arenite, cut by siltstone bands and a conglomerate. Zones of high mineralization correspond in both drillholes.

AP95 and TAP95: The mineralization in both drillholes is found with massive arkosic arenite.

AP115 and TAP115: In both drillholes the mineralized zone is capped by a thin conglomerate band, with correlatable zones of higher mineralization.

AP241 and TAP241: The mineralization in both drillholes is found within massive arenite, and zones of high mineralization roughly correspond in both drillholes.

Mineralization Tenor

The comparison of mineralized portions are tabulated in Table 10.1.2_1 and shown graphically in Figures 10.1.2_1 to 5.

Table 10.1.2_1 Tschudi Project

Summary of Twin Drilling Results

Length (m) Cu (%) Ag (g/t) Cu % m Ag g/t m

Mother Twin Mother Twin Mother Twin Mother Twin Mother Twin AP55 18.30 22.78 1.39 0.86 25.54 11.97 25.44 19.59 467.38 446.28AP95 4.30 7.24 1.16 1.35 7.91 15.67 4.99 9.77 34.01 70.76 AP115 14.90 13.95 2.27 2.83 14.36 14.88 33.82 39.48 213.96 550.73 AP241 12.30 12.14 1.59 0.71 23.76 5.50 19.56 8.62 292.25 104.64 Ave 12.45 14.03 1.60 1.44 17.89 12.01 20.95 19.37 251.90 293.10



Figure 10.1.2_1 Twin Drillhole Comparison Mineralized Length

Figure 10.1.2_2 Twin Drillhole Comparison Average Cu%



Figure 10.1.2_3

Twin Drillhole Comparison Average Ag g/t

Figure 10.1.2_4 Twin Drillhole Comparison Cu % m



Figure 10.1.2_5

Twin Drillhole Comparison Ag g/t m

10.1.3 Conclusion

In all cases lithology could be correlated between the pairs within the parameters of the geological modelling.

The style of mineralization was found to be similar in all pairs with mineralization dominantly confined to arenites or arkosic arenite. Mineralization consisted of bornite/chalcopyrite in the hypogene zones and assorted copper oxides, silicates and carbonates in the oxide zones.

Correlation of tenor of mineralization was undertaken on the broad mineralized units and on higher grade subunits within these. In isolated cases where definitive lithological markers could be seen in both holes these were preferentially used as boundaries.

Overall correlation is good. The best correlation is achieved on downhole mineralized length, followed by contained metal (Cu % m) and the greatest variability in Ag grade. There is no systematic bias evident in grades between twin pairs, recent drilling shows higher and lower grades than historic holes in all facies. Weighted averages (Table 10.1.2_1) show very little difference in overall weighted average grade, length and metal content. There is, however, significant variability with nearly 30% difference in grade in some holes.

Based on the comparison of lithology, style of mineralization and tenor of mineralization, the twin drilling demonstrates the validity of including the historical data into the mineral resource estimate.



10.2 Historic Database Reconstruction

10.2.1 Geology

Historic data was obtained from the Weatherly offices and archives in Tsumeb. The data comprised original hard copy logs of diamond drill holes and wagon drill holes. The data was captured electronically into a database and coded on the current lithological codes (Appendix 1) devised for the programme. An abridged version of the detailed descriptions recorded for the geological intervals was captured.

10.2.2 Downhole Survey

Downhole survey data was recaptured from hardcopy logs, where present, and integrated into the Micromine database. In some cases where these had not been captured the original Eastman survey photographic plates were available. These were used to run some validation checks on captured data too. A magnetic declination of 10°W was corrected for on magnetic azimuths.

10.2.3 Collars

Some 110 of the historic drillhole collars could be found in the field and were resurveyed by a professional surveyor using a differential GPS. Coordinates for collars not located in the field were picked up off the historical drillhole logs and plans which were georeferenced in Micromine. Coordinates from drillhole collars surveyed in the field were compared against the historically recorded figures and were within 3m accuracy. All data was captured in the Namibian Lo17 system.

10.2.4 Assays

Assay results were recaptured from original hardcopy logs or historical reports where available for Cu, Pb, Zn, Ag, Mn, Fe, Cd and S. Minimal original assay certificates were available, as most of the testwork was carried out at the Tsumeb Laboratory. Commonly the entire drillhole was sampled, but in a few cases only the lower mineralized zones were sampled. The analyses were taken on split half core.

