CNI 43-101 Technical Report Las Cruces Copper Project, Southern Spain Prepared for MK Resources Company July 8, 2004 9033.05

CNI 43-101 Technical Report

Las Cruces Copper Project,

Southern Spain

Prepared for

MK Resources Company

July 8, 2004

9033.05

Prepared by

Pincock, Allen & Holt

Richard Addison, P.E.

Darrel L. Buffington, P.E.

Gerald D. Crawford, P.E.

Nelson D. King

Mark G. Stevens, C.P.G.

A Division of Hart Crowser

274 Union Boulevard, Suite 200

Lakewood, Colorado 80228-1835

Fax 303.987.8907

Tel 303.986.6950

CONTENTS

				Page
1.0	EXECUTIVE SUMMARY			1.1
	1.1	Introduction		1.1
	1.2	Project Summary		1.1
		1.2.1	Geology, Exploration and Resource Estimation	1.1
		1.2.2	Mining	1.3
		1.2.3	Water Management	1.5
		1.2.4	Metallurgy and Processing	1.6
		1.2.5	Infrastructure and Ancillary Facilities	1.7
		1.2.6	Environmental, Permitting and Restoration	1.7
		1.2.7	Capital Costs	1.8
		1.2.8	Operating Costs	1.9
		1.2.9	Project Execution and Schedule	1.10
		1.2.10	Metal Marketing	1.12
		1.2.11	Exchange Rate Effects	1.12
		1.2.12	Financial Model	1.12
2.0	INTRODUCTION AND TERMS OF REFERENCE			2.1
	2.1	Qualified Person and Participating Personnel		2.1
	2.2	Terms and Definitions		2.1
	2.3	Units		2.2
	2.4	Source Documents		2.2
3.0	DISCLAIMER			3.1
4.0	PROPERTY DESCRIPTION AND LOCATION			4.1
	4.1	Location		4.1
	4.2	Tenure		4.1
	4.3	Mineral Zone Location		4.4
	4.4	Environmental Liabilities		4.4
	4.5	Permitting		4.4
	4.6	Royalties and Encumbrances		4.5
5.0	ACCESSIBILITY, CLIMATE, LOCAL RESOURCES AND INFRASTRUCTURE			5.1
	5.1	Topography, Elevation and Vegetation		5.1
	5.2	Access and Transport		5.1

Pincock, Allen & Holt 9033.05 July 8, 2004 Page i

CONTENTS (Continued)

			Page
	5.3	Infrastructure	5.2
	5.4	Sufficiency of Surface Rights	5.2
6.0	HISTORY		6.1
7.0	GEOLOGIC SETTING		7.1
	7.1	Regional Geologic Setting	7.1
	7.2	Deposit Geologic Setting	7.1
8.0	DEPOSIT TYPES		8.1
9.0	MINERALIZATION		9.1
	9.1	Secondary Sulfide Mineralization	9.1
	9.2	Primary Sulfide Mineralization	9.2
10.0	EXPLORATION		10.1
	10.1	Deposit Exploration	10.1
11.0	DRILLING		11.1
	11.1	Core Drilling	11.1
	11.2	Reverse Circulation Drilling	11.3
12.0	SAMPLING METHOD AND APPROACH		12.1
	12.1	Core Sampling	12.1
	12.2	Core Recovery	12.2
13.0	SAMPLE PREPARATION, ANALYSES AND SECURITY		13.1
	13.1	Sample Preparation and Analyses	13.1
	13.2	Density Testwork	13.1
	13.3	Sample Security	13.2
14.0	DATA VERIFICATION		14.1

Pincock, Allen & Holt 9033.05 July 8, 2004 Page ii

CONTENTS (Continued)

				Page
	14.1	Assay Checks		14.1
	14.2	PQ Twin Drilling		14.2
	14.3	Other Verification		14.2
	14.4	Drillhole Database Management		14.2
15.0	ADJACENT PROPERTIES			15.1
16.0	MINERAL PROCESSING AND METALLURGICAL TESTING			16.1
		16.1	Metallurgical Testing Programs	16.1
		16.2	Metallurgical Samples	16.1
		16.3	Batch Metallurgical Testing	16.2
		16.4	Pilot Leaching and Solvent Extraction Testing	16.5
		16.5	Pilot Effluent Testing	16.6
		16.6	Ore Variability	16.6
		16.7	Copper Recovery Predictions	16.7
		16.8	Copper Production Estimate	16.7
		16.9	PAH Comments	16.9
17.0	MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES			17.1
	17.1	General Information		17.1
	17.2	Drill Hole Sample Database		17.1
		17.2.1	Database Content	17.1
		17.2.2	Database Checking	17.3
		17.2.3	Drill Hole Core Recovery	17.4
		17.2.4	Sample Statistics	17.5
	17.3	Composites		17.8
		17.3.1	Composite Calculation	17.8
		17.3.2	Composite Statistics	17.8
		17.3.3	Composite Variography	17.9
	17.4	Rock Model		17.10
	17.5	Grade Models		17.10
	17.6	Resource Estimate		17.12
		17.6.1	Resource Statement	17.12
		17.6.2	Resource Classification	17.13
		17.6.3	Previous Resource Estimates	17.14
	17.7	Additional Exploration Potential		17.15

Pincock, Allen & Holt 9033.05 July 8, 2004 Page iii

CONTENTS (Continued)

				Page
	17.8	Reserve Development		17.16
		17.8.1	Mine Design	17.16
		17.8.2	Cutoff Grade	17.18
		17.8.3	Pit Design	17.19
		17.8.4	Mining Dilution and Loses	17.22
	17.9	Ore Reserves Statement		17.22
		17.9.1	Effects on Reserves by Other Factors	17.23
		17.9.2	Recoverability	17.24
18.0	OTHER RELEVANT DATA AND INFORMATION			18.1
19.0	INTERPRETATION AND CONCLUSIONS			19.1
	19.1	Geologic Evaluation Conclusions		19.1
	19.2	Exploration Program Conclusions		19.1
	19.3	Resource Estimation Conclusions		19.1
20.0	RECOMMENDATIONS			20.1
21.0	REFERENCES			21.1
22.0	ADDITIONAL REQUIREMENTS FOR DEVELOPING OR PRODUCING PROPERTIES			22.1
	22.1	Mining Operations		22.1
		22.1.1	Ore Processing	22.2
	22.2	Production Schedule		22.18
	22.3	Recoverability		22.19
	22.4	Markets		22.19
	22.5	Contracts		22.21
	22.6	Environmental Considerations		22.22
		22.6.1	Permitting	22.22
		22.6.2	Restoration	22.24
		22.6.3	Groundwater Management	22.26
	22.7	Taxes and Royalties		22.27
	22.8	Capital and Operating Cost Estimates		22.27
		22.8.1	Operating Costs	22.29
	22.9	Economic Analysis		22.31

Pincock, Allen & Holt 9033.05 July 8, 2004 Page iv

CONTENTS (Continued)

Page

23.0 CERTIFICATES OF QUALIFICATION 23.1

TABLES
1-1	Mineral Resource Summary (1.00% copper cutoff, 0.97 density adjustment)	1.3
1-2	Diluted Mineable Reserves	1.4
1-3	Project Capital Cost Estimate	1.9
1-4	Total and Unit Operating Costs, Year 5	1.10
12-1	Core Sample Size Distribution	12.1
16-1	Copper Production Schedule	16.9
17-1	Sample Copper Data Statistics	17.7
17-2	Sample Density Data Statistics	17.7
17-3	Composite Copper Data Statistics	17.9
17-4	Composite Density Data Statistics	17.9
17-5	Block Model Copper Data Statistics	17.11
17-6	Block Model Density Data Statistics	17.11
17-7	Mineral Resource Summary (1.00% copper cutoff, 0.97 density adjustment)	17.13
17-8	Resource Confidence Classification Criteria	17.14
17-9	Comparison of Historical Resource Estimates	17.15
17-10	Basic Cutoff Grade Calculation	17.18
17-11	Design Basis – Mine Operating Cost Assumptions	17.19
17-12	Pit Design Parameters	17.20
17-13	Effective Dilution Summary	17.22
17-14	Diluted Mineable Reserves	17.23
22-1	Major Process Plant Equipment List	22.6
22-2	Principal Processing Parameters	22.7
22-3	Mine and Mill Production Schedule	22.19
22-4	Historic Refined Copper	22.20
22-5	Summary of Permit Status	22.23
22-6	Annual Restoration Areas and Budgets	22.26
22-7	Project Capital Cost Estimate	22.29
22-8	Cash Flow Summary	22.30
22-9	Cash Flow Model Parameters	22.32
22-10	Sensitivity of NPV to Copper Prices at a Fixed Exchange Rate of 1.00 euros per US dollar	22.34
22-11	Sensitivity of IRR to Copper Price and Euro Conversion Rates at a Fixed Copper Price of $1.00/lb, $ Millions	22.34
22-12	Sensitivity of NPV @10% Copper Price and Euro Conversions, $Millions	22.35

Pincock, Allen & Holt 9033.05 July 8, 2004 Page v

CONTENTS (Continued)

		Page
22-13	Sensitivity of IRR to Copper Price and Euro Conversion	22.35
FIGURES
4-1	General Location Map – Las Cruces Copper Project	4.2
4-2	Mine Status Year +10, Main Infrastructure Items	4.3
7-1	General Geologic Section	7.3
7-2	General Geologic Plan of Deposit Lenses	7.4
11-1	Drill Hole Locations with Topography	11.2
16-1	Metallurgical Sampling Composites, Planview	16.3
16-2	Metallurgical Sampling Composites, Section	16.4
16-3	Copper Extraction vs Copper Grade	16.8
17-1	Model Boundary with Drill Holes & Ultimate Pit Outline	17.2
17-2	Number of Samples & % Copper Above Core Recovery Cutoff	17.6
17-3	Ultimate Pit Design	17.21
22-1	Las Cruces Process Flow Diagram	22.4
22-2	LME Spot Copper Price	22.5
22-3	Financial Model – Cash Flow Analysis, Economic Sensitivity	22.16
22-4	Historical Copper Prices	22.21
22-5	Cash Flow Economic Sensitivity	22.33
22-6	Historical Exchange Rate	22.36

Pincock, Allen & Holt 9033.05 July 8, 2004 Page vi

1.0 EXECUTIVE SUMMARY

1.1 Introduction

Pincock, Allen & Holt (PAH) conducted an Independent Technical Audit of the Las Cruces Copper Project (Las Cruces or the Project) located near Seville in southern Spain for the benefit of potential financing parties (the Lenders). The audit resulted in the PAH report titled “Independent Technical Audit of the Las Cruces Copper Project” dated April 29, 2004.

Following the submission of the Technical Audit, MK Resources Company requested that PAH produce a Technical Report Under the format of CNI 43-101F1. This report is a summary of the original Technical Audit and is in compliance with the format and content requirements set forth under Canadian National Instrument 43-101.

The Technical Audit and the Technical Report are based upon the project designs, development plans, costs, and predicted performance of the Project as depicted in the November 2003 Feasibility Study prepared by DMT-Montan Consulting GmbH (DMT-MC) and Outokumpu Technology Group (OTG). Bechtel International, Inc. (Bechtel) had prepared a Feasibility Study of the project in March 2001. Some information from the Bechtel study was reviewed and incorporated by DMT-MC and OTG into the recent study.

1.2 Project Summary

Mineable reserves for the Las Cruces Copper Project, estimated at 16 million tonnes of ore, will be processed to produce 966,000 tonnes of cathode copper over the 15-year life of the mine. The reserves were estimated at a Base Case copper price of $0.76 per pound. The initial project capital cost is estimated at 280 million euros. The project operating costs, over the 15-year mine life, are estimated to be 0.327 euros per pound of copper produced.

No material deficiencies were identified during the audit that would preclude the Project from meeting the designed production and cost objectives within the range of the cost estimates presented in the Feasibility Study. The Las Cruces project has been thoroughly evaluated and reasonable and achievable development and operating plans have been prepared.

1.2.1 Geology, Exploration, and Resource Estimation

Mineralization at Las Cruces, as in most other Iberian Pyrite Belt deposits of southern Spain and Portugal, consists of syngenetic massive sulfides containing polymetallic mineralization. Post depositional secondary copper enrichment occurred in the upper part of the massive sulfide deposit, forming the mineralization of interest. Subsequently the deposit was buried under 100 to 150 meters of sandstone and calcareous mudstone (marl).

Pincock, Allen & Holt 9033.05 July 8, 2004 1.1

PAH finds that the degree of geologic evaluation of the Las Cruces deposit is generally reasonable for supporting a feasibility study. Geologic interpretation for the resource model was consistent with reasonable correlations between drill holes. The majority of the geologic data is from previous exploration conducted by Rio Tinto. The deposit geometry is reasonably defined from drill holes averaging around 50 meters spacing. The understanding of the copper grades, however, has been somewhat impacted by poor core recovery. The secondary sulfide mineralogy has been reasonably evaluated and is subject to wide grade variations over short lateral distance, as would be expected given the nature of the mineralization.

PAH’s evaluation of the drilling and sampling programs by Rio Tinto on the Las Cruces deposit was reasonably conducted using industry standard procedures and techniques. Drill hole coverage is adequate for feasibility study, with some infilling recommended. Abundant evidence exists of the tendency to preferential loss of chalcocite during the drilling and sampling, particularly in lower core recovery intervals, resulting in lower copper grades than is realistic. As a result, the confidence in the copper grades reported for samples in the lower core recovery zones is in question. The evidence indicates, however, that the copper grades should tend to be higher than reported. Analyses were conducted using standard procedures and were accompanied by reasonable check assay programs for quality control, which showed no significant grade bias or systematic analytical problems.

DMT-MC created a revised resource model for the Las Cruces deposit in 2003, largely to incorporate changes in what data was included or excluded as a result of sample recovery issues. The modeling approach otherwise was very similar to the previous 2000 model created by Independent Mining Consultants. The estimated resource includes all secondary sulfide material in the model at the given cutoff grade, without consideration to any mineable limits. Table 1-1 shows that at a 1.0 percent copper cutoff, the measured + indicated resource is 15.6 million tonnes of HCH, HCL and HC4 ore types, averaging 6.89 percent copper. There is an additional inferred resource of 0.40 million tonnes of primarily HCH averaging 8.66 percent copper. Incidental primary sulfide (CZ) and gossan that occurs in the planned pit will be stockpiled separately and is not considered part of this resource estimate.

PAH believes that the DMT-MC secondary sulfide resource model was carefully created using standard engineering methods appropriate for this deposit. The model provides a reasonable representation of the distribution of the secondary sulfide mineralogic zones. PAH believes that the resource model, due to natural deposit variability and due to the excluded sample and composite data (about 30 percent of composite data excluded), may be a conservative estimator of grade locally, but believes that on a global basis the estimation is reasonable.

Pincock, Allen & Holt 9033.05 July 8, 2004 1.2

TABLE 1-1 / 17-7

MK Resources Company

Technical Report for the Las Cruces Project

Mineral Resource Summary (1.00% copper cutoff, 0.97 density adjustment)

Ore Type	Measured		Indicated		Measured +Indicated		Inferred
	Tonnes (000)	Grade (% Cu)	Tonnes (000)	Grade (% Cu)	Tonnes (000)	Grade (% Cu)	Tonnes (000)	Grade (% Cu)
HCH	3,860	7.25	2,670	8.00	6,530	7.56	274	9.48
HCL	5,020	6.56	3,370	5.77	8,390	6.24	37	4.86
HC4	440	8.61	230	8.55	670	8.59	49	6.93
Total	9,320	6.94	6,270	6.82	15,590	6.89	360	8.66

CLC plans for infill drilling over the life of the mine to increase the resource confidence for short-term mine planning. Fourteen of these holes are planned to be drilled prior to the start of mining, to supply local in-fill data and will also serve to help validate the model. The resource estimate compares reasonably with those from previous resource models and grade differences are due to intentional changes in modeling methodologies. The model provides an acceptable basis for which subsequent mine engineering work can be conducted in order to delineate mineable reserves acceptable to the SEC or any other financial institution.

1.2.2 Mining

Las Cruces will be mined using conventional open pit mining methods, based upon hydraulic shovels and trucks, with drilling and blasting in the lower marls and ore zones. The project is unusual in respect to the relatively high stripping ratio, and correspondingly high grade of the ore zones. CLC plans to use contract miners for both prestripping and production over the life of the mine.

The mine plans for Las Cruces have evolved over time to reflect differing equipment requirements and slope constraints. Notably, the cutoff grades for the deposit have not been altered substantially, as various investigations have shown that changes to the cutoff grade has only a minimal effect on the reserves. As with the stripping ratio, this characteristic can be attributed to the high average grade of the ore.

The ultimate pit plans and production schedule were developed by Independent Mining Consultants (IMC) of Tucson, Arizona, with due regard for the requirements of the equipment and the geotechnical attributes of the slopes, with a minor exception on the east side. IMC developed a series of seven phases or pushbacks for the scheduling effort, which were subsequently used to develop both tonnes and grade by year and the annual mine advance maps. PAH believes IMC followed accepted engineering practice in the effort.

The reported reserves for the Las Cruces project are presented below as Table 1-2.

Pincock, Allen & Holt 9033.05 July 8, 2004 1.3

TABLE 1-2 / 17-14

MK Resources Company

Technical Audit of the Las Cruces Project

Diluted Mineable Reserves

Description	Ore			Waste Tonnes x1000	Total Tonnes x1000	Strip Ratio w/o
	Tonnes x1000	Percent Copper	Lbs. Cu mm
Proven	13,938	6.897	2,119.3	—	13,938	n/a
Probable	1,432	6.416	202.6	—	1,432	n/a
Other Dilution	602	0.574	7.6	—	602	n/a
Waste	—	—	—	232,013	232,013	n/a
Total	15,972	6.62	2,329	232,013	247,985	14.53

Note: Mineable reserve is included in the measured and indicated mineral resource estimate.

The above reserves were developed at a nominal 1 percent copper cutoff; however; within the realm of reasonable cutoff grades (1.0 to 2.0 percent copper) the reserves change very little. Lower cutoffs have a similarly modest effect. Current plans are to take essentially the entire oxide copper zone to the plant. Selective mining will be practiced if the ore is sufficiently predictable, but highly selective mining is not required to achieve the above reserve.

The upper layers of the marl zone are anticipated to be free-digging, which imparts a significant cost savings on the project mining costs. Overburden of this soft nature is unusual in metals mining; therefore, considerable attention was focused on the assumption. PAH examined several cores and was able to cut the material with a knife. PAH also discussed the marls with two equipment vendors, both of which felt that their larger mining units were capable of digging the material without blasting. Furthermore, the characteristics of the marl indicate that avoiding the blasting step may have benefits on the overall slope stability.

Overall, the mine equipment selection for the operation appears appropriate. DMT-MC has specified the use of 1.8 cubic-meter shovels and 35-tonne haul trucks to develop the ore. There is no evidence that equipment of this small size is necessary to mine the deposit: PAH feels that the relatively low operating costs estimated would not be possible were these small units employed. PAH further notes that the contract mining estimates did not include equipment of this size for production.

Drilling and blasting specifications are generally reasonable with regards to drill productivities and spacing. Overburden will be mined in 10-meter lifts, while ore zones will be mined in 5-meter lifts. The DMT-MC specification for 3.5 inch drill bits for the 5-meter bench is unlikely, as the resulting powder factor of 0.1 kg/tonne would be insufficient, even in the relatively soft rock expected at Las Cruces.

On an overall basis, however, the costs and productivities are reasonable and are adequately supported by the contract mining estimates.

Pincock, Allen & Holt 9033.05 July 8, 2004 1.4

1.2.3 Water Management

Control of groundwater is an essential aspect of the mine plan due to the potential impacts on the mining schedule and costs, pit slope stability, environmental protection and regional groundwater resources. To provide for a safe and efficient mining operation, excavation of the proposed open pit requires advance dewatering of the aquifer. The groundwater management system will consist of eight separate sectors, each sector equipped with pit wells, collection lines, a booster pump station, distribution lines and re-injection wells. Up to 28 dewatering wells will be drilled around the perimeter of the open pit with about 5 in-pit wells to be drilled from stable in-pit benches. Approximately 26 re-injection wells will be drilled with capacities ranging from 2 to 15 liters per second (lps), depending upon the sector and well location. The site water balance indicates that the average pumping rate will be about 140 lps.

The hydrogeologic analyses performed to support the groundwater management plan designs are more than sufficient to demonstrate the feasibility of the groundwater management plan. Well-established methods of field-testing, analysis and computer modeling have been employed to develop the conceptual design. The risks that groundwater problems could pose to efficient mine operations are well mitigated by the dewatering and re-injection scheme. A key to success will be adaptive management of the scheme as the project develops, through the planned interactive use of a computerized groundwater flow simulation model. This will serve to progressively enhance understanding of site hydrogeology and how it is being modified by each stage of the proposed scheme. The Guadalquivir River Authority issued an approval for the de-watering re-injection management system in October 2003.

The Las Cruces project will modify the actual surface water flows that currently exist. The Molinos, Garnacha and two other small streams will be diverted so that the mining and ore processing operations will have minimal impact upon the existing watershed. Runoff from the site will be maintained as either contact or non-contact water. Any water which enters the pit, waste dumps, tailings facility or haul roads, or may otherwise come in contact with potentially acid generating rock, will be maintained as contact water. This water will be collected and will report to either a holding pond and added to the process water stream or the contact water treatment pond, and treated prior to reuse or release. Non-contact water will report to sedimentation ponds prior to discharge. All runoff management systems have been designed based on a 100-year storm event. After mine closure, all potentially reactive surfaces will be covered and there will no longer be any contact water. The contact water treatment plant, however, will be kept operational as a contingency measure. PAH believes that the surface water management plan is reasonable and well conceived and should achieve the planned goals for diversion, collection and treatment, as required.

Process water will primarily be obtained from the San Jeronimo wastewater treatment plant (WWTP). Since the discharge from this WWTP is required to maintain the “ecological flow” in the Guadalquiver River downstream of the mine during dry periods, CLC will only be allowed to draw plant effluent, at the rate of about 120 lps, for approximately 7 months per year. To provide water

Pincock, Allen & Holt 9033.05 July 8, 2004 1.5

for the operation during the remainder of the year, a primary water supply pond will be constructed. Because of the volume of the pond, the retention dam is considered to be a large dam; therefore a spillway will be constructed. Spanish regulations require that a detailed operating and contingency plan be submitted for approval with the dam permit application. This document will need to be developed by CLC as part of the final dam design. The wastewater will be chlorinated prior to placement in the pond and will be pre-treated by a reverse-osmosis plant for specific process uses.

1.2.4 Metallurgy and Processing

The DMT-Montan feasibility study reflects modifications to the process plant design based on technology studies done by Outokumpu Research Oy (OPRC), a world leader in base metal processing technology. These technology studies and recently completed test works demonstrated that improvements could be made to the process plant design to reduce capital costs. The process will be simplified by using solely atmospheric leaching and eliminating expensive autoclaves for pressure leaching without sacrificing copper recovery.

PAH believes that the metallurgical testing performed by Outokumpu to develop the process flowsheet, material balances and plant design criteria was well conceived, complete and conclusive. The metallurgical samples used were representative of the types and grades planned for processing. Results from bench-scale testing were properly used to establish copper recovery relationships for the three primary ore types. These relationships were used for mine modeling and to predict the project copper recoveries based on the ore type and grade processed. Pilot-scale test results confirmed the copper recovery estimate and provided design criteria for the process plant design. Some additional testing is planned prior to final plant design to confirm recovery projections and provide additional design information, particularly for the leach reactors.

Metallurgical testwork indicates that copper recoveries should exceed 90 percent over the project life. The copper production schedule and estimate is reasonable. Recoveries are based upon individual ore types and ore grades being processed.

Ferric-leach/SX-EW technology is becoming more common at copper heap leach operations that treat chalcocite-rich ores; however, the agitated ferric-leach technology that will be used for treating the Las Cruces ore is a relatively new technique in the copper industry. A related process had been used at a copper mine in Australia and will be used at a similar sized facility being constructed in Laos. In addition, the zinc industry in Finland has utilized the agitated ferric-leach technology successfully for years. The SX-EW portion of the planned facility is proven technology.

The process plant has been designed in a logical and complete manner. The design is well conceived and should be capable of meeting the design throughput rate of 1.3 million tonnes per year and copper production rate of 66,000 tonnes per year. Aside from the atmospheric leach reactors, the processing equipment and techniques are conventional for the minerals processing industry. Ore will be crushed in three-stages, ground in a ball mill, and then leached with ferric sulfate, sulfuric acid and oxygen in agitated tanks. After thickening, the copper-rich solution is

Pincock, Allen & Holt 9033.05 July 8, 2004 1.6

passed through a solvent extraction circuit and then the copper is recovered as cathodes by electrowinning. Leach residues will be filtered, stacked and encapsulated in a mine waste dump. The disposal design will exceed regulatory requirements. A process bleed stream will be treated to remove heavy metals and then discharged. The effluent quality should meet discharge standards as evidenced by the laboratory testwork.

1.2.5 Infrastructure and Ancillary Facilities

As the project is in an area with well-developed infrastructure, no problems are anticipated meeting the infrastructure requirements. Grid power is available nearby, process water will be pumped from a water treatment plant, potable water will be supplied from nearby wells, good highways pass by the project site, port facilities are near to the project, and ample housing and labor support facilities are available nearby. Support facilities to be constructed at the site are sufficient to administer, secure, supply, maintain and operate the project.

Land purchase agreements are progressing with over 60 percent of the land purchases being arranged. Additional purchases being made in timely manner are critical to meeting the project schedule. No foreseen problems in purchasing or leasing properties have been identified.

1.2.6 Environmental, Permitting and Restoration

The permitting process and administration for a mining project in Spain is very complex and being comprised of state, regional, provincial and municipal levels of government. Based on PAH’s review of the EIS, it appears complete and in compliance with Spanish, EU and World Bank standards. The document contains adequate baseline data describing flora, fauna, soil resources, and water and air quality. Of importance in the permitting process for the Project was the issuance of a positive Declaration of Environmental Impact (DEI), which was issued by the Provincial Delegation of the Regional Ministry for the Environment on May 9, 2002. CLC was granted the Mining Concession on August 6, 2003. There are a number of permits outstanding in order for the project to proceed. To avoid potential delays, PAH recommends that CLC maintain the current emphasis on permitting, including land acquisition and supporting detailed engineering.

The environmental restoration plan for the Project is based on the plan submitted as part of the Project Environmental Impact Study (EIS) dated March 2001. The current plan incorporates the results of the study plus modifications for project optimizations and incorporation of all permit conditions specified in the DEI. Two principal modifications to the original restoration plan are: (1) backfilling of the open pit with inert waste materials, and (2) surface areas to be restored have been modified in line with the new mining plan. PAH has reviewed the other required modifications stated in the DEI and found them to be achievable within the revised restoration plan. Per the DEI directives, at the appropriate time, CLC must present for authorization a specific Performance Project plan for backfilling the pit with inert materials. As the backfilling of the pit comprises a significant issue within the environmental restoration plan, PAH has concerns as to the timing for submission of the plan and its subsequent authorization by environmental authorities, especially if any changes would occur in the environmental administration of the project.

Pincock, Allen & Holt 9033.05 July 8, 2004 1.7

PAH believes that the Las Cruces project has properly incorporated environmentally responsible practices into the design of their operations. Waste dumps will be continually contoured and reclaimed to enhance the appearance of the Las Cruces project. The area’s water resources affected by the mine have been studied and plans have been developed to protect these resources. Water Management Consultants, an international hydrological engineering and design company, developed the water management plan, including the extraction system for dewatering the overlying sand aquifer around the mine site. Water will be extracted from the aquifer to prevent it from entering the mine pit. The extracted water will be re-injected into the aquifer away from the mine. Water needed to process the ore will be pumped 15 kilometers from a sewage treatment plant and further treated before use.

1.2.7 Capital Costs

The total project capital estimate of 292 million euros is thorough, developed using sound engineering principles, and appears reasonable and complete. The cost of the project is higher than projects of similar throughput capacity; however, the high cost is because of the complexity of the process facility, the fact that cathode copper will be produced on site, the high initial waste stripping requirement, amount of land purchases, and the extensive water management systems necessary for the project to be successful.

The initial capital costs of the project are estimated at 280 million euros, including working capital, land purchases and contingencies, but excluding bonding requirements, interest during construction and other financing costs. The estimated capital costs reflect an overall reduction in costs over the Bechtel Feasibility as a result of the atmospheric leach process plant improvements described above.

The capital cost estimates were prepared first by obtaining quotations from contractors and equipment suppliers and then applying those costs to the feasibility-level design plans developed for earthworks and structures. A summary of the initial and sustaining project capital requirements is presented in Table 1-3.

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TABLE 1-3 / 22-7

MK Resources Company

Las Cruces Project Technical Report

Project Capital Cost Estimate

Area	Initial Capital (euros in 000’s)	Sustaining Capital (euros in 000’s)	Total Capital (euros in 000’s)
Geotechnical Management	550	0	550
Mining	37,527	288	37,815
Surface Water Management	2,302	4,315	6,617
Ground Water Management	7,449	4,883	12,332
Process Plant	149,991	4,929	154,921
Tailings and Mine Rock Management	6,376	7,856	14,232
Water Supply & Storage, Effluent Discharge	4,974	0	4,974
Electrical Power Supply	2,966	50	3,016
Infrastructure	2,116	155	2,271
Environmental Evaluation & Monitoring	2,441	0	2,441
Environmental Restoration Plan	3,130	0	3,130
Permits	2,428	0	2,428
Land	20,832	0	20,832
Owner’s Costs	36,908	143	37,051
Total Capital Cost	279,992	22,619	302,611
Bond, Working Capital & Closure Recoveries	0	-11,105	-11,105
Total	279,992	11,514	291,506

1.2.8 Operating Costs

Operating costs include cost estimates for all labor, consumables, utilities, contractor costs and equipment maintenance and repairs. PAH believes that the estimate is thorough, complete and reasonable for a project of this type located in Spain. An example of the annual operating costs, costs for Year 5 are presented in Table 1-4. Based on the DMT-Montan feasibility study, the Las Cruces project has the potential to be a low-cost producer. Cash operating costs per pound of copper produced are expected to average 0.33 euros over the life of the project, which would place the project among the lowest cost copper producers in the world. Cash operating costs include charges for mining ore and waste, processing ore through milling and leaching facilities, royalties and other cash costs. Charges for reclamation work done during mining operations are also included in cash costs. Cash operating costs do not include sustaining capital costs over the life of the mine or charges for reclamation work to be performed after production ceases. Based on the DMT-Montan feasibility study, sustaining capital costs are expected to average approximately 1.5 million euros per year, and will be funded by cash provided by operations. Post-production closure and reclamation expenses are estimated at approximately 43 million euros, excluding salvage values and other factors that could partially offset these expenses.

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TABLE 1-4

MK Resources Company

Las Cruces Project Technical Report

Total and Unit Operating Costs, Year 5

Area/Category	SUBTOTAL (1)	Contingency	TOTAL
TOTAL COSTS (euros 000’s)
General & administrative	6,366	438	6,804
Mining	15,064	753	15,817
Ore processing	24,087	1,233	25,320
TOTAL	45,517	2,424	47,941
UNIT COSTS (euros/tonne ore)
General & administrative	5.76	0.40	6.16
Mining	13.63	0.68	14.31
Ore processing	21.80	1.12	22.91
TOTAL	41.19	2.19	43.39
UNIT COSTS (euro cents/lb copper)
General & administrative	4.4	0.3	4.7
Mining	10.4	0.5	10.9
Ore processing	16.6	0.9	17.5
TOTAL	31.4	1.7	33.1
Tonnes ore milled (000’s)	1,105
Pounds of copper produced (000’s)	145,000

⁽¹⁾ Subtotal category includes: Labor, Power, Maintenance Parts, Oxygen, Sulfuric Acid, Natural Gas, Lime, Minor Consumables and Other

1.2.9 Project Execution and Schedule

Execution of the project will be divided between MK Resources Company/CLC and an engineering/Procurement/ Construction (EPC) contractor. MK Resources Company/CLC will have responsibility for the following functions:

• Permitting

• Financing

• Land acquisition

• Mine pre-stripping

• Environmental monitoring.

