Hydrometallurgy 106 (2011) 84–92

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Conventional and electrochemical bioleaching of chalcopyrite concentrates bymoderately thermophilic bacteria at high pulp density

Ali Ahmadi a,b, Mahin Schaffie c, Jochen Petersen d, Axel Schippers e, Mohammad Ranjbar a,b,⁎a Department of Mining Engineering, Shahid Bahonar University of Kerman, Iranb Mineral Industries Research Centre (MIRC), Shahid Bahonar University of Kerman, Iranc Department of Chemical Engineering, Shahid Bahonar University of Kerman, Irand Centre for Bioprocess Engineering Research, University of Cape Town, South Africae Federal Institute for Geosciences and Natural Resources (BGR), Hannover, Germany

⁎ Corresponding author. Department of Mining EUniversity of Kerman, Iran. Tel./fax: +98 341 2113663.

E-mail address: m.ranjbar@web.de (M. Ranjbar).

0304-386X/$ – see front matter © 2010 Elsevier B.V. Adoi:10.1016/j.hydromet.2010.12.007

a b s t r a c t

a r t i c l e i n f o

Article history:Received 22 September 2010Received in revised form 8 December 2010Accepted 8 December 2010Available online 16 December 2010

Keywords:ElectrochemistryBioleachingElectro-bioreactorChalcopyriteModerately thermophilic bacteria

Conventional and electrochemical bioleaching were investigated to extract copper from Sarcheshmehchalcopyrite concentrate at high pulp densities. Experiments were conducted in the presence and absence of amixed culture of moderately thermophilic iron- and sulphur oxidizing bacteria using a 2-L stirred electro-bioreactor at 20% (w/v) pulp density, an initial pH of 1.4–1.6, a temperature of 50 °C, a stirring rate of 600 rpmand Norris nutrient medium with 0.02% (w/w) yeast extract addition. The results of 10 day leaches showedthat, when using electrochemical bioleaching in an ORP range of 400 to 430 mV, copper recovery reachesabout 80% which is 3.9, 1.5 and 1.17 times higher than that achieved in abiotic electrochemical leaching,conventional bioleaching, and electrochemical bioleaching at 440–480 mV ORP, respectively. It appears thatapplying current directly to the slurry optimises both, the biological and chemical subsystems, leading to anincrease in both, the dissolution rate and the final recovery of copper from the concentrate. Mineralogicalanalysis of the solid residues of electrochemical leaching in both, biotic and abiotic media, showed theformation of chalcocite and covellite minerals on the surface of not leached chalcopyrite. It is postulated thatthe reduction of refractory chalcopyrite to more soluble minerals such as chalcocite and covellite is achievedthrough both, electron transfer upon electrode contact and by ferrous reduction at the low ORP of the slurry.These secondary minerals are then rapidly dissolved through bioleaching, while at the same time a formationof a passive layer of jarosites is minimised. This process also appears to promote an increased bacteria–solidratio due to favourable growth conditions.

ngineering, Shahid Bahonar

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© 2010 Elsevier B.V. All rights reserved.

1. Introduction

Currently, pyrometallurgical processes are the predominant routeto treat chalcopyrite flotation concentrates, which from an environ-mental perspective have some major problems, especially theemission of SO2 (Dimitrijevic et al., 2009). Over the last 30 years,many researchers in universities and industry have focused theirefforts on finding ways to extract copper by bioleaching. However, forchalcopyrite due to the slow dissolution kinetics, caused primarily bypassivation of the mineral surface, bioleaching has not beenimplemented at the full commercial scale as yet.

A number of recent studies have reported that the oxidationreduction potential (ORP) is one of the main parameters governingthe chemical and biological leaching rate of chalcopyrite. In thisregard, several authors (Hiroyoshi et al., 1997, 2000, 2001; Pinches

et al., 2001; Third et al., 2002; Sandström et al., 2005; Cordoba et al.,2008) have reported that during the chemical leaching of chalcopyritein ferric sulfate media, both the rate and yield of copper dissolution isat a maximum in a narrow range of ORP around 400–450 mV (vs. Pt,Ag/AgCl), whereas at ORPs above this range, the surface passivation ofthe mineral could occur. Moreover, it has been found that applyingdirect current into the bacterial slurry significantly enhances both theactivity and growth of microorganisms (Natarajan, 1992; Nakasonoet al., 1997; Ahmadi et al., 2010a).

It should be mentioned that in conventional tank bioleaching ofchalcopyrite, extremely thermophilic microorganisms (temperaturesas high as 70–80 °C) are required to extract copper rapidly whilstmaintaining economic viability. The leading process is BioCOP™which operates at 78 °C and 12% (w/w) pulp density (Batty andRorke, 2006). At these high temperatures, difficulties such as lowsolubility of oxygen, high rate of evaporation, high corrosion of reactorconstruction materials, and high sensitivity of the thermophilic cellsto metabolic stress caused by excess turbulence (in the contact ofincreased pulp density) occur. These are less of a problem at 50 °C orbelow (Rawlings et al., 2003; Olson et al., 2003; Okibe et al., 2003).

Cu

mu

lati

ve p

assi

ng

(%

)

Particle size (micrometer)

100

80

60

40

20

01 10 100

Fig. 1. Particle size distribution of copper concentrate.

85A. Ahmadi et al. / Hydrometallurgy 106 (2011) 84–92

A previous study found some promising results to extract copperfrom chalcopyrite concentrates by electrochemical bioleaching at lowertemperatures (≤50 °C)(Ahmadi et al., 2010a). In that research,electrochemical and conventional bioleaching experiments werecarried out on a chalcopyrite concentrate using a mixed culture ofmesophilic bacteria at 35 °C and a mixed culture of moderatelythermophilic bacteria at 50 °C, operating at 10% (w/v) pulp density.The results showed that the control of solution ORP around 425 mV (vs.Pt, Ag/AgCl), by applying current directly to the slurry, significantlyincreases both, the cell concentrations and copper recovery in bothcultures, especially in the presence ofmoderately thermophilic bacteria.