All samples were analysed for Cu and Ag, with sporadic Pb, Zn, Mn, Fe, S and Cd analyses.

10.2.5 Historic Core

A significant amount of the core from the first historic drilling campaign has been kept, while the core from later historic campaigns has been discarded. The core that was kept is stored in a core shed on Weatherly premises in Tsumeb. Although it has been stored correctly on shelves under a roof, much of the core has weathered away, or has been vandalised, so as to render it largely unusable. Only isolated portions of core within some drillholes are useable (Figure 10.2.5_1).

The remaining historical core has been re-visited, and portions have been checked against the original hardcopy logs. The recorded information correlates well with the core.



Figure 10.2.5_1 Photograph showing Condition of Historical Core

10.2.6 Comparison with Recent Drilling

In areas of tight spaced drilling around the underground workings it was possible to partially validated historic drillholes by comparison between the data sets. Five drillholes in this area where the collars could not be found (AP38, AP102, AP134, AP182, AP244) have been discarded from the database. This was due to very poor correlation between the Otavi/Mulden contact position between historic and recent drilling, likely due to incorrect collar positions. All holes are from the same historic drilling campaign.

Five drillholes (AP129, AP131, AP170, AP249, AP250) have been discarded from the database, as there was no information found for their collar positions.



11 MINERAL RESOURCE ESTIMATE

11.1 Drillhole Database

The nature of the data included in the drillhole database has been described above. The mineral resource estimation was performed by creating Datamine block models and interpolating Cu and Ag grade values from the drillhole database. Assay data compiled from the most recent and historical surface drilling programmes were exported from the drillhole database in ASCII format and consisted of the following database files:

• Surveyed drillhole collars

• Downhole survey data

• Drillhole assay data consisting of Cu values measured as a weight percent, Ag grade measured as g/t

Both Cu and Ag assays contained extremely low values and values of 0% for Cu and 0g/t for Ag. The zero values were replaced with a grade of half the assay detection limit i.e. 0.005% for Cu and 0.5g/t for Ag.

The methodology of creating the structure of the mineralized volumes has been described in Section 6.5. Grade shells were created on section lines at a minimal cut-off of 0.15% Cu. Wireframes solids of the primary, continuous mineralized zone at the base of the Mulden Group sandstones (the Lower or OB0 Zone) and the 14 Upper Zones were constructed. A sample database of the mineralized zones was created by extracting samples within these wireframes and assigning an Zone code to them (OB0 to OB14).

The majority of the drillholes were sampled at a 1m interval, however, the database included sample data from samples ranging from 0.05m – 18m in length. All samples were composited to a constant 1m length, with a composite minimum of 0.5m being permitted at the mineralized zone boundaries.

The sample statistics of the 1m composite data are tabled in Tables 11.1_1 and 11.1_2.

Statistics for the total intersected length per mineralized zone are tabled in Table 11.1_3.

Assay frequency histograms are illustrated in Figures 11.1_1 and 11.1_2

Examination of the assay statistics revealed that there were a number of very high Ag values in the data. These samples were identified and were removed by cutting the sample distribution at a value of 89.23 g/t, the 99.5 percentile.



