2009 Proceedings_Processing of Refractory Sulphides at Mantos de or[877175]

$: 2009 Proceedings_Processing of Refractory Sulphides at Mantos de or[877175]$
361

PAPER 22

Processing of Refractory Sulphides at Mantos de Oro, Chile

Kevan J. R. Ford 1 John G. Peacey 2

Leoncio Sevilla G. 3 Erling Villalobos C. 3

1 Kinross Gold Corporation 52nd Floor Scotia Plaza, 40 King Street West,

Toronto, Ontario Canada M5H 3Y2 E-mail: [email protected]

2 Queen’s University

Department of Mining Engineering Professor, NSERC-Xstrata Research Chair

Kingston, Ontario Canada K7L 3N6 E-mail: [email protected]

3 Compania Minera Mantos de Oro Los Carrera 6651, Copiapo Chile

E-mail: [email protected] [email protected]

Key Words: Enargite, Refractory gold, Flotation, Oxidation, Cyanide, Arsenic

January 20 to 22, 2009 Ottawa, Ontario, Canada

41st Annual Meeting of the Canadian Mineral Processors

Proceedings of the 41st Annual Canadian Mineral Processors Conference – 2009

362

ABSTRACT Kinross Gold Corporation’s Mantos de Oro (MDO) gold/silver mine in northern Chile currently processes 15,000 tpd of oxide ores through a mill, whole ore cyanide leach plant followed by a Merrill-Crowe refinery. With the expected depletion of these oxide ores in a few years a new sulphide mineral deposit, Ladera-Farellón has been identified to potentially replace existing ore reserves. The deposit contains approximately 0.5g/t Au, 6g/t Ag and 0.5% Cu. The principal sulphide mineral in the deposit is enargite (Cu3AsS4) which hosts the majority of gold and silver and the copper. Direct cyanidation testwork has shown the sulphide mineral is refractory with a low gold recovery of 36% and silver recovery of 45%. Oxidation of the sulphides is required to liberate gold and silver and allow recovery of copper as an important pay-metal. Therefore, at MDO a new metallurgical process route with oxidation as a pre-treatment step is required for viable treatment of such new sulphide mineralized material with acceptable recoveries of pay-metals. Analysis of flowsheet options was completed based on bench-scale and pilot scale metallurgical testwork as well as with reference to existing studies and operations. The process options investigated assume concentration of sulphides by flotation, followed by either pressure oxidation, bacteriological oxidation or roasting. Preliminary evaluation of flowsheet options focused on process risk, operability, technology availability and the economics of the project with emphasis on use of the existing metallurgical plant to reduce capital cost. Operating costs for each process option were also included in the relative analysis. The challenges of treatment of MDO sulphides, principally enargite and pyrite for recovery of gold, silver and copper center on fixation of arsenic and sulphur capture with environmentally acceptable disposal of residue tailings. This environmental focus was included in the process option selection and comparative analysis. This paper presents a review of the process options considered and highlights comparative advantages/disadvantages of each process with associated environmental issues. Preliminary conclusions are drawn on roasting as a potentially favorable process option for the treatment of MDO sulphides.

INTRODUCTION The La Coipa mine is located in the Atacama Region III of Northern Chile, 800 kilometers north of Santiago and 140 kilometers north-east of the city of Copiapó, Chile. Figure 1 shows the location of the mine. The mine is operated by Compañía Minera Mantos de Oro (MDO), a wholly owned subsidiary of Kinross Gold Corporation of Canada. MDO Operation The mine historically has mined and processed oxide ores from five deposits consisting of Ladera-Farellón, Coipa Norte, Brecha Norte, Can-Can and Puren. Coipa Norte, Brecha Norte and Puren are currently being mined by open pit methods. The current flowsheet to process oxides ores at MDO is shown in Figure 2.


363

Oxide ores are trucked to a primary crusher and ore stockpiled. Crushed ore (-150mm) at 15,000 t/d is fed to a SAG - 2 x Ball mill grinding circuit which produces a cyclone overflow at a P80 of 130 microns. Milled oxide ores are leached in cyanide and pregnant Au, Ag leach solution (PLS) recovered by three counter-current decantation (CCD) thickeners. The PLS is fed to a Merrill-Crowe, zinc precipitation circuit with removal of mercury via retorts and smelting of the calcine to produce Au, Ag dore bars for sale. CCD tailings solids are belt filtered to produce solids tailings of 15 - 18% moisture for disposal onto a dry stack tailings facility. Details of the existing MDO plant equipment and operation are given in Table 1.

Figure 1: Location of La Coipa Mine in Chile

Figure 2: Current MDO Mill Flowsheet


364

Table 1: Existing MDO Mill Plant Equipment and Operation Mill Plant Item Parameter

RoM Ore Crusher – Svedala Cone x 1 CSS 12 inches, 6000 HP installed power SAG Mill, variable motor drive x 1 Fuller 28' x 14', installed power 5,000 kW SAG Mill power draw 2,864 kW SAG Trommel screen 11mm slot Pebble crusher Nordberg Omnicone 1560 Ball Mills x 2 Fuller 16' x 24.5', installed power 3,000 kW Ball Mill power draw 3,000 kW each Ball Mill hydro-cyclones 2 x cluster of 5 cyclones, Cavex 26“ diameter Cyclone overflow P80 130 microns Cyanide leach 8 x 2600m3 leach tanks, continuous overflow CCD 3 x Delkor Thickeners 21.3m diameter Belt Filtration 12 x Delkor vacuum belt filters Merrill Crowe Zinc ppt. 4 x disc filters (56 discs), 4 x Eimco filter presses Refinery 8 x Summit Valley Hg retort-calciner Smelting furnace 2 x Rotary Reverberatory furnace (oil fired) MDO Dore composition 90% Ag, 5% Au

MDO is currently facing a decrease in the oxide ore reserves at La Coipa, with an expected remaining mine life of 3-4 years. A sulphide deposit estimated at approximately 81Mt is located under the Ladera-Farellón (LF) pit, with gold and silver associated mainly with enargite (Cu3AsS4) and some with the significant levels of pyrite which also occur in the sulphide ore. Possible extension of mine life at MDO may be potentially achieved by future exploitation and processing of the LF sulphide deposit.