One of the main limitations of bioleaching processes in stirredtanks, especially in the case of extreme thermophiles, is the negativeeffect of high pulp densities on both extraction rate and final recoveryof metals. Oxygen and carbon dioxide availability, low bacteria–solidsratio, metabolic stress by high shear stress and abrasive conditions,inhibition of bacterial attachment, and the build-up of toxic leachproducts or other detrimental substances (such as some flotationreagents) have been reported as the most significant problems for asuccessful operation of bioleaching at high solid contents (Bailey andHansford, 1993; Acevedo and Gentina, 2007). To overcome theseproblems and meet the requirements of industry, microorganismsmust be adapted to high pulp densities (Mishra et al., 2005). However,to date, there is no information in the literature on the electrochem-ical control of the bioleaching process at high pulp densities of stirredreactors (at 10% ; Ahmadi et al., 2010a). For this reason, the presentresearch work investigates the process of electrochemical bioleachingat high pulp density and compares its efficiency to that ofconventional and electrochemical leaching processes in the presenceand absence of moderately thermophilic bacteria. The work wasexecuted by leaching of a Sarcheshmeh chalcopyrite flotationconcentrate in a stirred electrobioreactor.

2. Experimental

2.1. Materials

A mixed culture of moderately thermophilic bacteria supplied bySarcheshmeh Copper Mine, Kerman, Iran, was used. The microorgan-ismswere grown at 50 °C on Norrismedium (0.4 g/L (NH4)2SO4, 0.4 g/LK2HPO4, 0.5 g/L MgSO4.7H2O) with the copper concentrate at pulpdensities from 2% to 20% (w/v) replacing the energy source. Theflotation concentrate was obtained from Sarcheshmeh Copper Mine,and contained 44.02% chalcopyrite (CuFeS2), 23.99% pyrite (FeS2), 6.87%covellite (CuS), 5.84% chalcocite (CuS2), 13.61% non-metallic mineralsand 4.79% copper oxide minerals as shown by mineralogical analysis.The chemical analysis of the representative sample is presented inTable 1.Mineralogical investigations on the feed and solid residueswereperformed by optical microscopy using a Leica phase contrastmicroscope (DMLP). Since during the quantitative determination ofphases, the iron present in the iron hydroxide precipitates is ascribed tochalcopyrite and pyrite minerals, the quantitative determination ofthese residue analyse are not scientifically reliable; hence in this studythe results are reported only qualitatively. The transformation isvisualised in order to underline our understanding of the mechanismof dissolution.

The particle size distribution of the concentrate was determined bywet sieving and cyclosizer and showed that 80% is passing 76 μm(Fig. 1).

Table 1Chemical analysis of the copper sulphide concentrate.

Elements Cu Fe S Si Al Zn Mg K

(wt.%) 27.50 23.03 14.82 3.87 1.45 0.99 0.40 0.24

2.2. Electrobioreactor

The leaching experiments were performed in a three-electrode, 2 Lglass electro-bioreactor with 4 baffles, thermostated at the desiredtemperature by circulating water from a constant temperature baththrough the double-wall jacket (Fig. 2). The reactor had a mediumvolume of 1.3 L with a medium height/diameter ratio equal to 1.1. Theleach slurry was mechanically stirred by a pitched blade impeller(diameter=5 cm) mounted on a rotating shaft. A titanium–platinummesh (15 cm×9 cm×0.1 cm), acting as the cathodicworking electrode,was immersed into the reactor solution. A platinum foil was used as acounter electrode and was put into a separate small anodic compart-ment, separated from the cathode chamber by a glass frit. An Ag/AgClreference electrode was in contact with the electrolyte in the mainchamber through a Luggin capillary, which ended just short of theworking electrode. Air was supplied through a stainless-steel ringsparger underneath the impeller.

2.3. Leaching experiments

Batch experiments were carried out in the electro-bioreactor at 20%(w/v) pulp density, an initial pH of around 1.5, a temperature of 50 °C, astirring rate of 600 rpm, an aeration rate of 1 vvm (volume of air/volumeof slurry/min) and Norris nutrient medium with 0.02% (w/w) yeastextract addition. The intense agitation is needed both for maintaining ahomogeneous suspension and increasing the rate of mass transfer(especially of oxygen and carbon dioxide from the gas phase). Duringthe leaching experiments, the pH of the suspensions was monitoredperiodically and adjusted to around 1.5 by addition of H2SO4 (6 M). Thevariations of ORP were recorded daily throughout the leaching period.ThepHandORPvaluesweremeasuredwith a Jenway3540pHmeter anda Pt electrode in reference to an Ag/AgCl electrode (+207 mV vs. SHE at25 °C), respectively. Samples were periodically taken from the slurry andfiltered through Whatman No.41 filter paper. After that the filtrate wasused for copper and iron analysis by the atomic adsorption method. Theremaining solidswere returned to the reactor. The evaporated liquidwasperiodically replaced by adding acidified distilled water (pH=1.5).

The biotic experiments were inoculated with 20% (v/v) of a culturepreviously adapted to 20% pulp density. To maintain the ORP in thedesired range (400–430 or 440–480 mV) during electrochemicalbioleaching, the potential of the working electrode was controlled withrespect to the reference electrode with a Solartron Sl 1287 potentiostat/galvanostat. To keep the ORP in the desired range, the applied potentialwas manually set slightly lower than the set point value during theexperiments, however it was always higher than 250 mV.

The initial solutions of the conventional bioleaching, electrochem-ical bioleaching at 400–430 mV and electrochemical bioleaching at440–480 mV experiments contained 2.46, 3.15 and 2.35 g/L iron,respectively, which originated from their inoculum solutions.

