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The economic viability of processing tailings to reduce environmental liability Sollner, Diana Donna Delia 2004

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T H E E C O N O M I C V I A B I L I T Y O F P R O C E S S I N G T A I L I N G S T O R E D U C E E N V I R O N M E N T A L L I A B I L I T Y b y D I A N A D O N N A D E L I A S O L L N E R , P . E n g . B . A . S c , T h e U n i v e r s i t y o f B r i t i s h C o l u m b i a , 1 9 9 3 A T H E S I S S U B M I T T E D I N P A R T I A L F U L F I L L M E N T O F T H E R E Q U I R E M E N T S F O R T H E D E G R E E O F M A S T E R O F A P P L I E D S C I E N C E i n T H E F A C U L T Y O F G R A D U A T E S T U D I E S ( D e p a r t m e n t o f M i n i n g a n d M i n e r a l P r o c e s s E n g i n e e r i n g ) W e a c c e p t t h i s t h e s i s a s c o n f o r m i n g t o t h e r e q u i r e d s t a n d a r d T H E U N I V E R S I T Y O F B R I T I S H C O L U M B I A A p r i l , 2 0 0 4 © D i a n a D o n n a D e l i a S o l l n e r , P . E n g . , 2 0 0 4 Library Authorization In presenting this thesis in partial fulfillment of the requirements for an advanced degree at the University of British Columbia, I agree that the Library shall make it freely available for reference and study. I further agree that permission for extensive copying of this thesis for scholarly purposes may be granted by the head of my department or by his or her representatives. It is understood that copying or publication of this thesis for financial gain shall not be allowed without my written permission. Title of Thesis: The Economic Viability of Processing Tailings to Reduce Environmental Liability " ~ Diana Sollner Name of Author (please print) Date (dd/mm/yyyy) Degree: Master of Applied Science Year: 2004 Department of Mining and Mineral Process Engineering The University of British Columbia Vancouver, BC Canada Page ii ABSTRACT Many mining operations produce tailings that contain sulphides. Sulphidic tailings are disposed into a tailings facility and measures are taken to control the release of deleterious seepage resulting from the oxidation of sulphide materials. Often this control needs to be maintained in the long term, which presents a long term environmental risk. This thesis examines the viability of processing tailings to reduce the long term environmental risk of tailings impoundments. A spreadsheet model was developed to calculate project life costs of two tailings disposal methods - conventional disposal and disposal of autoclaved tailings. A generic mine site in British Columbia and assumed operating parameters were used as a basis for the model. Unit operation designs for the two flowsheets were based on recently designed or constructed equipment or on accepted design methodology. Capital costs and unit operating costs were obtained from recently completed pre-feasibility studies. Monte Carlo simulations were run while varying selected parameters to derive project costs for a range of situations. The simulation results indicate that for smaller operations where the processing rate is less than 5000 tonnes per day, the mine life is less than 12 years and the sulphur content is less than 12% it may be economically viable to autoclave tailings in order to produce material that would be more geochemically stable in the long term. The combination of parameter values at which autoclaving tailings is economically viable follows a curve as sulphur content decreases in conjunction with processing rate and mine life increases. The economics of autoclaving tailings is sensitive to the amount of solids reporting to the autoclave. Page iii TABLE OF CONTENTS ABSTRACT ii TABLE OF CONTENTS iii LIST OF TABLES iv LIST OF FIGURES v 1. INTRODUCTION 1 1.1 Objective 1 1.2 Background 1 1.3 Methodology 9 1.4 Assumptions 10 1.5 Thesis Layout 10 2. THE MODEL 11 2.1 Input Parameters 13 2.2 Sulphide Separation 16 2.3 Sulphide Oxidation 22 2.4 Water Treatment 26 2.5 Tailings Facility 33 2.5.1 Tailings Characteristics 33 2.5.2 Facility Design 36 2.6 Tailings Facility Closure 40 2.7 Security and Bonding 42 2.8 Project Cost Summary 46 3. SENSITIVITY ANALYSES 50 3.1 Generic Simulations 50 3.2 Case Studies 62 4. DISCUSSION 64 4.1 Model Results 64 4.2 Model Construction 65 5. CONCLUSIONS 69 6. RECOMMENDATIONS FOR FURTHER STUDY 70 7. REFERENCES 73 APPENDIX A Example Calculation of an HDS Water Treatment Plant Design Using Unpublished Data by Humber (1996) 81 APPENDIX B Example Calculation of a Simple Lime Neutralization Water Treatment Plant Design Using Unpublished Data by Humber (1996) 90 APPENDIX C Dam Height Calculation Macro 98 APPENDIX D Simulation Results 102 Page iv LIST OF TABLES Table 1.1 Security Posted Over Time at Sixteen British Columbia Mines 3 Table 1.2 Base Case Parameters 10 Table 2.1 Summary of Flotation Test Results 17 Table 2.2 Sulphur Content of Desulphurized Tailings in the Model 18 Table 2.3 Assumptions for Flotation 19 Table 2.4 Retention Times for the Autoclave Circuit Components 24 Table 2.5 Look-Up Table for Sizing HDS Water Treatment Plants 28 Table 2.6 Look-Up Table for Sizing Simple Lime Neutralization Water Treatment Plants 32 Table 2.7 Calculated Acid Base Account for Flotation Tailings - Base Case 35 Table 2.8 Assumptions for Tailings Management Facility Design 37 Table 2.9 Sources for Cost Data 47 Table 2.10 Project Cost Summary Sheet - Conventional Case 48 Table 2.11 Project Cost Summary Sheet - Autoclaved Tailings Alternative 49 Table 3.1 Scenario Definitions 62 Table 3.2 Total Cost of the Autoclaved and Conventional Alternatives in Defined Scenarios 62 Page v LIST OF FIGURES Figure 1.1 Security Values over Time at Selected British Columbia Mines 5 Figure 2.1 Flowsheet for Conventional Tailings Disposal 12 Figure 2.2 Flowsheet for Processing Tailings 14 Figure 2.3 Input Parameters Sheet 15 Figure 2.4 Sulphide Separation Sheet 21 Figure 2.5 Sulphide Oxidation Sheet 25 Figure 2.6 Water Treatment Sheet - Conventional Tailings 29 Figure 2.7 Water Treatment Sheet - Autoclaved Tailings Alternative 31 Figure 2.8 Tailings Facility Sizing Sheet - Conventional Tailings 38 Figure 2.9 Tailings Facility Sizing Sheet - Autoclaved Tailings 39 Figure 2.10 Security and Bond Calculation Sheet - Conventional Tailings 44 Figure 2.11 Security and Bond Calculation Sheet - Autoclaved Tailings 45 Figure 3.1 Capital Cost ($millions) vs. Mi l l Tonnage 53 Figure 3.2 Total Cost (Smillions) vs. Mil l Tonnage 54 Figure 3.3 Total Cost ($millions) vs. Mine Life 55 Figure 3.4 Total Cost ($millions) vs. Sulphur Content 56 Figure 3.5 Total Cost ($millions) vs. Mil l Tonnage 58 Figure 3.6 Total Cost (Smillions) vs. Mine Life 59 Figure 3.7 Total Cost (Smillions) vs. Sulphur Content 60 Figure 3.8 Sulphur Content vs. Mil l Tonnage for Simulations where the Total Cost for the Autoclaved Alternative is Less Than the Conventional Alternative 61 Page 1 1. I N T R O D U C T I O N 1.1 Objective The idea of processing tailings as a means of improving the environmental performance of tailings typically elicits the response "it is too expensive!" This thesis calculates and compares the mine life cost of a conventional tailings disposal system and a selected processed tailings system to determine if there are situations in which processing tailings is an economically viable way of reducing the long term environmental liability of a tailings impoundment. 1.2 Background Many mining operations, particularly base metal mines, produce waste materials containing sulphides. Most of these sulphides will react to produce a low pH, metal enriched drainage that can contaminate watersheds if allowed to enter the receiving environment. This acidic drainage is one of the most significant environmental issues the mining industry must address (Tremblay, 2000). The Intergovernmental Working Group estimated that the liability for acidic drainage at mine sites in Canada is approximately $5.2 billion (MEMPR, 1995). Mine sites that have tailings facilities that are currently producing acidic drainage include Mt. Nansen (Yukon), Duthie (British Columbia), Faro (Yukon), Kam Kotia (Ontario), and Poirier (Quebec). To date, only Faro is treating water. Tailings facilities at numerous mines, such as Heath Steele, could produce acidic drainage if current control strategies were not maintained. Seepage from the tailings dam at the Poirier site contains 38,600 mg/L S04, 20 mg/L Zn, 1.4 mg/L Cu and 17,300 mg/L Fe at pH 3.2 (Lewis et. ai, 2000). It is reported that surface water Page 2 quality is affected for more than 21 km downstream from the Poirier site. Clearly, acid rock drainage control in tailings facilities is necessary to protect the environment. The financial consequences of acidic drainage and the growing importance of adequate reclamation funding can be seen in the security bonds levied against British Columbia mines. In a security policy discussion paper (MEMPR, 1995), the Government states that its approach to reclamation is to set broad objectives and then negotiate mine-specific requirements. This approach enables the Government to address a property's unique features. This flexibility is particularly important when acidic drainage is an issue at a mine site. The policy discussion paper also states that the Province will be requiring full security prior to a mine's closure to provide reasonable assurance that all reasonably foreseeable reclamation activities are covered. This policy is reflected in the security bonds posted for sixteen British Columbia mines over time (see Table 1.1). The change in security values over time are presented in Figure 1.1 for selected mines. Table 1.1 Security Posted Over Time at Sixteen British Columbia Mines (In SMillions) Year Permit # Kemess M-206 Huckleberry M-203 Gibraltar M-40 Myra Falls M-26 Sullivan M-74 Island Copper M-9 Highland Valley Copper M - l l Table Mountain M-127 1970 0.11 1971 0.1 1976 1978 0.335 1980 0.2 1983 1984 0.65 1989 1990 1991 1992 1993 0.0035 1994 14 0.11 1995 4.11 1996 6.11 1997 12 2 10.8 8.11 1998 9.11 10.25 1999 29.5 10.11 4 2000 11.11 2001 12.11 2002 13.11 Current Value 12 2 29.5 10.8 13.11 4 10.25 0.11 Source: Reclamation permit. Table 1.1 (cont'd) Security Posted Over Time at Sixteen British Columbia Mines (In $Millions) Year Golden Bear Snip Eskay Creek Endako Mt. Polley Blackdome Equity Silver Premier Gold Permit # M-187 M-190 M-197 M-4 M-200 M-171 M-114 M-179 1970 1971 1976 0.125 1978 1980 1983 0.425 1984 1989 10 1990 21 1991 1 1992 37.5 1993 1994 3.7 1995 1.15 38.3 1996 22.7 1997 1.9 0.1 3 1998 25 1999 6 2000 1.545 2001 3.774 2002 Current Value 1.545 1 3.774 6 1.9 0.1 25 3 Source: Reclamation permit. Figure 1.1 Security Values over Time at Selected Brit ish Columbia Mines (Constant 2001 dollars) Page 6 Security is required before the construction of a new mine. The amount of security is based on the reclamation liability created by the mine design. Reclamation typically includes the decommissioning and demolition of all structures, burying foundations, decommissioning roads, resloping waste rock piles, restricting access to underground mine workings or open pits, stabilizing and covering tailings impoundments, and revegetating all disturbed areas where practical (pit walls generally are not required to be revegetated). Cost estimates for these physical works are predictable (the number, size and shape of structures and waste rock dumps are known and contractors will give cost estimates) and generally straightforward. The challenge in requiring full security is estimating the liability associated with acidic drainage. As a rule, mines do not go into production expecting an acidic drainage problem. Reclamation plans and security bonds are established with the understanding that potential sulphide oxidation will be mitigated and managed such that acidic drainage is not created. Difficulties arise when acidic drainage develops after operations begin. Costs for constructing a water treatment plant are estimated easily enough. The challenge lies in estimating the annual operating costs of the plant, namely predicting the amount of acidity that will be generated, which in turn dictates the amount of lime that will be consumed. It is with this issue that the Equity Silver Technical Committee had to grapple for several years (Equity Silver, 1996). In the end, lime consumption predictions had to be revised after several years of monitoring data was available. The uncertainty in estimating long term water treatment costs at the Equity Silver mine can be seen in the up and down security value shown in Figure 1.1. The initial security value Page 7 in Figure 1.1 reflects the policy of the day, namely a fixed dollar amount per disturbed hectare. In the mid 1980s acidic drainage developed at the site and the government realized that the posted security would be insufficient to address the issue should the company abandon the site. Negotiations began between the government and the company, with extensive discussion focused on the rate of acidity generation, and consequently lime consumption and long-term water treatment plant operating costs. Water treatment plant data at the time indicated an increasing rate of lime consumption. Therefore, increasing amounts of security were set for the mine, as shown in Figure 1.1. In the mid 1990s, data indicated a leveling off and decrease in lime consumption. The security value was decreased to reflect this change. The Mine Environment Neutral Drainage (MEND) program was initiated by the federal and provincial governments and the mining industry of Canada in 1988 to co-ordinate research in order to reduce the liabilities associated with acidic rock drainage. Research is in the mechanisms of sulphide oxidation, potential methods for treating existing acidic drainage sites and methods to prevent the formation of acidic drainage. Tremblay (2000) summarizes the results and observations from the MEND program. One of the results is that prevention is the best strategy. Once sulphides start to react and to produce contaminated runoff, the reaction is very difficult to stop. Prevention methods have focused on the isolation of sulphide minerals from oxygen and/or moisture. This has resulted in the development of engineered covers and subaqueous or underwater disposal. Another approach to prevention has been Page 8 desulphurization and separate disposal and management of the sulphide concentrate. These methods have shown to be effective when designed correctly; however, the sulphide material remains in the tailings impoundment and poses a risk to the environment if any of the remedial/containment measures fail. This thesis explores a slightly different interpretation of prevention - that is the prevention of sulphides from entering a tailings impoundment. Remove the primary material that produces acidic drainage and the potential for producing acidic drainage is removed, thereby reducing the long-term liability of the mine. One method of prevention is to sufficiently desulphurize tailings to produce a net neutral product for disposal. The sulphide concentrate can then be oxidized in an autoclave to produce hematite (Fe203) and sulphuric acid, which is neutralized to produce gypsum and metal hydroxide sludge. The end result is a material that is stable in the physico-chemical environment of a tailings impoundment. With the tailings being inert, the tailings impoundment will not need to be lined and only a simple cover will be required at closure. It is argued that this method of tailings disposal can be comparable in cost to a more conventional tailings disposal method, but have a lower environmental risk, when whole project costs are considered. There are a number of methods to oxidize or isolate a sulphide concentrate, including roasting, bio-oxidation and encapsulation (in cement, cement derivatives and bitumen). Autoclaving was arbitrarily selected for this thesis based on the reasonably well understood technology. Page 9 1.3 Methodology A spreadsheet model was developed to calculate scoping level designs and the associated costs of two tailings disposal alternatives - conventional tailings disposal and autoclaved tailings disposal. Analysis was made on a generic open pit mine located in British Columbia using assumed operating parameters. The effect of these assumptions on the cost estimate were evaluated by conducting simulations in which input parameters are randomly varied within reasonable lower and upper bound values. In this fashion the situations in which processing tailings are a reasonable alternative are better defined. Unit operation designs were based on the design parameters of equipment recently designed and constructed or on accepted design methodology. Estimated costs were obtained from recently completed pre-feasibility studies. For simplicity, capital expenditures were assumed to take place in one year at the beginning of the operation although in reality some of these expenditures may be staggered. Security deposits were also assumed to be made in one year at the beginning. It was assumed that the security deposit is a one time expense covering the cost of closure. In reality, the security deposit is returned to the proponent as reclamation and closure is completed. This assumption was made for the sake of simplicity. Post-closure bonds were assumed to be posted two years prior to closure. The evaluation of the viability of tailings processing was based on the comparison of the costs between a conventional tailings disposal method and the autoclaved tailings method. It was assumed that the two mine scenarios are identical in every way except Page 10 for the tailings disposal method. Therefore, only those items associated with tailings disposal are costed in the model. While it can be argued that some cost numbers may be imprecise, all items are costed to the same level of accuracy. The value of the model is not in the absolute numbers but in the comparison of the subtotal cost of the two tailings disposal methods. 