DRAFT REPORT ON GRANGE RESOURCES LIMITED · PDF fileGRANGE RESOURCES LIMITED SOUTHDOWN...

DRAFT

REPORT ON

GRANGE RESOURCES LIMITED SOUTHDOWN MAGNETITE SCOPING STUDY FOR PER

Submitted to:

Grange Resources Limited 221 St Georges Terrace

PERTH, WA 6000 DISTRIBUTION: 1 Copy - Grange Resources Limited 1 Copy - Golder Associates Pty Ltd November 2005 05641009-R08

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TABLE OF CONTENTS SECTION PAGE 1.0 INTRODUCTION......................................................................................... 1

1.1 Scope of Work....................................................................................... 1 1.2 Conventions .......................................................................................... 1

2.0 GEOLOGICAL BLOCK MODEL.................................................................. 3 3.0 OPEN PIT OPTIMISATION MODEL........................................................... 5 4.0 OPEN PIT OPTIMISATIONS ...................................................................... 6

4.1 Mining and Processing Costs ................................................................ 6 4.2 Slope Angles ......................................................................................... 7 4.3 Dilution and Mining Recovery ................................................................ 7 4.4 Process Recovery ................................................................................. 7 4.5 DTR Cut-off ........................................................................................... 7

5.0 OPTIMISATION RESULTS......................................................................... 9 6.0 PRELIMINARY MINE SCHEDULES......................................................... 12

6.1 Base Case Mine Schedule .................................................................. 13 6.2 Alternative Two-Stage Mine Schedule................................................. 17

7.0 DUMPING.................................................................................................. 20 7.1 All External Dumping ........................................................................... 20 7.2 Backfill Dumping.................................................................................. 21 7.3 Waste Material Characteristics ............................................................ 23

8.0 POTENTIAL FOR UNDERGROUND MINING ......................................... 25 9.0 CONCLUSIONS AND RECOMMENDATIONS ........................................ 26 LIST OF TABLES Table 2-1: Geological Block Model Parameters...........................................................................3 Table 2-2: Major Variables in Geological Models .......................................................................3 Table 2-3: Grade Tonnages within model ....................................................................................4 Table 2-4: Model at 10% DTR Cut-off.........................................................................................4 Table 3-1: Whittle Model Dimensions .........................................................................................5 Table 3-2: Resources within Whittle Model at 10% DTR Cut-off ...............................................5 Table 4-1 : Ore Processing Costs & Assumptions.........................................................................6 Table 4-2 : DTR Cut-offs...............................................................................................................8 Table 5-1 : Optimisation Cases......................................................................................................9 Table 5-2: Summary of Results ..................................................................................................10 Table 5-3: Grades within Case 4 Pit 13 ......................................................................................10 Table 6-1: Tonnages and Grades within Smoothed Whittle Pit..................................................13 Table 6-2: Single Stage Schedule Summary...............................................................................14 Table 6-3: Two-Stage Schedule Summary .................................................................................18 Table 7-1 : External Dump Capacity ...........................................................................................20 Table 7-2 : Backfill Dumping Scenario.......................................................................................22

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Table 7-3 : Waste Material Breakdown.......................................................................................24 LIST OF FIGURES Figure 5-1 : Section 638 500m E .................................................................................................11 Figure 5-2 : Section 640 000m E .................................................................................................11 Figure 6-1: Pit Shell for Preliminary Scheduling .......................................................................12 Figure 6-2: Likely Mining Sequence ..........................................................................................13 Figure 6-3: Long Section (looking North) showing strip layout ................................................13 Figure 6-4: Approximate Face Positions – End of Year 1...........................................................15 Figure 6-5: Approximate Face Positions – End of Year 2...........................................................15 Figure 6-6: Approximate Face Positions – End of Year 3...........................................................15 Figure 6-7: Approximate Face Positions – End of Year 4...........................................................16 Figure 6-8: Approximate Face Positions – End of Year 5...........................................................16 Figure 6-9: Approximate Face Positions – End of Year 10.........................................................16 Figure 6-10: Approximate Face Positions – End of Year 15.......................................................17 Figure 6-11: Two Stage Pit/Strip/Block Layout .........................................................................17 Figure 6-12: Two-Stage Schedule - Approximate Face Positions – End of Year 5 ...................18 Figure 6-13: Two-Stage Schedule - Approximate Face Positions – End of Year 10 .................19 Figure 6-14 Two-Stage Schedule - Approximate Face Positions – End of Year 15...................19 Figure 6-15 Two-Stage Schedule - Approximate Face Positions – End of Year 20...................19 Figure 7-1 : External Dump Layout.............................................................................................21 Figure 7-2 : Layout of Dumps for Backfill Case .........................................................................23 Figure 7-3 : Cross-Section on Line AA’......................................................................................23 Figure 7-4 : Waste ARD Containment.........................................................................................24 LIST OF APPENDICES Appendix A Domain – Oretype Definition Appendix B Four-X Optimisation Explanation Appendix C Four-X Parameter Files Appendix D Optimisation Results

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1.0 INTRODUCTION

Golder Associates (Golder) were requested by Alex Nutter of Grange Resources Limited (GRL) to carry out mine planning work to provide input to the Public Environmental Review (PER) prior to the full Bankable Feasibility Study (BFS). The level of this planning work is considered to be conceptual to prefeasibility.

1.1 Scope of Work

The scope of work includes:

• Optimise the deposit using provided prices and concentrate costs.

• Choose pit shells to act as ultimate and cutback pits.

• Schedule the shells to achieve 6.6 Mt per year concentrate production rates.

• Develop initial dump and pit layout.

1.2 Conventions

Currency

All costs and prices are in Australian dollars unless otherwise specified.

Grid references

Easting, Northing and RL references used in this report relate to the GDA/MGA Zone 50 Grid, unless specifically stated otherwise.

Block dimensions

Three-dimensional entities in this report are described in the format x by y by z, where x refers to the Easting or across strike distance in metres, y refers to the Northing or along strike distance in metres and z refers to the RL or vertical distance in metres.

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Abbreviations used in this report are:

ARD Acid rock drainage

BCM Bank cubic metre

BIF Banded Iron Formation

DTR Davis Tube Recovery

Golder Golder Associates Pty Ltd

Grange Grange Resources Limited

Insitu Material in deposit before mining

LCM Loose cubic metre

LOI Loss on Ignition

ROM Run of Mine – includes dilution and allowance for mining recovery

Note: All references to “ore” in this report mean mineralised material estimated within conceptual pits and are not reportable as Ore Reserves.

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2.0 GEOLOGICAL BLOCK MODEL

The block model (sd_geol_050901.bmf) was created by Golder. This block model is aimed at providing a starting point to develop conceptual mining plans for the PER. It was created from the data available at the end of August using Inverse Distance Squared grade estimation. No resource classification had been carried out at that time.

Table 2-1 summarises the basic block model parameters.

Table 2-1: Geological Block Model Parameters

X (Easting) Y (Northing) Z (RL)

Parent Block size (m) 25 25 10

Model origin 636 150 6 176 100 -400

Extent (m) 6 550 1 375 580

Number of Parent Blocks 262 55 58

Minimum Sub-cell Size 5 5 2

Table 2-2 gives the major variables in the model.

Table 2-2: Major Variables in Geological Models

Variable Values – Use

Domain See Appendix A

Sg Density

Dtr Davis Tube Recovery

Dtfe Fe grade of Davis Tube concentrate

Dts S grade of Davis Tube concentrate

Dtsio2 SiO2 grade of Davis Tube concentrate

Dtal2o3 Al2O3 grade of Davis Tube concentrate

Dtcao CaO grade of Davis Tube concentrate

Dtmgo MgO grade of Davis Tube concentrate

Dtp P grade of Davis Tube concentrate

Oretype See Appendix A

Oxide 0 – Fresh, 1 – Oxidised, 2 – Pallinup

The mineralisation occurs in the fresh rock. Table 2-3 gives the grade and tonnages by Dtr cut-off for the model. Table 2-4 gives the quantities of mineralisation in the model by oretype. The details of the oretype definition are attached in Appendix A.

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Table 2-3: Grade Tonnages within model

Cut-off (Dtr%)

Tonnage (Mt)

Dtr (%)

10 492 35.8 15 440 38.6 20 428 39.2 25 418 39.6 30 386 40.5 35 320 42.2

Table 2-4: Model at 10% DTR Cut-off

Oretype Tonnage (Mt)

Dtr (%)

Dtfe (%)

Dtsio2 (%)

Dtal2o3(%)

Dcao (%)

Dmgo (%)

Dtp (%)

Dts (%)

hgo1 402.1 39.86 68.71 2.18 1.42 0.19 0.25 0.004 0.54 hgo2 15.2 29.95 67.59 2.46 1.89 0.18 0.26 0.002 0.62 hsul 4.7 14.39 62.03 3.63 1.81 0.16 0.36 0.009 5.39 mark 37.0 14.20 66.29 2.47 1.67 0.44 0.37 0.001 1.73 pcom 14.7 18.40 66.09 2.66 1.87 0.44 0.46 0.006 0.56 qgoo 17.9 15.08 64.88 3.74 2.38 0.23 0.49 0.006 1.83 qpoo 0.6 15.78 66.49 3.00 1.78 0.35 0.26 0.001 1.27 Total 492.2 35.80 68.21 2.29 1.50 0.22 0.28 0.004 0.72

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3.0 OPEN PIT OPTIMISATION MODEL

The geological model was used to create a Whittle model (Grange1.mod). Mining costs were added to the geological model prior to being exported to Whittle. The Whittle model had the same dimensions as the geological model (Table 2-1).

Table 3-1: Whittle Model Dimensions

X (Easting) Y (Northing) Z (RL)

Block size (m) 25 25 10

Model origin 636 150 6 176 100 -200

Extent (m) 6 550 1 375 580

Number Blocks 262 55 58

Table 3-2 summarises the in situ mineralisation within the Four-X optimisation model. A cut-off of 10% DTR was applied to the model.

