GALVANOX TM A NOVEL GALVANICALLY-ASSISTED ATMOSPHERIC LEACHING TECHNOLOGY FOR COPPER CONCENTRATES

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1 327 Canadian Metallurgical Quarterly, Vol 47, No 3 pp , 2008 Canadian Institute of Mining, Metallurgy and Petroleum Published by Canadian Institute of Mining, Metallurgy and Petroleum Printed in Canada. All rights reserved GALVANOX TM A NOVEL GALVANICALLY-ASSISTED ATMOSPHERIC LEACHING TECHNOLOGY FOR COPPER CONCENTRATES D.G. DIXON 1, D.D. MAYNE 1 and K.G. BAXTER 2 1 University of British Columbia, Vancouver, Canada 2 Bateman Engineering Pty Ltd, Perth, Australia dixon@interchange.ubc.ca (Received in revised form April, 2008) Abstract A novel technology for leaching copper from primary copper concentrates has been developed at the University of British Columbia (UBC). This patented technology takes advantage of the galvanic couple between pyrite and chalcopyrite to ensure rapid and complete oxidation of chalcopyrite under mild conditions in acidic iron sulfate solution, without the need for microbes, ultrafine grinding or chemical additives such as chloride, nitrate or surfactants. Copper recoveries of 98% or greater can be achieved at 80 C under atmospheric conditions in as little as 4 hours of residence time. The process is selective for chalcopyrite over pyrite, generates near-quantitative levels of elemental sulphur and is fully compatible with conventional solvent extraction and electrowinning of LME Grade A copper cathodes. In this paper, the fundamentals of the Galvanox TM process are reviewed and selected results of batch leaching tests on various concentrate samples are presented and discussed. Résumé À l université de la Colombie Britannique (UBC), on a développé une nouvelle technique de lixiviation du cuivre à partir de concentrés primaires de cuivre. Cette technique brevetée prends avantage du couple galvanique entre la pyrite et la chalcopyrite pour assurer une oxydation rapide et complète de la chalcopyrite en conditions bénignes dans une solution acide de sulfate de fer, sans avoir besoin de microbes, de broyage ultrafin ou d additifs chimiques comme le chlorure, le nitrate ou des agents de surface. On peut obtenir des récupérations de cuivre de 98% ou plus à 80 C sous conditions atmosphériques en aussi peu que 4 heures de résidence. Le procédé est sélectif pour la chalcopyrite par rapport à la pyrite, génère des niveaux presque quantitatifs de soufre élémentaire et est entièrement compatible avec l extraction conventionnelle par solvant et l extraction par voie électrolytique des cathodes de cuivre LME de qualité A. Dans cet article, on examine les principes fondamentaux du procédé Galvanox TM et l on présente et discute des résultats choisis d essais de lixiviation en lot de différents échantillons de concentré. GALVANOX TM HISTORY UBC researchers David Dixon and Alain Tshilombo developed a novel process for the galvanically-assisted atmospheric leaching of primary copper concentrates. A preliminary U.S. patent application was filed in June Since that time, several other patent applications have been filed [1] and a successful Patent Cooperation Treaty (PCT) examination in September 2006 spawned many other national phase applications. In October 2006, UBC entered into an exclusive marketing agreement with Bateman Engineering BV and the two parties are working closely to identify potential applications of the process. Batch testwork programs on a number of candidate concentrates have been completed and continuous pilot campaigns on two of them were completed in early The results of these runs will be the subject of a later paper. GALVANOX TM FEATURES Galvanox TM is the galvanically-assisted atmospheric leaching of copper from chalcopyrite (CuFeS 2 ) concentrates in a ferric/ferrous sulphate medium, which offers several potential advantages over existing processes. Specifically, the process:

2 328 D.G. DIXON, D.D. MAYNE and K.G. BAXTER 1. runs under atmospheric conditions (~80 C), thus requiring only low-cost leaching tanks (as opposed to high-cost pressure leaching autoclaves) 2. is strictly chemical and therefore requires no microbes (i.e., this is not bioleaching) 3. runs in a pure sulphate medium requiring no special solution additives (e.g., chloride, ammonia) or specialized equipment (i.e., the solutions are not particularly corrosive) 4. does not require fine grinding (i.e., a P 80 of 53 to 75 µm will typically suffice) 5. generates primarily elemental sulphur (>95%) and therefore, levels of oxygen consumption are near the theoretical minimum for chalcopyrite oxidation 6. operates below the melting point of sulphur and therefore requires no surfactants (e.g., lignin sulphonate, quebracho, OPD, fine coal) 7. is selective for chalcopyrite over pyrite and is therefore applicable to low grade or bulk concentrates, offering a distinct mill recovery advantage over smelting 8. achieves virtually complete copper leach extraction, typically in less than 24 hours 9. is fully compatible with conventional solvent extraction-electrowinning (SX-EW) technology for the production of LME Grade A pure copper cathode directly at the mine site. GALVANOX TM CHEMISTRY The key to the Galvanox process is the presence of ground pyrite (FeS 2 ) in the leaching reactor at a level of two to four times the mass of chalcopyrite. Chalcopyrite is a semiconductor and therefore corrodes electrochemically in oxidizing solutions, as shown schematically in Figure 1. In ferric sulphate media, the overall leaching reaction is as follows: CuFeS 2 (s) + 2 Fe 2 (SO 4 ) 3 (a) CuSO 4 (a) + 5 FeSO 4 (a) + 2 S (s) This reaction may be represented as a combination of anodic and cathodic half-cell reactions: (1) Anodic half-cell reaction: CuFeS 2 Cu 2+ + Fe S + 4 e (2) Cathodic half-cell reaction: 4 Fe e 4 Fe 2+ (3) The fundamental problem with chalcopyrite oxidation is that the mineral surface becomes passivated (i.e., resistant to electrochemical breakdown) at solution potentials above a certain level, typically between 440 and 520 mv (for a review of all available data, refer to [2]). It is widely held that this phenomenon results from the formation of a passivating film on the mineral surface, which most likely consists of an altered, partially Fe-depleted form of chalcopyrite [3]. As a result of this, most investigators have assumed that it is the anodic half-cell reaction (i.e., the mineral breakdown reaction) which limits the overall rate of leaching. We have determined, however, that it is primarily the cathodic half-cell reaction (i.e., ferric reduction) which is slow on the chalcopyrite surface [4]. Hence, the presence of an alternative, catalytic surface for ferric reduction in electrical contact with the chalcopyrite, as shown schematically in Figure 2, will alleviate the passive behaviour of chalcopyrite in a ferric sulphate solution. Pyrite is an effective and convenient provider of this alternative surface for ferric reduction: effective because pyrite mass additions roughly two to four times that of chalcopyrite ensure rapid and complete copper extraction and convenient because such pyrite levels are typically present in primary copper ores. Hence, the requisite pyrite level in the reactor may be achieved in many cases simply by floating a bulk pyrite/chalcopyrite concentrate and sending this directly to the leaching circuit. This has the added advantage of minimizing copper losses in the flotation circuit. If the pyrite levels in the ore are inadequate, then a leach residue recycle stream may be implemented, as shown in the simplified process flowsheet of Figure 3. In the leach reactors, chalcopyrite is leached selectively at relatively low potential in the presence of the pyrite catalyst, producing a solid sulphur residue according to the following overall reaction: Fig. 1. Schematic diagram of typical electrochemical leaching. Fig. 2. Schematic diagram of galvanically-assisted electrochemical leaching.

