A review of the structure, and fundamental mechanisms and kinetics of the leaching of chalcopyrite

    A review of the structure, and fundamental mechanisms and kinetics of the leaching of chalcopyrite Y. Li, N. Kawashima, J. Li, A.P. Chandra, A.R. Gerson PII: DOI: Reference:

S0001-8686(13)00019-5 doi: 10.1016/j.cis.2013.03.004 CIS 1267

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Advances in Colloid and Interface Science

Please cite this article as: Li Y, Kawashima N, Li J, Chandra AP, Gerson AR, A review of the structure, and fundamental mechanisms and kinetics of the leaching of chalcopyrite, Advances in Colloid and Interface Science (2013), doi: 10.1016/j.cis.2013.03.004

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ACCEPTED MANUSCRIPT A review of the structure, and fundamental mechanisms and kinetics of the leaching of chalcopyrite

Minerals and Materials Science & Technology, Mawson Institute, University of South Australia, Mawson Lakes 5095

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a.

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Y. Li a, N. Kawashima a, J. Li a, A.P Chandra a, A.R. Gerson a,*

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* Ph: +61 422 112 516, E-mail: [email protected]

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ABSTRACT

Most investigators regard CuFeS2 as having the formal oxidation states of Cu+Fe3+(S2-)2.

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However, the spectroscopic characterisation of chalcopyrite is clearly influenced by the considerable degree of covalency between S and both Fe and Cu. The poor cleavage of CuFeS2 results in conchoidal surfaces. Reconstruction of the fractured surfaces to form, from

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what was previously bulk S2-, a mixture of surface S2-, S22 and Sn2- (or metal deficient sulfide)

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takes place. Oxidation of chalcopyrite in air (i.e. 0.2 atm of O2 equilibrated with atmospheric water vapour) results in a Fe(III)-O-OH surface layer on top of a Cu rich sulfide layer

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overlying the bulk chalcopyrite with the formation of Cu(II) and Fe(III) sulfate, and Cu(I)-O on prolonged oxidation. Cu2O and Cu2S-like species have also been proposed to form on exposure of chalcopyrite to air.

S22-, Sn2- and S0 form on the chalcopyrite surface upon aqueous leaching. The latter two of these species along with a jarosite-like species are frequently proposed to result in surface leaching passivation. However, some investigators have reported the formation of S0 sufficiently porous to allow ion transportation to and from the chalcopyrite surface. Moreover, under some conditions both Sn2- and S0 were observed to increase in surface concentration for the duration of the leach with no resulting passivation. The effect of a number of oxidants, e.g. O2, H2O2, Cu2+, Cr6+ and Fe3+, has been examined. However, this is often accompanied by poor control of leach parameters, principally pH and Eh. Nevertheless, there is general agreement in the literature that chalcopyrite leaching is significantly affected by solution redox potential with an optimum Eh range suggesting the participation of leach steps that involve both oxidation and reduction.

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ACCEPTED MANUSCRIPT Three kinetic models have generally been suggested by researchers to be applicable: diffusion, chemical reaction and a mixed model containing diffusion and chemical components which occur at different stages of leaching. Passivation effects, due to surface diffusion rate control,

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may be affected by leach conditions such as pH or Eh. However, only initial conditions are generally described and these parameters are not controlled in most studies. However, at

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fixed pH, Eh and temperature, it appears most likely leaching in sulfuric acid media in the presence of added Fe3+ is surface reaction rate controlled with some initial period, depending

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on leach conditions, where the leach rate is surface layer diffusion controlled. Although bioleaching of some copper ores has been adopted by industry, bioleaching has yet

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to be applied to predominantly chalcopyrite ores due to the slow resulting leach rates. Mixed microbial strains usually yield higher leach rates, as compared to single strains, as different

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bacterial strains are able to adapt to the changing leach conditions throughout the leach process. As for chemical leaching, passivation is also observed on bioleaching with jarosite

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being likely to be the main contributor.

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In summary, while much has been observed at the macro-scale regarding the chalcopyrite leach process it is clear that interpretation of these phenomena is hampered by lack of

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understanding at the molecular or atomic scale. Three primary questions that require elucidation, before the overall mechanism can be understood are: 1. How does the surface of chalcopyrite interact with solution or air borne oxidants? 2. How does the nature of these oxidants affect the surface products formed? 3. What determines whether the surface formed will be passivating or not? These can only realistically be tackled by the application of near atomic-scale analytical approaches, which may include quantum chemical modelling, PEEM/SPEM, TEM, AFM etc.

Keywords: Chalcopyrite, Oxidation mechanisms, Chemical leaching, Bioleaching, Kinetics.

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ACCEPTED MANUSCRIPT 1. Introduction Chalcopyrite (CuFeS2) is one of the most abundant and widespread copper-bearing minerals [1], accounting for approximately 70% of the Earth’s copper [2]. Although 80 - 85% of

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copper is produced through pyrometallurgical processes, much attention has also been paid to hydrometallurgical processing routes [3]. Hydrometallurgical extraction of copper from

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chalcopyrite is considered to be both more economical and environmentally friendly than

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pyrometallurgical extraction, especially when the copper-bearing sulfide minerals are present at low grade in the ore [4]. In particular heap bioleaching of low-grade copper ore is a

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developing technology that has been successfully applied to copper oxides and secondary copper sulfides in a significant number of operations worldwide but is yet to be applied successfully on an industrial scale to chalcopyrite [5]. To date chalcopyrite leaching has not

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been widely adopted by industry due to the extremely slow leach kinetics [6]. Hence, there is a need to better understand the kinetics and mechanism of chalcopyrite leaching for

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successful industrial hydrometallurgical leaching implementation [7].

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Several studies, mainly using H2SO4 and HCl acidic media, have suggested that the leaching process is inhibited by the formation of a passivating layer [8-12], but there exists

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disagreement as to the nature of the surface layers formed during the leaching process [13] and there is no generally accepted theory as to the mechanism of their formation [2, 3, 14]. Sulfur (S0) [15-19], disulfide (S22-) [20-22], polysulfide (or alternatively named metal deficient sulfide, Sn2-) [8, 23] and Fe hydroxy-oxide [24, 25] identified on chalcopyrite surfaces leached in various media, including ferric sulfate acid [8, 22, 26-28], hydrochloric acid [18, 29, 30] and hypochlorite solution [31], have all been proposed as contributing to a passivating layer with different leach mechanisms proposed, based on the products found [10, 32-36]. However, in contrast Harmer et al. [3] found no evidence of passivation during chalcopyrite leaching when using specific leaching conditions. A number of related factors have been suggested to explain the slow rate of chalcopyrite oxidation [35, 37, 38] with various reagents having been adopted as aids for chalcopyrite oxidation during leaching, Fe3+ being the most common [2, 8, 26, 32, 39]. Other reagents such as, O2 [40, 41], H2O2 [32, 42], nitric acid [43, 44], Cu2+ [6, 45], Ag+ [6] and Cr6+ [16, 46] have also been studied.

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ACCEPTED MANUSCRIPT This review focuses on the chalcopyrite structure, oxidation under atmospheric and aqueous conditions with emphasis on leaching, the effect of galvanic interactions, as well as biooxidation and bio-leaching. In addition, oxidation kinetics and the rate controlling factors are

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discussed. Our aims are to identify the nature of chalcopyrite oxidation processes proposed to date, the factors affecting both oxidation and leaching and the aspects of these processes that

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are still unclear or controversial.

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2. Chalcopyrite Structure

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2.1 Crystal Structure

The crystal structure of CuFeS2 was initially determined by Burdick and Ellis in 1917 at which time they also named the mineral chalcopyrite [47]. Chalcopyrite is a covalent copper

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sulfide [48] that is isostructural with sphalerite (ZnS). However, the c-parameter of the tetragonal unit cell of chalcopyrite is about twice the length of the unit cell of sphalerite.

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A unit cell of sphalerite consists of four Zn atoms and four S atoms with the Zn atom located

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at the fractional coordinate (0, 0 ,0) and the S atom at (0.25, 0.25, 0.25). The distance between Zn and S atoms is 2.34 Å and the shortest distance between two equivalent Zn or S

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atoms is 3.83 Å [48]. In contrast the chalcopyrite unit cell contains four Cu, four Fe and eight S atoms. In each half of the chalcopyrite unit cell along the c-parameter, the four Zn atoms of sphalerite are replaced by two Cu atoms and two Fe atoms in the tetrahedral interstices of the S framework [1, 49]. Each S atom is coordinated by a tetrahedron of metal atoms (two Fe and two Cu) whilst each metal atom is coordinated by a tetrahedron of S atoms, but the S atoms are shifted slightly from the centre of the metal tetrahedra towards the Fe-Fe edge (Fig. 1). Fe occupies a regular tetrahedral arrangement with all Fe-S bond angles equal to 109.47°, whereas the coordination for both Cu and S is not a regular tetrahedron with angles varying from 108.68° to 111.06° [1, 50, 51] (Fig. 1). Chalcopyrite belongs to the I-III-VI2 type semiconductors with tetrahedral coordination [52] and has crystal structure space group I42d. Cu is located at the fractional coordinates of (0,0,0) and (0, 0.5, 0.25), S at (0.2575, 0.25, 0.125) [48, 50]. Fe at (0, 0, 0.5) has spin α while the Fe at (0, 0.5, 0.75) has spin β giving chalcopyrite its antiferromagnetic structure at room temperature [52]. The dimensions of the CuFeS2 unit cell are a = b = 5.2890 Å and c = 10.4230 Å. The Cu-S and Fe-S bond lengths are 2.3020 Å and 2.2566 Å respectively [49] as

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ACCEPTED MANUSCRIPT the S atoms are displaced slightly towards the Fe atoms, and the S-S separations are 3.6850, 3.7408 and 3.7956 Å [1, 49]. However, some variation in these values has been reported by

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Edelbro et al. [48]: dCu-S=2.30 Å, dFe-S= 2.26 Å and dCu-Fe=dCu-Cu=dFe-Fe=3.71 Å. 2.2 Chemical and Electronic Structure

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2.2.1 Oxidation States

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Chalcopyrite is most often considered to be in the Cu+Fe3+(S2-)2 valence state [53-56], rather than Cu2+Fe2+(S2-)2 as has been proposed on occasion [57, 58]). In addition the Cu L- and Fe

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K- and L-edge X-ray absorption near-edge structure (XANES) spectra of chalcopyrite have been interpreted to suggest that some share of the Cu and Fe may be divalent [57]. However,

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the likelihood of variable oxidation states for Cu and Fe in otherwise identical crystallographic positions seems unlikely. Hall and Stewart [49] proposed the effective oxidation state as being between Cu+Fe3+(S2-)2 and Cu2+Fe2+(S2-)2 resulting from significant

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covalent bonding, a similar finding to that by Sainctavit et al. [59].

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The valence band structure of chalcopyrite has been studied using X-ray photoelectron spectroscopy (XPS) [60] and X-ray absorption spectroscopy (XAS) [61] and its electronic

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band structure has been investigated using optical reflectance measurements with photon energies ranging from 0.0025 to 6 eV [62]. S 2p, Fe 2p and Cu 2p XPS peaks always occur as doublets (2p1/2 and 2p3/2) as a result of spin orbit splitting with the 2p3/2 peak having twice the intensity of the 2p1/2 peak. These peaks do not overlap for these elements and hence it is the binding energy of the larger 2p3/2 peak that is usually discussed. XPS has been used to determine the oxidation state of Fe for a range of relevant materials with the 2p3/2 binding energies of Fe2+ and Fe3+ reported to be in the ranges 706.6 - 712.2 eV and 708.7 - 714.4 eV, respectively [50]. In reality, the binding energy of Fe2+ is generally considered to be lower than that arising from Fe3+ [63-65]. The binding energy of Fe, its oxides and other selected compounds, including FeS2, FeO, and FeCO3, are provided in Table 1. Very few mineral sulfides are considered to contain Fe3+ with KFeS2 being an exception with a Fe 2p3/2 binding energy of 708.7 eV [50]. Other researchers have reported that the binding energy of Fe(III)-S is above 707.7 eV while 711-712 eV is indicative of Fe(III)-OOH [64, 66, 67]. Lab and synchrotron XPS Fe 2p spectra collected from chalcopyrite show a strong peak ranging across reported binding energies of 707.0-709.6 eV arising from fully coordinated Fe bonded to sulfur in the bulk chalcopyrite [17, 25, 68-70]. The Fe 2p3/2 binding Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT energy of 708.5 eV for CuFeS2 is considered to be indicative of Fe(III) [21] even though this binding energy is slightly smaller than that of KFeS2 discussed above. However, XPS analyse clearly indicate the presence of Cu(I) (see discussion below) and S2- thus also providing

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evidence for a Fe(III) state. Fujisawa et al. [71] applied XPS, resonance photoemission, Auger-electron, optical

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reflectance and electron-energy-loss (EELS) spectroscopies to the study of the electronic

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structures of CuFeS2 and chalcopyrite-type CuAl0.9Fe0.1S2, concentrating on the valence-band and Fe and Cu states. It was concluded that, compared with CuAl0.9Fe0.1S2, CuFeS2 has an appreciable tail toward larger binding energies in the core Cu 2p and S 2p XPS spectra, and

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that these tails are likely to be the result of energy-loss satellites due to S 3p → Fe 3d excitation, indicating that the S 3p - Fe 3d bonding is substantially covalent. This covalency

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may also contribute to the relatively small binding energy of the chalcopyrite Fe 2p. The valence band peak at the binding energy of 3 eV is recognised as a being due to the Cu

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3d states rather than S 3sp while the S 3s state is assigned to the weak hump around binding

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energy 14 eV [71]. From XPS, ultraviolet photoemission spectroscopy (UPS), EELS, constant initial-state (CIS) spectroscopy, Auger-electron and optical reflectance analyses, the predominant features around the binding energies of 1 and 5 eV were proposed to be mainly