LADERA FARELLON DEPOSIT Geology and Mineralization The LF deposit has been classified as an Au-Ag-Cu epithermal system of high sulphidation type, with advanced argillic alteration, located in volcanic and volcanoclastic rocks of the La Coipa Volcanic Complex. Figure 3 shows the overall lithological profile of the LF pit in which the sedimentary rocks occur. The sedimentary materials correspond to fine alternation of carbonaceous black lutite, rich in organic matter of lacustrine origin, interstratified with fine-grain feldesphatic sandstone, showing parallel lamination and normal gradation. The mineralization in the LF deposit shows strong vertical zoning with areas rich in silver in the higher levels, corresponding with the oxide zone, passing through a zone of silver and gold mineralization in the middle of the deposit, and ending in a sulphide zone with gold and copper mineralization at depth. The gold-copper mineralization at depth is principally controlled by structures and is generally found in the lutite and arsenic zones. Gold is found principally in its native state or as electrum. Common silver minerals present are native silver, cerargite, embolite, electrum argentite, and argetojarosite. The majority of the copper present in the system occurs as enargite, with smaller quantities of covellite and bornite


365

found with traces of chalcocite and chalcopyrite. Almost all of the arsenic is present as enargite. Iron mineralization in the sulphide zone is principally as pyrite and significant amounts occur with enargite in the sulphide zone.

Figure 3: Section 800NW of Ladera-Farellón Pit

Previous geological data indicated that the deposit is constituted by a relatively homogenous mineral mass relative to lithology and alteration, resulting in the existence of one single geological unit, characterized by a matrix of sedimentary rocks. An extensive drilling program was executed in order to develop the sulphide deposit block model and to provide large composite samples for bench-scale and pilot flotation testwork. Geometallurgical Characterization Geometallurgical characterization of multiple core samples was performed to characterize the variability of the ore and identify sub geological units in the deposit. Analysis of the grade distribution of Au, Cu and Fe from the drill core data base led to the definition of five sub geometallurgical units (UGM1 to 5). A selection of drill core composites was made for each UGM by applying grade range criteria for Cu and Fe. Ranges of Cu and Fe were used to indicate levels of enargite and pyrite and allowed evaluation of gold deportment associated with these sulphides. Assays of composites representing each UGM, as used in the testwork program are shown in Table 2.

Table 2: Composition of Ladera-Farellón Geometallurgical Units Geometallurgical Unit

Cu %

Fe %

As %

S %

Au g/t

Ag g/t

UGM1 0.43 – 0.48 5.74 – 5.87 0.14 11,6 0.41 5.0 UGM2 0.36 – 0.40 3.84 – 3.87 0.17 9.3 0.39 4.0 UGM3 0.66 – 0.66 5.65 – 5.78 0.23 11,6 0.59 4.0 UGM4 0.51 – 0.55 7.30 – 6.78 0.20 13,7 1.22 7.0 UGM5 0.69 – 0.72 5.62 – 6.04 0.27 12,4 0.40 5.0


366

The relative distribution of sulphide minerals in each UGM was determined based on Fe and Cu species. High Fe and S levels indicate significant proportions of pyrite with copper present mainly as enargite with minor amounts of chalcocite and covellite. Table 3 shows the relative proportions of minerals calculated for each UGM.

Table 3: Relative Distribution of Sulphide Species in Ladera-Farellón UGM’s

Sulphide Mineral UGM 1 UGM 2 UGM 3 UGM 4 UGM 5 Enargite Cu3AsS4 5.92 9.24 9.02 8.34 11.65 Pyrite FeS2 92.83 90.76 90.04 90.79 88.35 Chalcocite Cu2S 1.25 - - - - Covellite CuS - - 0.94 0.87 - Total 100.00 100.00 100.00 100.00 100.00

An extensive geometallurgical testwork program was conducted by SGS Lakefield Research Chile and included:

• Bond Work Index tests (BWI) • Drop Weight Test (DWT), abrasion index (Ai) determination and SMC tests to assess

existing crushing and grinding capacity • Batch rougher and cleaner flotation tests with tailings filtration • Pilot scale testwork to verify flotation performance and provide large concentrate

samples for downstream testwork and cyanidation tests

From this testwork process parameters were evaluated for milling the ore and concentrating the sulphides for downstream processing and cyanidation. SULPHIDE ORE CONCENTRATION Milling Characterization of each UGM ore was determined to assess whether the existing SAG - Ball mill circuit would be able to produce the required grind at a given throughput for downstream flotation and/or cyanidation of the harder LF sulphide ores. Refer to Table 1 and Figure 2 for details on the existing MDO circuit. Full PQ core samples were used to perform JK Drop Weight and SMC testwork on the 5 UGM composite samples [SGS Report #1 - Milling and Flotation, 2008]. Bond Ball Work Index (BWI) values for the five UGM samples were similar; ranging from 12.8 to 13.4 kWh/t. Measured BWI values indicated that the LF ore falls within a medium hardness range. Values for SMC tests on the 5 UGM samples were also similar, with Axb values within a range from 32.8 to 42.2, matching ores of a medium hardness range. In summary, the impact testwork classified the LF ores in the medium-hard range with regard to SAG milling. The existing MDO grinding circuit was evaluated with a plant survey and a JKSimMet


367

simulation model developed to predict the current operation. This model was used to simulate milling of LF sulphide ores with a target P80 grind size of 130 microns. Simulations indicated the following mill throughput performance, when milling current and future LF ores, as summarized in Table 4:

• The current SAG mill operates at very low speeds (50% of critical) and easily attains a throughput of 700 t/h or 15,000 t/d.

• Increasing the speed of the SAG mill to more normal speeds of 75% of critical, allows the mill circuit to achieve a required throughput rate of 700 t/h or 15,000 t/d when grinding the harder sulphide ores.