1.4 Assumptions Initial design and costing were based on a number of assumed operating parameters, selected from the range published in Mining Sourcebook (1998). Table 1.2 summarizes these assumptions. Table 1.2 Base Case Parameters Parameter Unit Value Milling rate tonnes/day 2500 Mine life years 20 Operating days days/year 344 Sulphide in tailings % 5 Neutralizing potential of tailings kg C a C 0 3 eqVtonne 100 1.5 Thesis Layout This thesis is organized into five main parts. The model is described in Section 2. Section 3 presents the simulation results. Section 4 discusses the simulation results and the construction of the model. Conclusions and recommendations are given in Sections 5 and 6, respectively. Page 11 2. T H E M O D E L The model is an Excel workbook containing one worksheet for each unit operation. Calculations for conventional and processed tailings disposal are carried out simultaneously. Conventional tailings disposal, typical at existing mine sites, is simply pumping the tailings to an impoundment after the marketable minerals have been recovered. Frequently the tailings supernatant is treated, with the water recycled to the mill and/or discharged to the receiving environment. The flowsheet for this process is shown in Figure 2.1. The pertinent issue associated with this disposal alternative is the geochemical behaviour of the tailings material and its influence on the tailings management facility and closure design requirements to safeguard the receiving environment. Figure 2.1 Flowsheet for Conventional Tailings Disposal Ore Flotation tailings Marketable Concentrate Pond Supernatant 51 If Water treatment Sludge to sludge pond Receiving Environment Tailings Facility Page 13 Processed tailings disposal is any combination of unit operations implemented to modify the tailings product. For this thesis, unit operations were selected to achieve a neutral tailings material. Several methods are available for each unit operation. However, specific methods were selected for this thesis based on reliability and industry acceptance. The flowsheet selected is shown in Figure 2.2. The methods selected for this thesis were flotation for desulphurizing tailings, followed by autoclaving to oxidize the final sulphide concentrate, and direct lime water treatment of the autoclave discharge. The model comprises seven worksheets: input parameters, sulphide separation, sulphide oxidation, water treatment, tailings management area, security and bonding (includes tailings facility closure), and project cost summary. Each model component is discussed in the following sections. 2.1 Input Parameters The "Input Parameters" sheet is the base sheet to which all other calculations refer. The user can input the values for parameters relating to flotation cell efficiency, site characteristics of the tailings impoundment area and financial considerations (cost of capital, interest rate and US dollar exchange rate). The parameters mill throughput, mine life, sulphide content in tailings and neutralization potential of tailings are varied randomly in the simulation. The inputs worksheet is shown in Figure 2.3. Figure 2.2 Flowsheet for Processing Tailings Ore Flotation tailings Marketable Concentrate Desulphurized tailings Sulphide Concentrate 1 r Storage tank Feed tank Autoclave Receiving Environment Tailings Impoundment Facility t Flash tank Acid Recycle 1 Water treatment Page 15 Figure 2.3 Input Parameters Sheet Project Parameters Fiscal Parameters Mill tonnage Sulpide in tailings NP in tailings Mine life Operating days/yr 2500 tonnes per day US$ exchange 5% %Closure bonded 100 kg C a C 0 3 eq./tonne tailings Bond interest rate 20 years 344 Climate Data 1.52 $CDN/$US 30% 3% Cost of Capital rate 10% Avg precip 1 m/yr Flotation Parameters %S in flotation tails %S in flotation concentrate Mass pull 0.35% 48% (assumes 90% pyrite recovery) 10% Water Treatment Lime efficiency 80% Tailings Disposal Consolidated density Floor width Max dam elevation Dam crest width Freeboard TMA wall slope Max TMA length 50% 150 m 100 m 10 m 5 m 30 degrees 20000 m Conventional Tailings Upstream dam slope Downstream dam slope Autoclaved Tailings 20 degrees 20 degrees Note: upstream dam slope for conventional tailings is shallower to allow the installation of a liner. Upstream dam slope Downstream dam slope 27 degrees 20 degrees Page 16 2.2 Sulphide Separation Sulphide separation is carried out in the processed tailings alternative only. As discussed in Section 2.0 above, flotation was selected as the separation method for this model. On this worksheet, the overall amount of tailings solids and the amount of sulphides reporting to the concentrate and desulphurized tailings streams are calculated. Equipment sizing is also calculated. The recovery of sulphides in the flotation circuit was calculated based on a fixed sulphur content in the concentrate and on a calculated mass pull rate (the percentage of the feed rate to the flotation cell that reports to the concentrate), where the sulphur content in the desulphurized tailings is dependent on the sulphur content of the feed. Values for these parameters are discussed below. A number of researchers have looked at general flotation of sulphides at neutral pH (Humber, 1995; Leppinen et al, 1997; Ityokumbul et al, 2000; Hodgkinson et al, 1994, and Benzaazoua et al, 2000). A l l of the researchers reported success in achieving reasonable (greater than 90%) sulphide recoveries. Test methodologies employed laboratory scale batch flotation cells, xanthates and copper sulphate. Most studies looked at various collectors and collector concentrations. The sulphide content of materials tested ranged from 2.35% to 21.4%. The sulphide content in the desulphurized tailings ranged from 0.06% to 4.15%. When floating cyanidation residues, a wash and repulp with fresh water was found to be necessary to achieve reasonable sulphide Page 17 recoveries at neutral to alkaline pH (Hodgkinson et al, 1994). A summary of reported results are presented in Table 2.1. Table 2.1 Summary of Flotation Test Results Test Number Leppinen et al, 1997 9 10 12 13 % S t o t starting 4.54 4.54 4.54 4.54 % S t o t tailings 0.71 0.66 0.48 0.61 % Recovery 93 88.2 96 93.4 Humber, 1995 1 2 3 4 5 6 7 8 9 Sample Selbaie 3.2* % S t o t starting 23.6 23.6 23.6 23.6 23.6 23.6 23.6 23.6 23.6 % S t o t tailings 21.10 0.70 0.76 1.18 0.97 0.73 1.43 1.55 1.19 % Recovery 14.52 98.31 98.30 97.31 97.71 98.35 96.74 96.27 96.9 Sample Selbaie 3.3 1 2 3 4 5 %S t o, starting 3.47 3.47 3.47 3.47 3.47 %S t o t tailings 0.49 0.32 0.41 0.71 0.39 % Recovery 87.89 92.31 90.67 81.95 90.10 Hodgkinson et al, 1994 6 7 13 14 16 % S t o t starting - - - - -% S t o t tailings** 0.25 0.24 0.35 0.39 0.44 % Recovery 95.31 95.37 93.41 92.67 91.56 Benzaazoua et al, 2000 P M G % S t o t starting 2.9 16.2 24.2 %S t o t tailings** <0.3 1.8 1.4 % Recovery 90 95 96 % S t o t = percentage of total sulphur in a sample. * Humber (1995) noted that significant oxidation of the tailings occurred prior to testing and the sample was difficult to treat. ** Values are calculated using published data and assuming 5% S in the test feed material. The data presented above show that the sulphur content of the desulphurized tailings is distinctly different for feed materials containing low or high sulphur contents. Based on this data, feed dependent sulphur contents were chosen for the desulphurized tailings. These values are presented in Table 2.2. Page 18 Table 2.2 Sulphur Content of Desulphurized Tailings in the Model %S in Feed %S in Desulphurized Tailings 2-5 0.35 5-10 0.46 10-15 0.57 15-20 0.68 20-25 0.79 The selection of sulphur content categories for feed material is arbitrary. The sulphur contents of desulphurized tailings for feed materials containing 2% to 5% S and 20% to 25% S are an average of data presented in Table 2.1. Intervening values were determined by interpolation. The total mass of the sulphide concentrate was calculated by using a mass pull rate (the percentage of the feed rate to the notation cell that reports to the concentrate). The mass pull rate is calculated based on the mass and sulphur balances: Mass: F = C + T (1) Sulphur: / F = cC + rT (2) where F, C and T is the mass of feed, concentrate and tailings, respectively /, c and / is the %S in feed, concentrate and tailings, respectively By rearranging equations (1) and (2), the mass pull rate can be calculated by the following equation: Page 19 MassPullRate = — = ^ — - (3) F c-t In the model, the %S in feed is randomly varied and the %S in tailings is determined according to Table 2.2. It is assumed that flotation will be reasonably efficient and the concentrate will contain 90% pyrite. This results in 48% S in the concentrate. In calculating equipment size additional assumptions were made regarding the water content of the various streams around the flotation circuit. Values were selected based on flotation design principles discussed by Arbiter (1985). Assumptions for flotation used in this model are summarized in Table 2.3. The mass balance around the flotation circuit is given in Figure 2.4. Table 2.3 Assumptions for Flotation Parameter Value Source Mil l tonnage Sulphide concentration Pulp density Residual sulphide Mass pull rate Water recovery in froth 500 - 20,000 tonnes/day 2% - 25% 30% 0.7% 50% 20% Randomly varied Randomly varied Arbiter (1985) median of Table 2.1 Benzaazoua et al (2000) Fig. 11 in Arbiter (1985) The volume of flotation cells required for a given tonnage and separation was determined using the following equation from Arbiter (1985): N V = 24 Page 20 QTEX where: N V = total effective cell volume (m ) Q = dry ore throughput (tonnes/day) T = retention time (minutes) E = pulp expansion factor due to aeration f 1 ^ 1 - % pulp volume as air (unitless) X = pulp flow rate (m3 pulp/min/tonne dry ore/hr) and X = 0.5338 1 100 , - + 1 S P% where: S = specific gravity of ore P% = pulp density 0.5338 = a constant It was assumed that retention time is 12 minutes (from Benzaazoua et al, 2000), retention time scale up factor is 2 (Arbiter, 1985), percent pulp volume as air is 15% (Arbiter, 1985) and the specific gravity of ore is 2700 kg/m3 (Arbiter, 1985). The sulphide separation worksheet in the model is shown in Figure 2.4. Figure 2.4 Sulphide Separation Sheet Input Parameters Mill tonnage 2500 tonnes per day 104 tonnes per hour Sulphide concentration 5% NP 100 kg CaC03 eq./tonne tailings Pulp density 30% Specific gravity of ore 2700 kg/m3 (from example in Flotation (Arbiter. 1985)1 % air in pulp 15% Collector addition 70 g/tonne Frothier addition 16 L/tonne Flotation time 12 minutes Scale factor 2 Scaled flotation time 24 minutes Residual S 0.4% Mass pull 10% Water recovery in froth 20% (from Figure 11 in Flotation (Arbiter, 1985)1 Sizing Flotation Cells NV = QTEX/24 Where NV = total effective cell volume Q = dry ore throughput, tonnes per day T = circuit retention time E = pulp expansion factor due to aeration X = pulp flow rate (m3 pulp/min/tonne dry ore/hr) And X = 0.5338[1/S + 100/P% -1] Where S = specific gravity of dry ore P% = solid content in pulp by weight Therefore: Q = 2500 tonnes per day T = 24 minutes E= 1.1765 X = 1.2457 m 3 pulp/min/tonne dry ore/hr NV= 3664 m 3 Mass Balance Parameter Units Ore Concentrate Tails Totals Solids mass tonnes/day 2500 243 2257 2500 Mass of S tonnes/day 125 117 8 125 Mass of NP kg CaC0 3 250000 24344 225656 250000 eq./day Water mass tonnes/day 5833 1167 4667 Pulp density % 30 17 33 Calculated ABA %S 5 48 0.35 AP kg/tonne 156 1503 11 NP kg/tonne 100 100 100 NP:AP 0.64 0.07 9.14 Page 22 2.3 Sulphide Oxidation Sulphide oxidation can be achieved by pressure oxidation or by atmospheric oxidation using aeration or bacterial catalyst. Pressure oxidation using an autoclave was selected for this model because of its more rapid process rate which better fits with the overall processing rate of a beneficiation plant. The solids and water mass in the flotation circuit concentrate is used as the input to the autoclave sizing. The autoclave design employed for this model is based on the design parameters utilized for the pressure oxidation unit at the Miramar Con Mine, a gold mine in Yellowknife, Northwest Territories (Fluor Daniel, 1999). The mine constructed an autoclave to treat pyrite and arsenic trioxide sludge, pyrite being the source of iron to convert arsenic trioxide to iron arsenate. The autoclave product is hematite and iron arsenate. In the model only pyrite is assumed to be treated; therefore, the autoclave product contains only hematite. The oxidation of pyrite (the predominant sulphide mineral in most tailings) to hematite occurs by the following reaction: 4 H 2 0 + 2FeS 2 +7^0 2 -> Fe 2 0 3 + 4 H 2 S 0 4 (4) The reaction is exothermic, spontaneous and generates heat. This generated heat enables the reaction to be self-sustaining when there is an uninterrupted supply of reactants, Page 23 namely pyrite and oxygen. Water is added to the autoclave to control the reaction. The oxygen demand by the autoclave is driven by the reaction equation shown above, namely 15 moles of oxygen is required to oxidize 4 moles of sulphur. An oxygen efficiency of 80% was assumed in the model (W.E. Norquist, pers. comm.). The products of oxidation will be hematite and sulphuric acid. The autoclave circuit includes the following components (see Figure 2.2): A storage tank for acidifying the flotation concentrate to remove any carbonates in the feed material, A feed tank for decreasing the pulp density and smoothing out the feed rate, A four compartment autoclave, for carrying out the oxidation reaction, An oxygen plant for supplying oxygen to the autoclave (not shown in Figure 2.2), A flash tank for releasing pressure and heat from the autoclave product, and A scrubber for removing acid from the steam vented from the flash tank (not shown in Figure 2.2). Sizing of the components is based on the feed rate, the retention time for each component and the operating parameters of the autoclave. The solids feed rate for each component is calculated by mass balance. Retention times are the same as those used at the Con Mine (W.E. Norquist, pers. comm.) and are summarized in Table 2.4. Operating parameters of the autoclave and pulp densities at various points in the autoclave circuit are from Fluor Daniel (1999). Pulp density of the flotation concentrate Page 24 entering the storage tank is 52% (see Figure 2.4). Water is added to the feed tank to bring the pulp density down to 26% in preparation for oxidation. The autoclave product exits at 19% solids, due to the dissolution of sulphides. The slurry loses water in the form of steam in the flash tank, thereby increasing the pulp density to 23%. The mechanical efficiency of the autoclave was assumed to be 85% (W.E. Norquist, pers. comm.). Table 2.4 Retention Times for the Autoclave Circuit Components Component Retention Time (hrs) Storage tank 10 Feed tank 24 Autoclave 1.5 Detailed calculations for sizing equipment are shown in Figure 2.5. Page 25 Figure 2.5 Sulphide Oxidation Sheet Input Parameters Autoclave feed rate S feed rate Water feed rate Pulp density Specific gravity of slurry Storage retention time Feed retention time Autoclave retention time Oxygen efficiency m 3 per cu.ft. Storage Tank Sizing 243.4392 tonnes/day 117 tonnes/day 1166.667 tonnes/day 17% 1580 kg/m 10 hrs. 24 hrs. 1.5 hrs. 80% (SG of Con Mine Flotation concentrate at 50% solids for autoclave) 0.02832 Feed Tank Sizing Volume of slurry 892 m3/day Pulp density for autoclave 26% 37 m3/hr Solid feed rate 243.4392 tonnes/day Retention time 10 hrs water feed rate 1166.667 tonnes/day Tank volume 372 m 3 added water -474 tonnes/day 13131 ft3 Total slurry mass 936 tonnes/day Specific gravity of slurry 1240 kg/m3 Autoclave Sizing Volume of slurry 755 m3/day Total slurry mass 936 tonnes/day 31 m3/hr 39013 kg/hr Retention time 24 hrs Specific gravity of slurry 1240 kg/m 3 Tank Volume 755 m 3 Volume of slurry 31 m3/hr 26663 ft3 Retention time 1.5 hr Autoclave Oxygen Demand Effective autoclave vol. 47 m 3 Converting FeS 2 to F e 2 0 3 Mechanical efficiency 85% Total autoclave volume 56 m 1960 ft3 2 FeS 2 + 7.5 0 2 + 4 H 2 0 -> F e 2 0 3 + 4 H 2 S 0 4 Therefore, 15 moles of oxygen is required for 4 moles of sulphide Autoclave Mass Balance S feed rate 117 tonnes/day 4879252 g/hr In Out S molecular weight 32 g/mol Total Solids tonnes/day 243.4392 126 Moles of S 152477 mol/hr S tonnes/day 117 0 . Water tonnes/day 693 539 Oxygen needed 571787 mol/hr Pulp density % 26% 19% Oxygen efficiency 80% Oxygen required 714734 mol/hr Flash Tank Mass Balance O molecular weight 16 g/mol In Out 0 2 required 11435746 g/hr Total Solids tonnes/day 126 126 11.4 tonnes/hr S tonnes/day 0 0 274 tonnes/day