Execution of these functions will require most of the General & Administrative (G&A) staff be hired early in the project; about a half dozen of the eventually total about 50 have already been hired and are working on accomplishing the tasks listed. In addition to these personnel, MK Resources Company/CLC will hire a temporary group of about six who will monitor and manage the EPC contractor.

Pincock, Allen & Holt 9033.05 July 8, 2004 1.10

The EPC contractor will be responsible for all engineering, procurement, and construction work associated with the project, with the exception of the mine pre-stripping. The EPC work will include the following items:

• Groundwater dewatering and re-injection

• Stream diversion and earthworks

• Infrastructure

• Process plant

As currently envisaged, the EPC contractor will work on a lump-sum basis.

PAH considers the execution plan appropriate.

The project schedule projects a period of 2 ¹/4 years, from the awarding of the EPC contract to the start of production. The schedule is summarized in Figure 10-3. The schedule assumes that much of the work will begin before land acquisition and financing are complete. Given that the financing and the land occupation strategies are acceptable, PAH considers the schedule reasonable and that it may well be completed in less time than forecast.

Assuming that the project is considered to start when the EPC contract is awarded, the essential milestones of the project are projected to occur at the end of the months stated below:

Final permits approved:	Month 8
Financing in place:	Month 8
Land occupation complete:	Month 8
Dewatering/re-injection system
Initiate construction:	Month 3
Start operation:	Month 7
Stream diversions and earthworks
Initiate construction:	Month 6
Finish construction:	Month 25
Mine pre-stripping
Award contract:	Month 3
Start pre-stripping:	Month 9
Process plant
Award contract:	Month 0
Initiate construction:	Month 8
Start production:	Month 27
Full production:	Month 37
Infrastructure
Initiate construction	Month 2
Finish construction	Month 26

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1.2.10 Metal Marketing

PAH believes that the approximate 66,000 tonnes per year of copper produced will be easily placed in the metal markets. Spain does not mine enough copper to be self-sufficient in refined production and relies heavily on the import of copper concentrates. Due to the concentrate imports, Spain is able to achieve a balance between refined copper production and consumption. Europe, however, needs metal and offers a ready market.

1.2.11 Exchange Rate Effects

Total development and operating costs for the Las Cruces project could be affected by a number of factors, including the exchange rate of the euro relative to the U.S. dollar. Although the DMT-Montan feasibility study reflects an overall reduction in capital costs, the estimated capital costs in U.S. dollars based on the current exchange rate would be higher than our prior estimates due to a significant decline in the exchange rate of the euro relative to the U.S. dollar. Exchange rates fluctuate, and we cannot predict what exchange rates will be over the course of development and operation of the Las Cruces project.

1.2.12 Financial Model

PAH has reviewed the Financial Model (V4(R2) Fin woSunk SchII FX1.00.xls) prepared by MK Resources Company and found it to be complete and the inputs were accurate and reflected the project costs, production levels and schedule presented in the Feasibility Study and revised mine plans. A Base Case was prepared using a copper price of 0.95 euros per pound of copper, unleveraged (without financing), without sunk costs and with subsidies. Note that the long-range models assume one dollar per euro as the exchange rate.

PAH then developed a summary cash flow model based on MK Resources Company’s financial model as a check for accuracy and to perform sensitivity analyses. PAH’s cash flow model is presented in Table 22-2, found at the end of this report. All analyses assumed a constant euros basis with no inflation or escalation included for costs or metal price.

The Base Case analysis results indicate that the project would produce an IRR of 22.3 percent and an NPV at a 10 percent discount rate of 203.5 million euros. PAH also performed a sensitivity analysis of the cash flow model at plus and minus 10 percent variations on copper price, capital costs and operating costs. As is common with most mining projects, the project economics are most sensitive to metal prices.

A copper cathode premium of 44 euros per tonne of copper (or approximately 0.02 euros per pound of copper) has been included for 95 percent of the copper produced in Years 2 through 15. Year 1 assumes that 75 percent of cathode copper would receive the premium. The premium adds, on average, approximately 2.7 million euros of revenue per year, or a total of about 40.0 million euros over the life of the project.

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Base Case net cash flow (unleveraged) during the first five production years averages approximately 78 million euros per year. The initial capital expenditures of 280 million euros for development of the project should be repaid halfway through the third production year. The projected cash flow for the remaining ten production years will average about 64 million euros per year.

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2.0 INTRODUCTION AND TERMS OF REFERENCE

MK Resources Company, of Salt Lake City, Utah, engaged Pincock, Allen and Holt (PAH) to prepare a Technical Report covering the Las Cruces Project near Seville, Spain, to meet the requirements of Canadian National Instrument 43-101. This report reflects the most recent estimates of resources and reserves, which were developed and evaluated in the PAH Diligence Report “Independent Technical Audit of the Las Cruces Copper Project, Southern Spain” dated April 29, 2004.

The objective of this report is to summarize both the PAH Diligence Report and the underlying feasibility study developed by MK Resources Company, in compliance with Canadian regulations. This report meets the requirement for Canadian National Instrument 43-101 and conforms to the recommended structure for form 43-101F1 for Technical Reports.

2.1 Qualified Person and Participating Personnel

The principal author of this report is Gerald D. Crawford, a registered professional engineer in the states of Colorado, Nevada and Florida. Mr. Crawford has been closely involved with the project since Fall 2003, and had cooperated with other engineers for an earlier diligence report in 2001. Mr. Crawford visited the site in January 2004. Mr. Crawford has reviewed the majority of the mining plans and operating costs, as well as the production components of the cash flow.

Other participating individuals include:

• Mark G. Stevens, C.P.G. – Geology and Resource Model

• Nelson D. King – Metallurgy, Process and Infrastructure

• Richard Addison, P.E. – Process and Operating Costs

• Darrel L. Buffington, P.E. - Environmental and Permitting

• Mark G. Madden, P.E. – Environmental and Geotechnical

Mr. Madden has visited the site as well.

2.2 Terms and Definitions

The following acronyms and abbreviations are used throughout the report:

CLC - Cobre Las Cruces

MK - MK Resources Company

PAH - Pincock, Allen and Holt

Cu - Atomic Symbol for copper

Pincock, Allen & Holt 9033.05 July 8, 2004 2.1

The resource and reserve definitions applied in this report are developed according to the standards set out in appendix 43-101CP, “Canadian Institute of Mining, Metallurgy and Petroleum – Definitions Adopted by CIM Council, August 20, 2000.

2.3 Units

PAH has reviewed all operating and capital costs in terms of 4^th quarter 2003 euros (€), unless otherwise noted. Copper prices have been stated both in terms of US dollars and euros, at an exchange rate of 1 euro per $US.

All units are carried in metric units of meters, liters and tonnes of 1,000 kilograms. Grades are described in terms of percentages on a weight basis. Densities are typically declared in terms of specific gravity, as a ratio of measured weight of the substance divided by the equivalent volume of water.

2.4 Source Documents

The source documents for this report are summarized in Section 21.

Pincock, Allen & Holt 9033.05 July 8, 2004 2.2

3.0 DISCLAIMER

This report was prepared for MK Resources Company by the independent consulting firm of Pincock, Allen and Holt (“PAH”) and is based in part on information not within the control of either the company or the consultant. While it is believed that the information contained herein will be reliable under the conditions and subject to the limitations set forth herein, PAH cannot guarantee the accuracy thereof.

Neither PAH nor the authors of this report are qualified to provide extensive comment on legal issues associated with the Las Cruces Project. Assessment of these aspects has relied upon the legal review by the Spanish counsel of MK Resources.

PAH has not independently verified environmental, geotechnical or hydrological testwork conducted by MK Resources, its employees or contractors, as discussed within sections 1, 4, 17 and 22. PAH has relied upon the content of the DMT-Montan Feasibility Study as the basis of our analysis.

Similarly, PAH has not independently verified the metallurgical testwork conducted by Outokumpu as discussed in sections 1, 16 and 22.

No warranty or guarantee, be it express or implied, is made by the consultant with respect to the completeness or accuracy of the legal, environmental, metallurgical, mineral processing or stockpile resources referred to in this document. Neither PAH nor the authors of this report accept any responsibility or liability in any way whatsoever to any person or entity in respect of these parts of this document, or any errors in or omissions from it, whether arising from negligence or any other basis in law whatsoever.

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4.0 PROPERTY DESCRIPTION AND LOCATION

4.1 Location

The Las Cruces Project is located in Andulasia, southern Spain, about 20 kilometers northwest of the city of Seville. A location map is presented in Figure 4-1. Access to the project site is excellent by well-maintained all-weather paved roads. Rail service is available in Seville, as is an international airport with connections throughout Europe. Port facilities are available in Huelva approximately 80 km to the southwest.

Las Cruces is situated in the Guadalquivir River basin. The Guadalquivir basin is relatively flat and extends approximately 420 km from the Alcaraz and Segura mountain systems in the east to the Costa de la Luz on the Atlantic Coast in the west. The Guadalquivir basin is set between a long range of low mountains to the north (Sierra Morena) and a group of formations making up the Cordilleras Beticas mountain range to the south.

The geographic coordinates of the property are N37°30´, W06°06´.

4.2 Tenure

Riomin Exploraciones, S.A. (Riomin) discovered the Las Cruces deposit in 1994. Subsequent various investigations and studies resulted in the preparation of an initial feasibility study in 1998. MK Resources, through its wholly owned subsidiary, Cobre Las Cruces, S.A. (CLC), acquired 100 percent of the project from Riomin in September 1999. Leucadia National Corporation, a diversified investment and holding company, has a 72.8 percent share in MK Resources Company.

CLC has been granted a mining concession for subsurface minerals through Mining Concession No. 7532, granted by the Regional Ministry for Employment and Technological Development of the Province of Andalusia. Other permits and surface rights are addressed separately.

CLC conducted an Environmental Impact Study that was published in April 2001. A Declaration of Environmental Impact was issued May 9, 2002 and a Mining Concession was granted August 6, 2003.

Land ownership within Spain is procedurally clear, and is established via survey. A map of the property is attached as Figure 4-2 (8-1 from DR). The property lines (fence lines) identified on the map have been provided by surveying services contracted by CLC.

No mineral development has taken place within the project area. Currently the land is used only for farming and livestock grazing.

Pincock, Allen & Holt 9033.05 July 8, 2004 4.1

FIGURE 4-1

Pincock, Allen & Holt 9033.05 July 8, 2004 4.2

FIGURE 4-2

Pincock, Allen & Holt 9033.05 July 8, 2004 4.3

While CLC owns the mineral rights to the Las Cruces orebody, they do not currently own surface rights to the land above the orebody or certain adjacent land that may be necessary to develop and operate the project. In the aggregate, CLC may need to purchase or lease up to 954 hectares, or approximately 3.7 square miles of land and associated rights-of-way.

CLC will need to purchase or lease approximately 835 hectares of land in order to complete the final processing of some of the water permit applications and bring the mine into production. The additional 119 hectares of land is not required to begin production, but the additional land might be necessary to accommodate operations in later years. CLC leased approximately 50 hectares of the land necessary to bring the mine into production in 2005. In March 2004, CLC entered into an agreement to purchase 500 hectares of land that are necessary to bring the mine into production. As a result, CLC has secured approximately 550 hectares (approximately 65%) of the 835 hectares of land required to bring the mine into production.

CLC is negotiating with other landowners to acquire the remaining 35 percent of the initial land needed for the project, and has also initiated proceedings to cause the regional government to expropriate the land if negotiations with landowners are unsuccessful. Mineral rights and surface land are separately transferable under Spanish law. Expropriation is commonly used by holders of mineral rights in Spain, in accordance with government laws, to acquire surface land necessary for the extraction of minerals. This procedure could prove time-consuming and could significantly delay commencing construction activities. If it becomes necessary to expropriate the land, this procedure could significantly delay development of the project.

4.3 Mineral Zone Location

The Las Cruces property limits will fully contain the deposit when acquisitions are completed. The ore deposit is located on the eastern side of the property, at a depth of approximately 120 meters. The pit outline visible in Figure 4-2 provides an approximate location of the deposit.

4.4 Environmental Liabilities

CLC is a Greenfields operation, comprised of agricultural land well away from populated areas. No existing liabilities are known to exist. As the project is developed, CLC will incur reclamation liabilities that are clearly identified in the planning, with appropriate allowances within the cash flow.

4.5 Permitting

Permitting for the Las Cruces project will include both environmental approvals for operation of the mine, processing plant and mine waste disposal as well as land use permits and approvals. A discussion of the environmental permitting status is presented in Section 22.6.1. Land use permits include obtaining authorization for construction of crossings of the water supply pipeline over railways, highways, and a water supply canal, as well as various crossing of powerline, telephone line and underground utility line rights-of-way. Works and activity permits will be required from local community councils.

Pincock, Allen & Holt 9033.05 July 8, 2004 4.4

CLC is well along in obtaining the necessary permits and approvals. Outstanding land use permits are primarily pending final preparation and processing until purchase or lease of surface rights is complete.

4.6 Royalties and Encumbrances

The cash flow model includes a royalty of 1.5 percent payable to Rio Tinto, negotiated as part of the original property acquisition. The royalty is payable only on the portion of the copper price at or greater than $0.80 per pound. No other royalties have been identified. No encumbrances specific to the property have been discussed within the documents provided to PAH. MK Resources Company, the owner, has identified certain loans from majority shareholder Leucadia in the securities filings; however, these are apparently to MK Resources rather than CLC.

Pincock, Allen & Holt 9033.05 July 8, 2004 4.5

5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES AND INFRASTRUCTURE

5.1 Topography, Elevation and Vegetation

The regional climate is Mediterranean and semi-arid with hot, dry summers (May through September) and warm, wet winters. Prolonged droughts periodically occur. Operations will be possible year-round.

Wind is generally from the southwest and brings hot and humid air into the area that generates autumn and winter rainstorms, and intense spring storms. Summer maximum temperatures average about 30 degrees Celsius (^oC), with the maximum occasionally reaching 40 ^oC in July and August. Winter maximum temperatures average about 15 ^oC and the minimum temperature may drop to as low as –5 ^oC in December and January.

Mean annual rainfall is about 550 millimeters and ranges from about 300 to 1,000 millimeters. The heaviest one-day rainfall on record was 100 millimeters.

The topography around the project is characterized by gently rolling hills of arable land. Elevations at the site range from 15 to 45 meters above sea level.

The site is characterized as having a moderate seismic activity. A seismic potential of a magnitude 5.4 (Richter Scale) event with a recurrence interval of 250 years has been defined for the Seville area by the Instituto Geografic Nacional and the Center Internacional de Sismologia.

5.2 Access and Transport

Access to the project site is excellent by well-maintained all-weather paved roads. Highway N-630 presently is in service to the east side of the property, while Highway SE-520 crosses the southern side of the property. The A66 freeway is presently under construction.

Rail service is available in Seville, with high-speed passenger rail service to Madrid. Seville offers an international airport with connections throughout Europe within 30 minutes of the site. Port Facilities are available in Huelva approximately 80 km to the southwest.

Seville (Sevilla, pop. 700,000), a major Spanish city, is located approximately 20 km to the south from the property, with a large and well-educated population. The village of Gerena (pop. 5,600) is located approximately 4 km to the northwest of the property.

Pincock, Allen & Holt 9033.05 July 8, 2004 5.1

5.3 Infrastructure

The power supply during construction will be provided through connections to the 15 kV overhead lines that end at the sites of the Render Sur factory and the Seroncillo farm, with additional capacity from a 400-volt diesel generator with a 1 MW capacity. During operation, the electrical power will be supplied through a new 220 kV branch line with a length of approximately 4.2 km, to be constructed for the Las Cruces project. On the Las Cruces project site a switchyard and a 40 MVA transformer will distribute the electric power to the different project locations.

The plant will require 60 liters per second of process water, 10 percent of which can be water supplied from contact water and the remainder of which must be supplied from elsewhere. The remaining required process water will be pumped 15 km from the San Jerónimo sewage treatment plant, located to the northwest of Seville. The process water will be pumped through a pipeline to the primary supply pond in the west of the Las Cruces project area. Water from sewage treatment plant will not be available during the dry season; therefore, the primary supply pond will require sufficient storage capacity (approximately 1.3 million cubic meters) to store water during the five-month dry season. Potable water will come from a well drilled into the Posadas-Niebla aquifer. CLC has applied to CHG for the concession to pump this water.

5.4 Sufficiency of Surface Rights

CLC has made a determined effort to minimize the land requirements for the property, and has developed a detailed plan for the land use at the property including adequate room for waste dumps, tailings ponds and pit slopes. PAH feels that sufficient room is available for the plans as currently presented.

It should be noted that CLC has the right to expropriate land under Spanish mineral law, but has avoided doing so for reasons of public relations. Nonetheless, the option does exist.

Pincock, Allen & Holt 9033.05 July 8, 2004 5.2

6.0 HISTORY

The Las Cruces deposit was original discovered by Rio Tinto in 1994, as a result of drilling on a gravimetric survey anomaly. Rio Tinto drilled the deposit between 1994 and 1999 and prepared a feasibility study in 1998.

In 1999, MK Resources Company (formerly known as MK Gold Company) purchased the project and established Cobre Las Cruces S. A. (CLC) as the local Spanish subsidiary company. The mining concession was granted in August 2003.

Bechtel, an international engineering and construction company, completed an independent feasibility study for the Las Cruces project in 2001. The Bechtel feasibility study incorporated results of an environmental impact study completed by FRASA Ingenieros Consultores, a team of national and international environmental engineering experts based in Madrid.

In 2003, DMT-Montan prepared a new independent feasibility study based upon Outokumpu technology. The DMT-Montan feasibility study incorporates the requirements from the DEI, the mining concession and various water permits into the development plan for the Las Cruces project. The DMT-Montan feasibility study has been reviewed by Pincock, an independent engineering company. In its report, Pincock stated that it did not identify any deficiencies that would preclude the Las Cruces project from meeting the designated production and cost objectives presented in the DMT-Montan feasibility study.

Pincock, Allen & Holt 9033.05 July 8, 2004 6.1

7.0 GEOLOGIC SETTING

7.1 Regional Geologic Setting

The Las Cruces deposit occurs near the eastern end of the Iberian Pyrite Belt, a 250-kilometer long and 40-kilometer wide geologic belt that extends eastward from southern Portugal into southern Spain. The belt is host to more than 100 mineral deposits, some of which were exploited for metals as long ago as pre-Roman times. Mineralization at Las Cruces, as in most other Iberian Pyrite Belt deposits, consists of syngenetic massive sulfides containing polymetallic mineralization. The massive sulfide is hosted by late Devonian to early Carboniferous Period volcanic and sedimentary rocks, deposited in a submarine setting within a narrow and relatively shallow intracontinental sea, characterized by bimodal volcanism and sedimentation.

Subsequent tectonism during the Hercynian Orogeny of the late Paleozoic Era resulted in the general uplift of the region, with variable deformation and faulting. This was followed by weathering and erosion of the overlying host rocks during an indeterminate period between the Hercynian Orogeny and the Miocene Epoch. Late in the period of erosion and weathering, the upper part of the original massive sulfide deposit was exposed. Near surface oxidation of the sulfide minerals to an iron oxide gossan occurred, along with the downward transport of copper leached from the oxidized gossan zone. The transported copper was precipitated and replaced unoxidized primary massive sulfide at depth, forming a secondary enrichment of the sulfide deposit. Relatively immobile gold and silver metals remained in the oxidized gossan, with local enrichments.

Transgression by a shallow sea in the middle Tertiary Period (Miocene Epoch) interrupted the weathering and oxidation process on the deposit. Subsequent submarine deposition buried the deposit under 100 to 150 meters of sandstone and calcareous mudstone (marl). Subsequent marine regression in Pliocene times again resulted in the subaerial exposure of the region. The current surface of the project area is on the calcareous mudstone sequence, with the Las Cruces deposit buried at depth below this sequence.

The sandstone unit that occurs at the base of the younger sedimentary sequence overlying the paleoerosion surface is an important regional aquifer. This aquifer consists of a basal conglomerate overlain by semi-consolidated sandstone and averages approximately five meters thick over the deposit area.

7.2 Deposit Geologic Setting

The Las Cruces massive sulfide is located within a host sequence of volcanic and sedimentary rocks, with a shale unit generally present as a 10 to 20 meter wide envelope around the massive sulfides. Mineralization occurs in one massive sulfide horizon that has a general strike to the east and a dip to the north at about a 35-degree angle, with a gradation change to the west to a northwesterly dip at 30 degrees. The overall dimensions of the massive sulfide deposit are 1,000 meters along strike,

Pincock, Allen & Holt 9033.05 July 8, 2004 7.1

500 meters or more down dip (still not completely drill delineated), and up to 100 meters thick (averages 30 to 40 meters). To the west, the massive sulfide mineralization thickens and is then truncated by a roughly north-south trending fault; with attempts to locate the offset mineralization not having been successful so far. Along strike to the northeast, the deposit is not completely drill delineated, however the grade of the mineralization is relatively low. Due to supergene enrichment processes in pre-Miocene time, secondary sulfide mineralization formed a generally horizontal zone in the up-dip part of the primary massive sulfide. Figure 7-1 shows a geology-drill hole cross section looking west and reflects the closely spaced drilling of a local statistical cross. Figure 7-2 shows a plan view of deposit.

The pre-Miocene paleosurface that cuts the up-dip part of the massive sulfide deposit is planar, with minor irregularities and a gentle dip to the south. The central and western part of the deposit, below the paleosurface, was most affected by the pre-Miocene weathering and oxidation. The uppermost part of the massive sulfide mineralization has been completely oxidized to an iron oxide gossan, with local enrichment of gold and silver. The gossan is also enriched in lead, reportedly as the mineral galena. Infilling of the porous gossan with variable carbonate occurred with the subsequent sedimentary marl deposition. PAH’s review found that the gossan is generally 10 to 20 meters in vertical thickness, with grades from all gossan drill hole samples averaging 5.9 g/t gold (up to 353 g/t Au) and 98 g/t silver (up to 1,733 g/t Ag). Average copper grade in the gossan is 0.20 percent. Apparently associated with the gossan formation was the strong silicification of the host rock in the hanging wall immediately above the massive sulfide deposit, containing erratic gold and silver enrichment.

Underlying the iron oxide gossan is the secondary sulfide zone in which the primary massive sulfide mineralization has been secondarily enriched by the downward migration and precipitation of copper leached from the gossan zone. The secondary sulfide mineralization is the most important economically, and the focus of the recent feasibility study. Secondary sulfide enrichment occurs in a roughly horizontal zone within the up dip part of the original primary massive sulfide deposit. The zone of secondary sulfide enrichment averages about 40 meters thick below the gossan, but in areas of intense fracturing secondary sulfide may extend up to 60 meters thick. The secondary sulfide mineralization is gradational downward into the primary sulfide zone. The footwall contact of the secondary mineralization with the underlying host rocks is usually sharp, as a result of the clay altered shales and volcanics being relatively impermeable. This permeability change tended to force secondary mineralization along the base of the massive sulfides, locally producing lenses of secondary mineralization that follow the contact zone down dip into the primary sulfides.

Secondary sulfide mineralization has been superimposed upon and has partially replaced the original primary sulfide mineralization below the gossan zone, increasing the quantity of copper minerals present. Secondary sulfide mineralization consisted predominantly of the deposition of chalcocite and to a lesser extent bornite and covellite by partial replacement of pyrite and other sulfides, and by infilling into any open space resulting from fracturing or original porosity. Chalcocite ranges from very fine grained (sooty) dark gray coatings and relatively unconsolidated

Pincock, Allen & Holt 9033.05 July 8, 2004 7.2

FIGURE 7-1

Pincock, Allen & Holt 9033.05 July 8, 2004 7.3

FIGURE 7-2

Pincock, Allen & Holt 9033.05 July 8, 2004 7.4

intergrowths to disseminated consolidated grains to consolidated veins and replacement bands. Chalcopyrite content is minor. PAH’s review found that grades from all secondary sulfide drill hole samples averaged 7.0 percent copper (up to 39.4% Cu), 0.2 percent zinc (up to 36.9% Zn), 0.6 percent lead (up to 40.0% Pb), 0.5 g/t gold (up to 62.9 g/t Au), and 26 g/t silver (up to 1,472 g/t Ag).

The secondary sulfide mineralization was divided into three main sub-lenses, and three minor sub-lenses. The boundaries between the six types are complex and gradational. The High Copper – High Density (HCH) lens consists of secondary enrichment of dense primary massive sulfides. The High Copper – Low Density (HCL) lens consists of secondary enrichment of primary semi-massive sulfides, footwall stockworks, or footwall lithologies of lesser density. The High Copper lens Number 4 (HC4) is a spatially distinct lens adjacent to the HCH secondary mineralization that follow the contact zone down dip into the primary sulfides zone. In addition, three minor sub-lenses occur: the High Copper Intrusive (HCI), the High Copper Footwall (HCF), and the Medium Copper Low Density (MCL); all three of which were grouped with the main lenses for resource estimation purposes.

Original primary massive sulfide mineralization (CZ lens) contains massive to semi-massive sulfide minerals (generally more than 80 percent sulfide). Pyrite is the predominate sulfide mineral, with lesser finely intergrown sphalerite, galena, and chalcopyrite, as well as minor enargite, tennantite, and tetrahedrite. PAH’s review found that grades from all primary sulfide drill hole samples averaged 3.2 percent copper (up to 18.9% Cu), 1.1 percent zinc (up to 11.9% Zn), 0.4 percent lead (up to 5.4% Pb), 0.5 g/t gold (up to 1.63 g/t Au), and 20 g/t silver (up to 174 g/t Ag). Most of the primary massive sulfide mineralization occurs down-dip to the northwest and are not considered for the current feasibility study. In the footwall below the primary massive sulfide is a stockwork of interconnected pyrite veins and veinlets, with local higher grades of copper and zinc.

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8.0 DEPOSIT TYPES

The Las Cruces deposit occurs near the eastern end of the Iberian Pyrite Belt, a 250-kilometer long and 40-kilometer wide geologic belt that extends eastward from southern Portugal into southern Spain. The belt is host to more than 100 mineral deposits, some of which were exploited for metals as long ago as pre-Roman times. Mineralization at Las Cruces, as in most other Iberian Pyrite Belt deposits, consists of syngenetic massive sulfides containing polymetallic mineralization. The massive sulfide is hosted by late Devonian to Early Carboniferous Period volcanic and sedimentary rocks, deposited in a submarine setting within a narrow and relatively shallow intracontinental sea, characterized by bimodal volcanism and sedimentation.

Volcanogenic massive sulfides are deposited as submarine hot spring exhalative metal sediment accumulations, typically in tectonic areas of active submarine volcanism, including rift spreading centers and island arc subduction zones. In these geologic environments the volcanic rocks may be interlayered to a variable degree with contemporaneous volcanic sediments. Massive sulfide bodies are deposited as metallic sulfide sediments, with a variable component of volcanic sediment and silica, in conformable layers and blanket surrounding the exhalative center. Massive sulfide layers consist predominantly of pyrite and/or pyrrhotite, with variable base metal sulfides commonly including chalcopyrite, sphalerite, and galena. Subsequent tectonism may result in variable uplift, folding, and/or faulting of the original massive sulfide deposit.

The Las Cruces massive sulfide was subjected to post-depositional tectonism during the Hercynian Orogeny of the late Paleozoic Era resulting in the general uplift of the region, with variable deformation and faulting. This was followed by weathering and erosion of the overlying host rocks during an indeterminate period between the Hercynian Orogeny and the Miocene Epoch. Late in the period of erosion and weathering, the upper part of the original massive sulfide deposit was exposed. Near surface oxidation of the sulfide minerals to an iron oxide gossan occurred, along with the downward transport of copper leached from the oxidized gossan zone. The transported copper was precipitated and replaced unoxidized primary massive sulfide at depth, forming a secondary enrichment of the sulfide deposit. Relatively immobile gold and silver metals remained in the oxidized gossan, with local enrichments.

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9.0 MINERALIZATION

9.1 Secondary Sulfide Mineralization

The pre-Miocene paleosurface that cuts the up-dip part of the massive sulfide deposit is planar, with minor irregularities and a gentle dip to the south. The central and western part of the deposit, below the paleosurface, was most affected by the pre-Miocene weathering and oxidation. The uppermost part of the massive sulfide mineralization has been completely oxidized to an iron oxide gossan, with local enrichment of gold and silver. The gossan is also enriched in lead, reportedly as the mineral galena. Infilling of the porous gossan with variable carbonate occurred with the subsequent sedimentary marl deposition. PAH’s review found that the gossan is generally 10 to 20 meters in vertical thickness, with grades from all gossan drill hole samples averaging 5.9 g/t gold (up to 353 g/t Au) and 98 g/t silver (up to 1,733 g/t Ag). Average copper grade in the gossan is 0.20 percent. Apparently associated with the gossan formation was the strong silicification of the host rock in the hanging wall immediately above the massive sulfide deposit, containing erratic gold and silver enrichment. While there is potential for economic value within the gossan, no studies have been completed. The project does not consider the gossan to have economic value.

Underlying the iron oxide gossan is the secondary sulfide zone in which the primary massive sulfide mineralization has been secondarily enriched by the downward migration and precipitation of copper leached from the gossan zone. The secondary sulfide mineralization is the most important economically, and the focus of this feasibility study. Secondary sulfide enrichment occurs in a roughly horizontal zone within the up dip part of the original primary massive sulfide deposit. The zone of secondary sulfide enrichment averages about 40 meters thick below the gossan, but in areas of intense fracturing secondary sulfide may extend up to 60 meters thick. The secondary sulfide mineralization is gradational downward into the primary sulfide zone. The footwall contact of the secondary mineralization with the underlying host rocks is usually sharp, as a result of the clay altered shales and volcanics being relatively impermeable. This permeability change tended to force secondary mineralization along the base of the massive sulfides, locally producing lenses of secondary mineralization that follow the contact zone down dip into the primary sulfides.

Secondary sulfide mineralization has been superimposed upon and has partially replaced the original primary sulfide mineralization below the gossan zone, increasing the quantity of copper minerals present. Secondary sulfide mineralization consisted predominantly of the deposition of chalcocite and to a lesser extent bornite and covellite by partial replacement of pyrite and other sulfides, and by infilling into any open space resulting from fracturing or original porosity. Chalcocite ranges from very fine grained (sooty) dark gray coatings and relatively unconsolidated intergrowths to disseminated consolidated grains to consolidated veins and replacement bands. Chalcopyrite content is minor. PAH’s review found that grades from all secondary sulfide drill hole samples averaged 7.0 percent copper (up to 39.4% Cu), 0.2 percent zinc (up to 36.9% Zn), 0.6 percent lead (up to 40.0% Pb), 0.5 g/t gold (up to 62.9 g/t Au), and 26 g/t silver (up to 1,472 g/t Ag).

Pincock, Allen & Holt 9033.05 July 8, 2004 9.1

The secondary sulfide mineralization was divided into three main sub-lenses, and three minor sub-lenses. The boundaries between the six types are complex and gradational. The High Copper – High Density (HCH) lens consists of secondary enrichment of dense primary massive sulfides. The High Copper – Low Density (HCL) lens consists of secondary enrichment of primary semi-massive sulfides, footwall stockworks, or footwall lithologies of lesser density. The High Copper lens Number 4 (HC4) is a spatially distinct lens of HCH secondary mineralization that follow the contact zone down dip into the primary sulfides zone. In addition, three minor sub-lenses occur: the High Copper Intrusive (HCI), the High Copper Footwall (HCF), and the Medium Copper Low Density (MCL); all three of which were grouped with the main lenses for resource estimation purposes.

