EX-3 126 techrep.htm PAMPA DE PONGO IRON PROJECT PRELIMINARY ECONOMIC ASSESSMENT 43-101 TECHNICAL REPORT DATED SEPTEMBER 30, 2008 Pampa de Pongo Iron Project Preliminary Economic Assessment Technical Report






Pampa de Pongo Iron Project

Preliminary Economic Assessment Technical Report

Caraveli Province, Peru





Report Prepared for

Cardero Resource Corp.








Report Prepared by

[techrep002.jpg]


September 30, 2008




















Pampa de Pongo Iron Project

Preliminary Economic Assessment

Technical Report

Caraveli Province, Peru



Cardero Resource Corp.

Suite 1920, 1188 West Georgia Street

Vancouver, BC V6E 4A2


SRK Consulting (Canada) Inc.

Suite 2200, 1066 West Hastings Street

Vancouver, BC V6E 3X2


Tel: 604.681.4196    Fax: 604.687.5532

E-mail: gdoerksen@srk.com

   Web site: www.srk.com


SRK Project Number 2CC029.000


The effective date of this report is September 30, 2008


This report was written by the following Qualified Persons:


Gordon Doerksen, P.Eng., SRK Consulting (Canada) Inc.

Marek Nowak, P.Eng., SRK Consulting (Canada) Inc.

George Wahl, P.Geo., SRK Consulting (Canada) Inc.

Leonard Holland, Chartered Engineer, Holland & Holland Consultants (UK)




Executive Summary

The Pampa de Pongo iron ore (magnetite) project consists of a block of 8 adjoining mining concessions (Concesión Minera)and 10 mining claims (Petitorio Minero) encompassing an area 19 km long and up to 10 km wide (15,300 hectares). The project is located on the southern coastal plain of Peru, 50 km south of the city of Nazca and 550 km southeast of Lima, in the province of Caraveli, in the governing jurisdiction of Arequipa. The geographic center of the property is located at 74° 50" W longitude, 15° 23" S latitude. The property is at approximately 400 m in elevation and is 20 km from the Pacific Ocean, 10 km from the Pan-American Highway and 38 km for the deep-sea port of San Juan. The property is located in a desert environment with minimal rainfall and very sparse vegetation.

Cardero Resource Corp. completed an option to purchase a 100% undivided interest in the property from Rio Tinto Mining and Exploration, Sucursal Peru (Rio Tinto). The Concesiones Mineras (8,000 ha) are presently in the process of being transferred from Rio Tinto to Cardero Peru, however, as of writing they are still held by Rio Tinto. Rio Tinto anticipates that the concession transfers will be completed imminently. The 10 Petitorios Mineros (7,300 ha) are held 100% by Cardero Peru S.A.C.

The Pampa de Pongo deposit was discovered by Rio Tinto in 1994 as part of a regional Iron Oxide Copper Gold (“IOGC”) exploration program. Rio followed up on the discovery with drilling in 1995 and 1996. The deposit does not outcrop and is buried beneath aeolian sand.

There are no historical mine workings, waste management facilities, tailings ponds or important natural features within the area of the Pampa de Pongo claims block.

Geology

The Pampa de Pongo deposit is located in the Marcona (Fe-Cu) District, named after the Marcona Iron Mine, which has operated continuously since 1953. The oldest rocks exposed are Precambrian in age and part of the Coastal Basal Complex, consisting of gneisses, potassium-rich granites and migmatites. This basement metamorphic complex is overlain by carbonates, pelitic sediments and a variety of intermediate to mafic igneous units of the Lower Paleozoic Marcona Formation, Middle to Upper Jurassic Rio Grande Formation and Bella Union Volcanics. Small remnants of the Upper Jurassic Jahuay Formation (mixed volcanics and sediments) and Lower Cretaceous Jauca Formation (sediments) cap the Jurassic succession.

Major intrusives include the San Nicolas Batholith and the more extensive Lower Cretaceous Coastal Batholith.

The majority of the mineralization at the Marcona mine is hosted by carbonates of the Marcona Formation. The Pampa de Pongo deposit is hosted by the younger Jurassic Juhuay Formation.

Mineralization

At Pampa de Pongo the mineralized bodies are semi-concordant and appear to have been controlled by a number of factors, primarily structure, but also host rock lithologies, primary porosity and secondary porosity related to fracturing or brecciation. Mineralization comprises semi-massive to massive magnetite replacement zones.

There are four main zones of mineralization and potential mineralization:

1.

North Zone – An untested exploration target indicated by a large magnetic anomaly.

2.

Central Zone – The Central zone is the main focus of this PEA. It is a two-layer mineralized body. The lower portion consists of a flat-lying, massive replacement lens up to 370 metres thick, measuring approximately 1,060 metres east-west by 1,000 metres north-south at the widest point. The resource remains open in all directions. Mineralization is consistent, averaging approximately 62% magnetite with multiple intersections greater than 80% magnetite. In the overlying upper portion (not included in the current resource), massive iron mineralization is cut with a minor component of hypabyssal porphyry sills. The lower, high-grade massive replacement at Pampa de Pongo, from which the Inferred Resource has been estimated, based on current data, is almost completely devoid of unmineralized intrusive rocks.

3.

East Zone – The East Zone is indicated by a 3D magnetic anomaly, the dimensions of which, suggest potential for 350-500 million tonnes of magnetite. A single drillhole on the southern edge of the anomaly intersected massive magnetite mineralization. Mineralization consists of 3 high-grade mineralized lenses grading 41% to 44% iron over tens of metres approximately 270 metres from surface. The total thickness of magnetite mineralization intersected is 292 metres.

4.

South Zone – This is a near-surface resource comprised of two separate zones of predominantly massive magnetite mineralization with a thickness of up to 120 metres, a combined length of  1,100 metres north-south and a width of 400 metres east-west at the widest point. This resource has not been included in the mine schedule presented in this report and represents future upside potential.

Metallurgy and Mineral Processing

Representative metallurgical samples were selected by SRK from four diamond drillhole cores from the Central Zone and beneficiation, magnetic concentration and pilot-scale pelletizing tests were conducted. The preliminary metallurgical testwork demonstrated that:

  • Wet magnetic separation would yield 93.4% Fe recovery;
  • Metallurgical quality exceeds industry standards for high quality blast furnace feed;
  • Commercially produced pellets made from Pampa de Pongo ore would be a suitable feedstock for the MIDREX ® Direct Reduction Process; and
  • Deleterious materials in the pellets would be at or below acceptable levels.

The mineral processing design for this study includes crushing – grinding – flotation – wet magnetic separation and pelletization. Economic values of copper and gold are anticipated to be extracted in the flotation stage with 50% recovery assumed for both metals.

Mineral Resource Estimate

The mineral resources were estimated by Ebi Ghayem (P.Geo.) and reviewed by Marek Nowak (P.Eng.) and George Wahl (P.Geo.) who are all Qualified Persons. The effective date of the Mineral Resource Statement is September 30th, 2008. SRK is not aware of any environmental, permitting, legal, title, taxation, socio-political, marketing or other relevant issues that may affect the mineral resource estimate.

The classified inferred mineral resource estimates at 15% Fe cut-off grade are tabulated in Table 1.

Table 1 : SRK Classified Mineral Resources for the Pampa de Pongo Deposit at 15% Fe cut-off

ZONE

Classification

Volume

(Mm3)

Density

(T/m3)

Tonnage

(Mt)

Fe

(%)

Au

(g/t)

Cu

(%)

Central

Inferred

203

3.69

748

41.7

0.059

0.093

South

Inferred

32

3.59

115

39.5

0.130

0.121

Total

Inferred

235

3.67

863

41.3

0.068

0.097

The resources represent all estimated blocks within the modeled zones. The economic cut-off used to generate mineral resources was assumed and based on experience with similar projects. The final cut-off required to produce a saleable product will need to be confirmed by future metallurgical testwork. This economic cut-off was applied to both the Central and South Zones. Although the South Zones contribute a relatively small tonnage, SRK is of the opinion that there are reasonable prospects for additional tonnage in this area which may then make these resources amenable to underground mining methods. Table 2 shows the mineral inventory at various cut-off grades.

Two exploration targets also exist on the property. The East Zone exploration target may potentially contain a conceptual tonnage of 350 to 500 Mt of magnetite mineralization. A single hole was drilled in the East Zone which intersected 292 m of semi-massive and massive magnetite mineralization at the extreme edge if the 3D magnetic anomaly. The second exploration target, the North Zone shows a magnetic anomaly but is untested.

Table 2 : Mineral Inventory at Various Cut-off Grades

ZONE

Cut-Off

Grade

Volume

(Mm3)

Density

(T/m3)

Tonnage

(Mt)

Fe

(%)

Au

(g/t)

Cu

(%)

CENTRAL

>30% Fe

163

3.80

618

45.1

0.061

0.098

>25% Fe

190

3.73

707

42.9

0.059

0.095

>20% Fe

200

3.70

739

42.1

0.060

0.094

>15% Fe

203

3.69

748

41.7

0.059

0.093

>10% Fe

204

3.68

752

41.6

0.059

0.093

> 5% Fe

205

3.68

753

41.6

0.059

0.093

> 0% Fe

205

3.68

753

41.6

0.059

0.093

SOUTH

>30% Fe

31

3.60

113

39.7

0.130

0.121

>25% Fe

32

3.59

115

39.5

0.130

0.121

>20% Fe

32

3.59

115

39.5

0.130

0.121

>15% Fe

32

3.59

115

39.5

0.130

0.121

>10% Fe

32

3.59

115

39.5

0.130

0.121

> 5% Fe

32

3.59

115

39.5

0.130

0.121

> 0% Fe

32

3.59

115

39.5

0.130

0.121

Total

> 15% Fe

235

3.68

863

41.4

0.068

0.097

Mining

The Central Zone massive magnetite mineralization is located between approximately 350 m and 800 m below surface and, as such, it was deemed potentially mineable by open pit or underground methods. An analysis of open pit and underground mining scenarios lead to the conclusion, based on the current data and assumptions, that an underground block cave mine would provide the most favourable economic results. The underground mine was designed on two sections, Block 1 and Block 2. The variability in the geometry of the bottom of the resource required the use of the blocks being mined from different levels. Block 1 extracts the highest grade, thickest and deepest material first. Block 2 is planned to extract the remaining resource from a higher undercut elevation, adjacent to Block 1.   

Based on the geometry, mining method and bulk density of the mineralization in the Central Zone, the production capacity of the mine was estimated to be 75,000 tpd or 27.4 MTPY. The mine production life is 24 years and includes a 5-year, straight-line production ramp-up. Table 3 shows the life of mine mill feed total by mining block.


Table 3 : LOM Mill Feed

Mining Block

Diluted Totals

MTonnes

Fe Grade

(%)

Cu Grade

(%)

Au Grade

(g/t)

Block 1

401

39.7

0.07

0.04

Block 2

160

37.7

0.11

0.09

Mineralized development muck

18

46.8

0.11

0.07

Total

580

39.4

0.09

0.05

Cost Estimates

Operating (“OPEX”) and capital (“CAPEX”) costs were estimated using a combination of first principles, reference projects and industry experience. Table 4 shows the calculated OPEX and Table 5 shows the CAPEX. A 20% contingency was used on capital costs to capture expenses not included in the estimates.

Table 4 : Unit OPEX Estimate Summary

Description

Unit

Cost

Mining

$/t milled

4.73

Beneficiation

$/t milled

1.70

Magnetic separation and Filtering

$/t milled

1.50

Flotation plant

$/t milled

0.30

Pelletizing plant

$/t milled

3.14

Site services

$/t milled

0.50

G&A

$/t milled

0.80

OPEX per tonne milled

$/t milled

12.67

OPEX per tonne of pellets produced

$/t of pellets

22.16


Table 5 : CAPEX Estimate Summary

Description

Unit

Pre-Production

Post Start-up

Total

Pre-construction

Construction

Mine Development

M$

136

396

656

1,188

Mine Mobile Equipment

M$

1

111

889

1,001

Mine Construction

M$

2

255

 

257

Beneficiation

M$

 

168

 

168

Mag Sep, Float, Slurry, Filter

M$

 

111

 

111

Pellet plant

M$

 

1,010

 

1,010

Tailings Dam

M$

 

59

129

188

Other

M$

95

40

 

135

EPCM

M$

 

354

 

354

Sustaining capital

M$

  

1,807

1,807

Capital cost w/o contingency

M$

234

2,504

3,482

6,219

Contingency (20%)

M$

46

501

696

1,243

TOTAL CAPITAL COST

M$

280

3,005

4,178

7,462

Preliminary Economic Assessment Results

A preliminary discounted cash flow analysis was conducted for the project using a range of iron pellet prices. The economic analysis used inferred mineral resources exclusively and, therefore, only provides a preliminary overview of the project economics based on broad, factored assumptions. Inferred mineral resources are considered too speculative geologically to have the economic considerations applied to them to be categorized as mineral reserves, and there is no certainty that the inferred resources will be upgraded to a higher resource category. There is also no certainty that the results of this preliminary economic assessment will be realized.

The main economic parameters used in the cash flow analysis are shown in Table 6. It was assumed that the project will be equity financed.

Table 6 : Main Economic Analysis Assumptions Common to all Cases

Item

Unit

Value

Copper Price

$/lb

2.00

Gold Price

$/oz

650

Iron recovery

%

93.4

Copper recovery

%

50

Gold recovery

%

50

Iron pellet grade

% Fe

64.5

Copper concentrate grade (Cu)

% Cu

22

Copper concentrate grade (Au)

 Au g/t

13.8

Payable iron (in pellets)

%

100

Payable copper (in Cu cons)

%

96.5

Payable gold (in Cu cons)

%

97

Offsite copper concentrate costs

  

Transport (all in)

$/wmt Cu concentrate

100

Treatment

$/dmt Cu concentrate

70

 Cu Refining

$/payable lb Cu

0.07

Au Refining

$/payable oz Au

6

Discount rate

%

10

Four cases were used in the cash flow analysis to demonstrate the variation of project economics with iron pellet price. All other variables were kept constant for all cases including the life of mine mill feed tonnes and grade. The four cases all used the assumption that pellets would be the final product. Case 1 assumed blast furnace pellets would be produced. The other cases assumed direct reduction pellets would be produced. Pellet prices were obtained from an independent market study, the 3-year average and reference public-domain reports and are shown in Table 7. Direct reduction pellets were assumed to have a 10% premium over blast furnace pellets.

Table 7 : Iron Pellet Price Assumptions

Case

Pellet Type

Pellet Price

(US¢/mtu)

Reference

1

Blast furnace

198

Independent market opinion for BF pellets

2

Direct reduction

169

3-year average (154 ¢/mtu)+ 10% DR pellet premium

3

Direct reduction

218

Independent market opinion (198 ¢/mtu)+ 10% DR pellet premium

4

Direct reduction

253

2008 public domain scoping study (230 ¢/mtu) + 10% DR pellet premium

The preliminary economic analysis results are shown in Tables 8 and 9.

Table 8 : NPV Results by Case

Taxation Assumption

Parameter

Unit

Net Present Value (NPV)

Case 1

BF Pellets

198 ¢/dmtu

Case 2

DR Pellets

169 ¢/dmtu

Case 3

DR Pellets

218 ¢/dmtu

Case 4

DR Pellets

253 ¢/dmtu

After Tax

0% discount rate

B$

         17.6

       13.7

       20.2

       24.9

8% discount rate

B$

            3.3

         2.2

         4.1

         5.4

10% discount rate

B$

            2.1

         1.3

         2.7

         3.7

12% discount rate

B$

            1.3

         0.6

         1.7

         2.5

Pre Tax

0% discount rate

B$

         27.3

       21.3

       31.4

       38.7

8% discount rate

B$

            5.8

         4.1

         7.0

         9.0

10% discount rate

B$

            4.0

         2.7

         4.9

         6.4

12% discount rate

B$

            2.7

         1.7

         3.4

         4.6

Table 9 : IRR and Payback period

Parameters

Unit

Case 1

BF Mid

Case 2 DR Lower

Case 3

DR Mid

Case 4

DR Upper

After tax IRR

%

18

15

20

23

Pre tax IRR

%

23

19

25

29

Payback Period (Post Tax, 10% DR)

Production years

7.6

10.0

6.8

5.7

A simplified sensitivity analysis was conducted and showed the project is most sensitive to metal price which yields a 45% variation in after-tax NPV10% for a 20% change in metal price. Mill feed grade is almost identical to the metal price sensitivity. The project is less sensitive to operating costs than capital costs. See Figure 1 below.

[techrep004.jpg]

Figure 23.1 : Case 3 Sensitivity Graph

Conclusions

The results of this preliminary economic assessment indicate that, based on the preliminary data available and assumptions used, the Pampa de Pongo Project is an economically robust project that warrants further exploration and study. The project is situated in a very favourable location within a short distance of infrastructure, including a deep-sea port facility, located 38 km west of the deposit. Drilling and magnetic surveys infer the presence of a large, massive magnetite mineralized Central Zone that is conducive to underground bulk mining. The South Zone resource and other exploration targets, particularly the East Zone, may enhance the total material available for exploitation. The metallurgy test work done on the Central Zone mineralization confirmed the production of pilot-scale direct reduction pellets with low levels of deleterious elements.

Recommendations

SRK recommends two stages of work to take Pampa de Pongo to the next level of development.  These recommendations are described below.  The Preliminary Economic Assessment will be contingent upon the success of the East Zone exploration drilling, while the Pre-feasibility Study will be contingent on the success of the definition drill programs of the East Zone and/or the Central Zone.

1. East Zone Exploration & Preliminary Economic Assessment

Exploration of the East Zone should be accelerated and prioritized to allow concurrent development with Central Zone, assuming positive results. Critical work items include:

  • Exploration Drilling – six holes to prove economic potential and determine Inferred Resources (4,500 metres).
  • Preliminary Economic Assessment – Scoping level assessment including conceptual mine design, resource estimate, metallurgy and economic value. The critical aspect would be to determine if the potential resource could be developed concurrently with Central  Zone, increasing the annual output of the operation.

2. Central and East Zone Pre-feasibility Study

The major components of the study should be as follows:

  • Central Zone Definition Drilling – approximately 17 drillholes, totalling 14,500m, resulting in a 200-metre drill spacing. This should be sufficient to upgrade existing resources from Inferred Resources to at least an Indicated Resource category. All permits are in pace to complete this work.
  • East Zone Definition Drilling - Rapid definition drilling to dovetail preliminary feasibility-level work with that being undertaken for Central Zone. This drilling would comprise approximately six additional drillholes for 4,500 metres.
  • Geotechnical Studies – concurrent with definition drilling, detailed geotechnical analysis including further structural definition and hydrogeology.
  • Environmental baseline – environmental monitoring, including groundwater assessment and preparation of Environmental Impact Assessment.
  • Metallurgy – Multiple bench-scale metallurgical tests to quantify metallurgical variation within the Central Zone. Further development of the furnace firing cycle to optimize and enhance blast furnace pellet quality. Additional grinding to reduce silica even further and conversely increase total iron grade in the pellets.
  • Central Zone Condemnation Drilling -  If not already defined through definition drilling, the outer edges of the Central Zone must be defined through up to 5 condemnation drillholes for a total of 4,250 metres.

Costs are based on a contractor quote of $120/m drilled and an estimated geotechnical logging and assaying cost of $30/m drilled.

Table : Summary of Recommended Diamond Drilling


Zone

No. of Holes

Est. Total Metres

Purpose

Estimated Cost ($M)

Priority 1

East

6

4,500

Confirmation of geophysical anomaly

0.7

East

6

4,500

Resource definition

0.7

Priority 2a

Central

17

14,450

Resource definition, technical data

2.2

Priority 2b

Central

5

4,250

Deposit boundary definition

0.6

TOTAL DRILLING

34

27,700

 

4.2

The cost of the East Zone mineral resource estimation and Preliminary Economic Assessment is estimated to be $200,000.

The cost of the Central and East Zone Preliminary Feasibility Study is estimated to be $2,600,000.

The total cost of the drilling, PEA and PFS is estimated to be $7M.










Table of Contents



1

Introduction


2

Reliance on Other Experts


3

Property Description and Location


4

Accessibility, Climate, Local Resources, Infrastructure and Physiography


5

History


5.1

Deposit Discovery


5.1.1

Rio Tinto Exploration


5.1.2

Previous Resource Estimates


6

Geological Setting


7

Deposit Types


8

Mineralization


9

Exploration


9.1

Transient Electromagnetic Survey


9.2

Ground Magnetic Surveys


9.3

Magnetic Inversion Models


9.3.1

Correlation of Inversions with Mineralization


10

Drilling


10.1

Iron Intersections


10.2

Copper-Gold Intersections


11

Sampling Method and Approach


12

Sample Preparation, Analyses and Security


12.1

Rio Tinto


12.2

Cardero


13

Data Verification


14

Adjacent Properties


15

Mineral Processing and Metallurgical Testing


15.1

Mineral Processing


15.1.1

Beneficiation


15.1.2

Pelletization


15.2

Metallurgical Testing


15.2.1

Testing Facilities


15.2.2

Sample Selection


15.3

NRRI Metallurgical Testing


15.3.1

Metallurgical Test Results


15.3.2

Blast Furnace End-User Considerations


15.3.3

Chemical Analysis and Quality Control


15.3.4

Future Work


15.4

MIDREX® Metallurgical Testing


15.4.1

Material Evaluation Test Results


15.4.2

DRI End-User Considerations


15.4.3

Detailed Results


15.4.4

Summary


15.4.5

Future Work


16

Mineral Resource and Mineral Reserve Estimates


16.1

Introduction


16.2

Geological Model


16.3

Data Used in Resource Estimation


16.4

Statistical Analyses


16.4.1

Bulk Density Data


16.5

Estimation Methodology


16.6

Block Models


16.7

Estimation Parameters


16.8

Block Model Validation


16.8.1

Comparison of Block Estimates with Composites


16.8.2

Swath Plots


16.9

Mineral Inventory


16.10

Mineral Resource Classification


16.11

Mineral Resource Statement


16.12

Comparison with Previous Resource Estimates


17

Other Relevant Data and Information


18

Additional Requirements for Technical Reports on Development Properties and Production Properties


18.1

Mining

18.1.1

Mining Context


18.1.2

Mining Method Selection


18.1.3

Mine Design


18.1.4

Mine Development Design


18.1.5

Mobile Equipment


18.1.6

Development and Production Schedules


18.1.7

Mining Support Services


18.2

Recoverability


18.3

Markets


18.3.1

Iron Ore Market Background


18.4

Contracts


18.5

Environmental Considerations


18.6

Taxes.