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associated with Fe d5L (L - ligand hole) and d6L2, with the feature around the binding energy of 11 eV assigned to be significantly of Fe d4 character [72]. The strong hybridisation between the Fe 3d and S 3p states in CuFeS2 has been demonstrated by the small chargetransfer energy (△) and large Coulomb energy (U). By comparison to the binding energy of Cu2O, the Cu oxidation state in CuFeS2 and CuAl0.9Fe0.1S2 has been found to be Cu+. In addition a weak divalent component has been detected in CuFeS2 and also Cu2O, which is due to the hybridisation between the occupied Cu 3d states and the empty conduction band [61, 71]. The divalent Cu component only appears in CuFeS2 rather than CuAl0.9Fe0.1S2 suggesting that the empty Fe 3d states are hybridised with the occupied Cu 3d states mediated by the intervening S 3p valence-band states [71, 73]. Ishii et al. [74] and Ghijsen et al. [75] suggest that following the Cu 3p → 4sp core excitation in Cu+ compounds, the M2,3 M4,5 M4,5 super Coster-Kronig-type Auger decay occurs instead of direct recombination as in Cu2+ compounds. The observed XANES chemical shift (K- edge) for Cu2+ compounds, as compared to Cu metal, is between 4.3 and 6.8 eV. The energy of the Cu K- edge of chalcopyrite is just 2.5 eV Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT greater than for the metal, which is in the shift range of 2.1 to 3.0 eV for Cu+ compounds [76]. The S K- and L-edge XANES imply that there is a strong mixing effect between Cu+ and Fe3+ 3d bands and S 3p and 3s states [71]. Petiau et al. [51] investigated the electronic structure of

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chalcopyrite by measuring K-edge XAS spectra of Fe, Cu and S and confirmed the hybridisation of Fe 3d with S 3p states on the basis of the shape of the strong S K-edge

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absorption peak formed (1.8 eV above Fermi level) the tail of which may be due to the hybridisation of cationic 4s orbitals and S 3p. The predominant absorption peak located

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between 7 and 10 eV above the Fermi level represents the empty 4p band [51]. It is generally accepted that Cu in chalcopyrite is predominantly of the 3d10 configuration [61,

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73]. Van der Laan et al. [77], using XAS, investigated the oxidation state of some Cu containing minerals by analysis of the Cu 2p spectra and concluded that for the monovalent

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compounds the ground state is mainly d10 with little d9 character and is mixed with ligand d10Ls character, where L is a ligand hole and s is a valence band electron. Compared with divalent compounds the Coulomb interaction between the 2p hole formed on photoelectron

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excitation and the 3d valence electrons does not play any role for monovalent compounds as

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the 3d shell is fully filled. The Cu+ compounds give rise to a sharp edged peak above 932 eV, due to excitation to the empty Cu s,d states, which is followed by a broad peak due to

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excitation to the empty ligand band. A greater 2p electron excitation energy is required as for Cu+, as compared to Cu2+, as excitation is to the 4s shell as the d10 shell is full, whereas for Cu2+ excitation is to the 3d9 shell. The maximum spectral intensity arising from Cu2+ compounds is around 931 eV.

The XPS binding energy of Cu 2p3/2 (Table 2) is in the range 932.0 - 932.9 eV for Cu+ containing materials and 933.1 - 936.6 eV for Cu2+ containing materials. Furthermore, electron transfer in the photoionised state from the ligand orbitals, in this case the 2p valence orbitals of the S, to the vacant 3d orbital of Cu, i.e. to form a 3d10L state, can result in the satellite peaks observed in the Cu2+ 2p spectra [50, 78]. The XPS spectra of Cu+ compounds do not contain charge-transfer satellites as the 3d shells are fully occupied. The reported binding energy of 932.4 eV for Cu 2p3/2 in CuFeS2 and the absence of satellite peaks both suggest the presence of Cu+ [50, 79]. Using an impurity Anderson model without multiplet structure, Karlsson et al. (1992a, b) calculated the 3d electron populations of the mono-, di- and trivalent Cu compounds of Cu2O, CuO and NaCu2O and concluded that there is large variation in the intensity of the d9- like Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT XPS satellite as well as in the width of the main line. This is despite there being less than a variation of 0.2 3d electrons (calculated) amongst these compounds, which is in contrast to the formal electron occupancies of 10, 9 and 8 3d electrons in the mono-, di- and trivalent Cu

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compounds, respectively. Specifically, they found 9.30, 9.13 and 9.12 3d electron populations for these three compounds. In simulated Cu2O Cu 2p XPS spectra, the d9 satellite is not

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obvious as Cu2O has predominant d10 character whereas in CuO, there is a satellite peak at around +10 eV as compared to the binding energy of the main peak. It was proposed that this

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is due to strong coupling to final states by the d9 ground state character and that it is the 2p3d9 state that gives rise to the satellite feature whereas the 2p3d10L state gives rise to the main

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peak. As expected, the calculated Cu 2p binding energy was observed to increase with the

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changes in valence [80].

Llanos et al. [53] investigated electron transfer phenomena by inserting alkali metals (Li and Na) into host chalcopyrite. The full width at half maximum (FWHM) height of 2 eV and

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binding energy of 932.5 eV for Cu 2p3/2 are in good agreement with the conclusion that only

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Cu+ is present in the chalcopyrite structure [24, 81]. The binding energies of Cu 2p3/2 and Cu 2p1/2 in LiCuFeS2 and NaCuFeS2 were found to be approximately the same as those for

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CuFeS2. However, when compared with CuFeS2, for both LiCuFeS2 and NaCuFeS2, the binding energies of Fe 2p3/2 and Fe 2p1/2 were shifted to lower energies. This result strongly indicates that Fe3+ in CuFeS2 was reduced to Fe2+ when alkali metals were inserted while the deintercalation process resulted in oxidation of Fe2+ to Fe3+, without the participation of the Cu+ in the chalcopyrite.

Based on these observations, Pearce et al.[61] and Fujisawa et al.[71] utilised 2p XPS and Ledge XAS along with Mössbauer data [56, 82] to study the Fe and Cu states in chalcopyrite indicating the presence of high-spin Fe3+ in chalcopyrite. The Cu d electron count was found to be between Cu d9 and d10. Specifically, Cu 2p3/2 XPS data showed there was no difference between CuFeS2, CuAl0.9Fe0.1S2 and Cu2O, and in contrast to the Cu2+ compounds there was no strong satellite structure. Fe 2p XPS cannot be used to determine the Fe d electron population as the d hole states (3d5 for Fe3+ and 3d6 for Fe2+) have electrostatic interactions in the ground state. But high-spin Fe3+ can be confirmed by Fe L2,3 XAS and Mössbauer studies [56]. The shape of the Fe L2,3 XAS spectra for CuFeS2 is similar to the spectrum of KFeS2 which contains high-spin Fe3+, and is different to the spectra arising from high-spin Fe2+ in troilite (FeS) or low-spin Fe2+ in pyrite (FeS2). It was concluded that the oxidation states in Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT CuFeS2 were Cu+Fe3+(S2-)2 [61], which is contrary to Cu2+Fe2+(S2-)2 which has also been proposed [58] but not generally accepted as [58] simplistically compared the Cu and Fe Ledge spectra of CuFeS2 with that of covellite (CuS) and mackinawite (FeS) [61].

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Examination of the electronic structure of chalcopyrite was conducted by Mikhlin et al. [57] who abraded chalcopyrite by using a steel file in a vacuum chamber and then carried out

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XANES analyses. Compared with bornite abraded in vacuum, the Cu L3-edge presented a

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strong pre-edge peak and a small post-edge peak shifted to higher energy. The pre-edge peak was proposed to be due to the electron transition Cu 2p3/2 to 3d at 931 eV indicating a minor occurrence of Cu 3d9, i.e. Cu2+, while the post-edge at 933 eV was proposed to result from

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the transition from Cu 2p3/2 to 4s [57]. The intensity of the Cu2+ main XAS peak is 25 times greater than that of Cu+, which means that even a minor amount of Cu2+ can be readily

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detected [61]. The Fe L2,3-edge energy was consistent with the Fe2+ oxidation state although the amount of Fe2+ was believed less than 20% of total Fe when compared with XPS and Mössbauer data [57]. For S L-edge spectra, a high binding energy tail accounting for 10-15%

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2.2.2 Magnetic Structure

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the excitation of S 3p to Fe 3d.

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of the total sulfur was observed. This may be due to a reduced negative charge resulting from

Compared with the Cu-based chalcopyrite-type compounds (i.e. I-III-VI2 type structures where I = Cu, Ag; III = Al, Ga, In; VI = S, Se, Te), CuFeS2 is considered to be a degenerate semiconductor, that is a semiconductor with such a high level of doping it begins to act more like a metal [71, 83-85] as demonstrated by its special optical, electrical and magnetic properties.

The Neel temperature (above which thermal energy causes the loss of magnetic ordering so that an antiferromagnetic material becomes paramagnetic) of CuFeS2 has been found to be extremely high at 823 K [62, 71, 86]. CuFeS2 can be decomposed to FeS2 plus a disordered face-centred cubic phase, which is close to chalcopyrite in composition and structure, at the slightly higher temperature of 830 K [86, 87]. Studies using neutron diffraction suggest that there exists a magnetic phase change close to 50 K [86], which is detectable by a change to the magnetic space group rather than by changes to the atomic positions [50]. However, based on the calculation of magnetic field strengths, the temperature for transition was suggested to be 190 K, rather than 50 K [88]. At Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT temperatures below this magnetic phase change Woolley et al. [86] and Rais et al. [88] found evidence of spontaneous magnetisation. The spontaneous magnetisation increases on cooling below 190 K where the mass magnetisation is at a minimum, and is predicted to be

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approximately 0.0015 J T-1 kg-1 at 0 K. This value suggests that the Cu magnetic moments remain in a paramagnetic state down to 190 K and order magnetically at lower temperatures

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[88].

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CuFeS2 is suggested to be antiferromagnetic from temperatures above this transition (50 K or 190 K) to 823 K. At room temperature, the magnitude of the Fe magnetic moment is 3.85 µ B, directed along the c-axis, while the Cu magnet moment is 0.2 µ B [50, 84]. By using General

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Structure Analysis System (GSAS) analysis [89] the average magnetic moment (temperature from 4.2 to 300 K) of the Fe atoms in chalcopyrite is calculated as 3.42 ± 0.07 µ B [86].

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Furthermore, when the temperature is below 50 K, a small Cu magnetic moment of 0.05 µ B is suggested by calculating the ratio of m(Cu)/m(Fe) of 0.015 using the same analysis

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simulation system [86, 89].

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Fe2+ has a 3d6 electron configuration and Fe3+ has 3d5 configuration with their high-spin states having 4 and 5 unpaired electrons, together with magnetic moments of 4 and 5 µ B,

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respectively [50]. These two magnetic moments for Fe ions are both greater than those derived from chalcopyrite neutron diffraction data, suggesting the participation of Fe 3d electrons in covalent bonding to the S atoms [84]. It is also possible that there exists low-spin Fe atoms, although this is extremely rare for tetrahedrally coordinated ions [50]. Such a configuration can probably be discounted for Fe in chalcopyrite, as this would yield atomic moments of approximately 2 µ B for Fe2+ and 1 µ B for Fe3+, both of which are much smaller than that measured.

Woolley et al. [86] demonstrated that, in the unit cell of CuFeS2 at temperatures greater than 50 K, the Cu-Fe coupling in chalcopyrite is with four Fe neighbours of Cu being in the form of a square with the spin opposite to that of Cu, with the Cu-Fe spacing of 3.740 Å and with a further four Fe neighbours of Cu in the form of a tetrahedron with spin parallel to that of the Cu with the Cu-Fe spacing of 3.714 Å (Fig. 2, solid line). Density functional theory (DFT) has been employed by de Oliveira et al. [54] to study the magnetic structure of chalcopyrite. Fe in each layer was found to have net spin up or down (Fig. 2) of about 3.26 electrons, which is close to the experimental value of 3.08 [48]. The covalency of the Fe-S bonds results in delocalisation of the spin and resultant reduced values. Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT From the view point of ligand field theory, the tetrahedral structure of the Fe and Cu centres and the weak field of the S2- ligand suggests the presence of Cu+ and Fe3+ as there was no Jahn-Teller distortion found.

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2.2.3 Band Structure

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The band gap of CuFeS2 is in the order of 0.53 - 0.6 eV [90-92]. Petiau et al. [51] proposed on the basis of XAS measurements and band structure calculations [93] that the Fermi level is

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only 0.15 eV greater in energy than the top of the valence band and is mainly composed of Cu 3d, and 0.3 eV lower than the bottom of the conduction band which is mainly composed

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of Fe 3d. The total value of 0.45 eV between the valance band and the conduction band is in reasonable agreement with other band gap measurements. It was proposed that the band 1.8

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eV above the Fermi level is due to hybridisation of Fe 3d and S 3p while the band 2.5 eV above the Fermi level is composed of Fe 4s character. Using data generated using the quantum chemical modelling software CRYSTAL 98 [94],

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Edelbro et al. [48] showed that the bands in the energy interval from -13.8 to -12.5 eV below

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the Fermi level resulted from the S 3s orbitals, which is similar to those of sphalerite. It should be noted that the usual practice is to set the Fermi level at 0 eV [51, 54]. The lower

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part of the valence band, from -6.0 to -3.5 eV, was proposed to be composed mainly of contributions from the S 3sp orbitals while the upper part of this region was found to be a combination of Cu 3d orbitals and S 3sp orbitals together with Fe 3d orbitals with spin α. Furthermore, the bands from -2.5 eV to the Fermi level are divided into three parts: the region closest to the Fermi level is a mixture of Fe 3d, Cu 3d and S 3sp orbitals, the lowest energy region is a mixture of Fe 3d and some Cu 3d orbitals, whereas the middle portion displays mainly Cu 3d features [48]. However, it was also subsequently proposed that the valence band (-2.75 eV up to the Fermi level) is populated by Cu 3d and S 3p orbitals whereas the bottom of conduction band up to 2.0 eV above Fermi level was composed of Fe 3d orbitals [54].