9.2 Primary Sulfide Mineralization

Original primary massive sulfide mineralization (CZ lens) contains massive to semi-massive sulfide minerals (generally more than 80% sulfide). Pyrite is the predominate sulfide mineral, with lesser finely intergrown sphalerite, galena, and chalcopyrite, as well as minor enargite, tennantite, and tetrahedrite. PAH’s review found that grades from all primary sulfide drill hole samples averaged 3.2 percent copper (up to 18.9% Cu), 1.1 percent zinc (up to 11.9% Zn), 0.4 percent lead (up to 5.4% Pb), 0.5 g/t gold (up to 1.63 g/t Au), and 20 g/t silver (up to 174 g/t Ag). Most of the primary massive sulfide mineralization occurs down-dip to the northwest and are not considered for the current feasibility study. In the footwall below the primary massive sulfide is a stockwork of interconnected pyrite veins and veinlets, with local higher grades of copper and zinc. The primary sulfide (CZ) is not considered as part of the current resource estimate.

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10.0 EXPLORATION

10.1 Deposit Exploration

The Las Cruces deposit was original discovered by Rio Tinto in 1994 as a result of drilling on a gravimetric survey anomaly. Rio Tinto drilled the deposit between 1994 and 1999. A total of 277 core holes, for 82,352 meters, were drilled by Rio Tinto, for exploration and deposit definition purposes and were included in the drill hole database for resource estimation (“CR” series of drill holes). Another 106 holes, for 24,279 meters, were drilled for geotechnical, hydrogeological and metallurgical purposes. The details of the drilling programs are discussed more completely in the next section on drilling.

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11.0 DRILLING

11.1 Core Drilling

A total of 277 core holes, for 82,352 meters, were drilled by Rio Tinto between 1994 and 1999, for exploration and deposit definition purposes and were included in the drill hole database for resource estimation (“CR” series of drill holes). Another 106 holes, for 24,279 meters, were drilled for geotechnical, hydrogeological and metallurgical purposes. Drilling of the “CR” series of holes was conducted by three different companies: Almaden (about 55 percent of holes), Insersa (about 40 percent of holes), and Iberica (about 5 percent of holes). Most of these holes were drilled vertically (70 percent of holes), with the remainder drilled at an angle (30 percent). Hole collar locations have been surveyed by a professional surveyor using a theodolite. All core holes were surveyed down-hole using a multi-shot or single-shot instrument.

Drilling of the “CR” holes by all companies consisted of rotary tricone drilling through the marl and sandstone overlying the deposit, followed by core drilling in the underlying host rock sequence. Core drilling was generally conducted to obtain larger diameter core for analysis, metallurgical purposes, and for reference. Core samples were taken at nominal 1-meter lengths (+50 percent) based on geological or mineralogical sampling breaks. The core was logged for geologic and mineralogic information and the percent core recovery recorded. PAH notes, however, that the first 23 holes (CR001 to CR023) did not have the percent core recovery recorded.

Drilling of the secondary sulfide deposits is on a variable grid with a drill hole spacing of 50 meters by 50 meters. Two statistical crosses were drilled, a complete cross in the western part of the deposit and a less complete cross in the central part of the deposit, with a nominal 12.5-meter spacing between these holes and inclined at a 60-degree angle. PAH believes that this drilling provides adequate coverage of the deposit, with some in-fill recommended near holes in which core recovery was an issue or were wider spacing occurs. IMC expressed a similar opinion in the 2001 Bechtel Feasibility and CLC concurs with these recommendations. Figure 11-1 shows the drill hole locations with some of the drill holes marked to indicate: 1: holes for which low core recovery from the secondary sulfide (<80%) excluded the entire drill hole from being used in the resource modeling; 2: holes for which low core recovery from the secondary sulfide (80%) excluded more than one half of the secondary sulfide; and 3: initial holes for which core recovery from the secondary sulfide was not recorded so a 100 percent core recovery was assigned.

Drilling of the secondary sulfides consisted of 88 percent PQ size core (85 mm core diameter), 9 percent HQ size core (64 mm core diameter), and 3 percent NQ size core (48 mm core diameter). Average copper grade for the three core sizes shows that the larger PQ core samples average about 15 percent higher in copper than the smaller HQ sized core and that the HQ sized core averages about 65 percent higher in copper than the smaller NQ sized core. Because of the increasing core surface area/core volume with decreasing core size, this provides evidence of the preferential loss of copper minerals from the core surface.

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FIGURE 11-1

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11.2 Reverse Circulation Drilling

Eighteen reverse circulation holes (311 mm diameter) were drilled in 1998 as twins of original PQ sized core holes as checks on the sampling and to provide material for metallurgical testwork. The reverse circulation sample results compared poorly on a sample-by-sample basis with the original core hole grades and were generally lower grade. On average, the reverse circulation samples were 43 percent lower in grade than the adjacent core hole (4.46 %Cu compared to 6.41 percent cu in the core holes). It is believed by CLC that the differences are attributable to short-range lateral variations in grade and also that significant amounts of fine-grained chalcocite were lost during the reverse circulation drilling and as a result the grades are not representative. These holes have not been included in the drill hole database for resource estimation and so these holes are not considered further in the resource modeling and estimation process.

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12.0 SAMPLING METHOD AND APPROACH

12.1 Core Sampling

Core drilling was generally conducted to obtain larger diameter core for analysis, metallurgical purposes, and for reference. Core samples were taken at nominal 1-meter lengths (+50 percent) based on geological or mineralogical sampling breaks. The core was logged for geologic and mineralogic information and the percent core recovery recorded. PAH notes, however, that the first 23 holes (CR001 to CR023) did not have the percent core recovery recorded.

Drilling of the secondary sulfides consisted of 88 percent PQ size core (85 mm core diameter), 9 percent HQ size core (64 mm core diameter), and 3 percent NQ size core (48 mm core diameter). Average copper grade for the three core sizes shows that the larger PQ core samples average about 15 percent higher in copper than the smaller HQ sized core and that the HQ sized core averages about 65 percent higher in copper than the smaller NQ sized core. Because of the increasing core surface area/core volume with decreasing core size, this provides evidence of the preferential loss of copper minerals from the core surface.

HQ diameter core was drilled from hole CR01 to CR028 and was largely halved for an analytical sample. Holes from CR029 to CR075 were largely drilled with larger PQ sized core and halved for an analytical sample. Holes from CR076 were largely drilled with larger PQ sized core and quartered for an analytical sample. As shown in Table 12-1, of the total of 4,407 secondary sulfide core samples, of which 231 were sampled in their entirety (5 percent), 632 samples consisted of a half of the core (14 percent), and 3,544 samples consisted of a quarter of the core (81 percent). As a result, the predominant sample by far was PQ sized core for which the sample consisted of a quarter of the core (3,466 samples or 79 percent of all of the secondary sulfide core). PAH finds that the core sample used generally provides an adequate sized sample. PAH notes that one quarter of the larger PQ sized core ideally provides a sample comparable in size to one half of the smaller HQ sized core. PAH questions the comparability of quarter samples of NQ sized core, as quarter samples of NQ sized core ideally only provides about 30 percent of the material of quarter samples of PQ sized core. PAH notes, however, that there are a relatively small number of this sample size.

TABLE 12-1

MK Resources Company

Las Cruces Project Technical Report

Core Sample Size Distribution

Sample Size	NQ Core (48 mm diameter)	HQ Core (64 mm diameter)	PQ Core (85 mm diameter)	All Core Sizes
Entire Core	0	0	231	231
Half Core	88	347	197	632
Quarter Core	26	52	3,466	3,544
All Sample Sizes	114	399	3,894	4,407

Note: Results are for the drill hole database provided to PAH.

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12.2 Core Recovery

Core recovery from the main mineralized lenses (HCH, HCL, and HC4) ranges from 0 to 100 percent, with an average of about 80 percent. With the application of an 80 percent core recovery cutoff, the average recovery was raised to about 93 percent. Geologic observation during the logging of the Las Cruces core provides good evidence of the variable loss of core material due to the abrasion and erosion of friable and sooty, chalcocite-bearing mineralization. The apparent preferential loss of this material during drilling and sampling has tended to decrease the relative content of copper minerals, with a proportionally less decrease in the relative content of the other associated minerals, including pyrite, quartz, sphalerite, and/or galena. As a result of this, the copper grade in the actual analyzed sample of the recovered core tends to be less than in the copper grade in the original in-place location. For an independent opinion, CLC contracted Dr. Stephen Henley of IMC (UK) to evaluate the issue and he came to a similar conclusion. Independent Mining Consultants, however, concluded in the 2001 feasibility study that higher-grade copper occurs preferentially in more competent rocks, whereby chalcocite acts as cement holding fragments together and improving recovery. Previously, Rio Tinto in their 1998 feasibility study, came to a similar conclusion to that of CLC and DMT in the current feasibility.

PAH finds that the majority of the geologic and statistical data provide evidence for the preferential loss of chalcocite, although reduction in copper grade is not a uniform systematic statistical relationship, but is somewhat irregular due to chalcocite distribution both as more durable masses and elsewhere as friable aggregates. It appears that both high core recovery and low core recovery zones have been subject to variable chalcocite loss. As such, core recovery itself may provide a good general indication that original in-place copper grades are higher than reported, but are not an absolute indication on a sample-by-sample basis that the original in-place copper grade was higher. The effect of the preferential loss of chalcocite tended to increase the difference between original in-place copper grade and the actual assayed copper grade at progressively lower core recoveries. This is significant in terms of generating a reasonable resource model, resulting in the question of what sample core recoveries result in reliable enough copper grades (see subsequent discussion in resource modeling subsection).

Samples of a nominal 1-meter length (+50 percent) were taken for analysis. The core (PQ sized core for the majority of samples) was sawed into quarters, with one quarter used for the chemical analysis, while one quarter was stored in the core box for reference. The remaining material was stored in a commercial freezer in Seville to minimize sample oxidation, for use in subsequent metallurgical testing.

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13.0 SAMPLE PREPARATION, ANALYSES AND SECURITY

13.1 Sample Preparation and Analyses

Sample preparation for up to hole CR128 was conducted at Anamet Laboratory in Avonmouth, United Kingdom. After this, a preparation facility was set up on site, with samples from CR129 to CR277 prepared on site under Anamet supervision, with the pulp then sent for analysis. For routine chemical analysis, the quartered core sample (PQ sized core for the majority of samples) was crushed to minus 2-millimeter size and then split into 250-gram quantities. The 250-gram samples were then pulverized to approximately to minus 200 mesh (-74 microns) for analysis. Coarse reject material was also stored in a freezer.

Most of the assaying was done at the Anamet Laboratory (Rio Tinto) in Avonmouth, United Kingdom. From hole CR257 to CR277 (the last 20 holes), OMAC Laboratory was used as the primary laboratory because the Anamet Laboratory was no longer available. Analysis for copper, lead, zinc, silver was by digestion of 0.5 grams of pulverized material in a four-acid solution (HCl, nitric, HF, and perchloric) with an atomic absorption spectophotometery determination. Gold analysis was generally by fire assay on 20 grams of pulverized material. For drill holes CR001 to CR081 and CR082 to CR106, a fire assay was conducted only if an initial atomic absorption spectophotometery determination indicated gold above 0.5 g/t or 0.25 g/t, respectively. Sulfur was analyzed by a Leco furnace procedure. Other trace elements were also analyzed, including bismuth, cadmium, tin, arsenic, antimony, barium, and mercury. PAH believes that the analytical procedures are generally reasonable, but that for gold it is more standard to use a 30-gram pulp for fire assay in order to provide more representative results. For potentially coarser gold particles present in the gossan zone, there is some evidence that the analytical protocol may underestimate the gold grade, which is being evaluated further by CLC. All logging and sample data were entered into an Access database, with analytical data imported from electronic files from the laboratory.

13.2 Density Testwork

Density tests were generally conducted by Rio Tinto on every assay sample in the Paleozoic age host rocks and mineral deposit, for a total of 15,787 tests. Density was measured on whole core prior to splitting, by measuring the sample weight in air and in water, with the density given by the weight in air divided by the difference between the weight in air and in water. This method works well for competent pieces of core, but is difficult for fragmented pieces. Where this was the case, a representative sub-sample was taken and shaped in such a way as to not allow trapped air to be caught under it. To insure quality control, a standard weight was measured and no weighing bias was found by Rio Tinto.

PAH notes that the samples were not sealed to preserve natural voids and the samples were not dried prior to measurement, as is standard practice. Due to one or two day delays in measuring core density, DMT has stated that the bulk of the contained moisture was evaporated and that the

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cores samples were considered to effectively be dry, except for water in sealed pores which will have a minimum effect of resource calculations. Rio Tinto checked the effect of moisture content for 219 pieces of core by measuring dry core density (naturally dry from two year storage) and then measuring wet core density (submerged in water for 3 days). The difference between dry and wet density was about a 1 percent difference, which PAH concurs with DMT that it does not believe is very significant.

A potentially greater source of error is related to open vugs and pore space, which fills with water immediately upon submersion during testing. Geologists and hydrologists working on the project have visually estimated that permeable porosity in the mineralization is about 3 percent, which is factored into the resource tonne totals. Larger vugs were reported by Rio Tinto to be insignificant and they estimated the open vug space to be a few percent, which is not factored into the resource tonne totals. For the HCH/HCL lens, the average core density was 3.62 tonnes per cubic meter, which was subsequently factored during the resource tabulation using a 3-percent porosity factor (average adjusted density of 3.51 tonnes per cubic meter).

The iron oxide rich gossan directly overlying the secondary sulfide mineralization is described as having a crumbly nature that has produced much broken core and core loss. DMT states that it is likely that a certain amount of void space is present, but is quite variable and difficult to estimate. For the gossan, the average core density was 2.46 tonnes per cubic meter, which included an average 5.1 percent moisture (average dry density of 2.34 tonnes per cubic meter).

For the Miocene age marl and sandstone overlying the deposit, density and moisture content was conducted on geotechnical core. The core was sealed in wax to preserve the moisture content of the sample. Moisture content was determined and then the dry density was determined. The average core density of the marl was 2.02 tonnes per cubic meter, which included an average 27 percent moisture (average dry density of 1.58 tonnes per cubic meter) based on 88 samples. The average core density of the sandstone was 2.27 tonnes per cubic meter, which included an average of 19 percent moisture (dry density of 1.74 tonnes per cubic meter) based on three samples.

PAH finds that the density determinations have been conducted in an acceptable manner and that the results are representative of the various rock types present.

13.3 Sample Security

PAH understands that reasonable and customary procedures were employed for the security of the samples. Samples were maintained under the control of the site geologists, shipping agents, and the analytical laboratory.

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14.0 DATA VERIFICATION

14.1 Assay Checks

Routine quality control procedures were used by the Anamet Laboratory throughout the analytical program, including the use of standard samples, duplicate checks on coarse reject materials, and duplicate checks on pulverized material. Check assays were conducted by OMAC for copper, lead, zinc, silver, and gold. The analytical procedure used a two-acid solution (HCl and nitric) for copper, lead, and zinc, which PAH notes is not as strong of an acid. For silver, an aqua-regia digestion was used. For gold a 30-gram pulp was used. All other analytical procedures were the same as Anamet Laboratory. Two reference standard samples consisted of a high grade and low-grade material from Las Cruces and were included in the sample submissions to Anamet Laboratories. The data from the quality control checks did not indicate any significant bias or quality control issues. PAH notes that a visit was not made to the Anamet Laboratory to see the operation firsthand, nor is PAH familiar with the general historical performance of the facility.

A second core sample by Rio Tinto was submitted to Anamet Laboratories for about one out of 20 samples (5 percent) as a check of sample grade reproducibility (1,223 check samples). The checks were conducted on a second quarter of the drill core (PQ sized core for the majority of samples) and as such reflects a combination of natural grade variability across the core, sample preparation error, and analytical variance. The results showed no significant copper grade bias between the original and the check; however, a fair amount of variability existed, believed to largely be the result of the natural irregular character of the chalcocite enrichment.

A round robin program, consisting of the analysis of 10 samples by five different laboratories, was conducted by Rio Tinto to assess variability between the labs. Some variability was observed in copper grades between labs, but in general, the results show that the Anamet copper results are comparable to those from the other laboratories.

An independent check was conducted in 2000 by Independent Mining Consultants, as part of the previous work for the 2001 Bechtel feasibility study. Fifty samples were randomly selected for copper and gold analysis from remaining core stored in the Sevilla refrigeration facility. The core was crushed to minus 2 centimeters and a 1,200-gram split sent to Actlabs-Skyline Laboratories in Tucson, Arizona. The results for copper showed these check samples to compare reasonably with original grade values. The results for gold were less consistent, with Actlabs-Skyline Laboratory consistently higher than the Rio Tinto values, generally on the order of 30 percent or so.

PAH finds that Rio Tinto had a reasonable sampling and assaying protocol in place for the analysis of drill hole samples, complete with a reasonable quality control program. PAH notes that the quality control program did not incorporate “blank” samples known to have no grade; however, this is considered a minor issue. Check assays indicate reasonable analytical precision (repeatability) and analytical accuracy (comparability to know reference material), but was affected by natural variability of the sample material.

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14.2 PQ Twin Drilling

Rio Tinto drilled four PQ sized core holes (CR-265, CR 266, CR 267, and CR268) as twins of original core holes in order to show that earlier holes with poor core recovery preferentially lost chalcocite, resulting in lower copper grades than was realistic. The twin PQ sized core holes were very carefully drilled (triple tube with steel inserts) to maximized core recovery and relatively few sample intervals fell below an 80 percent core recovery. Comparison of the original hole with the twin was affected by short-range lateral variability in copper grade that made for an irregular comparison on a sample-by-sample basis. The weighted averages of the 116 original samples to the new twin samples showed that on average the recovery increased by 21 percent (71% to 92% recovery) and the average copper grade increased by 8.5 percent (5.07% to 5.50% Cu). PAH believes that this provides further evidence of the tendency to have better copper grades with better core recovery (preferential loss of chalcocite). The four PQ twin holes were included in the drill hole database for resource estimation.

14.3 Other Verification

The project has been sampled and tested over several years by various geologists, operating under the ownership of two different mining companies. The mineralization can be observed visually, so there is no doubt of its existence. As such, PAH did not conduct any additional verification of the sample grades.

14.4 Drillhole Database Management

All data for the Las Cruces project is stored in a database. The raw lengths for the geotechnical sections of the log were stored and all calculations and outputs were carried out on reports or forms backed by queries. The different lithological units were coded for input in the digital database.

Sample intervals and sample type as marked for cutting by the geologist was entered manually from the sample control sheets filled out by the sampling technicians. Assay data was returned from the laboratory on spreadsheets. The data loading and merging with the assay table was carried out directly from the imported spreadsheets. Duplication, overwriting and missing data was all flagged as part of the import control routine. All collar information is entered manually from the survey report sheets. Downhole survey data was read from the film produced from the survey equipment and entered manually.

A wide range of checks were made to ensure parity. A number of spot checks were also made on the hanging wall and footwall surfaces and also on the data processing to produce the final estimation figures allocated to the block model. Random checking of the database was also done

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and only minor errors and missing values were found mainly within stockwork and CZ zones. A considerable amount of core recovery and core diameter entries were missing and were subsequently entered.

In addition to the work conducted by others at the various stages of Feasibility studies, PAH conducted independent checks for integrity and statistical bias as well.

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15.0 ADJACENT PROPERTIES

The Las Cruces deposit occurs near the eastern end of the Iberian Pyrite Belt, a 250-kilometer long and 40-kilometer wide geologic belt that extends eastward from southern Portugal into southern Spain. The belt is host to more than 100 mineral deposits, some of which were exploited for metals as long ago as pre-Roman times. Mineralization at Las Cruces, as in most other Iberian Pyrite Belt deposits, consists of syngenetic massive sulfides containing polymetallic mineralization.

The Las Cruces deposit is a blind deposit that does not outcrop, due to the 100 to 150 meters of sandstone and calcareous mudstone (marl) that was deposited on top of the deposit. No other deposits have been found in the immediate area, but exploration is difficult given the thickness of the overburden. The nearest deposits are Aznalcollar and Los Frailes, both occurring approximately 10 kilometers to the west in the area where the host rock assemblage outcrops at the surface. The Aznalcollar and Los Frailes deposits consist of a zinc-lead massive sulfides that have been in production over the last 10 to 20 years.

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16.0 MINERAL PROCESSING AND METALLURGICAL TESTING

16.1 Metallurgical Testing Programs

Metallurgical testwork began in 1996, directed by the previous owner, Rio Tinto. Rio Tinto found the ore difficult to process by flotation. The ore was amenable; however, to processing by a variety of hydrometallurgical methods, including pressure, bacterial, ferric sulfate, ammonium sulfate, and ammonium carbonate oxidation/leaching. Rio Tinto decided that processes using ferric sulfate were likely to be the most successful and investigated two processes using pilot plants that incorporated continuous oxidation and leaching and fully integrated extraction circuits: one used bacterial oxidation and the other pressure oxidation; results of these tests indicated that either option would be economically viable. Copper recoveries reported from the testwork ranged from 90 to 92 percent.

Following CLC’s purchase of the property in 1999, previous testwork was reviewed and it was decided that pressure-oxidation/atmospheric-leaching was the more attractive alternative. In January 2000, CLC contracted Bechtel to prepare a bankable feasibility study based on the pressure leaching technology developed by Dynatec. The study was completed in early 2001.

In January 2003, CLC contracted Lurgi Metallurgie GmbH/Outokumpu Technology Group (OTG) to prepare a revised technical and economic evaluation of the Las Cruces process plant to optimize costs for the facility, improve copper recovery, and simplify the process flowsheet. OTG reviewed the project information and recommended a revised flowsheet that incorporated atmospheric oxidation/leaching (using oxygen in a heated slurry), Outokumpu Compact SX technology, and other innovations to improve copper recovery and lower capital and operating costs. To confirm their predictions, OTG performed further metallurgical testwork (batch and mini-continuous pilot-scale) at their Outokumpu Research Center (ORC) in Pori, Finland to demonstrate the viability of their proposed flowsheet and supply design criteria for the envisioned process facility.

16.2 Metallurgical Samples

The mineralization of primary interest has been designated by CLC as High Copper (HC) with two sub-classifications, HCL (High Copper Low Density) and HCH (High Copper High Density). The copper in the ore is primarily chalcocite with minor amounts of covellite, bornite, chalcopyrite, tennantite-tetrahedrite complex and enargite. The main differences between HCL and HCH ore types are that the HCL has more silica and less pyrite than the HCH. Though there are wide variations in the copper content of the ore types, typical metal contents are 6 to 8 percent copper, zinc from 0.1 to 0.4 percent, lead from 0.3 to 1.4 percent, iron from 28 to 36 percent and sulfur ranges from 33 to 46 percent. Precious metals content of the ore types averages about 0.5 gpt gold and 26 gpt silver. Because copper accounts for over 90 percent of the metal value found in the ore types, metallurgical testing tended to focus mainly on recovery techniques for the copper.

Pincock, Allen & Holt 9033.05 July 8, 2004 16.1

For the OTG metallurgical testing program, a Master Composite (MC) weighing about 5.5 tonnes was prepared by CLC from over 1000 drill core intervals to represent the entire ore body. The MC consists mainly of the HCH and HCL ore types with minor amounts of the C-4 ore type, a small high-copper area located in the very western portion of the deposit. The C-4 ore represents only about 3 percent of the total ore planned for mining; however, the copper content in the C-4 ore is high at about 9.5 percent. The MC was made up in the weight ratios of HCH:HCL:C-4 of 7:7:1, nearly the same as the planned ore mining tonnage ratios. Approximately 1 tonne of MC remains in cold storage in Seville. The MC contains about 6.3 percent copper, nearly the same as the average ore grade planned for processing of 6.6 percent copper.

Figures 16-1 and 16-2 present the Metallurgical Sample Location Maps for the Master Composite sample drill holes and intervals. PAH believes that the MC fairly represents the ore body and the results from the testwork should be sufficient for projection of the metallurgical response of the various ore types planned for processing.

The mineralogical composition analysis of the MC indicated that over 60 percent of the material is pyrite, less than 1 percent is chalcopyrite, about 23 percent is chalcocite/covellite/bornite, 13 percent is quartz, and less than 2 percent as miscellaneous sulfide minerals (galena, sphalerite, tetrahedrite-tennantite). Sequential Copper Phase analysis (assaying using various solvents to determine mineral content) also indicated that less than 5 percent of the MC was made up of primary sulfide copper minerals; accordingly, recovery of copper by ferric sulfate leaching had the potential to be above 90 percent.

The grinding requirements were determined by Bond Ball Mill Work Index measurements on 15 different samples of HCH and HCL ore types during 1998 testing. The HCL sample averaged 11.88 kWh/t while the HCH samples averaged 13.9 kWh/t with test values ranging from 9.0 to a high 16.5 meaning that the samples have a high variability. Additional work index measurements were performed in 2000 on 17 different samples and the Master Composite. The MC measurement was 10.8 kWh/t with the average of all tests being 10.2 kWh/t with values ranging from 7.9 to 13.3. The grinding circuit design assumes that the work index will vary from 10 to 15 kWh/t. PAH believes that the grinding circuit design criteria utilized work index values that were reasonable based upon the results of the testing.

16.3 Batch Metallurgical Testing

In OTG’s review of the previous Dynatec pilot plant results, it was determined that atmospheric leaching was not completed after the 40 minutes of retention time. Also, based upon their experience in commercial leaching of zinc, cobalt, copper and nickel, OTG believed that further investigations using atmospheric oxidation/leaching were warranted.

Pincock, Allen & Holt 9033.05 July 8, 2004 16.2

FIGURE 16-1

Metallurgical Sampling Composites Planview

Pincock, Allen & Holt 9033.05 July 8, 2004 16.3

FIGURE 16-2

Metallurgical Sampling Composites Section

Pincock, Allen & Holt 9033.05 July 8, 2004 16.4

Eleven batch oxidation/leach tests were performed using the MC in lab-scale Outokumpu TOP (OKTOP) reactors under atmospheric pressure with oxygen sparged into the heated slurry. Copper recoveries and iron extractions in the batch tests were above 92 percent and about 3 percent, respectively. Optimum results were obtained with ferric iron concentrations at 20 to 25 gpl, a temperature of about 90 degrees centigrade and leach times of about 8 hours.

16.4 Pilot Leaching and Solvent Extraction Testing

Based on results of the batch oxidation/leach testing, two separate 10-day continuous pilot plant oxidation/leach tests were performed at ORC to demonstrate the technical operability of a combined atmospheric oxidation/leach and solvent extraction circuits on the MC, confirm key control parameters for the process, determine the amount of bleed solution, confirm iron extraction without lowering copper recoveries, confirm optimum oxidation/leach time, and determine solid/liquid separation parameters. The test results would also be used to generate design criteria for the process facility.

The first oxidation/leach test run found that the ground MC material was too coarse (-300 microns) for proper suspension in the leach reactors and for optimization of power requirements for agitation. Daily copper recoveries during the first run, however, were generally above 94 percent despite the suspension and high power consumption issues. The MC material used in the second test run was ground finer (-150 microns). Daily copper recoveries during both tests were above 94 percent, as long as the oxygen addition and ferrous to ferric iron ratios were optimized.

Overall, the continuous leach tests resulted in copper recoveries of over 94 percent, established an oxidation/leach retention time of about 7 hours, confirmed desirable iron extractions at about 3 percent, showed zinc extractions of about 45 percent that leveled off when the bleed rate to control iron levels was established, indicated solution acid concentrations of 20 to 60 gpl, resulted in oxygen consumptions ranging from 36 to 54 kg/t of ore and showed arsenic extractions of about 25 percent.

Settling and filtering characteristics were also obtained during the testing. The calculated unit area for thickening of the residues ranged from 0.169 to 0.192 square meters per tonne per day. The calculated filterability of the residue ranging from 400 to 1200 kilograms per square meter per hour with the resultant cake moisture being about 9 percent. These results have been used in establishing the plant design criteria.

ORC completed slurry-mix tests in larger pilot-scale vessels to better determine the requirements for agitation, wear materials and corrosion. Results of these tests indicate that 0.514 kW/cubic meter will be required for leach agitation and that special attention should be paid to the design of the leach agitator blades to minimize erosion. Corrosion tests on three types of corrosion-resistive metal were also performed with no corrosion observed. Additional tests are scheduled prior to final design to confirm the agitation requirements and reactor and agitator design parameters.

Pincock, Allen & Holt 9033.05 July 8, 2004 16.5

Continuous SX (solvent extraction) pilot tests, without electrowinning, were operated on PLS (pregnant leach solution) from the leach test runs and simulated the primary solvent extraction circuit. Raffinate from the SX pilot tests was returned directly to the oxidation/leach tests for use as leach solution. A brief secondary SX test was performed using lime-neutralized solution obtained from primary SX raffinate solution. The primary SX circuit consisted of two extraction stages, one wash stage and two stripping stages, similar to the plant design. The secondary SX circuit consisted of one extraction stage, one wash stage and two stripping stages, similar to the plant design.

The SX pilot operation demonstrated the combination of oxidation/leach and SX was successful and resulted in predicted copper extractions from PLS of 70 percent for the project, raffinate containing 10 gpl (grams per liter) copper and 70 gpl acid and iron concentrations of about 15 gpl. Improved removal of suspended solids in the plant operation should result in less “crud” formation. The secondary SX test indicated that copper extraction would be about 90 percent.

16.5 Pilot Effluent Testing

Raffinate bleed from the SX testing was neutralized with milk of lime to precipitate heavy metals. Selenium, arsenic, cadmium, copper, zinc, and lead concentrations in the treated effluent were well below CLC discharge limits. Resultant clarified solutions can be discharged to the environment after pH adjustment.

Thickening and filtering tests were performed on precipitates to generate design criteria for the process plant. Cake moistures experienced in the filtering tests ranged from 52 to 68 percent by weight.

16.6 Ore Variability

Thirty samples were tested to determine the variability of the mineralogy within the ore types. All samples were subjected to sequential copper phase analysis (assaying using various solvents to determine mineral content) while nine samples, representing different years of the Project, were subjected to batch leaching tests.

Most of the thirty sequential analyses indicated that primary copper sulfide minerals (low anticipated copper recovery) make up less than 10 percent of the material; therefore, the ore types should be amenable to atmospheric oxidation/leaching. The nine samples representing different years of the project had an average copper extraction of about 92 percent with values ranging from 88 to 95 percent and iron extractions averaging about 5 percent. The HC4 sample had an abnormally high iron extraction at about 14 percent.

Because only one oxidation/leach test was performed on HC4 material, additional batch oxidation/ leach tests will be performed on HC4 material prior to final design to analyze the behavior of the iron as well as obtain more information on copper extractions from this material. Additional batch oxidation/leach tests with samples representative of the different project years, based upon the current mining plans, will also be performed to expand the database on copper and iron extractions of the primary ore types.

Pincock, Allen & Holt 9033.05 July 8, 2004 16.6

16.7 Copper Recovery Predictions

The overall life-of-mine copper recovery is predicted to be 91.4 percent, with the ferric leach process, based on the current mine production plan. Batch oxidation/leach test results of nine variability samples (4 HCH, 4 HCL, 1HC4) were used to develop relationships between copper recoveries and copper grades by ore type.

The following copper recovery relationships were developed from the test data for each ore type and were used to estimate the copper recoveries for the project:

• HCH: Cu recovery = (0.0027 * Cu Grade) + 0.8873

• HCL: Cu recovery = (0.0027 * Cu Grade) + 0.9101

• HC4: Cu recovery = (0.0027 * Cu Grade) + 0.9241

Copper losses in the plant operation were also estimated and considered solution losses contained in leach residue and precipitate filter cakes. An overall copper loss adjustment factor due to these losses, for all ore types, is estimated at 0.70 percent.

Due to expected start up problems, a copper loss adjustment factor of 1.7 percent is also estimated to occur during the first year of operation.

Figure 16-3 presents the metallurgical recovery graphs by ore type and shows the predicted recoveries versus ore grades based on the existing information.

16.8 Copper Production Estimate

A first-year production ramp up factor of 71.7 percent cathode production capacity has been estimated for the Project based upon predicted losses in plant availability and lower than design plant efficiencies during that time period. Cathode quality is also anticipated to ramp up to the predicted 95 percent of LME Grade A quality by the fourth month of operation.

The first-year ramp up factor, the loss factors and the recovery prediction relationships discussed above were all used to develop the copper production schedule for the life of the Project.

Approximately 16 million tonnes at a grade of 6.6 percent copper will be processed over the 15-year project life. About 91.4 percent of the copper will be recovered to produce approximately 966,000 tonnes of cathode copper, or an average of 66,000 tonnes per year.

