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Acidic sulphate leaching of chalcopyrite concentrates in presence of pyrite S.M. Javad Koleini a , Valeh Aghazadeh a , Åke Sandström b,⇑ a Department of Mineral Processing, Faculty of Engineering, Tarbiat Modares University, Tehran, Iran b Division of Extractive Metallurgy, Department of Chemical Engineering and Geosciences, Luleå University of Technology, SE-971 87 Luleå, Sweden a r t i c l e i n f o Article history: Received 1 September 2010 Accepted 30 November 2010 Available on
  Acidic sulphate leaching of chalcopyrite concentrates in presence of pyrite S.M. Javad Koleini a , Valeh Aghazadeh a , Åke Sandström b, ⇑ a Department of Mineral Processing, Faculty of Engineering, Tarbiat Modares University, Tehran, Iran b Division of Extractive Metallurgy, Department of Chemical Engineering and Geosciences, Luleå University of Technology, SE-971 87 Luleå, Sweden a r t i c l e i n f o  Article history: Received 1 September 2010Accepted 30 November 2010Available online 30 December 2010 Keywords: Copper concentrateChalcopyritePyriteLeachingGalvanic effect a b s t r a c t Copper concentrates with mineralogy dominated by chalcopyrite have slow leaching kinetics at atmo-spheric pressure in sulphate media because of the formation of passivation layer on its surface duringtheleaching.Toenhancetheleachingrateofthecopperconcentrate,pyritewasaddedtoactasacatalyst.Pyrite and copper sulphide minerals then form a galvanic cell which increases both the copper leachingrate and yield. Effect of parameters such as solution redox potential, temperature, initial acid concentra-tion, solids content, total initial iron concentration and pyrite to copper sulphide minerals mass ratiowere investigated. Mineralogical analyses by XRD were performed on selected leach residues and thefeed materials. A copper recovery higher than 80% in 24h was achieved at a redox potential of 410mVvs Ag, AgCl, a temperature of 85  C, 15g/L of initial acid concentration, a solid content of 7.8% (w/v), atotal initial iron concentration 5g/L and pyrite to copper sulphide minerals mass ratio 2:1. XRD patternson leach residues showed that candidates for surface passivation, i.e. jarosite and elemental sulphur,were formed at high total initial iron concentrations.   2010 Elsevier Ltd. All rights reserved. 1. Introduction Presently, the smelting/converting technology is the preferredroute to produce copper from sulphide concentrates containingchalcopyrite as the main copper sulphide and sulphides of delete-rious minor elements such as arsenic, antimony, and bismuth(Padilla et al., 2010).The world production of refined copper, estimated by the Aus-tralian Bureau of Agricultural and Resource Economics (ABARE),for the years 2008–2009 was 19.987Mt and around 20% of thatwas produced by hydrometallurgical methods (Helle et al., 2010).Chalcopyrite(CuFeS 2 ),withvalence+1,+3and  2forCu,FeandS, respectively, is the most abundant copper sulphide mineral andalso it is the most refractory copper mineral regarding chemical aswell as bioleaching (Sandström et al., 2005). In order to acceleratethe dissolution of chalcopyrite, various leaching methods havebeen proposed fromlaboratory studies. There are different optionsto enhance chalcopyrite leaching in sulphate media; among theseare: autoclave leaching at high pressure and temperature, use of microorganisms, ultra fine grinding, and the addition of silver ionsor chloride ions (McDonald and Muir, 2007; Carranza et al., 2004; Dreisinger, 2006). In nature, most metal sulphide minerals havesemiconductor properties. The mineral, or the region, with thehighest rest potential will act as the cathode of the galvanic celland is protected whereas that with the lowest rest potential willserve as anode, and its rate of dissolution will be increased (Nicol,1975; Mehta and Murr, 1982; Natarajan and Iwasaki, 1983; Liuet al., 2007; Tshilombo, 2004). Galvanic effects, occurring betweenconductingandsemiconducting mineralsin aqueoussystems, playan important role in the aqueous processing of ores and minerals,such as in flotation and