18.6.1

Royalties


18.6.2

Value Added Tax


18.6.3

Corporate Income Tax


18.6.4

Depreciation and Amortization


18.7

Operating Cost Estimates (OPEX)


18.7.1

Mining OPEX


18.7.2

Mineral Processing OPEX


Beneficiation


18.7.3

General and Administration (“G&A”) and Site Services OPEX


18.7.4

OPEX Summary


18.8

Capital Cost Estimates (CAPEX)


18.8.1

Mining


18.8.2

Mineral Processing


18.8.3

CAPEX Summary


18.9

Economic Analysis


18.9.1

Sensitivity Analysis


18.10

Payback


18.11

Mine Life


19

Interpretation and Conclusions


19.1

Risk and Opportunities


20

Recommendations


21

References


22

Abbreviations and Acronyms


23

Date and Signature Page


List of Tables



Table 2: Mineral Inventory at Various Cut-off Grades


Table 3: LOM Mill Feed


Table 4: Unit OPEX Estimate Summary


Table 5: CAPEX Estimate Summary


Table 6: Main Economic Analysis Assumptions Common to all Cases


Table 7: Iron Pellet Price Assumptions


Table 8: NPV Results by Case


Table 9: IRR and Payback period


Table 10: Summary of Recommended Diamond Drilling


Table 3.1: List of Mining Claims Controlled by Cardero


Table 3.2: Royalty Rates


Table 3.3: List of Claim Corner Coordinates


Table 5.1: Historical Exploration


Table 5.2: Rio Tinto Diamond Drilling Iron Grade Highlights


Table 5.3: Rio Tinto Reverse Circulation Drilling Highlights


Table 5.4: Concentrate Chemical Composition


Table 5.5: Fired Pellet Chemical Composition


Table 5.6: Historic Mineral Resource Estimate by Helsen (2005)


Table 9.1: Exploration Work by Cardero


Table 10.1: Cardero Drilling Highlights


Table 11.1: Phase of Drilling and Description of Sample Type


Table 15.1: BF Grade Pellet Comparison and QA/QC


Table 15.2: Linder Test Results


Table 15.3: Hot Load Test Results


Table 15.4: Material Evaluation Results


Table 16.1: Modeled Domain Names, Location and Number of Drillholes


Table 16.2: Capped Au and Cu Assays


Table 16.3: Basic Statistics of Deleterious Substances in the Central Zone


Table 16.4: Block Model Extents


Table 16.5: Estimation Parameters


Table 16.6: Mineral Inventory at Different %Fe Cut-off Grades


Table 16.7: SRK Classified Mineral Resources for Pampa de Pongo at 15% Fe cut-off


Table 16.8: Comparison of SRK and Previous Resource Estimates of the Central Zone


Table 16.9: Comparison of SRK and Previous Resource Estimates of the South Zone


Table 17.1: Possible Pre-Production Project Schedule


Table 18.1: Central Zone Geotechnical Parameters


Table 18.2: South Zone Geotechnical Parameters


Table 18.3: South Zone Open Pit Geotechnical Parameters


Table 18.4: Initial Whittle™ Optimization Parameters


Table 18.5: PCBC Footprint Finder Parameters


Table 18.6: Mining Block Tonnages and Grades


Table 18.7: Primary Components of Mine Design


Table 18.8: Mobile Equipment Fleet


Table 18.9: Development and Equipment Productivity Assumptions


Table 18.10: Life of Mine Development Schedule


Table 18.11: Production and Development Assumptions


Table 18.12: Life of Mine Production Schedule


Table 18.13: LOM Plan Mineralized Material Sources


Table 18.14: Mining Royalty Rates


Table 18.15: Depreciation Assumptions


Table 18.16: Mine OPEX Assumptions


Table 18.17: Unit OPEX Estimate Summary


Table 18.18: Development Costs


Table 18.19: CAPEX Estimate Summary


Table 18.20: Iron Pellet Price Assumptions by Case


Table 18.21: Main Economic Analysis Assumptions Common to all Cases


Table 18.22: IRR and Payback Period


Table 18.23: NPV Results by Case


Table 18.24: Sensitivity Results by Case


Table 18.25: Payback Period by Case (Post tax NPV10% Model)


Table 20.1: Summary of Recommended Diamond Drilling



List of Figures



Figure 3.1: Pampa de Pongo Tenure


Figure 4.1: Location & Infrastructure


Figure 4.2   Site Infrastructure


Figure 6.1: Regional Geology


Figure 6.2: Stratigraphic Column (modified from Hawkes et. al. 2002)


Figure 8.1: TMI with Mineralized Zones and Diamond Drillholes


Figure 9.1: 600m Horizontal Depth Slice through 3D Magnetic Model


Figure 10.1: Drillhole Locations


Figure 10.2: Central Zone Composite Cross Section at 517400 E (looking east)


Figure 10.3: Central Zone Composite Cross Section at 8301600 N (looking north)


Figure 13.1: Fe Quarter Core Duplicates


Figure 13.2: SiO2 Quarter Core Duplicates


Figure 13.3: S Quarter Core Duplicates


Figure 13.4: Au Quarter Core Duplicates


Figure 13.5: Cu Quarter Core Duplicates


Figure 13.6: P and Mn ALS-Chemex Internal Pulp Duplicates


Figure 13.7: SiO2, Al2O3, MgO, TiO2 and CaO ALS-Chemex Internal Pulp Duplicates


Figure 13.8: Comparison of SG values of Inspectorate Laboratories vs. Field Values


Figure 15.1: Proposed SABC Circuit Process Flowsheet


Figure 15.2: Proposed PDP Straight-Grate Induration Circuit Process Flowsheet


Figure 16.1: 3D Magnetics and Drillhole Locations in the Central and East Zones


Figure 16.2: Plan View of Zone Shapes and Locations (Central: Green, East: Red, South S1: Cyan, South S2: Purple)


Figure 16.3: Statistics of De-clustered Composite Fe (%) Assays in the Four Mineralized Domains


Figure 16.4: Statistics of De-clustered Composite Au (g/t) Assays in the Four Mineralized Domains


Figure 16.5: Statistics of De-clustered Composite Cu (%) Assays in the Four Mineralized Domains


Figure 16.6: Correlation between Fe and S.G. Values


Figure 16.7: Comparison of block estimates of accumulation (Fe*SG) and SG with drillhole composite accumulation and SG contained within the blocks


Figure 16.8: De-clustered Average Composite Grades Compared to Block Estimates in all Zones


Figure 16.9: Estimated Grades for 8301687 E-W Section


Figure 18.1: Preliminary Geotechnical Domains for the Central Zone (section view looking east)


Figure 18.2: Preliminary Geotechnical Domains for the South Zones S1 (left) and S2 (right) (section view looking east)


Figure 18.3: Principles of Caving Methods


Figure 18.4: Central Zone Section View Showing a Single Mining Block with Excluded Mineralization and Excess Waste


Figure 18.6: Mine Development and Central Mineralized Zone Looking South (200 m gridlines)


Figure 18.7: Mine Development and the South Mineralized Zone Looking East (200 m gridlines)


Figure 18.8: Plan View of Mine Development Design


Figure 18.9: Isometric View of Mine Development Design


Figure 18.10: Ventilation Schematic


Figure 18.11: Case 1 Sensitivity Graph


Figure 18.12: Case 2 Sensitivity Graph


Figure 18.13: Case 3 Sensitivity Graph


Figure 18.14: Case 4 Sensitivity Graph


Figure 18.15: Percent Change to After-Tax NPV10% for Selected Case 3 Variables


Figure 18.16: Case 3 Project Post Tax Discounted Cash Flow Projection



List of Appendices


APPENDIX A: Geotechnical Information

APPENDIX B: Cash Flow Sheets

APPRNDIX C: Structural Geology Report



1

Introduction

This Preliminary Economic Assessment Technical Report was prepared by SRK for Cardero Resource Corp. The purpose of the report is to present SRK’s mineral resource estimate for the property and present preliminary economic findings for the Pampa de Pongo project based on preliminary mining and metallurgical investigations. The report also provides guidance for potential advancement of the project in the form of several recommendations.

A site visit to the Pampa de Pongo property and to Cardero offices in Lima was conducted by Qualified Persons (“QP”) Gordon Doerksen, P.Eng. and George Wahl, P.Geo, both of SRK, on March 17 to 20, 2008.

The project diamond drill core for the project is stored in Lima and all drill core from holes PP10, PPD-002, PPD-003, PP20A and PP21 was inspected by Doerksen and Wahl on March 17, 2008. Core from drillholes RTDDH-1, RTDDH-2, RTDDH-3, PPD-19 and RTDDH-9 was inspected by Wahl on March 20, 2008.

During the inspection of the Pampa de Pongo property on March 18, 2008, several drillhole markers (pipes encased in concrete bases) were inspected and coordinates verified by Wahl using a hand-held GPS unit.

QPs Marek Nowak and Leonard Holland did not perform a site visit as the project is in the early exploration stage and nothing applicable to Mr. Nowak’s or Mr. Holland’s work could be gained from visiting the project.

Drillhole and assay information from both the Rio Tinto and Cardero drill programs was provided by Cardero Resource Corp. 3D magnetic inversion data was provided by Quantec Geophyiscs Worldwide of Argentina. Metallurgical data was produced by

The metallurgical testing was undertaken by the Natural Resources Research Institute (NRRI) in Minnesota, United States.

Details of additional sources of information are included where appropriate in the body of the report.

All units in this report are based on the International System of Units (“SI”), except for some units which are deemed industry standards such as troy ounces (“oz”) for precious metals and pounds (“lb”) for base metals. All currency values are United States dollars (“$” or “US$” or “USD”) unless otherwise noted.

This report uses many abbreviations and acronyms common in the mining industry, most of which are defined in the body of the text. Further explanations can be found in Section 22.

The economic analysis conducted in this report uses Inferred Mineral Resources exclusively and only provides a preliminary overview of the project economics based on broad, factored assumptions. Inferred mineral resources are considered too speculative geologically to have the economic considerations applied to them to be categorized as mineral reserves, and there is no certainty that the inferred resources will be upgraded to a higher resource category. There is also no certainty that the results of this preliminary economic assessment will be realized.

 


2

Reliance on Other Experts

This report relies on marketing studies and pricing opinions for iron ore and iron pellets provided by Global Strategic Solutions Inc., who provided a detailed analysis of historical pricing and opinions of long-term pricing from multiple sources. The report was authored by Sara A. Hornby, C.Eng., Ph.D., President of Global Strategic Solutions, Inc. Ms Hornby’s experience includes 5 years in the UK steel industry, 18.5 years in the industrial gas arena in the metals, prime materials, mining, glass, cryogenics, welding and heat treatment markets, 2 years as Director of Operations and Technology for a start up company improving EAF steelmaking energy use, 4 years with Midrex Technology Inc. as their DRI/HBI steelmaking and marketing specialist before starting her own consulting company.

There is no guarantee that the market studies will prove to be reliable. The selection of accurate metal prices is made extremely difficult in times of high commodity price volatility as is current the case in the world markets. SRK has used a range of iron prices to demonstrate the implications of iron prices variations.

 


3

Property Description and Location

Pampa de Pongo is located on the southern coastal plain of Peru approximately 50 km south of the city of Nazca and 550 km southeast of the Capital, Lima (Figure. 3.1). The geographic center of the property is located at 74° 50" W longitude, 15° 23" S latitude. In detail, the property is located in the province of Caraveli in the governing jurisdiction of Arequipa.

According to Peruvian regulations, the right to explore for and exploit minerals is granted by way of mining concessions. In order to obtain a mining concession, a mining claim (a Petitorio Minero) must be presented before the INGEMMET (the Peruvian Mining Authority). A Petitorio Minero is only an application for a mining concession, therefore it does not grant any exploration or exploitation rights but concedes to its holder a preference over the area with respect to any other subsequent overlapping Petitorios Mineros. On acceptance of the application, the Petitorio Minero is converted to a mining concession (a Concesión Minera). The regulations governing the process of application for Petitorio Minero result in overlapping tenure until such times as the Concesión Minera is granted. In Figure 3.1, several of Cardero’s Petitorios Mineros are overlapping with previously existing tenure. Only those portions of the application which are on open ground will be considered for issue as Concesión Minera to Cardero.

The Pampa de Pongo property consists of 8 adjoining mining concessions and 10 mining claims. Together, they form an area 19 km long and up to 10 km wide (15,300 hectares). The mineral resources and mineralized zones are located within the 8 mining claims. There are no historical mine workings, waste management facilities, tailings ponds or important natural features within the area of the Pampa de Pongo claims block. The particulars of the claims are listed in Table 3.1 and a list of the claim corner coordinates is listed in Table 3.3. The claim corners have not been physically surveyed and marked.

A Concesión Minera gives the holder the right to explore for, and to exploit, such mineral resources which may be found within a body of ground of unlimited depth, the boundaries of which are defined by vertical planes corresponding to the sides of a cube, rectangle or closed polygon whose vertices are defined by UTM (Universal Transverse Mercator) coordinates. An Acumulación Derecho Minero Titulada is granted when two or more, contiguous or overlapping, mining concessions, both or all of which belong to the same title holder, are grouped together in the National Mining Register as a single mining concession, also under the same title holder, adopting the registration date of the oldest of the accumulated mining concessions.

There are no known environmental liabilities associated with the property. The property has almost no surface disturbance other than historical drillhole stand pipes mounted in small concrete blocks, survey markers and sporadic vehicle tracks.

Cardero Resource Corp. completed an option to purchase a 100% undivided interest in the property from Rio Tinto Mining and Exploration, Sucursal Peru (Rio Tinto). This was completed through staged payments totalling $565,000 and issuing 70,000 shares over a four year period ending on January 27, 2008.

The Concesión Mineras are presently in the process of being transferred from Rio Tinto to Cardero Peru, however, as of writing they are still held by Rio Tinto. Rio Tinto anticipates that the concession transfers will be completed imminently. The 10 Petitorios Mineros are held 100% by Cardero Peru.

Table 3.1 : List of Mining Claims Controlled by Cardero

CODE

CONCESION

TITLE

HECTARES

010327993

RETOZO-50

RIO TINTO MINING AND EXPLORATION LTD

1,000

010224594

RETOZO 85

RIO TINTO MINING AND EXPLORATION LTD

1,000

010224694

RETOZO 86

RIO TINTO MINING AND EXPLORATION LTD

1,000

010225094

RETOZO 90

RIO TINTO MINING AND EXPLORATION LTD

1,000

010225194

RETOZO 91

RIO TINTO MINING AND EXPLORATION LTD

1,000

010225294

RETOZO 92

RIO TINTO MINING AND EXPLORATION LTD

1,000

010226194

RETOZO 101

RIO TINTO MINING AND EXPLORATION LTD

1,000

010226294

RETOZO 102

RIO TINTO MINING AND EXPLORATION LTD

1,000

010528608

FELINO 1

CARDERO PERU S.A.C.

1,000

010528708

FELINO 2

CARDERO PERU S.A.C.

100

010528808

FELINO 3

CARDERO PERU S.A.C.

800

010528908

FELINO 4

CARDERO PERU S.A.C.

600

010529008

FELINO 5

CARDERO PERU S.A.C.

1,000

010529108

FELINO 6

CARDERO PERU S.A.C.

1,000

010529208

FELINO 7

CARDERO PERU S.A.C.

600

010529308

FELINO 8

CARDERO PERU S.A.C.

1,000

010529408

FELINO 9

CARDERO PERU S.A.C.

1,000

010529508

FELINO 10

CARDERO PERU S.A.C.

200

[techrep006.jpg]

Figure 3.1 : Pampa de Pongo Tenure

The Peruvian government collects royalties on behalf of all federal, provincial and municipal governments. The royalty rates are scaled according to gross revenue generated as shown in Table 3.2.

Table 3.2 : Royalty Rates

Gross Revenue

Royalty

$0 to $60M

1% of gross revenue

$60M to $120M

2% of gross revenue

above $120M

3% of gross revenue


Table 3.3: List of Claim Corner Coordinates

Claim Name

Corner

Easting

Northing

Claim Name

Corner

Easting

Northing

RETOZO-50

1

517,776

8,307,637

FELINO 3

1

513,776

8,305,637

2

517,776

8,305,637

2

513,776

8,299,638

3

513,776

8,305,637

3

512,776

8,299,638

4

513,776

8,306,637

4

512,776

8,301,637

5

511,776

8,306,637

5

511,776

8,301,637

6

511,776

8,307,637

6

511,776

8,302,637

RETOZO 85

1

517,776

8,305,637

7

512,776

8,302,637

2

517,776

8,302,637

8

512,776

8,303,637

3

514,776

8,302,637

9

511,776

8,303,637

4

514,776

8,304,637

10

511,776

8,304,637

5

513,776

8,304,637

11

512,776

8,304,637

6

513,776

8,305,637

12

512,776

8,305,637

RETOZO 86

1

522,775

8,304,637

FELINO 4

1

512,776

8,300,638

2

522,775

8,302,637

2

512,776

8,298,638

3

517,776

8,302,637

3

513,776

8,298,638

4

517,776

8,304,637

4

513,776

8,296,638

RETOZO 90

1

517,776

8,302,637

5

511,776

8,296,638

2

517,776

8,297,638

6

511,776

8,300,638

3

515,776

8,297,638

FELINO 5

1

515,776

8,298,638

4

515,776

8,302,637

2

515,776

8,294,638

RETOZO 91

1

519,776

8,302,637

3

512,776

8,294,638

2

519,776

8,297,638

4

512,776

8,296,638

3

517,776

8,297,638

5

513,776

8,296,638

4

517,776

8,302,637

6

513,776

8,298,638

RETOZO 92

1

521,776

8,302,637

FELINO 6

1

517,776

8,297,638

2

521,776

8,297,638

2

517,776

8,292,638

3

519,776

8,297,638

3

515,776

8,292,638

4

519,776

8,302,637

4

515,776

8,297,638

RETOZO 101

1

519,776

8,297,638

FELINO 7

1

518,776

8,292,638

2

519,776

8,292,638

2

518,776

8,290,638

3

517,776

8,292,638

3

515,776

8,290,638

4

517,776

8,297,638

4

515,776

8,292,638

RETOZO 102

1

521,776

8,297,638

FELINO 8

1

523,775

8,292,638

2

521,776

8,292,638

2

523,775

8,290,638

3

519,776

8,292,638

3

518,776

8,290,638

4

519,776

8,297,638

4

518,776

8,292,638

FELINO 1

1

515,776

8,309,637

FELINO 9

1

523,775

8,297,638

2

515,776

8,307,637

2

523,775

8,292,638

3

510,776

8,307,637

3

521,776

8,292,638

4

510,776

8,309,637

4

521,776

8,297,638

FELINO 2

1

513,776

8,306,637

FELINO 10

1

512,776

8,296,638

2

513,776

8,305,637

2

512,776

8,294,638

3

512,776

8,305,637

3

511,776

8,294,638

4

512,776

8,306,637

4

511,776

8,296,638

The Peruvian Ministry of Mines is responsible for issuing all permits relating to mining and exploration properties. There are currently discussions in Peru to establish an Environmental Ministry. It is not known how this potential change in government structure will affect the minerals industry. There are currently three levels of permit required at various stages of exploration.

Activity Type A:  Does not require a permit. Work includes geophysics, surface sample collection, and everything that doesn’t have an environmental impact.

Activities Type B:  Allows construction of up to 20 drilling pads as long as less than 5% of the land is disturbed. This permit has been granted to Cardero Peru S.A.C. Permit duration is typically for the drilling season.

Activity Type C:  Requires an Environmental Impact Assessment, which the government will review over 4 months. Cardero has initiated the necessary environmental work required to make this application.

Following a feasibility study, an exploitation permit is required to commence mining. This requires a second environmental report with all project details, after which applications are submitted for a milling concession, which will be reviewed by the government for 6-12 months.

Surface rights must be acquired from the government (at minimal cost) or from the appropriate community (on a lease or purchase basis).



4

 Accessibility, Climate, Local Resources, Infrastructure and Physiography

The Project area lies within the desert coastal tract of southern Peru, approximately 550 kilometres south of Lima (Figure 4.1), a northward continuation of the Atacama Desert of northern Chile. It is extremely arid, receives little annual rainfall (<1 cm/year) and is virtually devoid of vegetation, except in irrigated river valleys.

During the summer months, the climate is hot and dry, but there is often a moderating effect from the coast with an inflow of cool moist area in the morning and late in the day. During the winter months, the coastal areas are cool, and covered by sea mists which provide humidity of around 100%. Strong onshore (southeast, verging-to-south, verging-to-southwest) winds are generated at this time by solar heating of the hinterland and pull cooler air inland from the coastal belt. These winds carry the mist and a considerable burden of aeolian sand with them, the latter being deposited against rising ground as the strength of the winds diminishes.

The topography in the project area is essentially flat-lying. The areas overlying existing resources vary between 320 metres above sea level in the south (South Zone) to 420 metres above sea level to the north (Central Zone). Relief is more pronounced in the northeast portion of the claim block. However, there are no mineral resources or prospects thereof in this area at this time.

The closest major community is the town of Nazca (population 22,132 as at July 2005), which is located approximately 50 km north of the property. Nazca is located on the Pan American Highway about a 5-hour drive south of Lima and has most amenities and services available. The small rural community of Bella Union is located 18 km by road to the southeast. The coastal village of Lomas, from which all project–related field work has been conducted, and which hosts the core logging and storage facilities for the project, lies 19 km by road from the Pampa de Pongo project area.

The Pan American Highway traverses the western boundary of the property. The property is readily accessible by a series of secondary asphalted and graded roads that lead from the highway to the abandoned Acari mine site, which is located 3.5 km east of the claim area. All parts of the claim area are easy to reach either along the existing network of secondary roads or by short travel by 4x4 or on foot, off existing roads.

SRK is of the opinion that sufficient surrounding land is available for potential processing plant sites, mine and related infrastructure and tailings storage.

There is a 200 kV electrical transmission line running southward from Lima to the port of San Nicolas, which serves the Marcona Iron mine beneficiation and pelletizing plant. From there the line extends southeast to the rural community of Bella Union, traversing the Pampa de Pongo property. This is a lower capacity transmission line, rated between 33-66 kV. This existing line would require upgrading to service the proposed Pampa de Pongo mine.

Plans to extend the natural gas pipeline south to Nazca and onward to the port of San Juan have been finalized. In April 2008, following a tender process, the Peruvian authorities awarded the contract to Empresa Energía de Bogotá (EEB) and Transportadora de Gas del Interior (TGI). The contract stipulates a 30-month completion period (late 2010), which exceeds the requirements for the proposed Pampa de Pongo infrastructure

The deep-water port of San Juan is located 38 km by road from the Pampa de Pongo claim boundary. The Peruvian authorities are planning to develop the port of San Juan, which would become the largest port in Peru. The port is a naturally sheltered bay, with a current water depth of 8 m at a distance of 15 m from shore and a depth of 21 m at a distance of 380 m from shore. Proposed developments include multiple loading facilities. The tender documents for port development are expected to be released in the coming months.

[techrep008.jpg]

Figure 4.1 : Location & Infrastructure

Figure 4.2   Site Infrastructure

 


5

History

5.1

Deposit Discovery

Rio Tinto Mining and Exploration, Sucursal Peru (Rio Tinto) discovered Pampa de Pongo in 1994 while drill testing a large magnetic anomaly identified by an airborne magnetic-radiometric survey. It was a blind target, totally concealed by aeolian sand. The airborne survey was part of a regional exploration program designed to identify prospective Iron Oxide Copper Gold (IOCG) targets in the district. Follow-up drilling by Rio Tinto in 1995 and 1996 identified widespread significant magnetite mineralization hosted in sedimentary and volcanic rocks. Mineralization occurs in a northwest-trending structural corridor approximately one kilometre in width and in excess of 12 kilometres long.