Pincock, Allen & Holt 9033.05 July 8, 2004 16.7

Figure 16-3

Copper Extraction vs Copper Grade

Pincock, Allen & Holt 9033.05 July 8, 2004 16.8

The estimated life-of-mine copper production schedule is presented in Table 16-1.

TABLE 16-1

MK Resources Company

Las Cruces Project Technical Report

Copper Production Schedule

	Year 1	Year 2	Year 3	Year 4	Year 5	Year 6	Year 7	Year 8
K-Tonnes Ore Processed	618	1,043	1,065	758	1,105	1,075	953	1,138
Mill Feed Assay (% Copper)	8.597	6.967	6.819	9.467	6.552	6.683	7.477	6.340
Tonnes Cathode Copper Produced	47,354	65,934	66,053	65,952	66,055	66,129	66,087	65,996
	Year 9	Year 10	Year 11	Year 12	Year 13	Year 14	Year 15	Totals
K-Tonnes Ore Processed	1,277	933	1,094	1,296	1,168	1,138	1,311	15,972
Mill Feed Assay (% Copper)	5.694	7.718	6.608	5.544	6.137	6.403	4.997	6.616
Tonnes Cathode Copper Produced	66,060	66,043	66,037	66,053	66,072	66,092	60,138	966,055

16.9 PAH Comments

• The metallurgical testwork has been performed in a thorough and professional manner and it appears that the atmospheric oxidation/leaching technology selected for the project, though unique in the copper industry, has a high likelihood of being successful.

• Additional batch metallurgical testing is planned prior to final design to better determine the metallurgical response of HC4 ore and to confirm annual copper recoveries from the main ore types based on the current mine plan.

• In addition, more testing will be performed to finalize the oxidation/leach reactor agitator and tank designs while optimizing oxygen requirements.

• The copper recovery predictions are reasonable and consider the test results, and adjustment factors for plant losses during ramp up and equipment operation.

• The copper production schedule is reasonable and considers the recovery predictions by ore type.

• Design criteria for the process facility appears to have been properly extracted from laboratory test results.

• The operation of “batch-scale” and “continuous mini-pilot-scale” tests was prudent with the copper recovery results being comparable.

Pincock, Allen & Holt 9033.05 July 8, 2004 16.9

17.0 MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES

The discussion of the mineral resources and reserves for the Las Cruces project are segregated by the geology / resource discussion, followed by the mine design and reserve discussion.

17.1 General Information

DMT created a revised resource model for the Las Cruces deposit in 2003, largely to incorporate changes in what data was included or excluded as a result of sample recovery issues. The modeling approach otherwise was very similar to the previous 2000 model created by Independent Mining Consultants. The resource model was set up to reasonably cover the deposit area using Maptek’s Vulcan Software package. Block size was set up to be 10 by 10 meters in plan and a 5 meter bench height, which locally was subcelled to 5 by 5 meters in plan and a 2.5 bench height. Figure 17-1 shows the resource model limits along with the project area topography and drill holes.

17.2 Drill Hole Sample Database

17.2.1 Database Content

The drill hole data for the project is stored in an Access database that contains all “CR” series holes used for resource evaluation. Assay files include grade values for copper, zinc, lead, gold, and silver, as well as other trace metals. The majority of the grade data is from PQ sized core for which the sample consisted of a quarter of the core (3466 samples or 79 percent of all the secondary sulfide core). Analytical data from the laboratory was provided in electronic format and merged directly into the drill hole database. Where analytical values were missing, either due to non-assay or sample loss, they were assigned a 0 value in the databases. For compositing and other processing of data, the 0 values are treated as non-existing. Assays below detection limits were assigned a value of one half of that limit in the database. Geological and geotechnical logging data were entered into the database through a data entry form.

The lithology code in the database is a two-part identifier that indicates the group of rocks (i.e. sulfide, gossan, volcanics, etc.) and the principle rock type (massive sulfide, semi-massive sulfide, disseminated sulfide, intense hematite, etc.) The lens code indicates the mineralized grouping (HCH - high copper/high density with secondary sulfides, HCL – high copper/low density with secondary sulfide, HC4 – high copper western zone, AU – gold enriched gossan, C1 – chalcopyrite dominant primary sulfide). The drill hole data file provided contains 18,093 samples.

A total of 4,407 samples (24 percent of the sample data) are from the secondary sulfide mineralization. Of the secondary sulfide samples, 57 percent are from the HCL zone, 36 percent are from the HCH zone, 5 percent are from the HC4 zone, and 2 percent are from the HCF zone.

Pincock, Allen & Holt 9033.05 July 8, 2004 17.1

FIGURE 17-1

Pincock, Allen & Holt 9033.05 July 8, 2004 17.2

Drilling of the secondary sulfides consisted of 88 percent PQ size core (85 mm core diameter), 9 percent HQ size core (63.5 mm core diameter), and 3 percent NQ size core (47.6 mm core diameter).

Although not considered as an economic zone in the current feasibility, the AU zone consists of oxidized gossan with secondary enriched gold and silver grades. A total of 704 samples (4 percent of the sample data) are from the precious metal enriched gossan.

The original massive sulfide deposit consists of primary sulfides, dominated by pyrite and chalcopyrite, and is considered waste material in the current feasibility. A total of 625 samples (3 percent of the sample data) are from the primary sulfide zone (C1 and CB lenses).

PAH notes that no core recovery was recorded in the drill hole database for holes CR001 to CR023, and eleven sample intervals in CR258. This includes six holes (108 samples) in the secondary sulfide zone significant for resource estimation (2.5 percent of the secondary sulfide sample data contained in holes CR001, CR002, CR019, CR020, and CR021, and CR258). In order not to lose these holes for compositing and subsequent resource estimation after applying a core recovery cutoff, the core recovery for these holes was set to 100 percent. PAH recommends that an actual core recovery value be estimated from core photographs or remaining core reference material for future work.

DMT noted in their 2003 feasibility report that densities were not recorded for 58 secondary sulfide sample intervals. PAH found only 12 secondary sulfide intervals for which no density was recorded, all in drill hole CR233.

17.2.2 Database Checking

Independent Mining Consultants conducted a validation of drill hole database integrity by randomly selecting 15 holes (about 5 percent of the data) by comparing original assay certificates with the database grade entries for copper, gold, lead, zinc, silver, and sulfur. Independent Mining Consultants found that the error rate was extremely low. The few discrepancies found were minor in nature and none were related to copper. Physical examination of 10 of these holes also found that the geologic logging and analytical values appeared consistent with the core. Original films for the downhole surveys of these holes also found no problems. As a result, Independent Mining Consultants concluded that the database integrity was good and was adequate for feasibility study work.

Independent Mining Consultants also conducted a validation of drill hole database density values. The error rate for the initial check of select holes was found to be high enough to warrant a complete check of all the available density data. They verified the density calculations from the basic weight data and found that 9 percent of the density data could no longer be verified and that 1 percent of the density data was incorrect. The erroneous data was discarded to produce clean density data. No independent density measurements were taken as part of this work as the remaining core was found to be too degraded to produce meaningful results.

Pincock, Allen & Holt 9033.05 July 8, 2004 17.3

PAH previously (October 2001) searched the database with an automated checker that seeks out data inconsistencies and/or anomalous values. Based on this examination, PAH found no significant errors that would adversely impact the quality of the resource estimate.

DMT for the current feasibility (2003) checked 20 holes from the original down-hole photographs and found 9 errors out of 145 readings (6 percent error rate). Previous review of down-hole surveys for 15 holes by Independent Mining Consultants found no discrepancies with the database. DMT further identified that the magnetic declination for true north was incorrect. The original 6-degree declination was corrected to 4 degrees.

17.2.3 Drill Hole Core Recovery Cutoff

Samples from drilling of the Las Cruces deposit was subject to considerable variability in core recovery. This was discussed in the 2001 Bechtel feasibility study and has been evaluated by CLC (July 2003) and IMC – UK (August 2003) and well summarized in the current 2003 DMT feasibility study. Geologic observation during the logging of the Las Cruces core provides good evidence of the variable loss of core material due to the abrasion and erosion of friable and sooty, chalcocite-bearing mineralization. The apparent preferential loss of this material during drilling and sampling has tended to decrease the relative content of copper minerals, with a proportionally less decrease in the relative content of the other associated minerals, including pyrite, quartz, sphalerite, and/or galena. As a result of this, the copper grade in the actual analyzed sample of the recovered core tends to be less than in the copper grade in the original in-place location.

PAH finds that the majority of the geologic and statistical data provide evidence for the preferential loss of chalcocite, although reduction in copper grade is not a uniform systematic statistical relationship, but is somewhat irregular due to chalcocite distribution both as more durable masses and elsewhere as friable aggregates. It appears that both high core recovery and low core recovery zones have been subject to variable chalcocite loss. As such, core recovery itself may provide a good general indication that original in-place copper grades are higher than reported, but are not an absolute indication on a sample-by-sample basis that the original in-place copper grade was higher. The effect of the preferential loss of chalcocite appears to have tended to increase the difference between original in-place copper grade and the actual assayed copper grade (after core recovery affects) at progressively lower core recoveries.

This is significant in terms of generating a reasonable resource model, resulting in the question of what sample core recoveries result in reliable enough copper grades. If too many of the lower core recovery samples that potentially contain the lower copper grades are used in the resource modelling, the resulting effect is to potentially understate the copper grade of the deposit. If too few of the lower core recovery samples that potentially contain the lower copper grades are used in the resource modeling, the resulting effect is to have a reduced confidence in the grade estimation because too few samples were used in the estimation. This is shown graphically in Figure 17-2.

Pincock, Allen & Holt 9033.05 July 8, 2004 17.4

For resource modeling purposes, a limit on the core recovery percentage of each sample is required, to determine which samples should be included (relatively reliable copper grades) and which samples should be excluded (relatively unreliable copper grades) for the subsequent resource modeling. This core recovery cutoff was been applied in the modeling process. In previous work (Bechtel Feasibility), a 50 percent core recovery cutoff was applied, with no justification of why this cutoff was selected. For the current work by Cobre Las Cruces (CLC), the reliability of the sample grades at those recovery levels were questioned and investigated further. As a result, the percent core recovery cutoff grade was increased to an 80 percent core recovery. This clearly gives more reliability for the copper grades to those samples that meet this higher core recovery cutoff grade criteria.

As a result, however, there are fewer actual copper grade values that are available for use in the modeling process, which then implicitly relies on whatever adjacent, non-excluded copper grades are available for the copper grade modeling process. In other words, the more sample copper grades are excluded by the core recovery cutoff grade, the less the actual data remaining for copper grade modeling. For secondary sulfide lenses, at core recovery cutoff grades above 70 percent, the number of available samples drops off quickly (at 50 percent cutoff about 90 percent of the samples are available, at 80 percent cutoff about 75 percent of the samples are available, at 90 percent cutoff about 50 percent of the samples are available, at 100 percent cutoff about 15 percent of the samples are available). The 80 percent core recovery cutoff for secondary sulfide samples includes 3,275 samples out of a total of 4,407 samples (74%).

PAH finds that the selection of an 80 percent core recovery cutoff grade appears to be a reasonable compromise, whereby the core recovery cutoff is high enough that the remaining non-excluded samples have reliable grades, but that the core recovery is low enough that there are a sufficient number of the remaining non-excluded samples to effectively create a copper resource block model (discussed further in database subsection). Dr. Stephen Henley of IMC –UK (August 2003) evaluated this issue during the preparation of the current feasibility study and provided an independent opinion, concluding that the 80 percent core recovery cutoff grade is reasonable and still likely to lead to an underestimation of the resource grade. This reduced confidence was reflected in the DMT resource estimate by disallowing any of the resource that has been heavily influenced by neighboring samples substituting for missing samples from being classified as measured, as discussed in the subsequent report subsection.

17.2.4 Sample Statistics

Sample copper statistics are summarized in Table 17-1. The table compares sample copper statistics for all secondary sulfide samples (0% core recovery cutoff) and for those secondary sulfide samples at an 80 percent core recovery cutoff. A volume weighted mean has been calculated to compensate for different sample lengths and densities. DMT noted some 58 secondary sulfide

Pincock, Allen & Holt 9033.05 July 8, 2004 17.5

FIGURE 17-2

Pincock, Allen & Holt 9033.05 July 8, 2004 17.6

samples did not have density determinations so these were calculated by regression relationships based on the iron and sulfur content of these samples. PAH notes within the data provided for checking, hole number CR233 has 12 secondary sulfide samples for which a density value is not present in the database. It can be seen that when the 80 percent core recovery cutoff is applied to the secondary sulfide samples the mean copper grade is approximately 10 percent higher grade than when no core recovery cutoff is used. PAH notes that secondary sulfide samples below the 80 percent core recovery cutoff have a mean copper grade of 5.10 percent copper, or 35 percent lower than the mean copper grade for the samples above the 80 percent core recovery cutoff.

TABLE 17-1

MK Resources Company

Las Cruces Project Technical Report

Sample Copper Data Statistics

Zone (Lens)	Parameter	Samples (0% RECP cutoff)	Samples (80% RECP cutoff)
All Secondary Copper (HCH, HCL, HC4)	Number Minimum (%Cu) Maximum (%Cu) Vol. Wt. Mean (%Cu) Std. Dev. (%Cu) Coef. Of Var.	4,407 0.01 39.35 7.01 5.71 0.83	3,275 0.01 39.35 7.85 6.00 0.79

Note: Volume weighted mean = grade * length * density.

Sample density statistics are summarized in Table 17-2. Density values were determined for every sample interval. PAH notes that CR233 had 12 sample intervals for which no density value was recorded in the drill hole database. The table compares sample density statistics for all secondary sulfide samples (0% core recovery cutoff) and for those non-excluded secondary sulfide samples at an 80 percent core recovery cutoff. As with the copper data, a volume weighted mean has been calculated to compensate for different sample lengths and densities. It can be seen that for the density data there is little difference when the 80 percent core recovery cutoff is applied to the secondary sulfide samples.

TABLE 17-2

MK Resources Company

Las Cruces Project Technical Report

Sample Density Data Statistics

Zone (Lens)	Parameter	Samples (0% RECP cutoff)	Samples (80% RECP cutoff)
All Secondary Copper	Number	4,407	3,275
(HCH, HCL, HC4)	Minimum (g/cm3)	1.86	2.00
	Maximum (g/cm3)	5.04	5.04
	Vol. Wt. Mean (g/cm3)	3.66	3.71
	Std. Dev. (g/cm3)	0.65	0.63
	Coef. Of Var.	0.18	0.17

Note: Volume weighted mean = grade * length * density.

Pincock, Allen & Holt 9033.05 July 8, 2004 17.7

17.3 Composites

17.3.1 Composite Calculation

Drill hole sample data was composited into regular geological composites at a target 5-meter length. Basically, within a given geologic interval the composites were split into regular intervals as near as possible to the target length. PAH notes that this forces composites not to extend across a geologic interval, but results in some variability in composite lengths, especially for narrower geologic intervals. Additional dilution was subsequently incorporated into the reserve estimate to compensate for mining effects at the edges of the secondary sulfide mineralization (see mining section). The composite grades were generated by weighting the individual samples by assay length and by density. This provides a volume weighted mean that is more appropriate with variable density materials. All secondary sulfide composites were considered together, as any separation by lens for geostatistical purposes was considered to be impractical.

If any sample data contained within a composite interval was excluded because it was below the 80 percent core recovery cutoff, then the 5-meter composite value was calculated using only the remaining sample data with above 80 percent core recovery. There was no minimum sample length required for the composite to be calculated. However, only composites for which more than 50 percent of the length was “supported” by non-excluded sample data were used for resource estimation.

17.3.2 Composite Statistics

As a result of the application of an 80 percent core recovery cutoff to the samples and the subsequent use of a support cutoff of 50 percent, a total of 650 non-excluded composites were calculated, 71 percent of the total number without applying cutoffs. Acceptable composites were available for 134 holes out of the 153 holes that intercepted the secondary sulfide zone. Nineteen holes were entirely excluded because they contained no acceptable secondary sulfide composites. In addition, PAH finds that there are another 23 holes for which the application of an 80 percent core recovery cutoff and a 50 percent support cutoff, results in less than half of the secondary sulfide composites for those holes being accepted. In addition, there were five drill holes (plus part of a sixth) for which a 100 percent recovery was assigned because no core recovery value was available, otherwise these holes may have experienced a high exclusion rate as well as they were among some of the first holes drilled. PAH evaluated the distribution of these holes and found them to be scattered throughout the deposit area. This is a better situation for resource modeling than if they were clustered together. A comparison of sample and composite copper statistics are shown in Table 17-3. PAH finds this to be a reasonable comparison.

PAH evaluated statistical outliers in the 5-meter composites based on data population distributions and found some statistical evidence for special treatment of copper grades above 25 percent. DMT did not apply any special treatment. PAH notes, however, that this would have affected only four composites and as such would have a minimal effect on the global resource.

Pincock, Allen & Holt 9033.05 July 8, 2004 17.8

TABLE 17-3

MK Resources Company

Las Cruces Project Technical Report

Composite Copper Data Statistics

Zone (Lens)	Parameter	Samples (80% RECP cutoff)	5-M Composites (50% Support cutoff)
All Secondary Copper (HCH, HCL, HC4)	Number Minimum (%Cu) Maximum (%Cu) Mean (%Cu) Std. Dev. (%Cu) Coef. Of Var.	3,275 0.01 39.35 7.85 6.00 0.79	650 0.18 27.88 7.67 5.07 0.66

Note:	1) For samples - volume weighted mean = grade * length * density.
	2) For composites – length weighted mean = grade * length.

A comparison of sample and composite density statistics are shown in Table 17-4 and PAH finds a reasonable comparison. PAH does not find any evidence for special treatment of any of the higher density values.

TABLE 17-4

MK Resources Company

Las Cruces Project Technical Report

Composite Density Data Statistics

Zone (Lens)	Parameter	Samples (80% RECP cutoff)	5-M Composites (50% Support cutoff)
All Secondary Copper	Number	3,275	650
(HCH, HCL, HC4)	Minimum (g/cm3)	2.00	2.25
	Maximum (g/cm3)	5.04	4.80
	Vol. Wt. Mean (g/cm3)	3.71	3.71
	Std. Dev. (g/cm3)	0.63	0.56
	Coef. Of Var.	0.17	0.15

17.3.3 Composite Variography

DMT conducted a geostatistical evaluation, with variography based on composites generated in holes at an approximate 50-meter spacing. Two “statistical crosses” of more closely spaced (12.5 meters) inclined (60 degrees) were drilled to support the geostatistical evaluation, by providing sample pairs at a closer spacing to better define the shorter range sample variability. One cross was placed in the western part of the deposit, while the other was in the central part of the deposit and was not completed.

DMT found that for the copper grade the longest effective range was 220 meters at a 20-degree azimuth. The shortest effective range was 110 meters at a perpendicular azimuth of 110 degrees. These directions were consistent with the geologic observations for the deposit. The independent variance (nugget) was found to be approximately 40 percent of the total variance, which PAH notes is indicative of irregular mineralization.

Pincock, Allen & Holt 9033.05 July 8, 2004 17.9

It is PAH’s experience that primary massive sulfide mineralization typically is deposited with a good degree of continuity such that relatively high distances or ranges of influence could be expected (up to 100 meters or more) for primary chalcopyrite massive sulfide mineralization. Subsequent secondary enrichment with chalcocite mineralization, however, is a very irregular process, dependent on leach solution grades and primary sulfide porosity and permeability and disrupts the original statistical continuity. Twin drilling corroborates this, in that intervals even at very short distances do not compare well.

PAH conducted a check of variography and found that variography for the mineralization using a zero percent copper cutoff (all composites) found results similar to those of CLC, with maximum ranges of 200 meters and a moderate nugget variance. Using copper grade indicators, PAH found that copper grades at a five percent cutoff increasingly showed the influence of the more irregular secondary chalcocite mineralization, with the maximum variogram range dropping to around 100 meters and an increase in the nugget variance. At a 7.5 percent copper grade indicator the maximum variogram range is around 60 meters, with a further increase in the nugget variance. The nugget effect is relatively high for a copper deposit and provides a good statistical measure of the irregularity of the copper in higher-grade parts of the deposit, one that can be corroborated by geologic observation of the drill hole cross sections.

17.4 Rock Model

For the current feasibility study the Las Cruces deposit was completely reinterpreted, focusing on the secondary sulfide zones. Digital surfaces were created to delineate the three main secondary sulfide sub-lenses, including the High Copper – High Density (HCH), which included the minor HCI lens, the High Copper – Low Density (HCL) which included the minor MCL and HCF lenses, and the High Copper Lens Number 4 (HC4). The primary sulfide and gossan zones were also delineated, as well as divisions of the Paleozoic host rocks and overlying Miocene sedimentary cover.

17.5 Grade Models

A copper model was created by DMT using an ordinary kriging (OK) interpolation method. As the boundaries between the secondary sulfide lenses are complex and gradational, the HCH/HCL sub-lenses were grouped together for resource estimation. Interpolation was conducted using a search ellipsoid with the primary axis having a 200-meter radius oriented at a 20 degree azimuth as indicated by variography and geologic observation. The secondary axis had a 90-meter radius oriented at a 110-degree azimuth. The tertiary axis had a 10-meter axis oriented vertically. A minimum of two composites and maximum of 30 composites was specified for the interpolation. PAH believes that 30 composites probably result in more smoothing of grades during the interpolation process, however, this would be expected to have only a small affect on the global resource average grade.

Pincock, Allen & Holt 9033.05 July 8, 2004 17.10

A density model was created for all geologic units and mineralized lenses, with the exception of the Miocene marl and sandstone units, using an inverse distance squared interpolation method. Interpolation was conducted using a search ellipsoid with the primary axis having a 400-meter radius oriented at a 20 degree azimuth as indicated by variography and geologic observation. The secondary axis had a 300-meter radius oriented at a 110-degree azimuth. The tertiary axis had a 20-meter axis oriented vertically. A minimum of two composites and maximum of 12 composites was specified for the interpolation.

Primary sulfides have been considered as a single domain for the purposes of resource modeling, but are not part of the economic evaluation of the current feasibility.

A comparison of composite and block model copper statistics are shown in Table 17-5 and PAH finds a reasonable comparison. The slightly lower average copper grade for the block model is the result of declustering affects during the interpolation process and is to be expected by kriging.

TABLE 17-5

MK Resources Company

Las Cruces Project Technical Report

Block Model Copper Data Statistics

Zone (Lens)	Parameter	5-M Composites (50% Support cutoff)	Block Model
All Secondary Copper (HCH, HCL, HC4)	Number Minimum (%Cu) Maximum (%Cu) Mean (%Cu) Std. Dev. (%Cu) Coef. Of Var.	650 0.18 27.88 7.67 5.07 0.66	14,246 1.41 18.16 6.62 2.95 0.45

Note:	1) For composites – length weighted mean = grade * length.
	2) For block model – arithmetic mean.

A comparison of composite and block model density statistics are shown in Table 17-6 and PAH finds a reasonable comparison.

TABLE 17-6

MK Resources Company

Las Cruces Project Technical Report

Block Model Density Data Statistics

Zone (Lens)	Parameter	Samples (80% RECP cutoff)	Block Model
All Secondary Copper	Number	3,275	14,246
(HCH, HCL, HC4)	Minimum (g/cm3)	2.00	2.77
	Maximum (g/cm3)	5.04	4.69
	Vol. Wt. Mean (g/cm3)	3.71	3.67
	Std. Dev. (g/cm3)	0.63	0.45
	Coef. Of Var.	0.17	0.12

DMT conducted a number of checks of the reasonableness of the block model. DMT concluded from statistical integrity checks that the block copper grade and core density had been adequately

Pincock, Allen & Holt 9033.05 July 8, 2004 17.11

accounted for, with no inexplicable changes. DMT noted that Rio Tinto had previously investigated the effects of non-Gausian data transformations by comparing their ordinary kriged model with a multiple indicator kriged model and found that they gave an almost identical mean copper grade and concluded that the ordinary kriging reasonably interpolated the copper grade. DMT conducted a point validation check and found a slight bias that may tend to slightly overestimate block copper grades on a local basis; however, they concluded this was negligible on a global basis in the model. A point validation for core density showed no evidence of bias. Declustering investigations by DMT found that the overall effects on the grade interpolation from the drill holes clustered in the statistical crosses was very small. Investigation of stationary effects (mean and variance constant over distance) by DMT concluded that non-stationarity conditions do not have a major impact on the validity of the block model estimate. DMT also visually compared composite data with block model data for sections and plans through the resource model and found reasonable correspondence.

PAH notes that a reasonable amount of checking appeared to have been conducted by DMT. PAH conducted some further independent checks of the DMT resource model. PAH evaluated the statistics for each step of the modeling process and visually compared composite and block model data, checked the local estimation integrity of the copper grade model for grade bias by comparing individual composite values against the block value in which the composite occurred (backmarking) and by conducting a cross validation. PAH then checked the global integrity of the copper grade model for grade bias by creating a nearest neighbor model (computer polygonal model) and then compared the results against the PAH inverse distance model. As a result, PAH believes that the resource model, due to natural variability in the deposit and due to the exclusion of data with core recoveries of less than 80 percent, may be a modest estimator of grade locally, but believes that on a global basis the estimation is reasonable. In addition, CLC has noted that both major and minor faulting are present in the deposit, but are not well defined at this stage of exploration, and may locally influence copper grade distribution.

17.6 Resource Estimate

17.6.1 Resource Statement

The Las Cruces mineral resource was estimated from the grade and rock models, developed as discussed above. The resource includes all secondary sulfide material in the model at the given cutoff grade, without consideration to any mineable limits. The resource tonnage uses modeled core densities for each block, which are then adjusted in the resource tabulation to account for permeable pore space. Table 17-7 shows that at a 1.0 percent copper cutoff, the measured + indicated resource is 15.6 million tonnes of HCH, HCL and HC4 ore types, averaging 6.89 percent copper. There is an additional inferred resource of 0.4 million tonnes averaging 8.66 percent copper. Incidental primary sulfide (CZ) and gossan that occurs in the planned pit will be stockpiled separately and is not considered part of this resource estimate.

Pincock, Allen & Holt 9033.05 July 8, 2004 17.12

TABLE 1-1 / 17-7

MK Resources Company

Technical Report for the Las Cruces Project

Mineral Resource Summary (1.00% copper cutoff, 0.97 density adjustment)

Ore Type	Measured		Indicated		Measured +Indicated		Inferred
	Tonnes (000)	Grade (% Cu)	Tonnes (000)	Grade (% Cu)	Tonnes (000)	Grade (% Cu)	Tonnes (000)	Grade (% Cu)
HCH	3,860	7.25	2,670	8.00	6,530	7.56	274	9.48
HCL	5,020	6.56	3,370	5.77	8,390	6.24	37	4.86
HC4	440	8.61	230	8.55	670	8.59	49	6.93
Total	9,320	6.94	6,270	6.82	15,590	6.89	360	8.66

PAH checked the resource estimation from the DMT model by independently tabulating the tonnes and grade from PAH’s copy of the DMT model and found a good comparison. PAH notes that within the secondary sulfide resource there are very few blocks with less than a 2 percent copper grade, so the resource tabulation at cutoff grades up to 2 percent copper is very similar.

17.6.2 Resource Classification

DMT has classified the resource in accordance with the Australasian Code For Reporting Of Identified Mineral Resources And Ore Reserves (JORC Code). Under the JORC Code, only measured and indicated resources may be converted to mineable reserves with the application of appropriate economic factors. PAH believes that the classification is in compliance with the Canadian Institute Of Mining (CIM) standards required for National Instrument 43-101 reporting. PAH also believes that the classification is consistent with the development of United States Securities Exchange Commission (SEC) acceptable mineable reserves.

DMT has assigned confidence categories to the resource based on a combination of distance and number of drill holes. The strategy used is shown in Table 17-8. For the HCH and HCL zones, the search ellipsoid was horizontally oriented with the major axis oriented at an azimuth of 20 degrees. For the minor HC4 zone, the search ellipsoid was dipped into the dipping orientation of the zone, using the same classification criteria. In addition, the measured resource confidence category was locally adjusted to reflect the decreased estimation accuracy resulting from the exclusion of about 30 percent of the total drill hole data (from the application of sample and composite quality cutoffs). For each measured block, if the ratio of excluded:non-excluded composites, within a 25-meter search radius, was greater than 0.5, then it was reassigned to indicated. This requires that more than two thirds of the nearby composites be non-excluded composites in order to be considered as measured. This additional criterion reclassified 4.7 million tonnes averaging 6.9 percent copper of measured resource into an indicated category.

Pincock, Allen & Holt 9033.05 July 8, 2004 17.13

TABLE 17-8

MK Resources Company

Las Cruces Project Technical Report

Resource Confidence Classification Criteria

	Measured	Indicated	Inferred
Search Distance	100 x 45 x 10	200 x 90 x 10	200 x 90 x 10
Min. No. Of Composites	3	3	2
Min. No. Of Drill Holes	3	3	1
Excluded Sample Ratio	<0.5	—	—

It is PAH’s experience that primary massive sulfide mineralization typically is deposited with a good degree of continuity such that relatively high distances or ranges of influence could be expected (up to 100 meters or more) for primary chalcopyrite massive sulfide mineralization. Subsequent secondary enrichment with chalcocite mineralization, however, was a very irregular process, dependent on leach solution grades and primary sulfide porosity and permeability and has disrupted the original statistical continuity. Twin drilling corroborates this in that intervals even at very short distances do not compare well, as does observations from cross sections.

PAH conducted a check of variography to corroborate the search distances and found that using a zero percent copper cutoff (all composites) the results were similar to those of CLC, with maximum ranges of 200 meters and a moderate nugget variance. Using copper grade indicators at progressively higher cutoff grades, PAH found that the variogram range decrease to about 60 meters at a 7.5 percent copper grade indicator. The nugget effect at this indicator is relatively high for a copper deposit and provides a good statistical measure of the irregularity of the higher grade copper values in the deposit, which can be corroborated by geologic observation of the drill hole cross sections.

As a result, PAH believes that shorter ranges of influence may be appropriate, but because of the nominal 50 x 50 meter drill hole spacing, actual distances from composite points do not reach the maximum limit of the search ellipsoid. For measured blocks the maximum distance to the nearest composite was 80 meters, with 99 percent of the measured blocks within 60 meters of a composite, and the majority of the measured blocks being within 10 to 40 meters of a composite. The distribution of indicated blocks shows a similar distance distribution, with the maximum distance to the nearest composite of 90 meters, with 95 percent of the indicated blocks with 60 meters of a composite, and the majority of the blocks being within 10 to 50 meters of a composite. CLC finds that much of the measured and indicated blocks with higher distances to the nearest composite is a result of deposit and drill hole geometries, combined with a narrow vertical search, that prohibit closer composites from being identified.

17.6.3 Previous Resource Estimates

Previous resource estimates were prepared for Las Cruces and were compared by PAH to the current estimate, as shown in Table 17-9. These historical estimates have not been reviewed in detail by PAH and are superceded by the current feasibility Study resource estimate. No additional

Pincock, Allen & Holt 9033.05 July 8, 2004 17.14

resource drilling has occurred since 1999. The last two estimates, DMT in 2003 and Independent Mining Consultants (Tucson) 2000, were both conducted for CLC. The current DMT estimate has incorporated fewer, but higher quality composites (>80% core recovery) for resource estimation, which is the main resource model change since the previous estimate. DMT classifies the resource as measured (59%), indicated (39%), and inferred (2%). In 2000, Independent Mining Consultants (Tucson) prepared a resource model for use in the Bechtel feasibility study (2001) that planned for open pit mining. Independent Mining Consultants (Tucson) considered the entire resource to be at a combined measured + indicated (100%), but did not separate the resource into individual categories. In 1998, Rio Tinto prepared a resource estimate for their internal feasibility study that envisioned underground mining, prior to transferring the property to CLC. Rio Tinto classified the resource as measured (58%), indicated (36%), and inferred (6%).