• The simulation work indicated that no additional ball mill will be needed to achieve targeted LF ore feed rates of 700 t/h or 15,000 t/d.

• A third ball mill significantly increases the maximum possible mill throughput rates to 920 t/h or 20,000 t/d.

Table 4: MDO Mill Throughput Rates for Oxide and LF Sulphide Ores

Ore Feed to Mill Mill Ore Throughput t/hr Mill Operation (for P80 = 130 microns)

Current Oxide ores 500 - 800 Pre-crushing (45mm CSS), SAG Mill + 2 Ball mills, SAG at 50% of critical speed, 6” ball size

LF Sulphide ores 700 Pre-crushing (45mm CSS), SAG Mill + 2 Ball mills, SAG at 75% of critical speed and 12% ball charge, 5” ball size

LF Sulphide ores 920 SAG + 3 Ball mills

Open-pit mine modeling, at pre-feasibility study level was completed based on the LF block model. A mining schedule to supply ore to the existing crushing and grinding plant at 15,000 t/d was developed to give a potential life of mine of approximately 15 years. Average mill feed grades from the LF deposit were estimated at 0.54 g/t Au, 5.7 g/t Ag, 0.48 % Cu. Flotation With relatively low grades of Cu, Au and Ag in LF sulphide ores, it is expected that these paymetals would be concentrated and efficiently recovered by flotation of the sulphides. From mineralogical analysis of each UGM representing the LF sulphide deposit, the presence of enargite with high levels of pyrite suggested a flotation strategy of bulk rougher flotation followed by cleaner flotation to collect the majority of the copper with gold/silver into an enargite concentrate. Depression of pyrite into the scavenger tails would allow either direct cyanidation of the tails for gold recovery or refloating of a pyrite concentrate for sulphuric acid manufacture. Subsequent gold recovery from calcined pyrite could be added. Flotation testwork therefore focused on verifying the recovery and grades of Cu, Au and Ag in a rougher concentrate, followed by separation and cleaning of an enargite concentrate. Effective separation of enargite from pyrite was seen as important to allow separate treatment


368

of the enargite for copper recovery and arsenic removal. However, recovery of refractory gold associated with the pyrite in the scavenger tailing is considered as a key factor in the flowsheet. The proposed flotation flowsheet is shown in Figure 4. A first stage rougher concentrate is floated to maximize sulphide recovery. The bulk rougher concentrate is reground to a P80 = 38 microns and processed through 3 stages of cleaning. The 1st stage cleaner tail is fed to a scavenger float stage to further recover enargite for recycle and increase the rejection of pyrite to the scavenger tails.

Figure 4: Flotation Flowsheet for LF Sulphides Rougher flotation tests were applied to each UGM composite, and performed at pH 10, using a reagent suite comprising collectors AP 5745E and AX 343 with frother MIBC/D-250 [SGS Report #2 – Flotation Testwork, 2007]. Grind versus Cu, Au recovery was investigated across a grind range of P80 = 90 to 150 microns and showed an optimum grind of P80 = 130 microns. Cleaning testwork was completed at an optimized grind of 38 microns at pH 12.0. Rougher flotation kinetics were generally fast with over 93% Cu and over 90% Au recovery achieved in 10 minutes. Cleaner flotation testwork resulted in Cu recoveries ranging from 87.6% - 91.7% with grades from 3.7 %Cu to 8.7 %Cu. The full flowsheet in Figure 4 was evaluated with closed cycle float testwork on each UGM composite sample and gave the following results, summarized in Table 5.


369

Table 5: Closed Cycle Flotation Testwork Results for LF UGM's Grade Metal Recovery

UGM Product Cu %

Fe %

Ag g/t

Au g/t

Cu %

Fe %

Ag %

Au %

Cleaner Concentrate 13.2 25.8 72.6 7.4 84.7 12.3 43.9 50.4 Combined Ro + Scav Tail 0.1 5.01 2.5 0.2 15.3 87.7 56.0 49.7 UGM1 Rougher Concentrate 1.9 26.4 19.5 1.3 93.3 91.4 83.1 87.8 Cleaner Concentrate 15.7 21.8 97.1 14.0 87.1 11.1 51.1 57.4 Combined Ro + Scav Tail 0.1 3.3 1.8 0.2 12.9 88.9 48.9 42.6 UGM2 Rougher Concentrate 2.0 21.9 17.8 2.6 93.5 91.6 76.4 90.4 Cleaner Concentrate 24.9 14.9 145.1 15.5 89.4 9.5 54.2 61.4 Combined Ro + Scav Tail 0.1 5.6 4.8 0.4 10.6 90.5 45.8 38.6 UGM3 Rougher Concentrate 4.7 25.9 41.3 4.3 96.5 92.4 84.0 94.0 Cleaner Concentrate 11.5 24.5 108.5 25.6 84.8 14.4 52.9 73.5 Combined Ro + Scav Tail 0.1 5.7 3.8 0.4 15.2 85.6 47.1 26.5 UGM4 Rougher Concentrate 2.2 27.6 27.9 5.7 94.3 92.2 36.7 92.5 Cleaner Concentrate 12.7 26.1 74.7 4.2 91.8 23.0 58.9 65.6 Combined Ro + Scav Tail 0.1 4.4 2.6 0.1 8.20 77.0 41.1 34.4 UGM5 Rougher Concentrate 3.1 25.4 25.5 1.5 96.3 94.0 84.9 90.9

Pilot flotation was testwork was completed on 26 tonnes of LF core samples representing all UGM’s in the deposit. The ore samples head grade averaged 0.62% Cu, 0.73g/t Au, 9.64 g/t Ag with a sulphur grade of 11.3%. On average the pilot testwork produced final cleaner (3rd stage) concentrate grades of 22.9 - 28.8% Cu, 16.1 - 17.6g/t Au, 196 - 242 g/t Ag 32 - 36.0%S, 14.8 - 15.1 % Fe and 7.0 - 9.3% As. Sampling of the circuit gave average recoveries for Cu, Au and Ag that were lower than the closed cycle test work, and ranged from 89.7 % - 95.6% Cu, 66.3 - 85.4% Au and 55.2% - 93.5% Ag. The pilot plant work provided about 500kg of bulk cleaner enargite concentrates for downstream pressure oxidation testwork. In summary from the flotation testwork on LF Sulphides, the following average recoveries and grades in Table 6 were assumed across the ore concentration step in evaluating LF process flowsheet options. Table 6: Ore Grades and Flotation Recoveries applied to LF Sulphide Ore Concentration