5.1.1

Rio Tinto Exploration

History of exploration on the Pampa de Pongo property begins with regional exploration by Rio Tinto which led to its staking. Historical exploration work carried out by Rio Tinto is outlined in Table 5.1. Portions of this work were directed at identifying potential copper-gold mineralization. The phases of the exploration program relevant to what was to become the Pampa de Pongo iron-ore deposit are described in more detail below:

Table 5.1 : Historical Exploration

Exploration Technique

Scope

Geological mapping

Covering approximately 5,500 hectares

Soil Geochemistry

1,024 Enzyme Leach samples

Rock Samples

1,227 Rock Chip samples

Gravity

3,277 gravity stations (100-200m lines, 25m stations)

Airborne Geophysics

Heli-magnetic & Radiometric survey (200m line spacing)

Induced Polarization

7 x 100m spacing (28.3km), 3 x 200m spacing (12.8km)

Reverse Circulation Drilling

6 drillholes for 1,401m, maximum depth 253m

Diamond Drilling

9 drillholes for 4,883.6m, maximum depth 800.4m

Metallurgical

2 x 35kg samples for bench-scale metallurgical test work

Rio Tinto Airborne Geophysical Survey

Rio Tinto commissioned a regional helicopter magnetic and radiometric survey. Cardero has access to survey data coincident with the Pampa de Pongo property.

Original data has been reviewed, on behalf of Cardero, by Quantec Geoscience, who produced a series of maps including Total magnetic Intensity, Analytical Signal, and 1st Vertical Derivative. Magnetic data demonstrates good to excellent correlation with mineralized magnetite bodies at depth, consequently magnetic data has subsequently been the major driver in exploration at Pampa de Pongo. The use of magnetic methods is described more fully in Section 10; however it is suffice to say that several additional high-priority targets remain to be fully tested.

Rio Tinto Gravity Survey

Rio Tinto collected measurements at 3,277 gravity stations within the Pampa de Pongo claim block. The data was collected on 100-200 metre line spacing, with measurements on 25 metres spacing along lines. Original data has been reviewed, on behalf of Cardero, by Quantec Geoscience, who produced a series of maps including Bouguer Anomaly map and a Residual Gravity map. Gravity data also shows good correlation with mineralized magnetite bodies at depth, due to the high specific gravity of magnetite, but it does not have sufficient resolution to be used in detailed exploration, nevertheless it does highlight several additional target areas for future follow-up exploration programs

Rio Tinto Drilling

Results from 9 diamond drillholes and 6 reverse circulation drillholes are summarized in Tables 5.2 and 5.3.

Table 5.2 : Rio Tinto Diamond Drilling Iron Grade Highlights

Drillhole

Target

Comment

From (m)

To

(m)

Interval (m)

Grade

(%Fe)

PPD-001

Centre of a large magnetic anomaly (Central Zone).


including

396

546.8

763

715.5

367

168.7

45.1

52.9

PPD-002

Planned 600m step out to the south from PPD-01.


including

410

460

556

556

146

96

30.3

33.9

PPD-003

Planned 500m step out to the east from drillhole PPD-01.


and

468

518

502

756

34

238

25.3

43.2

PPD-004

2600m step out, southeast from PPD-01.


and

276

384

366

536

90

152

41.0

30.1

PPD-005

Targeting edge of anomaly

     

PPD-006

Hole abandoned at 216m.

     

PPD-007

Drilled 850m NE of PPD-006.

 

112.8

150.0

37.2

50.4

PPD-008

Drilled 150m west of PPD-006.

 

210

321

111

38.9

PPD-009

Drilled 50m east of PPD-006.

 

259.4

392

132.6

48.0

Table 5.3 : Rio Tinto Reverse Circulation Drilling Highlights

Drillhole

Target

Depth

Significant Results

PPR-001

Shallow copper-gold

  

PPR-002

Shallow copper-gold

 

21m @ 0.43% copper & 0.68g/t gold

PPR-003

Shallow copper-gold

  

PPR-004

Shallow copper-gold

  

PPR-005

Shallow copper-gold

  

PPR-006

Shallow copper-gold

  

Results of the Rio Tinto work at Pampa de Pongo are discussed in a technical paper on the Marcona IOCG district by Rio Tinto staff geologists and researchers from Queen’s University, (Hawkes et. al. 2002).

Rio Tinto Metallurgy

Met-Chem Canada Inc. conducted preliminary metallurgical testwork on behalf of Rio Tinto. The study investigated the feasibility of producing a pellet feed chemically suitable for Direct Reduction pellet manufacture and, if suitable, to evaluate the resulting pellets on a laboratory scale. The drillcore samples were taken from South Zone and East Zone. No samples from Central Zone (the focus of this report) were tested, although mineralization and metallurgical properties would be expected to be similar.

Following investigation of conventional beneficiation techniques (screening, magnetic separation and flotation), it was demonstrated that the Pampa de Pongo mineralized material could be concentrated up to 66-69% iron through a simple Low Intensity Magnetic Separation (LIMS) process, with the remaining gangue (waste) material consisting of 2.5-3.0% magnesium oxide (MgO) and 0.5-1.5% sulphur. Concentrates were obtained through 3 passes in a magnetic separator and recoveries are described by Met-Chem as excellent. Concentrates showed a high magnetite content that normally results in low energy consumption in the firing process. Results are shown in Table 5.4.

Table 5.4 : Concentrate Chemical Composition

Element/Compound

Sample 1  (4B/9B)

Sample 2  (4B/4B)

Head Grade

Concentrate

Head Grade

Concentrate

Fe %

56.2

69.0

53.0

66.7

SiO2 %

5.42

0.55

6.19

0.84

Al2O3 %

1.66

0.69

2.17

0.92

MgO %

8.00

2.60

9.62

2.89

S %

1.84

0.54

2.37

1.51

Fe3O4 %

64.3

82.5

61.0

80.2

Pellets produced from Sample 1 were of good quality and strength, meeting normal commercial specifications without further work. The pellets exhibited extremely low clustering indexes, which is highly desirable in the natural gas based Direct Reduction processes. Sample 1 contained a slightly higher than normal percentage of MgO but this did not affect the quality of the pellet produced. Additional slagging in the Electric Arc Furnace (EAF) could reduce the MgO content, however the high MgO content of the metalized pellet can be an advantage in the steelmaking process as most EAF refractory linings are MgO based, and the slag routinely needs to be conditioned to have 10~14% MgO so as to mitigate slag line refractory erosion.

Concentrate from Sample 2 contained higher MgO and higher sulphur content than Sample 1. The pellets produced from Sample 2 were weaker due to higher sulphur content in the concentrate. In this case, Met-Chem suggested that further test work, including a finer grind and more sophisticated flotation reagents, should further improve the concentrate chemistry.

Samples 1 and 2 were mixed with 0.8% bentonite, pelletized and fired. The fired pellet chemical compositions are shown in Table 5.5.

Table 5.5: Fired Pellet Chemical Composition

Parameter

Sample 1  (4B/9B)

Sample 2  (4B/4B)

Fe %

66.9

66.6

SiO2 %

0.81

1.22

Al2O3 %

0.74

0.95

MgO %

2.61

2.91

S %

<0.01

0.01

It was concluded that the strength and reduction properties (reducibility, linder and clustering) of the pellets are well within the acceptable range for typical commercial products. Finally, additional processing steps were recommended to reduce the sulphur content of Sample 2. Requirements for further work are typical of early-stage metallurgical test work.

5.1.2

Previous Resource Estimates

Helsen (2005) in a report entitled, Geological Valuation Report of the Pampa de Pongo Property, dated August 6, 2005 estimated mineral resources for the Central and South Zones. The available magnetic data and 3D magnetic modelling generated at that time resulted in an estimate that was more liberal than the current SRK resource estimate. The Helsen estimation methodology also used a less constrained grade envelope to define the mineralized zones used for resource estimation. Historic mineral resources reported by Helsen are included in the following Table 5.6. SRK is of the opinion that the estimate and classification of resources is in accordance with sections 1.2, 1.3 and 2.4 of the National Instrument 43-101 and was appropriately based on the data available and resulting interpretations made at that time.

Table 5.6 : Historic Mineral Resource Estimate by Helsen (2005)

Mineralized Zone

Inferred Resources (million tonnes)

Fe

(%)

Cu

(%)

Au

(g/t)

Central Zone

848

44.9

0.12

0.07

South Zone – East

100

43

0.15

0.22

South Zone – West

5

43.8

0.27

0.26

Total

953

44.7

0.12

0.09

 


6

Geological Setting

The Pampa de Pongo deposit is located in the Marcona (Fe-Cu) District, named after the Marcona Iron mine, which has operated continuously since 1953. The oldest rocks exposed are Precambrian in age and part of the Coastal Basal Complex, consisting of gneisses, potassium-rich granites and migmatites that underlie the western and southern parts of the Marcona area (Figure 6.1 Regional Geology). At the Mina Justa Deposit – Marcona mine area, this basement metamorphic complex is overlain by carbonates and pelitic sediments of the Lower Paleozoic Marcona Formation together with marine andesitic to basaltic volcanics (and associated, hypabyssal intrusives), volcaniclastics and sediments of the middle to Upper Jurassic Rio Grande Formation and Bella Union Volcanics. Small remnants of the Upper Jurassic Jahuay Formation (mixed volcanics and sediments) and Lower Cretaceous Jauca Formation (sediments) cap the Jurassic succession, as seen in the vicinity of Pampa de Pongo. A summarized stratigraphic column for the Marcona District is included (Figure 6.2 Stratigraphic Column).

Major intrusives include the San Nicolas Batholith, which predates the Jurassic sequence and underlies a large tract south and west of the Marcona mine area. The more extensive Lower Cretaceous Coastal Batholith underlies large areas east of Pampa de Pongo. Abundant intra-mineral and post-mineral andesitic dykes, sills and small stocks referred to as “ocöite” are intimately associated with the main deposits in the Marcona district and are inferred to signify long-lived thermal activity associated with high heat flow and transfer of hydrothermal fluids centered over magmatic sources at depth. The ocöite is generally coarsely porphyritic and contains abundant large plagioclase phenocrysts, commonly with reaction rims recording multiple magmatic dissolution/overgrowth events.

The majority of the mineralization at the Marcona mine is hosted by carbonates of the Marcona Formation. The Pampa de Pongo deposit is hosted by the younger Jurassic Juhuay Formation. Lithologies, seen in drill core at the fringes of the deposit, include coarse volcaniclastics, conglomerates, banded siltstones, shales, sandstones and carbonate-rich sediments. Within the massive magnetite replacement body, unmineralized remnants are extremely rare. The intersection of basement gneisses in some drillholes suggests that the underlying Marcona Formation and Rio Grande Formation may have been eroded in this part of the district.



[techrep009.jpg]

Geology taken from GMAP (Geological Mosaic for Andean Peru), a re-digitized and re-attributed national geological database supplied by Anglo Peruana, Lima, based on 1:100,000-scale national geological map quadrangles produced by the governmental “Instituto Geológico Minero y Metalúrgico” (INGEMMET) and modified locally from Cardero’s field mapping

Figure 6.1 : Regional Geology




[techrep011.jpg]

Figure 6.2 : Stratigraphic Column (modified from Hawkes et. al. 2002)

 

7

Deposit Types

It is not proposed here to discuss the Iron Oxide Copper Gold (IOCG) model in any detail, please refer to Sillitoe (2003), the Australian Mineral Foundation volume (2000) and references therein for additional information. It is suffice to state that IOCGs are capable of forming in a broad range of geotectonic settings (orogenic collapse, anorgenic magmatism and subduction related), that there is general consensus regarding their mineralogy and alteration facies but that debate is ongoing regarding their parental fluid sources. These may be paraphrased as an ‘external’ brine / evaporate model (Barton et al, 2003) versus a purely magmatic model as eloquently discussed by Sillitoe (2003), finally Hitzman and co-workers propose a hybrid mixing (magmatic and external evaporitic fluids) model (2003).

Globally, IOCG deposits can vary between vein-style deposits, hydrothermal breccias, replacement mantos or massive replacements. In general, IOCG deposits display close spatial and temporal relationships with plutonic complexes and spatially associated fault systems.

The mineral deposits and showings located in the Marcona District have many of the defining characteristics of the iron oxide copper-gold (IOCG) class of deposits. Within this broadly defined deposit-type, the Marcona District deposits can be sub-categorized as hosting both iron dominated / Kiruna-type mineralization (dominated by magnetite mineralization such as Marcona Mine and Pampa de Pongo) and copper rich IOCG deposits (Mina Justa). Large Mesozoic IOCG systems are actively mined along the western coast of South America and include Chilean deposits such as Candaleria, Punta Del Cobre and Manto Verde; and Peruvian deposits such as Raul Condestable, Mina Justa, Marcona and Pampa de Pongo.

Pampa de Pongo exhibits replacement of original host volcano-sedimentary lithologies by massive magnetite. At Pampa de Pongo, iron oxides are the principal ore-forming minerals (hematite, magnetite or martite) with anomalous secondary copper and gold. To date, no economic accumulations of copper have been discovered, however, the potential for future copper-gold discoveries has not been ruled out.



8

Mineralization

Magnetite mineralization at Pampa de Pongo was originally discovered following an airborne magnetic survey and subsequent drill testing. Since the initial discovery, detailed ground magnetic surveys and 3D magnetic inversion models have formed a critical component of exploration at Pampa de Pongo with a strong correlation between the highest magnetic susceptibility (as defined by the 3D inversion models) and significant magnetic mineralization intersected in drillholes. To date, two inferred resources (Central Zone and South Zone) have been defined through drill-testing of magnetic anomalies. A third target (East Zone) has had only one drill test, which intersected 292 metres of semi-massive and massive magnetite mineralization at the extreme edge of a 3D magnetic anomaly. The remainder of this high-priority anomaly remains untested but it is estimated to represent potential for an additional 350-500 Mt of magnetite mineralization. A fourth high-priority anomaly (North Zone) remains completely untested.

At Pampa de Pongo the mineralized bodies are semi-concordant and appear to have been controlled by a number of factors, primarily structure, but also host rock lithologies, primary porosity and secondary porosity related to fracturing or brecciation. Mineralization comprises semi-massive to massive magnetite replacement zones with minor sulphides (pyrite-pyrrhotite-chalcopyrite ± minor marcasite, bornite, arsenopyrite, sphalerite and galena) and minor specular hematite. Alteration minerals include apatite, serpentine, talc, epidote, chlorite, magnesite, dolomite, calcite, actinolite, albite, tourmaline, garnet, phlogopite, biotite, sericite and K-spar.

There are four main zones of mineralization and potential mineralization, which can be observed on the Total Magnetic Intensity (TMI) map (Figure 8.1). Details of magnetic interpretation and 3D magnetic modeling are fully explained in Section10. The four zones are:

North Zone –Untested Exploration Target

The North Zone is a high-priority magnetic exploration target. The large 3D magnetic anomaly measures 1,000 metres north-south by 1060 metres east-west at its widest point. Target depth is estimated to be 600 metres, but only drill testing will determine the true extent of the anomaly. Drilling to date in the North Zone of the property has been shallow, targeting copper-gold targets and completed prior to detailed magnetic surveying.

Central Zone

The Central Zone is a two-layer mineralized body. The lower portion consists of a flat-lying, massive replacement lens up to 370 metres thick, measuring 1,060 metres east-west by 1,000 metres north-south at the widest point. Mineralization remains open. Mineralization is consistent, averaging approximately 62% magnetite with multiple intersections greater than 80% magnetite. In the overlying upper portion, predominantly massive iron mineralization is cut with a minor component of hypabyssal porphyry sills. The lower, high-grade massive replacement at Pampa de Pongo, based on current drill results, is almost completely devoid of unmineralized intrusive rocks. Continuity of high grade mineralization within this zone is very strong down-hole and between drillhole intercepts.

East Zone –Advanced Exploration

The East Zone was first delineated following detailed ground magnetic surveying in 2008. The anomaly has a footprint measuring 1,860 metres north-south by 650 metres east-west. The southern edge of the anomaly was partially tested by a Rio Tinto drillhole in 1996 (drillhole PPD-004), which intersected very significant massive magnetite mineralization. Mineralization consists of 3 high-grade mineralized lenses, including 86 metres (from 276-362 metres) at 41.38% iron, 66 metres (from 470-536 metres) at 44.28% iron, and 23.9 metres (from 609.2-633.1 metres) at 42.36% iron. A broader, continuous 292-metre interval grading 30.85% iron was intersected from 268-558 metres. Mineralization is shallow at 276 metres from surface.

South Zone

The South Zone is comprised of two separate lenses of mineralization. These near-surface zones consist predominantly of massive magnetite mineralization. The South 1 (“S1”) Zone measures 700 metres north-south by 900 metres east-west at the widest point with a true thickness of up to 120 metres, while the South 2 (“S2”) Zone measures 250 metres north-south by 500 m east-west at the widest point with a true thickness of 30 metres. Mineralization remains open in all directions. The upper portion of the deposit has been eroded. Geologically younger Quaternary to recent cover material is comprised of semi-lithified sandstones and conglomerates, ranging from 80-300 metres thick. The resource has not been included in the mine schedule presented in this report and represents future upside potential. Additional exploration is planned, targeting potentially shallower mineralization, which contingent on results, may later be added to the Central Zone mine plan and schedule.


Note - only diamond drillholes of sufficient depth to test for mineralization are shown

Figure 8.1 : TMI with Mineralized Zones and Diamond Drillholes

 


9

Exploration

Exploration work carried out by Cardero to date is outlined in Table 9.1.

Table 9.1 : Exploration Work by Cardero

Technique

Scope

Geological mapping

Mapping of limited outcrop and float

Electromagnetic Survey

Transient Electromagnetic

Ground magnetic survey 2004

Central & South grids 119.5 line kilometres

Ground magnetic survey 2008

Three infill grids 162.3 line kilometres

magnetic modelling

3D Inversion models

Diamond Drilling

13 drillholes for 4,211.6 metres, max. depth 800 metres

9.1

Transient Electromagnetic Survey

Transient Electromagnetic (TEM) surveying was initially conducted targeting potential copper sulphide mineralization. It is not relevant to exploration for iron (magnetite) resources and is not further discussed here.

9.2

Ground Magnetic Surveys

Exploration and drilling for magnetite iron mineralization has been driven by ground magnetic surveys, which were designed to increase the resolution provided by the historical Rio Tinto helicopter regional magnetic survey.

Cardero initiated a detailed survey in 2004, covering the areas now termed the Central and South Zones. In 2008, a second survey was initiated to infill the remainder of the prospective ground, screening for additional magnetic anomalies. Data for both surveys was collected by Quantec Geoscience Peru S.A.C. using two GEM Systems - GSM-19 Overhauser magnetometers. The GSM-19 magnetometers measure the magnetic field of the earth to an accuracy of 0.01 nano-Tesla (nT).

The combined surveys covered 281.7 line kilometres and the resulting Total Magnetic Intensity (TMI) map is included in the previous section, Figure 8.1.

9.3

Magnetic Inversion Models

An inversion is a mathematical procedure, which utilizes measurements taken at surface, generating 3D sub-surface models based on certain assumptions. The inversion models characterize how the relevant physical property, in this case magnetic susceptibility, will be distributed in three dimensions below the ground surface. At Pampa de Pongo, the measurements being utilized are the surface magnetic data. The assumptions were based on hard information gathered from Rio Tinto’s previous drilling, which had intersected massive magnetite at known depths below surface. In the case of Pampa de Pongo, there is a very large geophysical contrast between the magnetic susceptibility of massive magnetite mineralization and unmineralized rock.

The magnetic data was inverted by Quantec Geoscience Peru S.A.C. using a 3D inversion package (MAG3D, version 3.1) supplied by the University of British Columbia, Canada. The 3D inversion returns a model of the subsurface magnetic susceptibility in SI units.

[techrep013.jpg]

Note only drillholes of sufficient depth to intersect mineralization are shown

Figure 9.1 : 600m Horizontal Depth Slice through 3D Magnetic Model

9.3.1

Correlation of Inversions with Mineralization

In Figure 8.1 the Central and South Zones do not appear to have a perfect correlation with Total Magnetic Intensity (TMI) map. This is because the TMI data was collected at surface, whereas the Central and South Zones are located at differing depths and have significantly different thicknesses of high grade iron mineralization, which all lead to differences in magnetic response.

In order to reveal a better correlation, the sub-surface mineralization must be overlain on ‘slices’ cut horizontally through the 3D Inversion model. The image in Figure 9.1 is a horizontal slice through the inversion model at a depth of 600 metres below surface and demonstrates excellent correlation between the Central Zone resource and the inversion model. This model also includes the area of the East Zone exploration target, where by analogy, magnetite mineralization is interpreted to be located within the area of the dashed line. Drillhole DDH-004 tested the edge of this anomaly and intersected a continuous 292 metres of semi-massive to massive magnetite mineralization.

Comparison of drill intercepts obtained to date indicate a high correlation between modeled magnetic anomalies using MAG3D and magnetite mineralization intersected in drillholes. As a result there is a good probability of intersecting further magnetite mineralization in the East and North Zones.

At Pampa de Pongo, magnetic inversions work extremely well imaging magnetic anomalies from surface to the base of mineralization. As with many mathematical models, the model is not resolved accurately at depth. Therefore, the 3D models appear to continue at depth, beyond the known limits of magnetite mineralization (Figure 9.2). For SRK’s resource estimate solids were terminated at the depth of drilling marked by the dashed line for the Central Zone. For conceptual tonnage estimate for the East Zone the depth extent of the average of the range in reported tonnage roughly approximates the dashed line in Figure 9.2.

Figure 9.2: 3D Magnetic Model (note Central Zone and East Zone only) looking East


10

Drilling

The 2004 Cardero drill program was primarily targeted towards copper-gold discovery at relatively shallow depth. The majority of drillholes were not drilled to sufficient depth to test what is now known to be a significant iron resource. All drillhole locations are included in Figure 10.1 with results summarized in Table 10.1.

All drillholes were collared in HQ-size core, reducing to NQ at a depth of approximately 300-400 metres from surface and reducing further to BQ as dictated by ground and drilling conditions.

10.1

Iron Intersections

Drillhole DDH04-PP019 (South Zone) and drillholes DDH04-PP020A and DDH04-PP021 (Central Zone) were drilled to test the 3D magnetic anomaly. All three drillholes intersected massive magnetite mineralization as predicted by the 3D magnetic model. The best mineralization was in drillhole DDH04-PP021, which intersected 303 metres at 51.2% iron. The focus of this report is the iron resource in Central Zone, consequently only iron intersections from Central Zone are described here.

Drillholes DDH04-PP020A and DDH04-PP021 tested the central portion of Central Zone (Figures 10.2 and 10.3). Drillhole DDH04-PP020A intersected a continuous interval of semi-massive to massive magnetite mineralization from 314.75 metres to 626 metres (311.25 metres total). Drillhole DDH04-PP021 intersected continuous semi-massive to massive mineralization in a 303-metre interval grading 51.2% iron.

Drillholes DDH04-PP020A and DDH04-PP021 both intersected significant iron mineralization in the zones overlying the massive mineralization as outlined above. These intersections are comprised primarily of massive mineralization, which has been intruded by unmineralized hypabyssal sills. The best mineralization intersected to date in the upper portion of the system was in drillhole DDH04-PP021 returning 347 metres at 22.5% iron.


Table 10.1 : Cardero Drilling Highlights

Drillhole

Target

Comment

From (m)

To

(m)

Interval (m)

Iron Grade

(%)

DDH04-10

EM Conductor - NW flank of the South Zone

Massive Fe

220.8

270

49.2

38.8

DDH04-11

Extension of EM conductor, 500 m South of DDH04-10

Abandoned

    

DDH04-12

EM Conductor West of South Zone

Negative

    

DDH04-13

South edge of South Zone

Strong alteration

    

DDH04-14

EM conductor - western flank of the Central Zone

Semi-massive Fe

133.4

154

20.6

30.3

DDH04-15

Major structural intersection associated with a weak magnetic anomaly

Negative

    

DDH04-16

Surface copper showings on NNE-trending structure.