Table 11.1_1

Tschudi Project

Descriptive Statistics for Cu % for 1m Composites per Mineralized Zone

OB FIELD NUMBER OF RECORDS

NUMBER OF SAMPLES MINIMUM MAXIMUM RANGE MEAN VARIANCE STANDARD

DEVIATION COEFFICIENT OF

VARIATION

0 Cu 7045 5338 0.01 12.44 12.43 1.02 1.13 1.06 1.04

1 Cu 7045 21 0.17 1.30 1.13 0.38 0.05 0.23 0.61

2 Cu 7045 668 0.01 4.90 4.90 0.80 0.63 0.79 0.99

3 Cu 7045 218 0.01 4.86 4.86 0.77 0.72 0.85 1.10

4 Cu 7045 22 0.10 2.55 2.45 0.47 0.26 0.51 1.09

5 Cu 7045 36 0.01 1.98 1.98 0.64 0.20 0.45 0.70

6 Cu 7045 140 0.06 4.60 4.54 1.03 0.87 0.93 0.90

7 Cu 7045 35 0.18 1.94 1.76 0.61 0.17 0.41 0.67

8 Cu 7045 92 0.03 5.40 5.37 0.96 1.09 1.04 1.08

9 Cu 7045 62 0.01 3.84 3.84 0.82 0.78 0.89 1.09

10 Cu 7045 39 0.01 2.51 2.51 0.48 0.30 0.54 1.13

11 Cu 7045 77 0.03 4.20 4.17 0.72 0.62 0.78 1.08

12 Cu 7045 168 0.01 6.80 6.80 0.70 0.80 0.90 1.29

13 Cu 7045 56 0.02 5.60 5.58 0.74 0.85 0.92 1.24

14 Cu 7045 7 0.15 1.30 1.15 0.41 0.14 0.37 0.90



Table 11.1_2

Tschudi Project Descriptive Statistics for Ag g/t for 1m Composites per Mineralized Zone