Solids Flowrate Grade % Metal Recovery Metal Production

Flotation Stream

t/d Au g/t

Ag g/t

Cu % Au Ag Cu Au

oz/d Ag

oz/d Cu t/d

Ore Feed 15,792 0.56 6.0 0.55 100.0 100.0 100.0 284 3,046 86.86

Rougher Concentrate 2,712 2.76 31.0 3.00 84.6 88.7 93.7 240 2,703 81.36

Cleaner Concentrate (Enargite)

295 19.9 153.6 25.00 66.4 47.9 85.0 189 1,458 73.80

Scavenger Tails (Pyrite) 2,424 0.91 11.3 0.32 24.9 28.8 8.9 71 877 7.76


370

Cyanidation Direct cyanidation of the scavenger tails from the cleaning flotation stage for recovery of gold and silver is envisaged in order to use the existing cyanidation agitation tanks used to treat MDO oxide ores. Gold and silver recovery from pregnant leach solutions (PLS) would be completed in the existing Merrill-Crowe refinery to precipitate the Au and Ag with zinc powder to produce dore metal. Any CN-Cu species in the PLS, leached from enargite and other copper sulphides unrecovered in the scavenger float step would be recovered in a SART-AVR plant and precipitated as Cu2S. This copper precipitate would join the cleaned enargite concentrate for downstream recovery of copper. The rougher tails, as the main mass flow stream from the flotation plant together with the leached scavenger tails would be filtered using the existing belt filters and discarded to tailings with about 18% moisture. Bottle-roll cyanidation leach tests on scavenger tails from flotation of each UGM composite gave low gold and silver recoveries. Gold leach recoveries ranged from 21 - 35% with silver recoveries at 19 - 41%. High CN consumptions were measured at 3.3 - 7.2kg NaCN/t indicated consumption by copper minerals residual in the scavenger tailings, with copper concentrations averaging 0.16 % Cu. Re-milling of the tailings even to a P80 of 12 microns did not significantly increase gold dissolution. These leach results clearly indicated the refractory nature of gold and silver in the LF sulphides. Mineralogical analysis of scavenger tailings and cleaner concentrates confirmed the refractory nature of gold occurring in both enargite and pyrite. The photo-micrographs in Figure 5 show the encapsulation of gold in enargite, pyrite and bornite at sizes below 10 microns. Geo-mineralogical work on LF ores indicates that gold predominantly occurs as fine colloidal gold locked in the sulphides, with some liberated gold occurring as liberated, un-encapsulated gold. Direct cyanidation of the scavenger tailings using the existing cyanide leach - Merrill-Crowe zinc precipitation circuits at MDO, albeit at low gold and silver leach recoveries, is a valuable enhancement of the overall recovery of gold and silver at MDO. The gold and silver in the scavenger tailings represent about 25% and 29% by mass of the gold and silver respectively in the mill feed. Copper deportment to these tailings is about 9% of copper in mill feed. In order to achieve overall high gold and silver recoveries from LF sulphides at MDO, a process flowsheet would require oxidation of sulphide minerals to liberate the gold and silver within the sulphide matrix. The gold and silver in the cleaner float concentrate represent about 60% and 43%, respectively by mass of these paymetals in the feed. Importantly, the sulphide mineral oxidation step would also focus on the recovery of copper, which represents about 45% by value of the paymetal content in the LF ore feed, as well as on fixation and disposal of sulphur and arsenic products from the oxidation step.


371

Figure 5: Gold Encapsulation in Ladera-Farellón Sulphides Overall MDO Flowsheet With flotation established as the concentration step to produce an enargite concentrate (3rd stage cleaner concentrate) with depression of pyrite to the scavenger tails, the potential overall flowsheet at MDO to treat LF sulphides is schematically represented in Figure 6. A milling rate of 15,000 t/d is assumed using the existing crushing and grinding, followed by rougher flotation, 3 stages of cleaning with scavenger float step to produce a pyrite tails. These tails would be leached in the existing continuous cyanidation tanks followed by 3 stages of counter-current decantation (CCD). The washed leached solids will be filtered using the existing belt filters and the final solids deposited on the dry stack tailings site. Belt filter filtrate with the PLS liquor from CCD no.1 and will feed the existing Merrill-Crowe zinc precipitation and gold /silver refinery plant. The enargite concentrate from the 3rd stage cleaners is fed to the oxidation process step for recovery of copper and fixation of sulphur and arsenic. Possible inclusion of a SART-AVR plant at MDO would allow recovery of any Cu in the CN leach PLS, precipitated as a Cu2S precipitate via the SART process. Such a Cu2S precipitate would be sold to a copper smelter.

Gold in Bornite

Gold in Enargite

Gold in Pyrite

Gold in Pyrite


372

At this stage pyrite in the scavenger tails, associated with gold and silver values (scavenger tails grades at 0.8 -1g/t Au and 12.5g/t Ag) would not be recovered for oxidation and fixation of sulphur as sulphuric acid. The tails sulphur concentration would be diluted by the float rougher tails and presently is not seen as an environmental limit to permitting of new process facilities at MDO for the treatment of LF sulphides. However, with the sulphur in the tails assumed in the form of pyrite FeS2 and as unrecovered enargite Cu3AsS4 total sulphur output at 600 t/d could be fixed and sold as sulphuric acid at approximately 2000 t/d. The operational viability and economics of acid production at these volumes at the altitude of MDO (3800 masl) is being investigated as part of the process options being considered for treatment of MDO sulphides.