3 GALVANOX TM A NOVEL GALVANICALLY-ASSISTED ATMOSPHERIC LEACHING TECHNOLOGY FOR COPPER Hence, the only significant reagent costs are electricity for electrowinning and the oxygen plant and a small amount of limestone to neutralize any excess acid. The overall Galvanox leaching process is very flexible in that the proportion of concentrate treated through the autoclave can be varied if there is a requirement for acid external to the process flowsheet for a secondary sulphide or oxide leaching process, such as a heap leach. The required quantity of acid can be set to meet the added demand. EXPERIMENTAL Fig. 3. Simplified block flowsheet of the Galvanox process. CuFeS 2 (s) + O 2 (g) + 2 H 2 SO 4 (a) CuSO 4 (a) + FeSO 4 (a) + 2 S (s) + 2 H 2 O (4) Pyrite is galvanically protected by the leaching mechanism; hence, pyrite oxidation is minimal in this process. This behaviour would appear to be unique amongst the chalcopyrite leaching technologies and will facilitate leaching of low grade concentrates economically. Following solid-liquid separation, the pregnant leach solution (PLS) is subjected to conventional solvent extraction and electrowinning (SX-EW) to produce pure copper cathodes: CuSO 4 (a) + H 2 O (l) Cu (s) + H 2 SO 4 (a) + ½ O 2 (g) A portion of the PLS from the atmospheric leach is sent through a small oxyhydrolysis autoclave at 220 C to oxidize and precipitate iron as hematite for safe disposal to tails and to regenerate acid while generating heat for the atmospheric leach step: FeSO 4 (a) + ¼ O 2 (g) + H 2 O (l) ½ Fe 2 O 3 (s) + H 2 SO 4 (a) In principle, the overall chemistry of the process is acid neutral: CuFeS 2 (s) + 5 / 4 O 2 (g) Cu (s) + ½ O 2 (g) + ½ Fe 2 O 3 (s) + 2 S (s) However, some make-up acid will always be required to account for losses and consumption by gangue minerals. This make-up acid is produced during iron oxyhydrolysis by feeding a small proportion of concentrate to the autoclave, which also ensures autothermal operation of the entire flowsheet, thus minimizing the need for steam addition: CuFeS 2 (s) + 17 / 4 O 2 (g) + H 2 O (l) ½ Fe 2 O 3 (s) + CuSO 4 (a) + H 2 SO 4 (a) (5) (6) (7) (8) Apparatus Batch leaching experiments are conducted in sealed 3 L jacketed glass reactors fitted with a single 6 bladed Rushton turbine impeller and three baffles made of 316 stainless steel (Applikon Dependable Instruments). The stirring shaft is attached to an exterior motor controlled to 1200 rpm. Each reactor is heated by a circulating hot water bath (Haake P14). Three probes are inserted into each reactor: a glass ph probe (AppliSens), a solution redox potential probe (Analytical Sensors) and a thermocouple (Omega). The probes are attached to a multi-channel digital controller (Applikon Dependable Instruments ADI 1030 Biocontroller) and their outputs recorded by a laptop computer. With the ADI 1030 Biocontroller, a solution potential set point is specified. When the measured solution potential falls below the set point, an analogue signal is sent to a digital gas mass flow meter (Aalborg) which allows oxygen to flow from a compressed gas cylinder into the reactor through a sparger. The oxygen flow is slowly decreased as the potential approaches the set point. Once the set point is reached, oxygen flow ceases. Excess oxygen exits the reactor through a condenser to limit the amount of water loss by evaporation. Procedure The reactor is filled with 1500 g of DI water and the desired mass of sulphuric acid and sealed. Next, the oxygen line, condenser and stirring motor are installed. The solution is mildly agitated while being heated to the operating temperature. Prior to use, the redox probe is immersed in Light s solution to ensure good working conditions before being inserted into the reactor. Similarly, the ph probe is calibrated using buffer solutions of ph 1.68 and ph 4.00 before being placed into the reactor and the thermocouple is placed into a temperature reference to ensure the ph measurements are corrected to the actual temperature of the standard buffer solutions. Once the solution has reached the desired temperature, iron salts (and copper salts if desired) are added, agitation is increased to 1200 rpm and the ore is introduced to the reactor. Finally, all open ports on the lid are plugged with rubber stoppers to seal the reactor. Samples are withdrawn from the reactor periodically using a 20 ml syringe. Approximately 8 ml of slurry is