Fig. 2. Schematic illustration of thermostated electrobioreactor.

Fig. 3. Copper recovery as a function of leaching time at 50 °C and 20% pulp density fordifferent experimental conditions: chemical leaching (CL), electrochemical leaching(ECL), bioleaching (BL) and electrochemical bioleaching (EBL).

86 A. Ahmadi et al. / Hydrometallurgy 106 (2011) 84–92

To evaluate the contribution of acid solution to copper recovery, anabiotic test of chemical leaching was carried out under the sameconditions as the bioleaching test, except the initial composition of theirsolutions (in the abiotic tests therewas no iron in the initial solution). Inaddition, an abiotic electro-leaching test was conducted by applying afixed 100 mA current (no ORP control), to investigate the effect ofapplied DC current to the slurry on the recovery of copper and iron,while other conditions were kept similar to the abiotic chemicalleaching test. It should bementioned that this electro-leaching test wasdifferent fromthatperformed in aprevious study (Ahmadi et al., 2010a),in which the working electrode was set as anode causing differentelectrochemical reactions expected here are different from those thatoccurred in the previous test. In the abiotic tests of this new study, themedium was sterilized with 2% (v/v) bactericide (2% (w/w) thymol inethanol) added to prevent microbial growth. In order to sterilize thereactor before each test, itwasoperated for 2 h at 85 °C and800 rpm in asolution of 10% (v/v) bactericide and 5% (v/v) HCl. Microbial growthwas periodically verified by observation under a Nikon opticalmicroscope (ECLIPSE, TE 2000-U). Free cells in solution were countedby direct counting using a Thoma chamber of 0.1 mm depth and0.0025 mm2 areawith the opticalmicroscope (magnification=1500×).

3. Results and discussion

In order to examine the influence of ORP and presence of bacteria onthe dissolution of chalcopyrite concentrate, various processes, namelychemical leaching (control test), bioleaching, electro-leaching andelectrochemical bioleaching were carried out in an electro-reactor at20% (w/v) pulp density. During evaluating the results and comparingthemwith those obtained from the previous research at the lower pulpdensity (Ahmadi et al., 2010a), it should be considered that theexperimental conditions of this study are different from those in the

previous work. One of the main differences relates to their differentstirring rate. In that research, because of fear of bacterial adaptation, thestirring rate was set at 300 rpm in the first 4 days with the result thatduring this period, a portion of the concentrate had settled at thebottomof the reactor anddidn't takepart in the leaching reactions. This problemwas solved in the high pulp density by consecutive adaptation tests (4times) at stirring rate of 600 rpmwhichwas constant at the rate duringthe main experiments.

3.1. Chemical leaching

Figs. 3 and 4 show the results of copper recovery (Fig. 3) and irondissolution (Fig. 4) from the concentrate by the various methods over

Fig. 5. Variation of pH as a function of leaching time at 50 °C and 20% pulp density fordifferent experimental conditions: chemical leaching (CL), electrochemical leaching(ECL), bioleaching (BL) and electrochemical bioleaching (EBL).

Fig. 6. ORP variation as a function of leaching time at 50 °C and 20% pulp density fordifferent experimental conditions: chemical leaching (CL), electrochemical leaching(ECL), bioleaching (BL) and electrochemical bioleaching (EBL).

Fig. 4. Total iron recovery as a function of leaching time at 50 °C and 20% pulp densityfor different experimental conditions: chemical leaching (CL), electrochemical leaching(ECL), bioleaching (BL) and electrochemical bioleaching (EBL).

87A. Ahmadi et al. / Hydrometallurgy 106 (2011) 84–92

10 days. As can be seen, the rate of chalcopyrite concentratedissolution in chemical leaching process (control test) is significantlylower than those obtained in the other processes in which final valuesof copper recovery and iron dissolution did not exceed 13% and 5%,respectively. The main portion of this copper extraction is attributedto the acid leaching of copper oxides and partial leaching of chalcocite(Eqs. (1) and (4)) and covellite (Eqs. (2) and (5)) minerals. It ispresumed that chemical dissolution of chalcopyrite (Eqs. (3) and (6))is negligible due to its high lattice energy (Habashi, 1978).

Cu2S þ 2Hþ→Cu2þ þ Cuo þ H2S ð1Þ

CuS þ 2Hþ→Cu2þ þ H2S ð2Þ

CuFeS2 þ 4Hþ→Cu2þ þ Fe2þ þ 2H2S ð3Þ

Cu2S + 1=2O2 + 2Hþ→Cu2+ + CuS + H2O ð4Þ

CuS + 1=2O2 + 2Hþ→Cu2+ + SB + H2O ð5Þ

CuFeS2 þ O2 þ 4Hþ→Cu2þ þ Fe2þ þ 2S∘ þ 2H2O ð6Þ

The smell of rotten eggs during the process could be related to theproduction of H2S through reactions (1) and (2).

Fig. 5 shows the variation of pH during various processes. The pHvalue rises during the chemical leaching of the concentrate, which isdue to acid consumption by the reactions of copper oxide andsulphide minerals (Eqs. (1)–(6)) as well as gangue minerals. It wascontrolled at around pH 1.5 by the addition of H2SO4 (during the first6 days). Fig. 6 shows that the ORP values remained low in the range of300–330 mV.

The low extraction ratio of Fe:Cu is probably due to the dissolutionof iron-free minerals i.e. copper oxides, covellite and chalcocite andthe insolubility of refractory iron bearing minerals i.e. pyrite andchalcopyrite under the conditions studied. Comparison of themineralogical analysis of the solid residue (after 10 days) (Fig. 7b)with that of the feed concentrate (Fig. 7a) clearly reveals thatsecondary copper bearing minerals i.e. covellite and chalcocite werenot dissolved.