TABLE 17-9

MK Resources Company

Las Cruces Project Technical Report

Comparison of Historical Resource Estimates

Estimate	%Copper Cutoff	Density Adjustment	Mineral Zones Included	Confidence Categories	Tonnes	Grade (%Cu)	Contained Tonnes Cu
DMT 2003	1.0	0.97	HCH,HCL,HC4	Meas+Ind+Inf	15,954,000	6.93	1,105,600
IMC 2000	1.0	0.95	HCH,HCL,HC4	Meas+Ind	15,511,000	6.07	941,1000
RioTinto 1998	1.0	0.97	HCH,HCL,C4	Meas+Ind+Inf	15,510,000	6.10	946,100

Note: Independent Mining Consultants (IMC) classified the entire resource as measured + indicated, with no inferred material.

It can be seen that the three feasibility study resource estimates are reasonably comparable in tonnes. The current DMT estimate, however, is about 10 percent higher in copper grade than the previous estimates due to the application sample and composite quality cutoffs, that tend to include higher grade samples and exclude lower grade samples, as discussed previously in this section.

17.7 Additional Exploration Potential

Additional exploration potential exists for the Las Cruces deposit in the secondary sulfide zone (the focus of this review), the primary sulfide zone (at depth), and the gossan zone (enriched precious metals overlying the secondary mineralization). In the secondary sulfide zone, potential exists for some upgrade to the copper grade based on the conclusions that copper tended to be preferentially lost during the drilling and sampling. Gold, silver, and minor base metal content also occurs in the mineralization, but will not be recovered through the planned processing methods. In addition to the copper grade of 7.0 percent copper (up to 39.4% Cu), the secondary sulfide also contains grades averaging about 0.2 percent zinc (up to 36.9% Zn), 0.6 percent lead (up to 40.0% Pb), 0.5 g/t gold (up to 62.9 g/t Au), and 26 g/t silver (up to 1,472 g/t Ag), based on averages for the drill hole data.

Below the Miocene unconformity, the uppermost part of the sulfide mineralization has been completely oxidized to an iron oxide gossan, with local enrichment of gold and silver. The gossan is also enriched in lead, as the mineral galena. PAH’s review found that the gossan is generally 10 to 20 meters in vertical thickness, with grades from all gossan drill hole samples averaging 5.9 g/t gold

Pincock, Allen & Holt 9033.05 July 8, 2004 17.15

(up to 353 g/t Au) and 98 g/t silver (up to 1,733 g/t Ag). Average copper grade in the gossan is 0.20 percent. Apparently associated with the gossan formation was the strong silicification of the host rock in the hanging wall immediately above the massive sulfide deposit, also containing erratic gold and silver enrichment. The economic contribution of the gossan zone is being evaluated further by DMT and is planned as an addendum to the current secondary sulfide feasibility study. Preliminary historic estimates by Rio Tinto at a 1.0 g/t gold cutoff indicate a mineralized material of 1.7 million tonnes averaging 4.25 g/t gold and 118 g/t silver, however Independent Mining Consultants found about the same tonnage, but with 2.52 g/t gold and 136 g/t silver. There is some evidence that the gold content has been under reported. None of the gossan mineralization is included in the current resource estimate.

Original primary massive sulfide mineralization contain massive to semi-massive sulfide minerals (generally more than 80% sulfide). Pyrite is the predominate sulfide mineral, with lesser finely intergrown sphalerite, galena, and chalcopyrite, as well as minor enargite, tennantite, and tetrahedrite. PAH’s review found that grades from all primary sulfide drill hole samples averaged 3.2 percent copper (up to 18.9% Cu), 1.1 percent zinc (up to 11.9% Zn), 0.4 percent lead (up to 5.4% Pb), 0.5 g/t gold (up to 1.63 g/t Au), and 20 g/t silver (up to 174 g/t Ag). Preliminary historic estimates by Independent Mining Consultants previously estimated the primary sulfide mineralized material to be 17.9 million tonnes averaging 1.85 percent copper, 2.97 percent zinc, 0.61 percent lead, 0.27 g/t gold, and 23.7 g/t silver, at a 1.0 percent copper cutoff. Most of the primary massive sulfide mineralization occurs down-dip to the northwest of the secondary sulfides. In the footwall below the primary massive sulfide is a stockwork of interconnected pyrite veins and veinlets, with local higher grades of copper and zinc. None of the primary sulfide mineralization is included in the current resource estimate.

The western part of the Las Cruces deposit has been cut off by a fault. Potential exists for finding the missing offset piece and some exploration drilling has been conducted. The Las Cruces deposit is a blind deposit, initially located by geophysical techniques, and subsequently drilled. Other anomalies are known in the area, and other historical producers (such as Los Frailes) are located within a 10-kilometer radius; therefore, other viable exploration targets are available.

17.8 Reserve Development

The reserves for Las Cruces are a subset of the resource model described earlier in this section, contained within an engineered mine plan. Several development options were considered, including underground methods; however, surface methods resulted in the best overall recovery at the lowest cost per tonne.

17.8.1 Mine Design

The Las Cruces project is planned to be a moderate-sized open pit mining operation using conventional truck and shovel operations. The use of underground methods was considered, however, the open pit method provided a more cost effective and controllable mining approach.

Pincock, Allen & Holt 9033.05 July 8, 2004 17.16

CLC plans to use a mining contractor to mine ore and waste, and will employ a technical group to supervise the mine contractor. After the first year of production ramp-up, the mine is expected to produce up to a relatively steady annual production of 66,000 tonnes of copper cathode, even during years of below-average feed grade, using on-site stockpiles from earlier, higher production years. Based on this processing capacity and the current reserves, the total life of the mine will be approximately 15 years.

Mine development will require a pre-production phase of 18 months for pre-stripping to expose sufficient ore to assure steady ore production. The open pit mining method will utilize drilling and blasting, loading with hydraulic excavators, and transport by trucks. The overall pit slope angle will be 28° in the upper and lower tertiary marl and sandstone, and 45° in the Paleozoic bedrock. Within the upper parts of the overburden, the marl will be mined without drilling and blasting. Trucks will haul the ore to the primary crusher located near the processing plant and will haul the rock waste to a surface storage facility where it will be covered by the marl overburden. Approximately 43 million tonnes of overburden will be mined during the development phase. In the later years of the project, partial backfilling of the pit with marl will occur.

The waste material handled in the course of the mining operations will be: (1), tertiary marl and sandstone; and (2) mine rock consisting of Paleozoic shale and volcanics. There will be four main dumps for waste: North, West, South and an in-pit backfill dump.

While the Las Cruces Project offers no unusual challenges from the mining perspective, there are some attributes of the property that influence some of the design decisions:

• The overburden layer, consisting primarily of marls, presents an unusually thick overburden for an open pit copper operation of this size.

• A good portion of the overburden is considered to be free-digging (i.e. does not require drilling and blasting).

• The average ore grades, at 6.6 percent copper, are unusually high.

• The total ore involved, at 16 million tonnes, is relatively low.

The open pit mineable reserves, pit designs, and production schedule were developed by Independent Mining Consultants of Tucson, Arizona (IMC). Equipment selection, labor staffing, and capital and operating costs were prepared for the open pit mine by DMT-Montan Consulting of Essen, Germany, based on the design work compiled by IMC. The mining plans include an 18-month pre-production plan and a 15-year life. The operation will commence backfilling the pit around the 9^th year of production and will continue until mine-out, followed by two years of partial backfill. Present plans call for the remainder of the pit to be used as a dump for inert fill.

Pincock, Allen & Holt 9033.05 July 8, 2004 17.17

17.8.2 Cutoff Grade

For the most part, the cutoff grade computations are derived from the original IMC estimates and assumptions. These design elements were not specified in the DMT report; however, use of the IMC values produces a breakeven cutoff grade that agrees well with that used in the Feasibility Study.

Typically, a cutoff grade is estimated by the engineers based upon the best state of knowledge at the time. A “first pass” is then developed on reserves, with the development of plans and more detailed operating costs. In most cases, the resulting operating costs are reasonably comparable to the initial assumptions, and the underlying mine plans require no modification. In cases where the estimated operating costs are materially different, (10 to 15 percent), it is prudent to consider a redesign of the pit based upon the more accurate numbers. Note that operating cost estimates are currently based on euros although the older design plan made use of dollar-based estimates.

Table 17-10 presents the basic economic assumptions used in the pit designs, along with a cutoff grade calculation. Note that specific refinements for haulage and pit slopes are added later.

TABLE 17-10

MK Resources Company

Las Cruces Project Technical Report

Basic Cutoff Grade Calculation

		Design & Reserve Basis				Feasibility Costs			Costs w/ $1.00 Cu
Description	Units	€		$		€		$	€		$
Mining Cost1.	per tonne ore	0.72		0.80		0.90			0.90
Processing Cost	per tonne ore	22.25		24.70		23.00			23.00
G&A Cost	per tonne ore	—		—		8.00			8.00
Breakeven Cost	per tonne ore	22.97		25.50		31.90			31.90
Internal Cost2	per tonne ore	22.25		24.70		23.00			23.00
Dollars / Euro	$/€	1.11
Copper Price	per pound Cu	0.72		0.80		0.80			1.00
Recovery	% Tot-Cu	91	%	91	%	91	%		91	%
Revenue per Pct.	per tonne per pct Cu	14.46		16.05		16.05			20.06
BE Cutoff3	%Cu			1.59		1.99			1.59
Internal Cutoff	%Cu			1.54		1.43			1.15

1. Mining costs for waste and overburden are variable as a function of depth.

2. The internal cost is the lowest cost per tonne needed to add to the contribution margin of the project (i.e. milling).

3. The value of the cutoff does not change as a function of the underlying currency

4. The euro is converted to dollars on a one-to-one basis.

It is apparent that the Design Basis differs considerably from the final costs estimated in the Feasibility (23.24 euros in the basis, versus 31.32 euros after detailed development, or 26 percent higher). While PAH notes that this level of difference would normally trigger a re-design of the mine plans, the difference is not material for the following reasons:

• The design of the mine is relatively insensitive to cutoff grade, due to the very high average grade and structurally constrained nature of the deposit.

Pincock, Allen & Holt 9033.05 July 8, 2004 17.18

• The mine operating costs used within the mine design are based upon first-principle cost estimates and contractor quotes, varying by rock-type and depth, rather than a simple average.

• The original price assumption of $0. 76 per pound of copper is lower than the current market price of $1.25 per pound (June 8 LME spot).

• The resulting design, while possibly sub-optimal as a result of the change in economics, is still feasible.

• The combined effect of higher operating costs and higher metal prices largely offset each other.

17.8.3 Pit Design

The mine production rate was established in an optimization study using an economic deposit model rather than a cutoff grade-based model. The basic underlying economic assumptions are presented in Table 17-10, with the exception of mining costs, which are calculated based on the formulae provided in Table 17-11. The ultimate pit design was based on a Lerchs-Grossman algorithm that produces an approximate pit design that is mathematically optimal for the design criteria.

TABLE 17-11

MK Resources Company

Technical Audit of the Las Cruces Project

Design Basis - Mine Operating Cost Assumptions

Rock Type	Elevation		Density dmt / m3	Cost			Formula
	Top msl	Bottom msl		At Top € / m3	At Bottom € / m3	Halfway € / dmt
Marl	100 15 -5 -55	15 -5 -55 -125	1.96 1.96 2.02 2.02	1.314 1.314 1.184 1.619	1.314 1.365 1.311 1.797	0.670 0.683 0.618 0.845	(0.1803 € /m3+1.06395 € /m3+ ((32.5m-15m)/10m+1)*0.02535 € / m3) (0.1803 € / m3+1.06395 € / m3+((32.5m-Block Elev.m)/10m+1)*0.02535 € / m3) (1.06395 € / m3+((32.5m-Block Elev.m)/10m+1)*0.02535 € / m3) (0.308 € / m3+1.06395 € / m3+((32.5m-15m)/10m+1)*0.02535 € / m3)
Sandstone	-100 -115	-115 -185	2.27 2.27	by dmt by dmt	by dmt by dmt	1.242 1.267	(0.547 € / dmt * Density dmt / m3) (0.558 € / dmt * Density dmt / m3)
Gossan	-125	-215	2.34	by dmt	by dmt	3.230	(1.378 € / dmt * Density dmt / m3)
Paleozoic Waste	-125	-215	2.71	3.652	3.880	1.390	(3.52545 € / m3+(1+(-95.1m-Block Elev.m)/10m+1)*0.02715 € / m3)
Paleozoic Ore	-125	-215	3.54	3.652	3.880	1.064	same as waste

The mining costs in Table 17-11 are developed from a combination of DMT calculations and contract mining vendor quotes. PAH has added approximate cost calculations to provide a reference for comparison to similar operations. All cost estimates were generated in euros.

Pit slopes have been developed by a series of consultants, with the most recent developed by Geocontrol (2003). The most recent variation is a more conservative design over the previous variations, requiring an overall slope of 28 degrees in the upper sediments and 45 degrees in the Paleozoic rocks. All studies reviewed by PAH have noted the need for extensive dewatering systems, which are incorporated into the design and cash flow.

Pincock, Allen & Holt 9033.05 July 8, 2004 17.19

Table 17-12 provides a summary of the design parameters used in the development of the ultimate pit.

TABLE 17-12

MK Resources Company

Technical Audit of the Las Cruces Project

Pit Design Parameters

Description	Value	Units	Comments
Slope Stability Parameters
Overall Slope in Marl	28	degrees	Geocontrol 2003
Overall Slope in Other Rock Units	45	degrees	Geocontrol 2003
Bench Height in Marl & Sandstone	10	meters	Design Decision
Bench Face Angle in Marl & SS	60-62	degrees	Geocontrol 2003
Bench Height in Ore Levels	5	meters	Design Decision
Bench Face Angle in Ore Levels	75	degrees	Geocontrol 2003
Ramp Parameters
Maximum Inclination in Marl & SS	8	percent	Better ramp angle for most trucks
Maximum Inclination in Paleozoics	10	percent	Good ramp angle for most trucks
Maximum Inclination for Temporary Ramps	15	percent	Acceptable for short distances only
Minimum Width for Two-Way Traffic	25	meters	3.0 x truck width + 5m (23.3 for Cat 777)
Minimum Width for One-Way Traffic	15	meters	1.5 x truck width + 5m (14.2 for Cat 777)
Blasting Requirements
Marl (assigned to lower 25%)	25	percent	Light blasting may improve production
Ore and Waste other than Marl	100	percent	Reasonable

The pit parameters are acceptable. The upper levels of the marl appear to be soft enough for free-digging, but may see somewhat lower productivities than would be achieved with blasted muck. PAH notes that 30 road widths have been incorporated into current to accommodate up to 135-tonne trucks.

The Lerchs-Grossman algorithm, while a powerful tool for determination of approximate pit geometry, does not produce a final design. The engineer must manually incorporate the roads and faces into the design, removing unstable slope structures from the design. DMT engineers developed the ultimate pit plan using the design parameters in Table 17-12. The DMT plan was subsequently refined by IMC, the results of which are presented in Figure 17-3.

Mining bench height will be 10 meters in the marl and 5 meters below the marl, to allow for improved grade control selectivity and minimization of dilution in the ore zone.

The haul road width, including berms, is 30 meters in the main haulage roads, narrowing to 25 meters in lower portions of the marl and primary rock, and narrowing further to 15 meters within small areas at the pit bottom. These widths are well within compliance with the industry accepted standard width of three to four times the haul truck width for the 135 tonne capacity trucks selected for waste mining. Primary haul road grades range from 7 percent, increasing to 12 percent over short distances in particularly tight areas at the deepest reaches of the pit. As with the road widths, these grades are compliant with industry standards.

Pincock, Allen & Holt 9033.05 July 8, 2004 17.20

FIGURE 17-3 Ultimate Pit Design

Pincock, Allen & Holt 9033.05 July 8, 2004 17.21

Crushed rock will be placed on the haul roads during their construction to provide a good road surface for the haul trucks. This practice will be particularly important in the marls, which are likely to become slippery when wet, or will work into impassable mud with the heavy traffic.

The ultimate pit, will be approximately 240 meters deep, 1.5 kilometers long (east-west), and 0.9 kilometer wide (north-south). The top of the HC mineralization is about 120 meters below the surface, accounting for the large amount of pre-production stripping required.

17.8.4 Mining Dilution and Losses

DMT has applied a correction factor of 0.97 to the ore tonnages expected from the pit to compensate for “porosity.” The effect of this correction is comparable to mining losses, although the correction is not described as such.

Dilution was incorporated into the mine model by adding 0.5 meter of hanging wall material and 1.0 meter of footwall material to the ore zone. The diluting grade was based upon the estimated grades of the diluting material. The hanging wall contact is visually obvious while the footwall contact will be determined by assay. No ore mining losses were included other than the porosity correction. The effective dilution for the in-pit material is provided in Table 17-13.

TABLE 17-13

MK Resources Company

Technical Audit of the Las Cruces Project

Effective Dilution Summary

Description	Total Tonnes x1000		Percent Copper	Tonnes Contained Copper
Undiluted In-Pit	15,016		6.999	1,051.0
Porosity Correction (.97)	15,110		6.920	1,045.6
Diluted Reserves	15,972		6.616	1,056.6
Diluting Material	862		1.279	11.0
Effective Percent Dilution	5	%

Given the high-grade and relatively well-defined ore-waste contacts, PAH feels that the dilution allowance is appropriate.

17.9 Ore Reserves Statement

Table 17-14 summarizes the resulting mining reserves, incorporating corrections and dilution.

Pincock, Allen & Holt 9033.05 July 8, 2004 17.22

TABLE 1-2 / 17-14

MK Resources Company

Technical Audit of the Las Cruces Project

Diluted Mineable Reserves

Description	Ore			Waste Tonnes x1000	Total Tonnes x1000	Strip Ratio w/o
	Tonnes x1000	Percent Copper	Lbs. Cu mm
Proven	13,938	6.897	2,119.3	—	13,938	n/a
Probable	1,432	6.416	202.6	—	1,432	n/a
Other Dilution	602	0.574	7.6	—	602	n/a
Waste	—	—	—	232,013	232,013	n/a
Total	15,972	6.62	2,329	232,013	247,985	14.53

Note: Mineable reserve is included in the measured and indicated mineral resource estimate.

The reserves presented in Table 17-14 are based upon:

• A cutoff grade of 1 percent copper,

• An engineered mining plan

• Allowances for both losses in mining and dilution.

The above reserves are based upon accepted engineering practice and, therefore, meet the standards for proven and probable reserves under the definitions of reserves as specified by both Canada and the United States. The mineral reserve is included in the measured and indicated mineral resource estimate.

The waste tonnage in Table 17-14 above does not include topsoil removal of approximately 499 thousand cubic meters. The cost of removing this material is considered within the cash flow.

PAH notes that the ultimate mine plans recover 95 percent of the available copper in the resource, which is much higher than the 60 to 70 percent average recovery generally experienced at other deposits. This recovery is understandable; however, given the high degree of structural control of the deposit and the very high grades present (6.5 % Cu). At the average grade of 6.62 Cu, an average ore block has a revenue potential of $146 per tonne ($1.00 copper price, no mill losses), or a net profit of $102 per tonne. Given these levels of potential profit, the open pit mining programs will extract virtually the entire leachable deposit.

Note that the primary sulfides below the pit bottom are not included within the reserve, as there is no metallurgical process currently planned to recover this material. There is some consideration for future underground mining of the sulfides; however, there are no current definitive plans to do so.

17.9.1 Effects on Reserves by Other Factors

A number of factors may affect the reserves of a given mining operation, but are not expected to unduly impact the plans for Las Cruces. These considerations are:

• Environmental Restrictions – the property is bounded on two sides by streams, which require setbacks or diversions. While the property plans must respect these requirements, the mine pit limits are not affected.

Pincock, Allen & Holt 9033.05 July 8, 2004 17.23

• Permitting Restrictions – No unusual permitting restrictions are in place that would change the reserves.

• Legal or Title Issues – Legal and Title issues are not anticipated to be troublesome under the Spanish Mining Law.

• Taxation – No unusual taxes are present.

• Socio-economic Issues – The project is located well away from population centers and other than increasing employment opportunities, no socio-economic issues are anticipated.

• Marketing – Copper is a commodity metal with numerous potential consumers within Europe.

• Other Issues – While dewatering is clearly a concern in the design of the mine, the engineering responses anticipated by CLC will not result in any material change to the mine plan or reserves.

The Las Cruces project is effectively a single-commodity mine, as is reflected in both the design and the cash flow projections. While there is potential for processing the gossan cap for gold, the plans are not sufficiently advanced to support a reserve, nor be included within the cash flow.

17.9.2 Recoverability

The Las Cruces project is effectively a single-commodity property recovering market-grade cathode.

Metal recoveries are a function of ore type and ore grade. The values for the three ore types, as derived from the test data, are as follows:

• HCH: Cu recovery = (0.0027 * Cu Grade) + 0.8873

• HCL: Cu recovery = (0.0027 * Cu Grade) + 0.9101

• HC4: Cu recovery = (0.0027 * Cu Grade) + 0.9241

The assumed price for copper is $0.76 per pound. This price was based on past long-term copper prices rounded to a conservative figure. Since the copper will be produced as cathode copper and sold as such, there will be no refining cost. The metal will be shipped directly to buyers and shipping costs are included, offset by cathode premiums.

Pincock, Allen & Holt 9033.05 July 8, 2004 17.24

18.0 OTHER RELEVANT DATA AND INFORMATION

PAH is not aware of any relevant data or information not already presented in this report.

Pincock, Allen & Holt 9033.05 July 8, 2004 18.1

19.0 INTERPRETATION AND CONCLUSIONS

19.1 Geologic Evaluation Conclusions

PAH finds that the degree of geologic evaluation of the Las Cruces deposit is generally reasonable for supporting a feasibility study. Geologic interpretation for the resource model was consistent with reasonable correlations between drill holes. The majority of the geologic data is from previous exploration conducted by Rio Tinto. The deposit geometry is reasonably defined from drill holes averaging around 50 meters spacing. The understanding of the copper grades, however, has been somewhat impacted by poor core recovery. The secondary sulfide mineralogy has been reasonably evaluated and is subject to wide grade variations over short lateral distance, as would be expected given the nature of the mineralization. A study is planned for the gold and silver mineralization in the gossan zone.

19.2 Exploration Program Conclusions

PAH finds that the drilling and sampling program by Rio Tinto on the Las Cruces deposit was reasonably conducted using industry standard procedures and techniques. Drill hole coverage is adequate for feasibility study, with some infilling recommended. Abundant evidence exists of the tendency to preferentially loss of chalcocite during the drilling and sampling, particularly in lower core recovery intervals, resulting in lower copper grades than is realistic. As a result, the confidence in the copper grades reported for samples in the lower core recovery zones is in question. The evidence indicates, however, that the copper grades should tend to be higher than reported. Analyses were conducted using standard procedures and were accompanied by reasonable check assay programs for quality control, which showed no significant grade bias or systematic analytical problems.

19.3 Resource Estimation Conclusions

PAH believes that the DMT secondary sulfide resource model was carefully created using standard engineering methods appropriate for this deposit. The model provides a reasonable representation of the distribution of the secondary sulfide mineralogic zones. PAH believes that the resource model, due to natural deposit variability and due to the excluded sample and composite data (about 30 percent of composite data excluded), may be a modest estimator of grade locally, but believes that on a global basis the estimation is reasonable. CLC plans for infill drilling over the life of the mine to increase the resource confidence for short-term mine planning. The resource estimate compares reasonably with those from previous resource models and grade differences are due to intentional changes in modeling methodologies. The model provides an acceptable basis for which subsequent mine engineering work can be conducted in order to delineate mineable reserves acceptable to the SEC or any other financial institution.

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20.0 RECOMMENDATIONS

While a number of recommendations have been made throughout the report, none are felt to be fatal flaws within the components of the feasibility study. The following matters were identified in the course of the review:

• PAH notes that CLC plans to drill an additional 40 to 50 core holes to improve definition and knowledge of the deposit. Of these, 14 holes were recommended prior to the commencement of mining, to improve the understanding of local variability and to validate the model. These holes have not been drilled at the time of this report. PAH recommends that this effort be expedited.

• Proceed on development of a resource and potentially a reserve for the gossan cap over the known ore zone.

• The ultimate pit as designed is somewhat steeper than the recommended slope on the eastern side of the pit. While there is a good possibility that the slopes will prove to be acceptable after mining has provided actual data, the pit should be modified to conform with the 28 degree slopes recommended by Geocontrol, or, alternatively, have Geocontrol reassess the recommended slopes.

• The mining plans envisioned by DMT require the use of 1.8 cubic meter loaders and 35 tonne trucks for ore mining. There is no demonstrated need for mining selectivity of this order, therefore PAH recommends that the ore mining be accomplished with 8.5 cubic meter loader and 91 tonne truck. This change should result in cost savings.

• The contract mining estimates should be resubmitted for binding estimates. While material changes in the rates are unlikely, binding quotes are normally required prior to project commencement.

• MK Resources Company has indicated that performance guarantees will be included within the EPC contract. PAH agrees that these guarantees should be in place.

• There are a number of permits outstanding in order for the project to proceed. To avoid potential delays, PAH recommends that CLC maintain the current emphasis on permitting, including land acquisition and supporting detailed engineering.

As noted above, none of these recommendations constitute undue risk, and are intended to clarify information and, in some cases, improve project performance.

PAH feels that the project feasibility study, as updated and described in this report, warrants further consideration, and should proceed with both financing and development.

Pincock, Allen & Holt 9033.05 July 8, 2004 20.1

21.0 REFERENCES

1. Bechtel International, Inc.; March 2001: Feasibility of the Las Cruces Deposit; internal consultant’s report.

2. DMT – Montan Consulting GmbH; 2003; Report on Geology & Resources of the Las Cruces Copper Deposit; internal consultant’s report.

3. DMT – Montan Consulting GmbH and Outokumpu – Technology Group; April 2004; Feasibility Of The Las Cruces Deposit; internal consultant’s report.

4. Doyle, Mike; July 2003; Chalcocite Loss From the Las Cruces Secondary Copper Orebody During Drilling and Sampling; internal company report.

5. IMC Group Consulting Ltd; August 2003; Review of Core Loss Procedures In Las Cruces Resource Modeling; internal company report.

6. Pincock, Allen & Holt; April 2004: Independent Technical Audit of the Las Cruces Copper Project, Southern Spain; internal consultant’s report.

7. Outokumpu Technology, Lurgi Metallurgie GmbH, DMT-Montan Consulting GmbH; November 2003: Las Cruces Copper Project Feasibility Study, internal consultants report.

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22.0 ADDITIONAL REQUIREMENTS FOR DEVELOPING OR PRODUCING PROPERTIES

CLC has a completed Feasibility Study that has been subjected to an independent diligence review. MK is in the process of securing funds to commence construction of the project, thus qualifying Las Cruces as a developing property.

22.1 Mining Operations

The Las Cruces deposit possesses several attributes that dictate the mining method employed. The ore body lies up to 150 meters below relatively flat-lying terrain. The upper 140 meters of the overburden is comprised of marls (relatively weak calcareous silts and clays), followed by a sandstone unit with substantial water. The ore deposit itself is comprised of high-grade partially-oxidized copper mineralization.

The depth of the deposit, in conjunction with the grade, suggests that underground methods be considered. CLC investigated the option, finding that the ground conditions and relatively poor mine recovery made underground mining unattractive.

The feasibility study is based upon open pit mining methods. The operation will be a typical truck and shovel operation, with drilling and blasting expected in the lower levels of the mine. The upper levels of the marl are expected to be free-digging (i.e. no blasting is required), an assumption that has been examined in detail both by equipment manufacturers and the independent engineer. The operating cost estimates assume that the upper 75 percent of the marl will be free-digging.

Successful development of the mine will require a substantial dewatering plan. Dewatering must be managed carefully at Las Cruces, as water is a precious commodity in southern Spain.

CLC plans to employ contract miners to develop the deposit. Spain supports a number of excellent contract mining companies, such as Cavosa, Peal and Sanchez Y Lago. The use of these companies is not unique to CLC; other mining companies, such as Rio Narcea Gold Mines have successfully employed a similar philosophy.

Operating cost estimates have assumed a mixed fleet of equipment, relying on 140 tonne trucks and 16 cubic meter hydraulic shovels for the overburden removal, followed by a combination of 91 tonne trucks and 8.5 cubic meter shovels. The feasibility has suggested a backhoe configuration for the shovels; however, the bench heights of 5 meters and 10 meters indicate a front-shovel configuration. The feasibility study further suggests a 1.9 cubic meter shovel coupled with 35 tonne trucks for ore movement, but there is no indication that this level of selectivity will be needed in ore mining.

Pincock, Allen & Holt 9033.05 July 8, 2004 22.1

The feasibility study suggested a 5-inch (127 mm) to 8-inch (203mm) capable drill of the D50 Tamrock class. This class of blasthole drill can readily drill a 5-meter bench, but will require multiple passes for a 10-meter bench.

22.1.1 Ore Processing

This section provides a brief description of the ore processing techniques, process plant equipment operation, process plant support facilities, and presents PAH’s comments regarding the plant design. Process plant capital and operating costs are discussed in other sections.

Summary

Metallurgical testwork began in 1996 by the previous owner, Rio Tinto, and then continued after CLC obtained the property in 1999. Rio Tinto found the ore difficult to treat by conventional flotation methods and it was determined that processes utilizing ferric sulphate chemistry to oxidize the copper minerals was likely to be the most successful technology. Rio Tinto investigated both atmospheric and pressure leaching technologies and determined that either option would be economically viable. CLC purchased the property in 1999 and conducted a pilot plant campaign and determined that the pressure leaching technology was more attractive. In January 2000, CLC contracted Bechtel to prepare a bankable feasibility study based on the pressure leaching technology developed by Dynatec. The study was completed in early 2001.

In January 2003, CLC contracted Lurgi Metallurgie GmbH/Outokumpu Technology Group (OTG) to prepare a revised technical and economic evaluation of the Las Cruces process plant and, in the process, optimize costs for the facility, improve copper recovery and simplify the process flowsheet. OTG reviewed the project information and then recommended a revised flowsheet that incorporated atmospheric leaching, Outokumpu Compact SX technology and other innovations to improve copper recovery and lower capital and operating costs. To confirm their predictions, OTG performed further metallurgical testwork (batch and mini-continuous pilot-scale) at their Outokumpu Research Center (ORC) in Pori, Finland to supply design criteria for the currently envisioned process facility. The OTG design was incorporated into the Feasibility Study of November 2003.

The ore processing facility is designed to operate 365 days per year, 24 hours per day and process ore at rates ranging from 2,000 to 3,000 tonnes per day (tpd) in years 1 through 3 and then up to 4,000 tpd beginning in year 4. The ore will come from the open pit mine and the ore grades will range from 5 to 10 percent copper. The copper in the ore is primarily found in chalcocite with some minor amounts found in chalcopyrite, tennantite-tetrahedrite complex and enargite. Approximately 66,000 tpa of copper cathode will be produced over the envisioned 15-year operation. The plant will be designed to have the ability to produce copper cathode at the instantaneous rate of up to 70,000 tpa.

The plant will utilize an atmospheric ferric-leach technology followed by solvent extraction and electrowinning; a processing system that is unique in the copper industry. The technology for this

Pincock, Allen & Holt 9033.05 July 8, 2004 22.2

project has been tested, developed and designed by OTG. A similar technology has been used over the past three years successfully in the zinc industry in Finland to directly leach zinc concentrates prior to electrowinning. In recent history, zinc concentrates have been subjected to oxidation by roasting prior to acid leaching and electrowinning.

The process plant envisioned for the Las Cruces project consists of unit processes and equipment that are commonly found in the mineral processing industry including crushers and screens, a grinding mill, agitation leach tanks, an oxygen plant, heat exchangers, cooling towers, thickeners, vacuum belt filters, a solvent extraction train, and an electrowinning tankhouse.

PAH understands that a ferric leach facility is currently being constructed in Laos for the Sepon Copper Project that is owned by Oxiana Minerals and Rio Tinto. The plant was designed by Bateman and Ausenco of Australia. OTG has signed a long-term copper “off-take” agreement with the owners and is also supplying the SX-EW equipment and technology package for that project. Oxiana will produce about 60,000 tpa of cathode copper from a chalcocite ore utilizing a technology that is similar to that which was initially planned for the Las Cruces project; autoclaves with an atmospheric ferric leach system.