4 330 D.G. DIXON, D.D. MAYNE and K.G. BAXTER Table I XRD Rietveld analysis of Cu concentrate 1 and Huanzala pyrite Mineral Formula Cu concentrate Mass % Pyrite Quartz SiO Muscovite KAl 2 AlSi 3 O 10 (OH) Plagioclase NaAlSi 3 O 8 CaAl 2 Si 2 O Gypsum CaSO 4 2H 2 O 1.1 Talc Mg 3 Si 4 O 10 (OH) Fluorite CaF Sphalerite (Zn,Fe)S 1.0 Marcasite FeS Pyrite FeS Chalcopyrite CuFeS Anilite Cu 7 S transferred to a centrifuge tube. Any slurry remaining in the syringe is returned to the reactor. The samples are filtered and assayed for copper content by atomic absorption (AA) or induction coupled plasma (ICP) spectroscopy. Once the test is complete, the contents of the reactor are emptied into a large flask. The walls of the reactor are washed with DI water to ensure that all solids are removed. The entire reactor contents (including wash water) are filtered through 5-µm filter paper and washed. After drying in warm air for two days, the filter paper and solids are weighed and then the solids are coned and quartered to obtain a representative sample for chemical analysis. Solid samples are submitted to a local laboratory for copper assay by multi-acid digestion and wet titration and/or ICP. RESULTS AND DISCUSSION Copper Concentrate % Cu A primary copper sulphide concentrate with 35.7% Cu was obtained from a mine in Zambia. Unfortunately at the time of the study, there was no local pyrite source that could be supplied and therefore massive pyrite samples from Huanzala Mine in Peru were used. Mineralogy and elemental composition of the copper concentrate and pyrite sample are summarized in Tables I and II, respectively. The copper concentrate was leached in an as-received condition. The massive pyrite samples were crushed in a series of cone crushers and then pulverized in 150 g lots for five minutes. Five variables were studied in the leaching of concentrate 1: initial acid content, solution potential, pyrite-to-chalcopyrite (Py:Cp) ratio, temperature and the effect of pyrite recycle. Only one variable was changed at a time in each experiment. Table 3 contains a test-matrix and summarizes the conditions for all the experiments. Pyrite Addition Experiments: In order to determine the effect of pyrite addition, experiments were conducted at various Py:Cp ratios. The results of these experiments are shown in Figure 4. Clearly, pyrite addition has a large effect on copper extraction. The test without added pyrite achieved only 57% extraction in 47 hours. The addition of 25 g of pyrite (Py:Cp ratio of 0.8) dramatically increased the extraction of copper; however, a significant induction period was observed. This induction period was also observed in the test in which 50 g of pyrite were added, although the length of the induction period was substantially diminished. Additions of 100 g and 150 g of pyrite (Py:Cp ratios of 2.7 and 3.8, respectively) displayed similar behaviour. There was no increase in the rate of copper extraction by increasing the pyrite content above 100 g. Both tests achieved >97% copper extraction in about 12 hours with no Table II Elemental analysis of concentrate 1 and Huanzala pyrite Element Mass % Cu concentrate Pyrite Cu Fe S Al 0.2 Ca Pb 0.2 Mg Ni 0.1 Zn 1.5 P 0.3

5 GALVANOX TM A NOVEL GALVANICALLY-ASSISTED ATMOSPHERIC LEACHING TECHNOLOGY FOR COPPER Fig. 4. Effect of pyrite addition on copper extraction. Fig. 6. Solution potential behaviour of potential experiments. discernable induction period. In an effort to conserve the Huanzala pyrite sample, all further tests were conducted using 100 g of pyrite. A second set of experiments was conducted in which the acid content was increased from 68 g to 90 g. The extraction curves (not shown) are similar showing little dependence on acid content. Potential Experiments: Tests were conducted at various potential set points from 425 to 485 mv (Ag/AgCl). The results are shown in Figure 5. Increasing the solution potential from 425 to 470 mv results in increasingly improved copper extractions. The two highest potential tests (470 and 485 mv) gave very similar results. The apparent decrease in copper extraction at 485 mv is an analytical artifact of pyrite dissolution. Solid assays confirm the higher extraction value (see Table III). The potential behaviour during a number of leach tests is shown in Figure 6. All of the tests began at a potential close to 495 mv. Within minutes of the concentrate being added to the solution, the potential dropped sharply, primarily due to the rapid consumption of ferric from the dissolution of secondary copper sulphides. The rate of ferrous oxidation could not keep up with the consumption of ferric during this period. Up to 455 mv the potential set point was attained within 30 minutes. The 470 mv set point was not attained until 4.1 hours into the test, while the 485 mv set point was not attained until 6.6 hours into the test. Interestingly, the potential behaviour of these two tests was identical for the first four hours, even though the oxygen flow rate was greater in the higher potential experiment. As a result, the extraction curves for those two tests are virtually identical. This indicates that the rate of leaching was limited by the rate of oxygen mass transfer during this period. Acid Addition Experiments: The results of the acid addition experiments are shown in Figure 7. Assuming that the copper Fig. 5. Effect of solution potential on copper extraction. Fig. 7. Effect of acid addition on copper extraction.