leaching. For semiconductive minerals,such as sulphides, direct contact of different minerals withdissim-ilar rest potentials initiates the galvanic effect. These interactionsoccur between sulphides, involving the flow of electrons fromgrains with a higher potential to grains with lower potentials,modifying the Fermi level of both minerals (Cruz et al., 2005;HolmesandCrundwell, 1995). Mehtaand Murr(1983) experimen- tally studied the galvanic interactions between a series of sulphideminerals, including the pyrite–chalcopyrite couple. The experi-mentalresultsrevealedthatthegalvanicinteractionsbetweenpyr-ite and chalcopyrite significantly affected leaching and extractionof chalcopyrite from a mixed-ore slurry of pyrite and chalcopyrite.The presence of pyrite in the slurry led to a 2–15 times increase inthe rate of dissolution of chalcopyrite. The rest potential of certainsulphide minerals are given in Table 1.Pyrite and chalcopyrite, the most common and exploitable sul-phide minerals, usually occur together and in contact with eachother.Duetothecompanionshipofpyritewithchalcopyriteinnat-ure the co-treatment of these minerals might be advantages froman economic point of view. A novel process for chalcopyrite leach-ing,Galvanox™, basedonagalvanically-assistedleachinghasbeenintroduced(Dixonetal., 2008).Inthisprocessprimarycoppercon-centrates are leached under atmospheric pressure in a ferric-fer-rous sulphate medium in the presence of pyrite. 0892-6875/$ - see front matter    2010 Elsevier Ltd. All rights reserved.doi:10.1016/j.mineng.2010.11.008 ⇑ Corresponding author. Tel.: +46 920 491290; fax: +46 920 491199. E-mail address: (Å. Sandström).Minerals Engineering 24 (2011) 381–386 Contents lists available at ScienceDirect Minerals Engineering journal homepage:  The dissolution reaction of chalcopyrite with ferric ions is writ-ten as CuFeS 2  þ 4Fe 3 þ !  Cu 2 þ þ 5Fe 2 þ þ 2S  ð 1 Þ Chalcopyrite dissolution in sulphate media with ferric sulphateisanelectrochemicalreaction,thereforeitcanbewrittenasanodicand cathodic half-cell reactions as Eqs. (2) and (3) (Mikhin et al., 2004; Misra andFuerstenau, 2005; Watling, 2006; Liu et al., 2007). Anodic half-cell reaction: chalcopyrite oxidation CuFeS 2  !  Cu 2 þ þ Fe 2 þ þ 2S  þ 4e  ð 2 Þ Cathodic half-cell reaction: reduction of ferric ions 4Fe 3 þ þ 4e  !  4Fe 2 þ ð 3 Þ According to Fig. 1a, in the absence of pyrite both the anodicand cathodic half-cell reactions take place on the chalcopyrite sur-face. Dixon et al. (2009) claim that the slow dissolution rate of chalcopyrite is because of a slow cathodic half-cell reaction atthesurfaceofchalcopyrite.Thatmeansthatiftherateofthecatho-dic half-cell reaction could be increased, the dissolution rate of chalcopyrite would be increased. In the presence of pyrite, thecathodic reaction on the pyrite surface is fast which in turn givesfasterdissolutionrateofchalcopyrite(Tshilombo,2004).Accordingto Fig. 1b both the cathodic reaction and the anodic reaction takesplace at the surfaces of pyrite and chalcopyrite, respectively.Instead of ferric ions oxygen can be the active oxidant. In thiscaseoxygenis reducedonthepyritesurface(Eq. (4)). Directoxida-tion of chalcopyrite by oxygen according to Eq. (5) is, according toLiu et al., (2007) negligible in comparison to Eq. (1). O 2  þ 4H þ þ 4e  !  2H 2 O  ð 4 Þ CuFeS 2  þ O 2  þ 4H þ !  Cu 2 þ þ Fe 2 þ þ 2S  þ 2H 2 O  ð 5 Þ In galvanic leaching of chalcopyrite in the presence of pyrite,ferric ions are consumed but are during the process regeneratedby oxidation with oxygen blowing according to Eq. (6) (Holmes and Crundwell, 2000). 4Fe 2 þ þ O 2  þ 4H þ !  4Fe 3 þ þ 2H 2 O  ð 6 Þ However, when the sum reaction of Eqs. (1) and (6) are consid-ered,Eq. (5)isderived,whichthusisthereactionwhichrepresentsthe complete chalcopyrite oxidation. From Eq. (5) it can be seenthat oxidation of chalcopyrite under conditions where elementalsulphur is formed, is an acid consuming reaction.Theredoxpotentialisimportantforeffectiveleachingofchalco-pyrite and in contrast to most other metal sulphides it has beenfound that chalcopyrite leaches better at relatively low redoxpotential (Sandström et al., 2005). Already in 1985, Kametani and Aoki (1985) established the effect of solution redox potential onchalcopyrite leaching in mixed solutions of ferrous and ferricsulphate. The authors reported that there is a critical potential atwhich the leaching rate is maximum and that the rate decreasesabove and below this potential, the optimum redox potential wasfound to be in the range 400–430mV vs SCE. Hiroyoshi et al.