NUMBER OF SAMPLES MINIMUM MAXIMUM RANGE MEAN VARIANCE STANDARD DEVIATION COEFFICIENT OF VARIATION

0 Ag 7045 5338 0.05 1386.29 1386.24 12.35 1080.57 32.87 2.66

1 Ag 7045 21 0.05 18.00 17.95 2.19 14.89 3.86 1.76

2 Ag 7045 668 0.05 80.00 79.95 8.16 107.28 10.36 1.27

3 Ag 7045 218 0.05 39.00 38.95 5.62 69.96 8.36 1.49

4 Ag 7045 22 0.05 32.10 32.05 5.97 71.07 8.43 1.41

5 Ag 7045 36 0.05 31.00 30.95 7.03 56.67 7.53 1.07

6 Ag 7045 140 0.90 111.40 110.50 11.84 240.90 15.52 1.31

7 Ag 7045 35 1.00 20.78 19.78 5.29 18.59 4.31 0.81

8 Ag 7045 92 0.05 73.00 72.95 8.80 208.78 14.45 1.64

9 Ag 7045 62 0.05 97.70 97.65 10.37 232.00 15.23 1.47

10 Ag 7045 39 0.05 44.00 43.95 5.00 72.34 8.51 1.70

11 Ag 7045 77 0.05 36.00 35.95 6.53 55.49 7.45 1.14

12 Ag 7045 168 0.05 70.60 70.55 6.43 100.03 10.00 1.56

13 Ag 7045 56 0.05 68.00 67.95 6.35 126.35 11.24 1.77

14 Ag 7045 7 1.00 10.10 9.10 3.01 9.06 3.01 1.00



Table 11.1_3

Tschudi Project

Descriptive Statistics for Total Composite Length (m) per Mineralized Zone


NUMBER OF SAMPLES MINIMUM MAXIMUM RANGE MEAN STANDARD

DEVIATION COEFFICIENT OF

VARIATION

0 LENGTH 618 387 0.70 46.00 45.30 13.85 8.35 0.60

1 LENGTH 618 10 0.70 5.00 4.30 2.06 1.61 0.78

2 LENGTH 618 68 1.00 21.50 20.50 9.86 5.75 0.58

3 LENGTH 618 48 0.40 14.00 13.60 4.52 3.49 0.77

4 LENGTH 618 6 1.00 8.70 7.70 3.67 2.49 0.68

5 LENGTH 618 7 0.50 9.00 8.50 4.97 3.15 0.63

6 LENGTH 618 13 1.00 24.00 23.00 12.11 6.43 0.53

7 LENGTH 618 4 6.00 11.00 5.00 8.75 2.28 0.26

8 LENGTH 618 10 1.00 14.90 13.90 9.20 3.53 0.38

9 LENGTH 618 11 2.00 10.00 8.00 5.56 2.33 0.42

10 LENGTH 618 5 1.00 13.30 12.30 7.86 4.24 0.54

11 LENGTH 618 12 0.90 27.00 26.10 6.87 7.88 1.15

12 LENGTH 618 23 1.00 16.20 15.20 7.28 4.60 0.63

13 LENGTH 618 7 1.00 17.00 16.00 7.87 4.99 0.63

14 LENGTH 618 7 0.50 2.00 1.50 1.21 0.52 0.43



Figure 11.1_1a

Frequency Histograms 1m Composite Samples Cu % per Mineralized Zone

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Figure 11.1_1b

Frequency Histograms 1m Composite Samples Cu % per Mineralized Zone

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Figure 11.1_2a

Frequency Histograms 1m Composite Samples Ag g/t per Mineralized Zone

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Figure 11.1_2b

Frequency Histograms 1m Composite Samples Ag g/t per Mineralized Zone

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11.2 Resource Block Models

The total width statistics in Table 11.1_3 indicates that there is considerable variation in the thickness of the mineralized zones. The robust Lower Zone is up to 46m thick but narrows to less than 1m. The Upper Zones are in general much thinner. The average thickness is often only a few metres. The drillhole spacing is at best 25m. In some areas, however, the drillholes are separated by hundreds of metres. In order to model the narrow width of some of the mineralized zones, it was necessary to use a parent block with a small vertical dimension. The horizontal dimensions in turn are small relative to the drillhole spacing. The dimensions of the parent block within the models of the mineralized zones are 10m x 10m in plan and 3m vertically. The wireframe volume for each mineralized zone was filled with model cells. Precise splitting of the parent block was permitted vertically. Where the wireframes representing the mineralized volumes reached the topographic surface, the block model was truncated using the topographic digital terrain model (DTM).

As previously described, a DTM was modelled to represent the sub horizontal oxide-sulphide boundary. Using this DTM, the cells in the block model were classified as representing oxide or sulphide material.

With the information available it was not possible to model the exact stope geometries for the underground mining to date. A wireframe solid was created enclosing the existing on and near reef development and it has been assumed that all material over the strike length of this development above its floor elevation has been mined.

11.3 Grade Interpolation

11.3.1 Lower Mineralized Zone (OB0)

The proposed process of mineralization indicates that there could be sub horizontal layering within the Lower Zone. This is supported by the fact that mineralized layering above this horizon has been identified and modelled. Due to the horizontal spacing between drillholes, it has not been possible to link individual mineralized bands between them within the broad >0.15% Cu mineralized zone. In the vertical direction, however, it is possible to measure the spatial continuity of grade values. A downhole variogram of the Cu % values was constructed. Using a 3m lag, the range of the average downhole experimental variogram (Figure 11.3_1) is between 6m and 12m.

In order to model any horizontal layering that may exist within the thick Lower Zone, a dynamic search strategy was used for grade interpolation. When estimating grade values into the parent cells within the block model, the orientation of the search volume changed and followed the dip and dip direction of the zone.



Figure 11.3_1 Downhole Variogram Cu%

To achieve this, the dip and dip direction of each triangle within the wireframe representing the Lower Zone was calculated. These values were assigned to a sample point, and positioned at the centre of gravity of each triangle. Using an inverse distance squared interpolation method and using only the closest two sample points, a dip and dip direction was estimated for each parent cell.

To estimate the Cu and Ag grade values from the 1m composite samples, a flat search volume was used. The initial search radii were set at 50m horizontally and only 5m vertically. This search volume was increased by a factor of two should a minimum of 5 samples not be found in the initial search volume. For estimation of each model cell, the orientation of the search ellipsoid was set equal to the dip and dip direction that had previously been estimated.

The drillholes are too widely spaced to allow horizontal variograms to be modelled using the 1m composite samples, hence an inverse distance squared interpolation methodology was used. The same search criteria and estimation method were used for both Cu and Ag.



11.3.2 Upper Zones (OB1 – OB14)

Descriptive statistics relating to the thickness of the Upper Zones are tabled in Table 11.1_3. For each borehole that intersected one of the Upper Zones, the assay values were composited across the total intersected length. Table 11.1_3 indicates that some of the mineralized lenses are intersected by only a few drillholes. The estimation of grade values using composite samples across the entire intersected width was considered appropriate for the Upper Zones.

Cu and Ag grades were interpolated by using ordinary kriging. Average variograms for both Cu and Ag grades were modelled (Figures 11.3.2_1 and 2). Sample pairs were only selected within each mineralized zone.