A simplified process flow diagram for the planned project is presented in Figure 22-1. A plot plan of the ore processing facility is presented in Figure 22-2.

The major process equipment is listed in Table 22-1 and the principal processing parameters are listed in Table 22-2.

Comminution

Run-of-mine (ROM) ore at minus 1 meter in size will be crushed in three stages to minus 15 millimeters and then ground in a single-stage ball mill to the size of approximately 95 percent passing 150 microns prior to leaching. This grind size will allow adequate suspension of the ore in the leaching system, provide reasonable settling and filtration characteristics, and provide optimum copper recoveries.

The ROM ore will be placed in several 1,000 tonne capacity surface stockpiles ahead of crushing so that blending can occur to achieve consistent plant feed grades. A front-end loader will reclaim the ROM ore from the stockpiles and/or mine trucks delivering ROM ore can dump directly into a hopper that is equipped with a stationary grizzly and rock-breaker. Ore will be conveyed from the hopper and passed through a jaw crusher. Crushed ore will be conveyed to a double-deck vibrating screen. Top-deck and bottom-deck oversize materials will pass through separate cone crushers. Cone crusher products will then be recycled to the vibrating screen. Screen undersize/fines will be conveyed to the fine ore bin for storage. Wet-scrubber dust collectors will be provided for the crushing plant and conveyor transfers to control dust emissions. The crushing plant will be operated two 8-hour shifts per day during the first two years and then three 8-hour shifts per day thereafter. Plant equipment availability is estimated at 70 percent of scheduled time. The remainder of the

Pincock, Allen & Holt 9033.05 July 8, 2004 22.3

FIGURE 22-1

Pincock, Allen & Holt 9033.05 July 8, 2004 22.4

FIGURE 22-2

Pincock, Allen & Holt 9033.05 July 8, 2004 22.5

TABLE 22-1

MK Resources Company

Las Cruces Project

Major Process Plant Equipment List

Equipment:	Qty:	Description:
Rock breaker	1	At Dump Hopper
Dump hopper with grizzly	1	Effective volume 30 m3
Steel plate conveyor	1	11 m long by 1,2m wide
Jaw crusher	1	900 mm x 1 150 mm jaw type
Belt conveyor	1	84 m horizontal center distance
Belt conveyor	1	46 m horizontal center distance
Vibrating screen	1	4 m x 2 m double deck type
Cone crusher 1	1	Cone type 250 t/h
Cone crusher II	1	Cone type 160 t/h
Belt conveyor	1	63 m horizontal center distance
Fine ore bin	1	1000 m3 effective volume
Fine ore belt conveyors	3	30 – 120 t/h
Ball mill feed belt conveyor	1	32,5 m horizontal center distance
Ball Mill, wet	1	4,7 m diameter x 7,4 m length
Ore thickener	1	20 m diameter
Leach feed tank	1	75 m3
Atmospheric leach tanks with heating coils	2	500 m3 effective capacity
Atmospheric leach tanks	3	500 m3 effective capacity
Raffinate heat exchanger	2	123 m2 per unit
Atmospheric leach thickener	1	24
Cooling tower	2	310 m3 capacity
PLS holding pond	1	2000 m3 volume
Emergency solution holding pond	1	2000 m3 volume
Leach solution polishing filter	2	8459 mm x 1800 mm
PLS storage tank	1	100 m3
Leach slurry filter (tailings)	4	125 m2 each
Gypsum slurry filter	1	9,2 m2
Gypsum slurry filter	1	21 m2
Filter discharge conveyor	1	70 m length
Primary SX OutoCompact settler	5	326 m2 settling area each
Secondary SX mixer settler	1	33 m2 settling area
Electrolytic cells	154	85 anodes and 84 cathodes each
Cathode stripping machine	1	350 cathodes per hour
Electrolyte filter	2	4880 mm x 1525 mm
Electrolyte heat exchanger	2	138 m2 each
Raffinate pond	1	2000m3 volume
Pre-neutralization reactor	2	24 m3
Neutralization reactor	2	40 m3
Gypsum thickener	1	10 m diameter
Tailings thickener	1	25 m diameter
Polishing filter	2	6450 mm x 1800 mm
Boiler, gas fired	1	30 t/h
Lime day bin	1	500 m3
Ball mill for lime	1	1400 mm diameter x 4000 mm
Lime mixing tank	2	30 m3 each
H2SO4 storage tank	1	250 m3
Reverse Osmosis plant	1	35,3 m3/h
Oxygen plant (supplied by others)	1	5500 Nm3/h at 93% O2
Air compressor	2	830 m3/h
Flocculant system	2	—

Pincock, Allen & Holt 9033.05 July 8, 2004 22.6

TABLE 22-2

MK Resources Company

Las Cruces Project Technical Report

Principal Processing Parameters

Precipitation (Ave.)		mm/year	543
Evaporation (Ave.)		mm/year	1,160
Plant Operation
	Hours per day		24
	Shifts per day		3
	Weeks per year		52
Plant Availability (Design)
	Crushing		70%
	Grinding		90%
	Leaching/SX		90%
	EW Tankhouse		98%
Crushing
	ROM ore size	mm	<1000
	Feed Rate	t/hr	162 to 212
	Final Crush Size	p100 in mm	15
Grinding
	Feed Rate	t/hr	85 to 165
	Power Consumption	kW/t (desgn)	15
	Grind Product size	p80 in microns	80
	Grind Thickener size	m diameter	20
Leaching
	Leaching time	hrs	7
	Leach density	g/l solids	300 to 500
	Oxygen consumption	Nm3/h (93% O2)	4,888
	Leaching temperature	degrees C	90
	Iron oxidation	%	3
	Sulfur oxidation	%	1.7
	Tailings Thickener	m diameter	25
PLS
	Copper content	gpl Cu	40
	Ferric iron content	gpl Fe 3+	25
	Ferrous iron content	gpl Fe 2+	25
	Acid content	gpl H2SO4	25
Tailings Filters	Moisture (final)	% by weight	10
PLS Pond Retention Time		hrs	6.4
Raffinate Pond Ret. Time		hrs	7.1
PLS Cooling Tower Temperature		degrees C (out)	38

Pincock, Allen & Holt 9033.05 July 8, 2004 22.7

TABLE 22-2 (continued)

MK Resources Company

Las Cruces Project Technical Report

Principal Processing Parameters

Solvent Extraction (Primary)
	No. of Extractors		2
	Organic	% extractant	45
	Organic	% diluent	55
	Org/Aqueous ratio		4 to 1
	Raffinate	gpl Cu	11 to 13
	Cu Recovery	%	70
	No. of Washing units		1
	No. of Stripping units		2
	Lean Electrolyte	gpl Cu	36
	Rich Electrolyte	gpl Cu	52
Electrowinning
	No. of EW cells		154
	Cathodes per cell		84
	Anodes per cell		85
	plating area per cathode	sq. m./side	1
	Current density (nominal)	amps/sq. m.	280
	Current efficiency (nominal)	%	92
	Cathode cycle	days/harvest	7
	Electrolyte temp.	degrees C	48
	Electrolyte flow	l/min/cell	387
	Copper depletion	gpl Cu	2
	Stripping Machine	blanks/hr.	250
	Tankhouse Capacity	tpa Cu cathode	66,000
	Tankhouse Capacity (max.)	tpa Cu cathode	70,000
Bleed Treatment
	Raffinate bleed Flow	cu. m./hr.	30
	Raff. iron content	gpl Fe	50
	Raff. acid content	gpl Sulfuric Acid	70
	Raff. zinc content	gpl Zinc	8
	CaO feed rate	tph	4.2
Neutralization
	Pre-neutralization	pH	2
	Gypsum thickener	m. diameter	10
	Gypsum filters	type	vacuum belt
	Neutralization	pH	10
	Neutralization	retention time Hrs	1
	Effluent acidification	pH adjustment	5.5 to 9
Gypsum filter	Moisture	% by wt.	50
Solvent Extraction (Secondary)
	No. of extractors		1
	PLS	gpl Cu	5
	Raffinate	gpl Cu	0.5

Pincock, Allen & Holt 9033.05 July 8, 2004 22.8

process facilities will be operated three 8-hour shifts per day with the equipment availability estimated at 90 percent.

Crushed ore will be withdrawn from the 2,200 tonne live capacity fine ore bin using variable speed belt feeders and belt conveyor, then combined with recycled process water and then ground in a ball mill. The ball mill will be rubber lined, charged with steel grinding balls, and equipped with a 2,650 kW motor and variable speed drive so that the required fluctuations in mill throughput can be accommodated. Mill discharge slurry will be pumped to hydro-cyclones for size separation. Cyclone underflow will be recycled to the ball mill for further grinding; while the cyclone overflow at 35 percent solids by weight and 95 percent passing 150 microns in size, will pass to the grinding thickener.

Thickener overflow water will be pumped to a process water head tank and then recycled to the grinding circuit or used as other process water makeup. Thickener underflow slurry will be leach circuit feed. The crushing and grinding circuits are designed to process ore at varying rates due to the variations in ore feed grades anticipated to come from the mine. The SX/EW systems are designed/limited to a nominal 66,000 tpa of copper cathode production, therefore, as the ore grade fluctuates, ore feed rates vary to maintain a relatively constant copper metal feed rate to the SX/EW system.

Atmospheric Leaching

Previous studies indicated that copper could be leached from the Las Cruces ore under atmospheric conditions and potentially without the installation of an autoclave. This led to the development of the OTG leaching process that was confirmed during the pilot plant testing at ORC in Finland. The atmospheric leaching system consists primarily of the leach reactors and the tailings (leach residue) thickening and filtration systems. Outokumpu has considerable experience in leach reactor design and operation beginning in the 1960s for cobalt, nickel matte, nickel concentrates, and zinc. More recently, in 2001, Outokumpu constructed a direct ferric leach plant for treating zinc concentrates using 900 cubic meter atmospheric leach reactors and vacuum belt filters for separating zinc solutions from the leached concentrates and from the gypsum residues. This technology uses similar equipment and control strategies to what Outokumpu has planned for Las Cruces.

Thickener underflow slurry will be pumped to the leach feed slurry tank where it will be mixed with sulfuric acid, defoamer, and a lean electrolyte bleed stream from the EW circuit. Mixed slurry will then be pumped to the first of five atmospheric leach reactors where it will be mixed with oxygen supplied from a third party installation and 90 degree centigrade heated raffinate solution that has been recycled from the SX circuit. The retention time in the leach reactors will be nominally 7 hours. The solids content in the slurry in the leach reactors will typically be about 500 gpl; however, as the ore grades increase, the solids contents will drop to as low as 300 gpl to maintain copper production at a constant level and maintain the 7-hours leach retention time.

Pincock, Allen & Holt 9033.05 July 8, 2004 22.9

Leaching of the ground ore to recover about 91 percent of the contained copper will take place in a series of five 500-cubic-meter-capacity reactors (OKTOP design) using an acidic oxidative iron solution with a ferric iron concentration of about 25 gpl and a total iron concentration of about 50 gpl. Ferrous iron will be oxidized back to ferric iron during leaching by the oxygen addition to the tanks. The first two reactors will be equipped with heating tubes to initiate or maintain adequate heat in the leaching system. Each steel reactor tank will be lined with rubber and have an overlay of ceramic bricks. This type of reactor lining has been successfully used for many years at operations in Finland. The reactors will be equipped with axial/radial double 6-blade agitators driven by 275 kW motors. This type of agitation is anticipated to achieve oxygen efficiencies/utilizations within the reactors at the design level of 90 percent.

Tailings Dewatering

The leached residue slurry will flow to a 25-meter diameter tailings thickener. Thickener overflow is pregnant leach solution (PLS) and will be pumped through two cooling towers to drop the temperature to less than 38 degrees centigrade prior to passing to a 2,500 cubic meter capacity PLS storage pond that is concrete and lined with HDPE with a retention capacity of above 6 hours. An emergency pond similar to the PLS pond will be constructed to retain plant spills and act as an emergency PLS pond, if required. Gypsum will be formed as the PLS is cooled. The gypsum sludges cleaned from the cooling towers and precipitated in the storage ponds will be collected and pumped back into the tailings thickener. Cooled PLS from the pond will be pumped through two polishing filters/clarifiers operating in parallel to remove remaining suspended solids and then directed to a PLS storage tank prior to being pumped to the solvent extraction circuit. Solids collected from the filters will be collected and pumped to the tailings thickener.

Tailings thickener underflow slurry will be pumped to a tailings filter feed tank where it will be combined with clay sludge from the organic pretreatment and crud treatment systems. Slurry will then be pumped to four 125 square meter vacuum rubber-belt filters operating in parallel that will have the capacity to wash and filter tailings at the highest (4,000 tpd) plant throughput rate. Filters will be designed and constructed of materials capable of handling hot slurries and the acidic conditions anticipated for the process. Filter cake, at about 10 percent moisture by weight, will be conveyed to the tailings filter cake storage area and then transported by truck to the disposal area. Filtrate waters will be collected and recycled to the PLS pond or pumped to the Secondary SX circuit.

Solvent Extraction

The Primary SX circuit will consist of one train having two extraction mixer-settlers operating in series, two stripping mixer-settlers, one organic washing mixer-settler and a loaded organic tank. The mixer-settlers will be efficient OutoCompact type units that are deeper than conventional units and consist of vertical uptake channels, guiding dispersion feed distributors, non-jetting picket fences, aqueous inner circulation channels, and efficient launders for the organic and aqueous phases. Outokumpu Vertical Smooth Flow (VSF) technology will also be incorporated into the mixer-settler design. The VSF technology consists of one Dispersion Overflow Pump (DOP) unit and two Spirok (double-helical) mixers with covers to minimize air entrainment.

Pincock, Allen & Holt 9033.05 July 8, 2004 22.10

PLS containing 40 gpl copper and 25 gpl sulfuric acid will be pumped from the storage tank to the first extraction unit where it will be mixed with semi-loaded organic (mixture of a kerosene-type diluent and a copper selective organic reagent) coming from the second extraction unit. The settler will then allow the organic and aqueous phases to separate. Aqueous (low-grade PLS) coming from the first extraction unit passes to the second extraction unit where it is mixed with stripped organic. The aqueous solution leaving the second extraction unit is raffinate solution that contains about 67 gpl sulfuric acid and 11 gpl copper meaning that about 70 percent of the copper that entered the extraction units initially in the PLS was removed/extracted by the organic. The aqueous, or raffinate, coming from the second extraction unit will flow by gravity to the 2,000 cubic meter capacity raffinate storage pond prior to recycle to the atmospheric leach circuit.

Loaded organic coming from the first extraction unit will pass to the Loaded Organic tank that will be equipped with scrub water circulation to enhance the removal of impurities, such as manganese and iron, from the organic. Scrubbed organic is then pumped to the Wash Settler that is a similar design to the extraction units. Lean electrolyte bleed solution and sulfuric acid are also added to the scrubbed organic prior to being treated in the Wash Settler. The washed-loaded organic is then pumped to the first stripping mixer-settler unit where it is contacted by semi-rich electrolyte solution coming from the second stripping unit. Copper is transferred from the loaded organic into the electrolyte due to the differences in acidity between the phases. The aqueous (rich electrolyte) contains about 50-gpl copper and 145-gpl sulfuric acid. The organic is pumped to the second stripping unit where it is contacted by the lean electrolyte that contains about 175-gpl acid and 35-gpl copper. The stripped organic is pumped back to the first extraction unit. Rich electrolyte flows by gravity to the Tank Farm area.

Raffinate bleed is required to control the amount of iron and zinc in the leach circuit solutions. This bleed solution stream will contain high levels (11 gpl) of copper; therefore, recovery of the copper prior to treatment and discharge is warranted. The treated bleed stream, called Secondary PLS, at a pH of 2.0 coming from the pre-neutralization circuit is pumped through polishing filters and then to the Secondary SX circuit that consists of a single mixer-settler unit. The bleed stream is mixed with a portion of the stripped organic coming from the stripping circuit to remove about 87 percent of the copper contained in the bleed stream. Loaded organic is then pumped to the Loaded Organic tank. Raffinate from this extraction circuit contains about 0.5-gpl copper and is pumped to the secondary raffinate storage tank prior to treatment.

Any crud (mixture of organic/iron/suspended solids) that is collected from the mixer settlers or the raffinate pond is collected and pumped to the organic treatment system. Crud will be mixed with bentonite, other organic requiring treatment and sulfuric acid and then pumped through a filter. Treated organic passes through the filter and is collected and recycled. Solids from the filter are collected and pumped to the leach residue filters for disposal.

Pincock, Allen & Holt 9033.05 July 8, 2004 22.11

Neutralization

Primary raffinate bleed will first be pre-neutralized to pH 2.0 using milk of lime, thickened, and thickener overflow pumped to the Secondary SX mixer-settler to recover copper as described above. After copper removal, the raffinate is pumped back to the neutralization circuit, the pH of the solution increased to about 10.0 with milk of lime and then pumped to the neutralization thickener.

The pre-neutralization thickener underflow is pumped to a vacuum rubber-belt filter to recover gypsum as a cake. The neutralization thickener underflow is also pumped to a separate filter to recover the metal hydroxides and gypsum as a cake. Filter cakes will be conveyed to the tailings residue storage area, combined with tailings and then trucked to the disposal area.

Neutralization thickener overflow is considered treated effluent and will be discharged via a holding pond and pump system.

SX-EW Tank Farm and Reagents

Rich electrolyte will flow to an after-settler that will be equipped with coalescing fences to separate the electrolyte from entrained organic. The organic will be collected from the top of the solution layer and pumped to organic the treatment system. Rich electrolyte will then be pumped through two polishing filters containing sand and anthracite to remove any remaining solids and organic. The filters will be backwashed periodically with lean electrolyte with the backwash reporting to the Filter Backwash Tank. Organic is skimmed from the tank and pumped to the organic treatment system. Electrolyte is collected and recycled to the process.

Filtered rich electrolyte, at 52-gpl copper and 38 degrees centigrade, then flows to the two-compartment Electrolyte Circulation Tank. In one compartment of the tank, rich electrolyte will be mixed with lean electrolyte, at 36-gpl copper and 50 degrees centigrade that will be returning from the electrowinning cells, at a ratio of about 1 part rich electrolyte to 6 parts lean electrolyte. The mixture (EW electrolyte feed solution), at 37-gpl copper and 48 degrees centigrade, is then pumped to the electrowinning circuit. Lean electrolyte is pumped from the circulation tank through water-cooled heat exchangers to decrease the solution temperature to about 48 degrees centigrade and then to the organic stripping portion of the solvent extraction circuit.

The EW circuit will require the addition of two reagents, Ammonium Lignosulphonate (Guar) and Cobalt Sulfate Heptahydrate. Guar will be used as a cathode surface smoothing and densifying agent and Cobalt Sulfate Heptahydrate will be used for preventing anode corrosion and will stabilize the lead oxide layer on the anode surfaces thus minimizing lead contamination of the copper cathodes. Both of these reagents will be received in bags and then mixed with reverse-osmosis water (RO) in separate mixing and feed systems. Both reagents will be added to the EW feed electrolyte stream with Guar being added at the rate of about 150 to 250 grams per tonne of copper produced and cobalt sulfate added to maintain a solution concentration of about 100 mgpl cobalt.

Pincock, Allen & Holt 9033.05 July 8, 2004 22.12

Crud is generally formed at the interface of organic and aqueous solutions in SX circuits. Suspended solids and other impurities add to the formation of crud. Crud will be removed from the settlers, on a batch basis, using pneumatic diaphragm pumps and pumped to the Crud Collection Tank where it will be mixed with acidic water to aid in phase separation. The phases will be allowed enough time to separate. Organic will be pumped from the tank through the clay filter that will also be used for organic treatment. Organic treatment will be part of the plant design with all organics recovered from the filter backwash and rich electrolyte settler tank being filtered through a clay filter before being recycled to the SX circuit. Acid activated bentonite clay will be used in the clay filter. Solids removed from the clay filter will be slurried and pumped to the leach residue filters for disposal.

Fresh diluent will be delivered by truck and stored. As needed, the diluent will be withdrawn from the storage tank and mixed with the recycled/treated organic.

A holding tank for various tank farm drainages/spills and will be equipped with pumps to recycle the materials to the appropriate circuits in the process.

A fire protection system will be provided and will consist two foam storage tanks, foam generators, piping around the SX equipment, sprinkler nozzles, flame detectors and manual system buttons.

Electrowinning

Copper will be recovered from the electrolyte solution in EW cells by passing a direct current between lead anodes and stainless steel cathodes. Electric current will be fed from two water-cooled transformer-rectifiers having a capacity of 53 kA in total. The transformer-rectifiers will be designed for a nominal load of 280 amps per square meter and a current efficiency of 92 percent. The peak load will be equivalent to 315 amps per square meter at a current efficiency of 90 percent.

The EW tankhouse will be designed in a “single-aisle” configuration with two rows of 77 cells each. A single electrical and single electrolyte circulation system will be provided. The cells will be constructed of polymer-concrete and will contain 85 anodes and 84 permanent stainless steel cathodes each. Each cathode will have a copper deposit area of 1 square meter and will be equipped with plastic edge strips to facilitate easy removal of the copper. Anodes will be made of an alloy of lead, calcium, and tin. Each cell will be covered by acid mist collection hoods to capture gases and acid mists generated during the electrowinning process. Off-gases and mists will be scrubbed.

Electrolyte will be pumped from the circulation tank to all of the cells, 2 to 3 gpl of the contained copper will be removed from the solution with the lean electrolyte gravity flowing back to the

Pincock, Allen & Holt 9033.05 July 8, 2004 22.13

circulation tank. Copper cathodes will be grown for 7 days then removed from the cells. During removal, the cathodes and cell bus bars will be washed automatically by an automated crane system. After washing, the cathodes will be transported to the 300-cathode per hour capacity stripping machine where they are washed again and then stripped automatically. Stripped steel cathodes are returned to the EW cells. Copper cathode is automatically stacked, sampled, weighed, bundled and marked prior to being transported by forklift to the cathode storage area.

Lime Slaking

Approximately 100 tpd of calcium oxide (CaO) will be delivered by truck and placed in a storage bin having 3 days capacity for the operation. CaO will be removed from the bin via a screw conveyor, combined with process water, and ground in a ball mill that is in closed-circuit with cyclones. Cyclone overflow will report to the Lime Mixing Tanks while cyclone underflow is recycled to the ball mill. Slaked lime slurry will be pumped to the neutralization circuits as required.

Utilities, Consumables, Reagents

A reverse-osmosis (RO) water treatment plant will be provided to supply de-mineralized water for process needs and will consist of filters, the RO unit, and a storage tank.

A gas-fired boiler will be provided to produce about 25 tph of steam for heating the leach reactors, raffinate, and electrode washing water in the EW tankhouse. RO water and condensate from the heat exchangers will be used to feed the boiler.

Approximately 150 to 200 tpd of 93 percent pure oxygen will be supplied for the process plant “over the fence” by a third-party from either a cryogenic or VPSA unit. The oxygen is required for the atmospheric leaching process and will be added into the OKTOP reactors. Space has been provided for the oxygen production facility adjacent to the leaching area. Power and water for the oxygen plant will be supplied to the contractor from the process facility. A budgetary quotation for oxygen supply was provided for the feasibility study by an oxygen supplier (Praxair) in October 2003.

Sulfuric acid and hydrochloric acid will be delivered by truck and stored in separate single tanks. Sulfuric acid will be pumped to the various leaching, washing and electrowinning processes as needed and hydrochloric acid will be used to adjust the treated effluent stream to a pH of 7.

Two flocculant mixing and distribution systems will be provided, one type for thickening of the ore and tailings slurries and the other type for neutralization.

Defoamer will be pumped directly from barrels to the leach feed slurry tank to eliminate foaming in the leach reactors.

Pincock, Allen & Holt 9033.05 July 8, 2004 22.14

Plant air will be supplied from two 830 cubic-meter per hour capacity air compressors, one operating and one on standby.

The process plant design also incorporates the installation of the Process Control Room, Process Plant Shop, and Laboratory located on the ground floor of the filter building. The facilities are masonry walls with concrete slabs for roofs. The Control Room, Shop, and Laboratory will be 150, 50, and 75 square meters each, respectively, and with the Control Room and Laboratory being air-conditioned.

Supplies for power and makeup water are discussed in Section 8 of this report.

Mass and Energy Balances

Outokumpu prepared process flow diagrams and detailed mass and energy balance data for the process plant design. The balances were prepared based upon the highest plant throughput (1.3 million tpa), plant availability of 90 percent, copper recovery of 92.5 percent, and copper cathode production of 66,000 tpa. The amount of solution bleed and treatment was set to maintain a maximum of 50-gpl iron in solutions and copper losses in solutions contained in filtered residues and neutralization products were set at 0.7 percent of input.

Oxygen transfer efficiency (from gas to dispersion in the slurry) in leaching was assumed at 90 percent. The oxygen will be used mainly to convert ferrous sulfate to ferric sulfate, which in turn leaches the chalcocite minerals to put copper in solution as copper sulfate and produce ferrous sulfate which is then converted to ferric sulfate by oxygen and sulfuric acid.

Sulfuric acid demand of about 2 tonnes per hour was calculated based upon leach circuit reactions and bleed streams from the process.

Steam requirement of 25 tph was estimated based upon the heat necessary to increase the leach slurry temperature to 90 degrees centigrade, the heats of reaction in the leach circuit, the temperatures of recycled raffinate solution, and miscellaneous plant heating requirements.

The heating requirements and startup times were also estimated for a plant “cold startup.” The analysis indicates that if the entire plant were starting up after a long shutdown, based upon full inventories of solutions and a steam plant that could produce 30 tph of steam, the startup time would be about 33 hours (1.5 days). The SX-EW portion of the plant would be able to return to design operating temperatures within several hours while the atmospheric leaching circuit would be the bottleneck and require about 24 hours to reach design operation temperatures. Because of this delay, there will be some process inefficiencies until the plant is stabilized. Typically, planned long-term (more than three or four days) shutdowns occur no more than two or three times per year.

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Tailings Storage Facility

Two types of tailings streams will be produced and will require disposal: filtered leach residue and filtered solids from the neutralization circuit. The leach residue will be about 92 percent of the total tailings and shows a strong potential for producing acid and metal-bearing leachates based on testing results. The leach residues are made up of about 80 percent sulfide minerals, primarily as pyrite with minor amounts of chalcocite and chalcopyrite along with constituents like arsenic and mercury. The neutralization solids will be about 8 percent of the total tailings, are inert, and contain mainly gypsum and magnetite with some metal hydroxides.

Geotechnical testwork performed on samples of tailings streams generated from pressure-leach metallurgical testing for the previous feasibility study indicated that the in-place density and moisture content of the combined tailings would be 2.3 tonnes per cubic meter and 7.5 percent moisture by weight, respectively. Preliminary calculations from tests on samples of tailings from the current processing scheme indicate that the bulk density and moisture of the combined tailings stream will be higher and at 2.89 tonnes per cubic meter and 16 percent moisture, respectively, due to design changes in the filtering and neutralization circuits. CLC states that additional geotechnical testwork on samples from other process related confirmation testing will be performed to confirm the tailings characteristics for the final tailings disposal area design effort; however, no adverse impacts on the tailings placement plans or design are anticipated to come from further laboratory work.

The Tailings Storage Facility (TSF) is embedded within the North Dump with the tailings material being confined by a compacted marl embankment, bottom and upper seals. The bottom seal will consist of 1.5 meters of compacted marl covered by a 1-millimeter thick HDPE liner and a 0.5-meter thick layer of sand and a geotextile material to allow moisture to pass into the sand layer. Any seepage collected in the sand layer will be directed to a seepage collection pond and then pumped to the Contact Water Pond at the process facility. The tailings will be delivered to the TSF by 40-tonne dump trucks, dumped, and then the material will be dozer-spread and roller-compacted in 0.5-meter thick lifts to an ultimate design density of 90 percent of standard Proctor maximum dry density and to a thickness of about 25 meters.

Once the tailings are stacked to their ultimate height, over 5 meters of compacted marl material will be placed on the top and then covered by geotextile material, 0.6-meters of sand, more geotextile, 5 meters of uncompacted marl and then topped with 0.5 meters of topsoil to grow vegetation.

The embankment or barrier berms will be built to a height of 25 meters and constructed of upper marl material from the mine pit development. The marl that will not come in contact with tailings will be compacted to a minimum of 95 percent of standard Proctor density while marl that will come into contact with tailings will be compacted to a minimum of 98 percent of standard Proctor density. Pore pressures within the berms are not expected to be an issue; however, piezometers will be installed to monitor pressures.

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PAH Comments

• PAH has reviewed the ore processing plans for the project and found no material deficiencies in the process plant design or process technologies selected. PAH believes that the plant layout and equipment selections are reasonable and adequate and that the plant should be able to perform and operate as planned. OTG has prepared process flow diagrams, mass and energy balances, plant layouts and drawings of sufficient detail and accuracy to enable the preparation of a capital cost estimate at the level of accuracy as stated in the feasibility study.

• In PAH’s review of the comminution circuit design and layout it was suggested that since only 13 to 18 hours worth of fine ore storage is available that consideration be given in the final design to having a crushed fine ore bypass directed to an emergency stockpile that can be reclaimed and put back into the fine ore bin so that the rest of the facility can be operated during extended crushing/screening plant repairs. Also, the possibility of utilizing similar sized cone crushers for the second and third stages of crushing might minimize the number of spare parts inventory required.

• PAH reviewed the equipment duplication and emergency bypass philosophy as well as retention times necessary for “process breaks” in case of emergency shutdowns for different plant circuits. PAH believes that there are sufficient numbers of parallel operating critical process equipment units (i.e., leach reactors, pumps, solution clarifiers, tailings residue filters, transformer-rectifiers, etc.), adequate solution storage capacities (6 to 8 hours) between the leaching circuit and the SX circuit, and equipment bypass piping to accommodate short-term circuit shutdowns.

• PAH toured the OMG Kokkala electrowinning facility located in Kokkola, Finland that produces up to 20,000 tpa copper cathode. The copper and cobalt leaching plant was designed and constructed by Outokumpu in 2001 and contains similar equipment to that being proposed for the Las Cruces project. The plant contains EW cells of similar size, cell covers with acid mist collection system, an automatic crane with cathode/bus bar washing system, and the automatic stripping machine technology. PAH was impressed with the few numbers of workers required to operate the plant, the control systems and the overall cleanliness of the facility and efficiency of operation. This technology would be more than adequate for the Las Cruces project.

• The ability of the OKTOP leach reactors to obtain a 90 percent oxygen transfer efficiency was reviewed and discussed with OTG and it was determined that this was an “optimum” figure. The final reactor design and selection of the optimum reactor agitators should optimize the transfer efficiency; however, it would not be out of the ordinary for the project to experience a figure in the 80 to 85 percent efficiency range which would increase the oxygen requirement and operating costs slightly.

• PAH understands that additional testwork will be performed prior to completing the basic engineering and final equipment selection to confirm the final reactor design and the power and materials of construction requirements for the reactor agitators due to the erosive and corrosive atmosphere within the leach reactors.

Pincock, Allen & Holt 9033.05 July 8, 2004 22.17

• Because of the high copper values expected in the PLS, the SX circuit has been designed to have extractant at 45 percent by volume in the organic (extractant and diluent combined). This level is higher than the 30 to 35 percent by volume typically run at other facilities that operate at much lower PLS grades; however, according to OTG, this percentage could be lowered somewhat without adversely impacting plant performance, if necessary.

22.2 Production Schedule

Las Cruces mining will begin with an 18-month pre-stripping program that should produce a nominal amount of ore for startup in the final months. Ore production will commence at the beginning of year 1 of operations and will continue for almost 15 years. The mining plan is somewhat unique in that later-phase waste stripping will be dumped at the west side of the pit, thus shortening the haulage distance and simplifying the reclamation process at the same time.

The mining production schedule is developed from a series of seven phases, designed such that ore will be exposed beneath the thick overburden without requiring excessive stripped inventories. The geometry of the deposit is sufficiently well known that ore exposure is not likely to be a problem.