Water tonnes/day 539 423 Pulp density % 19% 23% Page 26 2.4 Water Treatment Water treatment is a unit operation in the conventional and processed tailings alternatives. As indicated in Figures 2.1 and 2.2, water treatment would occur at different points in the process. The conditions at these two points would be distinct from each other, requiring different water treatment processes. Conventional Tailings Disposal Water treatment is almost always required for excess pond water prior to recirculation to the mill or discharge to the receiving environment. Tailings pond water frequently contains one or all of acidity, sulphate or elevated metal concentrations and need to be removed in order to meet water quality criteria for mill process water or for the environment. Volumes requiring treatment are determined by the amount of water discharged with the tailings solids and the consolidated tailings density, mine water pumped to the tailings facility, precipitation and background run-off entering the facility. Lime neutralization is applied extensively in the mining industry (Murdoch et al, 1995). Field application has shown the ability of lime neutralization to treat very acidic solutions, accommodate a wide range of flows, and with moderate capital and operating costs compared to other treatment technologies. Lime neutralization can achieve near complete precipitation of metals as metal hydroxides and sulphuric acid as gypsum by the following reactions: Page 27 M e 2 + + SO4 2" + Ca(OH)2 + 2H 2 0 -> Me(OH)2 + CaSCy2H 2 0 and (5) 2Me 3 + + 3S0 4 2" + 3Ca(OH)2 + 2H 2 0 -> 2Me(OH)3 + 3CaS0 4-2H 20 (6) Lime neutralization is typically implemented in one of three ways (Murdoch et al, 1995): 1. Lime is added to the effluent stream and mixed with the tailings discharge; 2. The effluent is aerated to oxidize iron and reacted with lime in a separate circuit; or 3. A modification of the second method. After the effluent is aerated and reacted with lime, the slurry is thickened and a portion of the underflow is recycled to the beginning of the water treatment circuit. The first method produces a voluminous sludge (2.5% solids) with questionable sludge stability. The second method produces a more chemically stable sludge, but the sludge density remains low. The third method, commonly known as the high density sludge (HDS) process, produces a chemically stable sludge with a thickened sludge density of about 20%. The sludge exhibits free draining properties and can rapidly achieve 40% to 50% solids density (Kuit, 1980, Kuyucak et al, 1991). Murdoch et al (1995) state that the improved sludge characteristics are a result of the formation of precipitates on the surfaces of the recycled particles. In other words, precipitates will preferentially form on existing solid surfaces rather than form new, smaller solid particles. The third method of neutralization was selected for this model. Page 28 The HDS water treatment plant was sized based on estimated flow and assumed solution chemistry typical of acidic drainage. Flow was estimated from the amount of process water that would be expressed during consolidation, an assumed precipitation rate of 1000 mm per year and run-off water. Run-off was assumed to be 20% of the process water and precipitation inflows, which is reasonable for a relatively wet climate. Tailings supernatant was assumed to contain sulphate and metals at a pH of 6 during operations. In the post-closure period, water quality was assumed to be 152 mg/L Fe and 3500 mg/L SO4 at a pH of 3.5. Example calculations are given in Figure 2.6. HDS water treatment design calculations for determining the size and cost of the plant was based on interpolating unpublished data by Humber (1996). In summary, tailings pond water would be neutralized to pH 9.3 with a residence time of 40 minutes. The size and cost of an HDS water treatment plant was calculated using Humber (1996) for five different scenarios. Results for these scenarios were placed in an Excel look-up table from which all other plant sizing and costing calculations were interpolated. An example calculation using the unpublished data by Humber (1996) is given in Appendix A. The look-up table in the model is given in Table 2.5. Table 2.5 Look-Up Table for Sizing HDS Water Treatment Plants Flow Capital Cost Operating Sludge (L/min) (US$) Unit Cost Volume ($/L/yr) (m3/L/yr) 1,200 1,000,000 235 3.054 4,000 2,300,000 206 3.054 5,000 2,600,000 198 3.054 20,000 6,700,000 169 3.054 35,000 9,400,000 158 3.054 Figure 2.6 Water Treatment Sheet - Conventional Tailings Input Parameters Water from processing Consolidated density Expressed water Infiltration Water (Operations) Precipitation 5,833 tonnes/day 50% 3,333 tonnes/day 3,333,333 L/day 1 m/yr 0.0027 m/day Tailings facility surface area 3,832,210 m Volume from direct precipitation 10,492 m3/day 10,492,018 Volume from run-off Total volume to treat (from tailings facility sizing sheet) L/day 2,765,070 L/day (assume 20% excess for run-on) 16,590,422 11,521 Infiltration Water (Post Closure) Infiltration through cover Tailings facility surface area Volume of infiltration thru cover HDS Plant Design flow pH (operations) pH (post closure) sulphate iron 1.00E-07 8.64E-05 3,832,210 331 331,103 230 12,700 6 3.5 3,500 152 L/day L/min cm/s m/day m m3/day L/day L/min L/min s.u. s.u. mg/L mg/L (typical rate through degraded liners) (from tailings facility sizing sheet) Canadian exchange rate capital cost operating cost (operations) 1.52 Cdn$/US$ 4,548,556 US$ 6,913,806 Cdn$ 206,375 US$/yr (1 order of magnitude less than cost at pH 3.5) 313,690 Cdn$/yr operating cost (post closure) 65,879 US$/yr 100,137 Cdn$/yr Sludge Pond 773,932 m 3 Page 30 Autoclaved Tailings Alternative The oxidized material produced by the autoclave will be comprised of insoluble material, predominately silicates, hematite and sulphuric acid. The calculated sulphuric acid concentration for the base case scenario is 98 g H2SO4 /L. This is consistent with the acid concentration at the Con Mine, where the flash slurry product contains 95 g H2SO4 /L (Fluor Daniel, 1999). The autoclave product will also contain dissolved metals, including an estimated 6 g/L of iron (Fluor Daniel, 1999). Clearly, this material needs to be neutralized prior to disposal to avoid affecting the receiving environment. As discussed above for conventional tailings disposal, the high density sludge water treatment system is the preferred treatment method due to the increased stability of the sludge and the lower sludge volumes. The existing solids in the autoclave discharge will allow the use of a simplified lime neutralization process and still achieve the sludge characteristics typical of an HDS process. In addition, the oxidized nature of the autoclave product makes aeration unnecessary. The lime neutralization water treatment plant was sized based on the flow and solution chemistry predicted for the autoclave discharge (Section 2.3). Example calculations are given in Figure 2.7. Page 31 Figure 2.7 Water Treatment Sheet - Autoclaved Tailings Alternative Input Parameters (Flash Tank Product) Pulp density of discharge Amount of solids Amount of liquid S oxidized Specific gravity of liquid Specific gravity of slurry Residence time Lime efficiency 23% 126 tonnes/day 423 tonnes/day 117 tonnes/day 1,040 kg/m3 1,220 kg/m3 30 minutes 80% (SG of flash tank product water at Con Mine (Fluor Daniel, 1999)) (SG of flash product slurry at Con Mine (Fluor Daniel, 1999) Sulphuric Acid Concentration in Flash Tank Product S oxidized S molecular wt Moles S 1 mole S = 1 mole H 2 S0 4 H 2 S0 4 molecular wt Amount of H 2 S0 4 Amount of liquid SG of water Volume of liquid 117 tonnes/day 117,102,037 g/day 32 g/mol 3,659,439 mol/day 98 g/mol 358,624,989 g/day 423 tonnes/day 1,040 kg/m3 406,687 L/day 282 L/min Concentration of H 2 S0 4 Volume of slurry Amount of slurry S.G. of slurry 882 g/L 549 tonnes/day 1,220 kg/m3 Volume of slurry 313 L/min Direct Lime Plant Design flow PH sulphate iron 300 L/min 1.0 s.u. 96 g/L 6 g/L Canadian exchange rate capital cost operating cost 1.52 Cdn$/US$ 1,465,078 US$ (includes 10% markup for 316 stainless steel) 2,226,919 Cdn$ 11,141,655 US$/yr 16,935,315 Cdn$ Sludge Pond 8,332,116 m 3 Page 32 The lime neutralization water treatment design calculations are interpolated from the unpublished data by Humber (1996), similar to the HDS plant design calculations described above. In the model, autoclave discharge slurry is neutralized to pH 9.3 with a residence time of 30 minutes. An example calculation using Humber (1996) is given in Appendix B. The look-up table in the model used for the lime neutralization water treatment plant sizing and costing is given in Table 2.6. The operating cost was calculated using the following equation: Operating cost(US$) = $40.41 * slurry volume(L/ j n ) * H 2 S 0 4 concentration j^^ Q (7) The unit cost of $40.41 was determined from Humber (1996). Table 2.6 Look-Up Table for Sizing Simple Lime Neutralization Water Treatment Plants Flow Capital Cost (L/min) (US$) 50 500,000 800 3,000,000 3,000 4,000,000 25,000 5,300,000 SENES (1994) presented HDS and conventional water treatment plant capital costs for various flow rates and acidity values. Comparing Humber (1996) to SENES (1994) for the HDS system at similar flow rates and acidity, the estimated capital costs are similar. For the conventional water treatment system, SENES (1994) presented data for acidity Page 33 levels that are two orders of magnitude less than the acidity levels predicted in this thesis. However, from the data presented in SENES (1994), capital cost approximately doubles as the acidity increases by two orders of magnitude. If SENES (1994) upper range capital costs are doubled, the Humber (1996) estimates are within the same order of magnitude. 2.5 Tailings Facility The design of a tailings facility is in part determined by the geochemical behaviour of the material that will be impounded. If the material has the potential to oxidize and produce acidic and metal laden drainage, the tailings facility will need to be lined to prevent the release of contaminated water into ground and surface water. However, lining will be unnecessary if the impounded material produces a neutral drainage. Therefore, an understanding of the tailings characteristics is required. 2.5.1 Tailings Characteristics Conventional Tailings Disposal The assumed tailings geochemical characteristics for the base case are 5% sulphide and the neutralization potential (NP) is 100 kg CaCC»3 equiv./tonne (see Table 1.2). In acid base account terms, the acidity potential (AP) is 156 kg CaCG*3 equiv./tonne for a net neutralization potential of the tailings material of -56 kg CaCC>3 equiv./tonne. The neutralization potential to acidity potential (NP:AP) ratio would be 0.64. In theory, each NP unit will neutralize each acidity potential (AP) unit to produce a net neutral drainage. At NP:AP ratios less than 1 material will eventually produce acidic drainage. At NP:AP ratios greater than 1, only neutral pH drainage should be produced. Page 34 However, the empirical nature of the acid base account test does not take into consideration all the factors affecting sulphide oxidation and carbonate dissolution. Data from mine sites indicate that one cannot be certain about the potential of a material to generate acidic drainage when the NP:AP ratio is between 1 and 3. Therefore, the criterion for classifying a material as non-acid generating is an NP:AP ratio greater than or equal to 3 (Price, 1997). The NP:AP ratio was calculated for 100 simulations where the AP and NP values were randomly varied. In all simulations the ratio was less than 3. Since all scenarios in this thesis result in a material that would be classified as acid generating, the tailings facility in the model for conventional tailings disposal was designed with a liner. Autoclaved Tailings Alternative In the processed tailings alternative two types of tailings would be produced - flotation tailings from the final sulphide separation step and neutralized autoclave tailings from the sulphide oxidation step. The flotation tailings will be comprised largely of gangue silicate material with a minor amount of sulphide. The mass balance on the flotation circuit for the base case, as discussed in Section 2.2, is presented in Table 2.7. The NP is expected to remain constant during the flotation tests (Catalan et al, 1999). The calculated acid base accounting (ABA) for the flotation tailings is also presented in Table 2.7. Page 35 Table 2.7 Calculated Acid Base Account for Flotation Tailings - Base Case Parameter Units Ore Concentrate Tails Mass Balance Solids mass tonnes/day 2500 243 2256 Mass of S tonnes/day 125 117 8 Mass of NP kg CaCOs eq./day 250000 24344 225656 Water mass tonnes/day 5833 1167 4667 Pulp density % . 30 17 33 Calculated ABA S % 5 48 0.35 AP* kg/tonne 156 1503 11 kg/tonne 100 100 100 NP:AP 0.64 0.07 9.14 * Acidity potential where A P = % S * 31.25 (Steffen Robertson and Kirsten, 1989) * * Neutralization potential. 100 kg C a C 0 3 eq./tonne tailings was assumed for the base case. Following the classifying criteria discussed in Price (1997) and discussed previously in this section under conventional tailings disposal, the flotation tailings produced under the base case scenario would be classified as non-acid generating. Therefore, a liner would not be required for the tailings facility. The sulphide concentrate from the flotation circuit would be the feed to the autoclave circuit. The presence of some sulphide in the flotation tailings dictates that a minimum amount of neutralization potential will be necessary in order to build an unlined tailings facility. This requirement means that processing tailings will only be a reasonable consideration if the neutralization potential of the ore meets this critical cut-off. The model does not include this cut-off; however, it must be taken into consideration when evaluating the model results. Page 36 The neutralized autoclave tailings, as discussed in Section 2.4, would contain silicate material that remained unaltered through the autoclaving process, hematite, metal hydroxides and gypsum. Silicate and hematite are all well known to be geochemically stable in oxidizing environments with low leaching rates. Studies completed by Zinck and Griffith (2000) demonstrate that high density water treatment sludges have low leachability at neutral to alkaline pH. The study results also show that metal concentrations in the treated effluent are low, indicating that lime precipitation is effective at metal removal from solution. With both the flotation tailings and the neutralized autoclave tailings likely to produce neutral, low metal drainage, it is expected that the combined tailings will not have a deleterious effect on the receiving environment. Therefore, an unlined tailings facility is used in the model for the autoclaved tailings alternative. 