3.2. Conventional bioleaching

To investigate the efficiency of the mixed culture of moderatethermophiles at high pulp density, a conventional bioleachingexperiment was carried out in the stirred bioreactor. The optimalconditions (temperature, 50 °C; initial pH, 1.5; nutrient medium,Norris; and yeast extract, 0.02% (w/w)) obtained from a previous

study (Ahmadi et al., 2010b) were employed. During preliminarybioleaching experiments (data not shown), the bacteria were adaptedto a high solid content and high degree of slurry agitation byincreasing the pulp density from 2 to 20% (w/v) and the stirring ratefrom 300 to 600 rpm and then maintaining these conditions. Theadapted culture (the 4th generation of a culture maintained at 20%pulp density and 600 rpm) was used to inoculate the bioleachingexperiment discussed here (inoculation volume=20% (v/v)).

Looking at Fig. 3, the copper recovery profile can be divided intothree distinct phases. Initially, approximately 21% of copper is rapidlyleached within the first day; this is mainly associated with thedissolution of copper oxides by H2SO4 and the first stage leaching ofchalcocite (Eq. (7)). Thehigher initial copper extraction in this biotic testcompared to the chemical leaching test (21.4%vs. 8.1%) ismainly relatedto their different initial solution compositions. In the bioleaching test,the iron concentration in the initial solution is 2.46 g/L (mostly as ferriciron), originating from the inoculation solution,which acts as a leachingagent for copper sulphides (Eqs. (7) to (9)) and pyrite (Eq. (10)).

Cu2S þ 2Fe3þ→Cu2þ þ 2Fe2þ þ CuS ð7Þ

CuS þ 2Fe3þ→Cu2þ þ 2Fe2þ þ S∘ ð8Þ

CuFeS2 þ 4Fe3þ→5Fe2þ þ Cu2þ þ 2S∘ ð9Þ

FeS2 þ 8H2O þ 14Fe3þ→15Fe2þ þ 2SO2−4 þ 16Hþ ð10Þ

Fig. 7.Mineralogical images of feed (a); solid residues of (b) chemical leaching, (c) bioleaching, (d ) electrochemical leaching, (e and f) electrochemical bioleaching, (Cv=covellite;Cc=chalcocite; Ccp=chalcopyrite; Py=pyrite).

88 A. Ahmadi et al. / Hydrometallurgy 106 (2011) 84–92

The initial copper extraction rate in this research is significantlyhigher than that obtained in the previous study (Ahmadi et al.,2010a), which could be attributed to the higher stirring rate employedhere (which prevents the setting of the solids.

The first phase is followed by another linear increase to about 51%,between days 1 and 7. This increase is probably associated with theleaching of chalcopyrite and covellite (natural mineral and theproduct of chalcocite leaching in Eq. (7)) according to Eqs. (9) and(8), respectively. This phase is followed by ceasing the copperdissolution in the remaining days to the end of experiment.Comparing the various processes in Figs. 3 and 4, it can be foundthat the final values of copper recovery (51.6%) and iron dissolution

(15.9%) are, respectively, around 3.8 and 3.2 times higher than thoseof the chemical leaching process. Mineralogical examinations showedthat secondary copper bearing minerals such as chalcocite andcovellite were not found in the bioleaching residue (Fig. 7c), althoughthey were present in the feed concentrate (Fig. 7a) and in the residueof the chemical leaching experiment as well (Fig. 7b).

In the bioleaching process, bacteria oxidize the insoluble metalsulphides, such as chalcopyrite, by indirect and/or contact mechanisms(Sand et al., 2001). Iron- and sulfur-oxidizing bacteria catalyticallygenerate the leaching agents of [Fe3+] and [H+] according to Eqs. (11)and (12), respectively, and then these agents, especially ferric iron,dissolve the sulphide minerals (Eqs. (1)–(10)). In this regard, sulphur

Fig. 8. SEM image and EDS analysis of the solid residue of conventional bioleaching.

89A. Ahmadi et al. / Hydrometallurgy 106 (2011) 84–92

oxidizing bacteria remove elemental sulphur, as a passivation layer,from the surface of sulphide minerals during the acid productionprocess (Eq. (12)).

4Fe2þ þ O2 þ 4Hþ þ Iron oxidizing acidophiles →4Fe3þ þ 2H2O

ð11Þ

SB + H2O + 3=2O2þ Sulphur oxidizing acidophiles →H2SO4 ð12Þ

Schippers and Sand (1999) proposed that the indirect mechanismoccurs via the thiosulphate route (primarily pyrite) or via thepolysulphides and sulphur route (most other sulphide mineralssuch as chalcopyrite). However, in the contact mechanism bacteriaattach to the mineral surface and prepare the medium and thenfacilitate the mineral attack through an electrochemical dissolutioninvolving ferric ions contained in the microbe's extracellular poly-meric substances (EPS) (Sand et al., 2001). Tributsch (2001)concluded that in practice suspended bacteria feed on chemicalspecies released by attached bacteria (cooperative mechanism).

The rapid increase of ORP to relatively high levels (Fig. 6) and thepH decrease (Fig. 5) indicate that both, activity and growth of thebacteria are very favourable under the conditions studied. It can beseen that ORP increases from 390 to 540 mV (after day 7) with the lagphase of bacterial growth being less than 1 day, whereas it wassignificantly longer in the bioleaching experiment of the previousstudy (Ahmadi et al., 2010a).