Table 3-2: Resources within Whittle Model at 10% DTR Cut-off

Model Tonnage (Mt) Con Fe% %DTR

Grange1 492.2 68.2 35.8

Appendix B provides a short explanation of the Whittle Four-X optimisation process.

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4.0 OPEN PIT OPTIMISATIONS

4.1 Mining and Processing Costs

The mining costs were estimated by Golder and the processing costs and pricing assumptions were provided by Grange and BurnVoir.

The mining costs are based on contractor mining. The contractor costs assumed for the optimisation are given in the formula below.

Waste Mining Cost ($/bcm) = (150-Bench RL)*0.06+3.42

Other mining costs which include supervision, administration, dewatering, grade control etc are estimated as $1.50/t ore for the purposes of the optimisation.

Table 4-1 : Ore Processing Costs & Assumptions

ROM Production 18.0 Mt/a Concentrate Production 6.6 Mt/a Pellet Production 6.8 Mt/a Exchange Rate 0.75 US$/$A $M/year Unit Cost Mining Costs – Supervision, GC & other 18.0 $1.00/t ROM Concentrator Processing Cost Concentrator Costs 50.6 $2.81/t ROM Royalty 11.6 $0.64/t ROM Transportation & Port 15.4 $2.33/t Conc Corporate Overhead 1.5 $0.23/t Conc Kemaman Plant Costs Plant Costs 55.2 $8.36/t Conc Concentrate Freight 52.8 $8.00/t Conc Concentrate unloading 13.2 $2.00/t Conc Pellet Loading 10.9 $1.65/t Conc Malaysia Overhead 1.3 $0.20/t Conc Total Cost 239.5 $36.29/t Conc Prices FOB Concentrate (A$/t Conc) 35.25 FOB Pellet (A$/t Conc) 79.14

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Two cases were developed based on production of concentrate or pellet. Each case has a different cost and pricing structure.

• Production of Concentrate

o Ore Cost ($/t ROM) = 4.45 + 2.56*DTR

• Production of Pellet

o Ore Cost ($/t ROM) = 4.45 + 22.77*DTR

4.2 Slope Angles

The overall angles for optimisation purposes were provided by Golder.

• South – 46°

• North – 43°

While these slopes are considered reasonable for this level of study, the geotechnical and hydrological study components of the BFS are still underway and slopes will be updated for the BFS.

4.3 Dilution and Mining Recovery

The geological model has no allowance for dilution or mining recovery. The ore will be grade controlled to create a diluted product of +10% DTR input to the plant.

4.4 Process Recovery

The process recovery uses the DTR in the model times 0.95 to allow for plant losses.

4.5 DTR Cut-off

The approximate marginal cut-offs for the two scenarios are given in Table 4-2. As part of the optimisation process a minimum of 10% DTR was applied to all cases.

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Table 4-2 : DTR Cut-offs

Scenario Cut-off DTR Concentrate 15%

Pellet 9%

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5.0 OPTIMISATION RESULTS

Table 5-1 summarises the cases run. The boundary constraint stops waste mining at the tenement boundary with Rio. Sample parameter files are attached in Appendix C.

Table 5-1 : Optimisation Cases

Case Pricing Boundary 1 Concentrate No 2 Pellet No 3 Concentrate Yes 4 Pellet Yes

The optimiser estimates a “best” case and a “worst” case discounted value. The best case requires that each shell be mined sequentially while the worst case mines the deposit on a bench by bench basis. The best case is generally impracticable as shell increments can be very small and therefore unmineable by themselves. The worst case is always achievable but gives much lower values. In practice, a compromise between the two cases is generally achieved by staging the pit using suitable cut-backs. An average of the best and worst discounted values (DAvg) has been calculated and used as a measure to compare optimisation results. A discount rate of 8% was used for this analysis.

The values returned by the optimiser do not include capital and are only used as a relative indicator of the sensitivity of the project to changes in costs, etc. Mine designs based on the shells will typically add 10% extra waste with some potential loss of ore. This is due to the requirement to take into account minimum mining width, access requirements and other practical mining constraints.

The results are summarised in Table 5-2 for the pits giving the highest undiscounted and discounted average cash flows. Full details are included in Appendix D.

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Table 5-2: Summary of Results

Case Rev Fact Pit Waste

(Mt) Ore (Mt)

Strip (t/t)

Con (Mt)

DTR %

Undisc CF ($M)

DAvg($M)

Highest Undiscounted Pits

Case1 1.0 18 699 308 2.3 112 38.3 897 418

Case2 1.0 20 1298 459 2.8 159 36.4 4160 1755

Case3 1.0 18 688 305 2.3 111 38.4 891 420

Case4 1.0 20 1241 447 2.8 154 36.4 4090 1758

Highest Average Value Discounted Pits

Case1 0.85 12 421 233 1.8 84 37.9 820 459

Case2 0.775 11 900 391 2.3 135 36.4 3897 1808

Case3 0.85 12 417 231 1.8 83 37.9 816 459

Case4 0.775 11 885 386 2.3 133 36.4 3856 1803

For the purposes of this study the effect of the boundary constraints are minimal. After discussions, Grange selected Pit 13 Case 4 as the Ultimate pit for PER purposes. The tonnages and grades of this shell as reported in Vulcan at 10% Dtr are given in Table 5-3.

Table 5-3: Grades within Case 4 Pit 13

ROM (Mt)

Conc (Mt)

Dtr (%)

Dtfe (%)

SiO2 (%)

Al2O3 (%)

CaO (%)

MgO (%)

Dtp (%)

Dts (%)

406 141 36.4 68.3 2.28 1.49 0.21 0.27 0.004 0.69

Two cross sections showing the ore body and pit outline are shown in Figure 5-1 and Figure 5-2.

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Figure 5-1 : Section 638 500m E

Figure 5-2 : Section 640 000m E

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6.0 PRELIMINARY MINE SCHEDULES

Preliminary life of mine schedules were developed for Southdown, based on the results of the pit optimisation process described in this report. The schedules are based on Whittle shells, and no detailed pit design work has been done at this stage. The main objectives of the preliminary mine scheduling process were to:

• Formulate mining method strategies and test the practicality of these strategies;

• Assess whether or not the nominated concentrate production rate of 6.6Mtpy is achievable;

• Estimate potential mining quantities for input to the mining equipment selection and cost estimation process; and

• Provide preliminary “ore” grade and tonnage profiles over the life of the mine for input to the treatment plant design process.

For the purposes of this study, it was decided to use a smoothed Whittle pit shell as the ultimate pit. As described in Section 5 the shell chosen was Case 4 pit 13. However as the smoothing process was likely to increase the pit size the next smaller shell (12) was used as the starting point.

The chosen shell was imported from Whittle into Vulcan and a smoothing process applied, in order to produce a less “jagged” pit outline and produce a better pit wireframe. The resultant pit shell is illustrated in Figure 6-1. Table 6-1 summarises the tonnes and grades within the smoothed pit shell, These are slightly different to the original Whittle shell.

Figure 6-1: Pit Shell for Preliminary Scheduling

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Table 6-1: Tonnages and Grades within Smoothed Whittle Pit

Waste (Mt)

ROM (Mt)

Conc(Mt)

Dtr (%)

Dtfe (%)

DtSiO2(%)

DtAl2O3 (%)

DtCaO (%)

DtMgO (%)

Dtp (%)

Dts (%)

1044 411 142 36.3 68.3 2.28 1.49 0.21 0.27 0.004 0.70

As is demonstrated in Table 5-3 the tonnages match very well with the originally selected Whittle Case 4 Pit 13.

6.1 Base Case Mine Schedule

The long, narrow Southdown pit will probably be mined in a series of blocks or “strips”, commencing at the Western end of the deposit and progressing to the East throughout the mine life. Figure 6-2 illustrates the likely mining sequence.

Figure 6-2: Likely Mining Sequence

The resultant pit wireframe was then spilt into a total of eleven strips – numbered 1 through 11, each strip being nominally 500m wide, with an endwall angle of 35° applied to the Eastern end of each strip. Each strip was subsequently divided into two “blocks” – designated “A” and “B” of nominal width 250m, in order to provide more flexibility in the scheduling process. Figure 6-3 shows a long-section along the pit’s East-West axis which demonstrates the strip design and strip/block nomenclature used.

Figure 6-3: Long Section (looking North) showing strip layout

Direction of MiningDirection of Mining

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A solid wireframe representing each of the 22 mining blocks (11 strips x 2 blocks per strip) was created in Vulcan software, and the Vulcan “Advanced Reserves” routine was then used to produce a spreadsheet containing the tonnes and grades of the different rock types by strip and block.

The block reserves spreadsheet was then processed into a conceptual mining schedule by assigning a mining period (year) to each rock parcel line in the spreadsheet. The results of the base-case schedule are summarised below in Table 6-2.

Table 6-2: Single Stage Schedule Summary

Note that for the purposes of this scheduling study, a recovery factor of 95% has been applied to the model DTR values to calculate recovered concentrate tonnage.

Concentrate Produced (kt) = Ore Mined (kt) * (%DTR/100) *0.95

Figures 6-4 through to 6-10 show a sequence of “end of period” maps for the base-case preliminary mine schedule.