Figure 11.3.2_1 Composite Variogram for Upper Mineralized Zones - Cu%



Figure 11.3.2_2 Composite Variogram for the Upper Mineralized Zones - Ag g/t

The search radii were initially set equal to the range of the variograms i.e. 90m for Cu and 114m for Ag. A number of the Upper Zones were intersected by only a few drillholes. In order to model any spatial variation in grade within each zone, a minimum of 2 and a maximum of 20 samples were required within the first search volume. If there were insufficient samples then the search radii were increased by a factor of 1.5. If there were still an insufficient number of samples then the search radii were increased by a factor of 20. This large search radius usually resulted in all the samples within each mineralized zone being used for kriging.

11.4 Resource Classification

A drillhole location plan shows that the mineral resource has generally been drilled at three different drill densities (Figure 11.4_1). As a result of the most recent drilling program, an area can be defined where the drillhole spacing is approximately 25m. Recent drilling, together with historical drillholes form an area where the drillhole spacing is approximately 50m x 50m or 50m x 100m. The drillhole spacing over the remainder of the resource area is much wider than this. The drillhole spacing thus provides a suitable method for the classification of the resource estimates for both the Lower and Upper Zones. Classification perimeters were hence constructed based on the spacing of the drillholes.

Despite the fact that the Upper Zones were defined by manually picking mineralized intervals from adjacent drillholes, the study of mineralization in underground workings and the understanding of the mineralization process, provide sufficient evidence to assume continuity of mineralization between boreholes.



Figure 11.4_1 Grade Shells and Drillhole Location Plan Showing Resource Classification Perimeters



The perimeters displayed on Figure 11.4_1 were constructed, and were based only on the location of the drillholes. The Measured Resource perimeter was constructed by snapping to the outermost drillholes drilled at approximately 25m spacing. The perimeter was then expanded by 12.5m. The Indicated perimeter was constructed by snapping to the outermost drillholes drilled at approximately 50m x 50m or 50m x 100m spacing. The perimeter was then expanded by 25m. The southern edge of this perimeter was further expanded to the outcrop position. The Inferred perimeter was constructed around the limits of the Datamine resource block model.

It has not been possible to accurately define which portions of the mineralized zones have previously been mined and hence estimate what has been left in situ within the strike length covered by the underground development. A wireframe volume was constructed that enclosed the majority of the historical mine workings. It is situated to the south of the Measured perimeter displayed on Figure 11.4_1 and is not included in the resource estimate.



11.5 Resource Statement

Table 11.5_1

Tschudi Project

Mineral Resource Estimate and Classification

Resource Category Tonnage (kt)

Cu Grade (%)

Ag Grade (g/t)

Density (g/cm3)

Cu Content (t) Ag (kg)

Total 50,654 0.82 10.23 2.57 413,641 518,160

Measured (Oxide) 81 1.11 10.68 2.46 896 866

Measured (Sulphide) 4,366 1.09 11.12 2.58 47,643 48,558

Indicated (Oxide) 5,245 0.65 7.07 2.46 34,241 37,056

Indicated (Sulphide) 21,263 0.89 11.27 2.58 188,653 239,631

Inferred (Sulphide) 19,699 0.72 9.75 2.58 142,207 192,049




Grade and Tonnage Estimates

Cut-off grade (Cu%) Tonnage (kt) Cu (%) Ag (g/t)

0.0 50,654 0.82 10.23

0.1 50,311 0.82 10.29

0.2 49,204 0.84 10.45

0.3 47,718 0.85 10.60

0.4 43,788 0.90 10.93

0.5 36,760 0.99 11.68

0.6 32,192 1.05 12.33

0.7 27,457 1.12 13.13

0.8 22,248 1.20 14.23

0.9 17,408 1.30 15.45

1.0 13,989 1.39 16.38

1.1 11,014 1.48 17.39

1.2 8,474 1.58 18.50

1.3 6,757 1.67 19.29

1.4 5,246 1.76 19.74

1.5 3,968 1.86 20.73

1.6 2,958 1.97 21.72

1.7 2,141 2.10 22.75

1.8 1,643 2.20 23.45

1.9 1,259 2.31 24.22

2.0 971 2.42 24.82



Table 11.5_3Tschudi Project


Domain Resource Category Tonnage

(kt) Cu (%)