Negative

    

DDH04-17

East edge of South Zone

Abandoned

    

DDH04-18

Central part of South Zone

Abandoned

    

DDH04-19

Central part of South Zone. Test of 3D magnetic model.

Massive Fe

212.5

400

187.5

36.2

DDH04-20A

Central Zone. Test of 3D magnetic model.

Massive Fe

308

600

292.0

47.4

DDH04-21

Central Zone. Test of 3D magnetic model.

Massive Fe

482

784

302.0

51.3

10.2

Copper-Gold Intersections

Exploration to date has identified several significant copper-gold intersections. Drillhole DDH04-PP010 intersected 49.2 metres at 0.32% copper and 0.34 g/t gold. Drillhole DDH04-PP016 intersected fault breccias over 33.2 metres (from 154.3 to 187.5 metres), within which 19.3 metres (from 154.3 to 173.6 metres) returned 0.88 g/t gold. In addition, Rio Tinto’s reverse circulation drillhole, PRC-002, intersected 21 metres at 0.43% copper and 0.68 g/t gold.

Other broad, low-grade copper intersections drilled by Cardero include drillhole DDH04-PP019 (187.55 metres at 0.17% copper and 0.28 g/t gold), drillhole DDH04-PP020A (292.0 metres at 0.16% copper and 0.11 g/t gold) and drillhole DDH04-PP021 (302.0 metres at 0.10% copper and 0.06 g/t gold).

The potential for higher-grade copper-gold mineralization is not fully understood at present but the potential for future copper-gold discoveries has not been ruled out. Regardless, broad, low-grade copper-gold intersections have very significant value in the Pampa de Pongo preliminary economic assessment as by-product credits.


[techrep015.jpg]

Figure 10.1 : Drillhole Locations



[techrep017.jpg]

Figure 10.2 : Central Zone Composite Cross Section at 517400 E (looking east)

[techrep019.jpg]

Figure 10.3 : Central Zone Composite Cross Section at 8301600 N (looking north)



11

Sampling Method and Approach

The HQ and NQ core from both the Cardero and Rio Tinto drill programs was sawn in half and sampled. The HQ core size was downsized at approximately the 300-400 metres depth as required by drilling conditions. As part of SRK’s due diligence sit visit, it was observed that the remaining half core from both drill programs was appropriately stored at a secure maintenance yard belonging to Geotec. Drilling Ltd in Lima, Peru. Drill core for both programs had been sampled over the entire drillhole extents. This sampling protocol reflected the nature of semi-massive and massive magnetite encountered throughout the drilling. In both drill programs sample lengths averaged 2.0 m with sample lengths ranging from 0.2 to 3.7 metres. Sample intervals selected were based on the presence of magnetite or alteration.

Although Rio Tinto had completed both diamond drilling and reverse circulation, only diamond drill results were used for the estimation of the Pampa de Pongo mineral resources. The reverse circulation drillholes generally failed to achieve sufficient depth to penetrate the mineralization. The Cardero drill program was comprised entirely of diamond drill core.

Drillhole spacing for the entire combined Rio Tinto and Cardero database ranges from 190 m to 1500 m, with a majority in the 200-400 m range. The density of drilling of the Rio Tinto and Cardero drill programs are roughly proportional and both cover the same Central and South Zone areas. Only one hole penetrates the southern extent of the East Zone. The distance from the South Zone to the North Zone represents an interpreted strike extent of approximately 9 km at an azimuth of 326 degrees. This trend roughly follows the interpretation of a deep seated fault structure which controlled the emplacement of mineralization.

A description of the type, number and metres of samples collected are included in the following Table 11.1.

Table 11.1 : Phase of Drilling and Description of Sample Type

Operator

Year

Sample Type

Metres of Drilling

No of Drillholes

No of Samples

Metres of Sample

Rio Tinto

1994-1996

Diamond Drill Core

4,883.7

9

1896

3406.2

Rio Tinto

1994-1996

Reverse Circulation

1,401

6

412

1230

Cardero

2004-2005

Diamond Drill Core

4,596.6

14

886

1771.1

Core recovery averaged 92%. No sampling issues were identified that would have a material impact on the accuracy of results.

SRK believes that the samples are representative and that there are no known factors which may have resulted in sample bias.

 


12

Sample Preparation, Analyses and Security

12.1

Rio Tinto

It is assumed by SRK that Rio Tinto employees were responsible for the collection of samples during the Rio Tinto phase of drilling. It is unknown what sample preparation, assaying or security protocols were followed by Rio Tinto during the 1994-1996 phase of exploration. It would appear from available assay certificates that Bondar-Clegg Laboratories and Anamet Services Laboratories of Bristol, UK were used to assay the Rio Tinto samples. It is not clear whether the Bondar Clegg laboratory was certified at the time the Rio Tinto program was in operation. The Anamet Services laboratory at the time of assaying was part of the Rio Tinto Group. Assay certificates indicate that assay methods FA01, G02 ICP04 and XRF04 were used however no other details were available. The assay certificates also indicate that assay procedures were adopted which followed the National measurement Accreditation Service (NAMAS). During the Rio Tinto program, it is reported that the assay laboratories did undertake quality control measures.

12.2

Cardero

Collection of samples from the Cardero 2004-2005 drill core was completed by employees of Cardero under the supervision of G.D. Belik who acted as Cardero’s on site QP. No sample preparation was completed on site other than the sawing of core. Belik indicates that the logging of core and sawing of samples were completed in a secure core logging facility in the local village of Lomas. Samples were delivered directly to the ALS-Chemex laboratory in Lima, Peru in sealed containers.

Sample preparation and gold analysis were completed by ALS-Chemex Laboratories in Lima. Sample preparation was comprised of weighing, crushing to 70% passing 2mm, followed by splitting and pulverizing the split to 85% passing 75 micron. Pulps were then transported to ALS-Chemex Vancouver, B.C. for multi-element ICP analysis. Gold analysis by ALS-Chemex (Lima) was comprised of fire assay of a 30 gram pulp followed by an atomic absorption spectroscopy finish. For higher grade samples (>10,000ppb) fire assay with a gravimetric finish was used.

Analysis at ALS-Chemex varied over the Cardero drill program. Analysis was comprised of ALS-Chemex sample method ME-CP81 whereby the sample undergoes a sodium peroxide fusion at 650°C followed by acid dissolution of the pellet and is finished with Inductively Coupled Plasma-Atomic Emission Spectroscopy (ICP-AES) reading. ALS-Chemex sample method ME-ICP41 was also used for more weakly mineralized intervals. This method adopts an aqua regia leach and may not be an optimal method as it only represents the leachable portion of the analyte. Incomplete acid digestion may result in the underreporting of some elements.

ALS-Chemex has developed and implemented at each of its locations a Quality Management System designed to ensure the production of consistently reliable data. The system is audited both internally and externally. Most ALS-Chemex laboratories are registered or are pending registration to ISO9001:2000 and a number of analytical facilities have received ISO 17025 accreditations for specific laboratory procedures. The ALS-Chemex Lima Laboratory has the Bureau Veritas Quality International ISO 9001:2000 Certification and the INDECOPI 17025 Accreditation. The ALS-Chemex Vancouver Laboratory is in compliance with ISO 9001:2000 for the provision of assay services according to QMI Management Systems Registration. The laboratory has also been accredited to ISO 7025 standards for specific laboratory procedures by the Standards Council of Canada (SCC).

In SRK’s opinion the sample preparation, security and analytical procedures were appropriate.



13

Data Verification

Data verification was evaluated by SRK for the purposes of this Technical Report. Verification was comprised of checking 5% of the electronic database against assay certificates, evaluating Rio Tinto and ALS-Chemex’s internal quality control measures during the course of assaying and assessing the results of a total of 39 quarter core splits from both the RT and Cardero drill programs. No pulps or coarse rejects were available for independent check assaying for either the Rio Tinto or Cardero sample programs.

No material issues were noted with Rio Tinto’s QA/QC results. Rio Tinto’s QA/QC program was comprised of the insertion of blanks and standards and analysis of duplicate pulps by Rio Tinto personnel. Blanks were inserted at an average rate of 1 per 100 samples, while standards were inserted at a rate of 1 per 33 samples and duplicates were assayed at a frequency of 1 per 12 samples. Rio Tinto used a total of 10 different standards. Blanks and standards showed no indication of contamination. Duplicate results indicated no material issues with the assay results.

No issues were identified with the Cardero assay program ¼ core duplicates or the ALS-Chemex internal QA/QC results, however, ¼ core duplicates for the Rio Tinto program did indicate some issues surrounding the quality of the deleterious grades such as P, MgO, Mn, and the alkalis. As a result, the Rio Tinto MgO, Al2O3, Mn and alkali grade seemed to consistently under-report grade when compared to the duplicate quartered core. It is suspected that the assay method used by Rio Tinto was likely an ICP assay method which suffered from incomplete acid digestion. As well, P values in the original Rio Tinto dataset were significantly higher than the ¼ core duplicates selected by SRK suggesting the potential for laboratory error. This finding was also supported by subsequent metallurgical testwork that showed that higher P values from Rio Tinto assaying were not replicated As a result of the analysis, SRK elected to not use the MgO, Mn, Al2O3, P, K2O, Na2O and TiO2 values from Rio Tinto for resource estimation. The Fe, S, SiO2, CaO, Au and Cu grades were unaffected and showed good reproducibility considering the samples were comparing duplicate quarter splits of core rather than duplicate pulps.

Results from the Fe, S, SiO2, CaO Au and Cu ¼ core duplicates are included below. The lower correlation of Cu and Au are interpreted to be acceptable because of the more variable grade distribution of copper and gold and typical variance expected with ¼ core duplicates.

[techrep021.jpg]

Figure 13.1 : Fe Quarter Core Duplicates

[techrep023.jpg]

Figure 13.2 : SiO2 Quarter Core Duplicates


[techrep025.jpg]

Figure 13.3 : S Quarter Core Duplicates

[techrep027.jpg]

Figure 13.4 : Au Quarter Core Duplicates

[techrep029.jpg]

Figure 13.5 : Cu Quarter Core Duplicates

ALS-Chemex provided detailed internal QA/QC results in PDF format as signed certificates for the period covering the Cardero assay program in 2004 and 2005. The internal QA/QC results were comprised of inserted blanks and standards as well as duplicates. All of the QA/QC data was reviewed in light of the minimum and maximum thresholds established by the laboratory. Apart from a few minor discrepancies at near detection limits, no systematic error was found and no discrepancies were identified.

The graphs in Figures 13.6 and 13.7 highlight good correlation between ALS-Chemex’s XRF to ICP analysis methods for 39 duplicates selected by SRK.

[techrep031.jpg][techrep033.jpg]

Figure 13.6 : P and Mn ALS-Chemex Internal Pulp Duplicates

[techrep035.jpg][techrep037.jpg]

[techrep039.jpg][techrep041.jpg]

[techrep043.jpg]

Figure 13.7 : SiO2, Al2O3, MgO, TiO2 and CaO ALS-Chemex Internal Pulp Duplicates

The QA/QC results indicated that no significant issues were encountered during verification other than some of the deleterious grades from Rio Tinto which were excluded from resource estimation. It is recommended that these holes be re-assayed in order to build an appropriate deleterious database. It is also recommended that for future drill programs, Cardero insert and monitor their own independent QA/QC program comprised of the insertion of blanks, duplicates and standards.

Cardero also undertook a verification of drillhole collar locations by resurveying all of the existing collars using differential GPS. Some corrections may have affected the reliability of the previous resource estimate.

A suite of 20 density samples were sent to Inspectorate Laboratories to check the onsite bulk density determinations. Figure 13.8 indicates that the results are well replicated by an independent laboratory.  Both Cardero and Inspectorate Laboratories used a water displacement method to determine SG.

[techrep045.jpg]

Figure 13.8 : Comparison of SG values of Inspectorate Laboratories vs. Field Values




14

Adjacent Properties

In this report there are no references to mineralization on adjacent properties. There are several references to the Marcona Mine, located 38 km to the northwest of Pampa de Pongo. This deposit is unrelated to Pampa de Pongo and does not have any known influence over any of the Pampa de Pongo mineralization discussed herein.

Marcona is referred to because it is an analogous deposit type, in terms of mineralization, resource size and proposed end-product iron oxide pellets. All information referred to is published and sourced from a technical paper (Hawkes et. al. 2002) or the Shougang Hierro Peru website (www.shp.com.pe/).  SRK has not been able to verify that the information regarding the Marcona Mine is necessarily indicative of the mineralization on the Pampa de Pongo property.


 


15

Mineral Processing and Metallurgical Testing

15.1

Mineral Processing

15.1.1

Beneficiation

The Pampa de Pongo concentrator will process approximately 27 million tonnes per year of mineralized rock (approximately 75,000 tonnes per day) at 40-45% average iron grade that will be used to produce 15 million tonnes per year of iron ore pellets.

Equipment sizing is based on information published from the Cerro Verde mine (Vanderbeek et. Al 2006), built in Peru and commissioned in January 2007. The Cerro Verde concentrator has 1/3 higher throughput than needed for the Pampa de Pongo concentrator and processes 108,000 mtpd of crude ore having a 15.3 kWhr/t Bond work index. In comparison, the Pampa de Pongo concentrator will process about the same amount or 75,000 mtpd of ROM rock with a somewhat higher Bond work index of 17.4 kWhr/t (measured 15.8 kWhr/st by NRRI, June 2008).

Primary (Gyratory) Crushing

Run of mine (ROM) mineralized rock will be fed to a 60-inch by 113-inch primary gyratory crusher with approximately a 1000 – 1200 hp motor that will be installed in a conventional fixed structure with a 500 tonne capacity dump pocket. Crushed ROM will be reclaimed from the primary crusher product surge pocket by three apron feeders per conveyor, installed in a separate concrete reclaim tunnels beneath the stockpile and hence to the coarse rock conveyor. The coarse rock conveyor will discharge to a ROM stockpile with 45,000 tonne live capacity, 190,000 tonne dead capacity and 300,000 tonne ultimate capacity (to be pushed via a bulldozer). The ROM conveyor will consist of two conveyors in series to allow for metal detection and removal without shutting down the entire secondary crushing plant.

Grinding

Ore will be reclaimed from the ROM stockpile by three apron feeders per conveyor, feeding two parallel SAG feed conveyors. The apron feeders and SAG feed conveyors are located in separate concrete tunnels installed below the ROM stockpile. Each of the two, parallel SAG feed conveyors in turn feeds one 40-by-22 foot SAG mill powered by a 22 MW gearless drive. Each SAG mill discharges onto a single 3.0-by-7.3 metre double deck, low head, banana type screen. SAG screen undersize material is combined with the discharge from two ball mills in a common sump and is pumped with two operating 28-by-26 inch variable speed pumps to separate ball mill cyclone clusters. Each SAG mill feeds two ball mills. Cyclone underflow from each cyclone cluster reports by gravity to a 24-by-35 foot ball mill powered by a 12 MW gearless drive. Cyclone overflow reports to rougher magnetic separation. SAG screen oversize is conveyed to the pebble crushing stockpile with 5,000 tonne live capacity (approximately four hours) and 46,000 tonne total capacity. Pebbles are reclaimed from beneath the pebble stockpile by three parallel conveyors installed in separate tunnels. The reclaimed pebbles feed three mP-1000 cone crushers. Pebble crusher product is conveyed back to the SAG feed conveyors with a common conveyor and proportioned to each SAG mill as desired.

[techrep047.jpg]

Figure 15.1 : Proposed SABC Circuit Process Flowsheet

Magnetic Separation

A typical magnetic separator concentrate line can produce about 1 million tons per year of concentrate. Therefore, producing the 15 million tonnes per year of concentrate will require 15 lines. Each of the fifteen concentrator lines will start with five 48-inch diameter by 10-foot long rougher wet magnetic separators. Rougher concentrate from the five units will feed two finisher magnetic separators of the same dimensions as the roughers.

Sulphur Removal

The magnetite concentrate contains 1.5% sulphur. During pilot-scale metallurgical work (detailed below), sulphur was not removed by flotation prior to pelletizing. However, sulphur levels were not found to be an issue for final pellet quality. Most or all of the sulphur was released during induration.

Potential overheating of pellets during induration due to sulphur oxidation is a possible future concern and sulphur may therefore need to be removed for metallurgical reasons.

Sulphur removal by flotation is part of the proposed beneficiation flowsheet (Figure 15.1). A sulphide froth flotation circuit will float off sulphide minerals, concentrating the magnetite in the underflow. This sulphide product will contain payable by-product copper and gold. The process flow sheet and equipment specifications are yet to be determined. However, based on typical copper residence time for a low-grade feed, and a total throughput of 27 million tonnes per annum, a flotation circuit of approximately 2500 cubic metres (total float capacity) would be anticipated. This float circuit would be based upon large Tank cells for the roughers with the cleaner circuits being based upon Tank, or conventional, cells depending upon the concentration ratio. Concentrates and tailings will be thickened for water reclamation and re-cycle, while the concentrates will be filtered on conventional filters to generate a copper-gold concentrate for marketing. Laboratory testing is underway and preliminary results suggest that a copper concentrate will be achievable.

Magnetic Concentrate

Finisher magnetic concentrate from each concentrator line will be pumped to one of two concentrate slurry thickeners to increase slurry density to required levels of approximately 60% solids. The overflow from the two thickeners will be recycled as concentrator process water. The underflow from each of the two thickeners will pass through a demagnetizing coil then travel several kilometres through a slurry pipeline from the concentrator to the pellet plant concentrate slurry tank. This slurry pipeline will therefore handle 15 million tonnes per year (mtpy) of concentrate at 60% solids. Flux addition will occur at the concentrate slurry tank and will come from a separate flux supply system. Filtering aids can be dosed into two concentrate slurry tanks. The concentrate slurry tank underflow will be split into two separate pipelines where each portion feeds a distributor to distribute feed to up to ten (thirty total) vacuum disk filters. The filtrate from the vacuum disk filters will be recycled via a second pipeline back to the concentrator’s slurry thickener to recover any concentrate that reports to the filtrate. This secondary pipeline will need to handle approximately 9 mtpy of water.

Tailings

Both the rougher and finisher magnetic separator tailings will be processed with a sulphide froth flotation circuit to concentrate the copper and gold (overflow froth product) that will be vacuum filtered to produce a copper-gold concentrate. The tailings from the copper / gold separation will be sent to a clarifier where the underflow will be sent to a tailings basin and the overflow water will be recycled to the mill.

15.1.2

Pelletization

Balling

Vacuum disk filters will be controlled to produce an iron ore concentrate that contains an appropriate level of moisture (approximately 7-9%) to produce acceptable quality green balls. The concentrate will be distributed by a bin supply system to any of twenty-four balling lines. Each balling line will consist of a filtered concentrate day bin each with a table feeder and a 12-foot diameter by 32-foot long balling drum equipped with roll screens. Bentonite (dosed at approximately 7 kilograms per tonne of concentrate) and/or an alternate binder will be added to the moist concentrate prior to balling. Roll screen undersize (“seeds”) will be recycled to the drum until they achieve product size (approximately three passes) and oversize will be pulverized and returned to the balling drum.

Induration

Four of the twenty-four balling lines will each feed one of six four-metre wide traveling grate induration machines that will each produce 2.5 million tonnes per year of indurated iron ore pellets. A roll feeder will distribute wet green pellets at an even depth across the grates and also allow for wet green pellet fines to be removed and recycled to the concentrate bin supply. In terms of fuel consumption, approximately 320,000 BTU of natural gas per metric ton of pellets will be consumed during induration (heat hardening). In terms of water consumption, approximately 91.5 gallons per minute wet wall electrostatic precipitator evaporative losses are anticipated.

[techrep049.jpg]

Figure .: Proposed PDP Straight-Grate Induration Circuit Process Flowsheet

15.2

Metallurgical Testing

The purpose of the pilot-scale pelletizing investigation was to demonstrate that the Pampa de Pongo mineralized rocks are suitable for production of a quality Blast Furnace (BF) grade and premium Direct Reduction (DR) grade pellet. DR pellets are premium products and typically attract higher value contracts than BF grade pellets. The price premium is typically 10% in terms of contract price and most industry analysts agree that this trend should continue in the long term. BF and DR metallurgy was undertaken at two separate laboratories.

15.2.1

Testing Facilities

The initial metallurgical testing was undertaken by the Natural Resources Research Institute (NRRI) in Minnesota, United States. NRRI was selected because their personnel bring previous experience beneficiating similar iron ore from the Marcona Mine, located approximately 35 kilometres to the northwest of Pampa de Pongo, and they have vast experience beneficiating and pelletizing iron ores throughout the world, including the magnetic taconites in Michigan and Minnesota in the United States and in Canada. NRRI are generally considered to be industry leaders in this type of test work.

The Direct Reduction material evaluation test work was undertaken by MIDREX® Technologies Inc., North Carolina, USA. MIDREX®  is the world leader in DRI technology, having built more than 60 modules in 19 countries, and has become the leading process technology for producing Direct Reduced Iron (DRI) with more than 60% of the world’s DRI produced using MIDREX® Technology.

15.2.2

Sample Selection

The sample for metallurgical testing was selected from existing drill core from within the Central Zone, which is the focus of this study. SRK determined the length-weighted mean grade of the zone and recommended sample intervals which would comprise a representative sample. Samples were taken from 4 drillholes (DDH-001, -003, -020A, and -021), totalling 359 metres of mineralized quartered drill core. The sample was shipped from Peru as two batches, each representative of the determined mean grade of the central zone. The first sample (producing 450 kilograms of concentrate) was used for initial bench-scale testing. The unused material was added to the second batch. The total weight of the sample was 1,090 kilograms.

15.3

NRRI Metallurgical Testing

Beneficiation, magnetic concentration and pilot-scale pelletizing was undertaken at NRRI.

15.3.1

Metallurgical Test Results

The initial step in the metallurgical testing process was to produce a suitable grade iron ore concentrate from drill core samples then subsequently produce iron ore pellets for further metallurgical testing. Magnetic separation produced a 65.5% iron concentrate with a total iron recovery of 93.4%. This was achieved with only a simple two stage (rougher-cleaner) wet magnetic separation.

Bench scale testing was used to establish conditions for pilot plant operation. Batch balling and mini-pot induration testing were used to provide a basis for pilot-scale full pot grate tests. Mini-pot test results are a good indicator of physical BF pellet quality and provide guidance for designing the full pot grate test plan. Full pot grate tests represent the conditions in a commercial plant and the physical, chemical, and metallurgical results are scalable to commercial operation; full pot grate tests also provide the opportunity to correlate pellet quality with the furnace firing cycle for quality optimization. It was noted that the Pampa de Pongo concentrate balled very well, meaning good growth rate and moisture control at all levels of additive. Since the amount of concentrate was relatively limited, a generic induration furnace temperature cycle was selected; however it was not completely developed to provide optimum pellet quality and NRRI fully expects further improvements would result from additional metallurgical test work.

Four full pot grate samples were fired, with the initial test being used to provide a baseline firing cycle for the subsequent three. The resulting fired pilot-scale pellet quality averaged 421 pounds compression strength and 94.3% of material sizing >1/4 inch after the tumble test. The metallurgical quality shows reducibility (0.94%/min.), swelling (10.8%) and porosity (27.0%) that all meet industry standards for high quality blast furnace feed.