% of DTR in Concentrate: 95%

Ore MinedWaste Mined

Total Mined Ore Mined

Waste Mined

Total Mined Strip ratio

Concentrate Produced

YEAR Mbcm Mbcm Mbcm Mt Mt Mt wst : ore %DTR %DTFE %DTS %DTSIO2 %DTAL2O3 %DTCAO %DTMGO %DTP kt1 5.2 19.9 25.2 18.4 52.6 71.0 2.9 37.8 69.1 0.37 1.94 1.54 0.18 0.26 0.011 6 5982 5.4 18.7 24.1 18.9 50.3 69.2 2.7 36.8 69.1 0.42 1.90 1.47 0.19 0.25 0.011 6 6073 5.7 18.6 24.2 19.8 50.5 70.3 2.6 35.1 68.7 0.49 2.00 1.59 0.18 0.28 0.008 6 6024 5.9 17.9 23.8 20.6 48.0 68.6 2.3 33.8 68.3 0.81 2.09 1.58 0.20 0.30 0.006 6 6035 5.6 19.1 24.7 19.7 50.2 69.9 2.5 35.3 68.3 0.83 2.12 1.54 0.20 0.28 0.006 6 6056 5.6 18.6 24.2 19.7 48.9 68.6 2.5 35.2 68.2 0.82 2.27 1.53 0.22 0.30 0.005 6 6057 5.7 17.5 23.1 19.8 48.9 68.6 2.5 35.2 67.9 0.76 2.53 1.51 0.24 0.33 0.005 6 6078 5.3 20.1 25.4 18.7 51.6 70.3 2.8 37.3 68.2 0.66 2.36 1.52 0.21 0.29 0.005 6 6179 5.8 17.1 22.9 20.2 48.8 69.0 2.4 34.5 67.9 0.68 2.55 1.47 0.23 0.31 0.004 6 60510 5.5 21.4 26.9 19.3 53.2 72.5 2.8 36.0 68.0 0.77 2.32 1.58 0.18 0.28 0.002 6 59711 5.5 18.8 24.3 19.4 53.4 72.8 2.8 35.8 67.8 0.94 2.37 1.60 0.18 0.27 0.002 6 60012 5.6 20.2 25.8 19.5 54.9 74.4 2.8 35.7 68.0 0.84 2.41 1.52 0.19 0.29 0.002 6 59813 5.1 19.9 25.0 18.1 54.0 72.1 3.0 38.5 68.4 0.63 2.22 1.47 0.19 0.27 0.002 6 61414 5.3 18.4 23.7 18.6 48.8 67.4 2.6 37.4 68.2 0.80 2.19 1.44 0.20 0.26 0.002 6 60215 5.3 17.6 22.9 18.6 46.9 65.5 2.5 37.5 68.0 0.82 2.32 1.44 0.21 0.25 0.002 6 60516 5.3 17.0 22.4 18.7 47.9 66.6 2.6 37.2 68.2 0.64 2.35 1.49 0.23 0.25 0.003 6 60817 5.2 16.8 22.0 18.3 47.0 65.3 2.6 38.0 68.6 0.54 2.15 1.38 0.23 0.23 0.002 6 59818 5.5 16.7 22.2 19.2 46.2 65.4 2.4 36.2 68.4 0.56 2.38 1.40 0.25 0.26 0.002 6 60919 5.2 16.6 21.8 18.4 45.4 63.7 2.5 37.8 68.5 0.57 2.35 1.38 0.24 0.25 0.002 6 60620 5.4 16.7 22.1 18.9 45.6 64.4 2.4 36.5 68.1 0.72 2.50 1.40 0.24 0.27 0.003 6 53321 5.0 14.7 19.7 17.7 43.1 60.8 2.4 37.1 68.2 0.76 2.49 1.43 0.24 0.26 0.002 6 22722 3.2 2.8 6.0 11.3 8.2 19.4 0.7 35.2 67.8 0.93 2.47 1.56 0.23 0.27 0.002 3 761

Total 117.4 385.0 502.4 411.5 1 044.3 1 455.7 2.5 36.3 68.3 0.70 2.28 1.49 0.21 0.27 0.004 142 006

Ore Grade

SOUTHDOWN DEPOSITPRELIMINARY LIFE OF MINE PRODUCTION SCHEDULEBASED ON PIT OPTIMISATION RUN 4, SEPTEMBER 2005

BLOCK STRIKE LENGTH 250 METRES

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Figure 6-4: Approximate Face Positions – End of Year 1



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6.2 Alternative Two-Stage Mine Schedule

A second life of mine Schedule case was also run, in an attempt to produce a more favourable strip ratio in the early years of the project. The ultimate pit was divided into two phases, with the South wall of Phases 1 being based on pit 5 of the Case 4 Whittle optimisation run. The North wall of Phase 1 was assumed to be the same as the ultimate pit. The Phase 2 pit was therefore assumed to be a South wall cutback to ultimate pit limits along the entire strike length of the pit, except for the Western extremity where the cutback width was too narrow to permit a two-phase approach. The conceptual phase/strip/block layout for this schedule is shown in Figure 6-11.

Figure 6-11: Two Stage Pit/Strip/Block Layout

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The results of the two-stage schedule are shown in Table 6-3.

Table 6-3: Two-Stage Schedule Summary

Comparison between Tables 6-2 and 6-3 shows a marginal improvement in the strip ratio in Years 3 and 4 of the two-stage schedule, however this is only a very small improvement over the base case schedule and the annual material movement profile is similar for both schedules in all other respects. This would indicate that there is probably not much benefit to be gained from adopting the two stage pit process, particularly when the extra access requirements are considered.

Figures 6-12 through to 6-15 show a sequence of “end of period” maps for the two-stage mine schedule.

Figure 6-12: Two-Stage Schedule - Approximate Face Positions – End of Year 5

% of DTR in Concentrate: 95%

Ore MinedWaste Mined

Total Mined Ore Mined

Waste Mined

Total Mined Strip ratio

Concentrate Produced

YEAR Mbcm Mbcm Mbcm Mt Mt Mt wst : ore %DTR %DTFE %DTS %DTSIO2 %DTAL2O3 %DTCAO %DTMGO %DTP kt1 5.3 20.1 25.4 18.6 53.7 72.3 2.9 37.4 69.1 0.35 1.95 1.51 0.18 0.26 0.011 6 5952 5.7 18.8 24.6 20.1 50.2 70.2 2.5 34.7 68.9 0.39 1.95 1.53 0.18 0.26 0.012 6 6123 5.8 14.0 19.8 20.1 38.3 58.5 1.9 34.5 68.6 0.72 2.00 1.55 0.19 0.27 0.008 6 5994 6.0 14.0 20.0 21.0 37.1 58.1 1.8 33.1 68.2 0.83 2.17 1.58 0.20 0.30 0.006 6 5975 5.6 19.2 24.9 19.7 51.4 71.1 2.6 35.2 68.3 0.83 2.24 1.53 0.21 0.30 0.007 6 5986 5.4 21.1 26.5 19.0 53.5 72.5 2.8 36.5 68.1 0.60 2.49 1.40 0.24 0.31 0.003 6 6057 5.7 18.6 24.3 19.8 53.0 72.8 2.7 35.1 67.8 0.77 2.50 1.48 0.22 0.29 0.002 6 5968 5.5 20.4 25.9 19.3 54.3 73.7 2.8 35.9 68.1 0.84 2.34 1.55 0.19 0.28 0.002 6 6009 5.4 20.0 25.4 18.8 52.4 71.2 2.8 36.9 68.3 0.69 2.28 1.49 0.21 0.28 0.002 6 60510 5.6 20.0 25.7 19.7 53.0 72.7 2.7 35.2 67.8 0.92 2.38 1.52 0.21 0.28 0.002 6 60011 5.5 20.2 25.7 19.2 54.5 73.7 2.8 36.2 68.1 0.78 2.34 1.45 0.22 0.25 0.003 6 60212 5.5 20.0 25.5 19.3 54.2 73.6 2.8 36.0 68.1 0.75 2.33 1.58 0.22 0.28 0.003 6 60313 5.3 20.8 26.1 18.7 56.5 75.1 3.0 37.2 68.3 0.60 2.31 1.48 0.22 0.27 0.005 6 59814 5.3 17.7 23.0 18.7 48.5 67.3 2.6 37.1 68.3 0.66 2.33 1.54 0.21 0.28 0.004 6 60615 5.4 17.1 22.5 18.9 46.2 65.1 2.4 36.7 68.0 0.80 2.31 1.57 0.19 0.28 0.002 6 60916 5.4 16.9 22.4 19.0 47.0 66.0 2.5 36.6 68.1 0.72 2.42 1.50 0.22 0.29 0.002 6 60617 5.3 16.4 21.7 18.6 46.3 64.9 2.5 37.4 68.3 0.63 2.38 1.42 0.23 0.27 0.002 6 60218 4.8 17.7 22.5 17.1 48.4 65.6 2.8 40.5 68.6 0.58 2.07 1.40 0.19 0.25 0.002 6 60119 5.1 18.1 23.2 17.9 50.9 68.8 2.8 38.8 68.5 0.61 2.23 1.37 0.23 0.24 0.002 6 59920 5.4 16.5 21.9 18.9 46.1 65.0 2.4 36.8 68.4 0.65 2.36 1.39 0.24 0.26 0.002 6 60221 5.3 15.2 20.5 18.8 42.2 60.9 2.2 36.9 68.2 0.71 2.47 1.42 0.24 0.26 0.002 6 57622 2.9 2.2 5.1 10.1 6.5 16.6 0.6 35.4 67.7 0.93 2.47 1.56 0.23 0.27 0.002 3 395

Total 117.4 385.0 502.4 411.5 1 044.3 1 455.7 2.5 36.3 68.3 0.70 2.28 1.49 0.21 0.27 0.004 142 006

Ore Grade

SOUTHDOWN DEPOSITPRELIMINARY LIFE OF MINE PRODUCTION SCHEDULEBASED ON PIT OPTIMISATION RUN 4, SEPTEMBER 2005

2-STAGE PIT APPROACH, BLOCK STRIKE LENGTH 250 METRES

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Figure 6-13: Two-Stage Schedule - Approximate Face Positions – End of Year 10

Figure 6-14 Two-Stage Schedule - Approximate Face Positions – End of Year 15

Figure 6-15 Two-Stage Schedule - Approximate Face Positions – End of Year 20

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7.0 DUMPING

Two different dumping scenarios were developed based on the single stage schedule (Table 6-2).

• All external dumping

• Mixed external and backfill dumping

7.1 All External Dumping

The direct waste produced is 385 Mbcm. Assuming a swell of 30% this material will require 500 Mlcm of dump space. Figure 7-1 shows a potential layout for this material. The overall slope of the outside face of the waste dump is 11°. Table 7-1 shows the capacity in the design by RL increment. This dump has capacity to contain all the mined waste plus the coarse rejects from the processing plant, if required.