2.5.2 Facility Design Two basic designs were used in the model - one for acid generating materials and one for materials that would produce neutral, low metal drainage. The fundamental difference between the two designs is the inclusion of a liner in the facility containing acid generating materials to minimize leachate release during operation. The liner assumed for the model comprises of a compacted foundation layer (clay), an impermeable liner and a protective sand layer. Page 37 For the purposes of this model it was assumed that the tailings facility would be located in a mountain valley of sufficient length with a tailings dam at one end. Sizing of the tailings facility was based on several assumptions, summarized in Table 2.8. Table 2.8 Assumptions for Tailings Management Facility Design Parameter Units Conventional Autoclaved Tailings Tailings Angle of mountain sides degrees 30 30 Max dam height m 100 100 Freeboard m 5 5 Impoundment floor width m 150 150 Upstream dam slope degrees 20 27 Downstream dam slope degrees 20 20 Dam crest width m 10 10 Tailings solids density 50% 50% The dam slope angles were selected to meet the stability requirements at closure (see Section 2.6) and the practicalities of installing a liner in the conventional tailings alternative (CC. Scott, pers. comm.). The tailings facility size was determined by extending the length of the facility and increasing the height of the dam until sufficient volume was achieved to store the complete inventory of tailings when the tailings had achieved a consolidated solids density of 50%. The dam height was then increased an additional 5 m to allow for freeboard. The macro written to calculate the dam height and tailings facility length is given in Appendix C. Example calculations determining tailings facility size is given in Figures 2.8 and 2.9. Figure 2.8 Tailings Facility Sizing Sheet - Conventional Tailings Tailings Discharge to Tailings Facility Tailings Solids Water Pulp density Specific gravity of slurry Volume inert slurry Operating days/yr Mine life Water treatment sludge Total tailings volume Consolidated density Specific gravity Consolidated volume Tailings Facility Size Plan View of tailings floor 2500 tonnes/day 5833 tonnes/day 30% 1233 kg/m 3 6759 m3/day 344 days/yr 20 yrs 773932 m 3 4.73E+07 m 3 50% 1459 kg/m3 2.36E+07 m 3 Tailings Facility Assumptions Walls of TMA are Max elevation is Freeboard is Floor width is U/S dam slope is D/S dam slope is Dam crest width Max TMA length 30 degrees 100 m 5 m 150 m 20 degrees 20 degrees 10 m 20000 m Floor width 150 m Minimum dam height 8 m Freeboard 5 m Dam Height 13 m Impoundment Length 19649 m Impoundment floor surface area 3,975,026 m 2 Dam volume 89,149 m 3 Surface area for reclaim - top 3,832,210 m 2 Surface area for reclaim - dam 5,701 m 2 Figure 2.9 Tailings Facility Sizing Sheet - Autoclaved Tailings Tailings Discharge to Tailings Facility Inert Tailings Solids Water Pulp density Thicken to Solids Water Inert slurry 2256.561 tonnes/day 4667 tonnes/day 33% 30% 2256.561 tonnes/day 5265 tonnes/day 7522 tonnes/day Specific gravity of solids Specific gravity of slurry 2650 kg/mJ 1220 kg/m3 Volume inert slurry Operating days/yr Mine life Inert tailings volume Consolidated density Specific gravity Consolidated volume Combined Tailings Inert tailings 6165 m /day 344 days/yr 20 yrs 4.24E+07 m 3 50% 1459 kg/m3 2.13E+07 m 3 2.13E+07 m Water treatment sludge 8.33E+06 m Total tailings volume 2.96E+07 m 3 Tailings Facility Size Plan View of tailings floor Tailings Facility Assumptions Walls of TMA are Max elevation is Freeboard is Floor width is U/S dam slope is D/S dam slope is Dam crest width Max TMA length 30 degrees 100 m 5 m 150 m 27 degrees 20 degrees 10 m 20000 m Floor width 150 m Minimum dam height 10 m Freeboard 5 m Dam Height 15 m Impoundment Length 19743 m Impoundment floor surface area 4,153,230 m 2 Dam volume 101,983 m 3 Surface area for reclaim - top 3,987,326 m 2 Surface area for reclaim - dam 6,579 m 2 Page 40 2.6 Tailings Facility Closure The reclamation code in British Columbia (MEMPR, 1992) requires that disturbed land be reclaimed to the level of productivity that existed prior to mining operations, that structures maintain long term stability and that long term water quality is maintained to an acceptable standard. In British Columbia, the most common land use around mine sites is wilderness/forest use. To meet this, reclamation will involve covering the tailings impoundment and revegetating with grass to stabilize the soil surface and planting trees and shrubs to restore a more typical vegetation cover. The primary structure of concern will be the tailings dam. To maintain stability and allow revegetation, slopes should be in the range of 2.5:1 horizontal to vertical to 3:1 H:V (CC. Scott, pers. comm. and CDA, 1999) or slope angles of 21.8° to 18.4°, respectively. The assumptions used in the tailings facility design (Section 2.5.2) take into account this stability requirement. Long term water quality is maintained typically by diverting clean surface runoff around the tailings facility and controlling the seepage from the facility itself. The degree of seepage control required would depend on the predicted quality of the seepage. The poorer the seepage quality, the higher degree of seepage control. As discussed in Section 2.5.1, tailings disposed in a conventional manner would be expected to oxidize and produce an acidic drainage containing high metal concentrations. Therefore, the closure plan in the model includes seepage control by Page 41 minimizing infiltration into the tailings facility. In some situations, long term water collection and treatment is required to lower receiving environment impacts (for example, Equity Silver Mine in British Columbia). The model includes a provision for long term water treatment. The closure plan in the model for the conventional tailings disposal case consists of a spillway in the dam, covering the surface of the impoundment with a low permeability cover, and revegetating the impoundment surface and the dam slope. The cover is comprised of a compacted foundation layer, a 2 mm HDPE liner and topped with growth medium. Post-closure monitoring includes an annual dam inspection and water sample collection. Some minor earthworks will likely be required in the first ten years after closure to repair the diversion ditch channels, spillway and cover until these structures have settled and stabilized. The model includes these provisions. An additional activity that needs to be considered for the post-closure period is water treatment of tailings impoundment seepages. Most mines are designed and permitted with the expectation that the waste management and the closure measures will be effective in preventing the release of deleterious water to the environment. However, the necessity of long term water treatment often becomes apparent during operations. The cost of water treatment during the post-closure period will depend on the acidity of the drainage(s) to be treated. Predictions of acidity levels (and costs) into the future are usually uncertain. This uncertainty will be the centre of discussions with regulators when negotiating the post-closure bond amount. The approach adopted for the Equity Page 42 Silver Mine was to regularly review the monitoring data and adjust the predicted water treatment costs accordingly (Equity Silver, 1996). Given the possibility of water treatment requirements for tailings facilities containing net acid producing tailings, the model includes a provision for post-closure water treatment in the conventional tailings disposal case. The closure plan in the model for the autoclaved tailings alternative reflects the benign geochemical nature expected of the tailings produced in this alternative. Seepage control and treatment are expected to be unnecessary due to the low metal concentrations predicted for tailings seepage. The closure plan consists of a spillway in the dam, covering the surface of the impoundment with a course granular material and revegetating the impoundment surface and the dam slope. Post-closure monitoring will include an annual dam inspection and water sample collection. Some minor earthworks will likely be required in the first ten years after closure to repair the diversion ditch channels, spillway and cover until these structures have settled and stabilized. These activities are included in the model. 2.7 Security and Bonding The model addresses the issue of reclamation bonding for the tailings facility only, as it was assumed that all other areas of the mine project would be the same. The model also addresses reclamation bonding in two parts - security and bonding. Security is the amount placed in trust at the beginning of the mine project to fund the necessary physical works at closure. The bond is the amount placed in trust prior to closure to generate sufficient interest income to fund the annual post closure costs. In the model it Page 43 is assumed that security is posted in the first year and that a bond is posted two years prior to closure. It is also assumed that the bond remains in perpetuity. The model calculates the amount of security from the estimated closure cost of the tailings management facility. Closure consists of constructing a spillway, cover (impermeable for the conventional case, coarse cover for the processed case) and revegetating the cover and the downstream dam face. The security amount was calculated to be one third of the closure cost estimate to be consistent with the 1997 level of funded liability in British Columbia (Errington, 1997). The bond cost is estimated from the expected post-closure activities. These activities were assumed to be an annual dam inspection, quarterly water quality monitoring, minor earthworks repairs every three years and maintenance seeding and fertilizing on 20% of the surface area every two years. In the conventional tailings disposal case it is also assumed that a water treatment facility will be operated. Due to the uncertainties associated with predicting post-closure water treatment costs, as discussed in Section 2.6, the water treatment costs estimated for the Equity Silver Mine was assumed for the model. The amount of bond is back-calculated from the estimated annual operating cost assuming the bond earns 3% interest. An example of the calculations for determining the security and bond amounts for conventional and autoclaved tailings is given in Figures 2.10 and 2.11, respectively. Figure 2.10 Security and Bond Calculation Sheet - Conventional Tailings Input Parameters Surface area - top Surface area - dam Bond interest rate Cover thickness 3,832,210 n r 5,701 m 2 3% 1.25 m Unit Costs Impermeable barrier Haul & place soil 10 $/m 2 5 $/m 3 Seeding - ground cover 0.075 $/m Seeding - erosion control 0.12 $/m 2 Closure Cost Spillway Cover Seeding Total closure cost $50,000 $62,273,408 $288,100 $62,611,508 Post Closure Cost Dam inspection Water quality monitoring Earthworks repair Maintenance seeding Water treatment plant Total post closure cost Cost per Yr Frequency $15,000 annual $25,000 annual $50,000 every 3 yrs $57,620 every 2 yrs Avg Cost per Yr $15,000 $25,000 $16,667 $28,810 $1,200,000 $1,285,477 Security & Bond Requirements Security amount Bond amount $18,783,452 $42,849,222 Figure 2.11 Security and Bond Calculation Sheet — Autoclaved Tailings Input Parameters Surface area - top 3,987,326 m 2 Surface area - dam 6,579 m 2 Bond interest rate 3% Unit Costs Simple cover Seeding - ground cover Seeding - erosion control 5 $/m 2 0.075 $/m 2 0.12 $/m 2 Closure Cost Spillway Cover Seeding Total closure cost $50,000 $19,936,632 $299,839 $20,286,471 Post Closure Cost Dam inspection Water quality monitoring Earthworks repair Maintenance seeding Total post closure cost Cost per Yr Frequency $15,000 annual $7,500 annual $50,000 every 3 yrs $59,968 every 2 yrs Avg Cost per Yr $15,000 $7,500 $16,667 $29,984 $69,151 Security & Bond Requirements Security amount Bond amount $6,762,157 $2,305,019 Page 46 2.8 Project Cost Summary The project cost summary compiles the capital and operating costs for each operating unit and assigns the costs to the year in which the expenditure is made. In the model it is assumed that all capital costs and security payments are incurred in the first year. Operating costs begin in year two and continue to the end of mine life. Bond payments occur two years prior to mine closure. The net present value (NPV) of the project is calculated to enable reasonable comparison between different scenarios. A discount rate of 10% was assumed to reflect the internal rate of return a mine company may use to make a financial decision. Cost figures for the various capital and operating components were obtained from existing mine operations or from recent feasibility studies. Cost estimates were scaled according to process rate where necessary. The unit cost data and sources for those data are listed in Table 2.9. The project cost summary for the conventional tailings disposal case (using the base parameters) is given in Table 2.10. The cost summary for the autoclaved tailings alternative is provided in Table 2.11. Page 47 Table 2.9 Sources for Cost Data Item Unit Rate Source Security (closure cost) calculation Errington, 1997 spillway $50,000 lump SRK Consulting, 2002 impermeable barrier $10/m2 SRK Consulting, 2002 place soil cover $5/m3 SRK Consulting, 2002 seeding $0.195/m2 Confidential Mine (1997) Bonding (post closure) dam inspection $15,000/yr P. Healey, pers. comm., 2002 water quality monitoring $25,000/yr Confidential Mine (1997) earthworks repair $50,000/3 yrs SRK Consulting, 2002 maintenance seeding $0.195/m2/2yrs Confidential Mine (1997) water treatment up to $1,200,000/yr Humber, unpublished data, 1996; Permit M- l 14 Flotation (capital) $1,152,000/2300 tonnes Benzaazoua et al, 2000 Flotation (operating) $0.40/tonne Mining Sourcebook, 1998 Autoclave (capital) $23,000,000/120 tonnes SRK Consulting, 2002 Autoclave (operating) $4/tonne thru mill Mining Sourcebook, 1998 Autoclave 0 2 plant* $100/tonneO2 SRK Consulting, 2002 Water treatment HDS (capital) US$1M-US$9.4M Humber, unpublished data, 1996 Direct lime (capital) US$3M - US$5.3M Humber, unpublished data, 1996 HDS (operating) US$158 -$235/L/yr Humber, unpublished data, 1996 Direct lime (operating) US$40.4 l /L /gH 2 S0 4 Humber, unpublished data, 1996 Tailings Facility grubbing $2/m2 SRK Consulting, 2002 foundation $l/m 2 SRK Consulting, 2002 liner $21/m2 SRK Consulting, 2002 liner fill $7/m3 SRK Consulting, 2002 dam $ll /m 3 SRK Consulting, 2002 * Oxygen is purchased from a third party rather than generated on-site. Table 2.10 Project Cost Summary Sheet - Conventional Case Item Total Cost Year 1 Year 2-17 Year 18 Year 19 Year 20 Capital Expenditures Bond & security Tailings facility Water treatment Piping & ancillary $61,632,674 $138,119,031 $6,913,806 $20,717,855 $18,783,452 $138,119,031 $6,913,806 $20,717,855 $0 $42,849,222 $0 $0 Subtotal Capital Discount Rate NPV Capital $227,383,366 10% $175,465,122 $184,534,144 $0 $42,849,222 $0 $0 Operating Expenditu res Tailings disposal Water treatment $653,600 $5,960,111 $34,400 $313,690 $34,400 $313,690 $34,400 $313,690 $34,400 $313,690 Subtotal Operating Discount Rate NPV Operating $6,613,711 10% $2,647,041 $0 $348,090 $348,090 $348,090 $348,090 NPV Project Cost $178,112,163 Table 2.11 Project Cost Summary Sheet - Autoclaved Tailings Alternative Item Total Cost Year 1 Year 2-17 Year 18 Year 19 Year 20 Capital Expenditures Bond & security Flotation cells Autoclave & ancillary Water treatment Tailings facility Piping & ancillary $9,067,175 $1,211,099 $14,484,150 $2,226,919 $9,428,270 $2,354,287 $6,762,157 $1,211,099 $14,484,150 $2,226,919 $9,428,270 $2,354,287 $0 $2,305,019 $0 $0 Subtotal Capital Discount Rate NPV Capital $38,771,901 10% $33,566,289 $36,466,882 $0 $2,305,019 $0 $0 Operating Expenditure s Flotation Autoclave 0 2 Plant Water treatment Inert tailings disposal $6,536,000 $65,360,000 $179,385,683 $321,770,984 $589,955 $344,000 $3,440,000 $9,441,352 $16,935,315 $31,050 $344,000 $3,440,000 $9,441,352 $16,935,315 $31,050 $344,000 $3,440,000 $9,441,352 $16,935,315 $31,050 $344,000 $3,440,000 $9,441,352 $16,935,315 $31,050 Subtotal Operating Discount Rate NPV Operating $573,642,622 10% $229,592,091 $0 $30,191,717 $30,191,717 $30,191,717 $30,191,717 NPV Project Cost $263,158,380 Page 50 3. S E N S I T I V I T Y A N A L Y S E S A number of assumptions were made to derive the costs presented in Section 2.8, specifically mine tonnage, mine life and sulphide content of the ore. The effect of these parameters on project cost was determined by conducting Monte Carlo simulations, varying each parameter simultaneously according to a triangular distribution. The convergence of the simulations was tested by computing the running average of the outputs as the number of trials increased. In each case, the running average exhibited considerable variation which decreased to insignificant levels as the number of trials increased. The simulations are discussed in the sections below. 3.1 Generic Simulations Mil l tonnage and mine life interact to impact several aspects of mine costs. The amount of material processed will have a direct impact on the size of equipment, the size of the tailings facility and the amount of reagents consumed. Mine life influences the size of the tailings facility without influencing the annual operating costs of the mill. The sulphur content of the tailings, in conjunction with mill tonnage, will influence the size of the autoclave in the autoclaved tailings alternative. The sulphur content will also determine the operating costs of the autoclave. These three parameters, namely mill tonnage, mine life and sulphur content, were varied simultaneously to examine the influence of these parameters on cost. Page 51 Mil l tonnage was varied from 500 tonnes per day to 20,000 tonnes per day with the most likely value, or peak point of the distribution, at 2,500 tonnes per day. The lower and upper bounds of the distribution were selected based on a review of mineral processing plant data in the Mining Sourcebook (1998), which indicated tonnages in the hundreds of tonnes to tens of thousands of tonnes. Mine life was varied between 10 years and 40 years with the most likely being 20 years. The parameters of the mine life distribution were arbitrarily selected but were reasonable for the industry. Flotation studies cited in Section 2.2 examined tailings samples containing a wide range of sulphide content. The same range (2% to 25%) was selected for the lower and upper bound for the triangular distribution. A most likely value of 5% was selected for the distribution. Sulphur content of tailings was varied independently from mill tonnage and mine life. Two hundred simulations were run. The capital costs of the autoclaved and conventional alternatives of the simulations are graphically presented in Figure 3.1. Figures 3.2 to 3.4 present the total costs against each varied parameter resulting from the simulations. Individual simulation results and the distribution of results are presented in Appendix D. Figure 3.1 shows that the capital cost for the autoclaved tailings alternative is consistently less than the capital cost for the conventional tailings alternative, irrespective of mine life or sulphur content. The main differences in cost are the tailings Page 52 facility and the security and bond for the conventional alternative. Costs associated with these two elements are significantly higher in the conventional alternative compared to the autoclave alternative. A liner is constructed in the conventional tailings facility, but not in the autoclaved tailings facility. The cost of an autoclave, while expensive, is less than the cost for a liner. Figure 3.1 also shows