6 332 D.G. DIXON, D.D. MAYNE and K.G. BAXTER Table III Experimental conditions and selected results for tests on concentrate 1 Cu Pyrite Initial Deionized Total Potential Calculated Final copper Pyrite Test concentrate mass acid water iron Fe 3+ /Fe 2+ set point Py/Cp Temperature extraction oxidation ID mass (g) content mass content ratio (mv vs ratio C (solids) (XRD) (g) (g) (g) (g) Ag/AgCl) % % K K K K K K % K K K K % K K K K K K16 A K % K K K K K K23 50 K8 B K24 D 50 K5 C K A K16 (in boldface type) is the baseline test B K23 used the residue from K8 as the pyrite source (after sampling) C K24 used the residue from K5 as the pyrite source (after sampling) D 7.5 g Cu was added to the initial solution for test K24 minerals leach completely to form elemental sulphur and that all other minerals are inert, the stoichiometric requirement of sulphuric acid is g for a 50 g charge of concentrate or kg acid/kg concentrate. The lowest acid content tested (45 g or 0.90 kg/kg) is near the stoichiometric requirement. Initial copper extraction is similar to the other acid tests, but the copper extraction begins to slow down after two hours, resulting in a final copper extraction of 87% after 24 hours. This suggests that 0.90 kg/kg is below the true stoichiometric requirement. Increasing the added acid to 55 g (1.10 kg/kg) improved copper extraction significantly. Further increases in acid addition increase the rate of copper extraction slightly. The most rapid copper extraction was obtained in the experiment with the highest acid addition. For that reason all subsequent tests were conducted with 90 g of acid (1.80 kg/kg). However, in practice the amount of acid added would be driven by the requirement for low acid levels (<10 g/l) in the PLS for efficient solvent extraction. Temperature Experiments: Figure 8 shows the results of otherwise identical tests conducted at 60, 70 and 80 C. A significant induction period was observed in the 60 C experiment. The length of the induction period decreased with increasing temperature. However, the final copper extraction was nearly the same for all three temperatures. These experiments show that higher temperatures (at least 80 C) are desirable for rapid copper extraction. Pyrite Recycle Experiments: Test solutions identical to those for K16 (baseline conditions) were prepared to examine the viability of recycling pyrite for subsequent charges of concentrate. 100 g of filtered and dried residue from tests K8 and K5 were used as the recycled pyrite source. Test K23 used the residue from K8 and test K24 used the residue from K5 with the addition of 5 g/l of copper to the initial solution. 50 g of concentrate were added in both tests. The results of these experiments as well as the extraction curve from the baseline experiment are shown in Figure 9.