Dry Stack Tailings

CN Leach Residue to Dry Stack Tailings

Crushing and

GrindingFlotation

Enargite Oxidation

LF Sulphide Ore

Cyanidation + SART

Au, Ag , Cu2S ppt

Copper Recovery

Sulphuric Acid

Production

Oxidized Enargite

Residue to Cyanidation

Rougher Tails

Pyrite Oxidation

As fixed as Scorodite

S fixed as Gypsum or Sulphuric acid

Enargite Cleaner

Conc

Pyrite Scavenger

Tails

Oxidized Pyrite Calcine

to Cyanidation

Figure 6: Overall MDO Flowsheet for Ladera-Farellón Sulphides

ENARGITE PROCESSING Process Selection Enargite as a complex copper-arsenic sulfide (Cu3AsS4) mineral, with refractory gold and silver values, poses many process challenges for the recovery of gold, silver and copper. Enargite concentrates are smelted directly by a few Cu smelters but most smelters limit their arsenic input for environmental reasons and only process clean copper concentrates containing less than about 0.3% As. Although penalties for excess levels of As in copper concentrates are high, potentially direct smelting of an MDO enargite concentrate by a smelter in Chile may be an attractive process route. Studies to date have not considered this route. Other methods for treating enargite concentrates are similar to those which are well proven


373

and used commercially for pre-treating sulphidic refractory gold ores and concentrates containing mainly pyrite and arsenopyrite [Marsden and House, 2006]. These processes are:

• Pyrometallurgical oxidation with air in roasters;

• High temperature pressure oxidation using tonnage oxygen in autoclaves;

• Biological oxidation, usually with air, in stirred tank reactors. Heap leaching is still under development.

Roasting The only process used commercially to-date for treating large quantities of enargite concentrate is partial roasting at temperatures in the range 600-7500C under either a slightly oxidizing or reducing atmosphere in either multiple hearth (El Indio) or fluidized bed furnaces (Boliden and Lepanto), with the resulting low-arsenic calcine being sold to Cu smelters for recovery of the Cu, Au and Ag values. Previous enargite roasting operations, such as Barrick’s El Indio mine produced arsenic trioxide that was sold as a commercial product [Smith, Foster, Minet and Cauwe, 1985]. Today the market for As2O3 has drastically diminished and the sale of large amounts of arsenic trioxide is no longer possible. The sulphur dioxide in the roaster gases was also discharged to the stack. Any future enargite roaster in Chile would be required to fix the arsenic and sulphur dioxide using the best available technology. The scrubbing of arsenic trioxide from copper and nickel smelter gases and its fixation in an environmentally acceptable manner is well-proven by various methods at several copper smelters with significant arsenic inputs (Horne, Chuquicamata, Saganoseki and Kennecott). The key issue in selecting the preferred roasting process flowsheet (Roasting) is minimizing the cost of arsenic fixation and disposal to satisfy environmental regulations. Using oxidizing roasting, the SO2 generated is usually recovered as sulphuric acid for sale. Pressure Oxidation Many hydrometallurgical processes have been developed for treating chalcopyrite concentrates and, although none are yet operating on a full commercial basis, most have also been proposed for the treatment of enargite concentrates. Enargite concentrates can be even more refractory than chalcopyrite concentrates and usually contain significant gold and silver values. The most developed copper hydrometallurgical process is the high-temperature total pressure oxidation process (HTPOX). This process is commercially-proven for refractory gold ores and has been successfully demonstrated by Phelps Dodge on clean chalcopyrite concentrates in Arizona. Although hydrometallurgical copper processes are more sensitive to changes in concentrate composition and mineralogy than pyrometallurgical processes, HTPOX is the least sensitive as it operates at the highest temperature and oxidizes virtually all the sulphides to sulphates. This makes HTPOX the hydrometallurgical equivalent of smelting but at the expense of high oxygen and acid neutralization requirements. Silver recovery by cyanidation of HTPOX leach residues is, however, low due to the formation of argento-jarosite and the proposed lime boil process to recover silver is not considered an attractive option.


374

The Platsol® process (Platsol) is the HTPOX process modified by the addition of up to 20 g/L NaCl, to solubilize gold and PGM values for recovery from the leach solution and to precipitate most of the silver as AgCl for recovery from the leach residue by cyanidation [Ferron, Fleming, O'Kane and Dreisinger, 2002]. Of the hydrometallurgical copper processes, Platsol was considered by Kinross as a favorable option because it alone offers high silver recovery as well as high copper and gold recoveries. HTPOX and Platsol processes also have the advantage over roasting of being able to leach sufficient iron to precipitate most of the arsenic directly as ferric arsenate in the leach step and to-date HTPOX leach residues after cyanidation have proven to be stable for disposal at several refractory gold HTPOX operations (e.g. Campbell Red Lake and Sao Bento). Bio-leaching As yet there are no commercial bio-leaching processes for enargite concentrates. BHP Billiton and CODELCO tested a large 25,000 tpa Cu demonstration plant, based on the BIOCOP process, near to the Chuquicamata mine in North Chile. However, the facility has since been closed down. Stirred tank bio-leaching was evaluated for MDO but high electrical energy demand to operate the reactors at altitude ruled out this option. For this study, it was decided that the “GEOCOAT®” process from GeoBiotics, LLC (Geocoat) looked the most promising bio-leaching process for enargite leaching and preliminary testing. Supporting this was the operation of a full scale continuous GeoCoat plant at the APM Agnes Mine, in South Africa in 2003 which treated 15,000 tonnes per annum of a gold/arsenopyrite concentrate. The mine is no longer in production due to uneconomical grades and production rates [Bath, Norton and Gunn, 2005]. In the Geocoat process, concentrate slurry is coated on to a crushed, sized support rock, which may contain paymetal values and/or be relatively unreactive to the acidic leach conditions. The coated support rock is stacked on a lined pad for bio-oxidation using either mesophilic, moderate or thermophilic bacteriological systems. Leached Cu in the PLS would be recovered by conventional SX/EW and the levels of Fe and As in solution would be controlled by bleeding solution to a neutralization stage to produce an iron-arsenic precipitate for disposal. On completion of the sulphide oxidation and Cu recovery, the oxidized concentrate would be recovered from the support rock and leached by cyanidation to recover gold and silver values. Process Flowsheets and Testwork Discussion on each process; Roasting, Platsol and Geocoat follows with process flowsheets for each shown in Figure 7, Figure 8 and Figure 9, respectively. Testwork was carried out for the Platsol and Geocoat options and preliminary results are included. Roasting testwork has not been completed at this stage, as it was considered to be sufficiently well proven based on previous commercial enargite concentrate roasting operations. Roasting In the Roasting option flowsheet in Figure 7, the enargite concentrate is partially roasted with air and fuel in a fluidized bed furnace. Most of the calcine will be recovered in the hot