Because concentrate quantity was limited, further development of the pot grate cycle was stopped after the fourth pot grate test. According to the NRRI report, if sufficient material had been available (from a larger sample of drill core), the preheat cycle time would have been modified and significantly higher pellet quality could have been obtained.

15.3.2

Blast Furnace End-User Considerations

Blast furnace operations generally source pellets globally through a network of purchasers. Operations are typically seeking the best pellet chemistry at the least cost to meet their needs and it is important that pellet chemistry be highly consistent. There are a number of specific needs that are relevant to the Pampa de Pongo pellet chemistry.

Deleterious Elements

Deleterious elements in iron ores and concentrates include silica, alumina, manganese, phosphorous, alkalis (such as sodium and potassium), and sulphur. All iron ore pellet producers have to contend with at least one of these elements, with almost all having a primary concern to reduce the silica content.

  • Silica is the deleterious element that a beneficiation plant is usually designed to reduce. Modern plants employ additional ore grinding and froth flotation procedures so as to reduce silica. Pampa de Pongo concentrate contains some of the lowest silica available within the industry, and achieves this without additional beneficiation steps. Typical commercial iron ore pellets have silica ranging from 2.7 to 5.4 %.
  • Pelletizing of Pampa de Pongo concentrate produced pellets with significantly lower silica levels of 1.65% (bench-scale) to 2.23% (pilot-scale). Bench-scale testing demonstrated that with limited additional grinding, the silica in the concentrate could be reduced even further to as low as 0.34% at a reasonable particle size of 60% minus 200 mesh (75 um) with only a simple two stage wet magnetic separation. The low-silica Pampa de Pongo pellet can allow a blast furnace operation to use higher silica (poorer quality) pellets or a higher proportion of their standard silica pellets without supplementing additional flux reagents.
  • Phosphorous content in iron at concentrations greater than 0.2% makes the product cold short (or brittle at low temperatures). In steel making, the product becomes brittle even with only 0.5% phosphorous. The phosphorous cannot be easily removed by fluxing, so it is critical that the iron ore be low in phosphorous. The Pampa de Pongo pellets contain only 0.004% phosphorous and should be highly desirable to most end-users.
  • Sulphur is a frequent contaminant in coal and sometimes in iron ore. Sulphur causes iron to hot short (or become brittle when hot). This means that the iron must be worked at lower temperature, which requires more energy. Commercial iron ore pellets should contain less than 0.05% sulphur. The Pampa de Pongo pilot-scale pellets contain only 0.008% sulphur and are a premium product in that respect.
  • Alumina is difficult to reduce so contamination of iron in the blast furnace is not a serious problem. However, aluminium does increase the viscosity of the slag, which can create difficulties in operation of the furnace. Viscous slag slows the descent of the charge in the furnace and prolongs the process, consuming more energy in the blast furnace. Al2O3 should typically be less than 2.5% and Pampa de Pongo pellets contain 1% Al2O3.

Fluxed Pellets

Fluxed iron ore pellets are made by adding magnesium and/or calcium in the form of limestone and dolomite to the pelletizing feed mix. The lime and /or dolomite additions are made to improve blast furnace operation, hot metal sulphur control and furnace refractory life.

Pellet producers are forced to pay extra for dolomite or limestone to produce fluxed pellets; therefore fluxed (basic) pellets are more costly to produce than standard (acid) pellets. Magnesium is naturally occurring in the Pampa de Pongo concentrate and behaves in a similar way to calcium in that it has a high affinity for sulphur at steelmaking temperatures and can also be beneficial in mitigating erosion of the magnesite steelmaking furnace refractory by liquid slag.

At lower induration furnace temperatures (<1300 o C), calcium has a higher affinity for sulphur than magnesium. Therefore, a high calcium pellet can trap sulphur and pass it along into the product pellet. However, a high magnesium / low calcium pellet will allow the sulphur contained in the iron ore concentrate to pass along with the furnace gases thus yielding a low sulphur iron ore pellet. This is important for blast furnace pellets, but particularly important for the higher premium Direct Reduction (DR) grade pellets since sulphur in these pellets can poison the catalyst required for the process and sulphur embrittles steel.

The magnesium level within the Pampa de Pongo concentrate will be viewed by most blast furnace operations as being a very desirable feature. Pampa de Pongo pellets can also be viewed as a premium blending feedstock material as it promotes the use of cheaper acid pellets (low ratio of percentages of [CaO+MgO]/[SiO2+Al2O3]) so that an optimum low cost feed to the blast furnace can be determined. Typical fluxed pellets contain 1.2 to 1.8% MgO. The Pampa de Pongo pellets contain a more favourable 3.6% MgO.

In addition, if a steelmaker so desires, the Pampa de Pongo pellets can be customized to produce designer pellets by controlling the silica level (from as low as 0.34% to 2.3% or higher) by simply controlling the amount of grinding of the concentrate and/or creating a pellet recipe and adding other additives or fluxes to the concentrate to provide the desired chemistry.

15.3.3

Chemical Analysis and Quality Control

A representative sample from each of the full pot grate fired pellets was tested to determine their physical qualities and characteristics. Based on these results, the remaining representative pellets from pot grate tests 2 through 4 were combined into one bulk representative sample and chemistry was determined from a representative sub-sample. Analysis from the final Pampa de Pongo pellets is presented in Table 15.1 below and compared to typical Blast Furnace chemical composition, where such estimations can be reasonably made. All of the analysis was undertaken at the NRRI laboratory. NRRI follow international (ISO) and North American (ASTM) procedures where such procedures exist for highly specialized iron pelletizing work.

In order to ensure quality control, a duplicate check sample was forwarded to ALS-Chemex in Vancouver for ISO-certified XRF analysis. Results from ALS-Chemex are also presented in Table 15.1 and were all found to be within acceptable limits of accuracy and precision.

Note that the iron grade was determined by NRRI by the wet chemical iron titration method, which is used for commercial trading of iron ore pellets and is more precise than the XRF determination after lithium metaborate fusion, as employed by ALS. For all other elements, the method of analysis was comparable.

Table 15.1 : BF Grade Pellet Comparison and QA/QC

Parameter

Units

Typical Blast Furnace Pellet

Pampa de Pongo Pellet

ALS Chemex QA/QC

Tumble test

% +1/4 inch

>95

94.3

 

Compression

lb

>400

421

 

Total Fe

%

>65

64.5

64.1

SiO2

%

<4.0

2.23

2.23

Al2O3

%

*

1.00

1.07

CaO

%

*

0.59

0.64

MgO

%

*

3.45

3.66

P

%

<0.02

0.004

0.01

S

%

<0.05

0.008

0.01

TiO2

%

*

0.075

0.08

Na2O

%

<0.03

0.092

0.08

K2O

%

<0.03

0.041

0.06

Porosity

%

>26.0

27

 

LTD

% +6.3mm

>90.0

81.3

 

R40

%/min

>0.90

0.94

 

Swelling

%

<15.0

10.8

 

* Fluxing compounds cannot be listed as typical - they are dependent on Blast Furnace specification.

15.3.4

Future Work

Further development of the furnace firing cycle, utilizing the mineralogy and chemistry, could be used to further optimize and enhance blast furnace pellet quality. Although Met-Chem previously undertook bench-scale metallurgical testing (on behalf of Rio Tinto), it was limited to mini pot testing only. The current study, conducted by NRRI, was the first attempt to produce commercially representative pilot-scale pellets from Pampa de Pongo iron concentrate. The resulting pellet quality is very good and there remains room for optimization.

NRRI commented that “preheat cycle time would have been modified and significantly higher pellet quality could have been obtained. It is believed that modification of the furnace cycle in preheat would have improved LTD (81.3% +6.3mm) that is known to be impacted by preheat ramp and production rate. Further development of the furnace firing cycle, utilizing the mineralogy and chemistry can also be used to further optimize and enhance pellet quality.”

A future metallurgical step would be to select and engineer commercial induration equipment and optimize firing cycles and iron ore pellet quality. This entails conducting a series of up to thirty, pilot-scale, full pot-grate tests to optimize pellet quality. This type of optimization is typically undertaken during a more advanced mine Feasibility Study.

15.4

MIDREX® Metallurgical Testing

Direct Reduced Iron (DRI) is a high-quality iron product created through the removal of oxygen (reduction) from iron oxide material, in solid state (without melting). Reduction is achieved when a reducing gas (typically hydrogen and carbon monoxide) is passed through the iron oxide feed material at a temperature between 760º and 970ºC. Direct Reduction Iron is commonly used in an Electric Arc Furnace (EAF) to produce steel, instead of, or in combination with, scrap iron. The DRI process requires specific physical and chemical quality feedstock, which must remain consistent as a product. Only a small percentage of pellets produced globally can meet the stringent requirements.

15.4.1

Material Evaluation Test Results

MIDREX® Technologies Inc. determined the suitability of the Pampa de Pongo pellets (as produced at NRRI) as Direct Reduction (DR) process feed, through industry standard Linder and Hot Load Tests.

Linder Test

The purpose of the Linder Test is to determine the iron oxide pellet’s susceptibility to generate unwanted fines during the reduction step accompanied by bed movement in the shaft furnace. Potential for fines generation is due to furnace rotation (simulating bed movement in the shaft furnace) and carburization of the iron oxide material during reduction (which can contribute to fines generation). Fines are defined as less than 3.36 mm material (6 mesh).

The Linder Test was performed on uncoated pellets. Results were positive, with more than 94% metallization and approximately 1.8% carbon in the DRI product. Fragmentation was low, at less than 1.5%.

Table 15.2 : Linder Test Results

Linder Test

Typical DRI

Uncoated

760ºC

% Metallization

>93.0

94.4

% Carbon

 

1.79

Fragmentation %< 3.36 mm

<2.0

1.4

Compression (kg)

>67.0

38

Metallization and fragmentation exceeded typical DRI product. Compression was lower than optimal. However, MIDREX® commented that “these were laboratory-made pellets, so the strength values are not indicative of a commercially-made pellet”.

Hot Load Test

The Hot Load Test simulates the increasing load experienced by the pellets as they descend through a shaft furnace. The test focuses on the physical durability of the pellet during reduction. Both uncoated and lime-coated pellets were tested and both performed well. Metallization was approximately 99%. The DRI product demonstrated reasonable strength as is typical for pilot-scale pellets. Critically, clustering did not occur with either uncoated (816ºC) or coated (927ºC) pellets, which is extremely positive.

Table 15.3 : Hot Load Test Results

Hot Load Test

Typical DRI

815ºC

Uncoated

815ºC

Coated

927ºC

% Metallization

>94.0

98.6

99.5

% <3.36mm after reduction

 

1.0

0.6

Tumble % >6.73mm

>90

85.8

93.4

Compression (kg)

>100

40

44

Clustering % >25mm

0

0

0

Metallization, fines generation and tumble results were all excellent. The lack of any clustering is outstanding and sets the Pampa de Pongo pellets apart as premium DR pellets. The lower compressive strength, as with the Linder Test, is not indicative of commercially-produced and optimized Pampa de Pongo pellets.

15.4.2

DRI End-User Considerations

There are a number of specific chemical needs in DRI operations that can be satisfied by the Pampa de Pongo DR grade pellet.



Fluxed DR Pellets

Fluxed DR-grade iron oxide pellets are made by adding magnesium and/or calcium (flux) in the form of limestone and dolomite to the pelletizing feed mix. Elevated magnesium oxide (MgO) in pellets is desirable to the steelmaker as it means less flux addition to the Electric Arc Furnace. Pellet producers are forced to pay extra for dolomite or limestone to produce fluxed pellets; therefore fluxed (basic) pellets are more costly to produce than standard (acid) pellets. On the other hand, the high MgO content in the Pampa de Pongo pellets may be viewed as leverage in price negotiations because the total iron content is lower than that of the highest quality DR grade pellets; therefore, a slight cost penalty may result. The Pampa de Pongo pellets contain elevated magnesium and are therefore saleable as self-fluxing DR grade pellets. Note that lime-addition to the pellet feed as flux should not be confused with lime-coating, the latter of which purely mitigates clustering of DRI product at high temperature.

Silica Content

Silica content of DRI should be as low as possible for steel-makers using an Electric Arc Furnace, preferably less than 2.0%. The Pampa de Pongo iron oxide pellets will meet this requirement, but further optimization of the processing can lower silica even further. Laboratory scale liberation grinding tests at NRRI have indicated that a concentrate can be produced with high-grade iron content (66.87% - 68.54%) and low silica content (0.18% - 0.57%) with a small amount of additional grinding. Low silica and the presence of MgO contribute to the zero-clustering observed in the Hot Load test undertaken at MIDREX®.

15.4.3

Detailed Results

The iron ore pellet sample sent from NRRI to MIDREX® was the product produced from three batches of full pot-grate pellets. MIDREX® conducted elemental analyses in triplicate. Sulphur and carbon are determined by LECO, while iron grade was determined by wet chemical titration, and remaining elements are determined by ICP.

Analysis and test results from the MIDREX® Material Evaluation are presented in detail (Table 15.4).

 

Table 15.4 : Material Evaluation Results

Test

Oxide Pellets as Received

Uncoated

Lime Coated

816ºC HL

+6 mesh

816ºC HL

-6 mesh

Linder

927 C HL

+6 mesh

928 C HL

-6 mesh *

Total Iron %

64.16

87.51

82.26

85.91

87.58

61.98

Metallic Fe %

-

84.51

79.81

81.07

87.12

58.57

Metallization %

-

96.6%

97.0%

94.4%

99.5%

94.5%

C

0.01

0.83

6.58

1.79

0.83

1.94

S

0.006

0.005

0.009

0.005

0.003

0.003

P

0.008

0.009

0.008

-

0.008

0.027

CaO

0.33

0.55

0.46

-

0.67

16.45

MgO

3.78

5.14

4.82

-

5.16

8.65

SiO2

2.14

2.88

2.84

-

3.30

7.55

Al2O3

1.19

1.56

1.48

-

1.6

1.73

TiO2

0.08

0.10

0.10

-

0.11

0.12

Gangue analysis total %

7.51

10.22

9.70

-

10.84

34.50

Actual total gangue %

8.4

10.8

10.4

10.8

11.4

35.0

gangue/Fe ratio

0.13

0.12

0.13

0.13

0.13

0.57

basicity

1.23

1.28

1.22

-

1.19

2.71

15.4.4

Summary

In its report, MIDREX® indicated that “The Cardero Pampa de Pongo DR pellets performed well in standard Midrex tests. Chemical and physical characteristics were comparable to commercially produced pellets that have been used successfully in the Midrex DR Process®. Based on these results, it is expected that commercially produced pellets made from this ore would be a suitable feedstock for the Midrex DR Process®”.

Maximum MIDREX® DR plant productivity is achieved by operating the shaft furnace with inlet gas temperatures ranging from 760-970°C (or higher). Lime coating of DR pellets is required when operating under higher temperature conditions, mitigating burden sticking and clustering. The results posted from the 927°C hot load test prove that the lime coated Pampa de Pongo pellets are a satisfactory feed material for a Midrex DR plant operating under the most severe conditions.

15.4.5

Future Work

Future work for DR grade pellets should be oriented at optimizing pellet quality. Areas where Pampa de Pongo pellet quality may be improved are:

1.

Reducing the pellet silica content is a high value quality improvement step. This can be accomplished by grinding the ore a little finer so as to target 0.34% silica content in the concentrate. Assuming that 0.66% bentonite addition is the binder, the resultant pellet silica content would fall from 2.14 to 0.67%, and hence the total iron content would increase from 64.16 to about 65.1%. Such an increase in pellet total iron content would be very well received by DR plant operators.

2.

Another method to further decrease the silica content could involve substituting the bentonite binder with an organic binder such as peridur. Additional grinding may also be employed with the use of peridur so as to further decrease the pellet silica content.

3.

Additional hot load testing of the Pampa de Pongo pellets at 927-980oC without the use of lime coating should be tested. The threshold temperature for clustering or sticking has yet to be determined, and if the Pampa de Pongo pellets can be processed in the hot load testing apparatus at these temperatures without lime coating, then this implies some cost savings.

4.

DRI which possesses high carbon content is in high demand, and the maximum carbon content for the Pampa de Pongo pellets has not been determined. Pellets produced from a finer grind concentrate for the purpose of decreasing the silica content may behave differently during carburization in the Linder test. The effect of concentrate grind size on carburization of the DRI product should be determined.


 


16

Mineral Resource and Mineral Reserve Estimates

16.1

Introduction

The primary objective of SRK’s work on the Pampa de Pongo’s deposit was to produce an independent, CIM compliant resource estimate for the deposit. This estimate supersedes a 2005 estimate for the deposit, produced by J.N. Helsen.

The SRK estimate was conducted on a dataset that combines both historical Rio Tinto data collected in the 1994-96 program and Cardero data collected in the 2004-05 drillhole program.

The principal components of SRK’s estimation work included:

  • Interpretation of the 3D magnetic inversion model;
  • Defining geological domains;
  • Construction of a 3D geological model in Datamine to constrain the data selection and interpolation processes;
  • Iron, gold, copper, mineral resource estimation by inverse squared distance methodology; and
  • Resource classification.

16.2

Geological Model

Four magnetite rich zones have been located within the Pampa de Pongo property: Central, East and two South Zones. Only the Central and two South Zones were solid modeled for resource estimation purposes. A range of tonnes and grade were estimated for the East Zone however due to the lack of drill density these were classified as conceptual.

The boundaries of the higher grade mineralization in the Central and South Zones were defined by relatively sharp drop offs in grade in both the hanging wall and footwall contacts. This hard boundary is roughly represented by a 20% Fe grade which drops to a 5-8% grade in the adjacent hanging wall and footwall as the intensity of replacement textures dramatically decreases. Within the mineralized zones a very continuous interval of >40% Fe grade was encountered. The main exception to the trend of a sharp drop off was the hanging wall of the Central Zone. In this area an area of highly mixed intervals of >5% Fe and <40% Fe overlies a >40% Fe interval of higher grade. A 20% Fe boundary was used to define solid boundaries for all zones in the solid modeling. A separate surface was created to separate the Central Zone high grade (>40% Fe) from the overlying mixed grade zone whose average grade met at least a 20% Fe threshold. SRK found this approach to solid modeling appropriate as the two areas were geologically distinct in that the overlying mixed grade zone contained a higher proportion of ocöite sills, and breccia zones as well as partial replacement textures. The underlying high grade zone was dominantly represented by massive magnetite replacement zones.

Quantec Geoscience Peru SA generated 3D magnetic models from ground magnetic surveys completed in 2008. This magnetic data was compared to vertical drillhole intercepts. A 0.6 SI susceptibility level correlated well with the drop off in Fe grades and was used to define the lateral extent of the Central Zone solids. Because the South Zone mineralization is considerably thinner than the Central Zone, a 0.45 SI susceptibility level was used to define the lateral extent of the South Zone solids. Figure 16.1 highlights the 0.6 SI magnetic susceptibility contour for the Central Zone located to the north and the East Zone located to the southeast.

[techrep051.jpg]

Figure 16.1 : 3D Magnetics and Drillhole Locations in the Central and East Zones

The East Zone is represented by an elongated magnetic anomaly similar to that encountered in the Central and South Zones. To date only one drillhole (PPD004) containing two intervals of high grade iron formation penetrates the southern edge of the East Zone. The relationship of magnetic intensity of the 3D magnetic inversion to thickness of the massive magnetite zone and associated high iron grades are reasonably well correlated in the Central and South Zone by drilling and were used to estimate the conceptual tonnes and grade range for the East Zone.

In the East Zone, in order to reflect the relative magnetic intensity to the high grade iron formation thickness, a 0.5 SI susceptibility contour was used to delimit the extents in plan along with approximately half the vertical thickness of the Central Zone to reflect the relative decrease in magnetic response between the two anomalies. The vertical thickness of the anomaly was adjusted by 50 m within the 0.5 SI susceptibility contour to generate two solids representing a minimum and maximum range of volume. This was then multiplied by the average density for mineralized iron formation. A range of conceptual grade was generated by averaging the two high grade intervals in drillhole PPD004 for an approximation of the higher grade limit and averaging both high grade intervals with the intervening low grade interval for an approximation of the lower grade range. The conceptual tonnes and grade are estimated to range between 350-500 million tonnes while the grade for the tonnage is estimated to range between 32% and 38% Fe.

Figure 16.2 shows diagrammatic shape and locations of the zones and Table 16.1 shows the Datamine domain coding, and description of the zones.

Figure 16.2 : Plan View of Zone Shapes and Locations (Central: Green, East: Red, South S1: Cyan, South S2: Purple)

16.2

Data Used in Resource Estimation

The data set used in the resource estimation of the Pampa de Pongo project consists of 1,277 Fe, Cu, and Au assays from six Rio Tinto and four Cardero drillholes (See Table 16.1). Based on high grade populations assessed from probability plots, the Au and Cu data were capped as presented in Table 16.2. Within the limits of the interpreted geologic solids, the assay data was composited to 2 m intervals, representing the most common raw sample length.

Table 16.1 : Modeled Domain Names, Location and Number of Drillholes

Estimation Domain

Zone Location

Number of Drillholes

C1_DWN

Central

5

C1_UP

Central

5

S1

South

3

S2

South

2

C2

East

1

Table 16.2 : Capped Au and Cu Assays

Zone

Au (g/t)

Cu (%)

Central

0.4

0.3

South – S1

0.6

0.5

South – S2

0.5

0.6

16.3

Statistical Analyses

Univariate statistics of declustered composite Fe, Au, and Cu grades are included in Figures 16.3 to 16.5. Overall, the Fe grades in the Central lower domain (C1_DWN) are higher than the Fe grades in the other domains. Both Au and Cu grades are generally higher in the S2 domain. The C1_Up zone represents part of the hanging wall mixed grade zone of the Central Zone and shows a broad range of grades corresponding to a larger coefficient of variation and lower median Fe grade. The C1_Up Zone extends to the overburden contact in two of the five Central Zone drillholes.

Basic statistics were also generated for deleterious elements in the Central Zone to provide an indication of potential economic quality. The results are based on assays from Cardero data. Table 16.3 indicates that there are no significant issues with deleterious substances.

Table 16.4 : Basic Statistics of Deleterious Substances in the Central Zone

Deleterious Substance

No. of data

Min

Max

Mean

Medium

StDev

CV

1st Quartile

3rd Quartile

Al2O3

439

0.07

16.65

3.53

1.12

4.22

1.20

0.91

5.22

MgO

439

1.29

28.00

8.68

7.37

4.49

0.52

5.59

10.40

Mn

439

0.02

0.26

0.11

0.10

0.04

0.34

0.08

0.13

P

439

0.00

0.49

0.04

0.04

0.04

0.95

0.01

0.06

SiO2

439

0.04

59.40

13.40

6.40

14.80

1.11

2.81

20.28

TiO2

439

0.01

1.55

0.21

0.05

0.28

1.39

0.03

0.30

S

439

0.02

21.90

2.45

2.42

2.03

0.83

1.60

2.96

[techrep053.jpg]

Figure 16.3: Statistics of De-clustered Composite Fe (%) Assays in the Four Mineralized Domains

[techrep055.jpg]

Figure 16.4 : Statistics of De-clustered Composite Au (g/t) Assays in the Four Mineralized Domains

Figure 16.5 : Statistics of De-clustered Composite Cu (%) Assays in the Four Mineralized Domains

16.1.4

Bulk Density Data

The bulk density (termed SG in the report) database is comprised of a set of 210 determinations, measured on-site by Cardero personnel. A small subset of 23 samples from the waste hanging wall was used to assign density value of 2.7 t/m3 to blocks outside of the mineralized domains.