The production schedule is somewhat unusual, in that the feed tonnage to the mill is not the driving factor in the schedule. For Las Cruces, the electro-winning section will be the primary constraint, limiting total production to 66,000 tonnes of cathode per year. A secondary limitation is the crushing and grinding section at 1.3 million tonnes of ore per year. Within these limitations, the mine may produce more tonnage at lower grades, or fewer tonnes at a higher grade.

Table 22-3 presents the production schedule.

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TABLE 22-3

MK Resources Company

Las Cruces Project Technical Report

Mine and Mill Production Schedule

	MINE PRODUCTION								MILL PRODUCTION
Prod. Year	Total Ore	Model Diluted Copper		Recov. Copper		Total Waste	Total Mined	Strip Ratio	Feed to Plant	Model Diluted Copper		Recov. Copper		Tonnes of Copper
	k-dmt	%Cu		%Cu		k-wmt	k-dmt	w/o	k-dmt	%Cu		%Cu		dmt
-2	—	—		—		17,000	17,000	—
-1	—	—		—		34,195	34,195	—
1	618	8.60	%	7.66	%	22,948	23,566	37.1	618	8.60	%	7.66	%	47,354
2	1,043	6.97	%	6.32	%	14,454	15,497	13.9	1,043	6.97	%	6.32	%	65,934
3	1,065	6.82	%	6.20	%	17,605	18,670	16.5	1,065	6.82	%	6.20	%	66,053
4	758	9.47	%	8.70	%	16,093	16,851	21.2	758	9.47	%	8.70	%	65,953
5	1,105	6.55	%	5.98	%	10,158	11,263	9.2	1,105	6.55	%	5.98	%	66,054
6	1,075	6.68	%	6.15	%	15,994	17,069	14.9	1,075	6.68	%	6.15	%	66,129
7	1,382	7.55	%	7.00	%	10,643	12,025	7.7	953	7.48	%	6.94	%	66,091
8	1,215	6.43	%	5.89	%	16,613	17,828	13.7	1,138	6.34	%	5.80	%	65,993
9	1,141	5.48	%	4.97	%	10,188	11,329	8.9	1,277	5.69	%	5.17	%	66,059
10	1,872	7.54	%	6.92	%	10,624	12,496	5.7	933	7.72	%	7.08	%	66,047
11	990	6.52	%	5.98	%	9,275	10,265	9.4	1,094	6.61	%	6.04	%	66,034
12	1,100	4.66	%	4.27	%	9,835	10,935	8.9	1,296	5.54	%	5.10	%	66,057
13	1,016	6.14	%	5.66	%	9,395	10,411	9.2	1,168	6.14	%	5.66	%	66,074
14	818	6.51	%	5.87	%	5,303	6,121	6.5	1,138	6.40	%	5.81	%	66,095
15	774	3.22	%	2.95	%	1,690	2,464	2.2	1,311	5.00	%	4.59	%	60,136
Total	15,972	6.62	%	6.05	%	232,013	247,985	14.5	15,972	6.62	%	6.05	%	966,061

22.3 Recoverability

The mineable recovery of the Las Cruces Deposit is unusually high for copper deposits as a result of the very high grades. Open pit copper mines typically average between 0.30 percent copper for deposits in the southwestern US to over 1 percent in the Chilean and Peruvian deposits. Las Cruces, by comparison, has an average diluted grade of 6.62 percent copper, recovering well over 95 percent of the resource.

Following review of the test-work and plant design, PAH feels that process recoveries in excess of 90 percent are achievable. Further discussion of the metallurgical recovery has been provided in Section 16 of this report. The copper production listed in Table 22-3 accounts for plant losses.

22.4 Markets

The Las Cruces project would produce approximately 66,000 tonnes of London Metal Exchange (LME) Grade A electrowon copper cathode annually over a projected mine life of 15 years. The cathode production would be suitable for all copper fabricators, including those that produce sheets, strip, tube, rod, wire, castings and forgings. Equally important, after a registration process, the copper cathode would be deliverable to terminal markets against copper futures contracts and be attractive to merchants.

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Spain does not mine enough copper to be self sufficient in refined production and relies heavily on the import of copper concentrate. By smelting imported concentrates, Spain is able to achieve a balance between production and consumption of refined copper. However, Europe is in a deficit and offers a ready market for any cathode CLC would not sell in Spain.

The largest end-use markets for refined copper are: heat exchange applications, plumbing, and electrical power generation and delivery. World demand for these products and services has increased, thus increasing demand for refined copper. World production and consumption of refined copper has shown an upward trend during the last five years. Table 22-4 shows historic production and consumption.

TABLE 22-4

MK Resources Company

Las Cruces Project Technical Report

Historic Refined Copper* (000 tonnes)

	1998	1999	2000		2001	2002
World Production	13,865	14,332	14,631		15,539	15,121
World Consumption	13,409	14,077	15,051		14,376	14,392
Difference	456	255	(420	)	1,163	189

* Source: CRU International

Copper cathode is priced on the basis of official commodity futures exchange quotes. CLC will receive the LME quote. The Monthly average of the spot LME closing price for the month of delivery would be used as the base price. As an alternative to the average pricing, customers would also be allowed to fix the price of all or part of their monthly shipment on any trading day during regular LME trading hours, using the spot LME price. Figure 22-3 presents a graph of LME spot metal prices since 1999.

The cash flow basis of $ 0.95 per pound is somewhat higher than the three-year historical average $0.796 per pound, but is within reason considering that the basis is substantially below the prevailing (June 8) copper price of $1.25 per pound. PAH feels that the metal prices are reasonably supported. Note that the conversion rate for euros to dollars is assumed to be an average of 1 to 1 by MK Resources Company.

Premiums are an amount charged over the base price of copper, and are typically negotiated annually. Since much of the copper consumed in Europe is imported, premiums are quoted on the basis of delivery through major Western European ports. As an example, cathode premiums in 2002 ranged from $0.017 to $ 0.024 per pound of copper. Freight from the port to the consuming location is absorbed by the customer. CLC would set its premium based on competitive conditions and freight costs.

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Off specification cathodes (off-spec) are typical for the industry. CLC estimates that 5 percent of their product will be off-spec (either bent or underweight cathodes or containing less than 99.5 percent copper). Off-spec cathodes can carry deductions ranging from $0.01 to $0.03 per pound of copper.

22.5 Contracts

CLC has not established firm arrangement for either mining or marketing.

CLC anticipates the use of contract miners for the stripping and extraction of ore. Budgetary-level cost estimates have been provided to CLC from three vendors, two of which were reviewed by PAH. While budgetary-level estimates have been provided to CLC, the contracts cannot be considered binding. The contract mining estimates are quite similar to the rates that are paid by other mining operations in Spain that have worked with PAH.

Concentrating costs are felt to be adequate by PAH, further discussion of operating costs are presented later within this section. As a direct producer of cathode copper, there will be no smelting and refining charges other than those incurred by CLC in the normal course of operations.

Transportation costs will be in metal cathode form. The cost of transportation and selling costs to the consumer is expected to be roughly equivalent to the cathode premium, thus offsetting each other. Explicit treatment of these costs has been included within the detailed cash flows, but the actual transportation costs can only be estimated until either contracts are signed or specific consumers can be identified. The present assumptions employed at the cash flows is an adequate treatment for this stage of analysis.

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22.6 Environmental Considerations

Environmental issues for Las Cruces are typical for the mining industry regarding noise, dust and on-going environmental management. Of these, the groundwater management and the reclamation process will result in material impacts on cash flow. Costs for both have been adequately factored into the economic analysis.

22.6.1 Permitting

Of importance in the permitting process for Las Cruces is the issuance of a positive Declaration Environmental Impact (DEI). This issuance gives the CLC the right to apply for a mining concession. A positive DEI for Las Cruces was issued by the Provincial Delegation of the Regional Ministry for the Environment on May 9, 2002. CLC was granted the Mining Concession on August 6, 2003.

As of April 2004, CLC has obtained the following key permits:

• DEI;

• Mining Concession;

• Dewatering - Reinjection System; and

• Effluent Discharge of DPH and DPMT.

Table 22-5 presents a summary of the status of the 49 permits CLC has identified for operation of the project. As of April 2004, nine permits were being processed for approval, and 15 permits were pending application to the regulatory authority. Of the pending permits, three require ownership or authorization from the landowner for submittal. CLC expects to receive approval for the permits being processed or pending within one year. There are a number of permits outstanding in order for the project to proceed. To avoid potential delays, PAH recommends that CLC maintain the current emphasis on permitting, including land acquisition and supporting detailed engineering.

Permits for diversion of surface waters will first require registering and platting of affected watercourses, including a policing zone (100 meters beyond the design flood stage). This is a relatively new Spanish regulation that postdates most property boundaries. Setting of the stream boundaries affects the total land area owned by the individual property owners. As such, approval by the affected property owners is required prior to acceptance of the final boundaries. CLC plans to wait until the lands have been purchased before they submit the permit application. This approach should mitigate any potential delays associated with this permitting.

Authorization for the withdrawal and reinjection of groundwater for pit dewatering was issued by the water use authority in November 2003. The proposed reinjection system was reviewed by the Spanish Geological Society. Previously, no similar system had been permitted in Spain. Pumping

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TABLE 22-5

MK Resources Company

Las Cruces Project Technical Report

Summary of Permit Status

Category		Permit	Regulatory Authority	Application Date	Status / Auth Date	Notification	PAH Comment
	1	Declaration of Environmental Impact	DPCMA	3/23/2003	5/9/2002	5/16/2002	Approved EIA serves as basis for project environmental status and underlies other Central government and regional agency approvals.
	2	Mining Concession	DPCEDT	3/23/2003	8/6/2003	8/6/2003	Granted by Central government with 30 year term and two possible 30-year extensions.
	3	Mining Project Optimization	DPCEDT		1/29/2004	2/6/2004
Mining— Hydrometallurgy— Project	4	Integrated Environmental Authorization	DPCMA	22/12/03	BP		Requirement for preparation of this document is a result of EU Directive 96/61 being incorporated into Law 16/2002. This will combine several permits/approvals including the DEI, DPMT, DPH and hazardous waste management into an integrated control of pollution approach.
	5	Special Plan	DPCOPT	4/15/2004	PA		Submittal to Regional Ministry of Public Works and Transport (DPCOPT)
Dewatering— Reinjection System	6	Dewatering—Reinjection Authorization	CHG	7/15/2002	10/30/2003	11/10/2003	Addresses pit dewatering and re-injection. Has four conditions as part of approval including establishing contingency plan prior to startup, analysis of water in each dewatering sector prior to startup, daily monitoring requirements, and monitoring and control plan
	7	Crossings over Regional Highways	DPCMA	2/11/2003	2/19/2003	3/17/2003
	8	Crossings over National Highways	Devpt. Min.	2/11/2003	5/10/2003
	9	Crossings over Livestock Trails	DPCMA	5/10/2001	BP		Process has been suspended until just before construction starts.
	10	Water Concessions, Consumptive Use	CHG	12/18/2000	BP		Concessions required from Ministry of Environment for three sources of water including Sewage Treatment Plant (STP) effluent, groundwater flowing to pit, groundwater for domestic use. CHG has issued report indicating this is compatible with the Guadalquivir Hydrologic Plan (PHG) and Regional Ministry of Health issued a favorable report on suitability of using STP effluent for industrial purposes. Final approval is pending change in agency personnel following recent change in government.
	11	Utilization of rainwater	CHG	12/19/2001	BP		Authorization from Ministry of Environment for collection and use of surface water runoff.
	12	Occupation DPMT by supply/Discharge Pipes	Coasts	5/27/2002	6/30/2003	7/8/2003
Water Concessions, Consumptive Use (Incl supply-discharge system and supply pond) and related permits	13	Occupation DPMT easement zone by supply/discharge pipes	DPCMA	5/14/2002	9/20/2002	10/3/2003
	14	Crossing & occupation of regional highways by supply/discharge pipes	DPCOPT	2/11/2003	2/19/2003	3/17/2003
	15	Crossing of national highways by supply- discharge pipes	Devpt. Min.	2/11/2003	5/26/2003	6/10/2003
	16	Crossing under railways by supply-discharge pipes	Renfe	4/29/2003	8/26/2003	10/29/2003
	17	Technical Report for supply/discharge pipes to cross Viar Canal land	CHG		6/2/2003		Technical report to support approval of authorization to cross Viar Canal land.
	18	Authorization for supply/discharge pipes to cross Viar Canal land	CHG		PA		Approval is pending final approval of construction details such as width of pipeline corridor and method of crossing canal.
	19	Crossing of Livestock Trails	CMA	5/10/2001	BP
	20	Sundry Crossings (power lines, underground telephone line, irrigation channels, Emasesa pipeline)	Various		PA
	21	Authorization for discharge into DPMT	CAN	5/14/2002	10/21/2003	10/28/2003
	22	Occupation of DPMT easement zone by discharge works	CMA	5/14/2002	10/21/2003	10/28/2003
Discharges and Related Permits	23	Occupation of DPMT by discharge works	Coasts	5/14/2002	9/23/2003	10/9/2003
	24	Authorization of discharge into DPH	CHG	4/26/2001	3/11/2003	3/20/2003
	25	Technical Feasibility Report	CHG	5/14/2001	11/6/2002
Demarcation of Public Water Domain (DHP) & Policing Zone (ZP)	26	Demarcation of DPH & ZP	CHG		PA		Demarcation of the Public Domain Waters (DPH) and the Policing Zone (ZP), which is a 100 meter zone on each side of the DPH, is required for final approval of discharge of unaffected runoff from the project to the Garnacha, Molinos and Almendrillos drainages. This is on hold pending land purchase negotiations to define the limit of land that CLC will own.
Demarcation of Public Inland- Maritime Domain (DPMT) & Easement Zone	27	Demarcation of DPMT of Rivera de Huelva	Coasts	7/4/2001	12/11/2002	2/4/2003
	28	DEI on power line	MMA	4/24/2002	10/9/2003	10/27/2003
	29	Authorization of Preliminary Project	Econ.Min.	4/25/2001	BP		Authorization for high voltage power line and substation
Mining—Hydrometallurgy—Project	30	Authorization of Construction Project	Econ.Min.		PA		Authorization to construct power line and substations including right of way.
	31	Crossing over DPH	CHG	2/21/2003	4/28/2003	6/10/2003
	32	Crossings over Livestock Trails	DPCMA		PA
	33	Crossings of SE-520 Regional Highway	CPCOPT		PA
	34	Environmental Report on Diversion	DPCMA	4/7/2003	7/29/2003	8/11/003
Diversion of Medium Voltage Power Line and Related Permits	35	Authorization of Preliminary Project	DPCEDT	4/7/2003	8/28/2003	9/1/2003
	36	Authorization of Construction Project (Year 7+)	DPCEDT		PA
	37	Crossings over DPH	CHG	4/7/2003	7/11/2003	11/4/2003
	38	Technical Feasibility Report	CHG	7/1/2001	5/28/2003	6/6/2003
Works & Installations in DPH & ZP	39	Authorization of works and installations in DPH and ZP	CHG		PA		Subject to demarcation of DPH and DPH-ZP as addressed under approval No. 26.
Stream Diversions	40	Technical Feasibility Report	CHG	5/25/2001	12/17/2002
	41	Authorization of Stream Diversions	CHG		PAL
	42	Disencumbrance/Encumbrance	Revenue Min.		PAL
SE-520 Underpass	43	Authorization of Project	DPCOPT		PA
SE-520 Roundabout	44	Authorization of Project	DPCOPT		PA
Livestock Trails	45	Demarcation	DPCMA	4/2/2001	BP
	46	Route Modification	DPCMA	5/4/2001	BP
	47	Authorization of Occupation by supply-discharge pipes and dewatering-reinjection system	DPCMA	5/10/2001	BP
Rural Tracks	48	Diversion / Abolition	Councils		PAL
Municipal Licenses	49	Works & Activity Licenses from Gerena, Guillena, Salteras, La Algaba, La Rinconda & Seville Town Councils	Councils		PA
KEY	BP PA PAL	Being Processed Application Pending Application Pending, Subject to Land Acquisition					Environmental-permit.x/s

and re-injection systems for aquifer management have been successfully used on mining and petroleum projects and is an accepted industry practice. For example, in the Carlin Trend in the state of Nevada, USA, re-injection of mine dewatering water is done at Jerritt Canyon mine, Cortez Joint Venture, Newmont’s Lone Tree mine, and Barrick Goldstrike mine re-injection is done by both injection wells and infiltration basins and is regulated by both the State of Nevada and the U.S. EPA under the Underground Injection Control Permit program.

While much of the project site is under single ownership, there are approximately 90 landowners involved with acquisition of the entire property (including water and power transmission corridors). CLC has been in contact with several owners, and has indicated that they would offer a price greater than the local fair market real estate value to acquire the properties. Under Spanish law, the mining concession allows the mine owner to acquire land through expropriation if no other avenue is available.

22.6.2 Restoration

The environmental restoration plan for Las Cruces is based on the plan submitted as part of the Project Environmental Impact Study (EIS) dated March 2001. The current plan incorporates the results of this study plus modifications for project optimizations and incorporation of all permit conditions specified in the DEI. Two principal modifications to the original restoration plan are:

1. The artificial lake to be created after the closure of the open pit has been replaced by a landfill for inert waste material, mainly from construction and demolition works. This alternative was analyzed and developed in a report prepared by INTECSA-INARSA entitled “Study of Post-Closure Inert Material Dry Seal, Las Cruces Project” dated December 2001. The new restoration plan for the pit will include partial backfill with marl, which will create a residual void completely lined with marl as a dry seal.

2. The surface areas to be restored have been modified in line with the new mining plan. Subsequent to the issuance of the November 2003 Feasibility Study, an optimized mining plan has been prepared by Independent Mining Consultants (IMC) of Tucson, Arizona. The surface areas in IMC’s mining plan differ slightly from those in the November 2003 Feasibility Study; in PAH’s opinion, it will not materially impact the restoration plan or costs.

PAH has reviewed the other required modifications stated in the DEI and found them to be achievable within the revised restoration plan. However, at this time, the Feasibility Study does not address in any detail the pit backfilling with inert materials, nor has CLC been required to submit a formal plan or restoration costs to the environmental authorities. It is estimated that under this scenario the pit backfilling would occur over more than 30 years. CLC recognizes that a specific Performance Project must be presented for authorization regarding the plan for backfilling with inert materials.

Pincock, Allen & Holt 9033.05 July 8, 2004 22.24

At this time, CLC has not made a decision as to whether the management of the backfilling operation (in effect, a controlled land fill) would be sold to an existing European landfill company, or under CLC management. PAH considers two options viable for the pit backfill: (1) with inert materials depending on availability, or (2) if sufficient inert materials are not available, as a worst case, backfilling with low-ARD mine waste materials (marl and sandstone) from the waste dumps.

The Las Cruces project will utilize on-going environmental restoration comprised of:

• Temporary planting for erosion protection;

• Final revegetation during various project phases; and

• Removal of infrastructure at project closure.

Temporary planting will be implemented for erosion prevention and landscaping primarily to dump surfaces and stockpiled materials that will be covered subsequently with other materials or moved elsewhere. Permanent restoration will be applied to the final reclaimed surfaces in dumps, tailings storage, mine waste storage, or areas that were used for temporary stockpiles.

CLC has identified nine main areas for environmental restoration. Restoration will be initiated at the outset of the two-year construction period, continue throughout the 15-year production period, and culminate during two years after production ceases. Approximately 29 percent of the permanent restoration will be completed during construction, and 32 percent during the two years after production ceases. The total restoration budget is €10.74 million. It should be noted that the Regional Ministry of Employment and Technological Development (CEDT) will require placement of a €13.75 million restoration bond, and an additional €5 million labor bond at start-up of the project.

Visual berms and waste piles will be graded at approximately 4 horizontal to 1 vertical slopes in all areas visible to the public. These slopes approximately match the rolling terrain in the project area. Slopes will be graded and seeded. The tailings disposal facility will be capped and seeded as placement of the tailings progresses across the facility. Similarly, the pit will be partially backfilled behind the mining operation as it proceeds across the proposed footprint of the pit. Backfilling of the pit will consume a portion of the marl produced over the life of the mine, limiting the size of surface waste dumps that will require reclamation.

Typically, the Spanish government requires posting of a bond based on the total disturbed area of the mine. Since CLC will be reclaiming the site throughout the life of the mine, it is anticipated that the largest bonding requirements will be early in the project, and will continuously decrease as mining and reclamation progresses.

Table 22-6 shows CLC’s annual budget for environmental restoration. Almost 50 percent of the environmental restoration will be completed during the first five years of production. Budgets in the initial years (Years –2 and -1) include costs for development and seeding of visual berms, revegetation, and habitat development in the diverted stream channels. Operational years include costs for ongoing restoration of waste dumps and tailings disposal areas that have reached final

Pincock, Allen & Holt 9033.05 July 8, 2004 22.25

proposed grades. Years 16-17 for post mining activities include final capping, grading, revegetation and other activities associated with reclamation of the waste dumps, livestock trails, power lines, tailings facility, mill site, water diversion and storage ponds, roads and other facilities.

The environment reclamation costs presented on a unit basis appear to be conservative. As an example, the unit cost associated with capping waste rock dumps and tailings facility with a marl cover is €15,400/ha. Restoration costs vary from €1,650/ha for temporary pasture on topsoil stockpiles to €25,900/ha for shrub – open Mediterranean forests.

TABLE 22-6

MK Resources Company

Las Cruces Project Technical Report

Annual Restoration Areas and Budgets

Project Year	Restoration Area (ha/year)	Percent Permanently Revegetated		Cumulative % Permanently Revegetated		Annual Budget Euros (000’s)
-2	78	11.6	%	11.6	%	1,316
-1	114	17.0	%	28.7	%	1,814
1	82	12.2	%	40.9	%	1,417
2	33	4.9	%	45.8	%	835
3	9	1.3	%	47.2	%	362
4	9	1.3	%	48.5	%	362
5	9	1.3	%	49.9	%	197
6	9	1.3	%	51.2	%	197
7	9	1.3	%	52.5	%	163
8	27	4.0	%	56.6	%	495
9	27	4.0	%	60.6	%	495
10	20	3.0	%	63.6	%	389
11	20	3.0	%	66.6	%	389
12	3	0.4	%	67.0	%	43
13	3	0.4	%	67.5	%	43
14	3	0.4	%	67.9	%	43
15	3	0.4	%	68.4	%	43
16	62	9.3	%	77.6	%	477
17	150	22.4	%	100.0	%	1,661
Totals	670	100.0	%	100.0	%	10,740

Note: Annual budget includes 10% contingency.

22.6.3 Groundwater Management

The groundwater pumping system is needed to depress the phreatic surface to elevations deeper than the mining zone. In the initial two years, this level must be depressed to an elevation below the sandstone aquifer (approximately 160 meters) to allow commencement of mining. The modeling reported in the Feasibility Study is based on drawing the groundwater down to the maximum mining depth as soon as possible. The model was used to evaluate the actual pumping required to depress the groundwater level at the rate that more closely approximates the proposed mining schedule.

Pincock, Allen & Holt 9033.05 July 8, 2004 22.26

CLC estimates that approximately 2.4 million cubic meters of water will need to be pumped in the initial 2 years to account for the water volume stored in the portion of the sandstone aquifer removed in preparation for mining. The model assumes that the water will be extracted and then re-injected, as the local aquifer is already depleted from local overuse.

Once operational, CLC will withdraw groundwater from the sandstone and Paleozoic layers at a rate of 150 l/s to 220 l/s, which will be reinjected. The proposed reinjection system appears feasible to minimize impacts to the Niblea-Posadas aquifer for most groundwater users in the area. Because the pumping and reinjection is a closed system, no adverse impacts to water quality within the aquifer are anticipated with slight rises in the aquifer levels occurring in all areas except for the area around the Render Sur. The Render Sur, a commercial rendering facility located relatively close to the proposed mill site, will be most impacted as groundwater levels will likely be depressed to below the plant’s current well depth. To mitigate this impact, water will be supplied to the Render Sur facility by the Las Cruces project.

22.7 Taxes and Royalties

The project is subject to a royalty of 1.5 percent of the copper prices greater than or equal to $ 0.80 per pound, payable to Rio Tinto as part of the original purchase agreement. At prices below $ 0.80 per pound, no royalty would be due. The Rio Tinto royalty is appropriately considered within the cash flow.

CLC will be subject to the standard suite of taxes payable under Spanish Law. Cash Flow estimates assume that income will be taxable at an average rate of 35 percent after deductions for depreciation and depletion.

Depreciation is calculated on a straight-line 10-year basis for initial capital, sustaining capital, project acquisition and financing costs. Losses can be carried forward under Spanish tax law, and income taxes for operating years . Depletion can be claimed as the lower of either 15 percent of the net revenue or 30 percent of the taxable earnings, whichever is lower.

Spain collects a Value Added Tax of 16 percent on all purchased goods, including those spent by the mine. VAT collected on construction and startup is refundable and is recaptured in the following year in the cash flow.

22.8 Capital and Operating Cost Estimates

The project capital cost estimates included in the November 2003 Feasibility Study report and supplemental information from DMT-MC received in April 2004 appear reasonable and complete. The estimates are based on quotations obtained in October 2003, from previous studies and recent adjustments by other consultants.

Pincock, Allen & Holt 9033.05 July 8, 2004 22.27

The capital estimates are based on an engineering, procurement, and construction management (EPCM) method of project execution. Installation costs include the cost of contractor’s labor, indirect costs for temporary facilities and services, construction equipment, supervision, overhead, and profit.

Contingencies have been included in the capital estimate to cover estimate errors, design improvements, pricing variations, schedule delays, equipment and material delays, and subcontractor’s claims.

All costs are expressed in fourth quarter 2003 euros. Price adjustments of previous quotations were made by using the CPI-Index for contracts with the public administration in Spain. The capital cost estimate excludes escalation, inflation, financial fees, interest on capitalized costs, income taxes, value added taxes (VAT), and loan interest. Costs prior to November 2003 are considered sunk costs and the pre-production period (Years –3 through –1) costs will be capitalized.

The total project capital cost estimate of 292 million euros includes initial capital of 280 million euros, sustaining capital of 23 million euros and bond, working capital recovery and closure recovery credits of 11million euros. The total project capital cost estimate is summarized by area in Table 22-5.

Project initial capital costs began in November 2003 and will continue through the 2½-year pre-production period. Initial costs also include startup and working capital expenditures during the first part of the first year of operation. Project sustaining capital costs cover the period after startup and continue through to project closure. Closure activities commence after the planned 15 years of operation and will last for six years. EPCM and contingency costs, each at about 22 million euros, appear reasonable and are included in the costs presented in Table 22-7. In PAH’s opinion, one area CMML should consider revising is the working capital projection. The method for estimating working capital is appropriate for the first production year at about 6.8 million euros (12 weeks of cash flow or 8 percent of revenue). Working capital is assumed to remain constant until the end of mine life when 90 percent is recovered. Copper production and revenue increase 40 percent the second year, which will likely increase working capital a similar amount. Total impact of this recommended working capital change is an increase of about 2.7 million euros the second year, recovering the same amount in year 15.

The project expects to receive subsidies of up to 53.7 million euros subject to approvals by the European Union, Spain and the province of Andalusia.

Pincock, Allen & Holt 9033.05 July 8, 2004 22.28

TABLE 1-3 / 22-7

MK Resources Company

Las Cruces Project Technical Report

Project Capital Cost Estimate

Area	Initial Capital (euros in 000’s)	Sustaining Capital (euros in 000’s)	Total Capital (euros in 000’s)
Geotechnical Management	550	0	550
Mining	37,527	288	37,815
Surface Water Management	2,302	4,315	6,617
Ground Water Management	7,449	4,883	12,332
Process Plant	149,991	4,929	154,921
Tailings and Mine Rock Management	6,376	7,856	14,232
Water Supply & Storage, Effluent Discharge	4,974	0	4,974
Electrical Power Supply	2,966	50	3,016
Infrastructure	2,116	155	2,271
Environmental Evaluation & Monitoring	2,441	0	2,441
Environmental Restoration Plan	3,130	0	3,130
Permits	2,428	0	2,428
Land	20,832	0	20,832
Owner’s Costs	36,908	143	37,051
Total Capital Cost	279,992	22,619	302,611
Bond, Working Capital & Closure Recoveries	0	-11,105	-11,105
Total	279,992	11,514	291,506

22.8.1 Operating Costs

Project operating costs presented in the Feasibility Study are summarized in Table 22-8. PAH has grouped these figures in three cost areas and by principal cost categories. The meaning of these terms, as used in this context, are as follows:

Cost areas: costs of physically- or operationally-defined areas

Cost categories: costs that are common to many areas, such as labor, power, etc.

The operating costs shown are for Operating Year 5, which is a fairly typical production year throughout the life of the project. The costs do not include IVA taxes.