375

cyclones with a smaller amount from the bed overflow. After the hot cyclones, the off-gases are after-burned with air to ensure complete oxidation of arsenic sulphide and sulphur to arsenic trioxide and SO2. The off-gas, at about 850

oC, is cooled in a water spray cooler to

about 350oC and then further cleaned in a hot electrostatic precipitator (ESP). Most of the

small amount of remaining calcine is recovered as a dust from the ESP. The calcine product is cooled and stored for shipping to copper smelters. The off-gas from the ESP is quenched with water to about 70

oC and then scrubbed in a high-pressure drop venturi scrubber. The

arsenic trioxide is dissolved into the venturi scrubber solution. The As-containing scrubber solution is treated in an effluent treatment plant to produce an environmentally-acceptable iron-arsenic precipitate for disposal. The SO2-containing gases from wet gas cleaning are sent to a sulphuric acid plant to recover sulphuric acid for sale. Platsol In the Platsol process flowsheet in Figure 8, the enargite concentrate slurry is pressure leached with oxygen injection in an autoclave in an acidified sulphate solution, containing small amounts of chloride ion (10-20 g/L Cl as NaCl), at a temperature of about 220

oC and

around 10atm oxygen pressure. The residence time for the slurry in the autoclave is 1-3 hours. Almost all the sulphides are oxidized to sulphate and almost all of the copper and gold are extracted into solution. Most of the arsenic is precipitated in the autoclave as a scorodite type precipitate. Part of the silver is precipitated as silver chloride enhancing silver recovery from the leach residue by cyanidation. Gold leached by the chloride ions is cemented out using copper powder added into the autoclave circuit. The copper-containing solution from solid-liquid separation is neutralized prior to recovery of copper by solvent extraction and electro-winning. Sulphates produced are neutralization with lime to form a gypsum precipitate which may be co-disposed with the float tails. The enargite concentrates obtained from pilot flotation tests were sent to Dynatec to conduct continuous HTPOX pilot tests using the Platsol process to improve the silver recovery. The tests included neutralization of the effluent, cyanidation of the residue and SX/EW of the copper in autoclave liquor. Key results are summarized in Table 7 and Table 8 and discussed below.

Table 7: Enargite Concentrate Sample Composition for Platsol Testwork Blended Concentrate Analysis Enargite

Concentrate Sample Cu % Fe % As % S % Au g/t Ag g/t

Representative 22.9 15.1 7.01 32.1 16.1 196

Bulk 24.1 12.0 7.82 29.1 12.1 143

The HTPOX product solids (thickener underflow) contained 10% As, 1.2 to 2.1% Cu, 15 to 21% Fe and 7 to 8% S principally as sulphate. The arsenic precipitated as a scorodite type phase in the HTPOX residue. After primary neutralization, the product solution (thickener overflow) contained on average 40 g/L Cu, 4.2 g/L Fe, 1.33 g/L As and 5.9 g/L H2SO4 with an average ORP of 460mV. Primary neutralization thickener underflow solutions averaged


376

<0.06% Cu, 1.1 to 3.0% As, 1.7 to 4.0% Fe and about 3.0% Si. Paymetal recoveries of copper, gold and silver from the Platsol testwork are summarized below in Table and show that high gold recoveries with moderate silver extraction via the Platsol process may be achieved.

Table 8: Platsol Testwork Paymetal Recoveries * Paymetal Process Step % Recovery Copper Autoclave Leaching 95 - 96 Gold HTPOX Residue Cyanidation 97 - 99 Silver HTPOX Residue Cyanidation 70 - 75

* At 98 % sulphur oxidation

Geocoat In the Geocoat process, shown in Figure 9, rougher float concentrate slurry is coated onto crushed, sized support rock and continuously stacked on a re-usable lined pad. The heap is leached under dilute sulphuric acid solution conditions for about 120 days. The oxidation leach process is aided by bacteria, which depending on bacteria type used operates favourably between temperatures of 33 – 65

oC. Levels of iron, arsenic and free acid in solution are controlled by

neutralization and the iron-arsenic precipitate is sent to disposal. The copper-containing PLS is subjected to conventional SX/EW steps to recover copper cathode. When the target copper extraction and sulphide oxidation is reached the heap material is reclaimed and washed. The oxidized concentrate is recovered and cyanided to recover the gold and silver values. Support rock is washed, recycled and recoated with fresh concentrate.

GeoBiotics was retained by MDO to evaluate the Geocoat bio-leach process to treat the rougher float concentrate from LF sulphide ores. Both batch amenability and column leach tests have been completed. Summary testwork results used to evaluate the Geocoat process are presented in Table 9 and Table 10.