that the capital cost difference between the two alternatives increases as daily tonnage increases. However, Figures 3.2 to 3.4 demonstrate the significant cost associated with operating costs, which makes the total cost for the autoclaved tailings alternative higher than the total cost for the conventional tailings alternatives in most simulations, particularly at larger mill tonnages and longer mine lives. It can also be seen that autoclaved tailings total cost increases at a higher rate compared to the conventional tailings total cost. Figure 3.1 Capital Cost (Smillions) vs. Mill Tonnage Al l data from first set of simulations are shown B Autoclaved O Conventional 0<S> o o 4>v 4?m fj 7E3 kJ'"ij' L*-J T 3 D B a m o • • 2000 4000 6000 8000 10000 12000 Mill Tonnage (tonnes/day) 14000 16000 18000 20000 Figure 3.2 Total Cost (Smillions) vs. Mill Tonnage All data from first set of simulations are shown H Autoclaved O Conventional b • "fays 4D m mm 2000 4000 6000 8000 10000 12000 14000 16000 18000 20000 Mill T o n n a g e ( tonnes/day) Figure 3.3 Total Cost (Smillions) vs. Mine Life A l l data from first set of simulations are shown $6,000 $5,000 f l Autoclaved O Conventional $4,000 E r $3,000 in o O $2,000 • • $1,000 $0 • B i 6 g g i o o S o o o o o o o o o o o I m B o D a B • 10 15 20 Mine Life (years) 25 30 35 40 Figure 3.4 Total Cost (Smillions) vs. Sulphur Content All data from first set of simulations are shown Page 57 The data presented in Figures 3.2 to 3.4 shows that the total cost for the autoclaved tailings alternative is less than for the conventional alternative in some simulations. The conditions under which this occurred can be broadly described as mill tonnage less than 5,000 tonnes per day, mine life less than 12 years and sulphur content less than 15% S. An additional 200 simulations were run where mill tonnage, mine life and sulphur content were simultaneously varied within this narrower distribution range, namely 500 to 5,000 tonnes per day for mill tonnage, 5 to 12 years for mine life, and 2% to 15% for sulphur content. The results for this second set of simulations are graphically presented in Figures 3.5 to 3.7. Individual simulation results are tabulated in Appendix D. This second set of data shows trends similar to the first set of data, namely the total cost for the autoclaved tailings alternative increases with increasing mill tonnage and mine life and at a rate greater than for the conventional alternative. However, this second set of simulations clearly shows that there are conditions where the autoclaved tailings alternative costs less than the conventional alternative. The simulation results were filtered to identify all simulations where the total cost for the autoclaved tailings alternative was less than the total cost for the conventional alternative. This subset of data is shown in Figure 3.8. The data indicate that total costs for the autoclaved alternative are less than the conventional alternative as sulphur content decreases in conjunction with mill tonnage and mine life increases. Figure 3.5 Total Cost (Smillions) vs. Mi l l Tonnage All data from second set of simulations are shown a Autoclaved O Conventional a B -o °f a Bn • — • — B -P B I 500 1000 1500 2000 2500 3000 Mill Tonnage (tonnes/day) 3500 4000 4500 5000 Figure 3.6 Total Cost (Smillions) vs. Mine Life Al l data from second set of simulations are shown a Autoclaved O Conventional 8 9 Mine Life (years) 10 11 12 13 Figure 3.7 Total Cost (Smillions) vs. Sulphur Content All data from second set of simulations are shown Figure 3.8 Sulphur Content vs. Mill Tonnage for Simulations where the Total Cost for the Autoclaved Alternative is Less Than the Conventional Alternative O 5 yr mine life • 6 yr mine life A 7 yr mine life X 8 yr mine life O 9 yr mine life X ^ X X X 500 1000 1500 Mi l l Tonnage (tonnes/day) 2000 2500 3000 Page 62 3.2 Case Studies Model runs were completed on three scenarios representing typical mining operations. One scenario was a small tonnage operation mining a deposit containing a relatively high percentage of sulphur. The second scenario was an underground operation mining a massive sulphide deposit. The third scenario was an open pit operation mining a relatively low sulphide deposit. The scenario definitions are given in Table 3.1. Scenario results are summarized in Table 3.2. Table 3.1 Scenario Definitions Parameter Scenario 1 Scenario 2 Scenario 3 Mill tonnage Sulphur content Mine life 1,000 tonnes/day 12% S 5, 10 and 20 years 4,000 tonnes/day 20% S 5, 10 and 20 years 20,000 tonnes/day 2.5% S 5,10 and 20 years Table 3.2 Total Cost of the Autoclaved and Conventional Alternatives in Defined Scenarios Costs in Smillions Scenario 1 Scenario 2 Scenario 3 Mi l l tonnage 1,000 t/day 4,000 t/day 20,000 t/day %S 12% 20% 2.5% Mine Life Autoclaved Conventional Autoclaved Conventional Autoclaved Conventional 5yr $107 $148 $596 $158 $432 $237 10 yr $174 $139 $1,027 $173 $742 $283 20 yr $239 $133 $1,463 $193 $1,069 $396 The scenario results suggest that autoclave alternative costs for a high tonnage, low sulphur operation (Scenario 3) can be comparable with an intermediate tonnage, high sulphur operation (Scenario 2) for a given mine life. This is largely a consequence of the Page 63 different mass pull rates at the sulphide separation step resulting in similar amounts of concentrate flowing to the autoclave. Following equation (3) in Section 2.2, the mass pull rates for Scenarios 2 and 3 are 40.6% and 4.5%, respectively. This translates to 1624 tonnes per day and 900 tonnes per day of material to the autoclave, respectively. However, the autoclaved alternative total cost remained higher than the total cost for the conventional alternative, which is consistent with the simulation results presented in Section 3.1. Page 64 4. D I S C U S S I O N 4.1 Model Results The model results indicate that for smaller operations under certain conditions it may be less costly to autoclave sulphidic tailings rather than dispose a reactive tailings material. This result was observed in the simulations where the mill tonnage was less than 3,000 tonnes per day, the mine life was less than 10 years and the sulphur content of the tailings was less than 12%. The autoclaving alternative is cost effective as sulphur content decreases in conjunction with mill tonnage and mine life increases. For all simulations, the capital cost associated with autoclaving tailings was less than the capital cost associated with conventional tailings disposal. This in itself may be very appealing to mine developers for the lower interest payments on the funds borrowed prior to production. However, autoclaving does have significantly higher operating costs. For all simulations, autoclaving operating costs were one to two orders of magnitude higher than the operating costs of conventional tailings disposal. The autoclaved tailings operating cost is dominated by the oxygen requirements of the autoclave and by the lime requirements to neutralize the autoclave discharge. Lime consumption was expected to be high due to the high acid concentration in the autoclave discharge. However, if a very concentrated acid solution was to be produced, such as is anticipated in the simulations containing very high (-20%) sulphide content, it may be more practical to separate the sulphuric acid from the autoclave discharge and sell it. This could potentially reduce the overall operating cost of the tailings processing Page 65 circuit by significantly reducing the lime consumption rate in the water treatment plant and by generating revenues to offset the cost of oxygen for the autoclave. In addition, the autoclave discharge would also likely contain a significant amount of dissolved metals that could possibly be recovered. Operating costs for a processed tailings alternative may also be reduced by using an alternate oxidation method. Oxidation methods that could be considered include bio-oxidation and atmospheric pressure leaching. Reagents associated with these processes are likely less expensive than oxygen. On the other hand, the oxidation rates achievable with these alternate methods are likely to be significantly less than the rate achieved in an autoclave. Although this could require a processing time that extends beyond the other mining activities, the operating costs may be covered by a bond - similar to the bonding of long term water treatment operations of today but with more predictable annual costs. The potential benefits of using an alternate oxidation method include reduced financial burden and good containment of problematic material prior to treatment while still achieving an inert tailings material for disposal. The main technical uncertainties that will need to be addressed are the efficiency of the oxidation process and the geochemical stability of the oxidized solids. 4.2 Model Construction Many of the design calculations in the model are extrapolations from existing designs. The limitation of this approach is that at the very low and very high ends of the Page 66 distribution range, the accuracy of the design calculations decrease. The areas of the model where this is an issue are discussed below. Autoclave Design The sulphide content of the material directly affects the size of the autoclave, where the amount of water required to maintain the heat balance is dictated by the sulphide content. The autoclave sizing in the model is based on the Con Mine autoclave (see Section 2.3) regardless of sulphide content in the autoclave feed. However, the Con autoclave was designed for a sulphide concentration of 12% in the autoclave feed, equivalent to about 6.5% sulphide in tailings in the model. The autoclave size (and cost) estimated for the simulations where pyrite content is greater than about 10% is likely an optimistic estimate. In other words, the autoclave size would likely be larger to accommodate the increased amount of water required to control the heat generated from the higher sulphide content. As a result the cost would be higher; however, this does not alter the comments in Section 4.1. Water Treatment Plant Design There are two limitations to the water treatment cost estimates in the model. One, the range of flows in the model exceed the range of data available. Second, there are no operational data on water treatment plant efficiency in neutralizing autoclave discharges in the manner proposed in the model. Page 67 As discussed in Section 2.4, cost estimates for water treatment were calculated by interpolating from a look-up table generated from Humber (1996). The lower and upper bounds of mill tonnage produce autoclave discharge flows that are outside the range of flows where operational data are available. At the low end, water treatment costs may be underestimated, due to a minimum cost that is required to build a plant regardless of treatment volume. Costs may be overestimated at the upper end because the proportion of incremental cost against treatment volume may change at the higher treatment volumes. Operational data for dilute acidic solutions was used as a basis for estimating the cost of neutralizing autoclave residue. However, field or test data were not found to verify the approach used in the model. Uncertainty Associated with Tailings with Uncertain Acidic Drainage Potential The model is designed for situations where there is a high sulphide content in the tailings. The lower bound for sulphide content assumed for this thesis was 2%. Even with sufficient neutralization to prevent acidic drainage, the oxidation of this much sulphide can be sufficient to produce unacceptably high metal concentrations in neutral drainage. In addition, the heterogeneous nature of tailings facilities allows for the possibility of localized acidic drainage although the overall ratio of neutralization potential to acidity production potential indicates neutral drainage. To accommodate these potential scenarios, tailings facility design in the model is done to the same Page 68 standard as if the entire tailings inventory would produce acidic drainage. The model does not calculate the probability of a geochemical characterization being incorrect and the cost of the remedial measures required to compensate. Page 69 5 . CONCLUSIONS A spreadsheet model was developed to compare the whole project costs of an autoclaved tailings disposal alternative to a conventional tailings disposal alternative. The purpose of autoclaving (or processing) tailings is to reduce the long term environmental risk of tailings facilities. Autoclaving was arbitrarily selected as the tailings processing method for its demonstrated ability to efficiently oxidize sulphides. This thesis demonstrates that processing tailings may be an economically viable way of reducing the environmental liability of mine sites under certain conditions. The conditions under which this holds true are an interplay between mill tonnage, mine life and sulphur content in the conventional tailings. In general terms, the economics of autoclaving tailings become attractive at lower sulphur contents and smaller tonnages. The major findings of this thesis are summarized below. 1. Autoclaved tailings disposal should be considered as a disposal option when the sulphur content of mill tailings is 14% or less, the mill tonnage is less than 5,000 tonnes per day and the mine life is less than 12 years. Cost competitiveness of autoclaved tailings follows a curve as sulphur content decreases in conjunction with mill tonnage and mine life increases. 2. Autoclaving tailings has a lower capital cost requirement compared to conventional tailings disposal. 3. The cost of autoclaving tailings is most sensitive to flotation efficiency. 4. The model would be improved by the addition of sub-models for water treatment and autoclave design. Page 70 6. RECOMMENDATIONS FOR FURTHER STUDY The results of running the model under various simulation scenarios suggest that under certain situations processing tailings may be an economically viable option to reduce the long term environmental risk of tailings facilities. The following studies are recommended to improve the model and to verify the observations discussed herein. 1. Develop a simplified version of MetSim®1 The model uses a single autoclave design as the basis for estimating the autoclave size for a range of scenarios. However, there are limits to the applicability of a single design. An autoclave design sub-model that utilized a more detailed mass balance and a heat balance would improve the accuracy of the model at higher sulphur contents. 2. Include a water treatment model Water treatment design and cost depends on the anticipated flowrates and the estimated chemistry of the flows. The model currently uses an assumed water quality for all scenarios and interpolates costs from a limited amount of operational data. The model would have wider applicability and be more accurate if a water treatment sub-model was included. The water treatment model should allow the user to input the predicted water quality of the flows to 1 MetSim® is a computer program that calculates the mass and energy balance of metallurgical processes. Recent versions of the program also include process control, equipment sizing, cost estimating and process analysis. MetSim® is sold by Proware. Additional information can be obtained from httj5://members.ozemail.corriyau/~<)zmetsim/metl/index.html. Page 71 be treated and contain a wider range of operational data from which to interpolate. 3. Verify lime neutralization of autoclave residue Autoclave residue neutralization in the model assumes direct neutralization of the residue, as opposed to the more typical approach in a processing plant where the solids and liquor of the residue are separated in a counter current decant circuit and neutralized separately. A bench scale test is recommended to evaluate the practicality of direct neutralization of autoclave residue that has not been separated. Parameters to study should include lime demand and sludge characteristics. 4. Add additional processing alternatives The model currently contains one alternative to conventional tailings disposal. Other alternatives could be equally effective in achieving a geochemically stable tailings material. Bio-oxidation, atmospheric pressure leaching, and autoclaving in multiple smaller vessels are some alternatives that could be included. 5. Expand the disposal and closure alternatives The model assumes a fixed tailings disposal and closure method for the processed tailings alternative. Other methods may also be suitable. For example, dry stacked tailings and direct vegetation may be a viable disposal and closure method when the run-off quality from tailings material is not a concern. Page 72 6. Cost benefit analysis of lower capital costs The capital cost of a project, and ways of reducing the amount of capital required, is an important issue due to the fact that capital is almost always borrowed money. In some cases, the alternative with a higher overall project cost will be selected for the lower capital cost. The model currently does not include the cost of capital and other financial considerations, such as tax issues. Inclusion of these items into the overall project cost summary may redefine the conditions under which processing tailings would be financially attractive. Page 73 7. REFERENCES Arbiter, N (ed), 1985. Flotation. In: SME Mineral Processing Handbook, N.L. Weiss (ed.). Society of Mining Engineers, New York, NY, 1985. Benzaazoua M . , Bussiere B., Kongolo M . , McLaughlin J., Marion P., 2000. Environmental Desulphurization of Four Canadian Mine Tailings Using Froth Flotation. Int. J. Miner. Process. 