Furthermore, Fig. 5 shows the variation of pH during the bioleachingexperiment. During the first day of the experiment, pH increases andsulfuric acid needs to be added to keep the reactor pH around 1.5. Afterthat the pH decreased gradually to a final value of 1.3, which indicatesthat the amount of acid-production is more than that of acidconsumption during the transition phase. It should be noticed that theinitial upward trendof pH is due to the acid consuming reactions suchasthe dissolution of copper oxides, chalcocite (Eqs. (1) and (4)), covellite(Eq. (2) and (5)), chalcopyrite (Eq. (3) and (6)) and gangueminerals aswell as the bacterial oxidation of Fe(II) to Fe(III) (Eq. (11)), while thesubsequent decrease of pH is due to the activity of bacteria to produceacid (Eq. (12)), the dissolution of pyrite (Eq. (10)) and the hydrolysis offerric iron to form jarosite (Eq. (13)). Jarosite is formed under theconditions of high pH and high ferric iron concentration, as might beexpected from Eq. (13) (Stott et al., 2000).

3Fe3þ þ Xþ þ 2HSO−4 þ 6H2O→XFe3ðSO4Þ2ðOHÞ6 þ 8Hþ ð13Þ

where X+=K+,Na+,NH4+and H3O+.

Despite using a dilute nutrientmedium (Norris), due to the high solidcontent, the concentration of alkali ions increased in the solution, whichis a favorable factor for precipitation of jarosite at the high solution ORPsfacilitated by iron oxidizing bacteria. Moreover, the precipitation is notreversible in chalcopyrite systems (Leahy and Schwarz, 2009); hence,once formed, the later increase of acid concentration doesn't dissolve theprecipitate. Jarosite formation causes the removal of Fe3+ and essentialbacterial nutrients, suchasK+orNH4

+, fromsolution, potentially resultingin a slowed-down process or even a complete stop. It restricts the flow ofbacteria, nutrients, oxidants and reaction products to and away from themineral surface (Hackl et al., 1995). The stoppage of copper dissolutionafter day 7 is likely related to the passivation of chalcopyrite by jarositeprecipitates. In our previous study (Ahmadi et al., 2010a), a significantamount of jarosite was found on the bioleaching residue, which wasconfirmed by SEM/EDS analyses (Fig. 8). The low leaching rate in thechemical leach could also be caused by the passivation of chalcopyritesurface by polysulphide compounds (Biegler and Horn, 1985). It is likelythat both jarosite and polysulphides passivate jointly themineral surface.

The increase of the rate of iron dissolution and the concomitantdecrease of the rate of copper dissolution in the final days of the

experiment (Fig. 4) are attributed to the leaching of pyrite and couldbe explained by the mechanism described by Petersen and Dixon(2006), in which the dissolution of chalcopyrite is preferentiallyaccelerated at low ORPs, while at higher ORPs, pyrite and covellite areleached much faster than chalcopyrite.

During the analysis of iron dissolution profiles, it should be bornein mind that these values do not take into account the portion of ironleached from the concentrate which was subsequently precipitated asiron hydroxides especially jarosite. This problem occurs to a asignificant extent during the conventional bioleaching test, wherethe ORP values surpass 500 mV and jarosite precipitation wouldoccur. As noted above, this phenomenon was encountered in ourprevious study (Ahmadi et al., 2010a). It should be noticed that irondissolution reported here represents theminimum iron recovery fromthe concentrate.

3.3. Electro-leaching

It can be assumed that the electrical charging of semiconductingmetallic sulphide minerals such as chalcopyrite could be done as aresult of periodic electrical contact between mineral particles and theworking electrode. These contact interactions may occur in theelectrochemical bioleaching experiment when current is applied tocontrol the solution ORP. Hence, to investigate the occurrence of thisphenomenon and its influence on the copper recovery and irondissolution from the concentrate, an electro-leaching experiment wascarried out at a current density of 1 mA cm-2 (total direct curren-t=100 mA) and 20% (w/v) pulp density. Such a low current densitywas chosen to minimize the evolution of hydrogen (Eq. (14)) and was

90 A. Ahmadi et al. / Hydrometallurgy 106 (2011) 84–92

in the range of the current passing in the electrochemical bioleachingtests discussed in the next subsection

2Hþ þ 2e−→H2 ð14ÞThe values of copper recovery and iron dissolution from the

concentrate by the electro-leaching method are also presented inFigs. 3 and 4. Fig. 3 shows that the final copper recovery by electro-leaching (~20%) is significantly higher than that in the chemicalleaching process, but substantially lower than those in both theconventional bioleach and the electrochemical bioleaching processesdiscussed below. However, iron dissolution is initially very low,whereas, the final dissolution is significantly higher than that obtainedin the chemical leaching test (Fig. 4). As Fig. 6 depicts, during the electro-leaching process, the ORP profile is slightly below that measured in thechemical leaching process. Because of low ORP, the dissolution of pyriteas the main iron bearing phase is very slow. Moreover, similar to whatwas measured in the other experiments, the solution pH rises initiallydue to the initial acid consuming reactions (Fig. 5).

Using the electro-leachingmethod, some reactionswould be expectedto occur as a result of passing direct current across the slurry. As noted byWarren et al. (1982), electrochemical reactions of a mineral are a directresult of the thermodynamic properties of the mineral, properties of theelectrolyte, and their interaction at the mineral-electrolyte interface.

The data in Figs. 3 and 4 suggests that during the electro-leachingprocess, the extraction ratio of Fe:Cu rises from0.28 in day 1 to 0.45 at theend of the test. This increase could be related to the electro-reduction ofchalcopyrite to chalcocite (Eqs. (15)and (16))and removing iron fromthechalcopyrite lattice as described by Biegler and Swift (1976).

2CuFeS2 þ 6Hþ þ 2e−→Cu2S þ 2Fe2 þ 3H2S ð15Þ

CuFeS2 þ 3Cu2þ þ 4e−→2Cu2s þ Fe2þ ð16Þ

This result was confirmed by mineralogical analysis (Fig. 7d), inwhich chalcocite was found around chalcopyrite particles in the solidresidue (blue-gray region). Chalcocite could also be producedbyelectro-reduction of covellite according to the following equation (Elsheriefet al., 1995).