2.3 Surface Structure Chalcopyrite displays poor cleavage resulting in a conchoidal surface rather than cleavage planes. Hence, traditional surface sensitive X-ray photoelectron spectroscopy (XPS) analyses represent an averaged combination of the different surfaces formed [55, 95]. Compared with conventional XPS, synchrotron XPS is well suited to the study of fresh fractured and oxidised Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT chalcopyrite surfaces due to the availability of brighter and more intense X-rays offering greater spatial and spectral resolution along with greater surface sensitivity [69, 96]. Since cleavage planes are not defined, upon fracture a variety of surface orientations exist which

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can contain both cationic and anionic dangling bonds. Three different sulfur species, fully coordinated S2-, S22- with low coordination and Sn2- are generally found to exist on fresh

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fractured chalcopyrite surfaces [3, 55, 95]. Klauber [97] suggests that the S22- species on fresh fractured chalcopyrite results from simultaneous surface reconstruction and redox process

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leading to the development of a pyrite like surface layer. The chalcopyrite surfaces of (001), (100), (111), (112), (101), (110), (012) and their reconstructions have all been studied [54,

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55, 95, 98].

de Oliveira and Duarte [54] examined the (001) S-terminated ((001)-S) and metal-terminated

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((001)-M) surfaces of chalcopyrite using plane wave density functional theory calculations. By calculation of the total density of states (DOS) and local density of states (LDOS) on Cu, Fe and S centres, it was possible to obtain insight into the chemical reactivity of chalcopyrite

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surfaces. It was proposed that surface Cu is likely to be oxidised and that Fe is mostly to be

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reduced to Fe2+, whilst the surface S can be easily oxidised or reduced as S contributes to both the valence and conduction bands [54]. It was observed that a significant reconstruction

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took place on the (001) S-terminated surface as the S-S distance was reduced to 2.158 Å, 1.501 Å shorter than the ideal bulk distance calculated by the same theory [54]. This distance indicates the surface formation of S22- as it is 0.118 Å shorter than the calculated value of 2.276 Å for the S22- molecule (gas phase, Perdew−Burke−Ernzerhof (PBE)/6-311 (d,p) level) and 0.223 Å longer than that of neutral S2 molecule [54, 98-100]. The difference of S-S length between S22- and S2 may be due to the fact that the π* orbitals of S22- are bonded with metal centres on the surface [54]. For the unreconstructed (001)-S surface, in a two unit cell supercell, there are eight S atoms in the first atomic layer and four Fe and four Cu atoms in the second layer [98]. The main difference found, in terms of DOS and LDOS, between the reconstructed surface and bulk is the Fe 3d orbitals were found to contribute to the DOS from -3 eV to the Fermi level, suggesting surface Fe3+ is reduced as S2- is oxidised to S22-. The bond length of S-S calculated in this instance was 2.15 Å before reconstruction and 2.12 Å after reconstruction while the bond length between S and metals on the second layer is 2.32 for the unreconstructed surface and 2.30 Å for the reconstructed surface [98].

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ACCEPTED MANUSCRIPT In contrast, for the (001)-M surface, in the two-unit cell model, there are four Fe and four Cu atoms in the first layer with eight S atoms in the next layer [98]. The metal atoms in the first and third layers of the unreconstructed surface move downwards and upwards on

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reconstruction, respectively, while the S atoms in the second layer move upwards. This increases the coordination number of the S and destroys the tetrahedral local symmetry. The

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Fe-Fe and Cu-Cu distances were calculated to be 2.61 – 2.64 Å while the distance of Fe-S and Cu-S was found to be 2.24 and 2.27 - 2.32, respectively, which are close to values also

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derived subsequently [98]. On this surface, a FeS2-type compound was suggested to form as the length of Fe-S is 2.226 Å and the angle of S-Fe-S is 93.4°, which are near to the values

Klauber [97] and Al-Harahsheh et al. [37].

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for FeS2 (2.008 Å and 113.9°). The formation of FeS2 on the surface was also confirmed by

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In contrast to the (001) surface the (100) surface contains Fe atoms with both spin up and down configurations. In a similar manner to the (001)-S surface, in a two-unit cell super-cell, the unreconstructed (100)-S surface contains eight S atoms in the uppermost layer and eight

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metals in the second atomic layer. In the most stable reconstructions, the length of the S-S

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bond formed in the first atomic layer is 2.11 Å when the S-S is bonded to three Cu and one Fe while it is 2.13 Å when the S-S is bonded to three Fe and one Cu [98]. The S atoms in

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(100)-M surface display similar characters to those of the (001)-M surface, on these two Mterminated surfaces, the alternate positions in the centre of squares of Cu or Fe atoms are occupied by S atoms [98]. The difference between these two M-terminated surfaces is that in the (100)-M, the metal atoms are located in alternating rows while in the (001)-M surface there are rows of Cu and rows of Fe [98]. In the unreconstructed (111)-S uppermost surface there are four S atoms in a one unit cell model with each S atom bonded to one metal atom which are in the second layer [98]. During the reconstruction process a four-S chain is produced by bonding two S-S together. In the (111)-M surface, there are two Fe and two Cu in the first layer in the one unit cell model, although these four metal atoms are not exactly in the same plane. Unlike (001)-M or (100)M, there is no infinite plane of metal atoms for the (111)-M surface, with only small units like lozenge formed by four Fe-Cu bonds (2.37-2.45 Å) and one Fe-Fe bond across the diagonal (2.47 Å). The angles of Fe-Cu-Fe in (111)-M are 58 ° and 61 ° while Cu-Fe-Cu angles are 116 ° and 120 °.

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ACCEPTED MANUSCRIPT Von Oertzen et al. [55] conducted molecular modelling on the (012) and (112) surfaces. The former surface is the only surface consisting of an equal number of metal atoms (Cu and Fe) and S atoms, and the latter surface would give rise to only S atoms on one side ((112)-S) and

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only metal atoms on the other side ((112)-M) of the cleavage. The under-coordinated S atoms on the (012) surface are likely to be reduced compared to the fully coordinated bulk S atoms,

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whereas for the (112)-S surface, after the relaxation, the monomer S atoms can form polymer chains. Compared with the fully coordinated S atoms in the bulk, the negative charges of

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these exposed S monomers was reduced to disulfide or polysulfide as observed through calculation of the Mulliken charges [55, 95].

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However, for the (110) and (101) surfaces, the metal atoms in the first layer relocate to the position between the S atoms remaining in the first layer and the S-M chains in second

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atomic layer. No new bonds are formed in these surfaces but the lengths of both Cu-S and FeS decrease when reconstruction happens [98].

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Based on all the reconstructed surfaces discussed above, three reconstruction mechanisms

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have been proposed [98]. The first mechanism is for the formation of S-S bonds, as for the surfaces of (001)-S, (100)-S. The S atoms in the first atomic layer lose two bonds and new S-

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S bonds are formed as evaluated through analysis of the electron localisation functions (ELF) and DOS [98]. This reconstruction is a physical relocation process [97] (Equation 1):  2
  
  2 

1

A freshly fractured surface was studied by Harmer et al. [95] using synchrotron and conventional XPS at low (250 eV) and high (1487 eV) photon energies, respectively. S 2p3/2 spectra revealed the main symmetric peak at 161.33 eV to be due to fully coordinated bulk S atoms. The presence of surface S2- was confirmed by the observation of a shift in S 2p3/2 peak position to 160.84 eV whilst the peak at 161.88 eV was suggested to be due to surface Sn2-. Sn2- species have been proposed to form on the S-terminated surfaces and can be regarded as a further reconstruction of S22- (Equation 2). 
  2
  2  4  

2

The second mechanism [98] consists of the formation of metal-metal bonds in M-terminated surfaces, such as (001)-M, (100)-M, (111)-M. The metal atoms in the bulk are coordinated to four S atoms in a tetrahedral geometry. After reconstruction the metal atoms in the first atomic layer lose two bonds while σ M-M bonds are formed by overlapping of dxy, dxz and Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT dyz orbitals. The distance of these M-M bonds are found to vary from 2.45 Å to 2.67 Å [98]. The third reconstruction mechanism is actually the metal atom relaxation process. During this process, there are no new covalent bonds found [98]. The metal atoms in the first layer move

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downwards to the position between the first two layers to increase the S-M-S angles, reducing the angular strain.

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According to Shuey [101], there exists substantial covalencey between the S anions and metal

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cations in the CuFeS2 structure. In order to break the covalent bonds, the electrons in the uppermost valence band need to be captured by a suitable redox couple, which is known as a hole formation process [102]. When electrons in CuFeS2 are extracted by Fe3+, or other

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suitable redox active species, these chemical bonds may be broken. These broken chemical bonds can be still present on the CuFeS2 surface leading to a greater interfacial reactivity and

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increased dissolution or oxidation rate [103]. 2.4 Impurity Elements

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Chalcopyrite-type semiconductors of the form I-III-VI2 have been the subject of much

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interest due to their potential for next-generation optical applications, such as solar batteries, nonlinear optical devices and luminescence diodes [104-106]. The band gaps (Eg) energies of

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Cu-III-VI2 (III=Al, Fe, Ga, In; VI=S, Se, Te) semiconductors range from 0.6 to 3.49 eV [104, 107], among which the materials with Eg in the range of 1.70-3.49 eV are viable for application as light-emitting devices in the visible and ultraviolet range [104]. Among these compounds, CuFeS2 is the only magnetic semiconductor [83]. Teranishi and Sato [92] investigated the role and contribution of 3d orbitals to the optical, electrical and magnetic properties of both n- and p-type CuFeS2. Liu et al. [108] suggested that it is easier to make p-type rather than n-type chalcopyrite through substituting cations/anions into the CuFeS2 structure with decreased/increased valencies. For Fe-doped CuAl(Ga)1-xFexS2 (x=00.008) strong optical absorption bands were observed at 1.3 and 2.0 eV using a Cary 14 spectrophotometer and infrared [92]. In addition, there was a rising absorption edge at the higher energy of 3.3 eV, which was 2.7 eV higher than that of CuFeS2 [92, 107]. This can be explained by a charge transfer transition between valence band and empty 3d states [92]. CuAlS2, with the widest band gap (3.49 eV) at room temperature, normally displays p-type mobility due to the propensity for defects such as Cu vacancies [104, 108]. It has an ordered tetrahedral array of flattened CuS4 and undistorted AlS4, giving rise to strong hybridisation Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT between Cu 3d and S 3p within the valence band maximum (VBM), with the conduction band being of Cu 4s character [108]. Liu et al. [108] enhanced hole conduction by substituting Zn for Al (CuAl1-xZnxS2, x = 0 - 0.10) as doping with Zn (3d10) slightly perturbs

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the Cu sub-lattice and induces the formation of Cu2+ resulting in a significant increase in conductivity. On Mn doping of CuGaS2 and CuAlS2 the Mn-S lengths are about 2.41 Å,

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which is a typical value for Mn-S tetrahedral bonds in Mn-bearing sulfides [109].

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The electronic structures of CuFeS2 and CuAl0.9Fe0.1S2 have been investigated by Fujisawa et al. [71]. Kinetic energies of 918.2 and 916.8 eV for Cu L3M4,5M4,5 were observed for CuFeS2 and CuAl0.9Fe0.1S2, respectively. For the Cu 3p (UPS) excitation region, the binding energy of

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CuAl0.9Fe0.1S2 was found to be 1.1 e V greater than that of CuFeS2. By comparison with other monovalent Cu-bearing compounds, the Cu oxidation state in both CuFeS2 and CuAl0.9Fe0.1S2

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is confirmed as Cu+ although some d9 configuration has been found, in some instances, along with the d10 ground-state configuration in CuFeS2. The appearance of a partial d9 configuration may be due to the occupied Cu 3d orbitals being hybridised with empty Fe 3d

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states via the intervening of S 3p [71].

Schneider et al. [106] observed that trace Fe impurities resulted in blackening of the CuGaS2 single crystals and resulted in very strong absorption of visible and near infrared light. This

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absorption may be ascribed to charge transfer transitions within Cu+-Fe3+ pairs. Also, electron paramagnetic resonance (EPR) has been applied for examination of point defects in CuAlS2, CuAlSe2 and CuGaSe2 with signals being assigned to Cu vacancies and residual Fe impurities [110].

The chemical and thermodynamic conditions applied during growth and doping play a key role in determining the resulting semiconductor type (n- or p-). Cu(In,Ga)Se2 usually forms as a p-type semiconductor under Cu-poor conditions whilst only CuInSe2 can be easily grown as an n-type semi-conductor under Se-poor and In-rich conditions [111]. Stephan et al. [112] established a mechanism by which to invert p-type into n-type conductivity by doping halogen atoms (Cl, Br and I) into CuInSe2. It has also been observed that p-type chalcopyrite undergoes a phase changing to n-type chalcopyrite at the pressure of 12.5 GPa [113]. Aguilera et al. [114] proposed the AgInS2 can be transformed from n-type to p-type by doping with Sn.

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ACCEPTED MANUSCRIPT 3. Atmospheric Oxidation 3.1 Oxidation Products

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There are a myriad of studies examining chalcopyrite oxidation in aqueous media but few on the oxidation of chalcopyrite by exposure to atmospheric O2 with natural atmospheric

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moisture or water. The nature and progress of this form of oxidation is important as the resultant layers may slow further oxidation or even subsequent aqueous leaching [8, 11, 21,

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33]. Studies [21, 23] have been conducted to investigate the products resulting from chalcopyrite oxidation. However, many aspects of the oxidation layer(s) formed are still not

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understood completely due to the complexity and variability of the interactions between chalcopyrite and the atmosphere.

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XPS has been used by Brion [24] to study the oxidation products of CuFeS2 in air or distilled water. They observed that little degradation of the first surface layer occurred after dry grinding and exposure to air for a few minutes. Fe hydroxide/oxyhydroxide was initially

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observed, followed by the formation of a basic iron sulfate which increased significantly with

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time. It was also found that on prolonged oxidation the surface Cu concentration decreased but with no change in chemical state. A broadening of S 2p peak to higher binding energy

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was observed and was suggested to be due to the –S2– groups formed on the Fe-depleted surface. In contrast, when the freshly ground chalcopyrite was exposed to distilled water, in the absence of O2, only Fe hydroxide/oxyhydroxide was detected within the first few atomic layers, without any evidence for sulfate or S0. These findings indicate surface oxidation products vary with exposure to different media (air or water). By conducting oxidation in air coupled with XPS measurements, Yin et al. [21] investigated the oxidation products formed on the surface of chalcopyrite electrodes, and proposed the formation of Fe2O3 and CuS. Auger electron spectroscopy (AES) depth profiling was employed to examine the oxidised CuFeS2 with the thickness of the new layer being about 4 nm, equivalent to approximately 7 monolayers [21]. Analogous results were obtained by Eadington [115] where the oxidation layer was found to be 1.2 ± 0.2 nm after exposure of CuFeS2 in dry air for 30 min and 1.5 ± 0.2 nm for in air-saturated deionised water for 1 min. Biegler and Horne [116] found that the thickness of the layer formed was about 3 nm and was composed of CuS and other S compounds.