7 GALVANOX TM A NOVEL GALVANICALLY-ASSISTED ATMOSPHERIC LEACHING TECHNOLOGY FOR COPPER supernatant leach liquor were removed from the reactor using a peristaltic pump and replaced with 500 g of de-ionized water and 70 g of make-up acid. A second 50 g charge of concentrate was added once the solution reached 80 C and the test was run for another 24 hours. The results are shown in Figure 10. Notwithstanding a small delay in copper extraction from the pyrite recycle test, most likely due to a small amount of pyrite consumption during the second half of the initial test, all three experiments behaved very similarly, indicating that recycled wet pyrite is just as effective as fresh pyrite. Fig. 8. Effect of temperature on copper extraction. Copper Concentrate % Cu A copper concentrate with 23.6% Cu comprising virtually all chalcopyrite was obtained from a mine in Australia. Typical leaching results for this concentrate are shown in Figure 11. The pyrite was sourced from a mine near Park City, Utah. Clearly, the pyrite recycle test with added copper sulphate (K24) out-performed the test with no added copper sulphate (K23). The results of test K24 were very similar to the baseline test, demonstrating the effectiveness of pyrite recycling, while the results of test K23 show a significant induction period in the absence of initial dissolved copper. The marked difference between these two tests can possibly be attributed to an alteration of the pyrite surface during drying or storage. It appears that this altered surface (an oxide coating, perhaps) diminishes the effectiveness of pyrite as a galvanic catalyst for chalcopyrite and that dissolved copper helps to reactivate the pyrite. However, it is worth noting that both pyrite recycle tests finished with complete copper extraction within 24 hours of retention time. A final experiment was devised to determine the effectiveness of recycling wet pyrite directly (as would naturally occur during actual plant operation). A baseline test was conducted for 24 hours at which point the agitation was stopped and the solids were allowed to settle. 500 ml of the Fig. 10. Effect of recycling wet pyrite on copper extraction. Fig. 9. Effect of recycling dried residue on copper extraction. Fig. 11. Effect of pyrite addition on copper extraction from concentrate 2 (30 g concentrate, 120 g pyrite, 30 g acid, 480 mv, 80 C).

8 334 D.G. DIXON, D.D. MAYNE and K.G. BAXTER Copper Concentrate % Cu A copper concentrate with 24.1% Cu comprising virtually all chalcopyrite was obtained from a mine in Ontario. Typical leaching results for this concentrate are shown in Figure 12. Park City pyrite was also used for this test. Copper Concentrate % Cu A copper concentrate with 20.1% Cu was obtained from a mine in Mongolia. Typical leaching results for this concentrate are shown in Figure 13. Huanzala pyrite was used for this test. Bulk Copper Concentrate 10.2% Cu A bulk copper concentrate with 10.2% Cu was obtained from a mine in British Columbia. This concentrate had a Py:Cp ratio of roughly 1.2 and was leached as-received, with no additional pyrite added. Leaching results are shown in Figure 14. While complete copper extraction was achieved within 24 hours, faster leaching could almost certainly have been achieved in this case by recycling pyrite to a higher Py:Cp ratio. Gold Recovery from Galvanox Residues Three leached residue samples of the bulk copper concentrate discussed above were combined and submitted to a local laboratory for gold leaching by bottle roll cyanidation. As shown in Figure 15, high gold recoveries were achieved with very modest levels of cyanide consumption. Apparently, the elemental sulphur which forms as a result of Galvanox leaching does not inhibit the subsequent recovery of gold from the residue, nor does it consume significant quantities of cyanide. As shown in Figure 16, the elemental sulphur layers which form on chalcopyrite particles during Galvanox leaching are very porous and tend to retain the outward shape of the original chalcopyrite particles. In addition to making the residues ideal for conventional gold cyanidation, this also gives the elemental sulphur particles a very low bulk density which renders them very easy to separate from the much denser pyrite particles in the residue, thus greatly facilitating pyrite recycle. SUMMARY Fig. 12. Effect of pyrite addition on copper extraction from concentrate 3 (10 g concentrate, 40 g pyrite, 15 g acid, 470 mv, 80 C). 1. Galvanox is robust and largely insensitive to the source of chalcopyrite or pyrite. 2. Process optimization is straightforward: a) Pyrite-to-chalcopyrite (Py:Cp) ratio: roughly 2:1 to 4:1 b) Acid concentration: stoichiometric + modest excess c) Solution potential: > 440 mv (Ag/AgCl), preferably 470 mv or higher d) Temperature: > 70 C, preferably 80 C Fig. 13. Effect of pyrite addition on copper extraction from concentrate 4 (57 g concentrate, 112 g pyrite, 60 g acid, 450 mv, 80 C). Fig. 14. Copper extraction from bulk concentrate (150 g bulk concentrate, Py/Cp ratio ~1.2, 75 g acid, 440 mv, 80 C).