There is high correlation between the Fe assay values and the SG data (see Figure 16.6). A formula was generated from the regression curve to assign SG values to all composites within the modelled domains.

[techrep057.jpg]

Figure 16.6 : Correlation between Fe and S.G. Values

16.5

Estimation Methodology

In order to appropriately weight the Fe grades by SG to account for the impact of density on Fe grade, SG values were used to interpolate Fe grades. Both the accumulation (SG x Fe %) and the SG were estimated by the inverse distance squared interpolation. The final block Fe estimates represent the ratio of accumulation and SG estimated values within each block. Au and Cu block grades were also estimated using inverse distance squared interpolation.

16.6

Block Models

The block model comprises blocks measuring 25 m x 25 m x 10 m in size aligned north-south and east-west. The basic block model geometry is summarised in Table 16.4.

Table 16.4 : Block Model Extents

Item

Easting

Northing

Elevation (masl)

Block Origin (m, UTM)

51,500

8,296,000

-400

Block Dimension (m)

25

25

10

Number of Blocks (each)

240

400

90




16.7

Estimation Parameters

Table 16.5 shows the estimation parameters used for estimating the blocks in all zones. The Datamine dynamic search routine was used for estimating the blocks in three passes. In the first pass, the blocks were estimated within a relatively small 400 m isotropic search ellipse. Progressively larger search ellipses were used to estimate grade in subsequent passes until all blocks in the interpreted solid were filled. A hard boundary was used to divide the grade interpolation for the lower grade C1UP and for the C1DWN portions of the Central Zone.

Table 16.5 : Estimation Parameters

Search Parameters

All Zones

Search volume shape

Sphere

Step I :search ellipse dimension

400,400,400

Step1: minimum and maximum sample

7,20

Step2: search ellipse dimension

800,800,800

Step2: minimum and maximum sample

7,20

Step3: search ellipse dimension

>800,>800,>800

Step3: minimum and maximum sample

7,20

Total samples used for estimation from each boreholes

7

16.8

Block Model Validation

The following validation exercises were carried out on the block model:

  • Comparison of local “well-informed” block accumulation and SG estimates with composites contained within those blocks.
  • Comparison of average drillhole accumulation and SG values with average block estimates along different directions – swath plots

16.8.1

Comparison of Block Estimates with Composites

Figure 16.8 shows how the “well-informed” block estimated accumulation (Fe∙SG) and block estimated SG compare with drillhole composites contained within those blocks. Each estimated block accumulation and SG were compared separately with drillhole average composite accumulation and SG within the block. Overall, average accumulation and SG values are almost identical and the correlation is very high. This is not surprising, considering large drillhole spacing and in turn large influence of drillhole assays on blocks intersected by the drillholes. The thick white line that runs through the middle of the cloud is the result of a piece-wise linear regression smoother.




[techrep059.jpg]

Figure 16.7 : Comparison of block estimates of accumulation (Fe*SG) and SG with drillhole composite accumulation and SG contained within the blocks

16.8.2

Swath Plots

The next check involved calculating polygonally de-clustered average composite Fe grades and comparing them with average block estimates along north-south and horizontal swaths. The average block estimates represent a ratio of accumulation and SG estimates. As shown in Table 16.8 the average composite Fe grades and the average estimated Fe grades are similar in both directions. This further supports the lack of any consistent spatial underestimation or overestimation of the block grades.

[techrep061.jpg]

Figure 16.8 : De-clustered Average Composite Grades Compared to Block Estimates in all Zones


16.9

Mineral Inventory

Mineral inventory represents all estimated blocks within the modeled zones. Because of potential for production at profit with underground mining methods there is no need in limiting the resources to an optimized Whittle shell.

Table 16.6 shows the mineral inventory for the Pampa de Pongo deposit at different Fe cut-off grades. The results of the estimates have been rounded to the nearest million tonnes to reflect the uncertainty in the estimation. Figure 16.9 shows an example of the resource shell outline and estimated block grades on the 8,301,687 E-W section.

Table 16.6 : Mineral Inventory at Different %Fe Cut-off Grades

ZONE

Cut-Off

Grade

Volume

(Mm3)

Density

(T/m3)

Tonnage

(Mt)

Fe

(%)

Au

(g/t)

Cu

(%)

CENTRAL

>30% Fe

163

3.80

618

45.1

0.061

0.098

>25% Fe

190

3.73

707

42.9

0.059

0.095

>20% Fe

200

3.70

739

42.1

0.060

0.094

>15% Fe

203

3.69

748

41.7

0.059

0.093

>10% Fe

204

3.68

752

41.6

0.059

0.093

> 5% Fe

205

3.68

753

41.6

0.059

0.093

> 0% Fe

205

3.68

753

41.6

0.059

0.093

SOUTH

>30% Fe

31

3.60

113

39.7

0.130

0.121

>25% Fe

32

3.59

115

39.5

0.130

0.121

>20% Fe

32

3.59

115

39.5

0.130

0.121

>15% Fe

32

3.59

115

39.5

0.130

0.121

>10% Fe

32

3.59

115

39.5

0.130

0.121

> 5% Fe

32

3.59

115

39.5

0.130

0.121

> 0% Fe

32

3.59

115

39.5

0.130

0.121

Total

> 15% Fe

235

3.68

863

41.4

0.068

0.097

16.10

 Mineral Resource Classification

The mineral resource for the Pampa de Pongo deposit has been classified entirely as inferred. The following factors contributed to the classification of the current resource estimate

  • Although assay data was largely validated by the QAQC analysis, much of the Rio Tinto deleterious assay database was found to be suspect and therefore not included in the resource estimate resulting in a lack of data to fully characterize potential economic quality;
  • Drillhole spacing of 250-500m;
  • Lateral extends of the zones are defined from “soft” geophysical data.
  • The classification of resources is based on the interpreted continuity of Fe grades only. The understanding of the geological controls on Au and Cu grade distribution are far less understood and therefore less confidence should be assigned to these grades.

[techrep063.jpg]

Figure 16.9 : Estimated Grades for 8301687 E-W Section

16.11

 Mineral Resource Statement

The mineral resources were estimated by Ebi Ghayem (P.Geo.) and reviewed by Marek Nowak (P.Eng.) and George Wahl (P.Geo.) who are all Qualified Persons. The effective date of the Mineral Resource Statement is September 30th, 2008. The mineral resources are not mineral reserves and do not have demonstrated economic viability. SRK is not aware of any environmental, permitting, legal, title, taxation, socio-political, marketing or other relevant issues that may affect the mineral resource estimate.

The classified inferred mineral resource estimates at 15% Fe cut-off grade are tabulated in Table 16.7. The resources represent all estimated blocks within the modeled zones. The economic cut-off used to generate mineral resources was assumed and based on experience with similar projects. The final cut-off required to produce a saleable product will need to be confirmed by future metallurgical testwork. This economic cut-off was applied to both the Central and South Zones. Although the South Zones contribute a relatively small tonnage, SRK is of the opinion that there are reasonable prospects for additional tonnage in this area which would then make these resources amenable to underground mining methods.

Table 16.7 : SRK Classified Mineral Resources for Pampa de Pongo at 15% Fe cut-off

ZONE

Classification

Volume

(Mm3)

Density

(T/m3)

Tonnage

(Mt)

Fe

(%)

Au

(g/t)

Cu

(%)

Central

Inferred

203

3.69

748

41.7

0.059

0.093

South

Inferred

32

3.59

115

39.5

0.130

0.121

Total

Inferred

235

3.68

863

41.4

0.068

0.097

16.12

Comparison with Previous Resource Estimates

Helsen (2005) estimated a total inferred mineral resource of 953 million tonnes at 44.7% Fe. This estimate was based on same drillhole dataset as the SRK resource estimate. A comparison between the current SRK and Helsen estimates is included in Table 16.8 and 16.9.

The main difference in tonnage between the two estimates results from the interpretation of the lateral extent of the 3D magnetic inversion data. SRK’s estimate was based on a 2008 magnetic survey and a recent interpretation completed by Quantec Geophysics while Helsen’s was based on 2004 magnetic data covering a much smaller survey area. . The end result was that the Helsen interpretation produced a much larger lateral surface area and resulting volume. Quantec Geophysics investigated this difference and concluded that it was possible to validate both the 2004 surface area used by Helsen and the 2008 surface area used for SRK’s resource estimate based on the magnetic database available for each estimate. As the 2008 magnetic data covers a larger area in more detail the SRK resource estimate is considered more reliable.

In addition, Helsen adopted an average thickness based on drill hole intercepts and multiplied this average thickness by the estimated surface area resulting in a volume represented by a vertical cylinder. The average thickness estimated by Helsen was biased by the drill hole intercepts which were targeted towards the core of mineralization and thickest portion of the 3D magnetic anomaly. SRK’s volume reflected the gradual tapering of the magnetic anomaly downwards around the periphery of the mineralization as suggested by the 3D magnetic inversion as well as the two drill holes intersecting the outer limits of the mineralization (drill holes RTDDH-2 and RTDDH-3). The 3D  dome-shaped solid used by SRK to estimate resources compared to Helsen’s cylindrical shape as well as differences in the interpreted lateral extent of mineralization explains a significant portion of the volume and tonnage differences between the two estimates.   

 Differences in estimated grade in both the Central and South Zones indicate that the SRK estimate also adopted a lower Fe grade to define the solid boundaries used to define composites for resource estimation. As a result a larger number of lower grade intervals were included in the SRK estimate.


Table 16.8 : Comparison of SRK and Previous Resource Estimates of the Central Zone

Estimate

Zone

Tonnes

(Mt)

Grade

(% Fe)

Helsen (2005)

Central

848

44.9

SRK Resource Model

Central

748

41.7

Table 16.9 : Comparison of SRK and Previous Resource Estimates of the South Zone

Estimate

Zone

Tonnes

(Mt)

Grade

(% Fe)

Helsen (2005)

South

105

43.0

SRK Resource Model

South

115

39.5

Table 16.9 indicates that the magnetic inversion interpretation generated an approximately similar tonnage for the South Zone.  With the SRK estimate, the South Zone tonnage is slightly larger and the grade was lower as a result of a lower cut off used to define the solid than was used by Helsen.   In both the Helsen and SRK estimates, the lack of low grade above the higher grade mineralization, the relatively narrow thickness of mineralization and the proximity of the mineralization to surface, all contributed to a much more consistent interpretation of the lateral extent and volume of the mineralization. In SRK’s opinion, the Helsen resource estimate reasonably reflects the 2004 magnetic data, the resource estimation method used and appropriately adopts an inferred resource classification.




17

Other Relevant Data and Information

A preliminary project schedule is shown Table 17.1. Only the major tasks have been highlighted and the assumptions made that positive results will continue, there are no unplanned delays and timely work is done. This schedule is for illustrative purposes only to show the long lead time (6 years) from present to production.

Table .: Possible Pre-Production Project Schedule

 

Year

2009

2010

2011

2012

2013

2014

Activity

Quarter

1

2

3

4

1

2

3

4

1

2

3

4

1

2

3

4

Drilling (res., geotech, hydro, metallurgy)

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

Pre-Feasibility Study

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

Development permit application

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

Environmental baseline and EA

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

UG dev., exploration, sampling, trials

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

Feasibility Study

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

Mining permit application

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

Financing

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

Mine and plant construction

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

Production

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 2015



18

Additional Requirements for Technical Reports on Development Properties and Production Properties

18.1

Mining

18.1.1

Mining Context

Geotechnical Evaluation

A preliminary geotechnical evaluation was conducted to assess and characterize the rock mass of the Central and South Zones for open pit and underground mining development options. Based on this evaluation, input recommendations for mining method selection and potential mine design parameters have been provided.

This evaluation is based on a limited number of exploration holes that have been drilled at the site by both Rio Tinto and Cardero. Of the 28 drillholes completed at the Pampa de Pongo project, the Central Zone and South Zones each have five holes that intersect the inferred resource.

Information used for the geotechnical evaluation included:

  • Drillhole logs for Rio Tinto and Cardero. Cardero drillholes which included RQD and recovery data, and collar co-ordinates. RQD values were not available for Rio Tinto drillholes;
  • Core photographs for all Cardero and Rio Tinto (split core only) drillholes;
  • The June 2008 structural report completed by SRK Consulting entitled “Evaluation of the Structural Geology of the Pampa de Pongo Fe Project, Peru.”  See Appendix C.

The geotechnical evaluation included a review of representative cores in the Lima core yard, followed by a comprehensive review of the drillhole photographs of all the available drillholes in the zones of interest. Rock Mass Rating (“RMR”, Laubscher 1990) geotechnical parameters were estimated for all the sections of the cores. From this assessment, broad geotechnical domains were determined and a representative range of RMR’s estimated for each of the domains.

A summary of the geotechnical evaluation is presented in this section. The detailed geotechnical descriptions and results of the rockmass evaluation are presented in Appendix A. All geotechnical descriptions have been interpreted from limited drillhole intersections of the mineralized zones and surrounding rockmass at the Pampa de Pongo project.

Geotechnical Domains

The geotechnical domains at the Pampa de Pongo project can be separated into four basic units: Overburden, massive mineralization, semi-massive mineralization, and country rock.

Figure18.1 and 18.2 show the summary results of the evaluation, with preliminary domains of ‘poor’ (brown), ‘fair’ (yellow), and ‘good’ (green) rock mass conditions assigned.

Geotechnical Description and Design Parameters – Central Zone

The Central Zone is characterized by shallow overburden of approximately 30 – 40 m depth consisting of conglomerates, gravel, sand and clay. Country rock beneath the overburden consist of volcaniclastic and dyke units, with rock mass conditions for all country rock types within the Fair range (estimated RMR 40 – 50).

The mineralization in the Central Zone consists of massive style mineralization beneath semi-massive style mineralization. In general the massive mineralization is of fair rock mass quality (RMR values estimated between 45 and 55) while the semi-massive is considered to be of poor to fair rock mass condition (RMR 30 – 45).


[techrep065.jpg]

Mineralization is outlined in solid red line, with faults represented by dashed red line.

Figure 18.1 : Preliminary Geotechnical Domains for the Central Zone (section view looking east)

Based on the estimated RMR values established in the geotechnical evaluation, the design parameters shown in Table 18.1 have been developed for the underground mining analysis.

Table 18.1 : Central Zone Geotechnical Parameters

Domain

RMR

Caving Hydraulic Radius

Country rock

40 - 50

20 - 27

Massive mineralization

45 - 55

24 - 30

Semi-massive mineralization

30 - 45

16 - 24

Hydraulic Radius (HR) is a parameter that characterizes the rock mass potential to cave. The hydraulic radius is a number derived by dividing the area of an excavation by the excavation perimeter. The hydraulic radius required to ensure propagation of the cave refers to the unsupported area of the cave back, that is, there is space into which caved material can move. No pillars can be left and caved material must be removed.

The geotechnical characteristics of the Central Zone are considered to be suitable for caving mining methods. The hydraulic radius of the Central Zone is markedly smaller than the mineralization footprint.

Geotechnical Description and Design Parameters – South Zone

The South Zone overburden consists of conglomerates, gravel, sand and clay with a thickness of between 100 m and 300 m. For a complete description of the overburden the reader should refer to the Memo entitled “Pampa de Pongo: Conglomerate Strength” prepared by SRK Consulting and contained in Appendix A.


[techrep067.jpg]

Mineralization is outlined in solid red line, with faults represented by dashed red line.

Figure 18.2 : Preliminary Geotechnical Domains for the South Zones S1 (left) and S2 (right) (section view looking east)

The country rock below the overburden primarily consists of volcanic units and, in general, rock mass conditions around the mineralized zone are considered to be Poor to Good (estimated RMR 35 – 55). Very poor rock mass conditions were observed immediately beneath the S1 Zone, with RMR estimated at <25.

The mineralization observed in drillholes from the South Zone is consistent with that seen in the Central Zone, having both massive and semi-massive style mineralization types, and generally the massive mineralization being located beneath the semi-massive mineralization.

Based on the estimated RMR values established in the geotechnical evaluation, the design parameters Presented in Table 18.2 have been developed.

Table 18.2 : South Zone Geotechnical Parameters

Domain

RMR

Stable Hydraulic Radius

Country rock

25 - 45

N/A

Massive mineralization

45 - 55

2.2 - 3.2

Semi-massive mineralization

30 - 45

1.2 - 2.2

Partial extraction mining methods could be considered in the South Zones. Cave mining methods are not considered feasible due to the irregular mineralization geometry, weak overburden materials, poor quality country rock mass, and the limited stoping dimensions achievable within the weak mineralization.

Based on the geotechnical review, an open pit scenario could also be considered for the South Zone (S2). The overburden pit wall slope for a potential South Zone pit would have to be at a very shallow angle (26o) due to conglomerate cementation that is predominantly weak, as seen as large zones of loose sand in drill core.  Additional holes in the overburden may be able to further define the overburden characteristics and strength variability. In this situation, the following slope angles presented in Table 18.3 are recommended.

Table 18.3 : South Zone Open Pit Geotechnical Parameters

Domain

Slope Angle

Overburden

26o

Country rock

45o

Hydrogeology

No information was available on hydrogeology for the Central or South Zones and it was assumed that any potential water inflow into the mines could be managed without excessive cost. The Pampa de Pongo project is located in a desert with extremely low precipitation and will not likely be influenced at all from precipitation. The existence and nature of any aquifers in the potential mining areas must be investigated if the project is to continue.

Grade Distribution

The grade distribution in the Central Zone appears to show an increase in grade with depth to the basal extent of the mineralized zone. Grades in the lower part of the Central Zone average about 46% Fe while grades in the upper part are in the 27% Fe range. Table 18.13 shows a distribution of grades in the upper (C1UP) and lower (C1DWN) of the Central Zone.

Structural Geology

The following excerpt is taken from the report entitled “Evaluation of the Structural Geology of the Pampa de Pongo Iron Project,Peru”, prepared by SRK Consulting in June 2008. This report is attached as Appendix C.

“Although data quality from Pampa de Pongo is relatively good, the data is sparse because of the current stage of exploration and lack of outcrop. Therefore, structural interpretations are tentative and likely to change with further drilling.

Fe-mineralization is hosted by the regional-scale NNW-trending Hucca fault system. In the vicinity of Pampa de Pongo ENE-trending faults affiliated to the Repiticion Group of Hawkes et al (2002) intersect this structure. In addition, a WNW-trending structural fabric is observed on magnetic survey maps (Analytic Signal). The deposit is therefore expected to be affected by brittle fault structures, including some significant fault zones.

Photographs of drillcore preserve evidence of significant numbers of brittle structures in the Pampa de Pongo area. The cores are affected by numerous broken and gougey zones associated with moderate to steeply-dipping slip surfaces. It has not been possible to validate the precise nature and style of faulting which have been reported elsewhere in the Marcona District.”

Further definition of the fault structures in the Pampa de Pongo area will be crucial for future mine development. The effect of structures on the possibility of mining has not been considered, as at this stage a caving direction has not been selected.

18.1.2

Mining Method Selection

The context of the Central Zone deposit did not make it readily apparent whether an open pit or underground mining method would provide the best economic outcome for the project. It was, therefore, decided to conduct a Whittle™ Optimization analysis to determine the potentially best approach based on preliminary parameters. The parameters selected for this early-stage study are shown in Table 18.4.

Table 18.4 : Initial Whittle™ Optimization Parameters

Item

Unit

Value

Fe Price (US$) FOB Peru coast

$/dmt pellets

130

Grade of pellets

% Fe

67

Mining Cost  - Open Pit (“OP”)

$/t rock

1.50

Mining Cost  - Underground (“UG”)

$/t ore

Varied from 5 to 50 (bulk to

Millhead grade

% Fe

45

Milling Cost

$/dmt pellets

13.00

Milling Cost

$/dmt Fe rec

19.40

Milling Cost

$/dmt millfeed

7.86

Process Recovery

%

90

Mining Recovery

%

99%

Mill/Production rate

tpa

18,169,000

Slope angles

degrees

45° Central,

Discount rate

%

10

Open Pit Analysis

The Whittle™ OP/UG cross-over study indicted the underground option for the Central Zones would yield better economic returns than the open pit scenario. This conclusion was based on the early assumption that the underground mining method would likely be a bulk mining method and, as such, an underground unit operating mining costs of $5.00/tonne was used. The verification of the suitability of a bulk mining (caving) method was later confirmed.

The study found that the Central deposit, with 300+ m of overlying waste was too deep and required too much pre-stripping to be economic. The Whittle™ optimization analysis did, in fact, generate an open pit shell, however, when a preliminary mining schedule was applied to the shell it became clear that the open pit option would not be economic. It must be noted, however, that the open pit scenario was investigated using zero grade material above the main Central Zone mineralized domain. If further drilling is done on the deposit the lower-grade, semi-massive mineralization above the main Central Zone should be estimated and an open pit re-evaluated.

Two other significant factors came into play in the Central Zone open pit analysis. As previously stated, Fe grade in the Central Zone increases with depth, so an open pit would only reach the higher grade material later in the mine life. Additionally, once it was determined that the central Zone had the potential to be mine with a low-cost block or panel cave method, the open pit option was eliminated as an option in this study.

Underground Method Selection

After compiling the available deposit context information, a review of underground mining methods was undertaken. The first and generally most desirable method investigated was a massive mining method. Block caving is generally the lowest cost, highest production bulk underground mining method and is appropriate for large deposits with consistent grades like Pampa de Pongo’s Central Zone. Caving operations have proven very successful in base metal mines in South America. The large physical footprint and vertical extent of the Central Zone, coupled with favourable preliminary ground conditions made the caving option the method of choice.

Caving Methology

Block Caving and Panel Caving are mining methods which do not use drilling and blasting to fracture production rock, but rather rely on an initial unsupported spans to fail and create broken rock, which flows by gravity to an extraction location, a drawpoint. This is achieved by opening up a large enough area beneath the mineralization to become geotechnically unstable and fail. Broken material in the cave is drawn out though openings beneath the undercut level (the start of the cave). As material is drawn from the cave, a void is opened and more material fails from the unstable span to fill the void. See  for a theoretical section view of how a cave works.

As mineralized material is drawn out of the cave zone, the cave will continue to expand until eventually waste rock outside of the mineralized zone will get mixed with the mineralized rock and create dilution. The pulling of muck from a drawpoint will continue until the rock becomes diluted to a point of not being economic (the cut-off grade). When this happens the drawpoint will be closed and another new one opened up in its place.

[techrep069.jpg]

Figure 18.3 : Principles of Caving Methods

18.1.3

Mine Design

Footprint Finder Analyses

Gemcom’s PCBC (Personal Computer Block Cave) Footprint Finder software was used to determine the most economic cave footprints or aerial extent for the Central Zone cave. The software works by splitting the deposit into individual columns and calculating the maximum economic value of each column based on the input parameters (see Table 18.5). The cave footprint is compiled by grouping regions of columns with positive economic values.