Table 7-1 : External Dump Capacity

RL (From To)

Incr. Volume (Mm3)

Cum. Volume (Mm3)

200 to 210 48.2 517.9 190 to 200 53.9 469.7 180 to 190 60.1 415.8 170 to 180 66.6 355.7 160 to 170 73.6 289.1 150 to 160 80.8 215.5 Base to 150 134.7 134.7

The dump as shown in Figure 7-1 covers about 950 ha. Although it is shown as only a single dump, internally there would be a number of cells designed to encapsulate the potentially ARD forming material.

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Figure 7-1 : External Dump Layout

This method of placing all waste externally was rejected by Grange.

7.2 Backfill Dumping

In the case of backfill dumping, a portion of the waste is returned to the mine on a progressive basis. However not all the material is able to be placed into backfill for geometric, scheduling and economic reasons. Table 7-2 gives a dumping schedule showing the potential split between external and backfill dumping. It should be noted that this is a conceptual schedule to enable a preliminary estimate of the amount of material that might be able to go to backfill.

Tailings will be co-disposed with the waste in the backfill from Year 7 onwards. The waste from the pit will create basins approximately 1000 m long by 60 m deep for filling with plant tailings. The basins will be separated by 50 m wide (at the crest) walls. Approximately half the tailings will be able to be placed into the backfill. The tailings will be progressively covered by a layer of oxide/Pallinup to seal them from the atmosphere. This will allow a progressive rehabilitation.

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Table 7-2 : Backfill Dumping Scenario

YEAR Waste Mined

(Mbcm)

Cum Waste

(Mbcm)

Cum Waste

(Mlcm)

External (Mlcm)

External (%)

Backfill (Mlcm)

Backfill (%)

1 19.9 19.9 25.9 25.9 100% 0.0 0% 2 18.7 38.6 50.2 24.3 100% 0.0 0% 3 18.6 57.2 74.4 24.1 100% 0.0 0% 4 17.9 75.1 97.6 23.3 100% 0.0 0% 5 19.1 94.2 122.4 22.3 90% 2.5 10% 6 18.6 112.8 146.6 19.3 80% 4.8 20% 7 17.5 130.2 169.3 17.0 75% 5.7 25% 8 20.1 150.3 195.4 17.0 65% 9.1 35% 9 17.1 167.4 217.7 12.2 55% 10.0 45%

10 21.4 188.8 245.5 12.5 45% 15.3 55% 11 18.8 207.6 269.9 8.5 35% 15.9 65% 12 20.2 227.8 296.1 6.6 25% 19.7 75% 13 19.9 247.7 322.0 3.9 15% 22.0 85% 14 18.4 266.1 345.9 2.4 10% 21.5 90% 15 17.6 283.7 368.8 2.3 10% 20.6 90% 16 17.0 300.8 391.0 2.2 10% 19.9 90% 17 16.8 317.6 412.8 1.1 5% 20.8 95% 18 16.7 334.3 434.5 1.1 5% 20.6 95% 19 16.6 350.8 456.1 1.1 5% 20.5 95% 20 16.7 367.5 477.8 1.1 5% 20.6 95% 21 14.7 382.2 496.9 1.0 5% 18.1 95% 22 2.8 385.0 500.5 0.2 5% 3.4 95%

Total 385.0 229.4 46% 271.1 54%

It is assumed that backfill can be placed at an overall angle of 25° (e.g. 30 m benches with slope of 37° and a berm of 25m). The earliest that backfill could start is towards the end of Year 5 with an increasing proportion going to backfill after that time.

Figure 7-2 shows the potential layout at the end of mine life for this dumping scheme. Figure 7-3 gives a cross-section through the backfill on Section AA’.

The external dump covers an area 620 ha and the top of the external dump is at 180 mRL. This is a significant reduction in both extent and height from the external case (950 ha and 210 mRL).

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Figure 7-2 : Layout of Dumps for Backfill Case

Figure 7-3 : Cross-Section on Line AA’

7.3 Waste Material Characteristics

At this time there is no detailed information on the ARD or other characteristics of the waste material. The comments below are preliminary and for the mineralized material are based on the DTS grades of the ore which may not be representative of the in-situ grades.

Table 7-3 summarises the waste material breakdown.

tailings basins

mine waste backfill

tailings basins

mine waste backfill

tailings basins

mine waste backfill

A

A’

A

A’

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Table 7-3 : Waste Material Breakdown

Waste Material Tonnage (Mt)

%

Waste external mineralised zone 787 75.3 Waste within mineralised zone

Oxide 38 3.6 Pallinup 29 2.8

Mineralised waste (10% Dtr <) 89 8.5 Waste with no grades – assume 50% mineralised 51 4.9

Waste with no grades – assume 50% unmineralised 51 4.9 Total 1045 100.0

Of the waste, the material external to the mineralised zone should have little sulphur and minimal ARD problems. Within the mineralised zone the Pallinup and oxide material should be leached of sulphur and have minimal ARD problems. The likely ARD producers will be the mineralised material below cut-off that will go into the dump (89 Mt). For the purposes of this study it is assumed that 50% of the waste with no DTR grades will have an ARD potential (51Mt).

Therefore a total of 140 Mt of the waste (14%) is likely to have to be encapsulated within the external dump or base of the backfill (Figure 7-2).

For the first 5-6 years of operation the ARD material will be encapsulated within the external surface dump (approximately 38 Mt or 18 Mlcm). The area set aside for this will the the North West Corner of the external waste dump.

After year 6 ARD material will be able to be placed in the pit backfill which will ensure the control of any acid drainage.

The oxide and Pallinup material will be used as an encapsulation material for the ARD waste, which will be enclosed in cells and sealed as dumping progresses (Figure 7-4).

Figure 7-4 : Waste ARD Containment

Oxide Layer

Non-ARD Waste ARD Material

ARD Cells

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Approximately 15 Mlcm of oxide and Pallinup material will be used as the tailings storage embankment construction material and to cover the external tailing storage facility and tailings basins within the pit backfill.

8.0 POTENTIAL FOR UNDERGROUND MINING

A review was carried out by Golder to determine if there was potential for underground mining the material below the base of the open cut or deeper extensions of the deposit.

Technically there appears to be no reason why an underground mine could not be established below the pit. Assuming a large scale sub-level operation (i.e. 3-4 Mtpa), a total mining operating cost including development could be of the order of $15/t.

The calculations below estimate the average revenue and total costs for a typical tonne of ore assuming a 10% DTR cut-off.

Revenue per tonne of pellet - $79.14 x 36% (DTR) x 0.95 (Recovery)

= $27.07/t

Processing and Administration Cost for pellets - $4.45 + 22.77 x 36% (DTR)

= $12.65/t

Underground Mining Cost (incl. development) = $15/t

Total Pellet Cost = $27.65/t

Net Loss per tonne = $0.58/t

At a 10% DTR cut-off mining the material underground is not cost effective.

If the average grade could be increased to 40% DTR by increasing the cut-off to 25% DTR, this would give a net profit of $1.52/t. However, this selectively may not be achievable using caving techniques.

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9.0 CONCLUSIONS AND RECOMMENDATIONS

• The optimisation work has shown that a 400 Mt ROM pit is possible.

• Two mining methods were investigated – A single and two stage approach. Schedules were developed for each of the mining methods.

• It is recommended to use the single stage in future work.

• Two different dumping scenarios were developed based on an all external approach and mixed external and backfill dumping approach.

• The backfill scenario was selected by Grange.

• A co-disposal solution of placing the tailings into basins in the backfill was developed as a concept for the PER. Approximately 50% of the tailings and waste rock will be able to be placed in the excavated pit.

• Underground mining appears to be technically feasible but uneconomic as an option.

GOLDER ASSOCIATES PTY LTD Ross Bertinshaw Principal J:\Jobs405\MINING\05641009-Grange_Southdown_BFS\REPORTS\R08-Scoping Study for PER\05641009-R08.doc

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APPENDIX A

ORETYPE DEFINITION

Domain Oretype Triangulation Geology Description Comment201 hgo1 CZ_T1.00t Quartz-magnetite-clinopyroxene gneiss and minor granulite Mineralisation211 hgo2 CZ_T1g.00t Coarse grained quartz-magnetite-garnet-orthopyroxene gneiss Mineralisation

221 hsul CZ_T2.00tFine grained quartz-sulphide-garnet-orthopyroxene-biotite-sillimanite mylonite and gneiss Very high sulphur