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ACCEPTED MANUSCRIPT Using X-ray absorption (XAS) and emission (XES) spectroscopy Todd et al. [117] explored the oxidation products formed under atmospheric conditions (T=293 K, PH2O ≈ 0.02 bar). After exposure for one week, they observed the Cu L-edge (XAS) at 931.4 eV, implying that

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there was a change in coordination from Cu(II)-S to Cu(II)-O as the absorption peak position is nearly the same as that of CuO. But, this explanation of the product formed (CuO) has been

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called into question by Pearce et al. [61] as the spectra of Cu2O used to draw this conclusion contained a small fraction of CuO. The intensity of Cu L- edge of Cu2+ is 25 times stronger

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than that of Cu+ and hence Cu(II)-O was identified rather than Cu(I)-O. The asymmetric peak found at approximately 951.0 eV showed a similar feature to the Cu L-edge spectra of Cu2O

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and Cu2S, which may be due to the 3d-4s hybridisation [118]. Furthermore, the Fe L3-edge at 709 and 710.5 eV showed the presence of Fe(III) oxides or Fe-O-OH while the S L-edge peak

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at 173.5 eV suggested the formation of sulfate [117]. XES analysis further supported the conclusion that there were both sulfate and (oxy)hydroxide species formed, probably including, FeOOH, Cu2O, Cu2S and some cupric and/or ferric sulfates [117]. However,

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sulfate is not generally detected in significant concentrations, unless as a secondary

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precipitate, upon chalcopyrite oxidation in water due to its considerable solubility [117]. Using XPS, Fe and Cu L2,3, and S and O K-edge XANES analyses Goh et al. [119] proposed

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that the chalcopyrite surface was initially covered by Fe oxide/hydroxyoxide species, rather than Cu oxide as suggested by Mikhlin et al. [57], on exposure to air for a few minutes. For extended exposure (less than one day), metal (Cu and Fe) 2p XPS and L-edge XAS in combination with O K-edge XAS confirmed the presence of Fe oxide species accompanied by rather small concentrations of Cu (I)-O (Cu 2p3/2 932.1 eV). Similarly, XPS (Fe 2p) was used by Buckley and Woods [25] who found evidence of iron hydroxy-oxide on the surface of CuFeS2 on exposure to air while Cu remained bonded to sulfur as CuS2. Furthermore, Cu sulfate (CuSO4) was confirmed to be present on extended atmospheric exposure [25]. By using AES, XPS and optical microreflectometry (OMR), Ruzakowski et al. [120] proposed the formation of the oxidation products of Cu5FeS4 and Cu2S between the outer layer of iron oxide and the bulk chalcopyrite. They also suggested the formation of iron hydroxide and iron sulfates in a thin outer layer of 0.5-1.0 nm depth. The formation of Fe3O4 or FeOOH, as the product of polishing chalcopyrite in air, was proposed by Holloway et al. [121]. They also found the evidence of formation of a copperrich sulfide of composition either Cu5FeS4 or Cu2S. After further exposure to air, Cu and S Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT components could not be examined due to the formation of a Fe oxide layer of a thickness of 20-40 nm. Recently, synchrotron scanning photoemission microscopy (SPEM) has been used to

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determine the surface speciation of CuFeS2 oxidised in air for 15 min [96]. 18 %S and 19 %S at S 2p3/2 binding energies of 161.7 eV and 162.9 eV respectively were attributed to disulfide

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and polysulfide after oxidation in air while the bulk monosulfide at 161.1 eV still contributed

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over half of the surface S. No sulfate was apparent. 3.2 Proposed Mechanisms

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Several possible oxidation pathways have been proposed for oxidation in air to give rise to the products described above. Equation 3 shows a possible reaction between chalcopyrite and

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H2O as presented by Gardner and Woods [122], showing the formation of Fe(OH)3, CuS and S0 on the chalcopyrite surface.

3

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 3  3   
  3    
   
  


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However, the concentration of CuS was always found to be greater than that of S0 although the quantities of both materials are equal according to Equation 3. This may be due to the

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sublimation of S0 during XPS analysis [25, 69]. Buckley and Woods [25] suggested the formation of sulfate and further secondary minerals are produced from the initial S rich Cu layer which they proposed to be CuS2 due to Fe migration to the surface to form an over layer of oxide (Equation 4 and 5):  
  2   
    


 
  3  
   
   


0    1

4 5

The S 2p3/2 peak near 168.3 eV indicated the presence of sulfate after extended exposure to air. The formation of sulfate was accompanied by Cu(II) which was characterised by a higher binding energy than that of Cu (I) as well as the presence of satellites in the 2p spectrum [25]. There is evidence to suggest that the freshly fractured chalcopyrite adopts a pyritic surface structure, implying Fe3+ is reduced to Fe2+ in the reconstruction process [37, 97, 98]. During this process, an electron is released:          
    
  

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ACCEPTED MANUSCRIPT When chalcopyrite is exposed to air, O2 may accept the released electron forming Cu+ or Cu2+ oxides (Equation 7 and 8): 4    4   2  


7

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       2 

8

2 
    2 
  2 


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2 
      
  2 


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In summary the overall oxidation process in air can be showed in Equation 9 and 10 [97]. 9 10

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As the OCP (open circuit potential) of pyrite is greater than that of chalcopyrite (Section 6, Galvanic Interactions), it is unlikely that the formation of pyrite as an oxidation product of

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chalcopyrite, as depicted in Equations 9 and 10, would occur. However, the oxidation of chalcopyrite in air involving the reduction of O2 only, or O2 in the presence of water, would give rise to a net positive standard potential for chalcopyrite reconstructing and pyrite

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production [97].

4. Aqueous Oxidation and Leaching

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4.1 Surface Species

Many lixiviants can be applied to copper leaching, such as sulfates, chlorides, ammonia and nitrates [8], of which ferric sulfate has been the most frequently used due to its operational simplicity and low cost [123].

Although considerable research has been carried in order to understand chalcopyrite leaching in various lixiviants, for instance H2SO4, HCl, HClO4 etc., with and without Fe3+ addition, the resulting surface chemistry has not been well understood. Specifically, no universal agreement has been reached on the composition of the new layers formed on the CuFeS2 surface [124] and the role of the new surface species in the leaching process. A number of surface species within the layer have been proposed, e.g. sulfur (S0) [15-19], disulfide (S22-) [20-22], polysulfide (Sn2-) [8, 23] and Fe hydroxy-oxide [24, 25]. Relevant S-containing species are listed in Table 3 along with their respective 2p3/2 binding energies. Dutrizac [15] reported that on the CuFeS2 surface, over 94% of the S is in the form of S0 and less than 6% is in the form of sulfate upon leaching in the presence of Fe2(SO4)3 at 95 °C.

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ACCEPTED MANUSCRIPT Hackl et al. [8] suggested the formation of copper rich polysulfide CuSn (n>2) in the surface layer and confirmed that the new layer of CuSn was formed due to the solid state changes at the high temperature range of 110-200 °C. A similar result has been found by Todd et al.

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[117] who proposed CuSn (n>2) was the main component in the surface layer resulting from the leaching process.

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Klauber et al. [124] and Klauber [27] reported the predominant surface species formed on

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CuFeS2 leaching with added Fe3+ in sulfuric acid are S0, and S22-, but no trace of Sn2- was identified. They identified the main component in the passivating surface layer on CuFeS2 particles as being S0, which may impact the identification of Sn2-. However, other studies

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have reported that the presence of a porous S0 layer may not impede the leaching rate to a

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significant extent [23, 125].

Based on XPS data, a surface composition of near Cu0.8S2 was suggested by Buckley and Woods [25] on leaching of freshly fractured chalcopyrite in 0.2 M acetic acid media for 40

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days. The presence of Cu1.3FeS2.7 was proposed when the freshly fractured chalcopyrite was

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immersed for 1 h whilst Cu1.4FeS3 was the product upon 24 h leaching. The specimen leached in acetic acid for 60 days was further immersed in hydrogen peroxide (H2O2) for 5 min,

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resulting in a surface layer containing S0 as well as Fe-deficient Cu sulfide. The use of air saturated acetic acid (a weak acid) medium for leaching of the fresh fracture surfaces of chalcopyrite was chosen to enable the examination of the properties of copper and sulfur products (or metal-deficient sulfide layers) in the absence of iron-oxygen compounds as these species formed dissolve in this medium. A sample oxidised in air for 6 days was leached for 18 h in hydrochloric acid (HCl), leading to the formation of Cu1.1FeS2.8. This acid-treated surface was further exposed to air for 3 days with a Fe-deficient Cu sulfide identified as the main surface component. During the electrochemical oxidation process, various passivating layers have been found, the formation of which depends on the electrolyte pH and the potential applied. Yin et al. [21] have employed XPS analysis and suggested that Fe was preferentially leached from the chalcopyrite surface, and a passivating film of CuS2 was formed at potentials of less than 770 mV SHE in 1 M HClO4. At greater potentials (> 940 mV SHE) S0 was formed in the strongly acidic media. Furthermore, in an alkaline solution (pH 9.2), the surface layer was found to consist of CuS2 and Fe2O3 (at potentials < 540 mV SHE) or S0, CuO and Fe2O3 (at potentials >

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ACCEPTED MANUSCRIPT 740 mV SHE). However, the passivating effect was less obvious in weak acidic (pH 6.8) or strongly alkaline solutions (pH 13) [21]. Harmer et al. [3] investigated the changes of surface speciation during CuFeS2 leaching in

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perchloric acid (HClO4) at pH 1 and 85 °C for a duration of 313 h (no Fe3+ was added to the leach liquor). The initial outermost CuFeS2 surface layer of Fe oxyhydroxide (2 nm Fe(III)-

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O-OH) was removed after 2 h of leaching, leading to the exposure of a Fe-poor but Cu-rich

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layer (Fe at.% from 59.8 to 18.1 % and Cu at.% from 14.9 to 42.6 %, all at.% normalised to Cu, Fe and S only). After 24 h, Fe(III)-S was observed and the relative amount of Fe increased from 18.1 % to 55 at.%. After 144 h of leaching, the Cu 2p3/2 X-ray photoelectron

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binding energy of 932.3 eV indicated the formation of Cu(I)-Sn2- and the concentrations of Cu and Fe became approximately the same at the end of the leach duration. From XPS

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spectra of S 2p3/2, the concentrations of S0 and Sn2- were found to increase dramatically after 24 h while the total S, increased from 21.7 % to 78.3 %, suggesting the major component in the surface layer is S compounds. Even though, no passivation effect was observed as the

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leach rate increased throughout the whole leaching process, which is in contrast to the

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passivation observed by other investigators [8, 15, 21, 25, 58].

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Using two-dimensional computer simulator, Nazari and Asselin [126] simulated the morphology of CuFeS2 leaching in acidic Fe2(SO4)3 solution based on the percolation model. In this model, CuFeS2 was presented as a composition of Cu, Fe and S atoms, distributed randomly in a ratio of 1:1:2. They controlled the reaction potential to be the leach active region of chalcopyrite (600-700 mV SHE) by controlling Fe3+ concentration in the media. The result showed that in the active redox region, Fe was selectively leached out prior to Cu, forming a Fe-deficient Cu polysulfide layer. This newly formed layer was confirmed by Holliday and Richmond [127] who pointed out that the dissolution rate of Fe was five times faster than that of Cu. The formation of S8 ring structures was also observed in the active condition in the model from Nazari and Asselin [126]. Acres et al. [69] observed that the S0 formed during CuFeS2 leaching in acidic solutions (pH 1 HCl) may sublime during XPS measurements if no cooling is applied. However, Harmer [128] found that the samples (leached for 144 h) stored in the introduction chamber of the XPS instrument at a pressure of 10-6 Torr for 12 h still contained a considerable percentage of S0. Sublimation of Sn2- species during XPS analysis was not reported.

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ACCEPTED MANUSCRIPT Acres et al. [69] utilised synchrotron XPS, NEXAFS and time-of-flight secondary ion mass spectrometry (ToF-SIMS) to determine the leach products from chalcopyrite, in the absence and presence of pyrite, exposed to pH 1 HCl solution for 2 h. it was found that surface

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concentrations of S22- and Sn2- are much greater on chalcopyrite particles when pyrite is present, suggesting the leaching rate was enhanced by pyrite. Meanwhile, the oxidation rate

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of a rough homogeneous sample was much greater than that of a smoother one, with an increase of 12 % in the surface area contribution of Sn2- (163.4 eV). On the smooth surface,

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nearly half of the S present was in the form of Sn2-, 35 % was bulk S2- and the rest was S22- or associated with the energy loss feature. Acres et al. [69] also found that there was no sulfate

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present on the leached surface. Furthermore, XPS and NEXAFS data suggests that the Cu present at the leached surfaces (both rough and smooth chalcopyrite) is predominantly in the

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form of Cu(I) sulfide, with very little Cu(I) oxide and no evidence of Cu(II) oxide, which is likely to dissolve upon formation at pH 1. Nevertheless, Fe 2p3/2 XPS spectral analysis supported the existence of hydrated Fe(III)-O-OH (at binding energy of 711-712 eV) at the

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surface indicating that the rate of formation rates is more rapid than the rate of dissolution.

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Thiosulfate (S2O32-), although not having been detected directly, has been suggested to be present as an intermediate product on leaching of CuFeS2 in sulfuric acid solution in the

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presence of Fe3+ [35]. S0, SO42- species and a S22- phase were identified with the SO42- being proposed to be a basic ferric sulfate akin to jarosite, causing passivation [35]. Meanwhile, the formation of S2O32- and Sn2- were found to have no direct effect on the leach rate [35]. Minerals with the same elemental composition but different crystal structure can be distinguished by Raman spectroscopy [129]. This technique was used by Parker et al. [130] to examine in situ the surface of CuFeS2 leached in the presence of FeCl3 in 0.1 M HCl solution. Results showed that the S-S bonds were formed resulting in a polymeric sulfur phase which was very stable at temperatures under 70 °C while over 70 °C it decomposed dramatically. However jarosite, as well as thiosulfate and polythionate were not detected [130]. When chalcopyrite was leached for a prolonged period in sulfate electrolytes (Fe2(SO4)3 with 0.1 M H2SO4, the leach rate became much slower than in HCl. The presence of polymeric sulfur was also found in the case of sulfate media, but it was different in nature to that found in HCl media, with shorter S-S bonds formed as indicated by a decrease in the wavenumber of the SS stretching band [130].

Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT Through analysis of XRD and SEM data Córdoba et al. [11] demonstrated that jarosites, as the products of hydrolysis from Fe3+ in ferric sulfate solution, contribute to the passivation of CuFeS2 rather than porous S0 in the pH range of 1.5 and 2.0 at 68 °C. Furthermore, O2 was

PT

determined to play a key role in the regeneration of the oxidizing agent (Fe3+) and the Cu extraction rate was accelerated significantly when compared to anaerobic conditions.

RI

Klauber [27] proposed that the chalcopyrite dissolution process in acidic ferric sulfate

(I)

SC

solution could be regarded as a 4-step process:

An initial reaction on the fresh chalcopyrite surface with high rate and low activation energy (Ea);

Formation of a thick S0 layer retards the dissolution and results in low rate and

NU

(II)

high Ea;

Removal of the thick S0 layer results in an increased linear rate which will

MA

(III)

continue if the formation of jarosite is hindered by the experimental conditions with controlled pH, temperature, concentration of Fe3+; Jarosite is produced in the absence of controlled pH and Fe concentration,

D

(IV)

TE

resulting in obscuration of the surface of chalcopyrite, reducing the reacting

AC CE P

surface area and resulting in a parabolic rate leach curve. Interestingly, through XPS and EDS, Kaplun et al. [131] found that there are no other significant Cu containing species on the surface of leached chalcopyrite while Fe(III)-S was detected as the Fe-containing species with some minor fractions of Fe(III)-O-OH. The addition of Fe3+ to the leach solutions increased the formation of Sn2- and SO42-. It was also found that the addition of Fe3+ resulted in high initial surface oxidation rate resulting in an unfavourable balance of S-containing species on the chalcopyrite surface, leading to a reduced leach rate in the later leaching stage [14, 131]. Redox potentials in the experiments carried out by Kaplun et al. [131] were proposed to be a very important factor and were controlled at the intermediate value of 750 mV (SHE) throughout to enable the simultaneous occurrence of oxidation and reduction reactions. 4.2 Factors Influencing Reaction Kinetics The dissolution rate of chalcopyrite has been reported, usually, to increase with increased oxidant concentration [32, 46] and/or temperature [8, 132, 133]. It is accepted that temperature plays an important key role in determining the dissolution kinetics of

Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT chalcopyrite [132, 134, 135]. The type and speed of agitation, particle size, acid concentration, oxidants and pulp density are also important considerations. Cu may be released into acid solution in the absence of O2:

PT

 
  4        2 

11

RI

Equation 11 is classified as a non-oxidative dissolution because the valence state of S remains the same and no net electrons transfer to the solution. As discussed previously, CuFeS2 has

SC

the oxidation states of Cu+Fe3+S2 [82, 119, 136], but Cu+ is readily oxidised upon exposure to air and water, hence, the status of Cu on the surface of CuFeS2 would be a mixture of Cu+ and

NU

Cu2+ prior to Cu2+ dissolution [135].

Normally, dissolved oxygen (DO) exists in solution in a relatively low concentration (<10 mg

MA

L-1) and the chalcopyrite dissolution rate is extremely slow under this condition (Equation 12):  
  4        2

12

D

O'Malley and Liddell [137] proposed that Cu2+ was likely to be mainly extracted by the

TE

dissolution process as compared to Cu+. After Cu2+ which has been oxidised on the surface is released, Cu+ on the unoxidised surface will be rapidly exposed to Fe3+ present in the acidic

AC CE P

solution, hence, Fe2+ and Cu2+ are expected to be the main aqueous components, depending on solution Eh. It has been determined by solution speciation calculation that free Fe2+ activity was more than ten times greater than Fe3+ activity at pH 1 in H2SO4, HClO4 or HCl solutions at 750 mV (SHE) and 75 °C while Cu2+ is the predominant Cu species in these solutions [14]. Most recently, solution speciation calculation using Phreeqc software with the LLNL (Lawrence Livermore National Laboratory) database showed that Fe2+ is the dominant Fe species up to an Eh value of 850 mV (SHE), above which free Fe3+ and Cu2+ will be the main free components in H2SO4 media. The most common reaction (Equation 13) in relation to chalcopyrite relating to acid rock drainage (ARD) is the oxidation due to the presence of Fe3+ which may be oxidised from Fe2+ by microorganisms [135]:  
  16    8      17    2  16

13

As the solubility of Fe3+ is relatively great in low pH media, the concentration of Fe3+ in ARD is expected to be high while the product solubility of Fe hydroxide decreases

Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT dramatically with increasing pH. Therefore, the extraction rate of chalcopyrite is suggested to be highly dependent on the solution pH values [135].

PT

4.2.1 Effect of Speed and Type of Agitation Antonijevic and Bogdanovic [32] found the oxidation rate of chalcopyrite in H2SO4 media

RI

with addition of Fe3+ was independent of stirring speed, suggesting the mechanism was not bulk diffusion controlled. Similarly, Dutrizac [26] observed chalcopyrite dissolution in either

SC

ferric sulfate or ferric chloride media is independent of stirring speed. In contrast, Sokic et al. [132] reported a slightly decreasing leach rate with increasing speed (100 to 450 rpm) in

NU

H2SO4 with added NaCl which they proposed was due to decreased contact between the particle and oxidant, as has also been proposed by other investigators [42].

MA

Adebayo et al. [105] conducted leaching in H2SO4 + H2O2, with occasional stirring as they considered that the decomposition rate of H2O2 increases with increasing stirring speed, which enhances the formation of molecular O2 resulting in bubble adsorption onto the

TE

D

mineral powder surface, impeding the contact between sample and peroxide. In an H2SO4 + K2Cr2O7 chalcopyrite leach system the dissolution rate increased with

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increasing stirring speed until 400 rpm, over which the leaching rate decreased dramatically [46]. Similar results have been found by Harmer [128] upon leaching of chalcopyrite in HClO4 with and without the addition of Fe3+. The greatest Cu dissolution rate was observed at 500 rpm. From 500 to 1000 rpm, the leaching rate decreased due to the reduced concentrations of reactants such as Fe3+, Fe2+, O2, H+ on chalcopyrite surface. When the speed was below 500 rpm, it was proposed that the mixing was not adequate, resulting in a lower leaching rate than that of the highest point value at 500 rpm [128]. Dong et al [138] observed that at stirring speeds less than 600 rpm, the stirring speed affected copper dissolution significantly in a Brønsted acidic ionic liquid, but there was no difference between 600 and 800 rpm. Nicol et al. [139] observed that in acidic chloride solution the Cu dissolution rate is much greater with magnetic agitation than mechanical agitation with a titanium impellor as the magnetically stirred reactor appeared to provide more abrasion amongst the particles, hence, mechanical agitation was adopted to reduce the abrasion effect on stirring [139].

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ACCEPTED MANUSCRIPT 4.2.2 Effect of Temperature The activation energy (Ea) of chalcopyrite dissolution can be derived by measuring the temperature dependence of the leach rate, which can also be used to determining the

PT

dominant leaching mechanism being undergone [131]. For chemical controlled processes, leaching rate can be enhanced by increasing temperature by only a few degrees. If the

RI

influence of temperature is weak, together a mild effect of agitation speed, it can be inferred that the kinetics of leaching is controlled by diffusion rather than chemical reaction [29]. It

SC

should be noted that a value of Ea over 40 kJ mol-1 implies chemical reaction (linear leaching) while less than 40 kJ mol-1 suggests the leaching process is controlled by diffusion (parabolic

NU

leaching) [29].

MA

A significant increase of reaction kinetics in acidic chloride solutions is obtained by increasing the reaction temperature from 60 to 70 °C, but from 70 to 90 °C, the leaching rate was stable [29]. Sokic et al. [132] found that dissolution was enhanced from 28 % to 70 %

D

within 240 min in 1.5 M H2SO4 solution when the temperature was increased from 70 to

TE

90 °C. Significant effects of temperature on leaching have also been observed by Córdoba et al. [2] in the temperature range of 35 - 68 °C where the Cu leached increased from less than

AC CE P

3 % to more than 80 % and Ea of 130.7 kJ mol-1 was derived. Padilla et al. [41] pointed out that Cu dissolution in H2SO4 solution at temperatures between 120 - 150 °C could reach over 90% within 45 min under the condition of high oxygen pressure (1216 k Pa).The Ea will be further reviewed in Section 4.3. 4.2.3 Effect of Particle Size

Models relating particle size to dissolution kinetics have been established, such as shrinking core models [32]. The selection of appropriate particle size ranges for industrial applications is important as this will determine the power consumption, reactor design and other preparation processes prior to leaching [29]. Dependence of reaction kinetics on particle size occurs when the reaction rate is diffusion controlled through a product layer, also called a transport process. In this case the diffusion rate is always proportional to the inverse square of initial particle radius while chemical reaction controlled processes are associated with the inverse of the initial radius [29] As expected, the smaller the particle size, the faster the leaching rate [105, 132, 139]. However, Dutrizac [26] observed that the rate of Cu extraction is independent of surface area Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT in either ferric chloride or ferric sulfate solutions. It was suggested that when the particle size was smaller than 75 µm, copper extraction becomes efficient in H2SO4 acidic media in the presence of K2Cr2O7 [46]. In contrast, Padilla et al. [41] reported no obvious effects on

PT

reaction kinetics occurred when the initial particle size was in the range of 49-89 µm.

RI

4.2.4 Effect of Acid Concentration and Type

SC

The effect of acid concentration on reaction kinetics requires thorough investigation since it directly determines process control and economics [29]. Small pH values can minimise

NU

hydrolysis and precipitation of Fe3+.

H2SO4 has been the most widely used lixiviant for chalcopyrite leaching [33, 41, 105, 140]. For copper leaching, the leach rate has been found to increase with increasing concentration

MA

of sulfuric acid [46]. The reasonable range of H+ concentration has been suggested to be 0.11.0 M [29]. A pH less than 0.5 was observed to promote passivation due to the Fe deficient

D

surface resulting from the competition between Fe3+ and H+ [32]. This effect becomes more

TE

obvious when the acid is very concentrated at 3 - 5 M [32]. However, when the acid concentration was increased to 6.0 M, Cu extraction proceeded dramatically to approximately 80 % within half an hour, which was explained as being due to the increased redox potential

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of H2O2 resulting from increased H+ ion concentration [105]. It was determined that the dissolution rate increased with increased H2SO4 concentration below 1.0 M [105]. Furthermore, a reaction order of 0.77 has been derived as a function of H2SO4 concentration [105].

Dreisinger and Abed [29] reported that HCl was more efficient than H2SO4 as in H2SO4 solution there results reduced activity of Fe3+ due to the formation of FeSO4+ [141]. Furthermore, the increasing concentration of Cl- would lead to a more direct participation of Cl- through particular adsorption or surface complexing. Nicol et al. [139] pointed out that the existence of Cl- can enhance the leaching rate but high concentration is not essential under the conditions of 35 °C and 560-600 mV SHE.

4.2.5 Effect of Oxidant Concentration Various reagents have been adopted as aids for chalcopyrite oxidation, such as, O2 [40], H2O2 [3, 14, 131], sodium nitrate [132], Cu2+ [45, 139], Cr6+ [16, 46] with Fe3+ being the most commonly applied [2, 8, 26, 32, 39]. Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT Nicol et al. [139] have found that a small amount of cupric ions (Cu2+ <0.1 g L-1) improved the Cu extraction rate in chloride solution but further increases in the initial concentration of Cu2+ had no obvious effects on leaching rate. This was confirmed by Hiroyoshi et al. [45]

PT

who also suggested that with the addition of Cu2+, an intermediate product, Cu2S, was formed at low potential region (about 560 mV SHE), resulting in greater leaching rate as Cu2S can be

RI

oxidised more rapidly than CuFeS2.

SC

Increasing the concentration of H2O2 can significantly enhance the oxidation rate of chalcopyrite in sulfuric acid media [32] while the concentration of K2Cr2O7 has no apparent effect on copper extraction with a pulp density of 10 g L-1 in the range of 0.1 - 0.15 M of

NU

oxidant [46]. However, Adebayo et al. [105] illustrated that in sulfuric acid solution there was a linear relationship between Cu leaching rate and H2O2 concentration between 10 and 15%.

MA

With increased initial concentration of H2O2, nonlinear deviation was observed due to the greater decomposition rate at greater concentration, resulting in reduced H2O2 concentration.

D

Fe3+, the most used oxidant, has been found to significantly impact chalcopyrite leach rates

TE

when the concentration of Fe3+ is at a low level (up to 100 mM) [14, 131, 142]. Below this value, increasing the concentration of Fe3+ has a positive effect on extraction rate while above this range increasing the concentration has little influence on leaching rate [142]. Li et al. [14]

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and Kaplun et al. [131] reported that 4-8 mM of Fe3+ leads to increased leach rates during the initial stages of chalcopyrite leaching in sulfuric acid solution but resulted in reduced leaching rates thereafter.

4.2.6 Effect of Redox Potential (Eh) Li et al. [14] and Kaplun et al. [131] propose that redox potential (Eh) impacts the leaching rate significantly. In order to clearly understand the passivating effects that take place during the leaching processes the passivation potential (Epp) has been used to distinguish the potentials where chalcopyrite leaches effectively and where it does not. Viramontes-Gamboa et al. [12, 143] demonstrated the active-passive behaviour of chalcopyrite in a wide range of Epp. Chalcopyrite leached in sulfuric acid media was always in the active state below 685 mV SHE no matter the content of impurities, acidity or temperature. Between 685 and 755 mV SHE, chalcopyrite was in a bistable system, implying it could be passive or active resting with the way it was brought to that potential. Specifically, a passive potential of 755 mV SHE was observed when the potential was increased from the active state at a temperature over 50 °C. However, if the chalcopyrite surface was already passivated, by decreasing the Minerals and Materials Science & Technology Mawson Institute, University of South Australia

29

ACCEPTED MANUSCRIPT potential to bring it back to the active region a reactive potential was found at 685 mV SHE. From 755 mV SHE upwards chalcopyrite leaching was slow due to a strong passivating effect.

PT

Sandström et al. [144] suggest that in sulfuric acid media a small Eh (620 mV SHE) is beneficial to chalcopyrite leaching when compared with a higher Eh, for instance 800 mV

RI

SHE. Passive regions have been observed over the potential range of 690-840 mV SHE at

SC

low temperature (25 °C) or at 840 mV at greater temperatures (65 °C) whilst the active to passive transition region at low temperature was suggested to be from 600 to 700 mV [126].