9 GALVANOX TM A NOVEL GALVANICALLY-ASSISTED ATMOSPHERIC LEACHING TECHNOLOGY FOR COPPER a b Fig. 15. Cyanide leaching of gold from Galvanox residue containing 22.8 g Au per tonne (77 g residue, 0.5 g/l NaCN, ph 11, 20 C) a) gold extraction, b) sodium cyanide consumption. a b Fig. 16. SEM micrograph of particle cross-sections prepared from a typical Galvanox leach residue a) partially leached, b) completely leached 3. Recycled pyrite is equally as effective as fresh pyrite and is a convenient way to provide higher Py:Cp ratios than are present in the concentrate. 4. Under the correct process conditions, Galvanox leaching is very rapid and is typically limited by the rate of gas-liquid mixing when all other conditions for successful leaching have been satisfied. 5. Conventional cyanide leaching of Galvanox leach residues gives high gold recoveries at reasonably low cyanide consumption levels.

10 336 D.G. DIXON, D.D. MAYNE and K.G. BAXTER A detailed scoping study conducted by Bateman Engineering in August 2006 suggested that a stand-alone Galvanox plant could be built for less than US $3,000 per annual tonne Cu and operated for less than US $0.13 per lb Cu. At a copper price of $1.20 per lb (far below current record high prices), this would generate an internal rate of return (IRR) exceeding 20%, rendering Galvanox the most economically attractive hydrometallurgical process option for primary copper concentrates by a significant margin. Projected IRRs from competing processes ranged from 7 to 14% for plants of the same capacity in the same location. This implies that the capital investment in a Galvanox plant could be paid off substantially faster than competing processes. In subsequent papers, the results of recent pilot campaigns on two concentrates will be presented, along with a discussion of several Galvanox flowsheet options and a detailed technical and economic evaluation of the Galvanox process. ACKNOWLEDGMENTS The authors gratefully acknowledge the Natural Sciences and Engineering Research Council of Canada (NSERC) for providing funding assistance for Galvanox process development through the Research Partnerships Program (Idea-to-Innovation Phase I grant number I2IPJ and Phase II grant number I2IPJ ). The financial support of Bateman Engineering BV and First Quantum Minerals Ltd. is also gratefully acknowledged. REFERENCES 1. D.G. Dixon and F.A. Tshilombo, Novel Leaching Process for Copper Concentrates World Patent Application WO 2005/118894, US Patent Application US 2005/ , Chile Patent Application , Peru Patent Application /OIN, PCT Patent Application PCT/CA2005/ G. Viramontes-Gamboa, B.F. Rivera-Vasquez and D.G. Dixon, Experimental Predictions of the Potential Range to Leach Chalcopyrite in Acidic Ferric Sulfate Media, 2007, The John E. Dutrizac Symposium on Copper Hydrometallurgy, Proceedings of the 6th International Copper-Cobre Conference, vol. IV(1) (P.A. Riveros, D.G. Dixon, D.B. Dreisinger and M.J. Collins, eds.), CIM, Montreal, pp R.P. Hackl, D.B. Dreisinger, E. Peters and J.A. King, Passivation of Chalcopyrite during Oxidative Leaching in Sulfate Media, Hydrometallurgy, 1995, vol. 39, pp A.F. Tshilombo, Mechanism and Kinetics of Chalcopyrite Passivation and Depassivation during Ferric and Microbial Leaching Solutions, Ph.D. Thesis, 2004, University of British Columbia.