231 pcom CZ_T3_1.00t Plagioclase-clinopyroxene-orthopyroxene-magnetite-quartz granulite and gneiss Dilution/internal waste232 pcom CZ_T3_2.00t Plagioclase-clinopyroxene-orthopyroxene-magnetite-quartz granulite and gneiss Dilution/internal waste241 mark CZ_T3a.00t Plagioclase-clinopyroxene-orthopyroxene-garnet-orthoclase-magnetite-quartz gneiss Outer marker.251 qgoo CZ_T4_1.00t Quartz-garnet-orthoclase-orthopyroxene-biotite-magnetite gneiss Dilution/internal waste High sulphur252 qgoo CZ_T4_2.00t Quartz-garnet-orthoclase-orthopyroxene-biotite-magnetite gneiss Dilution/internal waste High sulphur261 qpoo CZ_T5a.00t Quartz-plagioclase-orthoclase-orthopyroxene-garnet-epidote gneiss Dilution/internal waste301 hgo1 EZ_T1.00t Quartz-magnetite-clinopyroxene gneiss and minor granulite Mineralisation311 hgo2 EZ_T1g.00t Coarse grained quartz-magnetite-garnet-orthopyroxene gneiss Mineralisation321 hsul EZ_T2.00t gneiss Very high sulphur331 pcom EZ_T3_1.00t Plagioclase-clinopyroxene-orthopyroxene-magnetite-quartz granulite and gneiss Dilution/internal waste332 pcom EZ_T3_2.00t Plagioclase-clinopyroxene-orthopyroxene-magnetite-quartz granulite and gneiss Dilution/internal waste341 mark EZ_T3a.00t Plagioclase-clinopyroxene-orthopyroxene-garnet-orthoclase-magnetite-quartz gneiss Outer marker.351 qgoo EZ_T4.00t Quartz-garnet-orthoclase-orthopyroxene-biotite-magnetite gneiss Dilution/internal waste High sulphur361 qpoo EZ_T5a_1.00t Quartz-plagioclase-orthoclase-orthopyroxene-garnet-epidote gneiss Dilution/internal waste362 qpoo EZ_T5a_2.00t Quartz-plagioclase-orthoclase-orthopyroxene-garnet-epidote gneiss Dilution/internal waste401 hgo1 FEZ_T1.00t Quartz-magnetite-clinopyroxene gneiss and minor granulite Mineralisation411 hgo2 FEZ_T1g.00t Coarse grained quartz-magnetite-garnet-orthopyroxene gneiss Mineralisation421 hsul FEZ_T2.00t gneiss Very high sulphur431 pcom FEZ_T3_1.00t Plagioclase-clinopyroxene-orthopyroxene-magnetite-quartz granulite and gneiss Dilution/internal waste432 pcom FEZ_T3_2.00t Plagioclase-clinopyroxene-orthopyroxene-magnetite-quartz granulite and gneiss Dilution/internal waste Very high sulfur441 mark FEZ_T3a.00t Plagioclase-clinopyroxene-orthopyroxene-garnet-orthoclase-magnetite-quartz gneiss Outer marker.451 qgoo FEZ_T4.00t Quartz-garnet-orthoclase-orthopyroxene-biotite-magnetite gneiss Dilution/internal waste High sulphur461 qpoo FEZ_T5a_1.00t Quartz-plagioclase-orthoclase-orthopyroxene-garnet-epidote gneiss Dilution/internal waste462 qpoo FEZ_T5a_2.00t Quartz-plagioclase-orthoclase-orthopyroxene-garnet-epidote gneiss Dilution/internal waste

OXIDATION.00t Base of oxidation Above this surface is wastePALLINUP.00t Soil, laterite and Pallinup Formation

101 hgo1 WZ_T1.00t Quartz-magnetite-clinopyroxene gneiss and minor granulite Mineralisation111 hgo2 WZ_T1g.00t Coarse grained quartz-magnetite-garnet-orthopyroxene gneiss Mineralisation121 hsul WZ_T2.00t gneiss Very high sulphur131 pcom WZ_T3_1.00t Plagioclase-clinopyroxene-orthopyroxene-magnetite-quartz granulite and gneiss Dilution/internal waste132 pcom WZ_T3_2.00t Plagioclase-clinopyroxene-orthopyroxene-magnetite-quartz granulite and gneiss Dilution/internal waste133 pcom WZ_T3_3.00t Plagioclase-clinopyroxene-orthopyroxene-magnetite-quartz granulite and gneiss Dilution/internal waste141 mark WZ_T3a.00t Plagioclase-clinopyroxene-orthopyroxene-garnet-orthoclase-magnetite-quartz gneiss Outer marker.151 qgoo WZ_T4_1.00t Quartz-garnet-orthoclase-orthopyroxene-biotite-magnetite gneiss Dilution/internal waste High sulphur152 qgoo WZ_T4_2.00t Quartz-garnet-orthoclase-orthopyroxene-biotite-magnetite gneiss Dilution/internal waste High sulphur

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APPENDIX B

FOUR-X OPTIMISATION

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Four-X Optimisation

Introduction

Optimisation is the process of finding the pit shape that gives the highest value for the deposit. The Lerchs Grossman (LG) algorithm, first published in 1964, is designed to do just this.

The method works on a regular block model of an ore body, developing lists of blocks that should or should not be mined. The final list defines the outline of a pit that gives the highest value considering parameters such as slope angles, mining and processing costs, product prices and recoveries. The pit defined includes every block that is worth mining when waste stripping is taken into account.

The initial computer implementation of LG was the Whittle Programming “Three-D” program released in 1985. This allowed the optimisation of a single pit at a time.

In 1987 the Whittle “Four-D” program was introduced and allowed the calculation of a series of nested pits at a single run. Each pit was optimal for a set of cost/price conditions. The nested pits could then be used to simulate the scheduling of the mine. This then allowed the discounted value of the various optimum pits to be estimated.

In 1996 the “Four-X” program was introduced. This program is similar to Four-D but can handle multiple products within the same deposit.

Four-X Suite of Programs

The main programs in the Four-X suite of programs are:

FXUT - This is a utility program that allows the checking of tonnages and grade within models and result files.

FXST - This program generates the structure arcs that control the optimisation slopes.

FXOP - This program produces the nested pits.

FXAN - This program is used for analysing the nested pits by compiling the results and carrying out economic analyses for the nested pits.

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Input Data

Two main data input files are required to run the optimisation programs. These are a Whittle block model file and a parameter file.

Block Model

The block model is a specialised file created by a mine planning package or other method.

Each block in the model contains data on:

Location - Block position (IX, IY, IZ).

Rock - Tonnes of rock in the block (ore+waste).

CAF - Cost adjustment factor. The CAF times the reference mining cost gives the cost per tonne of mining material in that block.

PAF - Processing adjustment factor. The PAF times the reference processing cost is the cost per tonne of processing any ore that maybe in the block.

Blocks that have ore will also contain data on the ore parcels in the block:

ore parcel data - Name of ore type.

Tonnes of ore type.

Metal content (maybe multiple metals).

The data given above allows the optimisation program to calculate the costs of mining the block and processing the ore. It can then compare this block cost against the potential revenue from the ore in the block and determine if it is worth mining.

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Parameter File

This file provides input to the programs on aspects such as:

Block model - Origin, block sizes and number of block.

Cost - Reference mining cost.

Ore - Reference processing cost for each ore type.

Metal recovery for each processing method.

Price - Set of prices for each product for which optimum pits are to be found. Prices are controlled by a set of Revenue Factors. A base price of the product is provided (say $500/oz) and has a revenue factor of 1. A revenue factor of 0.8 will therefore represent a price of $400/oz.

Slope - Overall slope angles (slope angles including allowance for roads).

The initial output file that must be created is a “structure arc” file. This uses the slope data from the parameter file to control which blocks must be mined to allow another lower block to be excavated.

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APPENDIX C

FOUR-X PARAMETER FILES

* Grange1 - September 2005 - 6.6Mt conc/a 1 25.0 25.0 10.0 636150.0 6176100.0 -400.0 2 262 55 58 3 1 1 1 1 1 2.0 4 1 262 1 55 1 58 5 2 12 0.0 0.0 6 0.0 46.0 6 180.0 43.0 12 0 0 4 4 0 $ tonne 13 0 1.0 1.0 1 3 1 2 year 14 0.3 0.025 1.5 18 dtr 1 4 0 4 unit 18 dtfe 2 4 0 4 unit 18 dts 3 4 0 4 unit 19 EX0 I (4.45+2.56*dtr.G/100) 19 EX01 I (dtr.G/100) 20 dtr 0 0.3525 unit 1.0 20 dtfe 0 0.000 unit 1.0 20 dts 0 0.0 unit 1.0 21 hgo1 1 0 1.0 21 hgo2 1 0 1.0 21 pcom 1 0 1.0 21 qgoo 1 0 1.0 21 qpoo 1 0 1.0 21 mark 1 0 1.0 21 hsul 1 0 1.0 21 oxid 1 0 1.0 21 wast 1 0 1.0 25 mil1 hgo1 EX0 26 dtr C 0 0.95 0 10 26 dtfe N 0 1.0 0 0 26 dts N 0 1.0 0 0 25 mil2 hgo2 EX0 26 dtr C 0 0.95 0 10 26 dtfe N 0 1.0 0 0 26 dts N 0 1.0 0 0 25 mil3 pcom EX0 26 dtr C 0 0.95 0 10 26 dtfe N 0 1.0 0 0 26 dts N 0 1.0 0 0 25 mil4 qgoo EX0 0.95 0 10 26 dtfe N 0 1.0 0 0 26 dts N 0 1.0 0 0 25 mil5 qpoo EX0 26 dtr C 0 0.95 0 10 26 dtfe N 0 1.0 0 0 26 dts N 0 1.0 0 0 25 mil6 mark EX0 26 dtr C 0 0.95 0 10 26 dtfe N 0 1.0 0 0 26 dts N 0 1.0 0 0 25 mil7 hsul EX0 26 dtr C 0 0.95 0 10 26 dtfe N 0 1.0 0 0 26 dts N 0 1.0 0 0 35 mil1 mil2 mil3 mil4 mil5 mil6 mil7

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APPENDIX D

OPTIMISATION RESULTS

Strip Increm. Increm.Revenue Price Pit Bench Rock Waste Ratio Mined Conc DTFE DTR DTS Mining Process Mining Process Revenue Cashflow DB DW Davg Cost Strip