(2000, 2001, 2004, 2008) found that copper extraction during oxi-dativeleachingof chalcopyritebyferricionor bydissolvedoxygenin sulphuric acid solution is a function of the ferrous, ferric, cupricions and acid sulphuric concentrations. They proposed a modelconsisting of a two-step reaction where conversion of chalcopyriteis first achieved CuFeS 2  þ 3Cu 2 þ þ 3Fe 2 þ !  2Cu 2 S þ 4Fe 3 þ ð 7 Þ Thenewlyformedchalcociteiseasiertoleachthanchalcopyrite(Sandströmet al., 2005; Aleksandar et al., 1982) and is oxidized byferric ions Cu 2 S þ 4Fe 3 þ !  2Cu 2 þ þ 4Fe 2 þ þ S  ð 8 Þ whenEqs. (7) and(8) aresummedEq. (1) forferricleachingof chal- copyrite is derived. In this study, the galvanic effect between pyrite and chalcopy-rite was used. To enhance chalcopyrite leaching the effect of parameters such as the solution redox potential, temperature, ini-tial acid concentration, solids content, total initial iron concentra-tion and mass ratio of pyrite to copper sulphide minerals wereinvestigated. To study changes in mineralogy, XRD diffractogramswere taken on selected residues and the feed materials. 2. Material and methods  2.1. Material, analytical and instrumentation techniques Copper concentrate and pyrite samples fromSarcheshmeh cop-per mine in Iran were used in the experiments. Mineralogical andchemical analyses of samples were investigated by optical micros-copy and Inductively Coupled Plasma-Atomic Emission Spectrom-etry (ICP-AES)/Quadrupole Mass Spectrometry (ICP-QMS)/SectorField Mass Spectrometry (ICP-SFMS) respectively. The copper con-centrate was used without further preparation whereas the pyritesamplewasfirstcrushedwitharollercrusherandthengroundinarod mill. Size distribution analysis by CILAS 1064 Liquid showedthat particle size of copper concentrate was 90% less than 33.4 mi-cron and for pyrite it was 90% less than 27 l m.  Table 1 Rest potential of some sulphide minerals (Mehta and Murr, 1983). Mineral Chemical formula Rest potential (V vs SHE)Pyrite FeS 2  0.63Chalcopyrite CuFeS 2  0.52Chalcocite Cu 2 S 0.44Covellite CuS 0.42Galena PbS 0.28Sphalerite ZnS   0.24 Fig. 1.  Chalcopyrite leaching by ferric ions (a) without pyrite (b) with pyrite.382  S.M.J. Koleini et al./Minerals Engineering 24 (2011) 381–386   In the leaching experiments, sulphuric acid with purity of approximately 98%, ferrous hepta hydrate sulphate, FeSO 4  7H 2 Oandferricsulphate, Fe 2 (SO 4 ) 3  6.76H 2 Owithhighpuritywereused.AllchemicalsusedwasfromMerckCompany.Deionisedwaterwasused for leaching experiments. Oxygen and nitrogen with highpuritywereusedforcontrolofredoxpotentialintheleachingreac-tor. Copper and total iron was analyzed by AAS (Perkin ElmerInstruments, Analyst 100). A platinum electrode with a Ag, AgClreference electrode was used for redox measurements, while aLange LDO™/sc100 was used for the measurement of dissolvedoxygen in the reactor. All reported redox potentials in this articleare relative to the Ag, AgCl reference electrode. X-ray diffraction(XRD) was performed using a Siemens D5000 automatic diffrac-tometer provided with a continuous scanning device. Cu K a  radia-tion of 40kV and 30mA and a sample rotation of 30rpm wereused. Diffraction patterns were measured in the range from 10  to 90   in 0.02  /step by counting 2s/step and crystalline phaseswere identified using the Joint Committee for Powder DiffractionStandards (JCPDS) file of the instrument.  