Table 9: Rougher Float Concentrate Composition Rougher Concentrate Analysis

Cu % Fe % As % S % Au g/t Ag g/t

Rougher Float Concentrate Sample For Geocoat 4.8 9.2 1.7 18.0 3.1 41

Table 10 Geocoat Paymetal Recoveries

Paymetal Recovery % Bacteria (leach temperature oC) Au Ag Cu Base-line 66.9 7.2 19.2

Mesophile (33 - 35 oC) 87.8 16.0 46.8

Moderate (43 - 45 oC) 91.9 46.6 60.6

Thermophile (63 - 65 oC) 94.2 64.3 82.6


377

Arsenic in Scrubbing Solution

Calcine to Smelter

Arsenic Residue to Disposal

Sulphuric Acid

Off-Gas As, SO2 SO2

Roaster

Wet Gas Cleaning

Hot Gas Cleaning

Acid Plant

Effluent TreatmentPlant for Arsenic

Removal

Enargite Concentrate

Figure 7: Roasting Flowsheet for Oxidation of Enargite Concentrate

Enargite Concentrate

HTPOX Autoclave

S

L

CN and SART

CN Leach Residue (As)

Leach Residue

Limestone

Cu2+

SO42-

NeutralisationSX

EWCopper

Gypsum

Au, Ag Dore

Figure 8: HTPOX (Platsol) Flowsheet for Oxidation of Enargite Concentrate


378

Support Rock Recycle

Leach Residue

Cu Containing PLS

GEOCOAT Bioxidation Heap Leach

Neutralisation, As Precipitation

SX

Rougher Float

Concentrate

Support Rock

EW

Cyanidation and SART

Gold, Silver

Recovery

Au, Ag

CopperLimestone

Fe/As Precipitation

Residue to Disposal

Figure 9: Geocoat Heap Bio-leaching of LF Rougher Float Concentrates

TECHNO-ECONOMIC EVALUATION Based on an initial mining plan and ore grades with the testwork results, a preliminary techno-economic evaluation was carried out on each of the process option flowsheets. Process recoveries, operating requirements and capital cost estimates for the three process options were developed at a concept study to pre-feasibility level (+/- 35% capital costs and +/- 20% for operating costs). For the evaluation the following paymetal grades were applied for the enargite concentrate feed (Roaster and Platsol process options) and for the rougher concentrate feed (Geocoat), respectively. Enargite cleaner concentrate: 25.6% Cu, 6.5% As, 26.9 g/t Au and 336 g/t Ag and Rougher concentrate 3.0% Cu, 3.1 g/t Au and 41 g/t Ag). The design basis for each option, in terms of ore and concentrate tonnages and grades is summarized in Table 11. Overall copper, gold and silver recoveries assumed for the three options are summarized in Table 12. The recovered and payable metals for each option are summarized in Table 13. The estimated Geocoat metal recoveries are higher because the feed is the rougher float concentrate rather than the cleaner concentrate fed to the Roaster and Platsol processes. Note that in this techno-economic evaluation for Platsol a low 63% Ag recovery was assumed. Later testwork gave higher results reported in the range of 70 to 75% Ag recovery, as shown in Table 8.


379

Table 11: Ore and Concentrate Tonnages, Grades and Recoveries for Process

Options Units Platsol Roasting Geocoat Ore

Tonnage t/d 15,000 15,000 15,000 t/a 5,475,000 5,475,000 5,475,000

Copper % Cu 0.55 0.55 0.55 t/a Cu 30,113 30,113 30,113

Gold g/t Au 0.8 0.8 0.8 oz/y Au 140,836 140,836 140,836

Silver g/t Ag 13.7 13.7 13.7 oz/y Ag 2,411,817 2,411,817 2,411,817

Concentrate Treated Tonnage t/d 270 270 3,000

t/a 98,550 98,550 1,095,000 Copper % Cu 25.60 25.60 2.50

t/a Cu 25,229 25,229 27,375 Gold g/t Au 26.9 26.9 3.1

oz/y Au 85,241 85,241 109,148 Silver g/t Ag 336.4 336.4 41.9

oz/y Ag 1,065,988 1,065,988 1,475,257

Table 12: Overall Metal Recoveries Used for Each Option

Process Cu Recovery %

Au Recovery %

Ag Recovery %

Platsol 98.0 97.0 63.0 Geocoat 92.5 90.0 60.0 Roasting calcine 99.5 99.5 99.5 Calcine payables 96.8 95.0 93.2

Table 13: Payable Metals for Each Option Process Cu Payables

t/a Au Payables

oz/a Ag Payables

oz/a Platsol 24,724 82,684 671,572 Geocoat 25,329 98,233 885,154 Roasting calcine 24,287 80,574 988,609

In view of the current uncertainty in estimating capital and operating costs even on projects with much more advanced engineering, only relative magnitudes of the project’s estimated


380

capital and operating costs for each option are presented. The costs for each option are summarized in Table 14 and 15 using a basis of 100 for the Platsol option.

Table 14: Relative Estimated Total Project Costs for Each Option Area Platsol Roasting Geocoat Mine 40 40 40 Mill Modifications 6 6 4 Process 30 16 39 Indirects 8 5 11 Total Direct & Indirect Costs 84 67 94 Contingency (20% TDIC) 16 13 19 Total Project Cost 100 80 113

The Roasting option has significantly lower capital costs than the hydrometallurgical options of Platsol and Geocoat because there is no copper metal production component as the calcine is sold to copper smelters. Estimated Geocoat capital costs are higher because the feed is the rougher float concentrate rather than the smaller volume of cleaner concentrate fed to the Roaster and Platsol processes.

Table 15: Estimated Relative Annual Operating Costs for Each Option

Area Platsol Roasting Geocoat Mine and milling 74 74 73 Process operating cost 26 10 29 Calcine treatment charges - 19 - Total Processing Costs 100 103 102

However, in the evaluation total processing costs for each option, including calcine treatment charges for the Roaster option, appear similar as shown in Table 15. Estimated total annual pre-tax revenues for each option are summarized in Table 16. These were based on estimated long-term prices of US$2.00/lb Cu; US$650/oz Au; US$12.0/oz Ag and US$100/t acid.

Table 16: Estimated Total Annual Revenues for Each Option

Area Platsol Roasting Geocoat Copper 64 63 66 Gold 31 30 37 Silver 5 7 6 Acid - 3 - Total Annual Revenues 100 103 109

The Geocoat option gave the highest annual revenues because of the higher overall copper and gold recoveries from treating rougher float concentrates. Pre-tax economic parameters for the options are compared in Table 17 with Platsol as the base case.