Ag (g/t)

Cu Metal (t)

Ag Metal (kg)

Oxide

Measured 81 1.11 10.71 896 865

Indicated 4,546 0.73 7.82 33,004 35,533

Measured and Indicated 4,627 0.73 7.87 33,900 36,398

Inferred

Sulphide

Measured 4,347 1.09 11.15 47,594 48,494

Indicated 19,869 0.94 11.82 185,990 234,886


Inferred 18,874 0.74 9.85 140,482 185,966

Total

Measured 4,428 1.10 11.15 48,490 49,359

Indicated 24,416 0.90 11.08 218,994 270,419


Inferred 18,874 0.74 9.85 140,482 185,966




(kt) Cu (%)

Ag (g/t)

CuMetal (t)

Ag Metal (kg)

Oxide

Measured 79 1.12 10.80 890 857

Indicated 3,921 0.78 8.54 30,753 33,502


Inferred

Sulphide

Measured 4,293 1.10 11.25 47,404 48,295

Indicated 18,775 0.97 12.20 182,137 228,964


Inferred 16,719 0.79 9.99 132,857 166,949

Total

Measured 4,372 1.10 11.24 48,294 49,153

Indicated 22,696 0.94 11.56 212,890 262,466


Inferred 16,719 0.79 9.99 132,857 166,949






(kt) Cu (%)

Ag (g/t)

Cu Metal (t)

Ag Metal (kg)

Oxide

Measured 72 1.19 11.24 856 805

Indicated 3,196 0.86 9.30 27,488 29,717


Inferred

Sulphide

Measured 4,181 1.12 11.42 46,893 47,734

Indicated 16,978 1.02 12.79 173,987 217,224


Inferred 12,334 0.92 10.86 113,564 133,979

Total

Measured 4,252 1.12 11.42 47,748 48,540

Indicated 20,175 1.00 12.24 201,475 246,941


Inferred 12,334 0.92 10.86 113,564 133,979



Domain Resource Category Tonnage (kt)

Cu (%)

Ag (g/t)

Cu Metal (t)

Ag Metal (kg)

Oxide

Measured 60 1.32 12.20 792 732

Indicated 2,440 0.96 10.38 23,378 25,331

Measuredand Indicated 2,500 0.97 10.42 24,170 26,063

Inferred

Sulphide

Measured 3,977 1.15 11.69 45,757 46,500

Indicated 14,871 1.09 13.59 162,416 202,044


Inferred 10,844 0.97 11.28 105,405 122,321

Total

Measured 4,037 1.15 11.70 46,549 47,231

Indicated 17,312 1.07 13.13 185,795 227,375


Inferred 10,844 0.97 11.28 105,405 122,321





Domain ResourceCategory Tonnage (kt)

Cu (%)

Ag (g/t)

Cu Metal (t)

Ag Metal (kg)

Oxide

Measured 55 1.39 12.87 757 702

Indicated 1,744 1.08 11.67 18,882 20,355


Inferred

Sulphide

Measured 3,600 1.20 12.26 43,307 44,130

Indicated 12,994 1.16 14.40 150,236 187,118


Inferred 9,064 1.04 11.93 93,812 108,114

Total

Measured 3,654 1.21 12.27 44,064 44,832

Indicated 14,738 1.15 14.08 169,117 207,473


Inferred 9,064 1.04 11.93 93,812 108,114




(kt) Cu (%)

Ag (g/t)

Cu Metal (t)

Ag Metal (kg)

Oxide

Measured 49 1.46 13.55 713 660

Indicated 1,225 1.22 13.61 14,978 16,674


Inferred

Sulphide

Measured 3,175 1.26 12.92 40,122 41,008

Indicated 10,900 1.23 15.40 134,563 167,836


Inferred 6,899 1.12 13.10 77,517 90,397

Total

Measured 3,224 1.27 12.93 40,834 41,668

Indicated 12,125 1.23 15.22 149,540 184,510


Inferred 6,899 1.12 13.10 77,517 90,397






(kt) Cu (%)

Ag (g/t)

Cu Metal (t)

Ag Metal (kg)