2CuS þ 2Hþ þ 2e−→Cu2S þ H2S ð17Þ

This reduction increases the overall dissolution rate of covellite orstage (II) of chalcocite leaching (Eq. (8)) as the rate of stage (I)chalcocite leaching (Eq. (7) is more rapid.

Biegler and Swift (1976) reported that at high current densities(N10mAcm−2) metallic-copper could be deposited on the cathodeelectrode. This deposition may be as a result of a cathodic reactiondescribed by Eq. (18) or due to the direct electroplating of Cu2+ insolution (Eq. (19)).

Cu2S þ 2Hþ þ 2e−→2Cu∘ þ H2S ð18Þ

Cu2þ þ 2e−→Cu∘ ð19Þ

It should be noted that to check the formation of copper on thecathode during the process, a portion of the slurry was occasionallytaken out and returned instantly to the reactor. No copper deposit wasobserved at any stage during the test. However, if copper is deposited,it can be a reductive agent for chalcopyrite leaching according toEq. (20) as reported by Hiskey and Wadsworth (1981). Therefore thedeposit could not have been formed.

2CuFeS2 þ Cuo þ 2Hþ→Cu2S þ Fe2 þ H2S ð20Þ

In terms of galvanic interactions between different metallicsulphide minerals, Mehta and Murr (1983) observed that whenpyrite, chalcopyrite, chalcocite and covellite are in contact with eachother, due to their different rest potentials the rate of covellitedissolution would be the fastest of all (anodic corrosion), followed bychalcocite and chalcopyrite, with the rate of pyrite dissolution thelowest of all (cathodic protection).

Furthermore, it has been reported (Warren et al., 1982; Lu et al.,2000) that chalcopyrite may be electrochemically converted to othercopper sulphide phases such as talnakhite (Cu9Fe8S16) (Eq. (21)) orbornite (Cu5FeS4) (Eqs. (22) and (23)), which would be dissolvedfaster than chalcopyrite.

9CuFeS2 þ 4Hþ þ 2e−→Cu9Fe8S16 þ Fe2þ þ 2H2S ð21Þ

5CuFeS2 þ 12Hþ þ 4e−→Cu5FeS4 þ 4Fe2þ þ 6H2S ð22Þ

2CuFeS2 þ 3Cu2þ þ 4e−→Cu5FeS4 þ Fe2þ ð23Þ

H2S produced during the process could be oxidized in the presenceof Fe(III) (Eq. (24)) and/or oxygen (Eq. (25)) or undesirably led to theprecipitation of covellite in the presence of Cu(II) (Eq. (26)).

H2S þ 2Fe3þ→S∘ þ 2Fe2þ þ 2Hþ ð24Þ

H2S + 1=2O2→SB + H2O ð25Þ

Cu2þ þ H2S→CuS þ 2Hþ ð26Þ

3.4. Electrochemical bioleaching

Two electrochemical bioleaching experiments were carried out,one at an ORP in the range of 400–430 mV and one at an ORP in therange of 440–480 mV over 10 days.

The values of copper recovery and iron dissolution from theconcentrate by electrochemical bioleaching are shown in Figs. 3 and 4,respectively. As can be seen in Fig. 3, the highest copper recoveryamong the various processes was obtained in the experimentconducted at the ORP range of 400–430 mV, in which copper recoveryafter 10 days reached 77%. When the ORP was kept higher, in therange of 440–480 mV, Cu dissolution leveled off after the 3rd daydespite showing the highest initial rate of copper extraction in alltests. The final recovery (66%) was significantly lower than thatobtained during the experiment conducted in the range of 400–430 mV, however it was still higher than that obtained by conven-tional bioleaching (~51%), electro-leaching (~20%) and chemicalleaching (~13%). The higher initial copper extraction rate in the first3 days of electrochemical bioleaching at 440–480 mV could be relatedto the higher ORP values in this period which is more favorable for theleaching of chalcocite and covellite minerals. In these biotic tests, theinitial iron concentrations, originating from the inoculated solution,were 3.15 g/L (ferric iron concentration=2.21 g/L) and 2.35 (mostlyas ferric iron) g/L in tests conducted at 400–430 mV and 440–480 mV,respectively. Ferric iron acts as a leaching agent for copper sulphidesand would lead to a high initial rate of copper dissolution from theconcentrate. Fig. 4 illustrates that the dissolution of iron duringelectrochemical bioleaching in the range of 400–430 mV increaseslinearly up to around 30%, as compared to 23%, 16%, 9% and 5% duringelectrochemical bioleaching at the range of 440–480 mV, conventionalbioleaching, electro-leaching and chemical leaching processes, respec-tively. It should be noted that in electrochemical bioleaching thedissolution of iron is significantly lower than that of copper (29%against 77%), similar to what was observed in the other processes

91A. Ahmadi et al. / Hydrometallurgy 106 (2011) 84–92

studied. The main reason for the low Fe:Cu extraction ratio inelectrochemical bioleaching could be the low solubility of pyrite as themain iron bearing phase in the concentrate at the prevailing potential.Mineralogical analysis of the solid residue from the electrochemicalbioleaching at 400–430 mV (Fig. 7e) shows that the chalcopyrite wascorroded much more, while the pyrite surface remained almostunaffected.