2.2. Experimental procedure Leaching experiments were performed in a 2.5L glass reactorwith a working volume of 1L. An airtight lid with a condenserwas used to minimize the evaporation from the reactor. Homoge-nousmixingof thepulpwas achievedbyagitationwithapropellerstirrer at a rotation speed of 400rpm. To avoid vortex formationtwo baffles were mounted perpendicular to the vessel wall.Initial redox potential was controlled by adjusting the ferric toferrousionratio.Oxygenornitrogengaswasinjectedcontinuouslyinside the reactor for controlling the redox potential. The redoxpotential of leaching reactor was controlled manually within±15mV. At desired time intervals, 5ml solution was withdrawnfor analysis of copper and iron. Solution lost by sampling andevaporation was compensated by addition of deionised water. 3. Results and discussion  3.1. Mineralogical and chemical analysis of the pyrite and copper concentrate samples Chemical and mineralogical analysis of the copper concentrateand pyrite samples are shown in Tables 2 and 3, respectively. The mineralogical analysis is calculated based on the chemicalanalysis andknowledgefromoptical microscopyof the mineralog-ical phases present in the samples. As seen in Table 3 the majorcoppermineralinthecopperconcentrateischalcopyritewithmin-or amounts chalcocite and covellite. Apart from its presence inchalcopyrite and pyrite smaller amounts of iron were by opticalmicroscopyalsoseeninhematiteandmagnetite.Thepyritesamplewas relativelypure withapyritecontent of approximately90–95%with the remaining mainly as gangue minerals.  3.2. Effect of solution redox potential InFig.2therateofcopperrecoveryatdifferentredoxpotentialsat 85  C, initial acid concentration of 15g/L, solid content of 7.8%(w/v), total initial iron concentration of 5g/L and pyrite to coppersulphide minerals mass ratio of 4 is shown. It is evident that thereisaredoxpotentialwindowintherange,410–440mVinwhichthecopper recovery is decent. This is in close agreement with the re-sults froma number of studies were it has been seen that the opti-mum redox potential for good copper recoveries varies relativelymuch in the range between 400mV and 480mV vs Ag, AgCl(Hiroyoshi et al., 2000, 2001, 2004, 2008; Sandström et al., 2005,Dixon and Tshilombo, 2005, Dixon et al., 2008; Koleini et al.,2010). This relative big difference in optimum conditions is con-centrate dependent and the reasons for this might be differencesin crystal orientation, impurity content or slight variations in min-eral stoichiometry (Tshilombo, 2004).When the optimumredox potential is exceeded the passivationgradually increases and becomes severe at redox potentials above550mVvsAg,AgCl(DixonandTshilombo,2005;Tshilombo,2004). On the other hand oxidation of pyrite starts to be significant atredox potentials above 500mV vs Ag, AgCl (Ou et al., 2007;Tshilombo, 2004). This is of course an unwanted reaction, sincepyrite is needed to achieve the galvanic interaction but also dueto the costs associated with first the consumption of oxidationagent during leaching and later its removal by precipitation inthe down stream processing. Therefore, a redox potential below500mV (vs Ag, AgCl) is the best option since it minimizes bothpassivation of chalcopyrite as well as oxidation of pyrite.In Fig. 3 the XRD patterns of feed and leach residues from theexperiments with solution redox potententials of 410 and460mV are shown. The feed is a mix of the pyrite sample andthe copper sulphide minerals at a mass ratio of 4:1 and the majorphases found were pyrite and chalcopyrite together with quartz asthe gangue material. Chalcocite and covellite were not detected byXRD due to their small concentrations in the feed. The diffracto-gramof the leach residueobtainedafter leachingat 410mVshowsthatmostof thechalcopyritehasbeenleachedat thisredoxpoten-tial whereas the leach residue pattern from the experiment at460mV confirms the low recovery of copper. In addition, tracesof elemental sulphur were found in the residue from the high re-dox potential (460mV) experiment.  Table 2 Chemical composition of the copper concentrate and pyrite (%). Component SiO 2  Al 2 O 3  CaO K 2 O Na 2 O Zn Cu Fe S MoCopper concentrate 16.0 5.1 0.94 1.0 0.24 0.742 23.2 25.1 26.5 0.049Pyrite 1.95 1.0 0.07 0.3 0.03 0.010 0.035 44.1 45.7 0.000  Table 3 Calculated mineralogical content of the copper concentrate. Cu 2 S CuS CuFeS 2  FeS 2  MoS 2  ZnS6.3 2.7 47.7 15.8 0.1 0.8 Fig. 2.  