Factor $/t Conc mRL Mt Mt t/t Mt Mt % % % $/t rock $/t ore $ M $M $M $M $M $M $M $/t con Ratio0.575 0.35 1 70 4.7 2.6 1.23 2.1 0.9 69.3 42.4 0.14 1.15 5.53 5.4 11.7 30.0 12.9 12.7 12.7 12.7 20.12 1.2 0.600 0.35 2 40 11.2 6.3 1.26 5.0 2.0 69.4 42.1 0.16 1.19 5.53 13.3 27.4 69.9 29.2 28.5 28.5 28.5 20.86 1.3 0.625 0.35 3 30 25.2 14.0 1.26 11.1 4.3 69.4 40.8 0.26 1.21 5.49 30.4 61.0 151.8 60.4 57.4 57.4 57.4 21.82 1.3 0.650 0.35 4 10 71.1 40.1 1.29 31.0 11.7 68.8 39.6 0.46 1.23 5.46 87.6 169.4 410.9 153.8 138.1 137.9 138.0 22.53 1.3 0.675 0.35 5 0 118.6 68.0 1.35 50.5 18.7 68.8 39.0 0.44 1.24 5.45 147.3 275.4 660.8 238.1 204.3 201.7 203.0 23.37 1.4 0.700 0.35 6 -10 220.6 124.4 1.29 96.2 34.2 68.7 37.4 0.54 1.26 5.41 277.1 520.0 1,204.2 407.1 319.5 306.6 313.1 24.28 1.2 0.725 0.35 7 -50 298.0 172.9 1.38 125.1 44.6 68.7 37.5 0.53 1.28 5.41 381.4 676.8 1,570.6 512.4 382.2 355.4 368.8 25.12 1.7 0.750 0.35 8 -60 380.4 227.9 1.49 152.5 54.6 68.7 37.7 0.53 1.30 5.41 493.0 825.6 1,923.0 604.4 429.3 385.1 407.2 26.05 2.0 0.775 0.35 9 -70 465.8 285.4 1.58 180.4 64.5 68.6 37.7 0.53 1.31 5.41 609.8 976.9 2,275.3 688.6 468.2 403.7 435.9 26.82 2.1 0.800 0.35 10 -80 542.7 339.8 1.67 202.9 72.8 68.6 37.8 0.54 1.32 5.42 716.6 1,099.1 2,566.0 750.3 492.5 407.6 450.1 27.77 2.4 0.825 0.35 11 -90 611.7 389.9 1.76 221.8 79.7 68.6 37.8 0.55 1.33 5.42 812.5 1,202.0 2,811.1 796.5 510.7 403.6 457.2 28.60 2.7 0.850 0.35 12 -90 654.2 421.2 1.81 233.0 83.8 68.6 37.9 0.55 1.33 5.42 872.5 1,262.5 2,955.3 820.2 519.5 398.6 459.0 29.44 2.8 0.875 0.35 13 -100 702.2 457.8 1.87 244.4 88.2 68.6 38.0 0.55 1.34 5.42 941.3 1,325.2 3,107.8 841.3 526.5 389.4 458.0 30.39 3.2 0.900 0.35 14 -100 751.7 496.0 1.94 255.7 92.4 68.6 38.0 0.55 1.35 5.42 1,012.0 1,386.8 3,257.1 858.2 531.6 377.5 454.6 31.24 3.4 0.925 0.35 15 -110 806.7 539.0 2.01 267.6 96.9 68.6 38.1 0.55 1.35 5.43 1,092.1 1,452.0 3,416.1 871.9 535.9 362.5 449.2 32.20 3.6 0.950 0.35 16 -150 901.9 614.1 2.13 287.8 104.5 68.6 38.2 0.56 1.37 5.43 1,232.4 1,562.2 3,684.2 889.5 539.5 333.8 436.7 32.94 3.7 0.975 0.35 17 -160 959.7 660.3 2.21 299.4 108.9 68.6 38.3 0.56 1.37 5.43 1,317.4 1,625.5 3,838.1 895.1 540.3 313.4 426.9 33.97 4.0 1.000 0.35 18 -170 1,007.0 698.7 2.27 308.3 112.3 68.6 38.3 0.56 1.38 5.43 1,387.8 1,674.5 3,959.0 896.6 540.2 295.7 418.0 34.82 4.3 1.025 0.35 19 -180 1,051.3 734.6 2.32 316.6 115.4 68.6 38.4 0.56 1.38 5.43 1,453.8 1,720.1 4,069.2 895.2 539.8 278.5 409.2 35.69 4.3 1.050 0.35 20 -190 1,105.9 779.5 2.39 326.4 119.1 68.6 38.4 0.56 1.39 5.43 1,535.5 1,773.5 4,199.4 890.3 538.6 255.5 397.1 36.58 4.6 1.075 0.35 21 -190 1,151.4 817.1 2.44 334.3 122.1 68.6 38.4 0.56 1.39 5.43 1,603.7 1,816.8 4,304.5 883.8 537.0 236.0 386.5 37.40 4.7 1.100 0.35 22 -190 1,223.4 877.4 2.54 346.0 126.6 68.6 38.5 0.56 1.40 5.44 1,713.2 1,880.9 4,463.7 869.5 533.6 203.8 368.7 38.43 5.2 1.125 0.35 23 -200 1,262.1 909.7 2.58 352.3 129.0 68.6 38.5 0.56 1.40 5.44 1,772.0 1,915.4 4,547.8 860.4 531.4 186.4 358.9 39.10 5.2 1.150 0.35 24 -200 1,314.5 954.0 2.65 360.6 132.1 68.6 38.6 0.57 1.41 5.44 1,852.5 1,960.5 4,658.1 845.0 528.0 161.7 344.8 40.14 5.4 1.175 0.35 25 -200 1,341.1 976.6 2.68 364.5 133.7 68.6 38.6 0.57 1.41 5.44 1,893.2 1,982.2 4,711.8 836.2 526.2 149.1 337.7 41.00 5.7 1.200 0.35 26 -200 1,376.5 1,006.8 2.72 369.7 135.6 68.6 38.6 0.57 1.42 5.44 1,947.2 2,010.5 4,780.8 823.0 523.5 131.7 327.6 42.01 5.8 1.225 0.35 27 -210 1,407.7 1,033.5 2.76 374.2 137.3 68.6 38.6 0.57 1.42 5.44 1,995.5 2,035.2 4,841.1 810.2 521.0 116.2 318.6 42.72 5.9 1.250 0.35 28 -210 1,427.5 1,050.5 2.79 377.0 138.4 68.6 38.6 0.57 1.42 5.44 2,026.0 2,050.4 4,877.8 801.3 519.3 106.2 312.7 43.82 6.1 1.275 0.35 29 -210 1,450.7 1,070.6 2.82 380.1 139.6 68.6 38.7 0.57 1.42 5.44 2,062.3 2,067.6 4,920.0 790.0 517.0 94.8 305.9 44.70 6.4 1.300 0.35 30 -210 1,489.1 1,104.0 2.87 385.1 141.5 68.6 38.7 0.57 1.43 5.44 2,121.5 2,095.0 4,987.1 770.4 513.2 74.3 293.8 45.51 6.7

6.6Mt Con /year - Costs and Pricing to Concentrate - No Boundary ConstraintCostsOre