Oxide

Measured 43 1.54 14.19 664 610

Indicated 856 1.39 15.33 11,885 13,120

Measured and Indicated 899 1.40 15.28 12,549 13,730

Inferred

Sulphide

Measured 2,594 1.36 13.85 35,173 35,918

Indicated 8,981 1.32 16.35 118,289 146,812


Inferred 4,935 1.24 14.69 60,961 72,501

Total

Measured 2,637 1.36 13.85 35,837 36,528

Indicated 9,836 1.32 16.26 130,173 159,931


Inferred 4,935 1.24 14.69 60,961 72,501




(kt) Cu (%)

Ag (g/t)

Cu Metal (t)

Ag Metal (kg)

Oxide

Measured 36 1.66 14.86 597 535

Indicated 657 1.52 16.60 10,001 10,902

Measured and Indicated 693 1.53 16.51 10,597 11,437

Inferred

Sulphide

Measured 2,064 1.46 14.83 30,141 30,595

Indicated 7,414 1.39 17.21 103,418 127,591


Inferred 3,820 1.32 15.58 50,397 59,520

Total

Measured 2,100 1.46 14.83 30,737 31,130

Indicated 8,070 1.41 17.16 113,418 138,493


Inferred 3,820 1.32 15.58 50,397 59,520



Figure 11.5_1 Mineral Resource Estimate Grade Tonnage Curves



12 CONCLUSIONS AND RECOMMENDATIONS

A significant copper resource has been delimited at Tschudi. The primary mineral resource is present within the Lower Zone. The thickness of this zone varies significantly from 1m to 45m. Additional mineralisation within 14 discontinuous Upper Zones has been modelled. In order to assess the potential for extraction by open pit a Scoping Study should be undertaken.

The drillhole spacing across the mineralization at Tschudi varies considerably. Estimating the thickness of the Lower Mineralized Zone, and the continuity of mineralization within it, is better where more drillhole logging and assay data is available. The extent of the mineralization, however, has not been fully delineated. It has only been possible to model the presence and extent of the Upper Mineralized Zones in areas where there is relatively close spaced drilling. There is considerable potential for extensions to the delimited mineralization to the southwest and at depth. This will require additional drilling if it is to be tested. Further drilling will also be required for upgrading the present resource estimates.

The current base of oxide surface is based on a 50% of total copper sulphuric acid soluble surface. In conjunction with metallurgical test work more useful surfaces should be developed which will better reflect changes in critical criteria for metallurgy.



13 REFERENCES

Anderson, G. M. & Macqueen, R. W., 2003. Mississippi Valley-Type Lead-Zinc Deposits, 79-90. In Roberts, R. G.& Sheahan, P A., Eds., Ore Deposit Models, Geoscience Canada Reprint Series.

Brown, A. C., 2003. Sediment-hosted Stratiform Copper Deposits, 99-115. In Roberts, R. G.& Sheahan, P A., Eds., Ore Deposit Models, Geoscience Canada Reprint Series.

Cairncross, B., 1997. The Otavi Mountainland Cu-Pb-Zn Deposits, Namibia. Mineralogical Record, Mar/Apr 1997.

Emslie, D. P., 1979. The Mineralogy and Geochemistry of the Copper, Lead and Zinc Sulphides of the Otavi Mountainland, South West Africa. Unpublished PhD thesis.

Lombaard, A. F., Gunzel, A., Innes, I. & Kruger, T. L., 1986. The Tsumeb Lead-Copper-Zinc-Silver Deposit, South West Africa/Namibia, 1761-1787. In: Anhausser, C. R. & Maske, S., Eds., Mineral Deposits of Southern Africa, Vol 2. Geol. Soc. S. Afr., 1314pp.

Lonergan, J. E. & Saayman, A. F., 2002. Orebody Modelling & Starter Pit Design, Tschudi Deposit. Unpublished company report.

Melcher, F., 2003. The Otavi Mountainland in Namibia: Tsumeb, Germanium and Snowball Earth, 413-435. Mitt.Osterr.Miner.Ges. 148.

Misiewicz, J. E., 1988. The Geology and Metallogeny of the Otavi Mountainland, Damara Orogen, SWA/Namibia. Unpublished M.Sc. thesis.