Copper recovery as a function of solution redox potential. S.M.J. Koleini et al./Minerals Engineering 24 (2011) 381–386   383   3.3. Influence of temperature In Fig. 4 the temperature dependence oncopper recoveryat thetemperatures 48  C, 68  C and 85  C is illustrated with the otherleaching conditions constant, i.e. solution redox potential410mV, initial acid concentration 15g/L, solid content 7.8% (w/v), total initial iron concentration of 5g/L and pyrite to copper sul-phide minerals mass ratio of 4. The copper recovery was signifi-cantly elevated and increased from 20% at 48  C up to 84% at85  C. It has been shown in previous studies that leaching of chal-copyrite is controlled by the chemical reaction at the surface withan activation energy around 70–90kJ/mol (Dutrizac et al., 1969;Dutrizac, 1981; Munoz-Castillo et al., 1979; Koleini et al., 2010).Suchhighactivationenergyimpliesthatleachingrateandrecoveryis highly temperature dependent and it has earlier been shown byBerry et al. (1978) that a twofold increase in reaction rate was ob-tained for each 10  C rise in temperature. Similar results on tem-perature dependence were seen in a resent study on the effect of galvanic interaction between chalcopyrite and pyrite (Koleiniet al., 2010). The XRD diffractogram of the leach residue from theexperiment at 48  C confirms that a substantial amount of chalco-pyrite still remains in the residue (Fig. 5).  3.4. Dependence of initial acid concentration The influence of initial acid concentration on copper recoverywas investigated at various initial acid concentration (7.5, 15 and30g/L), solution redox potential 410mV, temperature of 85  C,solid content of 7.8% (w/v), total initial iron concentration of 5g/L and pyrite to copper sulphide minerals mass ratio of 4 is shownin Fig. 6. It can be seen that the initial acid concentration has noor little effect on the copper recovery under the conditions usedin these experiments. However, as discussed previously, leachingof chalcopyrite is an acid consuming process and enough acid forthereactionhastobeprovided.Intheexperimentswiththelowestinitial acidconcentration, pHat the start was around0.8andwhenpH was checked after the experiments it was usually in the rangeof 1.4–1.6, but in some experiments with good copper recoveriespH was as high as 1.8–1.9. These high pH values are at the limitwhere ferric iron starts to hydrolyse and might precipitate as jaro-site.Themainroleofacidistopreventhydrolysisofferricionsandin general an acid concentration of at least 0.1M is sufficient toavoid this (Dutrizac et al., 1969; Dutrizac, 1981). Dixon et al. (2008) showedina galvanicallyassisted copper concentrateleach-ing study that only stoichiometric amount of acid is required andthat further increase in acid concentration had little effect onleaching. Antonijevic and Bogdanovic (2004) showed that at lowpH values passivation occurs by the formation of an iron deficientsolid phase at the surface. Interestingly, it was shown in a thermo-dynamic study calculatedon a solutioncontaining 0.09MFe 3+ and0.2MSO 2  4  thatthemainspecieresponsibleforchalcopyriteoxida-tion was the ferric sulphate complex  ð Fe ð SO 4 Þ  2  Þ  and not the ferricion itself (Córdoba et al., 2008). Tshilombo (2004) claimed that at high initial acidconcentration, the recoveryof copper is higher be-causeof anincreasedgalvanic effect and non-oxidativedissolutionof chalcopyrite. However, in the present research no dependenceon the acid concentration was observed. One of the reasons canbethedifficultiesincontrollingtheredoxpotential.Inourpreviouswork on pure chalcopyrite leaching in the presence of pyrite, the Fig. 4.  Copper recovery as a function of temperature. Fig. 5.  XRD pattern of leach residue from the experiment at 48  C. Fig. 6.  Effect of initial acid concentration on copper recovery. Fig. 3.  XRD pattern of (a) leach residue at 460mV, (b) leach residue at 410mV and(c) feed.384  S.M.J. Koleini et al./Minerals Engineering 24 (2011) 381–386 
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