Grange Resources LimitedSouthdown Magnetite

September 2005 Optimisation

7/09/2005 11:08 AM grange1sso.xls Grange1


Factor $/t Conc mRL Mt Mt t/t Mt Mt % % % $/t rock $/t ore $ M $M $M $M $M $M $M $/t con Ratio0.525 0.79 1 70 4.7 2.5 1.09 2.3 0.9 69.2 41.4 0.19 1.15 13.87 5.4 31.3 70.2 33.5 33.1 33.1 33.1 41.41 1.1 0.550 0.79 2 20 44.2 23.3 1.11 20.9 7.6 68.5 38.5 0.59 1.23 13.21 54.1 275.8 603.7 273.9 253.1 253.1 253.1 43.48 1.1 0.575 0.79 3 0 185.8 98.2 1.12 87.6 30.0 68.5 36.0 0.63 1.25 12.65 231.4 1,107.9 2,371.1 1,031.8 833.5 826.7 830.1 45.20 1.1 0.600 0.79 4 -60 366.7 207.0 1.30 159.7 54.5 68.4 35.9 0.66 1.29 12.63 474.5 2,018.1 4,316.8 1,824.1 1,288.1 1,254.6 1,271.3 46.91 1.5 0.625 0.79 5 -80 543.1 323.9 1.48 219.2 75.0 68.4 36.0 0.66 1.32 12.65 717.1 2,772.9 5,935.8 2,445.8 1,562.3 1,487.3 1,524.8 48.75 2.0 0.650 0.79 6 -100 672.6 416.4 1.63 256.2 88.0 68.4 36.2 0.66 1.34 12.68 898.8 3,249.9 6,966.4 2,817.7 1,697.4 1,582.3 1,639.8 50.59 2.5 0.675 0.79 7 -110 777.8 494.6 1.75 283.2 97.4 68.3 36.2 0.67 1.35 12.70 1,049.9 3,595.6 7,711.3 3,065.7 1,776.4 1,625.5 1,700.9 52.78 2.9 0.700 0.79 8 -150 940.9 619.4 1.93 321.4 110.8 68.3 36.3 0.68 1.37 12.71 1,289.4 4,086.9 8,771.7 3,395.2 1,865.6 1,659.5 1,762.6 54.54 3.3 0.725 0.79 9 -180 1,051.8 707.0 2.05 344.8 119.1 68.3 36.4 0.68 1.38 12.73 1,454.5 4,389.6 9,427.2 3,582.9 1,910.4 1,663.9 1,787.1 56.48 3.7 0.750 0.79 10 -190 1,165.3 797.7 2.17 367.6 127.0 68.3 36.4 0.69 1.39 12.73 1,624.4 4,679.1 10,049.1 3,745.5 1,947.6 1,655.9 1,801.7 58.45 4.0 0.775 0.79 11 -200 1,290.3 899.8 2.30 390.5 135.0 68.3 36.4 0.69 1.41 12.74 1,815.2 4,974.5 10,687.1 3,897.3 1,978.5 1,637.2 1,807.8 60.31 4.4 0.800 0.79 12 -200 1,367.2 963.5 2.39 403.7 139.7 68.3 36.4 0.69 1.41 12.74 1,933.1 5,143.6 11,052.1 3,975.3 1,992.9 1,620.2 1,806.5 62.24 4.8 0.825 0.79 13 -210 1,426.6 1,013.3 2.45 413.4 143.0 68.3 36.4 0.69 1.42 12.74 2,024.6 5,267.1 11,317.5 4,025.7 2,002.1 1,604.7 1,803.4 64.10 5.1 0.850 0.79 14 -210 1,501.8 1,077.2 2.54 424.7 146.9 68.3 36.4 0.69 1.43 12.74 2,141.2 5,411.7 11,629.2 4,076.2 2,010.7 1,581.8 1,796.2 66.31 5.7 0.875 0.79 15 -220 1,569.1 1,134.7 2.61 434.4 150.3 68.3 36.4 0.69 1.43 12.74 2,246.8 5,536.1 11,896.3 4,113.4 2,016.7 1,560.6 1,788.6 68.15 5.9 0.900 0.79 16 -220 1,624.3 1,182.2 2.67 442.2 152.9 68.3 36.4 0.69 1.44 12.74 2,332.9 5,632.8 12,101.9 4,136.2 2,020.1 1,540.2 1,780.2 70.36 6.1 0.925 0.79 17 -220 1,661.3 1,214.3 2.72 447.0 154.6 68.3 36.4 0.69 1.44 12.74 2,391.0 5,693.9 12,232.6 4,147.7 2,021.9 1,526.5 1,774.2 72.20 6.6 0.950 0.79 18 -220 1,699.4 1,247.6 2.76 451.8 156.2 68.3 36.4 0.69 1.44 12.74 2,450.9 5,753.9 12,360.4 4,155.6 2,023.0 1,511.1 1,767.0 74.24 7.0 0.975 0.79 19 -230 1,727.0 1,271.8 2.79 455.2 157.3 68.3 36.4 0.69 1.45 12.73 2,494.7 5,795.9 12,449.2 4,158.6 2,023.4 1,499.9 1,761.7 76.47 7.1 1.000 0.79 20 -230 1,756.6 1,297.8 2.83 458.8 158.5 68.3 36.4 0.70 1.45 12.73 2,541.1 5,839.5 12,540.6 4,159.9 2,023.4 1,487.2 1,755.3 77.85 7.3 1.025 0.79 21 -230 1,779.5 1,318.3 2.86 461.2 159.3 68.3 36.4 0.70 1.45 12.73 2,577.0 5,870.5 12,606.9 4,159.3 2,023.3 1,476.5 1,749.9 79.93 8.3 1.050 0.79 22 -230 1,803.6 1,339.7 2.89 463.8 160.2 68.3 36.3 0.70 1.45 12.73 2,615.3 5,902.8 12,674.8 4,156.7 2,022.9 1,465.6 1,744.3 82.28 8.2 1.075 0.79 23 -230 1,820.8 1,355.1 2.91 465.7 160.7 68.3 36.3 0.70 1.45 12.72 2,642.4 5,924.9 12,721.1 4,153.8 2,022.5 1,457.3 1,739.9 84.10 8.5 1.100 0.79 24 -230 1,846.0 1,377.7 2.94 468.3 161.6 68.3 36.3 0.70 1.45 12.72 2,682.7 5,956.8 12,787.6 4,148.1 2,021.6 1,445.9 1,733.8 85.85 8.5 1.125 0.79 25 -230 1,855.8 1,386.5 2.95 469.3 161.9 68.3 36.3 0.70 1.45 12.72 2,698.1 5,968.4 12,811.8 4,145.3 2,021.2 1,440.9 1,731.1 88.52 9.2 1.150 0.79 26 -230 1,864.6 1,394.5 2.97 470.1 162.2 68.3 36.3 0.70 1.46 12.72 2,712.4 5,978.7 12,833.4 4,142.3 2,020.8 1,436.8 1,728.8 90.11 9.5 1.175 0.79 27 -230 1,882.9 1,411.2 2.99 471.7 162.7 68.3 36.3 0.70 1.46 12.72 2,741.3 5,998.6 12,875.5 4,135.5 2,019.8 1,427.3 1,723.6 91.73 10.2 1.200 0.79 28 -240 1,887.4 1,415.2 3.00 472.2 162.8 68.3 36.3 0.70 1.46 12.72 2,748.5 6,003.7 12,885.8 4,133.6 2,019.5 1,425.2 1,722.3 93.89 9.3 1.225 0.79 29 -240 1,893.9 1,421.1 3.01 472.7 163.0 68.3 36.3 0.70 1.46 12.71 2,758.9 6,010.6 12,900.1 4,130.6 2,019.0 1,421.8 1,720.4 96.11 10.1 1.250 0.79 30 -240 1,904.0 1,430.3 3.02 473.7 163.3 68.3 36.3 0.70 1.46 12.71 2,775.1 6,021.4 12,921.9 4,125.4 2,018.3 1,416.7 1,717.5 98.18 9.9

6.6Mt Con /year - Costs and Price to Pellet Stage - No Boundary ConstraintCostsOre

September 2005 OptimisationSouthdown Magnetite

Grange Resources Limited



Factor $/t Conc mRL Mt Mt t/t Mt Mt % % % $/t rock $/t ore $ M $M $M $M $M $M $M $/t con Ratio0.575 0.35 1 70 4.7 2.6 1.23 2.1 0.9 69.3 42.4 0.14 1.15 5.53 5.4 11.7 30.0 12.9 12.7 12.7 12.7 20.12 1.2 0.600 0.35 2 40 11.2 6.3 1.26 5.0 2.0 69.4 42.1 0.16 1.19 5.53 13.3 27.4 69.9 29.2 28.5 28.5 28.5 20.86 1.3 0.625 0.35 3 30 25.2 14.0 1.26 11.1 4.3 69.4 40.8 0.26 1.21 5.49 30.4 61.0 151.8 60.4 57.4 57.4 57.4 21.82 1.3 0.650 0.35 4 10 71.1 40.1 1.29 31.0 11.7 68.8 39.6 0.46 1.23 5.46 87.6 169.4 410.9 153.8 138.1 137.9 138.0 22.53 1.3 0.675 0.35 5 0 118.6 68.0 1.35 50.5 18.7 68.8 39.0 0.44 1.24 5.45 147.3 275.4 660.8 238.1 204.3 201.7 203.0 23.37 1.4 0.700 0.35 6 -10 220.6 124.4 1.29 96.2 34.2 68.7 37.4 0.54 1.26 5.41 277.1 520.0 1,204.2 407.1 319.5 306.6 313.1 24.28 1.2 0.725 0.35 7 -50 298.0 172.9 1.38 125.1 44.6 68.7 37.5 0.53 1.28 5.41 381.4 676.8 1,570.6 512.4 382.2 355.4 368.8 25.12 1.7 0.750 0.35 8 -60 380.4 227.9 1.49 152.5 54.6 68.7 37.7 0.53 1.30 5.41 493.0 825.6 1,923.0 604.4 429.3 385.1 407.2 26.05 2.0 0.775 0.35 9 -70 465.8 285.4 1.58 180.4 64.5 68.6 37.7 0.53 1.31 5.41 609.8 976.9 2,275.3 688.6 468.2 403.7 435.9 26.82 2.1 0.800 0.35 10 -80 542.7 339.8 1.67 202.9 72.8 68.6 37.8 0.54 1.32 5.42 716.6 1,099.1 2,566.0 750.3 492.5 407.6 450.1 27.77 2.4 0.825 0.35 11 -90 606.7 386.3 1.75 220.4 79.2 68.6 37.8 0.55 1.33 5.42 805.7 1,194.4 2,793.1 793.0 509.4 403.9 456.6 28.63 2.7 0.850 0.35 12 -90 648.6 417.2 1.80 231.4 83.3 68.6 37.9 0.55 1.33 5.42 864.8 1,254.1 2,935.4 816.4 518.2 399.2 458.7 29.44 2.8 0.875 0.35 13 -100 694.8 452.4 1.87 242.4 87.4 68.6 38.0 0.55 1.34 5.42 931.1 1,314.4 3,082.4 836.8 525.3 390.5 457.9 30.35 3.2 0.900 0.35 14 -100 744.5 490.7 1.93 253.8 91.7 68.6 38.0 0.55 1.35 5.42 1,002.2 1,376.3 3,232.4 853.8 530.6 378.9 454.7 31.26 3.4 0.925 0.35 15 -110 798.3 532.8 2.01 265.4 96.1 68.6 38.1 0.55 1.35 5.43 1,080.5 1,440.1 3,387.9 867.2 534.2 364.4 449.3 32.20 3.6 0.950 0.35 16 -150 893.5 608.0 2.13 285.6 103.7 68.6 38.2 0.56 1.37 5.43 1,220.9 1,550.3 3,656.1 884.8 537.9 335.8 436.8 32.94 3.7 0.975 0.35 17 -160 944.4 648.7 2.19 295.7 107.6 68.6 38.3 0.56 1.37 5.43 1,295.9 1,605.7 3,791.4 889.7 538.6 318.2 428.4 33.98 4.0 1.000 0.35 18 -170 992.4 687.6 2.26 304.8 111.0 68.6 38.4 0.56 1.38 5.43 1,367.3 1,655.5 3,914.3 891.4 538.7 300.5 419.6 34.75 4.3 1.025 0.35 19 -180 1,036.0 723.0 2.31 313.0 114.1 68.6 38.4 0.56 1.38 5.43 1,432.4 1,700.5 4,022.9 890.0 538.0 283.5 410.7 35.74 4.3 1.050 0.35 20 -190 1,091.1 768.2 2.38 322.9 117.9 68.6 38.4 0.56 1.39 5.43 1,514.8 1,754.4 4,154.4 885.1 536.4 260.5 398.5 36.55 4.6 1.075 0.35 21 -190 1,130.6 800.9 2.43 329.7 120.4 68.6 38.4 0.57 1.39 5.43 1,574.1 1,791.8 4,245.5 879.5 534.8 243.9 389.4 37.39 4.8 1.100 0.35 22 -190 1,203.4 861.8 2.52 341.6 125.0 68.6 38.5 0.57 1.40 5.44 1,684.9 1,856.8 4,406.7 864.9 531.1 211.3 371.2 38.46 5.2 1.125 0.35 23 -200 1,241.2 893.5 2.57 347.7 127.3 68.6 38.6 0.57 1.40 5.44 1,742.2 1,890.3 4,488.6 856.0 528.9 194.2 361.5 39.05 5.2 1.150 0.35 24 -200 1,291.1 935.6 2.63 355.5 130.3 68.6 38.6 0.57 1.41 5.44 1,819.0 1,933.3 4,593.8 841.3 525.4 171.3 348.3 40.16 5.4 1.175 0.35 25 -200 1,320.3 960.4 2.67 359.9 132.0 68.6 38.6 0.57 1.41 5.44 1,863.6 1,957.3 4,652.8 831.7 523.3 157.1 340.2 40.96 5.7 1.200 0.35 26 -200 1,349.3 985.2 2.71 364.1 133.6 68.6 38.6 0.57 1.41 5.44 1,908.1 1,980.2 4,709.3 821.0 521.0 142.9 332.0 42.05 6.0 1.225 0.35 27 -210 1,379.1 1,010.7 2.74 368.4 135.2 68.6 38.6 0.57 1.42 5.44 1,954.2 2,003.8 4,766.9 808.8 518.5 128.6 323.5 42.68 5.9 1.250 0.35 28 -210 1,393.9 1,023.4 2.76 370.5 136.0 68.6 38.6 0.57 1.42 5.44 1,977.2 2,015.0 4,794.4 802.1 517.1 121.2 319.2 43.79 6.2 1.275 0.35 29 -210 1,418.2 1,044.4 2.79 373.8 137.3 68.6 38.7 0.57 1.42 5.44 2,015.0 2,033.2 4,838.6 790.3 514.8 108.9 311.9 44.66 6.3 1.300 0.35 30 -210 1,445.7 1,068.4 2.83 377.3 138.6 68.6 38.7 0.57 1.42 5.44 2,057.7 2,052.4 4,886.5 776.4 512.1 94.6 303.4 45.55 6.9