NU

Velásquez-Yévenes et al. [145] discussed the effect of media potential on the chalcopyrite leaching rate. They conducted experiments at controlled potentials in 0.2 M HCl with 0.5 g

MA

dm-3 Cu2+ at 35 °C and found the leach products were highly dependent on the solution potential. The dissolution rate was improved significantly when potentials were between 540 to 580 mV SHE while the range of 550 - 620 mV SHE was suggested as the active potential

D

window. When the potential was greater than 630 mV SHE, the leaching rate slowed [145].

TE

From an electrochemical viewpoint, Ghahremaninezhad et al. [146] summarised the products

AC CE P

formed on progressive potential increase in acidic media (Table 4). A chalcopyrite electrode was used by Velásquez et al. [147] to confirm the leaching products formed in a borax alkaline solution at pH 9.2. Cyclic voltammetry (CV) and XPS data supported the formation of CuS2 and CuFe1-xS below the redox potential of 640 mV SHE, Fe2O3 and CuO in the range of 640-940 mV SHE and FeOOH, CuFeO2 from 940 to 1040 mV SHE.

Hiroyoshi et al. [6, 148, 149] and Holliday and Richmond [127] have proposed the following reactions of chalcopyrite oxidation (Equation 14 and 15) in sulfuric acid media with dissolved oxygen as the oxidant:

 
  4          2
  2 

4    4    4    2 

14 15

Koleini et al. [150] determined that the greatest Cu recovery could be obtained in the range of 610-640 mV (SHE). Similar conclusions are drawn by many authors [12, 143, 144, 150, 151] although the overall reaction (Equation 16) would suggest that a high redox potential (high concentration of Fe3+) would result in the most rapid copper leaching [6, 8, 26]. Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT  
  4       5    2


16

Equation 16 can also be represented by the following half-cell reaction pair (Equation 17 and  
        2
  4 

PT

18):

18

RI

4    4   4  

17

From Equations 14 and 16, we can see that chalcopyrite oxidation may be able to be

SC

enhanced by the presence of both Fe3+ and O2. By carrying out chalcopyrite leaching under oxidative conditions with dissolved O2 and Fe3+ in acid sulfate solution, Hiroyoshi et al. [149]

NU

found the leach rate in acidic ferric sulfate solutions was significantly enhanced by the addition of significant concentrations of Fe2+ (0.50 mol dm-3) and Cu2+ (0.01 mol dm-3)

MA

together with 0.03 mol dm-3 Fe3+. This observation cannot be inferred from Equation 16 as Fe2+ and Cu2+ are products of this reaction and the leaching rate should be reduced by increasing their concentration. In order to understand this phenomenon, a two-step reaction

D

model was proposed by Hiroyoshi et al. [152]:

TE

 
  3   3    4    2  


AC CE P

 4  

  
  4    2   


19 20

Equation 19 indicates that in the presence of Fe2+, thus requiring low redox potential, and Cu2+, chalcopyrite is reduced to chalcocite (Cu2S). Compared to chalcopyrite, the intermediate Cu2S is more easily oxidised by Fe3+ (Equation 20) and this results in the improved copper leaching rate observed at the low potentials of 560-600 mV (SHE) [139] while a maximum leaching rate was obtained at 630 mV (SHE) [152]. Thus, an optimum Eh range is implied by this model since both reduction and oxidation processes happen successively. However, when the concentration of Cu2+ is small, the presence of Fe2+ may impede CuFeS2 oxidation [149]. The criterion for enhanced copper extraction by Fe2+ and Cu2+ can be represented as Equation 21:     

21

where Ec is the critical potential of Cu2S formation and Eox is the oxidation potential of Cu2S. Under the conditions of 25 °C and 1 atm, these potentials can be expressed as Equation 22 and 23 [149]: Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT   0.681  0.059 log

    

.


22

.

  0.561  0.059 log# .

23

PT

where # is the activity of species of i.

RI

Although this model illustrates the copper extraction process, changes to the surface chemistry were not taken into account. Hence, by applying X-ray photoelectron spectroscopy

SC

(XPS), time of flight secondary ion mass spectrometry (ToF-SIMS) and scanning electron microscopy (SEM), Harmer et al. [3] proposed a three-step reaction pathway: 15  16

 

NU

3 
  12  12

   $4   6 %
  3  4    &'(  ) 2

MA

$3   6 %
  12  12

  12  12  

$3   12  12  6  %
  12  12

 

TE

12  2 

D

$3   12  12  6  %
  12    6
  3  

24

25

26

AC CE P

During the first oxidation step (Equation 24) Cu2+ and Fe2+ are released into solution and monosulfide is polymerised to polysulfide. This is followed by a reduction step (Equation 25), where S2- or other short chain Sn2- are formed on the surface. Fe2+ is oxidised to Fe3+ during this second step and H+ is adsorbed from solution to balance the surface charge. The second oxidation step (Equation 26) may result in the formation of crystalline Sn0, which is produced from the shorter chain Sn2- rather than long-chain Sn2- [3]. The three-step reaction mechanism consists of both oxidation and reduction reactions and thus a cycle occurs that recreates Fe3+ during the leach process. If a suitable Eh range is chosen, both oxidation and reduction can occur simultaneously, or alternatively. Greater and smaller Eh favours oxidation and reduction, respectively. 4.3 Rate Laws Under non-oxidative conditions, the dissolution of CuFeS2 in acidic solutions can be presented as Equation 27 and 28 [15, 23, 139]:

 
  4        2     4    2      2

Minerals and Materials Science & Technology Mawson Institute, University of South Australia

27 28 32

ACCEPTED MANUSCRIPT The maximum rate of H2S production in Equation 27 can be presented as * 

+ 4, / $ %, where km is the mass transport coefficient for H2S from the surface of the

mineral; K is the equilibrium constant for Equation 27 and [H+] is the surface proton

PT

concentration. H2S is consumed in Equation 28 so that Equation 27 can continue [139]. However, Equation 27 is not thermodynamically spontaneous as K is only 2.8×10-19 at 35 °C

RI

[139], hence the following reaction (Equation 29) is more thermodynamically favourable under non-oxidative acidic conditions due to the higher K value (2.8×10-3 at 35 °C) [139]:

SC

 
  2   
      

29

NU

CuS was detected at potentials < 600 mV SHE in chloride solutions [139]. The flux of copper ions from the surface of CuFeS2 can be expressed as Equation 30 [139]:

MA

 . -  $ % +. . coth 2 +. 3+ 4 

30

D

where J is the flux, [H2S]s is the equilibrium concentration of H2S at the surface; k is the

TE

pseudo 1st order rate constant in 0.1 M acid; D is the diffusion coefficient and km is the mass transfer coefficient for Fe3+ at the surface. Based on Equation 30, Nicol et al. [139] deduced

AC CE P

that when the concentration of Fe3+ increased in the range over 0.1 M, the rate of CuFeS2 dissolution is unlikely to rise rapidly. The rate would not be accelerated with increasing acid concentration and adding Fe2+ should reduce the Cu extraction rate. The Arrhenius equation (Equation 31) can be used to calculate Ea for leaching processes from leach rate constants (k) measured as a function of temperature (T in Kelvin): +  + 

  

31

where k0 is the pre-exponential factor and R is the gas constant (kJ mol-1 K-1), T is temperature (K). Based on this equation, Acero et al. [153] proposed the following chalcopyrite dissolution rate law for pH 3 (Equation 32): (  10. .  


 

32

where r is the dissolution rate of chalcopyrite (mol m-2 s-1). The leach rates were measured using a flow-through apparatus with variable concentrations of dissolved O2 over the temperature range of 25 to 70 °C. Compared with the initially pristine chalcopyrite surface, the leached samples were covered by an enriched S and Cu layer as Fe was released Minerals and Materials Science & Technology Mawson Institute, University of South Australia

33

ACCEPTED MANUSCRIPT preferentially to Cu for all the experimental conditions examined. However, this layer did not hinder dissolution. As a result of the activation energy derived of 32 5 5 kJ mol-1 and the lack

of rate variation with stirring, it was proposed that the dissolution rate was controlled by

PT

surface reactions and was independent of the non-stoichiometric surface layer [153]. In order to determine a ‘universal’ chalcopyrite dissolution rate law Acero et al. [134]

RI

conducted another set of flow-through experiments using different acidic conditions (HCl and (  10. .  $ % .

 . 




  

SC

H2SO4, pH from 1 to 3) and derived a further leaching rate law (Equation 33): 33

NU

where [H+]is the activity of H+. This rate law suggests that the dissolution rate is dependent on temperature and activity of H+, with the rate increasing with decreasing pH. Using HCl

MA

aqueous media at pH 2 and 3 at 298 K, r was found to be (1.5 ± 0.2) × 10-11 and (8.4 ± 1.2) × 10-12, respectively. The Ea for leaching for these conditions was determined to be 31 ± 4 kJ

D

mol-1, which is consistent with a surface controlled dissolution reaction [134, 153].

TE

O'Malley and Liddell [137] proposed that during the chalcopyrite leaching process in the presence of added aqueous Fe3+, the leached Cu+ was likely to be oxidised by the Fe3+. This

AC CE P

has been subsequently confirmed by Orth and Liddell [34] who conducted similar leach experiments in HCl solution. A rotating ring disk electrode was used and the rate equation (Equation 34) was derived.

(  69.7 7 10  7.0 7 10

!" # !$  #

8 $ %$   %

34

This rate was found to be first order with respect to both [Cu+] and [Fe3+] and also it would appear that the rate would increase with respect to increasing pH (in contrast to Equation 33) and Cl- concentration. An empirical dissolution rate law has been proposed by Kimball et al. [135]: (  + 9

   : ;

<  

35

where r is the rate of dissolution (mol m-2 s-1), k0 is the pre-exponential factor in the Arrhenius rate law, Ea is the activation energy (J mol-1), R is the gas constant (J mol-1 K-1), T is temperature (K), mi is the concentration of H+, Fe3+ and Cl- in molarity or partial pressure of O2, ni is the rate order with respect to that reactant. This equation can be transformed into:

Minerals and Materials Science & Technology Mawson Institute, University of South Australia

34

ACCEPTED MANUSCRIPT log (  log + 

> 2.303*=  ∑  log <

36

log[H+], log pO , log [Fe3+], log [Cl-] and 1/T are regarded as the regressors (or independent 2

PT

variables) and log r is the response with a significance level <0.0001.

Kimball et al. [135] compiled numerous rate expressions for CuFeS2 leaching in H2SO4, HCl

RI

and HNO3 media in the presence of O2 with or without Cl- and also derived a more general (  10 .  % ⁄ $ % .

SC

leaching rate equation:

%

37

NU

Equation 37 indicates that the rate of non-oxidative dissolution of CuFeS2 is only dependent on temperature and [H+]. This rate law also suggests that the effect of pO was not significant. 2

MA

This was in accord with previous findings which showed that at 25 °C the rates of dissolution were nearly the same although pO was varied (0.21, 0.05 and 0.005 atm) [153]. 2

TE

further rate law was obtained:

D

Further experiments were conducted with the addition of Fe3+ by Kimball et al. [135] and a

38

AC CE P

(  10 .%%  % ⁄ $ % .% $   % .

The Ea for this model was 48±10 kJ mol-1. This rate law suggests that the dissolution rate was highly dependent on the combined effects of temperature, [H+] and [Fe3+]. In these two rate laws (Equation 37 and 38) if the total reaction rate (rtotal) represents the combination of nonoxidative rate (rH+ ) and oxidation rate by Fe3+ (rFe3+ ), then rFe3+ can be effectively equal to rtotal because at a given condition of pH, temperature and [Fe3+], rH+ was two orders of magnitude smaller than that of rtotal [135]. These two rate laws (Equation 37 and 38) show that [Cl-] is not a significant rate affecting factor, which may be due to the experimental conditions used of small pH. Adebayo et. al [105] studied chalcopyrite leach rate kinetics in H2O2 and H2SO4 media. They derived an Ea of 39 kJ mol-1 and proposed that the rate controlling step was a chemical reaction at the chalcopyrite surface, which was further confirmed as there was a linear relationship between the rate constant and the reciprocal of particle size. In addition, the dissolution rate increased with increasing concentrations of H2O2 and H2SO4. A negative

Minerals and Materials Science & Technology Mawson Institute, University of South Australia

35

ACCEPTED MANUSCRIPT effect on leach rate was observed with increased stirring speed which may due to the accelerated decomposition of H2O2. Similar to the results found by Acero et al. [134] and Antonijvec et al. [16] established a rate

PT

law with a positive correlation of [H+]0.80-0.92 with leaching conducted in H2SO4 with the oxidant potassium dichromate. In the presence of dissolved O2 and Cl-, the rate was found to

RI

be dependent on temperature and [H+] while on the addition of aqueous Fe3+, the rate law

SC

was highly dependent on the combined effects of temperature, [H+]0.8 and [Fe3+]0.42 [135]. Li et al. (2010) deduced a contrasting rate law (Equation 39) by assuming shrinking spherical particles of the same initial volume of 1.

NU

 %

3$ % .

39

MA

2.0 5 0.2 $  @> @A 

S is the relative surface area, C is the Cu concentration in solution (M), t is the time (h), 2.0 is the rate constant (with units M0.7 h-1), and [Fe3+] and [H+] are activities (both in M). This rate

D

law was derived for the conditions of pH 1 and 2 (H2SO4, HClO4, HCl), Eh 750 mV (SHE),

TE

75 °C in the absence of added Fe3+. They did not find any correlation between leach rate and surface speciation, which may arise either directly due to leaching and/or indirectly due to

AC CE P

secondary precipitation, e.g. Fe(III)-O-OH species. A further rate law [154], derived under contrasting conditions, found the leach rate to be first order with respect to Fe2+ activity (Equation 40) under the specific conditions of fixed pH 1 (H2SO4, HClO4, HCl) at 75 °C, but this time, Eh was set at 900 mV (SHE) so that the relative proportion in solution of Fe3+ was much greater than Fe2+. 

@> @A  3.3 5 0.1 $  %

40

These two contrasting equations (Equations of 39 and 40) support the conclusion that in the absence of added Fe3+, solution speciation is rate controlling and an optimum Eh range exists as they suggest contrasting rate determining species whose relative predominance is based on Eh. Further work has subsequently been carried out by Kaplun et al. [131] who derived Ea for Cu and Fe dissolution (Table 5) at the fixed conditions of pH 1 and Eh 750 mV (SHE). However, at this time Equations 39, 40 and the Ea derived have not been incorporated into a single rate law enabling rate prediction as a function of Eh and temperature.

Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT An interesting, and on first analysis contradictory, phenomena has been observed from the previous work conducted by Li et al. [14] and Kaplun et al. [131] and centres on the effect of aqueous Fe3+. At the beginning of leaching experiments with the addition of Fe3+, the rate is

PT

rapid as compared to when no Fe3+ is added (Fig. 3) but is not maintained over the period of the leach. The initial rapid rate can be easily understood from Equation 39 [131] and the

RI

slower subsequent rate can be rationalised on the basis of contrasting surface speciation between the two systems. However, what cannot be explained is why there is not a similar

SC

acceleration in the latter stages of leaching, where Fe3+ was not added initially, due to a significant and comparable build-up of Fe3+ in solution (both experiments were conducted

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with controlled Eh of 750 mV SHE).

MA

4.4 Aqueous Leaching Mechanisms

Chalcopyrite leaching mechanisms can be influenced by varying temperature, stirring speed, Ea, particle size, etc. The overall effects of these variables determine the pattern of dissolution

D

of the Cu. The shrinking core model is the most popular approach to modelling the

TE

dissolution processes. There are essentially three kinetic models for CuFeS2 leaching: diffusion, surface reaction and a mixed kinetic model containing components of diffusion and

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surface reaction [29, 41, 46, 131, 153]. By assuming different controlling steps various kinetic models have been proposed. These and there related mechanisms as summarised in Table 5.

When the reaction is conducted under different conditions, the reaction rate appears to be controlled via a number of different mechanisms. Ikiz et al. [31] studied chalcopyrite leaching in hypochlorite solutions and the pH was adjusted by HCl. They fitted the leaching data with fluid film diffusion, surface chemical reaction and diffusion through the product layer (Table 6) and concluded that the leaching process was controlled by diffusion through an elemental S layer. As suggested by Dreisinger and Abed [29], when a new layer is produced, the resistance of fluid transport through the new layer will be much greater than through the fluid film surrounding the particles. Hence, during chalcopyrite leaching, the influence of the fluid film can be ignored because the formation of a surface layer is inevitable. The particle size of chalcopyrite in Ikiz et al. [31] experiments was found to be the most influential factor in terms of rate of leaching and the Ea for Cu dissolution was calculated as 19.88 kJ mol-1, which further confirmed the mechanism of diffusion through the produced layer. Minerals and Materials Science & Technology Mawson Institute, University of South Australia

37

ACCEPTED MANUSCRIPT In reality, the diffusion between the solid and the leach media is more complex than it is usually considered to be and can involve the following consecutive steps [155]: (I)

Diffusion of the fluid reactants through the fluid film surrounding the solid

PT

particle; Diffusion of the fluid through the porous solid layer;

(III)

Adsorption of the fluid products from the solid reaction surface;

(IV)

Chemical reaction with the solid surface;

(V)

Desorption of the fluid reactants from the solid reaction surface;

(VI)

Diffusion of the formation away from the reaction surface through the porous

SC

RI

(II)

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solid media and then through the fluid film surrounding the solid. If any of the above steps is much slower than others then that step then becomes rate-

MA

determining.

Aydogan et al. [46] found that the kinetics of dissolution of chalcopyrite in acidic potassium dichromate (K2Cr2O7) at 50 - 97 °C can be fitted by the shrinking core model with the

D

diffusion through a porous S0 layer considered to be the rate controlling. The Ea for Cu

TE

leaching was calculated to be 24 kJ mol-1. Through investigation of the effects of concentrations of H2SO4 and K2Cr2O7, stirring speed, temperature and particle size, they

AC CE P

derived the following kinetic model (Equation 41): +' A  1   B  1  B



 



2C( .) A 3D #(  *

41

where X is the fraction oxidised, MB is the molecular weight of chalcopyrite (kg mol-1), D is the diffusion coefficient in the porous product layer (m2 s-1), CA is the concentration of the dissolved lixiviant A in the bulk of the solution (mol m-3), ρB is the density of chalcopyrite (kg m-3), a is the stoichiometric coefficient of the reagent, r0 is the initial radius of chalcopyrite (m), t is the leaching time (s), kd is the rate constant and is directly proportional to 1/r20 . By considering the effect of agitation, temperature, particle size, initial acid concentration Dreisinger and Abed [29] proposed one leaching rate law for sulfate media (Equation 42) and one for chloride media (Equation 43). 1  31  B(

 

 21  B(  ,  $ % exp I33,880>*=K A +

Minerals and Materials Science & Technology Mawson Institute, University of South Australia

42

38

ACCEPTED MANUSCRIPT 1  31  B(

 

 21  B(  ,  $ % exp I22,423>*=K A +"

43

where Xb is the fractional conversion with respect to chalcopyrite particles, k0’and k0’’ are

PT

intrinsic parabolic leaching rate constants for sulfate and chloride solutions, respectively, r is the initial particle radius, [H+] is hydrogen ion or acid concentration, R is the gas constant, T

RI

is the temperature, t is time.

SC

The activation energies were found to be 33.9 and 22.4 kJ mol-1 for sulfate and chloride media respectively. In these two parabolic models, the rate controlling step was proposed to

NU

be associated with the diffusion of H+ through the new formed surface layer. The leaching kinetics can be enhanced by increasing temperature (<65 °C) and the concentration of H+,

MA

and by decreasing particle size. The leach rate in chloride solution was found to be faster than that in the sulfate media, which is in agreement with other authors [141]. However, using the same shrinking core model Adebayo et al. [105] established a rate law for +, A  1  1  B



+ - 

TE

 

D

chalcopyrite leaching with H2O2 in H2SO4 (Equation 44). . ,

A

44

AC CE P

There is a linear relationship between the rate constant and the reciprocal of the particle size, and the activation energy was found to be 39 kJ mol-1 implying that the controlling step is a chemical surface chemical reaction. In addition, the dissolution rate increased with increasing concentrations of H2O2 and H2SO4 and their reaction orders were 1.45 and 0.77, respectively. Similar conclusions were made by Antonijevic and Bogdanovic [32] when they carried out leaching with H2O2 in H2SO4. Compared with the negative effect of stirring speed on the dissolution rate [105], the stirring speed was found to have no effect on dissolution rate, suggesting it was not a liquid-phase diffusion controlled process. The reaction activation energy was found to be 60 kJ mol-1 and the reaction has an order of 0.3 with respect to the concentration of H2SO4. Based on the method of reduced half time of reaction [156], Sokic et al. [132] leached chalcopyrite in H2SO4 with NaNO3 from 70 to 90 °C and found that the leaching rate could be enhanced by increasing temperature, concentrations of H2SO4 with NaNO3, or decreased stirring speed or particle size. Their data fit well to a mixed control model: +A    ln1  B

Minerals and Materials Science & Technology Mawson Institute, University of South Australia

45 39

ACCEPTED MANUSCRIPT The activation energy derived from this model was 83 kJ mol-1. It was suggested the initial leaching step is determined by the surface chemical reaction. But with increasing temperature and Cu extraction, the leaching rate decreases gradually and the leach rate is controlled by

PT

surface reaction and diffusion. When Cu extraction rate is over 60%, the leaching process is determined by diffusion through the S0 layer formed and the leaching rate is very slow at this

RI

stage. Similar results have also been found by Saxena and Mandre [30] who also reported mixed control kinetics in processing of Cu dissolution with ferric chloride solutions between

SC

30 and 90 °C and found Ea was in the range of 15 - 28 kJ mol-1.

In contrast to the literature discussed above Kaplun et al. [131] derived Ea for both Fe and Cu.

NU

They suggested that the increased activities of Fe3+ (initially added) and H+ could stimulate the Cu extraction at the beginning of leaching. The Ea derived from the Arrhenius plot is

MA

21±5 kJ mol-1 for Cu within the first 10 h of leaching, but increases to 83 ± 10 kJ mol-1 after 10 h while for Fe, it is 76±10 kJ mol-1 throughout. They also suggested the leaching rate for Cu is firstly diffusion controlled (that is controlled by mass transport of a dissolved reactant

D

from the bulk solution to the solid-solution interface or product ions diffusion from the

TE

interface to the bulk solution, <10 h) in presence of Fe3+ and then becomes chemical reaction controlled (>10 h). On the other hand, Fe leaching is chemical reaction-controlled with or

AC CE P

without the presence of added Fe3+ [131]. Without the addition of Fe3+, the activation energies for Cu and Fe are 80 ±10 kJ mol-1 and 84 ± 10 kJ mol-1, respectively, implying chemical reaction controlled processes for both Cu and Fe throughout the entire leach. Dong et al. [138] studied chalcopyrite leaching by using a Bronsted acidic ionic liquid ([bmim]HSO4) from 40 to 90 °C. They found that copper extraction is very slow at lower temperatures (<70 °C). However, the rate increases dramatically from 70 to 90 °C, due to the high Ea (69.4 kJ mol-1). Two mechanisms were proposed to test the dissolution kinetics. It was proposed the process is firstly controlled by diffusion of oxidant (DO or Fe3+) through a product layer of S0: +' A  1  B  1  B

 

 



/!0 # . ,

A

46

Then the process is controlled by interfacial chemical reaction: +, A  1  1  B

 

+

 !$  #  !0 #

. ,

Minerals and Materials Science & Technology Mawson Institute, University of South Australia

A

47

40

ACCEPTED MANUSCRIPT where X is the fraction reacted at time t, DS is the effective diffusion coefficient of oxidant through the solid layer formed, [OX] is oxidant concentration in the bulk solution, [H+]S and [OX]S are H+ and oxidant concentrations on the unleached surface, respectively, and m and n

PT

are the reaction orders. Padilla et al. [41] studied chalcopyrite pressure leaching (5 - 12 atm) at 125-150 °C in a H2SO4 - O2 system and proposed a modified shrinking core model:  

 +1

2 ,

A

48

RI

+A  1  1  0.45B

SC

where X is the fraction reacted, k’ is a kinetic constant depending on the temperature, m is the reaction order with respect to the partial pressure of O2, r0 is the radius of chalcopyrite, k is

NU

the apparent kinetic constant and t is the leaching time. The activation energy was found to be 93.5 kJ mol-1, suggesting a surface chemical reaction was the rate determining step. In

MA

addition, the rate was proposed to be first order with respect to oxygen partial pressure. Hackl et al. [8] suggested a mixed diffusion and chemical reaction model to illustrate the oxidation, leaching and the passivation in sulfate solution under O2 at 110 °C:

D

 
     3 
     M    2  M  

(fastest)

TE

M N ,   M ≈ 1

AC CE P

   3  4    5 
  O   1  M

   2O  1  M     5 
  1    O    2
  21    O  

49

(slower) 50 (slowest) 51

In the first reaction (Equation 59), Fe is leached out preferentially as a result of diffusion through a passivating layer of Cu1-xFe1-yS2, and then this intermediate sulfide is decomposed to form Cu1-x-zS2, which is alternatively expressed as CuSn which could impede further leaching. Hackl et al. [8] view was that the latter chemical reaction control is determined by the leaching rate of CuSn. The S0 layer would be sufficiently porous to not inhibit reactant/product diffusion. The decomposition rate of CuSn increases with increasing temperature, and if the temperature is greater than 200 °C, oxidation and leaching will no longer be impeded by CuSn. The formation of S0 decreases at high temperatures as the Sn2polymeric species may fragment and oxidise to sulfate before S8 rings can be produced.

5. Chalcopyrite Bioleaching

Minerals and Materials Science & Technology Mawson Institute, University of South Australia

41

ACCEPTED MANUSCRIPT Commercial application of bioleaching of ores began with copper bioleaching at the Kennecott Copper Bingham Mine in the 1950s and has since become a viable low cost option for extracting metals, especially from low grade ores [157]. Since this time, research and

PT

development into biohydrometallurgical applications has enabled improved economically and environmentally friendly designs, from uncontrolled copper dump leaching to bioheap and

RI

stirred tank bioleaching adopted in current operations [158]. Commercial bioleaching for Cu has so far only been confined to copper oxide and secondary copper sulfides, such as

SC

chalcocite, with parallel applications for chalcopyrite being hampered by its refractory nature. Bioleaching of chalcopyrite has so far only been confined to laboratory based bench scale

NU

studies and small-to-large scale pilot plant tests [5]. Bioleaching requires simple equipment and simple operational procedures, and has low setup and operation costs; however, it can

MA

return slow leach rates, restricting its range of commercial applications [5]. 5.1 Role of Microbes

D

A number of different microbes have the ability to oxidise or assist in the oxidation of

TE

chalcopyrite. Most belong to the phylogenetic branch of archaebacteria and bacteria, and comprehensive listing of the different autotrophic and heterotrophic bacteria are provided

AC CE P

elsewhere [4, 159]. However, common microbes extracted from acidic mine sites consist of Acidithiobacillus, Leptospirillum, Sulfolobus and Sulfobacillus bacteria [160]. Laboratory based studies are usually conducted using single strains or mixed strains of bacteria whilst commercial application uses a mixed (consortium) of bacterial strains. Mixed strains, under optimum growth conditions, usually yield higher leach rates, as compared to single strains, as different bacterial strains are able to adapt to the changing leach conditions throughout the leach process [161]. Acres et al. [69] used a mixed culture of microbes from an acid mine drainage site and found that Leptospirillum was dominant during the initial leach stages whilst later leach stages resulted in stable concentrations of Leptospirillum and Acidthiobacillus. Detailed bioinformatics study to map the gene content and metabolic potential of different bacterial strains can lead to genetic manipulation for enhanced physiology of certain strains that allows more efficient bioleaching in pre-set conditions [162]. The microbes used in bioleaching are able to survive in extreme conditions of temperature (mesophiles 30-40 °C, moderate thermophiles ≈50 °C and extreme thermophiles > 65 °C), pH and metal concentrations. Such microbes are usually able to oxidise Fe and/or S and the following equations indicate the roles that may be played by microbes in chalcopyrite oxidation: Minerals and Materials Science & Technology Mawson Institute, University of South Australia

42

ACCEPTED MANUSCRIPT  
  4  
6        2
  2 

4    4  
6 PQQQQQQQQQQQQQQQQR 4    2  78 '9 (:8,

2
  3
6  2  PQQQQQQQQQQQQQQQR 2  2

53 54 55

RI

 
  4       5    2


PT

; '9 (:8,

52

These equations indicate that microbes assist in leaching chalcopyrite through generation of

SC

Fe3+ and oxidation of S, which has been indicated in passivation of chalcopyrite surfaces

NU

during leaching [163].