381

Table 17: Relative Pre-Tax Economic Comparison for Each Option Economic Parameter Platsol Roaster Geocoat Capital Costs 100 80 113 Processing Costs 100 103 102 Revenues 100 103 109 Annual Pre-tax Operating Margin 100 103 115 Pre-tax NPV @ 10% discount rate 100 125 117

Based on this preliminary evaluation, the Roasting option appears to provide the highest pre-tax NPV followed by the Geocoat option and then Platsol. Lower Roasting capital costs and a higher pre-tax margin combine to give the highest relative NPV of the options evaluated. PROJECT RISKS AND OTHER CRITERIA Project risk and key other criteria for each of the processes are compared in Table 18. The Roasting option is the only one that has been operated commercially to-date on copper concentrates and therefore the Geocoat and Platsol processes represent a higher level of project risk. Based on experience with other new metallurgical processes, it will take longer to develop efficient operating and maintenance practices and to ramp up production to design capacity for these processes. CONCLUSIONS Based on the preliminary techno-economic evaluation, the Roasting option appears to offer the lowest project risk as well as the highest estimated pre-tax NPV. MDO is currently investigating the Roasting option further with regard to roasting testwork on enargite concentrates and further process flowsheet development. Engineering study work is planned to estimate more accurately the capital and operating costs of the Roasting option. The current MDO crushing and grinding plant is suitable to mill LF sulphide ores to the required grind at current throughput rates and feed a new flotation plant to separate enargite from the pyrite tails stream. Further flotation optimization testwork is required to improve the enargite /pyrite separation and economic evaluation is required to investigate the production of sulphuric acid from the significant levels of pyrite present in LF sulphide ores. Coupled with flowsheet engineering for the Roasting option will be significant study and evaluation of the environmental permitting requirements at MDO. Areas for environmental study will be disposal of flotation tails streams, handling and disposal of a Fe-As scorodite type precipitate and calcine from the possible production of sulphuric acid from roasting of the pyrite.


382

Table 18: Process Options - Qualitative Risk Analysis, Advantages, and Disadvantages Criteria Platsol Geocoat Roasting

Process Risk

High - never operated commercially. Materials of construction are an issue

High - never operated commercially on copper concentrates

Low - proven commercial operation

Operability

Integrated process with several unit operations. More complex flowsheet than gold POX. Sensitive to feed mineralogy.

Heap leach pads are simple to operate but can give lower recoveries than testwork. Relatively insensitive to feed mineralogy.

Fluid-bed roasters (FBR) have high on-stream time. (95%). Need good gas scrubber design & operation to avoid blockages. Less insensitive to feed mineralogy than hydro options.

Maintainability

Autoclaves are usually high maintenance. Exotic materials of construction costly to maintain

Should be high based on other heap leach operations

Fluid-bed roasters have low maintenance requirements

Availability of Technology

Patented license terms to be negotiated

Patented license terms to be negotiated

Available from several providers

Recoveries/Payables

Cu High but usually lower than pilot tests

High but usually lower than pilot tests

High minus smelter deduction

Au High but usually lower than pilot tests

High but usually lower than pilot tests


Ag Moderate & variable depending on feed & POX conditions

Low & variable depending on feed & heap leach conditions


Arsenic Fixation

As mostly fixed as scorodite in POX residue. Stability in alkaline disposal conditions after cyanidation may be an issue (Sao Bento, Campbell Red Lake experience)

Amount and stability of As precipitated in heap not established.

As in scrubber solution fixed as scorodite in HTPOX residue, similar to Platsol but disposal at lower pH as not cyanided.

Environmental Permitting

Best Available Technology - should meet requirements

Less certain - testwork required

Best Available Technology - should meet requirements

Economics

Time to reach design capacity is uncertain & represents a major project risk. Preliminary capex higher than roaster with similar operating margin

Time to reach design capacity is uncertain & represents a major project risk. Preliminary capex is highest with similar operating margin to other options.

FBR’s generally ramp up to design capacity quickly. Proven process & potential for operator training at existing operations. Preliminary capex is lowest with similar operating margin to other options

Other

Autoclave materials of construction POX campaign life LME cathode certification

Geocoat process long leach time LME cathode certification

Arsenic level in calcine (<0.5%) Long-term smelter TC/RC’s


383

ACKNOWLEDGEMENT The authors wish to thank Kinross Gold Corporation for permission to publish this paper and release the contained information. REFERENCES Bath M., Norton A. and Gunn M., 2005. Commercialization of the GEOCOAT® Bio-oxidation Process, ALTA 2005. Rendezvous Observation City, Scarborough Beach, Perth WA, 16-18. Ferron C. J., Fleming C. A., O'Kane T. and Dreisinger D., 2002. High temperature chloride assisted leach process to extract simultaneously Cu, Ni, Au and the PGM’s from various feedstocks. Proceedings of 32nd Hydrometallurgy Meeting: Chloride Metallurgy 2002. Vol. 1, Montreal, Canada, 19-23. Marsden J. and House I., 2006. The Chemistry of Gold Extraction. Metallurgy and Exploration: Society for Mining, pp. xix, 651. SGS Report #1, July 2008. “Programa de Ensayos de Laboratorio de Flotación y Molienda. Proyecto 4099, Informe #1", SGS Minerals Services, Julio 2008”. SGS Report #2, November 2007. “Ensayos Piloto de Flotación con Mineral Sulfurado de La Coipa. Proyecto 4099, Informe # 2, SGS Minerals Services, Noviembre 2007”. Smith E. H., Foster J. W., Minet P. H. and Cauwe P. H., 1985. Selective Roasting to De-Arsenify Enargite/Pyrite Concentrate From St. Joe's El Indio Mine - From Pilot Plant to Commercial Operation, Complex Sulfides--Processing of Ores, Concentrates and By-Products; San Diego, California; USA, 10-13.

2009 Proceedings_Processing of Refractory Sulphides at Mantos de or[877175]

Documents

Transcript of 2009 Proceedings_Processing of Refractory Sulphides at Mantos de or[877175]