Approximate annual costs for each of the principal cost areas for Operating Year 5 are as follows:

G & A:	7 million € (vary +/- 1)
Mining:	16 million € (vary 11 to 17, average 15)
Ore processing:	25 million € (vary +/- 1)
TOTAL:	48 million €

Pincock, Allen & Holt 9033.05 July 8, 2004 22.29

TABLE 22-8 Cash Flow Summary (11x17)

TABLE 22-8 MK Resources Company Las Cruces Project Technical Report Cash Flow Summary				Year ending April of:
7/22/2004 16:23	Unit Year	Total		2004      -3		2005      -2		2006      -1		2007       1		2008       2		2009       3		2010       4		2011       5
Mine Production
Ore mined (dry)	000's Tonnes	15,972				—		20		598		1,043		1,065		758		1,105
Ore grade (copper)	% Cu	6.6%				0.0%		3.3%		8.8%		7.0%		6.8%		9.5%		6.6%
Waste mined (dry)	000's Tonnes	188,298				13,776		27,575		17,806		13,025		12,538		12,747		11,255
Total material mined	000's Tonnes	204,270				13,776		27,595		18,404		14,068		13,603		13,505		12,360
Plant Production
Ore to plant	000's Tonnes	15,972								618		1,043		1,065		758		1,105
Ore grade	% Cu	6.6%								8.6%		7.0%		6.8%		9.5%		6.6%
Contained copper	000's Tonnes	1,056.7								53.1		72.7		72.6		71.8		72.4
Copper recovery	000's Tonnes	91.4%								89.1%		90.7%		91.0%		91.9%		91.2%
Cathode copper produced	000's Tonnes	966.1								47.4		65.9		66.1		66.0		66.1
Cathode copper produced	000's Pounds	2,129,783								104,398		145,360		145,622		145,399		145,626
Revenue
Copper price	€ per pound									0.950		0.950		0.950		0.950		0.950
Cathode premium	€ per pound									0.015		0.019		0.019		0.019		0.019
Copper realization	€ per pound									0.965		0.969		0.969		0.969		0.969
Total revenue	000's €	2,063,258								100,741		140,848		141,102		140,885		141,106
Sales cost
Marketing, freight, & royalty	000's €	(76,576	)							(3,748	)	(5,227	)	(5,236	)	(5,228	)	(5,236	)
Operating Costs
Mine	000's €	(251,614	)							(17,341	)	(14,878	)	(13,541	)	(18,832	)	(16,209	)
Process plant	000's €	(362,107	)							(18,697	)	(23,480	)	(24,319	)	(22,274	)	(24,680	)
Ancillary (G&A)	000's €	(129,425	)							(8,940	)	(8,443	)	(8,163	)	(7,523	)	(7,581	)
Total operating cost	000's €	(743,146	)							(44,978	)	(46,801	)	(46,023	)	(48,629	)	(48,470	)
Operating income (before income tax and depreciation)	000's €	1,243,536								52,015		88,820		89,843		87,029		87,399
Adjustments to Income
Depreciation	000's €	(333,736	)							(31,724	)	(32,323	)	(32,436	)	(32,528	)	(32,668	)
Interest from environmental fund	000's €	18,995						416		416		564		650		805		980
Taxes	000's €	(305,178	)					—		—		—		—		—		(21,581	)
Net Income after tax	000's €	623,617						416		20,707		57,061		58,057		55,306		34,131
Noncash	000's €	333,736								31,724		32,323		32,436		32,528		32,668
Operating Cash Flow	000's €	957,352						416		52,431		89,384		90,493		87,834		66,799
Capital Cost
Mine	000's €	(36,514	)	—		(13,980	)	(22,259	)	—		(23	)	—		—		(103	)
Process Plant	000's €	(141,859	)	—		(48,619	)	(86,893	)	(1,682	)	(40	)	(145	)	(200	)	(190	)
Ancillary	000's €	(93,992	)	(5,931	)	(47,096	)	(26,132	)	(4,902	)	(671	)	(911	)	(663	)	(1,023	)
Contingency	000's €	(21,829	)	(505	)	(8,113	)	(11,861	)	(426	)	(53	)	(69	)	(57	)	(87	)
Subtotal capital	000's €	(294,194	)	(6,436	)	(117,808	)	(147,145	)	(7,010	)	(787	)	(1,125	)	(920	)	(1,403	)
Subsidies	000's €	53,748		—		—		24,161		29,587		—		—		—		—
Capital and VAT Taxes	000's €	(5,421	)	(248	)	(10,568	)	(4,770	)	12,876		—		—		—		—
Environmental fund payment/ refund	000's €	—		—		(13,860	)	—		(4,950	)	(2,860	)	(5,170	)	(5,830	)	(1,870	)
Closure recoveries	000's €	3,360		—		—		—		—		—		—		—		—
Working capital	000's €	(678	)	—		—		—		(6,782	)	—		—		—		—
Total net capital cost	000's €	(243,186	)	(6,685	)	(142,236	)	(127,754	)	23,721		(3,647	)	(6,295	)	(6,750	)	(3,273	)
Net Cash Flow	000's €	714,167		(6,685	)	(142,236	)	(127,338	)	76,151		85,737		84,198		81,084		63,526
Cumulative Net Cash Flow	000's €			(6,685	)	(148,921	)	(276,259	)	(200,107	)	(114,370	)	(30,172	)	50,912		114,438
Net present value (as of May 1, 2004)
@ 5%	000's €	389,947		Internal rate of return =						22.3%
@ 10%	000's €	203,927		Payback Period						3.4
@ 15%	000's €	92,610
@ 20%	000's €	23,269

Pincock, Allen & Holt

9033.05 July 8, 2004

TABLE 22-8 Cash Flow Summary (11x17)

TABLE 22-8 MK Resources Company Las Cruces Project Technical Report Cash Flow Summary	Year ending April of:
7/22/2004 16:23	2012       6		2013       7		2014       8		2015       9		2016     10		2017     11		2018     12		2019     13		2020     14		2021     15
Mine Production
Ore mined (dry)	1,075		1,382		1,215		1,141		1,872		990		1,100		1,016		818		774
Ore grade (copper)	6.7%		7.5%		6.4%		5.5%		7.5%		6.5%		4.7%		6.1%		6.5%		3.2%
Waste mined (dry)	10,848		12,082		10,943		8,709		7,888		8,715		7,309		7,050		5,591		441
Total material mined	11,923		13,464		12,158		9,850		9,760		9,705		8,409		8,066		6,409		1,215
Plant Production
Ore to plant	1,075		953		1,138		1,277		933		1,094		1,296		1,168		1,138		1,311
Ore grade	6.7%		7.5%		6.3%		5.7%		7.7%		6.6%		5.5%		6.1%		6.4%		5.0%
Contained copper	71.8		71.3		72.2		72.7		72.0		72.3		71.8		71.7		72.9		65.5
Copper recovery	92.0%		92.7%		91.5%		90.9%		91.7%		91.3%		91.9%		92.2%		90.7%		91.8%
Cathode copper produced	66.1		66.1		66.0		66.1		66.0		66.0		66.1		66.1		66.1		60.1
Cathode copper produced	145,788		145,697		145,495		145,636		145,600		145,587		145,621		145,663		145,707		132,582
Revenue
Copper price	0.950		0.950		0.950		0.950		0.950		0.950		0.950		0.950		0.950		0.950
Cathode premium	0.019		0.019		0.019		0.019		0.019		0.019		0.019		0.019		0.019		0.019
Copper realization	0.969		0.969		0.969		0.969		0.969		0.969		0.969		0.969		0.969		0.969
Total revenue	141,263		141,175		140,979		141,116		141,080		141,068		141,101		141,142		141,185		128,467
Sales cost
Marketing, freight, & royalty	(5,242	)	(5,239	)	(5,232	)	(5,237	)	(5,235	)	(5,235	)	(5,236	)	(5,238	)	(5,239	)	(4,767	)
Operating Costs
Mine	(13,093	)	(19,433	)	(14,587	)	(14,850	)	(13,923	)	(17,609	)	(12,871	)	(10,953	)	(15,196	)	(6,709	)
Process plant	(24,478	)	(23,633	)	(24,919	)	(25,887	)	(23,499	)	(24,617	)	(25,998	)	(25,179	)	(24,927	)	(25,520	)
Ancillary (G&A)	(7,608	)	(7,549	)	(7,877	)	(7,969	)	(7,641	)	(7,757	)	(7,517	)	(7,431	)	(7,422	)	(7,411	)
Total operating cost	(45,179	)	(50,615	)	(47,383	)	(48,705	)	(45,063	)	(49,984	)	(46,386	)	(43,563	)	(47,545	)	(39,640	)
Operating income (before income tax and depreciation)	90,842		85,320		88,364		87,174		90,782		85,850		89,479		92,341		88,400		84,059
Adjustments to Income
Depreciation	(32,773	)	(32,983	)	(33,194	)	(33,373	)	(33,452	)	(1,299	)	(1,308	)	(1,265	)	(1,243	)	(1,167	)
Interest from environmental fund	1,036		1,165		1,201		1,274		1,304		1,304		1,313		1,313		1,313		1,313
Taxes	(26,897	)	(25,014	)	(26,102	)	(25,704	)	(26,979	)	(23,372	)	(24,644	)	(25,644	)	(24,263	)	(8,376	)
Net Income after tax	32,207		28,488		30,269		29,370		31,655		62,481		64,839		66,746		64,208		75,829
Noncash	32,773		32,983		33,194		33,373		33,452		1,299		1,308		1,265		1,243		1,167
Operating Cash Flow	64,981		61,471		63,463		62,744		65,106		63,781		66,148		68,010		65,451		76,996
Capital Cost
Mine	—		(23	)	—		—		—		(103	)	—		(23	)	—		—
Process Plant	(220	)	(1,090	)	(1,100	)	(1,060	)	(190	)	(190	)	(190	)	(50	)	—		—
Ancillary	(770	)	(869	)	(899	)	(629	)	(542	)	(573	)	(631	)	(573	)	(655	)	(605	)
Contingency	(64	)	(114	)	(112	)	(103	)	(50	)	(57	)	(56	)	(45	)	(45	)	(41	)
Subtotal capital	(1,054	)	(2,096	)	(2,111	)	(1,792	)	(782	)	(923	)	(877	)	(691	)	(700	)	(646	)
Subsidies	—		—		—		—		—		—		—		—		—		—
Capital and VAT Taxes	—		(2,711	)	—		—		—		—		—		—		—		—
Environmental fund payment/ refund	(4,290	)	(1,210	)	(2,420	)	(990	)	—		(330	)	—		—		—		—
Closure recoveries	—		—		—		—		—		—		—		—		—		—
Working capital	—		—		—		—		—		—		—		—		—		4,883
Total net capital cost	(5,344	)	(6,017	)	(4,531	)	(2,782	)	(782	)	(1,253	)	(877	)	(691	)	(700	)	4,237
Net Cash Flow	59,637		55,454		58,932		59,961		64,324		62,528		65,271		67,320		64,751		81,233
Cumulative Net Cash Flow	174,075		229,529		288,461		348,422		412,746		475,274		540,544		607,864		672,615		753,849

					01/00/1900
	Year ending April of:
7/22/2004 16:23	2022     16		2023     17		2024     18		2025     19		2026     20		2027     21
Mine Production
Ore mined (dry)	—		—
Ore grade (copper)	0.0%		0.0%
Waste mined (dry)	—		—
Total material mined	—		—
Plant Production
Ore to plant	—		—
Ore grade	0.0%		0.0%
Contained copper	—		—
Copper recovery	0.0%		0.0%
Cathode copper produced	—		—
Cathode copper produced	—		—
Revenue
Copper price	0.950		0.950		0.950		0.950		0.950		0.950
Cathode premium
Copper realization
Total revenue	—		—
Sales cost
Marketing, freight, & royalty
Operating Costs
Mine	(17,088	)	(14,501	)	0		0		0		0
Process plant	0		0		0		0		0		0
Ancillary (G&A)	(4,213	)	(4,126	)	(1,208	)	(1,056	)	(1,016	)	(972	)
Total operating cost	(21,301	)	(18,628	)	(1,208	)	(1,056	)	(1,016	)	(972	)
Operating income (before income tax and depreciation)	(21,301	)	(18,628	)	(1,208	)	(1,056	)	(1,016	)	(972	)
Adjustments to Income
Depreciation	—		—		—		—		—		—
Interest from environmental fund	1,313		1,313		0		0		0		0
Taxes	8,090		—		(1,166	)	(8,712	)	(8,474	)	(36,338	)
Net Income after tax	(11,897	)	(17,314	)	(2,373	)	(9,769	)	(9,491	)	(37,310	)
Noncash	—		—		—		—		—		—
Operating Cash Flow	(11,897	)	(17,314	)	(2,373	)	(9,769	)	(9,491	)	(37,310	)
Capital Cost
Mine	—		—		—		—		—		—
Process Plant	—		—		—		—		—		—
Ancillary	188		(105	)	—		—		—		—
Contingency	40		(11	)	—		—		—		—
Subtotal capital	228		(116	)	—		—		—		—
Subsidies	—		—		—		—		—		—
Capital and VAT Taxes	—		—		—		—		—		—
Environmental fund payment/ refund	—		—		43,780		—		—		—
Closure recoveries	—		—		3,360		—		—		—
Working capital	—		—		1,221		—		—		—
Total net capital cost	228		(116	)	48,361		—		—		—
Net Cash Flow	(11,669	)	(17,430	)	45,987		(9,769	)	(9,491	)	(37,310	)
Cumulative Net Cash Flow	742,179		724,749		770,736		760,968		751,477		714,167

The quantity of copper produced by the project is fixed at 66,000 tonnes (145,000 pounds) per year, a figure that is set by the capacity of the electrowinning section of the ore processing plant. There will, however, be considerable annual variation in both tonnes of material mined and tonnes of ore processed. While large variations in tonnages of ore mined are normal for most mines, large variations in ore processed are not. This variation in the plant ore processing rate is because it is governed by the copper production rate rather than by the milling rate and, as the ore grade changes, the quantity of ore that can be processed changes.

Approximate unit operating costs (in euros per tonne milled) for Operating Year 5 are as follows:

G & A:	6 €/t
Mining:	14 €/t (90 € cts/t material mined)
Ore processing:	23 €/t
TOTAL:	43 €/t

Unit operating costs also in euro cents per pound of copper produced are as follows:

G & A:	5 € cts/lb.
Mining :	11 € cts/lb
Ore Processing:	17 € cts/lb
TOTAL:	33 € cts/lb

The unit cost figures per tonne are considerably higher than normal for most mining operations; however, because cathode copper will be produced on site, and no smelting will be required, there is a benefit to the overall operating cost. This higher cost is expected given the proximity to a major urban area, the high mine stripping ratio, the complexity of the plant, and the high ore grade. Fortuitously, the high ore grade offsets the high operating costs, resulting in very reasonable and competitive unit costs per pound of copper produced.

PAH believes that the cost estimates have been carefully and thoroughly prepared and that the values presented fairly represent the probable costs. However, even mines which have been in operation for several years have actual costs vary by as much as 10 percent from budgeted values; accordingly, the operating cost estimates could vary by as much as 20 percent from the stated figures.

Operating costs for each of the three primary cost areas are discussed in the following section followed by a discussion of the principal cost categories.

22.9 Economic Analysis

MK Resources has developed an economic model to an after-tax level. Operating and capital costs are consistent with the summaries provided earlier and are felt to be a reasonable representation of the performance of the project. Assumptions underlying the development of the cash flow are provided in Table 22-9. The cash flow for the project is presented as Table 22-8. Only proven and probable reserves have been included within the cash flow.

Pincock, Allen & Holt 9033.05 July 8, 2004 22.31

TABLE 22-9

MK Resources Company

Las Cruces Project Technical Report

Cash Flow Model Parameters

Parameter	Assumption
Copper Sales Price	0.95 dollars per pound of copper
Cathode Premium	44 euros per tonne of cathode copper
Mine Life	15 Years
Total tonnes ore mined (dry basis)	15,972,000
Average mined ore grade	6.6 % copper
Total tonnes waste mined (dry basis)	188,298,000
Overall plant copper recovery	91.40%
Average cathode copper production	66 k-tonnes per year
Average cash operating cost	0.327 euros per pound copper
Initial capital	280.0 million euros
Sustaining capital	22.6 million euros
Total Subsidies	53.7 million euros
Working capital	6.8 million euros in year 1
Effective tax rate	35% of net income

PAH notes that the cash flow includes €53.7 million in subsidies from the government. These subsidies are deducted from the initial capital prior to depreciation calculations.

The copper price of $0.95 per pound used in the cash flow differs from the design criteria of 0.76 euros per pound used in the mine design. PAH believes that given the Las Cruces grade distribution that this increase in copper price does not materially affect the mine design, ultimate pit or material below cutoff grade. Any mine design change resulting from the increased copper price assumption would improve projected economic results.

A copper cathode premium of 44 euros per tonne of copper (or approximately 0.02 euros per pound of copper) has been included for 95 percent of the copper produced in Years 2 through 15. Year 1 assumes that 75 percent of cathode copper would receive the premium. The premium adds, on average, approximately 2.7 million euros of revenue per year, or a total of about 40.0 million euros over the life of the project.

Base Case net cash flow (unleveraged) during the first five production years averages approximately 78 million euros per year. The initial capital expenditures of 280 million euros for development of the project [should be] repaid halfway through the third production year. The projected cash flow for the remaining ten production years will average about 64 million euros per year.

Pincock, Allen & Holt 9033.05 July 8, 2004 22.32

The Base Case net cash flow discounted at 10 percent (NPV) totals 203.5 million euros., Base Case internal rate of return (IRR) is a robust 22 percent. Payback occurs after 3.4 years.

Project Sensitivity

Project sensitivities are similar to other commodity metal mines. Sensitivity to price is the highest, with a 10 percent drop in price (to $0.855 per pound) resulting in an NPV drop of about €61 million. Conversely, an increase of 10 percent in price (to 1.045 per pound) results in a similarly large increase to the NPV. Project economics are roughly half as sensitive to operating and capital costs. Figure 22-5 presents a summary and chart of the sensitivities.

FIGURE 22-5

MK Resources Company

Las Cruces Project Technical Report

Cash Flow Economic Sensitivity

Sensitivity analysis is typically provided on both metal prices and currency conversion rates as they affect the Net Present Value at a range of discount rates. Table 22-10 summarizes the sensitivity of the project NPV in millions of dollars to copper price. The Base Case for the sensitivity analysis is shown in bold at a copper price of $0.95 per pound copper, 10 percent discount rate and 1:1 US$:euro exchange rate.

Pincock, Allen & Holt 9033.05 July 8, 2004 22.33

TABLE 22-10

MK Resources Company

Las Cruces Project Technical Report

Sensitivity of NPV to Copper Prices at a Fixed Exchange Rate of 1.00 euro per US dollar

$ Millions

Discount Rate	Copper Price, ($ / lb)
	0.80	0.90	0.95	1.00	1.10	1.20	1.30
6%	215.8	301.7	344.3	386.6	471.2	555.1	638.9
8%	155.1	229.6	266.5	303.0	376.2	448.3	520.6
10%	106.3	171.3	203.5	235.2	299.0	361.6	424.3
12%	66.8	123.9	152.1	180.0	235.9	290.6	345.4
14%	34.5	85.1	110.0	134.6	183.9	232.1	280.4

The design basis currency of the project is the euro, consequently fluctuations in the value of the euro-versus-dollar exchange rate will have an impact on project performance. Tables 22-11, 22-12, and 22-13 demonstrate the effects. Note that the conversion rate is stated in terms of euros per dollar, rather than dollars per euro and the NPV is expressed in millions of dollars. A rate of 1.20 euros per dollar indicates that the dollar is more valuable than the euro.

TABLE 22-11

MK Resources Company

Las Cruces Project Technical Report

Sensitivity of NPV to Euro Conversion Rates at a Fixed Copper Price of $0.95/lb

$ Millions

Discount Rate	Euro Conversion (euros / $)
	0.80	0.90	1.00	1.10	1.20
6%	223.9	291.1	344.3	387.1	422.7
8%	154.2	216.9	266.5	306.3	339.4
10%	98.3	157.1	203.5	240.7	271.7
12%	53.1	108.4	152.1	187.1	216.3
14%	16.3	68.7	110.0	143.1	170.7

Table 22-12 provides a synopsis of the effects that changes to both currency conversion rates and copper prices have on the NPV at a fixed discount rate of 10 percent. Table 22-13 shows the same effect on IRR.

Pincock, Allen & Holt 9033.05 July 8, 2004 22.34

TABLE 22-12

MK Resources Company

Las Cruces Project Technical Report

Sensitivity of NPV @ 10% to Copper Price and Euro Conversion, $US millions

		Copper Price $ / lb
Euro Conversion (euros / $ )		0.80		0.90	0.95	1.00	1.10	1.20	1.30
	1.20	176.2		239.8	271.7	302.9	365.5	428.3	490.9
	1.10	144.5		208.7	240.7	272.5	335.3	397.9	460.7
	1.00	106.3		171.3	203.5	235.2	299.0	361.6	424.3
	0.90	58.3		124.5	157.1	189.7	253.4	317.2	379.9
	0.80	(2.7	)	64.9	98.3	131.1	196.3	260.2	324.0

TABLE 22-13

MK Resources Company

Las Cruces Project Technical Report

Sensitivity of IRR% to Copper Price and Euro Conversion

		Copper Price $ / lb
Euro Conversion (euros / $ )		0.80		0.90		0.95		1.00		1.10		1.20		1.30
	1.20	23	%	27	%	29	%	30	%	34	%	37	%	41	%
	1.10	20	%	24	%	25	%	27	%	31	%	34	%	37	%
	1.00	17	%	20	%	22	%	24	%	27	%	30	%	33	%
	0.90	13	%	17	%	19	%	20	%	24	%	27	%	29	%
	0.80	10	%	13	%	15	%	17	%	20	%	23	%	25	%

Currency Exchange Rate History

The euro has been in existence as a currency since January 1, 1999, when the various European currency exchange rates were officially established as part of the consolidation of the European Union. The formal roll-out of the euro occurred on January 1, 2000. Trading rates have fluctuated within a range of +/- 20 percent with an overall average of 1 euro per US dollar. The present conversion rate is 0.81 euros per US dollar ($1.23 per euro) as of June 8, 2004. Figure 22-6 provides a chart of the exchange rate history. The MK Resources model assumes a one-to-one long-term conversion rate.

Pincock, Allen & Holt 9033.05 July 8, 2004 22.35

Pincock, Allen & Holt 9033.05 July 8, 2004 22.36

23.0 CERTIFICATE OF QUALIFIED PERSON

As an author of the report entitled “CNI 43-101 Technical Report Las Cruces Copper Project, Southern Spain”, dated July 8, 2004 (the “Technical Report”) and prepared on behalf of MK Resources Company (formerly known as MK Gold Company) (the “Issuer”), I, Richard Addison, P.E., C. Eng., Eur. Ing., do hereby certify that:

1. I am currently an associate Principal Process Engineer of:

Pincock, Allen & Holt

274 Union Blvd., Suite 200

Lakewood, CO 80228

USA

2. My residential address is: 10555 West Jewell Avenue, Apt. #2-104 Lakewood, Colorado 80232.

3. I graduated from the Camborne School of Mines in England as an Honors Associate in 1964 and subsequently obtained a Master of Science degree in metallurgical engineering from the Colorado School of Mines in 1968. I have practiced my profession continuously since 1964.

4. I am a Registered Professional Engineer (#3198) in the state of Nevada, USA, a Charted Engineer in the U.K., and a registered European Engineer in the EEC. I am a member of the American Institute of Mining, Metallurgical, and Petroleum Engineers and a member of The Institute of Materials, Minerals and Mining in the U.K.

5. I have worked as a metallurgical engineer for a total of 37 years since my graduation from university and have been involved in the evaluation and operation of mineral properties for gold, silver, copper, lead, zinc, tin, aluminum, iron, gypsum, limestone, barite, sulfur, pyrite, oil shale, coal, and diamonds in the United States, Canada, Mexico, Dominican Republic, Honduras, Nicaragua, Costa Rica, Panama, Venezuela, Guyana, Peru, Ecuador, Bolivia, Argentina, Chile, Spain, Portugal, Britain, Bulgaria, Indonesia, Papua New Guinea, the Philippines, Japan, Tunisia, Ghana, Zambia, South Africa, Russia, Kyrghyzstan, and Australia.

6. I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

Pincock, Allen & Holt 9033.05 July 8, 2004 23.1

7. I am responsible for the preparation of Section 4.4, Environmental Liabilities, the preparation of the Mineral Processing and Metallurgy Testing (Section 16), the Flow Scheme (part of Section 22) and Section 22.6, Environmental Considerations, of the Technical Report.

8. I have had prior involvement with the property that is the subject of the Technical Report, contributing to the preparation of a Technical Audit in 2004.

9. I am not aware of any material fact or material change with respect to the subject matter of the Technical Report that is not reflected in the Technical Report, the omission to disclose which makes the Technical Report misleading.

10. I am independent of the Issuer in accordance with Section 1.5 of NI 43-101.

11. I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

12. I consent to the filing of the Technical Report with any securities regulatory authority, stock exchange or other regulatory authority and any publication by them, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

Dated in Lakewood, Colorado, this 8^th day of July, 2004.

/s/    Richard Addison

Richard Addison, P.E., C. Eng., Eur. Ing.

Pincock, Allen & Holt 9033.05 July 8, 2004 23.2

CERTIFICATE OF QUALIFIED PERSON

As an author of the report entitled “CNI 43-101 Technical Report Las Cruces Copper Project, Southern Spain”, dated July 8, 2004 (the “Technical Report”) and prepared on behalf of MK Resources Company (formerly known as MK Gold Company) (the “Issuer”), I, Darrel L. Buffington, P.E., do hereby certify that:

1. I am employed as Principal Geotechnical Engineer by:

Pincock, Allen & Holt

274 Union Blvd., Suite 200

Lakewood, CO 80228

USA

2. My residential address is: 6170 Arapahoe Drive, Evergreen, Colorado 80439

3. I graduated from the University of Missouri-Rolla with a B.S. in Geological Engineering in 1978 and obtained a M.S. degree in Civil Engineering from the same university in 1980. I have practiced my profession continuously since 1980.

4. I am a Registered Professional Engineer in the states of Colorado (#23705), Nevada (#07739) and Arizona (#34415) and a member of the Society of Mining, Metallurgy and Exploration (SME), the Canadian Institute of Mining, Metallurgy and Petroleum and the American Society of Civil Engineers.

5. Since 1980, I have been involved in broad-scale civil and geotechnical engineering practice, including substantial experience in environmental review and permitting issues of mining properties and operations. This includes reviewing environmental management systems as part of due diligence evaluation of operating mines, providing technical analysis of mine waste containment facilities, review of regulatory compliance issues, and developing strategies for addressing environmental impacts in the mine planning process.

6. As a result of my education and experience, I am a “Qualified Person” as defined in National Policy 43-102.

7. I am presently Manager of Environmental Services and Process Engineering with the international consulting firm of Pincock, Allen and Holt, and have been so since March of 1995.

Pincock, Allen & Holt 9033.05 July 8, 2004 23.3

8. I am responsible for the preparation of Section 4.4, Environmental Liabilities, and Section 22.6, Environmental Considerations, of the Technical Report.

9. I have not had prior involvement with the property that is the subject of the Technical Report.

10. The sources of all information are noted and referenced in the Technical Report. The information provided by the various parties is to the best of knowledge and experience correct.

11. I am not aware of any material fact or material change with respect to the subject matter of the Technical Report that is not reflected in the Technical Report, the omission to disclose which makes the Technical Report misleading.

12. I am independent of the Issuer in accordance with Section 1.5 of NI 43-101.

13. I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

14. I consent to the filing of the Technical Report with any securities regulatory authority, stock exchange or other regulatory authority and any publication by them, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

Dated in Lakewood, Colorado, this 8^th day of July, 2004.

/s/    Darrel L. Buffington

Darrel L. Buffington, P.E.

Pincock, Allen & Holt 9033.05 July 8, 2004 23.4

CERTIFICATE OF QUALIFIED PERSON

As an author of the report entitled “CNI 43-101 Technical Report Las Cruces Copper Project, Southern Spain”, dated July 8, 2004 (the “Technical Report”) and prepared on behalf of MK Resources Company (formerly known as MK Gold Company) (the “Issuer”), I, Gerald David Crawford, P. Eng., do hereby certify that:

1. I am currently employed as a Principal Mining Engineer by:

Pincock, Allen & Holt

274 Union Blvd., Suite 200

Lakewood, CO 80228

USA

2. My residential address is 11040 W. Park Range Road, Littleton, Colorado 80127.

3. I graduated from New Mexico Institute of Mining and Technology with a Bachelor of Science degree in Mining Engineering in 1977 and subsequently obtained a Master of Business degree from the University of Nevada-Reno in 1985, and I have practiced my profession continuously since 1977.

4. I am a Registered Professional Engineer (#30847) in the state of Colorado, USA, a Registered Professional Engineer (#9918) in the state of Nevada, USA, a Registered Professional Engineer (#47421) in the state of Florida, and a member of the American Institute Of Mining, Metallurgical, and Petroleum Engineers, Inc.

5. I have worked as a mining engineer for a total of 27 years since my graduation from university and have been involved in the evaluation and operation of mineral properties for gold, silver, copper, lead, zinc, palladium, uranium, coal, and industrial minerals in the United States, Canada, Mexico, Costa Rica, Panama, Peru, Chile, Spain, Indonesia, Ghana, Mali, Russia, India and Australia.

6. I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

7. I am responsible for the preparation of the Summary (Section 1), Introduction, Disclaimer, Description and Accessibility (Sections 2-5), reserve (part of Section 17), Conclusions, Recommendations, and Additional Information (Sections 18 through 22) sections of the Technical Report.

Pincock, Allen & Holt 9033.05 July 8, 2004 23.5

8. I have had prior involvement with the property that is the subject of the Technical Report in a previous diligence review in 2002.

9. I have personally visited the Las Cruces site in Andalusia, Spain on January 13 to 15, 2004.

10. I am not aware of any material fact or material change with respect to the subject matter of the Technical Report that is not reflected in the Technical Report, the omission to disclose which makes the Technical Report misleading.

11. I am independent of the Issuer in accordance with Section 1.5 of NI 43-101.

12. I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

13. I consent to the filing of the Technical Report with any securities regulatory authority, stock exchange or other regulatory authority and any publication by them, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

Dated in Lakewood, Colorado, this 8^th day of July, 2004.

/s/  Gerald D. Crawford

Gerald D. Crawford, P.E.

Pincock, Allen & Holt 9033.05 July 8, 2004 23.6

CERTIFICATE OF QUALIFIED PERSON

As an author of the report entitled “CNI 43-101 Technical Report Las Cruces Copper Project, Southern Spain”, dated July 8, 2004 (the “Technical Report”) and prepared on behalf of MK Resources Company (formerly known as MK Gold Company) (the “Issuer”), I, Nelson D. King, Chief Process Engineer, do hereby certify that:

1. I am currently employed as Chief Process Engineer by:

Pincock, Allen & Holt

274 Union Blvd., Suite 200

Lakewood, CO 80228

USA

2. My residential address is: 12849 W. 55^th Place, Arvada, Colorado 80002.

3. I graduated from Colorado School of Mines with a B.S. in Metallurgy Engineering in 1972 and have practised in my profession continuously since 1973. I am an active member of the Society for Mining, Metallurgy and Exploration, Inc. (SME).

4. Since 1973, I have been involved in mineral processing plan design and construction, metallurgy consulting, and mine and plant operations management for open pit and underground mining operations and ore processing facilities. The projects I have been involved in were for the recovery of copper, gold, silver, lead, zinc, platinum, palladium, and molybdenum, and located in the United States, Canada, Mexico, Peru, Chile, Spain, Bulgaria, Russia, Indonesia, Australia, Argentina and Ghana.

5. I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

6. I am responsible for the reviews of the metallurgy testwork, metallurgical assumptions of future operations, future plant capital and operating costs, and production forecasts. I have prepared and/or assisted in the preparation of the mineral processing and metallurgical testing, metal recoverability, and product markets portions of the Technical Report.

7. The sources of all information are noted and referenced in the Technical Report. The information provided by the various parties is correct, to the best of my knowledge and experience.

Pincock, Allen & Holt 9033.05 July 8, 2004 23.7

8. I am not aware of any material fact or material change with respect to the subject matter of the Technical Report that is not reflected in the Technical Report, the omission to disclose which makes the Technical Report misleading.

9. I am independent of the Issuer in accordance with Section 1.5 of NI 43-101.

10. I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

11. I consent to the filing of the Technical Report with any securities regulatory authority, stock exchange or other regulatory authority and any publication by them, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

Dated in Lakewood, Colorado, this 8^th day of July, 2004.

/s/   Nelson D. King

Nelson D. King, Chief Process Engineer

Pincock, Allen & Holt 9033.05 July 8, 2004 23.8

CERTIFICATE OF QUALIFIED PERSON

As an author of the report entitled “CNI 43-101 Technical Report Las Cruces Copper Project, Southern Spain”, dated July 8, 2004 (the “Technical Report”) and prepared on behalf of MK Resources Company (formerly known as MK Gold Company) (the “Issuer”), I, Mark G. Stevens, C.P.G., P.G., do hereby certify that:

1. I am currently employed as a Chief Geologist by:

Pincock, Allen & Holt

274 Union Blvd., Suite 200

Lakewood, CO 80228

USA

2. My residential address is 4229 E. 106^th Place, Thornton, Colorado 80233.

3. I graduated from Colorado State University with a Bachelor of Science degree in geology in 1977 and subsequently obtained a Master of Science degree in geology from the University of Utah 1981, and I have practiced my profession continuously since 1981.

4. I am a Professional Geologist (PG-651) in the state of Wyoming, USA, a Licensed Geologist (PG-477) in the state of Washington, USA, a member of the American Institute Of Professional Geologists (CPG-08388), a member of the American Institute Of Mining, Metallurgical, and Petroleum Engineers, Inc. (SME), and a member of the Society Of Economic Geologists (SEG).

5. I have worked as a geologist for a total of 23 years since my graduation from university and have been involved in mineral exploration and evaluation of mineral properties for gold, silver, copper, lead, zinc, coal, and industrial minerals in the United States, Canada, Mexico, Costa Rica, Panama, Peru, Chile, Spain, Sweden, Portugal, Philippines, Kazakhstan, Russia, India and Australia.

6. I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

7. I am responsible for the preparation of the history (Section 6), geology (Sections 7-9), exploration (Sections 10-15), and resource (part of Section 17) sections of the Technical Report. I have not visited the Las Cruces property.

Pincock, Allen & Holt 9033.05 July 8, 2004 23.9

8. I have not had prior involvement with the property that is the subject of the Technical Report.

9. I am not aware of any material fact or material change with respect to the subject matter of the Technical Report that is not reflected in the Technical Report, the omission to disclose which makes the Technical Report misleading.

10. I am independent of the Issuer in accordance with Section 1.5 of NI 43-101.

11. I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

12. I consent to the filing of the Technical Report with any securities regulatory authority, stock exchange or other regulatory authority and any publication by them, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

Dated in Lakewood, Colorado, this 8^th day of July, 2004.

/s/    Mark G. Stevens

Mark G. Stevens, C.P.G., P.G.

Pincock, Allen & Holt 9033.05 July 8, 2004 23.10