Table 18.5 : PCBC Footprint Finder Parameters

Item

Block 1

Block 2

Comment

Iron Price ($/T)

$194.03

$194.03

 

Iron Process Recovery

90%

90%

 

Milling Cost  ($/T)

$7.86

$7.86

 

Mining  Costs ($/T)

$3.00

$3.00

 

Z1

0

0

FROM height (below TIN)

Z2

600

600

TO height (above TIN)

HIZ

60

60

Height of Interaction zone (Laubscher)

FIRST_DIL

0.75

0.75

First dilution entry (Laubscher)

DEV_COST

1000

1000

Development cost per unit area

DISCOUNT

0.1

0.1

Discount rate (Eg 0.1 or 10%)

VMINING

73

73

Vertical mining rate (eg 80m/y)

HMAX

600

600

Maximum allowable HOD

Caving Blocks

The shape of the base of the Central Zone deposit prevents a single, flat footprint from caving the entire deposit. With a single footprint significant quantities of waste would have to be mined from the south end of the deposit, or significant quantities of the mineralized zone would need to be left behind on the north end of the deposit. (See )  

[techrep071.jpg]

Figure 18.4 : Central Zone Section View Showing a Single Mining Block with Excluded Mineralization and Excess Waste

The deposit was analyzed in multiple parts to maximize the deposit recovery and prevent waste from being initially mined. Figure 18.5 presents the two mining blocks determined by this analysis. Block 1 was determined to be the largest economic footprint developed entirely in mineralized material at a single elevation. This initial block was designed to provide the quickest economic payback and encompasses the highest grade material and the greatest vertical cave height. The higher grade boosts initial revenue and the high cave height minimizes initial development costs.

Figure 18.5 : Central Zone Section View Outlining Two Cave Blocks

 The shape of Block 2 was determined to be the most economic footprint of the remaining mineralization. Block 2 was set back 70° from the base of Block 1, to accommodate geotechnical instability as a result of caving activities below. A pillar between the Blocks 1 and 2 was taken back at an angle of 70o and was excluded from the mining plan, as it will not likely be recovered by the mining of either block.

LOM Tonnes and Grade

Footprint Finder was used to calculate the insitu tonnages and grades of mineralized material in each cave block. The height of each column in each block was stopped at the point where external dilution (material not in the mineralized zone model) is first introduced to the column. An additional 10% was added to the total tonnes mined to account for dilution. This corresponded with an assumed grade factor of 90% over the life of the mine. Total diluted tonnes and grades of each block are listed in  and do not include mineralized development muck.

Table 18.6 : Mining Block Tonnages and Grades

Mining Block

Diluted Totals

MTonnes

Fe Grade (%)

Cu Grade (%)

Au Grade (g/t)

Avg. Cave Height (m)

Block 1

401

39.7

0.07

0.04

265

Block 2

160

37.7

0.11

0.09

215

Total

561

39.1

0.09

0.05

250

Production Rate

Using a vertical cave advance rate of 100 mm per day, a drawpoint area of 250 m² per drawpoint and average cave heights for each block, maximum production rates were estimated. These factors were determined based on SRK experience with other caving projects. Although 100 mm per day has been used for the entire life of the mine, cave advance rates of 200-250mm per day may be achievable once a drawpoint has been open for some time. Based on the draw rate and the number of drawpoints available for extraction a mining rate of 75,000 tpd was selected. This rate provides a 24 year mine life, using approximately 800 drawpoints to meet full production.

Based on experience, it was assumed that production in Block 1 would take five years to ramp up to its steady target rate of 75,000 tpd. For this study, the ramp-up rate was assumed to be linear, although the actual rate of production in the initial years will start off slow and increase as the development of the undercut progresses.

Caving Methods

At 250m² per drawpoint, Block 1 will have a total of 1740 drawpoints and Block 2 will have a total of 844 drawpoints. 

Block 1 will be mined as a panel cave, since production targets can be met without having to open all the drawpoints ate one time. There will be opportunity to distribute the capital cost of and undercut development and drawpoint construction over the life the mine with a panel cave method.

Block 2 will be mined as a block cave since the majority of drawpoints must be open in order to meet production targets. All undercut development and drawpoint construction must be completed by the time that Block 2 reaches full production.

Dilution

It is typical of a caving operation to keep pulling from a drawpoint until the material is no longer economic – it falls below the cut-off grade. Depending in the grade of the mineralization this may support a high factor of dilution. For this study grade factors were explicitly defined. It was assumed that the mine would have a Grade Factor of 95% (95% of the value of insitu grade will reach the mill) for the first 3 years of mine life. A Grade Factor of 90% has been assumed for the remaining years of mine life. The higher grade factor in the initial 3 years of cave development was based on the assumption that, while growing, the cave would not reach sufficient height to encounter excessive external dilution. It also assumed that the mineralized development muck, about 3% of total LOM mill feed, would have no dilution.

The grade factor was held constant for the year 4 and beyond and assumed that all high dilution mineralization encountered will be offset by low dilution mineralization from newer drawpoints.

Cut-off Grade

Because grade factor was explicitly defined, a cut off grade was not used in the Footprint Finder mining simulation. Mining of a column was cut off as soon as any dilution external to the C1UP and C1DWN regions defined in block model was encountered.

18.1.4

Mine Development Design

Upon obtaining the basic caving block shapes from Footprint Finder, mine development planning commenced using Mine 2-4D mine planning and scheduling software. All capital development was planned in 3D to determine development types, lengths and quantities as well as infrastructure required to support a 75,000 tpd Central Zone mine.

Figures 18.6 and 18.7 show section views of the planned development in relation to the mineralized zone. Figures 18.8 and 18.9 show the mine development in plan and isometric view respectively.

[techrep073.jpg]

Figure 18.6 : Mine Development and Central Mineralized Zone Looking South (200 m gridlines)

 [techrep075.jpg]

Figure 18.7 : Mine Development and the South Mineralized Zone Looking East (200 m gridlines)

[techrep077.jpg]

Figure 18.8 : Plan View of Mine Development Design

[techrep079.jpg]

Figure 18.9 : Isometric View of Mine Development Design

Primary Development

Table 18.7 summarizes the primary development components of the mine design including dimensions and length. A total of approximately 115,000 m of primary development was estimated for the LOM plan.  

Table 18.7 : Primary Components of Mine Design

Description

H (m)

W (m)

Length (m)

Access

   

Main Decline

5.5

6

             6,100

Conveyor Decline

4.5

5

             6,100

Conveyor Access Ramp X-Cuts (15m every 100)

4

4

             1,800

Remuck bays (20m every 150m)

5.5

5.5

                800

Safety bays (2m every 30m)

2

2

                400

Sump and Misc Dev.

5.5

5.5

                500

Block 1

   

Ventilation Drifting

6

6

           11,400

Haulage Level Drives

5.5

6

             6,100

Haulage Level X-Cuts

5.5

6

                600

Extraction Level Drives

5.5

5.5

             6,100

Extraction X-Cuts

5.5

5.5

           15,900

Access Ramps

5.5

5.5

             1,200

Undercut Access (Ramps and Drifts)

4.5

4.5

             3,300

Conveyor Drift

4.5

5

             1,200

Misc. Infrastructure (shops, crushers, etc)

5.5

5.5

           10,000

Block 2

   

Ventilation Drifting

6

6

           11,300

Haulage Level Drives

5.5

6

             4,900

Haulage Level X-Cuts

5.5

6

                400

Extraction Level Drives

5.5

5.5

             4,700

Extraction X-Cuts

5.5

5.5

             7,900

Access Ramps

5.5

5.5

             1,900

Undercut Access (Ramps and Drifts)

4.5

4.5

             2,600

Subtotal Development Metres

  

         105,200

Raisebore Development Metres

   

Block 1

   

Ventilation Shafts

8.5

 

             4,200

Ore Passes

5

3.925

             1,100

Exhaust Raises

4

3.925

             1,700

Block 2

   

Ore Passes

5

5

             1,100

Exhaust Raises

4

28

             1,400

Subtotal Raisebore Development Metres

  

             9,500

Secondary Development

Secondary development was comprised of undercut and drawpoint excavations and was not planned in detail. Instead, factors for other caving operations were used, summarized by an effective area per drawpoint of 250 m2. Common dimensions and costs were assumed for the undercut and drawpoint and are explained in the OPEX section of the report.

Individual vent raises between extraction level cross cuts and exhaust drifts were not modeled, but their cumulative development length was estimated and added to the development schedule.

Ventilation Development

Construction Ventilation

Auxiliary vent fans and tubing will be used to ventilate the main and conveyor declines while they are under construction. When both declines are complete (access and conveyor) they will provide a loop to supply all ventilation to the mine until vent shafts to the surface are completed with fresh air traveling down the access ramp and return air traveling up the conveyor decline.

Auxiliary vent fans will be used to provide adequate air to dead end working faces during all development activity. During construction there will be a maximum of 30 active faces at one time.

Production Ventilation

Once the ventilation shafts to surface are completed, fresh air will be taken in though the three 8.5 m diameter intake vent shafts on the west side of the mine and exhausted through three matching exhaust vent shafts on the east side of the mine. Each intake and exhaust shaft will have fans on surface forcing air down it or drawing air up it respectively.

On the haulage levels, fresh air will pass through the length of all primary tunnels from west to east. On the extraction level, fresh air will enter the west end, travel through the ring drifts around each side of the mineralized zone, through both ends of each cross cut, and down central vent raises to the exhaust level. (See )  Air to the undercut will be boosted and distributed to the crosscuts with auxiliary fans.

In all sections of the mine, ventilation doors, bulkheads and regulators will be maintained and adjusted to optimize the efficiency of the ventilation system.

[techrep081.jpg]

Figure 18.10 : Ventilation Schematic

De-watering Development

Development for a pumping level is included at the base of the main ramp. This development will contain sumps to collect water, facilities to settle solid particles out of the water and a permanent pumping station to transport the water to surface. No de-watering development was planned other than sump and pump station excavations.

Extraction Development

Construction Haulage

Initial mineralized development rock and development waste rock will be loaded onto haul trucks and transported out of the mine through the main decline. Once the conveyor is commissioned there will be opportunity to use excess convey capacity to transport development waste to surface. Provisions will be made on surface to redirect waste to a separate stockpile on surface.

Production Haulage

Perimeter haulage excavations were designed around the Blocks 1 and 2. ROM rock will be transported from the drawpoints to orepasses on the extraction level by LHDs. Trucks will haul the rock from the orepasses to one of two crushing stations located at each end of the deposit. The crushers will feed the primary conveyor to surface where the ROM rock will be stockpiled for processing. The stockpile will have a reclaim system that transports ROM rock from the stockpile to the mill as required.

Undercut Development and Drawpoint Construction

Block 1 will be a panel cave, requiring the caving panels to advance at a rate that will allow the mine to maintain full production. The undercut level will expand and new extraction drawpoints constructed at the required rate of panel advance. Block 2 will be a block cave, requiring the entire undercut to be developed and all of the drawpoints to be constructed before full production can be achieved. The total undercut areas and drawpoint requirements over the life of the mine are listed in . At a 75,000 tpd production rate, and a draw of 93 tonnes/drawpoint/day, approximately 800 drawpoints are required to be in production at any given time.

18.1.5

Mobile Equipment

Mining activities will be conducted with conventional underground mobile mining equipment. LHDs will be used for the transportation of muck from drawpoints to orepasses and trucks will be used to transport muck from orepasses to the crushers. Development and construction equipment was assumed to be done conventional diesel-powered underground equipment with electric-hydraulic jumbos, longhole rigs, rockbolters and other ancillary equipment. The equipment fleet estimated for this study is listed in Table 18.8.

Production Fleet Selection

To meet the production requirements of this study, 21-tonne capacity LHDs and 80-tonne haul trucks were selected. LHDs and haul trucks were assumed to be operating an average of 14 hours per day each. The approximate productivity of one unit was determined by applying the maximum haulage distance required of the unit to the productivity curves provided by the equipment manufacturer. This productivity was further reduced by 1/3 to reflect realistic working conditions.

Quantities of each machine were by dividing the required daily production by the calculated productivity of one unit and adding 15% for spare units.

Production Support Equipment

A rock breaker will be required at each active orepass to ensure that large material is broken to pass through orepass grizzlies. It has been assumed that as the caving panels advance in Block 1 not all 24 orepasses will be active at one time but that all 18 ore-passes will be active at all times in Block 2.

Drawpoint hang-ups due to oversized rock will be drilled and blasted to free up the drawpoint. One hang-up drill per six production LHDs was assumed.

Table 18.8 : Mobile Equipment Fleet

Equipment Type

Equipment Function

Total

Production

Development

Undercutting

Construction

Maintenance

Management

Production

10m³ LHDs

34

     

34

80t Trucks

18

     

18

Rock Breakers

20

     

20

Development and Construction

6m³ LHDs

 

5

3

5

  

13

40t Trucks

 

6

3

2

  

11

Shotcrete Sprayers

 

3

 

2

  

5

Concrete Trucks

 

4

 

3

1

 

8

Graders

1

  

1

1

 

3

Scissor Lifts

 

1

1

5

2

1

10

ANFO Loaders

 

3

1

1

  

5

Flat Deck/ Crane Trucks

   

5

2

 

7

Drilling

Long-Hole Drills

  

7

   

7

Raise Bores

   

3

  

3

Rock Bolters

 

5

3

2

1

 

11

Cable Bolters

   

1

  

1

1 Boom Jumbos

   

5

  

5

Oversize Drills

6

     

6

2 Boom Jumbos

 

5

3

   

8

Support Equipment

Lube/Fuel Trucks

    

6

 

6

Large Personnel Carriers

2

3

2

2

2

 

11

Small Personnel Carriers

1

3

2

2

6

8

22

Development and Construction Equipment

Table 18.9 outlines the basic productivity assumptions that were used to determine primary development and construction requirements. Remaining development equipment was factored based on the chosen quantities of primary equipment.

Table 18.9 : Development and Equipment Productivity Assumptions

Equipment

Productivity

Longhole Drill

525 m³ of undercut drilled per day.

2 Boom Jumbo

11.1 m of development drilled per day. (3 x 3.7m long rounds)

Construction Crew

12 days construction for a drawpoint per crew

Support Equipment

Support equipment was selected to service underground development and production operations. The quantities of transport and maintenance equipment were factored from the total equipment in the mine.

Crushers

Two underground crushers, each capable of crushing the full mine production rate to a 0.3 m nominal size were selected. The two crushers will feed a central single conveyor. Crushers were planned at either end of the haulage level to reduce the maximum haulage distance.

Conveyors

Conveyors were selected for the transport of rock to surface project due to their low operating costs per tonne of rock, the flexibility, and the speed at which a conveyor decline can be constructed from the exploration access decline.

Three methods were briefly considered in this study for transporting material out of the mine.

  • Haul Trucks
  • Hoisting up a shaft with skips.
  • Conveyors

Due to the depth of this mine the distance that trucks would have to haul to surface makes the haulage cost per tonne far greater than with shafts or conveyors. The quantity of trucks that would be required to move 75,000 tpd makes truck haulage impractical due to development and ventilation requirements and the number of haulage ramps required.

It was estimated that at least two production shafts would be required to transport 75,000 tpd of material out of the mine plus a men and materials shaft. It was estimated that a single conveyor decline could be used to transport the full mine production coupled with an access decline for men and materials. It was envisioned that the construction of a conveyor decline would be stared from multiple points along the length of the main decline. This will allow the conveyor decline to be constructed in less time than the shafts that would be required for this mine.

18.1.6

Development and Production Schedules

Development Schedule

The LOM schedule is shown in 0. The schedule is based on early access to the Central Zone deposit with an exploration decline, that will initially allow for sampling and testing of the mineralized zone, and that will later serve at the principle access to the mine. All subsequent development in the mine will start from the exploration decline. 

Critical Milestones in the development plan are at Years 1, 5 and 18 and are outlined below.

Year 1 - Primary development and infrastructure required for initial production:

  • Sufficient Block 1 undercut, extraction, and haulage level development to support initial production of 4.3 Mt.
  • Block 1 undercut opened to 38,000 m² to support initial production.
  • 271 Block 1 drawpoints constructed to support initial production.
  • Conveyor decline completed to transport first production mill feed rock.
  • One crusher online.
  • One Intake Vent Shaft, one Exhaust Vent Shaft, and 4 vent drifts to provide adequate air to support production equipment

Year 5 - Primary development and infrastructure required for full production

  • Block 1 extraction and haulage levels completed
  • Block 1 undercut expanded to 237,000m² support full production.
  • 811 Block 1 drawpoints constructed to support full production.
  • >Conveyor drift and all crushers online.
  • All vent shafts and Block 1 vent drifts completed.

Tear 18 - Primary development and infrastructure required for full Block 2 production

  • All Block 2 undercut, extraction, and haulage levels completed.
  • All Block 2 drawpoints constructed.
  • All Block 2 vent drifts completed.

Table 18.10: Life of Mine Development Schedule

Development Activity

Pre-Production Years

Production Years

-6

-5

-4

-3

-2

-1

1

2

3

4

5

6

7

8-18

Exploration / Main Decline

 

 

 

 

 

 

 

 

 

 

 

 

 

 

Sump and Misc Dev.

 

 

 

 

 

 

 

 

 

 

 

 

 

 

Haulage Level Drives

 

 

 

 

 

 

 

 

 

 

 

 

 

 

Extraction Level Drives

 

 

 

 

 

 

 

 

 

 

 

 

 

 

Extraction X-Cuts

 

 

 

 

 

 

 

 

 

 

 

 

 

 

Access Ramps

 

 

 

 

 

 

 

 

 

 

 

 

 

 

Ventilation Drifting

 

 

 

 

 

 

 

 

 

 

 

 

 

 

Ventilation Shafts

 

 

 

 

 

 

 

 

 

 

 

 

 

 

Conveyor Decline

 

 

 

 

 

 

 

 

 

 

 

 

 

 

Undercut Access Ramps & Drifts

 

 

 

 

 

 

 

 

 

 

 

 

 

 

Ore Passes

 

 

 

 

 

 

 

 

 

 

 

 

 

 

Internal Exhaust Raises

 

 

 

 

 

 

 

 

 

 

 

 

 

 

Misc. Infrastructure

 

 

 

 

 

 

 

 

 

 

 

 

 

 

Initial Drawpoints & Undercut

 

 

 

 

 

 

 

 

 

 

 

 

 

 

Continuing Drawpoints & Undercut

 

 

 

 

 

 

 

 

 

 

 

 

 

 

Conveyor Drift

 

 

 

 

 

 

 

 

 

 

 

 

 

 

Block 2 Construction

 

 

 

 

 

 

 

 

 

 

 

 

 

 

Production Schedule

The assumptions in Table 18.11 were used to determine the development and operating estimates for Blocks 1 and 2. These assumptions are based specific characteristics of the Central Zone deposit combined with typical operating parameters for other caving mines.

Table 18.11 : Production and Development Assumptions

Parameter

Unit

Block 1

Block 2

Ore bulk density

t/m3

3.7

3.7

Daily vertical advance

m

0.1

0.1

Daily production per drawpoint

tpd

93

93

Draw area per drawpoint

m2

250

250

Tonnage per vertical m per drawpoint

t

925

925

Cave height

vertical m

265

215

Total tonnage per drawpoint

t

245,125

198,875

Tonnes per m2

t

981

796

Daily production

t

75,000

75,000

Annual production

t

26,250,000

26,250,000

Active DPs

each

811

811

Mining was designed to start in Block 1 and ramp to full production in 5 years. Mining in Block 2 was scheduled to meet production requirements when Block 1 is depleted.  outlines the basic production schedule of the mine. Production estimates in Years 6-17 and 18-22 are assumed to be constant. Mineralized rock mined through development activities is included in the scheduled feed to the mill. Table 18.13 shows the LOM production tonnages by source.

Table 18.12 : Life of Mine Production Schedule

Parameter

Units

Pre-Prod.

Production Years

-2

-1

1

2

3

4

5

6-17

18-22

23

24

Block 1

 

 

 

 

 

 

 

 

 

 

 

 

Production required

Mt 

 

 

4.6

9.1

13.7

18.3

22.8

27.4

 

 

 

Minimum undercut area

 ‘000m2

 

30

38

102

135

169

203

237

 

 

 

Minimum drawpoints required

ea. 

 

 

271

406

541

676

811

811

 

 

 

Block 2

 

 

 

 

 

 

 

 

 

 

 

 

Production

Mt 

 

 

 

 

 

 

 

 

27.4

23.4

 

Minimum undercut area

 ‘000m2

 

 

 

 

 

 

 

 

203

203

 

Minimum drawpoints required

ea. 

 

 

 

 

 

 

 

 

811

574

 

Development

 

 

 

 

 

 

 

 

 

 

 

 

Mill feed from development

Mt 

 

 

2.7

1.5

1.1

1.5

1.5

 

 

4

8.6

MILL FEED

Mt/year

  

7.3

10.6

14.8

29.7

24.3

27.4

27.4

27.4

8.6

Mining Block Sequence

Block 1 will be mined first because of its larger size and greater potential return on capital. To maximize initial return, the first panels of Block 1 will begin in the central southern area of the undercut. This is where the highest columns and grades will be mined. Two panels will continue out from this point, one progressing to the north-west and the other to the north-east.

Construction and ramp-up production from of Block 2 will take place during the final years of Block 1 production, so that Block 2 will be capable of meeting production targets when Block 1 is depleted. 

Calculating Tonnes and Grades

The Footprint Finder module of PCBC is suitable for providing rough scheduling for this level of study.

Assumptions used when running Footprint Finder are as follow:

  • Full production rate is 67,500 tpd. This allows for explicit grade factor of 90% to be applied as described in Section 18.1.2.
  • Production for 5 year ramp up time is calculated linearly based on full production.
  • Columns are mined at defined vertical mining rate of 100mm per day.

The process that Footprint Finder follows when producing a schedule is as follows:

  • Highest value column is mined first.
  • Production is taken from adjacent columns as necessary to achieve required production.
  • Once production starts on a column, it continues until the column intersects external dilution.
  • Mining progresses out from the initial columns radically.
  • Total tonnes, footprint area required, and average grade of material mined is output for each year.

Mineralized Development Rock

Overbreak from all development and undercutting was assumed to be 10% and all development muck was assumed to have an insitu density of 3.7 t/m3.

For the calculation of total tonnage, undercut height was assumed to be 5m. The grade of each undercut level was determined from the average grade of all mineralized blocks at the undercut elevation.

Total mineralized development rock tonnages were calculated from individual development lengths and cross sectional areas. The quantities of each development type that are mined as mill feed or waste are based on a visual approximation and the average development grade is based on the average grade of areas of the block model occupied by development.

Table 18.13 : LOM Plan Mineralized Material Sources

Block

Source

Mtonnes

Fe Grade (%)

Cu Grade (%)

Au Grade (g/t)

1

C1DWN Mineralized Zone

210

51.4

0.10

0.06

C1UP Mineralized Zone

151

34.0

0.06

0.03

Total Undiluted Production

361

44.1

0.08

0.05

Production Dilution

40

-

-

-

Total Diluted Production

401

39.7

0.07

0.04

Mineralized Development Muck

11

47.6

0.11

0.07

2

C1DWN Mineralized Zone

113

46.0

0.14

0.10

C1UP Mineralized Zone

31

26.9

0.09

0.07

Total Undiluted Production

144

41.9

0.13

0.10

Production Dilution

16

-

-

-

Total Diluted Production

160

37.7

0.11

0.09

Mineralized Development Muck

6

45.4

0.11

0.07

Diluted Totals

Total Diluted Production

562

39.1

0.09

0.05

Total Mineralized Development Muck

118

46.8

0.11

0.07

 

Total LOM Mill Feed

580

39.4

0.09

0.05

18.1.7

Mining Support Services

Ventilation

Total ventilation requirements were determined using a conservative factor, based on SRK experience, of 1 m3/s of air for each Kt of ROM rock produced per month. For a 75,000 tpd production rate, 2,250 m3/s of ventilation air will be required. The total pressure drop through the mine was estimated to be 3.2 KPa. Using an assumed motor efficiency of 75% it was calculated that a total of 9.5 MW of fan power is required for the mine.