60 (2000) 57-74. British Columbia. Ministry of Energy, Mines and Petroleum Resources (MEMPR). Health, Safety and Reclamation Code for Mines in British Columbia. 1992. British Columbia. Ministry of Energy, Mines and Petroleum Resources (MEMPR). "Mine Reclamation Security Policy in British Columbia. A Paper for Discussion." February, 1995. Canadian Dam Association (CDA), 1999. Dam Safety Guidelines. Catalan, L.J.J., L i , M.G., McLaughlin, J., Nesset, J., St-Arnaud, L., 1999. 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Kuyucak, N . , Sheremata, T. and Wheeland, K., 1991. Evaluation of Improved Lime Neutralization Processes. Part I Lime Sludge Generation and Stability. In: Proceedings of the Second International Conference on the Abatement of Acidic Drainage, Montreal, PQ, September 16-18, 1991. Leppinen J.O., Salonsaari P. and Palosaari V., 1997. Flotation in Acid Mine Drainage Control: Beneficiation of Concentrate. Canadian Metallurgical Quarterly, vol 36, no. 4, pp. 225-230. Page 76 Lewis B.A., Gallinger R.D., and Wiber M . , 2000. Poirier Site Reclamation Program. In: Proceedings from the Fifth International Conference on Acid Rock Drainage, Denver, CO, 2000. Mining Sourcebook, 1998. Mineral Processing Plants - Costs. Can. Min. J., pp. 65-83. Ministry of Energy, Mines and Petroleum Resources (MEMPR), 1995. Mine Reclamation Security Policy in British Columbia. A Paper for Discussion prepared for the Province of British Columbia. 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Acid Mine Drainage - Status of Chemical Treatment and Sludge Management Practices. Prepared for The Mine Environment Neutral Drainage (MEND) Program, June 1994. Steffen Robertson and Kirsten (SRK), 1989. Draft Acid Rock Drainage Technical Guide. Prepared for the BC Acid Mine Drainage Task Force, Volume I. SRK Consulting, 2002. Study of Management Alternatives - Giant Mine Arsenic Trioxide Dust. Report 1CI001.10 prepared for Department of Indian Affairs and Northern Development, December 2002. Page 80 Tremblay, G.A., 2000. The Canadian Mine Environment Neutral Drainage 2000 (MEND 2000) Program. In: Proceedings from the Fifth International Conference on Acid Rock Drainage. Denver, CO. Vol. 1, pp. 33-40. Zinck, J.M. and Griffith, W.F., 2000. An Assessment of HDS-Type Lime Treatment Processes - Efficiency and Environmental Impact. In: Proceedings from the Fifth International Conference on Acid Rock Drainage. Denver, CO. Vol. II, pp. 1027-1034. Page 81 APPENDIX A E X A M P L E C A L C U L A T I O N OF AN HDS W A T E R T R E A T M E N T PLANT DESIGN USING UNPUBLISHED DATA BY HUMBER (1996) Use; 1. Complete solids generation w/ recycle estimate 2. Complete operating parameters 3. Complete process design 4. Set pond requirements 5. Complete Process Flowsheet and Equipment List Enter the Process Data on this sheet: HDS Process Design Base Conventional 4000 Umin 11 August, 2002 Pre-Feasibility Rev. #1 General Design Information Flocculant Dosing System Design Flowrate: Solids Generation Recycle Ratio Solids SG Feed pH Rapid Mix Tank pH Reactor pH Clarifier U/F Density Clarifier Overflow Solids 4,000 Umin 5.9 g/L plant feed 11 (?:1) 2.8 3.5 13.5 pH Units 9.3 pH Units 25 % 0 mg/L Flocculant Dose Rate Flocculant Addition Rate Undiluted Floe Concentration Lime Dosing System Lime Addition Rate (as Ca(OH)2) Lime Slurry Concentration Slurry pH Solids SG Storage Requirements 100 mg floc/kg solids (range from 50 to 200) 7.0 mg floc/L plant feed (range from 1 to 10) 0.5 % 3.4 g lime/L plant feed 15 % 14 pH Units 2.4 6 hours Aeration Requirements Feed Iron Content 152 mg/L Percentage Ferrous Iron 100 % Average Density of Air 1.201 kg/m3 Oxygen Transfer Efficiency 20 % Available CaO Lime use Ferrous/Ferric Iron Dosing System Fe27Fe3+Dosage (15%) Available Iron 90.0 % 2.9 g lime (CaO)/L plant feed 0 mg Fe2+/Fe3+/L Feed 23 % Vessel Residence Times: Operating Costs Reactor Residence Time Lime Sludge Mix Tank Rapid Mix Tank Flocculant Tank Clarifier Upflow Ratio Recycle Water Tank 40 minutes 4 minutes 3 minutes 2 minutes 1.200 (m3/hr)/m2 0.5 minutes Lime Cost Flocculant Cost Ferric/Ferrous Sulphate Cost Power Cost Manpower Cost O&M Capital 100 US$/tonne 3600 US$/tonne 140 US$/tonne 200 hp 8 man-hours/day 3 % of capital cost 0.05 US$/kw-hour 24 US$/man-hour 2,300,000 US$ total capital Page 83 Water Quality and Sludge Generation Prediction HDS Process Design Base Conventional 4000 L/min 11 August, 2002 Hydroxide Mass of Mass of Mass of Ion lonWt Hydroxide Weight Ion Present OH- Precip. (g/mol) Formula (g/mol) (mg/L) (mg/L) (mg/L) Al 26.98 AI(OH)3 78.01 10.00 18.91 28.91 Ag 107.87 AgOH 124.88 0.10 0.02 0.12 As 74.92 As(OH)3 125.95 0.57 0.39 0.96 Bi 208.98 Bi(OH)3 260.01 0.00 0.00 0.00 Ca 40.08 Ca(OH)2 74.1 55.00 0.00 0.00 Cd 112.41 Cd(OH)2 146.43 0.15 0.05 0.20 Cu 63.55 Cu(OH)2 97.57 35.00 18.74 53.74 Fe 55.85 Fe(OH)3 106.88 152.00 138.88 290.88 Pb 207.2 Pb(OH)2 241.22 26.00 4.27 30.27 Mg 24.31 Mg(OH)2 58.33 14.00 19.59 33.59 Mn 54.94 Mn02 86.94 25.00 0.00 39.56 Ni 58.71 Ni(OH)2 92.73 4.20 2.43 6.63 S* 32.06 CaS0„.2H20 172.18 0.00 0.00 0.00 Sb 121.75 Sb(OH)3 172.78 0.00 0.00 0.00 Se 78.96 Se(OH)4 147 0.00 0.00 0.00 Si 28.09 Si(OH)2 62.11 0.00 0.00 0.00 Zn 65.38 Zn(OH)2 99.4 54.00 28.10 82.10 SO/'* 96.06 CaS04.2H20 172.18 3500.00 0.00 3853.71 C032" 59.98 CaC03 100.06 123.00 0.00 205.19 TSS n/a n/a n/a n/a n/a 263.00 Total 231.37 488S.85 Residual S042" concentration 1350 mg/L (pure solubility range from 1240 - 1435 mg/L) * Use either (S) or (S04). Lime Requirements Based on calcium requirements Based on hydroxide requirements Solids Generation = 5.9 g/L (includes 10.0 % lime enerts) (includes 20.0 % unreacted lime solids) 2.75 0.00 0.50 g Ca(OH)2/L effluent OR g Ca(OH)2/L effluent g Ca(OH)2/L effluent (S042" based) (S based) Lime Utilization = Available CaO = Lime use = Lime use = 80.0 % 90.0 % 3.4 g Ca(OH)2/L effluent 2.9 g lime (CaO)/L effluent Page 84 Sludge Quality Prediction HDS Process Design Base Conventional 4000 L/min 11 August, 2002 Mass of Mass of Mass of Mass of Sludge Ion Ion Present OHT Precip. Metal Composition (mg/L) (mg/L) (mg/L) (mg/L) (%) Al 10.00 18.91 28.91 10.00 0.17 Ag 0.10 0.02 0.12 0.10 0.00 As 0.57 0.39 0.96 0.57 0.01 Bi 0.00 0.00 0.00 0.00 0.00 Ca 55.00 0.00 0.00 0.00 0.00 Cd 0.15 0.05 0.20 0.15 0.00 Cu 35.00 18.74 53.74 35.00 0.60 Fe 152.00 138.88 290.88 152.00 2.59 Pb 26.00 4.27 30.27 26.00 0.44 Mg 14.00 19.59 33.59 14.00 0.24 Mn 25.00 0.00 39.56 25.00 0.43 Ni 4.20 2.43 6.63 4.20 0.07 CaS0 4 . 2H 2 0 0.00 0.00 0.00 n/a 0.00 Sb 0.00 0.00 0.00 0.00 0.00 Se 0.00 0.00 0.00 0.00 0.00 Si 0.00 0.00 0.00 0.00 0.00 Zn 54.00 28.10 82.10 54.00 0.92 CaSO„.2H20 3500.00 0.00 3853.71 n/a 65.70 C a C 0 3 123.00 0.00 205.19 n/a 3.50 TSS n/a n/a 263.00 n/a 4.48 Lime Inerts n/a n/a 976.61 n/a 16.65 Total 231.37 5865.47 321.02 95.81 Balance Check: 100.00 % Solids generation • Ultimate drained percent solids • Sludge pond lifetime : 5.9 g/L 55 % 20 years Annual Average Data: Plant feed rate : Total dry solids production : Sludge volume purged : Volume at ultimate density : Pond volume required : 3,750 L/minute 31.7 tonnes/day 106.3 m3/day 37.2 m3/day 272,000 m 3 11560.8 tonnes/year 38811.4 m3/year 13587.7 m3/year Vessel Sizes Lime Sludge Mix Tank: 4 m 3 = 1014 USgal Rapid Mix Tank: 15 m 3 = 3932 USgal Reactor Vessel: 198 m 3 = 52427 USgal Flocculation Tank: 10 m 3 = 2658 USgal Clarifier Diameter: 16 m = 53 ft Lime Storage Tank: 33 m 3 = 8839 USgal Recycled Water Tank: 2 m 3 = 540 USgal Aeration Requirements Total Iron Content = 152 mg/L Percent Ferrous Iron = 100 % Oxygen Transfer Efficiency = 20 % Total Flow In = 4000 L/min Total Ferrous Iron = 0.6 kg/min = 36.5 kg/hr Aeration required = 104 m3/hour 61 S C F M Tank Dimensions (no freeboard included) aspect ratio D= 2.0 m or 6.6 ft H - 1.2 m or 4.0 ft 1.64 D= 3.0 m or 9.8 ft H - 2 .1m or 6.9 ft 1.42 D= 7.0 m or 23.0 ft H= 5.2 m or 16.9 ft 1.36 D= 2.5 m or 8.2 ft H= 2.0 m or 6.7 ft 1.22 H= 4 .0m or 13.1ft W=L= 2.9 m or 9.5ft 1.38 D= 1.5 m or 4.9ft H= 1.2 m or 3.8 ft 1.30 Sludge and Reagent Flowrates Sludge Purge and Recycle Sludge Purge Data Sludge Purge = Solids Generation = 23 kg/min = 52 Ibs/mln Solids Volume = 8 L/mln = 2 USgpm Water Flow = 70 L/min = 19 USgpm Total Flow = 79 L/min = 21 USgpm SG Slurry = 1.19 pH Slurry = 9.3 pH Units SG Solids = 2.8 Slurry % Solids = 25.00 % Lime Circuit Lime Dosina Solids Mass = 15 kg/min = 34 Ibs/min Solids Volume = 6 L/min = 2 USgpm Water Flow = 87 L/min = 23 USgpm Total Slurry Flow = 93 L/mln = 25 USgpm Slurry SG = 1.10 pH Slurry = 14 pH Units SG Solids = 2.4 Slurry % Solids = 15.00 % Flocculant Dosing Floo Dosing Rate = 7 mg/L effluent treated Flow Into Floo Tank = 4961 L/min = 1311 USgpm Undiluted Floo Flowrate = 7 L/min = 2 USgpm Diluted Floo Flowrate = 70 L/min = 18 USgpm Floo Consumption = 50 kg/day = 111 lbs/day Sludge Recycle Data Solids Recycled = 258 kg/min = 5S9 Ibs/min Solids Volume = 92 L/mln = 24 USgpm Water Flow = 774 L/mln = 205 USgpm Total Flow = 866 L/mln = 229 USgpm SG Slury = 1.19 pH Slurry = 9.3 pH Units SG Solids = 2.8 Slurry % Solids = 25.00 % Lime Loop Out Of Storage Tank Lime Loop Return To Storane Tank . Solids Mass = 61 kg/min = 135 lbs/mm Solids Mass = 46 kg/min = 101 Ibs/min Solids Volume = 25 L/min = 7 USgpm Solids Volume = 19 L/mln = 5 USgpm Water Flow = 346 L/mln = 91 USgpm Water How = 260 L/mln = 69 USgpm Total Slurry Flow = 372 L/mln = 98 USgpm Total Slurry Flow = 279 L/min = 74 USgpm Slurry SG = 1.10 Slurry SG = 1.10 pH Slurry = 14 pH Units pH Slurry = 14 pH Units SG Solids = 2.4 SG Solids = 2.4 Slurry % Solids = 15.00 % Slurry % Solids = 15.00 % Iron Sulphate Dosina Lime Dosing Fe 2+/3+ Dosing Rate = 0 mg Fe2*/Fe '*/L effluent treated Lime Dosing Rate = 3.4 g llme/L effluent treated Plant Feed Rate = 4000 L/min = 1057 USgpm Lime Dosing Rate = 2.9 g lime (CaO + inerts)/L \treated Solution Concentration = 15 % Average Plant Feed = 4000 L/minute Solution Flowrate = 0 L/min = 0 USgpm Daily Consumption = 16.7 tonnes/day Iron Sulphate Consumption = 0 kg/day = 0 lbs/day Annual Consumption^ 6078 tonnes/year quicklime Oo Tank Flows Out Of Lime/Sludae Mix Tank Solids Mass = 273 kg/min = Solids Volume = 99 L/min = Water Flow = 861 L/min = Total Slurry Flow = 959 L/min = Slurry S G = 1.18 pH Slurry = 13.5 pH Units SG Solids = 2.77 Slurry % Solids = 24.10 % Out Of Reactor Tank Solids Mass = 281.54 kg/min = Solids Volume = 101 L/min = Water Flow = 4861 L/min = Total Slurry Flow = 4961 L/min = Slurry SG = 1.04 pH Slurry = 9.3 pH Units SG Solids = 2.80 Slurry % Solids = 5.47 % Out Of Rapid Mix Tank 603 Ibs/min Solids Mass = 282 kg/min = 621 Ibs/min 26 USgpm Solids Volume = 101 L/min = 27 USgpm 227 USgpm Water Flow = 4861 L/min = 1284 USgpm 253 USgpm Total Slurry Flow = 4961 L/min = 1311 USgpm Slurry SG = 1.04 pH Slurry = 9.3 pH Units SG Solids = 2.80 Slurry % Solids = 5.47 % Out Of Flocculant Tank 620.80 Ibs/min Solids Mass = 282 kg/min = 621 Ibs/min 27 USgpm Solids Volume = 101 L/min = 27 USgpm 1284 USgpm Water Flow = 4931 L/min = 1303 USgpm 1311 USgpm Total Slurry Flow = 5031 L/min = 1329 USgpm Slurry SG = 1.04 pH Slurry = 9.3 pH Units SG Solids = 2.80 Slurry % Solids = 5.40 % Clarifier Flows Clarifier Overflow Solids Mass = 0 kg/min = 0 Ibs/min Solids Volume = 0 Umin = 0 USgpm Water Flow = 4086 L/min = 1079 USgpm Total Slurry Flow = 4086 L/min = 1079 USgpm Slurry SG = 1 pH Slurry = 9.3 pH Units SG Solids = n/a Slurry % Solids = 0 % Balance Check (Overall) Total Solids In = 23.46 kg/min Total Solids Out = 23.46 kg/min % Deviation = 0.00 % Balance Check (Clarifier) Total Solids In = 282 kg/min Total Solids Out = 282 kg/min % Deviation = 0.00 % Clarifier Underflow Solids Mass = 282 kg/min = 621 Ibs/min Solids Volume = 101 L/min = 27 USgpm Water Flow = 845 Umin = 223 USgpm Total Slurry Flow = 945 Umin = 250 USgpm Slurry SG = 1.19 pH Slurry = 9.3 pH Units SG Solids = 2.80 Slurry % Solids = 25.00 % Total Water In = 4156 Umin Total Water Out = 4156 Umin % Deviation = 0.00 % Total Water In = Total Water Out = % Deviation = 4931 Umin 4931 Umin 0.00 % Page 89 Operating Cost Estimate Base Conventional 4000 L/min 11 August, 2002 Reagent Dose Rate (mg/L plant feed) Annual Average Plant Flow Rate (L/min) Annual Reagent Consumption (tonnes/year) Reagent Unit Cost (US$/tonne) Annual Reagent Cost (US$/year) Quicklime 2891 3,750 5698 100 570,000 Flocculant 7 3,750 14 3600 50,000 Iron Sulphate 0 3,750 0 140 0 Sub-total: $620,000 Item Annual Consumption Unit Cost (US$) Annual Cost (US$/year) Electric Power 1.31 million kW-hours 0.05 65,000 O & M Capital 3 % of capital cost 2300000 69,000 O & M Manpower 8 man-hours per day 24 70,000 Sub-total: $204,000 Total Annual Operating Cost: $824,000 /year (US dollars) Normalized Annual Operating Cost: $0.42 /m 3 (US dollars) $1.58 /1000 USgal (US dollars) Discount Interest Rate: 10% Expected Plant Lifetime: 20 years Present Value of Plant Operating Costs: $7,015,000 US dollars Net Present Value of Plant: $9,315,000 US dollars Page 90 APPENDIX B E X A M P L E C A L C U L A T I O N OF A SIMPLE L I M E NEUTRALIZATION W A T E R TREATMENT PLANT DESIGN USING UNPUBLISHED DATA B Y H U M B E R (1996) Simple Lime Neutralization Process Design Use: 1. Complete solids generation 2. Complete operating parameters Base Process 3. Complete process design 3000 L/min 4. Set pond requirements 5. Complete Process Flowsheet and Equipment List 11 August, 2002 Pre-Feasibility Enter the Process Data on this sheet: Rev. #1 General Design Information Design Flowrate: Solids Generation Solids SG 3,000 L/min 487.8 g/L Effluent Treated 2.8 Flocculant Dosing System Flocculant Dose Rate Flocculant Addition Rate Undiluted Floe Concentration 100 mg floc/kg solids (range from 50 to 200) 48.8 mg floc/L effluent treated (range from 1 to 10) 0.5 % Feed pH Flash Mix Tank pH Reactor Vessel pH Settling Pond Solids Density Aeration Requirements -7.0 pH Units 7 pH Units 9.3 pH Units 50 % Lime Dosing System Lime Addition Rate (as Ca(OH)2) Lime Slurry Concentration Slurry pH Solids SG Storage Requirements 93.9 g Lime/L Effluent Treated 15 % solids 14 pH Units 2.4 6 hours Feed Iron Content Percentage Ferrous Iron Average Density of Air Oxygen Transfer Efficiency 6000 mg/L 0 % 1.201 kg/m3 20 % Available CaO Lime use Operating Costs 90.0 % 78.9 g lime (CaO)/L effluent Vessel Residence Times: Flash Mix Tank Res. Time Reactor Vessel Res. Time Floccuiator Tank 10 minutes 30 minutes 3 minutes Lime Cost Flocculant Cost Power Cost Manpower Cost O&M Capital 100 US$/tonne 3600 US$/tonne 100 hp 8 man-hours/day 3 % of capital cost 0.05 US$/kw-hour 24 US$/man-hour 4,000,000 US$ total capital Page 92 Water Quality and Sludge Generation Prediction Simple Lime Neutralization Process Design Base Process 3000 L/min 11 August, 2002 Hydroxide Mass of Mass of Mass of Ion lonWt Hydroxide Weight Ion Present OH" Preclp. (g/mol) Formula (g/mol) (mg/L) (mg/L) (mg/L) Al 26.98 AI(OH)3 78.01 0.00 0.00 0.00 Ag 107.87 AgOH 124.88 0.00 0.00 0.00 As 74.92 As(OH)3 125.95 0.00 0.00 0.00 Bi 208.98 Bi(OH)3 260.01 0.00 0.00 0.00 Ca 40.08 Ca(OH)z 74.1 0.00 0.00 0.00 Cd 112.41 Cd(OH)2 146.43 0.00 0.00 0.00 Cu 63.55 Cu(OH)2 97.57 0.00 0.00 0.00 Fe 55.85 Fe(OH)3 106.88 6000.00 5482.18 11482.18 Pb 207.2 Pb(OH)2 241.22 0.00 0.00 0.00 Mg 24.31 Mg(OH)2 58.33 0.00 0.00 0.00 Mn 54.94 MnOz 86.94 0.00 0.00 0.00 Ni 58.71 Ni(OH)2 92.73 0.00 0.00 0.00 S* 32.06 CaS04.2H20 172.18 0.00 0.00 0.00 Sb 121,75 Sb(OH)3 172.78 0.00 0.00 0.00 Se 78.96 Se(OH)4 147 0.00 0.00 0.00 Si 28.09 Si(OH)2 62.11 0.00 0.00 0.00 Zn 65.38 Zn(OH)2 99.4 0.00 0.00 0.00 s o 4 2 * 96.06 CaS04.2H20 172.18 96000.00 0.00 169670.61 C032" 59.98 CaC03 100.06 0.00 0.00 0.00 TSS n/a n/a n/a n/a n/a 280000.00 Total 5482.18 461152.79 Residual S042" concentration * Use either (S) or (S04). Lime Requirements Based on calcium requirements 1340 mg/L (pure solubility range from 1240 -1435 mg/L) Solids Generation = 487.8 g/L (includes 10.0 % lime enerts) (includes 20.0 % unreacted lime solids) 75.09 0.00 g Ca(OH)2/L effluent OR g Ca(OH)2/L effluent (S04 based) (S based) Based on hydroxide requirements 11.94 g Ca(OH)2/L effluent Lime Utilization • Available CaO = 80.0 % 90.0 % Lime use = Lime use = 93.9 g Ca(OHyL effluent 78.9 g lime (CaO)/L effluent Page 93 Sludge Quality Prediction Simple Lime Neutralization Process Design Base Process 3000 L/min 11 August, 2002 Mass of Mass of Mass of Mass of Sludge Ion Ion Present OH" Precip. Metal Composition (mg/L) (mg/L) (mg/L) (mg/L) (%) AJ 0.00 0.00 0.00 0.00 0.00 Ag 0.00 0.00 0.00 0.00 0.00 As 0.00 0.00 0.00 0.00 0.00 Bi 0.00 0.00 0.00 0.00 0.00 Ca 0.00 0.00 0.00 0.00 0.00 Cd 0.00 0.00 0.00 0.00 0.00 Cu 0.00 0.00 0.00 0.00 0.00 Fe 6000.00 5482.18 11482.18 6000.00 1.23 Pb 0.00 0.00 0.00 0.00 0.00 Mg 0.00 0.00 0.00 0.00 0.00 Mn 0.00 0.00 0.00 0.00 0.00 Ni 0.00 0.00 0.00 0.00 0.00 CaSO„.2H20 0.00 0.00 0.00 n/a 0.00 Sb 0.00 0.00 0.00 0.00 0.00 Se 0.00 0.00 0.00 0.00 0.00 Si 0.00 0.00 0.00 0.00 0.00 Zn 0.00 0.00 0.00 0.00 0.00 C a S 0 4 . 