The amount of current passing through the reactor varied from100 to 450 mA. As has been shown, both, applying current directlyinto the slurry and controlling the ORP, have a positive effect on boththe chemical and biological sub-systems. In this regard, the increasedcopper extraction rate is postulated to be due to the three followingreasons: firstly, by electrochemical reduction of ferric iron, theconcentration of ferrous iron which is a source of energy for bacteria,increases, so their growth and activity would be enhanced. Enumer-ation of bacterial populations in the final solutions showed that thecell density increased from about 4×108 cells/ml for conventionalbioleaching to about 9×108 cells/ml for electrochemical bioleachingat 400–430 mV. Previous work done by the authors (Ahmadi et al.,2010a) also showed that after 5 days the cell concentrations of both,mesophilic and moderately thermophilic cultures, were about 3–4fold higher in electrobioleaching slurries than those counted in therelated conventional bioleaching processes. Natarajan (1992) andNakasono et al. (1997) also found that applying a current into asolution containing Acidithiobacillus ferrooxidans increases the cellconcentrations substantially. Industrially, this result is very important,because low bacteria–solids ratios have been cited as one of the mainproblems of the bioleaching processes at high pulp densities (Baileyand Hansford, 1993). The second reason is that the highest dissolutionrate of chalcopyrite occurs at low solution ORPs around 400 mVwhichhas been explained by Hiroyoshi and co-workers (Hiroyoshi et al.,1997, 2000, 2001). This effect has also been confirmed by otherauthors (Pinches et al., 2001; Third et al., 2002; Sandström et al., 2005;Cordoba et al., 2008; Gericke et al., 2010). Moreover, SEM/EDAXexaminations of our previous research (Ahmadi et al., 2010a) showedthat electrochemical bioleaching of chalcopyrite concentrate ataround 425 mV, significantly reduces the amount of jarosites on thechalcopyrite surface. Conner (2005) also demonstrated that applyingDC current into the bioleaching solutions significantly reduces theformation of jarosite in the solid residues. The third reason is that, asobserved in the electro-leaching experiment, periodic electricalcontact between chalcopyrite and the working electrode electro-chemically reduces chalcopyrite to more soluble minerals such aschalcocite and covellite. Mineralogical observations of the residue ofelectrochemical bioleaching at 400–430 mV clearly visualized thecoverage of chalcocite and covellite minerals on the surface of notleached chalcopyrite which is attributed to the reduction ofchalcopyrite to these secondary copper sulphides through both directelectron transfer to the mineral and indirect reduction by Fe(II) at thelow solution ORP. This reason was verified by electrochemicalanalyses (Ahmadi, 2010c) and was also reported previously byBiegler and Swift (1976).

4. Conclusions

Conventional and electrochemical (bio)-leaching were explored toextract copper from Sarcheshmeh chalcopyrite flotation concentrateat high pulp density using a mixed culture of moderately thermophilicbacteria adapted to 20% (w/v) pulp density. Leaching of copper wasfound to be most efficient during bioleaching with the ORPelectrochemically controlled between 400 and 430 mV. The finalcopper recovery of electrochemical bioleaching at 400–430 mV (77%)was higher by a factor of 5.9, 3.9, 1.5 and 1.17 relative to that ofchemical leaching (control test), conventional bioleaching, electro-leaching and electrochemical bioleaching at 440–480 mV, respectively.Mineralogical observations of the sold residues of electrochemical

processes showed that chalcopyrite is converted to chalcocite andcovellite minerals which is attributed to both direct electron transfer tothe mineral and indirect reduction by Fe(II) at the low solution ORP.

Prevention of formation of passive layers such as jarosites as aresult of low slurry ORP, increase of the bacterial concentration andthe electro-reduction of chalcopyrite to less refractory minerals suchas chalcocite and covellite are considered to be the main reasons forenhancing both, the dissolution rate and final copper recovery of theconcentrate in the electrochemical bioleaching process.

From the results of this research work, it can be concluded that theelectrochemical bioleaching could be considered as one of the mostpromising alternatives to extract copper from high grade chalcopyriteconcentrates. In this regard the development of the process byestablishing a continuous system and the assessment of economicalparameters are needed to justify conducting the process in largerscales.

Acknowledgments

The authors would like to appreciate the support of the NationalIranian Copper Industries Company (NICICO) especially Mr. RezaAtashdehghan, Mr. Saeid Ghasemi and Mrs. Zahra Manafi.

References

Acevedo, F., Gentina, J.C., 2007. Bioreactor design fundamentals and their application togoldmining. In: Donati, E.R., Sand,W. (Eds.), Microbial Processing of Metal Sulfides.Springer, Netherlands, pp. 151–168.

Ahmadi, A., 2010c. Process design and kinetics modeling of copper electro-bioleachingfrom sulphide minerals. Ph.D. thesis (in English), Shahid Bahonar University,Kerman, Iran.

Ahmadi, A., Schaffie, M., Manafi, Z., Ranjbar, M., 2010a. Electrochemical bioleaching ofhigh grade chalcopyrite flotation concentrate in a stirred tank reactor. Hydromet-allurgy 104, 99–105.

Ahmadi, A., Ranjbar, M., Schaffie, M., Manafi, Z., 2010b. Optimization of copperconcentrate bioleaching by a mixed culture of moderately thermophilic bacteria.Proceeding of Copper 2010 Conference. GDMB, Hamburg, pp. 2575–2589.

Bailey, A.D., Hansford, G.S., 1993. Factors affecting bio-oxidation of sulfide minerals athigh concentrations of solids: a review. Biotechnology and Bioengineering 42 (10),1164–1174.

Batty, J.D., Rorke, G.V., 2006. Development and commercial demonstration of theBioCOP™ thermophile process. Hydrometallurgy 83, 83–89.

Biegler, T., Horn, M.D., 1985. The electrochemistry of surface oxidation of chalcopyrite.Journal of the Electrochemical Society 132, 1363–1369.

Biegler, T., Swift, D.A., 1976. The electrolytic reduction of chalcopyrite in acid solution.Journal of Applied Electrochemistry 6, 229–235.

Conner, B. D., 2005. Bioleaching and electrobioleaching of sulfide minerals. M.Sc. thesis,West Virginia University, Morgantown, USA.

Cordoba, E.M., Munoz, J.A., Blázquez, M.L., González, F., Ballester, A., 2008. Leaching ofchalcopyrite with ferric ion. Part IV: The role of redox potential in the presence ofmesophilic and thermophilic bacteria. Hydrometallurgy 93, 106–115.