Maintenance

Equipment maintenance has been included in the development and mining costs used in this study. Maintenance of underground infrastructure such as drawpoints and concrete roadways has been included in the operating costs used in this study.

Adequate mobile equipment for all maintenance activities has been included in the mobile equipment fleet.

Ancillary Services

Capital has been included for underground services including an electrical distribution system, communication system, water collection and pumping system, underground maintenance shops, and refuge stations.

18.2

Recoverability

This report assumes a magnetic separation recovery of 93.4% of total iron. This estimate is based on metallurgical test work on Pampa de Pongo samples. Recovery assumptions for copper and gold are both 50% and are only based on experience from other operations. No metallurgical testing of copper and gold recoveries was done to support the assumed recoveries and is therefore unreliable. The payable copper and gold contribution to the cash flow is only about 3% of the total value.

18.3

Markets

Unlike most metals which are openly traded, the iron ore market is a competitive market with the majority of producers and consumers trading with each other directly under closed contractual conditions. As iron ore is an internationally traded bulk material, with the transport costs a significant proportion of the delivered cost to customers, the relative locations of the suppliers and consumers forms a very important component in the commodity’s economics.

Pampa de Pongo is in an advantageous location to ship iron ore products to Asian or Indian markets. The Pampa de Pongo property lies 38 kilometres from an existing deep-water port on the Pacific coast.

18.3.1

Iron Ore Market Background

Most mining companies sell iron ore on a ‘Free On Board’ (‘FOB’) basis, at the port of origin. The cost of shipping from the port of origin to the customer’s destination port is usually added to the FOB price and paid for by the customer and will vary for each individual customer. The basic prices at each destination port are effectively controlled by the FOB price at the major production centers plus shipping costs. Delivered cost is usually referred to as ‘Cost, Insurance and Freight’ (“CIF”).

It is assumed that Pampa de Pongo will sell its iron ore products (blast furnace and potentially direct reduction grade iron ore pellets) for similar prices that major companies such as Vale, Rio Tinto and BHP Billiton realize, at the same destination ports. Any advantage arising from differences in shipping costs (due to the more favourable geographic location of the deposit) are a ‘plus’ to the overall costing structure.

The price at destination port (CIF) is calculated as FOB Price plus shipping costs to destination. Pampa de Pongo’s FOB price would be Price at Destination minus Pampa de Pongo shipping costs to Asia. Location is therefore extremely important. For example, the shipping cost associated with Brazilian iron ore pellets would be expected to be higher than that for Pampa de Pongo. Hence, the FOB price for Brazilian iron ore pellets would be lower than that for Pampa de Pongo (on an equivalent percentage iron-grade basis).

Shipping costs for Pampa de Pongo iron ore pellets to the target markets of Asia and India are uncertain at this time. Pricing for this study was therefore determined by reference to historical, current and long-term trends for world blast furnace pellets FOB.

Current Iron Ore Markets

The years 2002-2003 saw a shift in the iron ore markets with sharply increasing demand, fuelled primarily by economic and infrastructural growth in China. The iron industry responded with increased iron ore production in tandem with much-publicized price increases for all products, including blast furnace and direct reduction pellets.

In September 2008, Cardero retained Global Strategic Solutions Inc. to conduct an iron industry market study. The results were encouraging, generally forecasting sustained demand and market growth. Taking known future capacity into account, there will continue to be shortages of iron ore (42 Mt in 2010, reduced to 30 Mt by 2017) and a continuing year-on-year 3.9% growth in the steel industry. Iron ore shortfalls are predicted despite expected increases in supply of 7% per year.

The overall iron ore market is still forecast to grow and as a result the overall price trend is forecast to continue rising. Aside for continuing market growth, factors contributing to iron ore prices include the cost of energy (natural gas and fuel oil), labour costs, and inflation.

Pampa de Pongo Pricing Assumptions

FOB pricing is usually expressed as US¢/dmtu (US cents per dry metric tonne unit); in other words, as the price per tonne for material grading 100% iron. To convert this figure to a more intuitive US$/tonne, the dmtu figure is multiplied by the actual grade of the product being sold.

The long-term blast furnace iron ore pellet FOB price used for this study is presented as a low, long-term price, a higher price reflecting expected market growth and a third price, between the two.

The low price is a calculated mean of 2006-2008 pricing from Vale’s Sao Luis operation in Brazil. The result is US¢154/dmtu ($99.35/tonne at 64.5% iron pellet grade). This figure is in line with long-term forecasting from financial institutions, ranging between US¢135/dmtu to US¢164/dmtu, which have historically been conservative prices.

The middle price is US¢198/dmtu ($128/tonne) and is sourced from an independent market study, prepared for Cardero Resource Corp. by Global Strategic Solutions Inc.

The higher price, US¢230/dmtu ($150/tonne @ 64.5% iron grade) is sourced from recent and comparable iron project preliminary economic assessments. This price is considered to be most realistic as a prediction of future blast furnace pellet prices.

Following a positive Material Evaluation from MIDREX Technologies Inc., the Pampa de Pongo pellets have been assessed as suitable for Direct Reduction process feed. DR-grade iron ore pellets typically carry a premium of 10% over blast furnace grade pellets as the processing requirements for gas based direct reduction plants are more stringent than that of the blast furnace. Pricing assumptions in the discounted cash-flow reflect this price premium.

18.4

Contracts

There are no contracts applicable to this report.

18.5

Environmental Considerations

No significant environmental work has been conducted on the project. Based on there being no communities which would be impacted directly by mining operations and the almost complete lack of surface water, flora and fauna on the property, SRK does not envision any environmental fatal flaws.

The Government of Peru does not require payment of an environmental bond ahead of mine development.

18.6

Taxes

Peru has an 8% corporate income tax established to help distribute corporate profits to workers of all levels. The tax is used to compensate workers according to their specific wage categories as defined by the Peruvian government. Contractors must contribute the 8% tax for their own workers. This tax is applied to taxable income prior to the calculation of regular corporate income tax.  

18.6.1

Royalties

Peru’s royalty payment requirements are shown in Table 18.14.

Table 18.14 : Mining Royalty Rates

NSR Value Range

Royalty

(% of NSR Value)

NSR <  $60 M

1%

$60 M < NSR < $120 M

2%

NSR > $120 M

3%

18.6.2

Value Added Tax

There is a 19% value added tax in Peru for all goods and services. There are means to recovery this tax back from the government and, as a result, the cash flow model does not show VAT as it is assumed that full recovery of the value added tax will be realized.

18.6.3

Corporate Income Tax

Peruvian corporate income tax has been assumed to be 30% in this study. The actual income tax paid will likely vary when detailed tax assessments are done, but 30% is considered appropriate for this level of study.

18.6.4

Depreciation and Amortization

Depreciation and amortization were taken into account in the simplified cash flow calculation. Depreciation schedules were assumed to be as per Table 18.15. The capital costs in the cash flow model were only broadly categorized into depreciation classes and scheduled accordingly.

Table 18.15 : Depreciation Assumptions

Depreciation Class

Depreciation Period

Depreciation method

Machinery

5 years

Straight line

Equipment

10 years

Straight line

Buildings

30 years

Straight line

Amortization calculations were done over 10 years beginning in the first year of commercial production.

18.7

Operating Cost Estimates (OPEX)

Operating costs for the PEA were based on factored estimates from other operations and well-established rules of thumb adapted to the specifics of the Pampa de Pongo project. The OPEX assumptions and their source are given in this section.

18.7.1

Mining OPEX

Mining costs were estimated based on SRK’s experience from other large South American block caves. The base parameters for the estimation of the mine OPEX are shown in Table 18.16.

Table 18.16 : Mine OPEX Assumptions

Parameter

Unit

Block 1

Block 2

Production development costs

$/m2

1,500

1,500

Mine OPEX

$/t for extraction

$3.50

$3.50

 

$/t for development

$1.60

$1.97

Total Mine OPEX

$/t

$5.10

$5.47

Less Secondary Development assigned to Capital in Year -1

$

(45,000,000)

Average LOM Mine OPEX

$/t

$4.73

18.7.2

Mineral Processing OPEX  

Design and cost estimates are based on published technical papers as well as plant operating information regarding the OPEX and CAPEX of comparable plants and processes.

Beneficiation

The total OPEX for a comparable Peruvian Verde crushing and grinding circuit was estimated to be $1.695 per tonne. The Pampa de Pongo OPEX is assumed to be the same on this cost per tonne basis. An OPEX of $0.30/tonne has been estimated for the flotation circuit.

In addition to the liberation OPEX, the magnetic separation, vacuum disk filtering, and slurry pipeline OPEX is estimated at $1.50/tonne of concentrate.

Pelletization

The average OPEX for two proposed 5 million tonne per year pellet plants (recent and comparable study) is $9.22 and was used as the basis for the Pampa de Pongo estimates after adjustment for Peruvian cost conditions.

Electricity in Peru is $0.045 per kWhr compared to US$0.075 per kWhr in the reference host-country, allowing a $1.42 deduction from the Pampa de Pongo OPEX. Mean labour costs in Peru are $13,746 per labourer year. Therefore, the labour OPEX per ton of pellets in Peru is US$0.344., compared to $2.65 in the reference project country. This allows a $2.31 deduction for the Pampa de Pongo OPEX. The total Pampa de Pongo pelletizing OPEX is estimated at $5.49 per tonne of pellets.

18.7.3

General and Administration (“G&A”) and Site Services OPEX

G&A costs were estimated at $0.80/t milled based on a study for a similar sized mine in Peru.

Site services costs of $0.50/t milled will cover the operation and maintenance of site infrastructure including waste management, road maintenance, primary power maintenance, water supply and treatment.

18.7.4

OPEX Summary

A summary of the on-site LOM unit OPEX is shown in Table 18.17.

Table 18.17 : Unit OPEX Estimate Summary

Description

Unit

Cost

Mining

$/t milled

4.73

Beneficiation

$/t milled

1.70

Magnetic sep/Filtering/etc

$/t milled

1.50

Flotation plant

$/t milled

0.30

Pellet plant*

$/t milled

3.14

Site services

$/t milled

0.50

G&A

$/t milled

0.80

OPEX per tonne milled

$/t milled

12.67

OPEX per tonne of pellets

$/t of pellets

22.16

18.8

Capital Cost Estimates (CAPEX)

18.8.1

Mining

Development

The costs used for each type of primary development in the mine model are listed in Table 18.18 with total costs estimated by using the primary development meters shown in Table 18.7.


Table 18.18: Development Costs

Development Item

Dimensions

(m)

Cost

($/m advance)

Comments

Ramps and Drifts

Main Decline

5.5 x 6

$10,000

 

Conveyor Decline and Drift

4.5 x 5

$15,000

Includes Installed Conveyor

Ventilation Drifting

6 x 6

$7,200

 

Haulage Level Drives

5 x 6

$8,000

Concrete Floor

Extraction Level Drives

5.5 x 5.5

$7,500

Concrete Floor

Access Ramps

5.5 x 5.5

$7,500

 

Undercut Access Ramps & Drifts

4.5 x 4.5

$7,000

 

Raises and Shafts

Ore Passes

5 mØ

$7,000

Raise Bored and Lined

Vent Passes

5 mØ

$5,000

Raise Bored

Vent Shafts

8.5 mØ

$52,000

Raise Bored, Stripped & Lined

Other Excavation and Construction

Drawpoint & Undercut

$1,500/m²

excluding access or extraction cross cuts

Development costs of drifts and ramps were estimated from first principle analysis of materials and labour for a standard sized development of 5 m wide by 5 m high. Costs of other development were extrapolated from this based on the difference of excavated volume. A rate of $600 per metre was added to drifts requiring a concrete floor.

Conveyor costs were based on another South American study at an installed cost of $15,000 per metre for the conveyor drift and decline.

The costs of a raisebored and stripped shafts, lined “ore” passes and unlined vent raises are based on SRK data from similar projects. 

A cost of $1500/m² for constructing and expanding the extraction and undercut levels of a caving mine is based on costs being realized by current operating caving mines. This reflects the total costs of increasing a mine's footprint by 1m² and includes in it the excavation of an undercut and the construction of required drawpoints. 

All undercut and drawpoint construction required prior to Year 1 is considered capital expense. All undercut and drawpoint construction after Year 1 is factored into the operating costs of the mine.

A total of $1,180 M in mine development capital will be spent over the life of the mine.

Stationary Equipment

  • Ventilation:  Initial exploration ventilation has been estimated at $0.5 M in Year -5 and an additional $0.5 M in Year -4. These capital costs will cover primary and auxiliary vent fans and ducting required to access the underground and support development work until ventilation shafts have been sunk and permanent ventilation fans have been installed in them.
  • For the permanent ventilation system, an installed ventilation plant cost, based on typical costs from other projects, of $2,000 per installed kW of required fan power has been used. The Pampa de Pongo Mine will require 9.5 MW of power and is therefore estimated to cost $19 million.
  • Crushers:  Two installed 75,000 tpd crushers including underground construction and infrastructure, (power, chutes, grizzlies, rockbreakers, etc.) were estimated to be $50M.
  • Material Transport:  Conveyors were costed as part of capital development but the associated items such as transfer points, chutes, grizzlies, rockbreakers were estimated separately at $10 M.
  • Surface Stockpile/Reclaim:  The surface stockpile and reclaim system for the ROM stockpile were estimated to cost a total of $100 M.
  • Other:  Shop equipment, pumps, and all other ancillary capital was estimated to be $56 M.   

Mobile Equipment

Mobile equipment costs were provided by a major equipment manufacturer. Equipment costs not supplied by this manufacturer were taken from the 2007 Infomine Mine Cost Service List and were increased by 20% to adjust for inflation. All purchase costs had 15% added to cover the costs of capital spares and transportation to the mine site.

Mobile equipment was purchased the year it is required in the mine. The lifespan of loading and haulage equipment was based on an operating life of 15,000 hours for an LHD and 25,000 hours for a haul truck. All remaining equipment was assumed to have a 5 or 10 year lifespan based on expected utilization. Replacements were scheduled over the entire life of the mine. It was assumed that no equipment will be rebuilt and that there is no salvage value at the end of equipment life. The total LOM mining equipment capital was estimated to be $1,001M.

It was assumed that all equipment required constructing the ventilation raises would be supplied by the contractor doing the construction.

Surface Earthworks

Major earthworks operations on surface will include the construction of a tailings impoundment facility, portals for the underground declines and casings for the collars of the ventilation raises. The tailings impoundment will be built using a combination mine waste rock and rock quarried from the local area. The topography of the Pampa de Pongo property does not provide a valley-fill tailings option. The impoundment will be built on relatively flat ground with a long perimeter waste rock dam. The dam construction has been costed in three phases over the life of the mine and would be built using a mining/construction contractor. The total surface earthworks were estimated to be $188M over the mine life for tailings and $8M for other mine surface excavations.

Surface Buildings

Mine related shop, offices, a dry, warehouse, explosive and cap magazines and a core storage facility were included in pre-production capital for a total cost of $14 M.

18.8.2

Mineral Processing

Beneficiation

The published capital cost of a comparable Peruvian crushing and grinding plant was $184.8 (1.3 times the capacity required for Pampa de Pongo). The entire total installed cost of the concentrator was calculated with a 0.77 power economy of scale and an escalation factor of 12%, between 2006/2007 and current 2008 dollars. The Pampa de Pongo crushing and grinding circuit is therefore estimated to be $148 million and has been escalated to $166 million.

Each of the fifteen concentrator lines (including rougher and cleaner magnetic separators and a vacuum disk filter) has been estimated at $1 million, totalling $15 million. It is estimated that the concentrate slurry pipeline will cost $1.2 million per km and will cost approximately $40.8 million. The smaller pipeline that will return water back to the mine has been estimated at approximately $0.6 million per km and will cost approximately $20.4 million. This will total $76.2 million in capital expenditures.

A capital cost of $35 million has been estimated as a requirement for construction of flotation circuits for copper-gold recovery.

Pelletization

The estimated total installed cost of a South American pellet plant is $66 per annual ton based on an extensive and detailed economic feasibility study conducted by CAEMI of the Itabiritos plant in Brazil. Therefore, a 15 million tonne per year pellet plant will cost approximately US$990 million with no economies of scale factored into the estimate. The addition of a pellet stacker-reclaimer, estimated at $20 million, brings total CAPEX to $1,010 million.

18.8.3

CAPEX Summary

A summary of the capital expenses by phase are shown in Table 18.19.

Table 18.19 : CAPEX Estimate Summary

Description

Unit

Pre-construction

Construction

Post Start-up

Total

Mine Development

M$

136

396

656

1,188

Mine Mobile Equipment

M$

1

111

889

1,001

Mine Construction

M$

2

255

 

257

Crush/Grind

M$

 

168

 

168

Mag Sep, Float, Slurry, Filter

M$

 

111

 

111

Pellet plant

M$

 

1,010

 

1,010

Tailings Dam

M$

 

59

129

188

Other

M$

95

40

 

135

EPCM

M$

 

354

 

354

Sustaining capital

M$

  

1,807

1,807

Capital cost w/o contingency

M$

234

2,504

3,482

6,219

Contingency

M$

46

501

696

1,243

TOTAL CAPITAL COST

M$

280

3,005

4,178

7,462

18.9

Economic Analysis

The economic analysis conducted in this report uses inferred mineral resources exclusively and only provides a preliminary overview of the project economics based on broad, factored assumptions. Inferred mineral resources are considered too speculative geologically to have the economic considerations applied to them to be categorized as mineral reserves, and there is no certainty that the inferred resources will be upgraded to a higher resource category. There is also no certainty that the results of this preliminary economic assessment will be realized.

A simplified cash flow model was made to determine the pre and post-tax NPVs at various discount rates for each case. Copies of the cash flow models are shown in Appendix B. No accommodation has been made for the deduction of interest expenses prior to taxation. The model assumes that the project would be 100% equity financed.

The economic analyses were performed on four different cases. Each case maintained the same basic parameters except for pellet prices which varied between 169 ¢/dmtu and 253 ¢/dmtu. Case 1 used a blast furnace pellet price of 198 ¢/mtu while the other cases used a range of direct reduction pellet prices. Sources for and reasoning behind the selected pellet prices are shown in the Table 18.20.

Table 18.20 : Iron Pellet Price Assumptions by Case

Case

Pellet Type

Pellet Price

(US¢/mtu)

Reference

1

Blast furnace

198

Independent market opinion for BF pellets

2

Direct reduction

169

3-year average (154 ¢/mtu)+ 10% DR pellet premium

3

Direct reduction

218

Independent market opinion (198 ¢/mtu)+ 10% DR pellet premium

4

Direct reduction

253

2008 public domain scoping study (230 ¢/mtu) + 10% DR pellet premium

The BF pellet price was determined from an iron pellet market study provided to Cardero by Global Strategic Solutions Inc. In Cases 2, 3 and 4, pellet quality was assumed to achieve DR standards and, as such, the unit pellet price was increased by a 10% premium above 64.5% Fe blast furnace pellet estimates.

The main assumptions used in the economic analysis are shown in Table 18.21 and the economic analysis results shown in Tables 18.22 and 18.23.

Table 18.21 : Main Economic Analysis Assumptions Common to all Cases

Item

Unit

Value

Copper Price

$/lb

2.00

Gold Price

$/oz

650

Iron recovery

%

93.4

Copper recovery

%

50

Gold recovery

%

50

Iron pellet grade

% Fe

64.5

Copper concentrate grade (Cu)

% Cu

22

Copper concentrate grade (Au)

 Au g/t

13.8

Payable iron (in pellets)

%

100

Payable copper (in Cu cons)

%

96.5

Payable gold (in Cu cons)

%

97

Offsite copper concentrate costs

  

Transport (all in)

$/wmt Cu concentrate

100

Treatment

$/dmt Cu concentrate

70

 Cu Refining

$/payable lb Cu

0.07

Au Refining

$/payable oz Au

6

Discount rate

%

10



Table 18.22 : IRR and Payback Period

Parameters

Unit

Case 1

Case 2

Case 3

Case 4

After tax IRR0%

%

18

15

20

23

Pre tax IRR0%

%

23

19

25

29

Payback Period (Post Tax, 10% DR)

Production years

7.6

10.0

6.8

5.7

Table 18.23 : NPV Results by Case

Taxation Assumption

Parameter

Unit

Net Present Value (“NPV”)

Case 1

BF Pellets

198 ¢/dmtu

Case 2

DR Pellets

169 ¢/dmtu

Case 3

DR Pellets

218 ¢/dmtu

Case 4

DR Pellets

253 ¢/dmtu

After Tax

0% discount rate

B$

         17.6

       13.7

       20.2

       24.9

8% discount rate

B$

            3.3

         2.2

         4.1

         5.4

10% discount rate

B$

            2.1

         1.3

         2.7

         3.7

12% discount rate

B$

            1.3

         0.6

         1.7

         2.5

Pre Tax

0% discount rate

B$

         27.3

       21.3

       31.4

       38.7

8% discount rate

B$

            5.8

         4.1

         7.0

         9.0

10% discount rate

B$

            4.0

         2.7

         4.9

         6.4

12% discount rate

B$

            2.7

         1.7

         3.4

         4.6

18.9.1

Sensitivity Analysis

Sensitivity analysis was done using metal prices, mill head grade, capital costs and operating costs as variables. Each variable was changed independently while all other variables were held at the constant.

The results of the sensitivity analysis are shown on graphs in Figures 18.11 to 18.14 and Table 18.24.

As with many mining projects, the net present value (“NPV”) of the project is most affected by the price of metal. This holds true for all the cases in this study. In Case 3, a 20% increase in metal prices leads to an increase in the post-tax net present value using a 10% discount rate (“PT-NPV10%”) from $2.690M to $3,967M, a 48% increase. The mill head grade and recovery also have a similar large impact with a $1,220M PT-NPV10% gain for a 20% increase in grade or recovery. Conversely, decreases in metal prices, mill head grades or recoveries have a correspondingly large negative effect on the NPV10%.

All cases indicate they are more sensitive to capital cost than operating cost. For Case 3, a 20% increase in capital cost drops the PT-NPV10% by $521M or 19%. An increase in Case 3 OPEX yields a $218M drop (8%) in PT-NPV10%.



Table 18.24 : Sensitivity Results by Case

Case

Variable

% Change

After-tax NPV10% (M$)

% Change

-20%

0%

20%

1

Capital Cost

25%

2,641

2,120

1,594

-25%

Operating Cost

10%

2,338

2,120

1,901

-10%

Metal Price

-55%

943

2,120

3,283

56%

Grade

-52%

1,001

2,120

3,226

53%

2

Capital Cost

42%

1,814

1,286

750

-41%

Operating Cost

17%

1,511

1,286

1,062

-17%

Metal Price

-78%

263

1,286

2,291

80%

Grade

-74%

322

1,286

2,234

75%

3

Capital Cost

19%

3,211

2,690

2,169

-19%