Murphy, G. C., 1980. Tschudi Prospect – Progress Report of Exploration and Geology to December 1979. Unpublished TCL company report.

Viviers, G. J., unknown. Geological Report on the Tschudi Copper/Silver Deposit. Unpublished Gold Fields Namibia company report.

Viviers, G. J., 1992. Geological Report on the Tschudi Copper/Silver Deposit. Unpublished TCL company report.

Appendix A Page: 1

Appendix A Geological Logging Codes and Templates

Appendix A Page: 2

Geological Logging Lithology Codes:

CODE NAMECALC Calcrete SOIL Soil LOSS Core loss PCOL Precollar / pilot hole VQTZ Quartz vein VQCB Quartz carbonate vein CAVT Cavity / Karst feature DOLO Undefined dolomite DOOL Oolitic/pisolitic dolomite DMAS Massive dolomite DBED Bedded dolomite DMIC Micritic dolomite DALT Altered dolomite DREC Recrystallised dolomite DSTR Stromatolitic dolomite DBRE Brecciated dolomite (frags in any matrix) DCSB Dolomite with chert/silica beds DASB Dolomite with argillite/sandstone beds CHER Massive chert COOL Oolitic/pisolitic chert CBRE Brecciated chert (frags in any matrix) CALG Algal chert AREN Arenite - Standard sandstone term if not arkosic or quartz-rich ARKO Arkose - Feldspar rich arenite - generally more weathered, Fe / K alteration, can be softer ARQZ Quartz arenite - Grey, generally glassy, hard, med grained v. quartz rich arenite ARBL Black arenite - Dark grey to black colour, fine-medium grained ARCA Calcareous arenite ARBR Brecciated sandstone (frags in any matrix) GRYW Greywacke MUST Mudstone SIST Siltstone CGLM Conglomerate BCGM "Basal" conglomerate

Appendix A Page: 3

Geological Logging Mineral Codes:

Minerals Code

Aurichalcite (Zn, Cu)5(CO3)2(OH)6 AC

Azurite Cu3(CO3)2(OH)2 AZ

Bornite Cu5FeS4 BN

Calcite CaCO3 CAL

Cerussite PbCO3 CR

Chalcocite Cu2S CH

Chalcopyrite CuFeS2 CY

Chrysocolla (Cu,Al)2H2Si2O5(OH)4.nH2O CS

Copper Cu CU

Covellite CuS CV

Cuprite Cu2O CP

Descloizite PbZn(VO4)(OH) DC

Digenite Cu9S5 DG

Dioptase CuSiO2(OH)2 DP

Dolomite CaMg(CO3)2 DOL

Duftite PbCu(AsO4)(OH) DF

Epidote Ca2(Al,Fe+3)(SiO4)3(OH) EP

Fluorite CaF2 FL

Galena PbS GA

Goethite FeO(OH) GO

Gold Au AU

Graphite C GR

Gypsum CaSO4.2H2O GY

Hematite Fe2O3 HE

Ilmenite FeTiO3 IL

Magnetite Fe3O4 MG

Malachite Cu2(CO3)(OH)2 MC

Mottramite PbCu(VO4)OH MO

Plancheite Cu8Si8O22(OH)4.H2O PL

Pyrite FeS2 PY

Silver Ag AG

Sphalerite (Zn,Fe)S SP

Sulphur S SU

Tennanite Cu10(Zn,Fe)2As4S13 TN

Appendix A Page: 4

Map and Section Legends

Appendix A Page: 5

Geological Logging Template

BH ID: Geologist: Date:

From

To

Sam

p N

o.

Gra

in s

ize

Har

d

Col

our

Wea

th

Lith

cod

e

Description

Bed

°

BC

BC

°

Stru

ctur

e

Stru

ct °

Alteration Veining

Min Style

Mineralization

Alt

1

Int

Alt

2

Int

QZ

QZ

CA

L

CA

L

PY

MC

AZ

CS

CH

CY

BN

/CV

CP

/CU

GA

SP

TOTA

L

Appendix B Page: 1

Appendix B Twin Drillhole Comparisons

Appendix B Page: 0

Striplog Comparison Drillholes AP55 and TAP55

Appendix B Page: 1


Appendix B Page: 2


Appendix B Page: 3


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