Dimitrijevic, M., Kostov, A., Tasic, V., Milosevic, N., 2009. Influence of pyrometallurgicalcopper production on the environment. Journal of Hazardous Materials 164,892–899.

Elsherief, A.E., Saba, A.E., Afifi, S.A., 1995. Anodic leaching of chalcocite with periodiccathodic reduction. Minerals Engineering 8 (9), 967–978.

Gericke, M., Govender, Y., Pinches, A., 2010. Tank bioleaching of low-grade chalcopyriteconcentrates using redox control. Hydrometallurgy 104 (3-4), 414–419.

Habashi, F., 1978. Chalcopyrite; its chemistry andmetallurgy. Mac-GrawHill, New York.Hackl, R.P., Dreisinger, D.B., Peters, E., King, J.A., 1995. Passivation of chalcopyrite during

oxidative leaching in sulfate media. HydrometaIlurgy 39, 25–48.Hiroyoshi, N., Hirota, M., Hirajima, T., Tsunekawa, M., 1997. A case of ferrous sulfate

addition enhancing chalcopyrite leaching. Hydrometallurgy 47, 37–45.Hiroyoshi, N., Miki, H., Hirajima, T., Tsunekawa, M., 2000. Amodel for ferrous-promoted

chalcopyrite leaching. Hydrometallurgy 57, 31–38.Hiroyoshi, N., Miki, H., Hirajima, T., Tsunekawa, M., 2001. Enhancement of chalcopyrite

leaching by ferrous ions in acidic ferric sulfate solutions. Hydrometallurgy 60,185–197.

Hiskey, J.B., Wadsworth, M.E., 1981. Electrochemical processes in the leaching ofmetal sulfides and oxides. In: Kuhn, K.C. (Ed.), Process and FundumentalConsideration of Selected Hydrometallurgical Systems. Society of MiningEngineering, New York, pp. 303–323. Chapter 26.

Leahy, M.J., Schwarz, M.P., 2009. Modelling jarosite precipitation in isothermalchalcopyrite bioleaching columns. Hydrometallurgy 98, 181–191.

Lu, Z.Y., Jeffrey, M.I., Lawson, F., 2000. An electrochemical study of the effect of chlorideions on the dissolution of chalcopyrite in acidic solutions. Hydrometallurgy 56,145–155.

92 A. Ahmadi et al. / Hydrometallurgy 106 (2011) 84–92

Mehta, A.P., Murr, L.E., 1983. Fundamental studies of the contribution of galvanicinteraction to acid-bacterial leaching of mixed metal sulfides. Hydrometallurgy 9,235–256.

Mishra, D., Kim, D.J., Ahn, G.J., Rhee, Y.H., 2005. Bioleaching: amicrobial process of metalrecovery; a review. Metals and Materials International 11 (3), 249–256.

Nakasono, S., Matsumoto, N., Saiki, H., 1997. Electrochemical cultivation ofThiobacillus ferrooxidans by potential control. Bioelectrochemistry and Bioener-getics 43, 61–66.

Natarajan, K.A., 1992. Effect of applied potentials on the activity and growth ofThiobacillus ferrooxidans. Biotechnology and Bioengineering 39, 907–913.

Okibe,N., Gericke,M., Hallberg,K.B., Johnson,D.B., 2003. Enumerationand characterizationof acidophilic microorganisms isolated from a pilot plant stirred-tank bioleachingoperation. Applied and Environmental Microbiology 69, 1936–1943.

Olson, G.J., Brierley, J.A., Brierley, C.L., 2003. Bioleaching review part B: progress inbioleaching: applications of microbial processes by theminerals industries. AppliedMicrobiology and Biotechnology 63, 249–257.

Petersen, J., Dixon, G., 2006. Competitive bioleaching of pyrite and chalcopyrite.Hydrometallurgy 83 (1-4), 40–49.

Pinches, A., Myburgh, P.J., Merwe, C., 2001. Process for the rapid leaching of chalcopyritein the absence of catalysis. US patent No: 6, 277, 341 B1.

Rawlings, D.E., Dew, D., Plessis, C.D., 2003. Biomineralization of metal-containing oresand concentrates. Trends in Biotechnology 21 (1), 38–44.

Sand, W., Gehrke, T., Jozsa, P.G., Schippers, A., 2001. (Bio) chemistry of bacterialleaching—direct vs. indirect bioleaching. Hydrometallurgy 59, 159–175.

Sandström, A., Shchukarev, A., Paul, J., 2005. XPS characterisation of chalcopyritechemically and bio-leached at high and low redox potential. Minerals Engineering18, 505–515.

Schippers, A., Sand, W., 1999. Bacterial leaching of metal sulfides proceeds by twoindirect mechanisms via thiosulfate or via polysulfides and sulfur. Applied andEnvironmental Microbiology 65 (1), 319–321.

Stott, M.B., Watling, H.R., Franzmann, P.D., Sutton, D., 2000. The role of iron-hydroxyprecipitates in thepassivationof chalcopyriteduringbioleaching.MineralsEngineering13 (10–1), 1117–1127.

Third, K.A., Cord-Ruwisch, R., Watling, H.R., 2002. Control of the redox potential byoxygen limitation improves bacterial leaching of chalcopyrite. Biotechnology andBioengineering 78 (4), 433–441.

Tributsch, H., 2001. Direct versus indirect bioleaching. Hydrometallurgy 59, 177–185.Warren, G.W., Wadsworth, M.E., El-Raghy, S.M., 1982. Anodic behavior of chalcopyrite in

sulfuric acid. In: Osseo-Asare, K., Miller, J.D. (Eds.), Proceedings of the III InternationalSymposium on Hydrometallurgy. AIME, Atlanta, Georgia, pp. 261–275.