6.6Mt Con /year - Costs and Pricing to Concentrate - Boundary ConstraintCostsOre





Factor $/t Conc mRL Mt Mt t/t Mt Mt % % % $/t rock $/t ore $ M $M $M $M $M $M $M $/t con Ratio0.525 0.79 1 70 4.7 2.5 1.09 2.3 0.9 69.2 41.4 0.19 1.15 13.87 5.4 31.3 70.2 33.5 33.1 33.1 33.1 41.41 1.1 0.550 0.79 2 20 44.2 23.3 1.11 20.9 7.6 68.5 38.5 0.59 1.23 13.21 54.1 275.8 603.7 273.9 253.1 253.1 253.1 43.48 1.1 0.575 0.79 3 0 185.8 98.2 1.12 87.6 30.0 68.5 36.0 0.63 1.25 12.65 231.4 1,107.9 2,371.1 1,031.8 833.5 826.7 830.1 45.20 1.1 0.600 0.79 4 -60 366.7 207.0 1.30 159.7 54.5 68.4 35.9 0.66 1.29 12.63 474.5 2,018.1 4,316.8 1,824.1 1,288.1 1,254.6 1,271.3 46.91 1.5 0.625 0.79 5 -80 543.1 323.9 1.48 219.2 75.0 68.4 36.0 0.66 1.32 12.65 717.1 2,772.9 5,935.8 2,445.8 1,562.3 1,487.3 1,524.8 48.75 2.0 0.650 0.79 6 -100 667.0 412.4 1.62 254.7 87.5 68.4 36.2 0.66 1.34 12.68 891.1 3,229.5 6,921.8 2,801.1 1,691.7 1,578.5 1,635.1 50.62 2.5 0.675 0.79 7 -110 768.6 487.8 1.74 280.8 96.6 68.3 36.2 0.67 1.35 12.69 1,037.1 3,563.8 7,641.9 3,040.9 1,769.4 1,621.5 1,695.5 52.78 2.9 0.700 0.79 8 -150 926.7 608.8 1.91 317.9 109.6 68.3 36.3 0.68 1.37 12.71 1,269.5 4,041.2 8,672.4 3,361.6 1,857.8 1,656.6 1,757.2 54.51 3.3 0.725 0.79 9 -180 1,036.6 695.5 2.04 341.1 117.8 68.3 36.3 0.68 1.38 12.73 1,433.1 4,341.4 9,322.5 3,547.9 1,903.2 1,661.9 1,782.6 56.46 3.7 0.750 0.79 10 -190 1,144.5 781.7 2.15 362.8 125.3 68.3 36.4 0.69 1.39 12.73 1,594.8 4,617.5 9,915.6 3,703.2 1,937.1 1,655.0 1,796.1 58.42 4.0 0.775 0.79 11 -200 1,270.5 884.5 2.29 386.0 133.4 68.3 36.4 0.69 1.41 12.73 1,787.1 4,915.4 10,558.6 3,856.0 1,968.2 1,637.1 1,802.6 60.33 4.4 0.800 0.79 12 -200 1,346.4 947.4 2.37 399.0 138.0 68.3 36.4 0.69 1.41 12.74 1,903.5 5,082.5 10,918.9 3,932.8 1,983.0 1,620.9 1,801.9 62.27 4.8 0.825 0.79 13 -210 1,394.5 987.7 2.43 406.8 140.7 68.3 36.4 0.69 1.42 12.74 1,978.0 5,182.5 11,134.1 3,973.6 1,990.5 1,609.3 1,799.9 64.18 5.1 0.850 0.79 14 -210 1,465.4 1,047.8 2.51 417.5 144.4 68.3 36.4 0.69 1.43 12.74 2,088.1 5,319.3 11,428.7 4,021.3 1,998.7 1,588.3 1,793.5 66.32 5.6 0.875 0.79 15 -220 1,519.3 1,094.1 2.57 425.2 147.1 68.3 36.4 0.69 1.43 12.74 2,172.8 5,417.6 11,641.0 4,050.5 2,003.6 1,571.6 1,787.6 68.23 6.1 0.900 0.79 16 -220 1,559.2 1,128.4 2.62 430.8 149.0 68.3 36.4 0.69 1.43 12.74 2,235.1 5,487.8 11,790.1 4,067.2 2,006.3 1,557.6 1,781.9 70.33 6.1 0.925 0.79 17 -220 1,598.2 1,162.3 2.67 436.0 150.7 68.3 36.4 0.69 1.44 12.74 2,296.4 5,552.7 11,928.5 4,079.4 2,008.1 1,543.0 1,775.5 72.20 6.6 0.950 0.79 18 -220 1,627.2 1,187.6 2.70 439.6 152.0 68.3 36.4 0.70 1.44 12.73 2,342.0 5,598.4 12,025.8 4,085.3 2,008.8 1,531.4 1,770.1 74.23 6.9 0.975 0.79 19 -230 1,657.5 1,214.1 2.74 443.4 153.2 68.3 36.4 0.70 1.44 12.73 2,390.0 5,644.7 12,123.6 4,088.8 2,009.4 1,519.6 1,764.5 76.29 7.1 1.000 0.79 20 -230 1,687.8 1,240.8 2.78 447.0 154.4 68.3 36.4 0.70 1.44 12.73 2,437.5 5,689.2 12,216.9 4,090.2 2,009.6 1,506.9 1,758.3 78.10 7.3 1.025 0.79 21 -230 1,707.1 1,258.1 2.80 449.1 155.1 68.3 36.3 0.70 1.45 12.73 2,467.9 5,715.3 12,272.7 4,089.5 2,009.5 1,498.1 1,753.8 80.03 8.3 1.050 0.79 22 -230 1,728.2 1,276.8 2.83 451.4 155.8 68.3 36.3 0.70 1.45 12.72 2,501.3 5,743.7 12,332.3 4,087.2 2,009.1 1,488.6 1,748.9 82.07 8.1 1.075 0.79 23 -230 1,746.3 1,292.9 2.85 453.4 156.5 68.3 36.3 0.70 1.45 12.72 2,530.2 5,767.5 12,381.8 4,084.1 2,008.6 1,480.3 1,744.4 84.19 8.1 1.100 0.79 24 -230 1,764.5 1,309.3 2.88 455.2 157.1 68.3 36.3 0.70 1.45 12.72 2,559.1 5,790.1 12,429.4 4,080.1 2,008.0 1,471.8 1,739.9 85.83 8.9 1.125 0.79 25 -230 1,771.2 1,315.3 2.88 455.9 157.3 68.3 36.3 0.70 1.45 12.72 2,569.6 5,798.3 12,446.1 4,078.2 2,007.6 1,468.6 1,738.1 88.21 8.7 1.150 0.79 26 -230 1,779.4 1,322.7 2.90 456.7 157.5 68.3 36.3 0.70 1.45 12.72 2,583.0 5,807.8 12,466.2 4,075.4 2,007.2 1,464.7 1,736.0 90.16 9.7 1.175 0.79 27 -230 1,792.0 1,334.1 2.91 457.9 157.9 68.3 36.3 0.70 1.45 12.72 2,602.8 5,821.8 12,495.3 4,070.7 2,006.5 1,458.4 1,732.4 91.85 9.8 1.200 0.79 28 -240 1,798.9 1,340.4 2.92 458.5 158.1 68.3 36.3 0.70 1.45 12.71 2,614.1 5,829.7 12,511.5 4,067.7 2,006.0 1,455.0 1,730.5 94.12 9.2 1.225 0.79 29 -240 1,805.0 1,345.9 2.93 459.1 158.3 68.3 36.3 0.70 1.45 12.71 2,623.6 5,836.2 12,524.8 4,064.8 2,005.5 1,451.8 1,728.7 95.24 9.6 1.250 0.79 30 -240 1,812.5 1,352.7 2.94 459.8 158.5 68.3 36.3 0.70 1.45 12.71 2,635.9 5,844.4 12,541.3 4,060.9 2,004.9 1,448.2 1,726.5 98.09 9.5

6.6Mt Con /year - Costs and Price to Pellet Stage - Boundary ConstraintCostsOre




DRAFT REPORT ON GRANGE RESOURCES LIMITED · PDF fileGRANGE RESOURCES LIMITED SOUTHDOWN...

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