2 H 2 0 96000.00 0.00 169670.61 n/a 34.78 C a C 0 3 0.00 0.00 0.00 n/a 0.00 TSS n/a n/a 280000.00 n/a 57.40 Lime Inerts n/a n/a 26664.52 n/a 5.47 Total 5482.18 487817.32 6000.00 98.88 Balance Check: 100.00 % Solids generation = 487.8 g/L Ultimate sludge percent solids = 50 % Sludge pond lifetime = 20 years Annual Average Data: Plant feed rate = 2,800 L/minute Total dry solids production = 1966.9 tonnes/day 717911.0 tonnes/year Sludge volume accumulated = 2669.3 m3/day 974307.8 m3/year Volume at ultimate density = 2669.3 m3/day 974307.8 m3/year Pond volume required = 19,486,000 m 3 Vessel Sizes Flash Mix Tank: 53 m 3 = 13990 USgal Reactor Tank: 159 m 3 = 41969 USgal Flocculation Tank: 16 m 3 = 4192 USgal Lime Storage Tank: 685 m 3 = 181007 USgal Aeration Requirements Total Iron Content = 6000 mg/L Percent Ferrous Iron = 0 % Oxygen Transfer Efficiency = 20 % Total Flow In = 3000 L/min Total Ferrous Iron = 0.0 kg/min 0.0 kg/hr Aeration required = 0 m 3/hour 0 S C F M Tank Dimensions (no freeboard included) aspect ratio D= 4.6 m or 15.1ft H = 3.2 m or 10.5 ft 1.44 D= 6.5 m or 21.3 ft H= 4.8 m or 15.7 ft 1.36 D= 3.0 m or 9.8 ft H= 2.2 m or 7.4 ft 1.34 H= 10.0 m or 32.8 ft V\=L= 8.3 m or 27.2 ft 1.21 Sludge and Reagent Flowrates Sludge Pond Accumulation Rate Solids Generation = 1463 kg/mln = 3227 Ibs/min Solids Volume = 523 L/min = 138 USgpm Interstitial Water Flow = 1463 L/min = 387 USgpm Total Accumulation Rate = 1986 L/min = 525 USgpm SG Slurry = 1.47 pH Slurry = 9.3 pH Units SG Solids = 2.8 Slurry % Solids = 50.00 % Lime Circuit Lime Dosina Solids Mass = 313 kg/mln = 690 Ibs/min Solids Volume = 130 L/min = 34 USgpm Water Flow = 1773 L/min = 468 USgpm Total Slurry Flow = 1903 L/min = 503 USgpm Slurry SG = 1.10 pH Slurry = 14 pH Units SG Solids = 2.4 Slurry % Solids = 15.00 % Flocculant Dosing Floe Dosing Rate = 49 mg/L effluent treated Flow Into Floo Tank = 5296 L/min = 1399 USgpm Undiluted Floe Flowrate = 52 L/min = 14 USgpm Diluted Floo Flowrate = 517 L/min = 136 USgpm Floe Consumption = 372 kg/day = 820 lbs/day Lime Loop Out Of Storage Tank Lime LOOP Return To Storage Tank Solids Mass = 1251 kg/min = 2759 Ibs/min Solids Mass = 939 kg/min = 2070 Ibs/min Solids Volume = 521 L/min = 138 USgpm Solids Volume = 391 L/min = 103 USgpm Water Flow = 7092 L/min = 1873 USgpm Water Flow = 5319 L/min = 1405 USgpm Total Slurry Flow = 7613 L/min = 2011 USgpm Total Slurry Flow = 5710 L/mln = 1508 USgpm Slurry SG = 1.10 Slurry SG = 1.10 pH Slurry = 14 pH Units pH Slurry = 14 pH Units SG Solids = 2.4 SG Solids = 2.4 Slurry % Solids = 15.00 % Slurry % Solids = 15.00 % Lime Dosing Lime Dosing Rate = Lime Dosing Rate = Average Plant Feed = Daily Consumption = 93.9 gllme/L effluent-treated 78.9 g lime (CaO + inerts)/L \treated 3000 L/minute 341.0 tonnes/day Annual Consumption= 124452 tonnes/year quicklime ra Tank Flows Out Of Flash Mix Tank Solids Mass = 1463 kg/min = 3227 Ibs/min Solids Volume = 523 L/min = 138 USgpm Water Flow = 4773 L/min = 1261 USgpm Total Slurry Flow = 5296 L/min = 1399 USgpm Slurry SG = 1.18 pH Slurry = 7 pH Units SG Solids = 2.80 Slurry % Solids = 23.47 % it Of Flocculator Tank Solids Mass = 1463 kg/min = 3227 Ibs/min Solids Volume = 523 L/min = 138 USgpm Water Flow = 5290 L/min = 1397 USgpm Total Slurry Flow = 5812 L/min = 1535 USgpm Slurry SG = 1.16 pH Slurry = 9.3 pH Units SG Solids = 2.80 Slurry % Solids = 21.67 % Out Of Reactor Tank Solids Mass = 1463 kg/min = 3227 Ibs/min Solids Volume = 523 L/min = 138 USgpm Water Flow = 4773 L/min = 1261 USgpm Total Slurry Flow = 5296 L/min = 1399 USgpm Slurry SG = 1.18 pH Slurry = 9.3 pH Units SG Solids = 2.80 Slurry % Solids = 23.47 % Out Of Settlina Pond Solids Mass = 0 kg/min = 0 Ibs/min Solids Volume = 0 L/min = 0 USgpm Water Flow = 5290 L/min = 1397 USgpm Total Slurry Flow = 5290 L/min = 1397 USgpm Slurry SG = 1.00 pH Slurry = 9.3 pH Units SG Solids = 2.80 Slurry % Solids - 0.00 % Operating Cost Estimate Base Process 3000 L/min 11 August, 2002 Reagent Dose Rate (mg/L treated) Annual Average Plant Flow Rate (L/min) Annual Reagent Consumption (tonnes/year) Reagent Unit Cost (US$/tonne) Annual Reagent Cost (US$/year) Quicklime 78927 2,800 116154.9 100 11,615,000 Flocculant 49 2,800 71.8 3600 258,000 Sub-total $11,873,000 Item Annual Consumption Unit Cost (US$) Annual Cost (US$/year) Electric Power 0.65 million kW-hours 0.05 0 O & M Capital 3 % of capital cost 4,000,000 120,000 O & M Manpower 8 man-hours per day 24 70,000 Sub-total: $190,000 Total Annual Operating Cost $12,063,000 /year (US dollars) Normalized Annual Operating Cost $8.20 /m 3 (US dollars) $31.03 /1000 USgal (US dollars) APPENDIX C D A M HEIGHT C A L C U L A T I O N M A C R O Macro to Calculate Dam Height Sub Conventional_TMA() 'Declaring Variables for Conventional Tailings Dim TailsVolume As Single 'Volume of tailings calculated in spreadsheet Dim TMA Volume As Single 'Volume of T M A calculated in Macro Dim DamHeight As Single Dim FloorWidth As Single 'Width of T M A bottom - specified by user Dim TMALength As Single Dim WallSlope As Integer 'Slope angle of T M A sides - specified by user Dim MaxHeight As Integer 'Maximum height of dam Dim MaxLength As Integer 'Maximum length of T M A Dim bEnough As Boolean Dim n As Integer 'Counter for dam height Dim i As Integer 'Counter for T M A length 'Declaring Variables for Autoclave Tailings Dim TailsVolumea As Single 'Volume of tailings calculated in spreadsheet Dim TMAVolumea As Single 'Volume of T M A calculated in Macro Dim DamHeighta As Single Dim FloorWidtha As Single 'Width of T M A bottom - specified by user Dim TMALengtha As Single Dim WallSlopea As Integer 'Slope angle of T M A sides - specified by user Dim MaxHeighta As Integer 'Maximum height of dam Dim MaxLengtha As Integer 'Maximum length of T M A Dim bEnougha As Boolean Dim j As Integer 'Counter for dam height Dim k As Integer 'Counter for T M A length 'Initializing Variables TailsVolume = Worksheets("TMA").Range("C24").Value FloorWidth = Worksheets("TMA").Range("F33"). Value WallSlope = Worksheets("TMA").Range("F30").Value MaxHeight = Worksheets("TMA").Range("F31"). Value MaxLength = Worksheets("TMA").Range("F37"). Value bEnough = False T M A Volume = 0 DamHeight = 0 TMALength = 0 n = l i = 0 'Initializing Variables for Autoclaved Tailings TailsVolumea = Worksheets("TMA").Range("L34").Value FloorWidtha = Worksheets("TMA").Range("044"). Value WallSlopea = Worksheets("TMA").Range("041"). Value MaxHeighta = Worksheets("TMA").Range("042"). Value MaxLengtha = Worksheets("TMA").Range("048"). Value bEnougha = False TMAVolumea = 0 Page 100 DamHeighta = 0 TMALengtha = 0 j = l k = 0 'Determining Dam Height and TMA Length for Conventional Tailings For n = 1 To MaxHeight If bEnough = False Then DamHeight = n For i = 1 To MaxLength TMALength = i T M A Volume = (DamHeight * TMALength * FloorWidth) + (DamHeight * Tan((90 - WallSlope) * Pi / 180) * DamHeight * TMALength) + (0.5 * DamHeight * Tan((90 - WallSlope) * Pi / 180) * DamHeight * FloorWidth) If T M A Volume > TailsVolume Then bEnough = True Exit For End If Nexti End If Next n 'Determining Dam Height and T M A Length for Autocalved Tailings For j = 1 To MaxHeighta If bEnougha = False Then DamHeighta =j For k = 1 To MaxLengtha TMALengtha = k TMAVolumea = (DamHeighta * TMALengtha * FloorWidtha) + (DamHeighta * Tan((90 -WallSlopea) * Pi / 180) * DamHeighta * TMALengtha) + (0.5 * DamHeighta * Tan((90 -WallSlopea) * Pi / 180) * DamHeighta * FloorWidtha) If TMAVolumea > TailsVolumea Then bEnougha = True Exit For End If Nextk End If Next j Worksheets("TMA").Range("D45"). Value = DamHeight Worksheets("TMA").Range("D49"). Value = TMALength Worksheets("TMA").Range("M56").Value = DamHeighta Worksheets("TMA").Range("M60"). Value = TMALengtha Worksheets("Record").Range("A6"). Value = Worksheets("Input Parameters").Range("C3").Value Worksheets("Record").Range("B6"). Value = Worksheets("Input Parameters ").Range("C4"). Value Worksheets("Record").Range("C6"). Value = Worksheets("Input Parameters").Range("C5").Value Worksheets("Record").Range("D6"). Value = Worksheets("Input Parameters").Range("C6"). Value Worksheets("Record").Range("E6"). Value = Worksheets("TMA").Range("C24").Value Worksheets("Record").Range("F6"). Value = TMA Volume Worksheets("Record").Range("G6").Value = Worksheets("Cost Summary-conventional").Range("C14").Value Worksheets("Record").Range("H6").Value = Worksheets("Cost Summary-conventional").Range("C22"). Value Page 101 Worksheets("Record").Range("I6").Value = Worksheets("Cost Summary-cx>nventional").Range("C24"). Value Worksheets("Record").Range("J6").Value = Worksheets("Cost Summary-autoclave").Range("Cl6").Value Worksheets("Record").Range("K6").Value = Worksheets("Cost Summary-autoclave").Range("C27").Value Worksheets("Record").Range("L6").Value = Worksheets("Cost Summary-autoclave").Range("C29").Value zt = (1750 - 500) / (5000 - 500) 'daily tonnage rate zm = (8-5) / (12-5) 'mine life pt = (0.07 - 0.02) / (0.15 - 0.02) 'sulphide content in tailings R = Rnd() 'generate a random number for dailing tonnage and mine life Rp = Rnd() 'generates a random number for sulphide in tailings IfR<=ztThen Worksheets("Input Parameters").Range("C3").Value = 500 + (5000 - 500) * Sqr(R * zt) Else Worksheets("Input Parameters").Range("C3").Value = 500 + (5000 - 500) * (1 - Sqr((l - R) * (1 - zt))) End If IfR<=zmThen Worksheets("lnput Parameters").Range("C42").Value = 5 + (12 - 5) * Sqr(R * zm) Else Worksheets("Input Parameters").Range("C42").Value = 5 + (12 - 5) * (1 - Sqr((l - R) * (1 - zm))) End If IfRp<=pt Then Worksheets("Input Parameters").Range("C4"). Value = 0.02 + (0.15 - 0.02) * Sqr(Rp * pt) Else Worksheets("Input Parameters").Range("C4").VaIue = 0.02 + (0.15 - 0.02) * (1 - Sqr((l - Rp) * (1 - pt))) End If End Sub APPENDIX D SIMULATION RESULTS Page 103 Distribution of Simulation Results Distribution of Conventional Total Cost in First Set of Simulations 35 -3 25 -E $109 $142 $176 $209 $243 $276 $310 $343 $377 $410 $444 $477 $511 $544 Conventional Total Cost ($millions) Distribution of Autoclave Total Cost in First Set of Simulations E CO 25 £ 20 X L JSL S473 $876 $1,279 $1,682 $2,084 $2,487 $2,890 $3,293 $3,696 $4,098 $4,501 $4,904 $5,307 $5,709 Autoclave Total Cost ($milIions) Page 104 D i s t r i b u t i o n o f C o n v e n t i o n a l T o t a l C o s t in S e c o n d S e t of S i m u l a t i o n s 40 35 30 $86 $92 $99 $106 $113 $119 $126 $133 $139 $146 $153 $160 $166 $173 Conventional Total Cost ($milions) D i s t r i b u t i o n o f A u t o c l a v e T o t a l C o s t in S e c o n d S e t o f S i m u l a t i o n s in 20 -I 15-n n m $ 106 $164 $222 $280 $338 $397 $455 $513 $571 $629 $687 $745 $803 $862 Autoclave Total Cost ($millions) Page 105 Table D - l Simulation Results Mine Mill %S in Conventional Cost Summary Autoclave Cost Summary Life Tonnage Tailings Capital Operating Total Capital Operating Total (Smill ions) ($millions) ($millions) ($mill ions) ($millions) ($mill ions) 5 741 5.14% $94.1 $0.6 $94.7 $18.5 $31.1 $49.6 5 650 7.27% $80.2 $0.5 $80.8 $19.0 $38.1 $57.1 5 597 10.40% $74.7 $0.5 $75.2 $21.4 $49.2 $70.7 5 730 12.80% $92.4 $0.6 $93.0 $22.1 $74.0 $96.1 6 1166 3.70% $87.5 $0.8 $88.3 $18.1 $42.2 $60.3 6 976 3.88% $133.3 $0.8 $134.2 $15.8 $32.1 $47.9 6 825 5.15% $107.6 $0.7 $108.3 $20.4 $34.7 $55.1 6 1079 5.98% $79.1 $0.7 $79.8 $20.8 $60.1 $80.9 6 1069 6.23% $78.1 $0.7 $78.8 $21.0 $62.0 $83.0 6 1144 6.37% $85.4 $0.7 $86.1 $22.4 $67.7 $90.2 6 783 6.38% $100.7 $0.7 $101.4 $21.1 $40.4 $61.5 6 1076 6.50% $78.8 $0.7 $79.5 $21.5 $65.0 $86.5 6 809 6.69% $105.0 $0.7 $105.7 $22.2 $43.8 $66.0 6 992 7.14% $136.0 $0.9 $136.9 $20.9 $57.1 $78.0 6 993 7.50% $136.3 $0.9 $137.1 $21.4 $60.0 $81.4 6 1224 8.10% $93.4 $0.8 $94.2 $26.8 $91.6 $118.4 6 1233 8.68% $94.3 $0.8 $95.1 $27.8 $98.7 $126.5 6 963 8.95% $131.1 $0.8 $131.9 $22.6 $69.2 $91.8 6 769 8.97% $98.6 $0.7 $99.2 $24.5 $55.3 $79.9 6 907 9.57% $121.2 $0.8 $122.0 $22.3 $69.5 $91.8 6 899 10.77% $119.9 $0.8 $120.6 $23.7 $76.7 $100.4 6 997 11.56% $137.0 $0.9 $137.9 $26.6 $91.3 $117.9 7 1386 3.99% $110.5 $0.9 $111.4 $21.8 $53.7 $75.5 7 1751 4.38% $105.5 $1.1 $106.5 $24.2 $82.7 $106.9 7 1497 4.66% $122.9 $1.0 $123.9 $21.6 $67.0 $88.7 7 1688 5.25% $145.4 $1.3 $146.6 $25.0 $92.8 $117.8 7 1599 5.72% $134.7 $1.2 $135.9 $24.7 $95.4 $120.1 7 1518 5.92% $125.2 $1.0 $126.2 $24.0 $83.9 $107.8 7 1428 6.00% $115.1 $0.9 $116.1 $26.6 $79.9 $106.5 7 1403 6.09% $112.4 $0.9 $113.3 $26.3 $79.6 $106.0 7 1519 6.29% $125.3 $1.0 $126.3 $25.1 $88.9 $114.0 7 1286 6.35% $99.8 $0.8 $100.6 $24.8 $75.9 $100.7 7 1581 6.87% $132.5 $1.2 $133.7 $26.9 $112.6 $139.5 7 1780 6.87% $108.0 $1.1 $109.1 $29.8 $126.9 $156.7 7 1758 6.89% $106.1 $1.1 $107.2 $29.5 $125.7 $155.2 7 1555 6.94% $129.6 $1.2 $130.7 $26.7 $111.9 $138.6 Page 106 Table D - l (cont'd) Mine Mill %S in Conventional Cost Summary Autoclave Cost Summary Life Tonnage Tailings Capital Operating Total Capital Operating Total (Smillions) ($millions) ($millions) ($millions) ($millions) ($millions) 7 1759 7.21% $106.2 $1.1 $107.2 $30.5 $131.3 $161.8 7 1312 7.27% $102.5 $0.9 $103.4 $27.2 $88.4 $115.5 7 1526 7.76% $126.2 $1.0 $127.2 $27.6 $109.5 $137.0 7 1728 7.85% $103.6 $1.0 $104.6 $31.2 $140.2 $171.3 7 1705 8.16% $147.4 $1.3 $148.7 $31.4 $143.6 $174.9 7 1621 8.41% $137.3 $1.2 $138.5 $30.5 $140.7 $171.1 7 1651 8.49% $140.9 $1.2 $142.2 $31.1 $144.7 $175.7 7 1578 8.57% $132.2 $1.2 $133.4 $30.0 $139.4 $169.4 7 1438 8.91% $116.3 $1.0 $117.2 $28.4 $118.1 $146.4 7 1728 9.22% $103.6 $1.0 $104.6 $33.5 $164.0 $197.5 7 1665 9.29% $142.6 $1.3 $143.9 $32.6 $159.2 $191.8 7 1491 9.36% $122.1 $1.0 $123.1 $29.9 $128.5 $158.4 7 1700 9.55% $146.8 $1.3 $148.1 $33.9 $167.0 $200.9 7 1609 9.71% $135.9 $1.2 $137.1 $32.3 $160.7 $193.1 7 1775 9.93% $107.6 $1.1 $108.6 $32.9 $181.1 $214.0 7 1684 10.52% $144.9 $1.3 $146.2 $35.1 $180.4 $215.5 7 1772 10.87% $107.3 $1.1 $108.4 $34.2 $196.2 $230.4 7 1681 12.73% $144.6 $1.3 $145.8 $35.7 $217.6 $253.3 8 2007 2.42% $127.4 $1.2 $128.7 $21.8 $55.3 $77.1 8 2203 3.44% $144.4 $1.5 $145.9 $24.4 $91.3 $115.6 8 1958 3.86% $123.2 $1.2 $124.4 $25.6 $82.2 $107.8 8 2312 3.88% $119.1 $1.4 $120.4 $26.6 $107.0 $133.6 8 1800 4.17% $109.7 $1.1 $110.8 $24.4 $81.3 $105.7 8 2076 4.18% $133.4 $1.3 $134.7 $24.9 $93.9 $118.9 8 1852 4.19% $114.1 $1.1 $115.3 $25.1 $84.0 $109.1 8 1842 4.97% $113.3 $1.1 $114.5 $27.1 $97.9 $125.0 8 1983 4.99% $125.4 $1.2 $126.6 $26.2 $105.8 $132.0 8 2314 5.36% $119.3 $1.4 $120.6 $30.4 $142.2 $172.6 8 1982 5.50% $125.3 $1.2 $126.6 $27.0 $113.9 $140.9 8 1930 5.64% $120.8 $1.2 $122.0 $26.7 $113.7 $140.5 8 2428 5.66% $127.3 $1.4 $128.7 $30.6 $157.3 $187.9 8 1355 5.94% $130.4 $1.2 $131.6 $23.5 $92.0 $115.5 8 1828 5.96% $112.1 $1.1 $113.2 $28.7 $113.4 $142.2 8 1826 6.38% $111.9 $1.1 $113.0 $29.5 $121.2 $150.7 8 1964 6.61% $123.8 $1.2 $125.0 $29.4 $134.8 $164.2 8 2090 6.67% $134.5 $1.3 $135.8 $31.0 $144.8 $175.8 8 2196 6.76% $143.8 $1.5 $145.3 $32.5 $168.9 $201.4 Page 107 Table D - l (cont'd) Mine Mill %S in Conventional Cost Summary Autoclave Cost Summary Life Tonnage Tailings Capital Operating Total Capital Operating Total (Smillions) ($milIions) ($millions) ($mill ions) ($millions) ($millions) 8 1850 6.80% $114.0 $1.1 $115.1 $30.7 $130.5 $161.2 8 1979 7.11% $125.0 $1.2 $126.2 $30.5 $145.7 $176.3 8 2294 7.15% $152.5 $1.6 $154.1 $32.6 $186.1 $218.8 8 2373 7.27% $123.4 $1.4 $124.8 $33.8 $195.8 $229.6 8 1876 7.44% $116.2 $1.1 $117.3 $29.8 $144.5 $174.4 8 1977 7.45% $124.8 $1.2 $126.0 $31.1 $152.3 $183.4 8 2278 7.46% $151.1 $1.6 $152.6 $33.1 $192.6 $225.7 8 1990 7.68% $126.0 $1.2 $127.2 $31.7 $158.0 $189.8 8 2068 7.86% $132.7 $1.3 $134.0 $33.4 $168.1 $201.5 8 2418 8.07% $126.6 $1.4 $128.0 $35.9 $220.9 $256.8 8 2088 8.15% $134.4 $1.3 $135.7 $34.2 $175.6 $209.8 8 1931 8.18% $120.9 $1.2 $122.1 $31.9 $163.2 $195.0 8 2035 8.23% $129.8 $1.3 $131.1 $33.7 $172.8 $206.5 8 2321 8.28% $119.7 $1.4 $121.1 $35.2 $217.2 $252.4 8 2059 8.34% $132.0 $1.3 $133.2 $34.2 $177.2 $211.4 8 2347 8.41% $121.5 $1.4 $122.9 $36.1 $223.2 $259.3 8 2035 8.42% $129.8 $1.3 $131.1 $34.0 $176.8 $210.8 8 2095 8.69% $135.1 $1.3 $136.4 $35.3 $187.8 $223.1 8 1282 8.82% $119.5 $1.1 $120.6 $27.0 $116.6 $143.6 8 2426 9.18% $127.2 $1.4 $128.6 $38.6 $251.2 $289.7 8 1906 9.27% $118.8 $1.2 $119.9 $33.7 $181.8 $215.6 8 2167 9.40% $141.3 $1.5 $142.7 $35.7 $229.8 $265.6 8 2129 9.60% $138.0 $1.4 $139.4 $35.6 $230.5 $266.1 8 2068 9.88% $132.6 $1.3 $133.9 $37.4 $209.9 $247.3 8 1863 10.00% $115.1 $1.1 $116.2 $34.2 $189.8 $224.0 8 2459 10.23% $129.5 $1.4 $131.0 $39.3 $280.7 $320.0 8 2080 10.41% $133.7 $1.3 $135.0 $36.2 $220.4 $256.5 8 2449 10.48% $128.8 $1.4 $130.2 $39.6 $286.4 $326.0 8 2061 10.50% $132.1 $1.3 $133.4 $38.2 $220.4 $258.6 8 1439 10.51% $143.5 $1.3 $144.8 $32.8 $168.8 $201.6 8 2024 10.71% $128.9 $1.2 $130.1 $38.0 $220.7 $258.7 8 2028 10.85% $129.3 $1.2 $130.5 $38.3 $224.0 $262.4 8 2120 10.87% $137.1 $1.4 $138.6 $37.5 $257.1 $294.5 8 2213 11.29% $145.2 $1.5 $146.8 $39.7 $278.7 $318.4 8 2054 12.73% $131.4 $1.3 $132.7 $39.8 $265.7 $305.5 8 1796 13.19% $109.3 $1.1 $110.4 $38.6 $240.7 $279.3 8 1918 13.29% $119.8 $1.2 $121.0 $40.8 $259.1 $299.9 Page 108 Table D - l (cont'd) Mine Mill %S in Conventional Cost Summary Autoclave Cost Summary Life Tonnage Tailings Capital Operating Total Capital Operating Total (Smillions) ($mill ions) ($millions) ($mill ions) ($mill ions) ($mill ions) 8 1833 13.60% $112.5 $1.1 $113.6 $39.8 $253.3 $293.2 9 2749 2.97% $151.0 $1.6 $152.7 $25.9 $99.7 $125.6 9 2575 3.40% $137.9 $1.5 $139.5 $25.7 $105.5 $131.2 9 2501 3.80% $132.6 $1.5 $134.0 $28.7 $113.6 $142.3 9 3181 4.13% $152.6 $1.9 $154.5 $32.2 $168.3 $200.5 9 2663 4.39% $144.6 $1.6 $146.1 $29.6 $138.2 $167.8 9 2627 4 .41% $141.9 $1.6 $143.4 $29.3 $136.8 $166.1 9 2880 4.68% $132.6 $1.7 $134.3 $31.1 $171.1 $202.1 9 2751 5.16% $151.2 $1.6 $152.9 $32.6 $162.9 $195.5 9 3022 5.25% $142.0 $1.8 $143.7 $33.7 $196.7 $230.3 9 2602 5.52% $140.0 $1.5 $141.5 $32.0 $164.4 $196.4 9 3016 5.54% $141.5 $1.8 $143.3 $34.7 $206.4 $241.0 9 2475 5.64% $130.6 $1.5 $132.1 $31.0 $159.6 $190.6 9 2908 5.72% $134.5 $1.7 $136.2 $34.1 $205.4 $239.6 9 2944 6.10% $136.8 $1.7 $138.5 $35.4 $221.1 $256.5 9 2913 6.18% $134.8 $1.7 $136.5 $35.3 $221.5 $256.8 9 2637 6.46% $142.5 $1.6 $144.1 $35.0 $194.1 $229.1 9 3122 6.48% $148.7 $1.8 $150.5 $36.9 $248.6 $285.6 9 2476 6.52% $130.7 $1.5 $132.2 $33.0 $183.9 $216.9 9 2571 6.63% $137.6 $1.5 $139.2 $34.6 $193.9 $228.5 9 2503 6.73% $132.7 $1.5 $134.1 $34.1 $191.6 $225.7 9 3172 6.82% $152.1 $1.9 $153.9 $38.3 $265.7 $304.0 9 2730 6.86% $149.6 $1.6 $151.2 $35.1 $212.9 $248.0 9 2708 7.27% $148.0 $1.6 $149.6 $35.8 $223.4 $259.1 9 2509 7.31% $133.1 $1.5 $134.6 $35.4 $208.2 $243.6 9 1599 7.35% $116.9 $1.3 $118.2 $27.8 $143.9 $171.7 9 2914 7.37% $134.8 $1.7 $136.5 $38.5 $262.8 $301.3 9 3026 7.44% $142.2 $1.8 $144.0 $38.3 $275.8 $314.1 9 3082 7.48% $146.0 $1.8 $147.8 $39.0 $282.3 $321.3 9 2741 7.48% $150.4 $1.6 $152.0 $36.9 $232.6 $269.5 9 1561 7.65% $112.5 $1.2 $113.7 $27.7 $135.3 $163.0 9 2723 7.91% $149.1 $1.6 $150.7 $37.6 $243.8 $281.4 9 2591 7.92% $139.1 $1.5 $140.7 $38.0 $232.2 $270.2 9 2556 8.02% $136.5 $1.5 $138.0 $37.8 $232.1 $269.9 9 2624 8.03% $141.6 $1.6 $143.2 $36.8 $238.6 $275.4 9 2508 8.37% $133.0 $1.5 $134.5 $37.9 $237.2 $275.2 9 3140 8.39% $149.9 $1.8 $151.7 $42.0 $321.7 $363.7 Page 109 Table D - l (cont'd) Mine Mill %S in Conventional Cost Summary Autoclave Cost Summary Life Tonnage Tailings Capital Operating Total Capital Operating Total (Smillions) ($mill ions) ($millions) ($mill ions) ($millions) ($mill ions) 9 2583 8.53% $138.5 $1.5 $140.1 $37.4 $248.9 $286.3 9 2551 8.57% $136.2 $1.5 $137.7 $38.9 $247.1 $286.0 9 2524 8.60% $134.2 $1.5 $135.7 $38.6 $245.2 $283.8 9 2492 8.61% $131.9 $1.5 $133.3 $38.2 $242.3 $280.5 9 2519 8.80% $133.8 $1.5 $135.3 $39.0 $250.2 $289.2 9 2993 9.04% $140.0 $1.7 $141.7 $42.6 $329.7 $372.4 9 2980 9.35% $139.1 $1.7 $140.9 $43.1 $339.3 $382.5 9 2612 9.36% $140.7 $1.6 $142.3 $39.7 $275.7 $315.3 9 3154 9.56% $150.8 $1.9 $152.7 $44.5 $366.7 $411.3 9 2526 9.70% $134.4 $1.5 $135.9 $39.3 $276.0 $315.3 9 3046 9.95% $143.5 $1.8 $145.3 $45.5 $368.7 $414.2 9 2861 10.01% $131.4 $1.7 $133.0 $43.0 $345.1 $388.1 9 2786 10.14% $153.9 $1.7 $155.6 $43.9 $315.4 $359.3 9 2744 10.53% $150.7 $1.6 $152.3 $44.2 $322.3 $366.5 9 2529 10.53% $134.6 $1.5 $136.1 $40.7 $297.2 $337.9 9 2687 10.90% $146.3 $1.6 $147.9 $44.2 $326.8 $371.0 9 2620 11.10% $141.3 $1.6 $142.9 $43.7 $324.4 $368.2 9 2795 11.62% $154.6 $1.7 $156.3 $45.7 $362.1 $407.7 9 3121 12.75% $148.6 $1.8 $150.4 $51.1 $478.5 $529.6 9 2665 13.05% $144.7 $1.6 $146.3 $46.9 $387.6 $434.5 9 2776 13.08% $153.2 $1.7 $154.8 $48.4 $404.6 $453.0 9 2722 13.60% $149.0 $1.6 $150.6 $48.9 $412.2 $461.1 10 3254 3.52% $157.7 $1.9 $159.6 $30.5 $148.4 $179.0 10 3519 4.52% $151.1 $1.9 $153.0 $34.7 $202.4 $237.1 10 3667 4.90% $160.4 $2.2 $162.6 $36.2 $242.9 $279.1 10 3270 5.37% $158.8 $1.9 $160.7 $35.1 $217.2 $252.3 10 3389 5.69% $143.1 $1.8 $144.9 $37.1 $238.2 $275.3 10 3785 5.76% $147.7 $2.1 $149.8 $39.8 $287.2 $327.0 10 3377 5.78% $142.3 $1.8 $144.2 $37.2 $240.8 $278.0 10 3767 5.80% $146.7 $2.1 $148.7 $39.7 $287.8 $327.5 10 3817 6.29% $149.6 $2.1 $151.7 $40.7 $315.2 $355.9 10 3324 6.49% $139.1 $1.8 $140.9 $38.9 $265.0 $304.0 10 3803 6.55% $148.7 $2.1 $150.8 $41.3 $326.6 $368.0 10 3530 6.55% $151.8 $1.9 $153.7 $39.7 $284.3 $324.0 10 3788 6.70% $147.9 $2.1 $150.0 $41.6 $332.7 $374.3 10 8021 6.92% $205.2 $3.1 $208.2 $64.2 $681.3 $745.5 10 3477 7.16% $148.5 $1.9 $150.4 $41.1 $305.3 $346.4 Page 110 Table D - l (cont'd) Mine Mill %S in Conventional Cost Summary Autoclave Cost Summary Life Tonnage Tailings Capital Operating Total Capital Operating Total ( S m i l l i o n s ) ( $ m i l l i o n s ) ( $ m i l l i o n s ) ( $ m i l l i o n s ) ( $ m i l l i o n s ) ( $ m i l l i o n s ) 10 3416 7.17% $144.7 $1.9 $146.6 $40.5 $300.1 $340.7 10 3265 7.67% $158.5 $1.9 $160.4 $41.6 $306.4 $348.0 10 3628 7.96% $157.9 $2.1 $160.1 $44.2 $376.4 $420.6 10 3779 7.98% $147.4 $2.1 $149.5 $45.7 $393.4 $439.1 10 3412 8.19% $144.5 $1.9 $146.3 $43.8 $341.3 $385.1 10 1830 8.58% $145.6 $1.5 $147.1 $31.9 $191.6 $223.5 10 1824 9.57% $144.8 $1.5 $146.3 . $33.8 $212.4 $246.2 10 3398 9.59% $143.6 $1.9 $145.5 $47.2 $396.6 $443.9 10 3848 10.11% $151.4 $2.1 $153.5 $51.3 $500.2 $551.5 10 3220 10.34% $155.4 $1.9 $157.3 $46.7 $401.1 $447.9 10 9976 10.35% $217.9 $3.5 $221.5 $88.2 $1,244.0 $1,332.2 10 3739 10.70% $145.1 $2.1 $147.1 $51.5 $514.0 $565.5 10 3263 11.15% $158.3 $1.9 $160.2 $49.2 $437.9 $487.2 10 1945 11.31% $124.3 $1.5 $125.8 $37.1 $282.4 $319.5 10 3846 12.61% $151.2 $2.1 $153.3 $56.8 $622.5 $679.4 10 1870 13.55% $150.9 $1.7 $152.5 $39.5 $324.9 $364.4 10 1239 14.51% $166.0 $1.4 $167.5 $34.9 $216.1 $250.9 10 4566 14.82% $174.3 $2.2 $176.5 $68.4 $812.9 $881.4 11 2103 4.06% $141.5 $1.7 $143.2 $26.1 $116.8 $142.9 11 2198 4.70% $152.4 $1.9 $154.3 $29.3 $147.8 $177.1 11 3956 6.09% $157.6 $2.2 $159.8 $41.4 $316.7 $358.1 11 4607 6.10% $161.6 $2.5 $164.1 $45.7 $390.3 $436.0 11 4445 6.29% $168.2 $2.5 $170.7 $44.9 $388.2 $433.2 11 4241 6.61% $156.7 $2.3 $159.0 $45.1 $367.5 $412.6 11 2041 8.25% $134.7 $1.6 $136.3 $35.0 $219.5 $254.5 11 2244 8.53% $129.8 $1.7 $131.5 $36.8 $263.5 $300.3 11 2241 8.85% $129.4 $1.7 $131.1 $37.7 $272.5 $310.2 11 4517 10.02% $156.8 $2.4 $159.3 $55.7 $615.0 $670.7 11 4373 10.10% $164.1 $2.5 $166.6 $55.0 $600.4 $655.5 11 4453 11.35% $168.7 $2.5 $171.2 $58.7 $686.3 $745.0 11 4261 12.36% $157.9 $2.3 $160.1 $59.4 $676.0 $735.5 11 4489 13.11% $170.7 $2.5 $173.2 $63.4 $798.2 $861.5 11 4104 13.74% $166.5 $2.3 $168.8 $61.8 $722.9 $784.7 12 2518 5.64% $157.7 $2.1 $159.8 $32.5 $207.6 $240.1 12 2535 6.75% $159.4 $2.1 $161.6 $35.6 $248.6 $284.2 12 4645 7.59% $163.6 $2.5 $166.2 $50.3 $486.3 $536.6 12 2274 9.21% $132.7 $1.7 $134.4 $37.3 $287.7 $325.0 12 4670 10.92% $165.0 $2.5 $167.5 $59.1 $692.4 $751.4 12 2370 14.44% $142.2 $1.8 $144.1 $48.1 $463.8 $511.9 Page 111 Table D - l (cont'd) Mine Mill %S in Conventional Cost Summary Autoclave Cost Summary Life Tonnage Tailings Capital Operating Total Capital Operating Total (Smillions) ($millions) ($millions) ($millions) ($millions) ($millions) 12 2504 15.38% $156.2 $2.1 $158.3 $50.9 $544.7 $595.6 12 2537 24.48% $159.7 $2.1 $161.8 $63.9 $876.5 $940.5 13 2626 5.50% $144.7 $2.0 $146.7 $33.4 $211.5 $244.9 13 2547 15.03% $137.5 $1.9 $139.4 $51.0 $541.4 $592.4 13 2701 20.41% $151.6 $2.2 $153.8 $60.6 $810.2 $870.8 14 3148 5.39% $152.6 $2.4 $155.0 $35.8 $268.6 $304.5 14 3026 9.45% $160.0 $2.4 $162.4 $45.7 $445.2 $490.9 14 2972 10.29% $155.5 $2.3 $157.7 $46.9 $454.6 $501.6 14 2986 11.32% $156.7 $2.3 $159.0 $48.6 $501.8 $550.4 14 3098 12.25% $166.0 $2.5 $168.5 $51.4 $583.9 $635.4 14 2969 13.34% $155.2 $2.3 $157.5 $51.8 $587.2 $639.0 14 3087 15.60% $165.1 $2.5 $167.5 $57.6 $736.0 $793.6 15 3578 4.33% $169.1 $2.7 $171.8 $35.9 $261.2 $297.1 15 3431 7.39% $158.6 $2.6 $161.2 $43.0 $410.3 $453.3 15 3333 11.16% $166.8 $2.5 $169.3 $51.1 $572.7 $623.9 15 3394 12.85% $156.0 $2.5 $158.6 $55.4 $692.5 $747.9 15 3540 17.11% $166.3 $2.7 $169.0 $65.1 $955.8 $1,020.9 15 3504 20.54% $163.8 $2.6 $166.5 $71.2 $1,132.8 $1,204.0 16 3724 7.39% $165.4 $2.7 $168.1 $46.1 $445.4 $491.4 16 3956 8.09% $168.2 $2.9 $171.1 $49.3 $531.3 $580.7 16 3662 8.10% $161.2 $2.7 $163.9 $47.3 $478.5 $525.8 16 3664 9.05% $161.4 $2.7 $164.0 $49.3 $533.6 $583.0 16 3651 9.36% $174.3 $2.8 $177.1 $50.2 $549.3 $599.5 16 3846 9.54% $173.5 $2.9 $176.4 $52.2 $606.6 $658.8 16 3937 10.36% $167.1 $2.9 $169.9 $54.7 $667.4 $722.1 16 3719 11.18% $165.0 $2.7 $167.7 $55.0 $660.9 $715.9 16 3767 11.76% $168.2 $2.8 $171.0 $56.6 $704.3 $761.0 16 3854 17.52% $174.0 $2.9 $176.9 $69.7 $1,096.3 $1,166.1 16 3807 20.20% $170.9 $2.9 $173.8 $74.4 $1,244.7 $1,319.1 16 3936 21.90% $167.0 $2.9 $169.9 $79.3 $1,395.9 $1,475.2 17 4374 3.77% $182.6 $3.2 $185.8 $37.7 $296.5 $334.3 17 4214 3.96% $173.0 $3.0 $175.9 $37.8 $291.1 $328.9 17 4073 4.49% $175.6 $3.0 $178.6 $38.8 $316.4 $355.2 17 4306 5.35% $178.6 $3.1 $181.7 $43.0 $397.0 $440.0 17 4104 5.56% $177.6 $3.0 $180.6 $42.6 $383.3 $425.9 17 4150 6.79% $180.5 $3.0 $183.5 $47.2 $470.1 $517.3 17 4051 7.58% $174.2 $3.0 $177.2 $48.9 $510.8 $559.6 17 4498 8.14% $179.3 $3.2 $182.5 $53.4 $622.9 $676.3 17 4221 11.12% $173.4 $3.0 $176.4 $59.2 $767.6 $826.8 Page 112 Table D - l (cont'd) Mine Mill %S in Conventional Cost Summary Autoclave Cost Summary Life Tonnage Tailings Capital Operating Total Capital Operating Total (Smill ions) ($millions) ($millions) ($millions) ($millions) ($millions) 17 4080 11.93% $176.0 $3.0 $179.0 $59.9 $795.7 $855.6 17 4249 12.75% $175.0 $3.0 $178.1 $63.5 $884.8 $948.3 17 4310 13.31% $178.8 $3.1 $181.9 $65.4 $960.7 $1,026.0 17 4431 18.91% $175.5 $3.1 $178.6 $79.2 $1,394.9 $1,474.2 18 5072 2.42% $192.4 $3.6 $196.0 $35.3 $235.0 $270.3 18 4876 4.57% $191.0 $3.5 $194.5 $43.6 $403.3 $446.9 18 4898 4.97% $183.2 $3.4 $186.6 $45.6 $439.0 $484.6 18 4572 6.96% $183.6 $3.3 $186.9 $50.3 $544.0 $594.4 18 4985 13.46% $187.8 $3.5 $191.3 $72.3 $1,148.9 $1,221.2 18 4807 14.73% $187.1 $3.4 $190.5 $73.7 $1,211.6 $1,285.3 18 4681 15.38% $180.1 $3.3 $183.4 $73.9 $1,198.3 $1,272.2 18 4555 20.14% $182.7 $3.2 $185.9 $83.2 $1,522.2 $1,605.5 19 5081 5.22% $192.9 $3.6 $196.5 $47.1 $467.6 $514.7 19 5641 6.82% $197.7 $3.8 $201.5 $57.2 $686.0 $743.2 19 5146 8.45% $187.9 $3.5 $191.5 $59.3 $756.1 $815.4 19 5182 8.86% $189.8 $3.6 $193.3 $60.8 $796.7 $857.5 19 5279 9.68% $194.8 $3.6 $198.5 $64.1 $885.5 $949.7 19 5658 10.24% $198.5 $3.8 $202.3 $68.8 $1,014.3 $1,083.1 19 5566 16.99% $201.5 $3.8 $205.3 $87.7 $1,642.0 $1,729.7 19 5262 17.45% $193.9 $3.6 $197.5 $85.4 $1,562.5 $1,647.9 19 5678 18.32% $199.5 $3.9 $203.3 $89.1 $1,806.4 $1,895.4 19 5379 20.82% $192.0 $3.7 $195.7 $91.6 $1,938.1 $2,029.7 20 5710 3.06% $201.0 $3.9 $204.9 $41.6 $333.3 $374.9 20 6034 4.60% $209.6 $4.0 $213.6 $51.0 $512.5 $563.6 20 6224 6.71% $211.8 $4.2 $215.9 $60.8 $758.4 $819.3 20 5715 6.74% $201.2 $3.9 $205.1 $57.5 $687.1 $744.6 20 6353 7.80% $211.4 $4.2 $215.6 $66.0 $895.9 $962.0 20 5812 8.56% $206.0 $4.0 $210.0 $64.6 $881.9 $946.6 20 5713 9.27% $201.2 $3.9 $205.0 $66.3 $936.3 $1,002.6 20 6023 9.97% $209.0 $4.0 $213.1 $71.3 $1,060.7 $1,132.0 20 5967 10.50% $206.3 $4.0 $210.3 $72.5 $1,096.6 $1,169.1 20 6162 14.49% $208.8 $4.1 $212.9 $87.0 $1,586.1 $1,673.1 20 5749 15.69% $202.9 $3.9 $206.8 $86.2 $1,565.9 $1,652.0 21 6470 4.87% $216.8 $4.3 $221.1 $54.8 $590.2 $645.0 21 6982 5.10% $223.8 $4.6 $228.4 $58.6 $662.5 $721.0 21 6899 5.29% $220.1 $4.5 $224.7 $58.9 $678.6 $737.5 21 6911 5.67% $220.6 $4.6 $225.2 $60.9 $726.8 $787.7 21 6654 5.70% $220.2 $4.4 $224.6 $59.1 $692.8 $751.9 21 6836 5.91% $222.7 $4.6 $227.3 $61.3 $747.9 $809.3 Page 113 Table D - l (cont'd) Mine Mill %S in Conventional Cost Summary Autoclave Cost Summary Life Tonnage Tailings Capital Operating Total Capital Operating Total (Smill ions) ($millions) ($millions) ($mill ions) ($mill ions) ($mill ions) 21 6679 6.05% $221.4 $4.4 $225.8 $61.0 $736.1 $797.1 21 6535 7.03% $213.7 $4.3 $218.0 $64.3 $833.1 $897.4 21 7017 8.26% $225.4 $4.6 $230.0 $72.8 $1,062.3 $1,135.1 21 7065 8.47% $227.6 $4.7 $232.2 $74.2 $1,096.7 $1,170.9 21 6875 8.57% $224.5 $4.6 $229.1 $73.1 $1,079.1 $1,152.2 21 6805 12.88% $221.3 $4.5 $225.9 $88.3 $1,583.3 $1,671.7 21 6421 14.36% $214.5 $4.3 $218.8 $89.3 $1,637.3 $1,726.6 21 6834 14.70% $222.6 $4.6 $227.2 $95.0 $1,812.8 $1,907.8 21 6864 21.47% $224.0 $4.6 $228.6 $112.6 $2,638.5 $2,751.1 22 7511 4.36% $232.3 $4.9 $237.2 $58.5 $636.3 $694.8 22 7499 5.30% $231.8 $4.9 $236.7 $62.8 $749.4 $812.3 22 7164 7.45% $226.8 $4.7 $231.4 $70.6 $981.1 $1,051.7 22 7447 8.48% $234.3 $4.9 $239.2 $77.4 $1,173.1 $1,250.5 22 7118 9.28% $224.7 $4.6 $229.4 $78.0 $1,208.3 $1,286.3 22 7376 10.79% $231.2 $4.8 $236.0 $85.8 $1,439.4 $1,525.3 22 7761 13.15% $238.7 $5.0 $243.7 $99.3 $1,868.3 $1,967.5 22 7175 14.43% $227.2 $4.7 $231.9 $94.2 $1,868.0 $1,962.2 22 7267 14.65% $231.4 $4.8 $236.1 $95.9 $1,920.7 $2,016.6 22 7712 18.54% $236.6 $5.0 $241.5 $114.7 $2,602.5 $2,717.2 22 7576 19.05% $235.2 $4.9 $240.1 $114.5 $2,627.0 $2,741.5 22 7165 21.01% $226.8 $4.7 $231.5 $115.2 $2,694.1 $2,809.3 23 7960 4.63% $242.9 $5.1 $248.1 $62.7 $712.9 $775.6 23 8169 5.46% $247.8 $5.3 $253.1 $68.1 $850.6 $918.7 23 7955 6.37% $242.8 $5.1 $247.9 $71.4 $948.8 $1,020.2 23 8470 8.04% $252.7 $5.4 $258.2 $83.9 $1,282.9 $1,366.8 23 8152 14.38% $247.1 $5.3 $252.3 $104.8 $2,172.9 $2,277.7 23 8260 16.23% $247.6 $5.3 $252.9 $113.1 $2,471.1 $2,584.3 23 8300 18.31% $249.3 $5.3 $254.7 $121.5 $2,801.2 $2,922.7 23 8113 19.39% $245.4 $5.2 $250.5 $123.0 $2,863.9 $2,986.9 23 8159 22.47% $247.4 $5.3 $252.7 $135.0 $3,371.8 $3,506.8 24 8812 5.91% $259.9 $5.6 $265.5 $75.3 $990.2 $1,065.5 24 9008 7.68% $264.7 $5.7 $270.5 $86.7 $1,320.3 $1,407.0 24 8830 8.53% $260.7 $5.7 $266.3 $89.6 $1,433.0 $1,522.6 24 8790 11.91% $259.0 $5.6 $264.5 $101.5 $1,943.4 $2,045.0 24 8653 12.27% $256.8 $5.5 $262.3 $101.7 $1,970.4 $2,072.2 24 8833 13.51% $260.8 $5.7 $266.5 $109.2 $2,238.1 $2,347.3 24 8522 14.99% $255.0 $5.5 $260.4 $111.6 $2,366.8 $2,478.4 24 8963 16.67% $262.8 $5.7 $268.5 $124.1 $2,785.9 $2,910.0 24 9108 18.23% $269.0 $5.8 $274.8 $132.7 $3,096.2 $3,228.9 Page 114 Table D - l (cont'd) Mine Mill %S in Conventional Cost Summary Autoclave Cost Summary Life Tonnage Tailings Capital Operating Total Capital Operating Total (Smillions) ($millions) ($millions) ($mill ions) ($mill ions) ($millions) 25 9488 3.04% $275.2 $6.0 $281.1 $62.2 $591.5 $653.7 25 9764 4 .11% $283.8 $6.2 $290.0 $71.7 $810.3 $881.9 25 9803 5.57% $282.4 $6.2 $288.5 $81.2 $1,063.0 $1,144.2 25 9816 8.63% $282.9 $6.2 $289.1 $99.5 $1,629.0 $1,728.5 25 9185 8.97% $268.8 $5.8 $274.6 $95.0 $1,565.7 $1,660.7 25 9517 9.87% $276.4 $6.0 $282.4 $103.3 $1,782.3 $1,885.5 25 9182 10.52% $268.7 $5.8 $274.5 $102.5 $1,814.8 $1,917.4 25 9427 13.69% $275.8 $6.0 $281.8 $117.1 $2,420.1 $2,537.1 25 9617 14.22% $280.6 $6.1 $286.7 $122.0 $2,590.2 $2,712.1 25 9466 18.39% $274.2 $6.0 $280.2 $138.9 $3,246.9 $3,385.8 25 9800 21.97% $282.2 $6.2 $288.4 $160.8 $4,044.8 $4,205.6 26 10481 6.77% $299.8 $6.6 $306.4 $94.8 $1,384.9 $1,479.6 26 10099 7.39% $292.0 $6.4 $298.3 $95.1 $1,441.4 $1,536.4 27 10690 4.47% $306.0 $6.7 $312.7 $81.2 $968.5 $1,049.7 27 10613 8.92% $302.8 $6.7 $309.4 $109.5 $1,834.6 $1,944.1 27 10877 13.26% $309.0 $6.8 $315.8 $134.1 $2,758.8 $2,893.0 27 10696 14.33% $306.3 $6.7 $313.0 $137.2 $2,928.9 $3,066.2 27 10623 16.44% $303.2 $6.7 $309.9 $147.6 $3,321.3 $3,468.9 28 11424 3.86% $325.1 $7.2 $332.3 $82.5 $911.0 $993.5 28 11537 7.02% $327.7 $7.2 $334.9 $106.7 $1,593.3 $1,700.0 28 11752 7.38% $332.7 $7.4 $340.0 $111.7 $1,716.9 $1,828.5 28 11439 8.10% $325.7 $7.2 $332.9 $113.3 $1,814.8 $1,928.2 28 11599 8.11% $330.3 $7.2 $337.6 $115.3 $1,842.5 $1,957.7 28 11683 13.18% $331.8 $7.3 $339.1 $145.4 $2,991.8 $3,137.2 29 12326 4.38% $349.3 $7.7 $357.0 $94.7 $1,119.5 $1,214.3 29 12494 8.15% $354.6 $7.8 $362.4 $126.1 $2,022.8 $2,148.9 29 12047 10.47% $341.2 $7.5 $348.7 $132.7 $2,455.0 $2,587.7 29 12044 13.43% $341.1 $7.5 $348.6 $152.5 $3,142.4 $3,294.9 29 12379 16.02% $351.5 $7.8 $359.3 $170.0 $3,855.5 $4,025.6 30 13092 4.27% $373.5 $8.2 $381.7 $101.2 $1,170.2 $1,271.4 30 13300 6.54% $379.3 $8.3 $387.6 $122.9 $1,749.5 $1,872.4 30 12755 8.97% $362.3 $8.0 $370.3 $132.5 $2,267.7 $2,400.2 30 13153 15.18% $374.6 $8.2 $382.8 $170.4 $3,906.7 $4,077.1 30 12780 16.06% $363.4 $8.0 $371.3 $172.0 $3,990.8 $4,162.8 31 13377 3.97% $381.2 $8.3 $389.5 $100.8 $1,116.5 $1,217.3 31 14027 4.77% $402.4 $8.7 $411.2 $116.1 $1,397.5 $1,513.6 31 13779 7.98% $394.3 $8.6 $402.9 $138.6 $2,212.7 $2,351.3 31 14060 8.01% $403.8 $8.8 $412.6 $142.6 $2,266.4 $2,409.0 31 13489 8.62% $384.6 $8.4 $393.0 $139.9 $2,321.3 $2,461.1 Page 115 Table D- l (cont'd) Mine Mill %S in Conventional Cost Summary Autoclave Cost Summary Life Tonnage Tailings Capital Operating Total Capital Operating Total (Smillions) ($mill ions) ($millions) ($mill ions) ($millions) ($mill ions) 31 13405 14.29% $382.4 $8.4 $390.7 $168.4 $3,768.0 $3,936.4 32 14488 6.14% $417.7 $9.0 $426.8 $134.7 $1,812.6 $1,947.3 32 14328 6.57% $413.0 $8.9 $421.9 $136.3 $1,903.2 $2,039.5 32 14651 8.81% $423.7 $9.1 $432.9 $149.4 $2,603.3 $2,752.7 32 14574 16.28% $421.4 $9.1 $430.5 $180.6 $4,691.6 $4,872.2 33 15030 4.60% $437.2 $9.4 $446.5 $126.1 $1,453.7 $1,579.8 33 15039 7.46% $437.6 $9.4 $447.0 $143.6 $2,271.4 $2,415.0 33 15070 10.30% $438.1 $9.4 $447.4 $157.5 $3,093.5 $3,251.1 33 14889 11.78% $432.0 $9.3 $441.3 $163.7 $3,492.1 $3,655.7 33 15421 12.39% $450.8 $9.6 $460.4 $168.2 $3,819.9 $3,988.1 34 15474 4 .81% $453.1 $9.7 $462.8 $132.4 $1,567.3 $1,699.7 34 15667 4.90% $460.0 $9.8 $469.8 $133.5 $1,613.8 $1,747.3 34 16136 6.34% $477.9 $10.1 $488.0 $143.5 $2,100.2 $2,243.6 34 15851 7.90% $467.4 $9.9 $477.3 $148.1 $2,552.6 $2,700.8 35 16325 4 .81% $485.1 $10.2 $495.3 $134.4 $1,661.4 $1,795.8 35 16272 10.49% $483.3 $10.2 $493.5 $162.2 $3,429.3 $3,591.5 36 17376 8.41% $528.0 $10.9 $538.9 $155.0 $2,995.3 $3,150.3 36 16949 15.51% $510.4 $10.6 $521.0 $187.1 $5,261.8 $5,448.9 37 17682 12.01% $532.5 $11.0 $543.5 $174.1 $4,292.0 $4,466.1 37 17766 15.47% $532.5 $11.0 $543.6 $190.0 $5,519.3 $5,709.3 38 18537 5.00% $532.7 $11.3 $544.0 $140.4 $1,975.3 $2,115.6 38 18950 7.57% $532.8 $11.5 $544.3 $154.2 $2,968.2 $3,122.3 

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