It is generally accepted that microbes mediate bioleaching through contact, indirect and cooperative mechanisms [160, 164]. The indirect mechanism occurs through planktonic cells

MA

which make Fe3+ available in solution for mineral surface oxidation. Attachment of bacterial cells onto the mineral surface is essential for the contact and

D

cooperative mechanisms. It is believed that bacterial contact occurs through secretion of

TE

extracellular polymeric substance (EPS) by the bacteria [160, 164, 165]. The EPS has been shown to consist mainly of proteins, lipids, sugars and Fe3+ [165]. Recently, Bobadilla et al.

AC CE P

[166] identified a new single lipoprotein named licanantase in the EPS of Acidithiobacillus thiooxidans and Acidithiobacillus ferrooxidans, and further showed that licanantase exerted an influence in increasing the rate of chalcopyrite bioleaching. The EPS also increases the reaction sphere of the bacteria and is the site of mineral dissolution. It is believed that the contact mechanism proceeds through an enzymatic oxidation involving Fe3+ which is a constituent of the EPS [160]. Bacterial attachment and attack is possibly site specific with cracks and defects on the surface attacked first [167]. The contacted bacteria mediate the Fe and S oxidation as shown in Fig. 4a. Rohwerder et al. [164] suggests that the Fe3+ (within EPS) attack extracts electrons from the mineral surface which reduces molecular oxygen through a complex redox chain within bacterial membranes. According to Rodríguez et al. [160] there may be two modes of bacterial attachment, reversible and irreversible (Fig. 4b). The reversible interaction results in the formation of a metastable complex between the mineral and bacterial surfaces and can yield a free cell again, while the irreversible contact form upon secretion of EPS. Koshini et al. [168] suggested through correlation of equilibrium adsorption data from bioleach experiments and Langmuir isotherm measurements that

Minerals and Materials Science & Technology Mawson Institute, University of South Australia

43

ACCEPTED MANUSCRIPT attached cells play a more dominant role in the leaching process compared to the indirect mechanism. The cooperative mechanism (Fig. 4.c) is established between the contacted and planktonic

PT

cells, when the contacted cells supply Fe2+ and S (or other S products) to the planktonic cells for their energy source. The net result is an efficient supply of oxidants (Fe3+) for oxidation

RI

and removal of surface S products. Gautier et al. [169] found that when there was a

SC

cooperative relationship between attached and planktonic cells, the leaching efficiency of chalcopyrite was much improved. Similar results were also found by Zhang et al. [170]. Rodríguez et al. [160] further showed that bacterial attachment was critical for high rates of

NU

leaching and that bacterial leaching consisted of 3 stages: the first involved extensive bacterial attachment; the second stage involved a decrease in bacterial attachment due to

MA

surface saturation; whilst in the third stage a balance between free and attached cells (cooperative mechanism) is reached. Chen et al. [161] used Acidithiobacillus ferrooxidants and Acidithiobacillus caldus to leach chalcopyrite. The surface properties of chalcopyrite

D

were charactered by using zeta-potential, FT-IR, and contact angle measurements. The data

TE

supported the proposal that during the initial stage, the direct mechanism played a dominant role since after the treatment from bacteria, the zeta-potential studies showed that the iso-

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electric point (IEP) of chalcopyrite moved towards the IEP of pure cells. 5.2 Bacterial Health and Adaptation Bacterial health in a bioleaching system is central to an efficient leaching process. Bacterial growth is affected by environmental, biological and physico-chemical factors. Rapid and healthy communities occur under optimum conditions of humidity, pH, Eh, temperature, energy and nutrient source, and the absence of inhibitors. Furthermore, adaptation to the leaching environment by bacterial communities is needed for an optimum leaching outcome. Bacterial adaptation can be a lengthy process where bacterial communities are progressively pre-adapted to the leaching environment such as temperature, mineral pulp densities, pH and redox potential prior to leaching. Xia et al. [171] showed that mineral surface attachment, shearing stress, and Cu tolerance were affected by the degree of bacterial adaptation to the leaching environment. They used pre-adapted (to the mineral leaching environment) and unadapted Acidithiobacillus ferrooxidans to leach chalcopyrite and found that leaching with adapted bacteria produced better leach rates. This effect was attributed to the lower lag period of adapted bacteria and their ability to form better and faster adsorption bonds with the Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT mineral surface, and better ability to withstand shearing stress and higher Cu tolerance. Recently, Karimi et al. [172] also showed that adapted At. ferrooxidans had a better tolerance to higher temperatures and produced better leaching rates.

PT

5.3 Effect of Microbes on Chalcopyrite Dissolution

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While the effect of using microbes for leaching minerals such as chalcocite, covellite and pyrite has been shown to have definite beneficial effect, this has been a subject of some

SC

discussion in the past with regard to chalcopyrite. Third et al. [173] conducted a series of experiments with and without bacteria and concluded that the redox potential (Eh) is more

NU

significant for determining chalcopyrite leach rates than the number or activity of bacteria. A mixed culture of mesophilic bacteria was used to conduct a series of leaching tests with

MA

different initial concentrations of bacteria (Third et al. [173]. The bacteria were inoculated into the leach solution either with or without washing. It should be noted that the bacteria was grown in a high Fe3+ broth and inoculation without washing also introduced Fe3+ into the

D

leach solution. For experiments conducted without washing a high bacterial concentration

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(50 % v/v) was found to be detrimental to the leaching rate, while all experiments conducted with washing resulted in a rapid rate of leaching. This suggested that the effect was due to a

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high initial Fe3+ concentration (and hence a high Eh) which caused poor leach rates in unwashed experiments. Third et al. [173] also conducted a series of chemical experiments under the same conditions as the bioleach experiments, but with different concentrations of either Fe2+ or Fe3+. Similar conclusions were reached as for the bioleach experiments, that high initial Fe3+ and Eh were detrimental to chalcopyrite leach rates, whilst Fe2+ containing leach media returned faster rates. Maximum bioleaching and chemical leaching rates were similar. Third et al. [173] further concluded that while Eh played a more significant role in determining leach rates, bacteria were needed for continued leaching through regeneration of optimal Fe3+ concentrations. If the concentration of Fe3+ being produced exceeds the concentration of Fe3+ being consumed, then bacterial use can become detrimental. Third et al. [173] suggests that the use of minimal bacterial activity coupled with Fe3+ (or redox) control would be the most efficient leaching approach. Watling [5] and Third et al. [173] also note that previous studies reporting a beneficial effect of either S or Fe oxidising bacteria on chalcopyrite leach rates may primarily be due to low redox potential, which in some studies was not measured or controlled. Gómez et al. [167] reported a definite beneficial effect of using bacteria for chalcopyrite leaching from use of a Minerals and Materials Science & Technology Mawson Institute, University of South Australia

45

ACCEPTED MANUSCRIPT thermophilic archaebacteria, Sulfolobus rivotincti. They also noted a detrimental effect of high initial Fe3+ concentration and high initial Eh in both chemical and bioleaching experiments. Furthermore, Gómez et al. [167] also noted that Fe2+ concentrations can have

PT

detrimental effect. They found that above 750 mV (SHE) bacterial growth was poor if the Fe2+ concentrations were low. They also suggested that if initial Fe2+ concentrations were

RI

high then bacteria may only oxidise solution Fe2+ and ignore chalcopyrite as an energy source. Leaching in such cases would proceed through the indirect mechanism. Recently, Ahmadi et

SC

al. [7] reported, through a series of bioleaching (using moderate thermophiles) and chemical leaching experiments conducted at pH 1.5, temperature 50 °C, and controlled Eh that

NU

electrochemical bioleaching in the Eh range of 600 to 630 mV (SHE) had an 80 % Cu recovery within 10 days which was 3.9 times higher than electrochemical leaching under

MA

same conditions. A number of reviews have also shown that there is a general acceptance that bacterial leaching of chalcopyrite does have a beneficial effect [4, 5, 157]. It has generally been found that high temperature and use of moderate thermophiles and thermophiles are

D

beneficial for chalcopyrite leaching [157, 174].

TE

Brierley and Brierley [157] further suggest that for cases where less acid (H2SO4) is needed to reduce downstream neutralisation costs, the bacteria should be prevented from contacting

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the mineral surface. Under this condition S oxidation produces elemental S and not H2SO4. Similar observations have also been made by Olsen et al. [158]; who suggest that mesophiles are not efficient at removing surface passivation and thermophiles and extreme fine grinding is needed to achieve optimum results. Sample mineralogy can also have an effect on the rate of chemical and bio-leaching as acid consumption by the gangue minerals can add to the cost of sulfuric acid. Depending on the mineralogy, reaction conditions can be optimised to generate more acid from within the ore; for example greater pyrite oxidation would produce more acid for the system. However, there is currently a lack of sufficient published data on ‘real’ ore samples relating ore mineralogy and solution leach chemistry. Petersen and Dixon [175] conducted a series of experiments using mesophiles, moderate thermophiles and extreme thermophiles to leach a copper-gold concentrate from Telfer mine in Western Australia. It was found that high temperature and low redox potential promoted chalcopyrite oxidation while low temperature and high redox potential promoted pyrite oxidation. Secondary copper sulfides (chalcocite and covellite) were found to leach preferentially under all conditions. It appears that to control the amount of chalcopyrite dissolution and the amount of pyrite dissolution (if needed for acidity in the system) a Minerals and Materials Science & Technology Mawson Institute, University of South Australia

46

ACCEPTED MANUSCRIPT balance between redox potential, temperature and a suitable consortium of microbes that grow and perform well in the set temperature range is needed. The selection of temperature (and hence microbes) and redox potential will be heavily dependent on the sample

PT

mineralogy. Bacterial activity can therefore be beneficial for chalcopyrite leaching if the corresponding

RI

electrochemical conditions are optimal. Key factors for optimum bioleaching of chalcopyrite

SC

are low initial Fe3+, low redox potential (best if controlled), sufficient initial Fe2+, use of high temperature and moderate to extreme thermophiles, fine grinding [176, 177] and use of

5.4 Surface Products and Passivation

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catalysts such as silver [6, 178-180] which is discussed later in the section 5.7.

MA

Over the past decades, many authors have addressed the issue of the slow kinetics of bioleaching chalcopyrite and the passivation layer formed in the process. There are various products that are suggested by different researchers for different leaching conditions,

D

including jarosite, chalcocite, covellite, elemental S [4, 5, 172, 181, 182]. Of these elemental

TE

S and jarosite are particularly important. These two products have been suggested to significantly influence the leaching rate by passivating the chalcopyrite surface. Chalcopyrite

AC CE P

surface oxidation leads to the formation of an intermediate (in terms of oxidation state) sulfide layer on the surface, which through progressive reaction forms elemental S [183]. Elemental S on chalcopyrite surfaces can also form through reduction, especially in low redox (low [Fe3+]/[Fe2+]) conditions [184]. As shown in Equations 56 to 57 chalcopyrite reduction may lead to the formation of Cu2S and H2S which are readily oxidised to elemental S.

 
        2    
  2        
  4    2   4    


   2    2    2  


56 57 58

The elemental S coating on the chalcopyrite surface may hinder the leaching process by inhibiting the flow of electrons and oxidants to and from the chalcopyrite surface [4]. Bacterial action leads to S oxidation which tends to produce sulfate. Bacterial action also produces more Fe3+(aq) and tends to increase solution Eh. Under such conditions with high solution sulfate and monovalent alkali cations, jarosite precipitation occurs (Equation 59) Minerals and Materials Science & Technology Mawson Institute, University of South Australia

47

ACCEPTED MANUSCRIPT which also coats the surface of chalcopyrite. Jarosite precipitation is pH dependent and is predominant in the range of pH 1.9 – 2.2 [4]. Jarosite precipitation has been suggested to reduce the interaction between the chalcopyrite surface and oxidants/bacteria.


59

RI

where M = K+, Na+, H3O+ or NH4+

 6

PT

3    2  6   C  C    

Sandström et al. [185] conducted a series of chemical and bioleach experiments and

SC

concluded that S0 and jarosite form under different electrochemical conditions and it is the jarosite that passivates the surface. Chemical and bioleaching experiments conducted at the

NU

reduced Eh of 620 mV (SHE) were found to have faster kinetics and better yield than leaching conducted at the greater Eh of 800 mV (SHE). Using XPS and XRD, they found large

MA

amounts of jarosite on leach residues from high Eh leaches, which according to them passivated the chalcopyrite surface. Large amounts of elemental S was found on leach residues from low Eh chemical leach experiments. Bioleach samples had relatively smaller

D

amounts of elemental S as bioleach allowed more complete oxidation of surface S to sulfate.

TE

Since large amounts of elemental S was found on leach residues from the fastest leach experiments (low Eh), elemental S was not considered as a passivating species.

AC CE P

Using XRD, Raman spectrometry and XANES, Zhu et al. [186] also concluded that jarosite appeared to hinder the bioleaching of chalcopyrite while elemental S had no effect. Zhu et al. [186] also found chalcocite and covellite (although these were not identified by XRD) on leach residues using S 1s XANES fitting and concluded that the oxidation mechanism proceeded through these secondary sulfide intermediates and that elemental S may partially be derived from these intermediates. Intermediate sulfides and S0 is suggested to hinder the dissolution process only at low temperatures (30 °C), while at higher temperatures (68 °C), there is no hindrance due to increased dissolution [160]. Recent studies by Yu et al. [187] also found that below 650 mV (SHE) Eh, jarosite effects on bioleaching were negligible. Jarosite is readily produced in the presence of high concentrations of Fe3+ and SO42- in the bioleaching system containing cations like K+, Na+, NH4+ or H3O+, which are common nutritional species for microorganism growth. Among the jarosite species, potassium jarosite is the most stable and generally the first jarosite formed in the bioleaching process. The elevated temperature of over 65 °C and pH 1.7 - 2.7 can boost the formation of jarosite while

Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT acidophilic microorganisms catalyse the formation of jarosite by oxidising of Fe2+ to Fe3+ [102]. Experiments with contacted and non-contacted Acidianus manzaensis by Zeng et al. [163]

PT

found that elemental S formation on chalcopyrite surfaces and possible surface passivation (by S0) occurred only in experiments where bacteria were prevented from contacting the

RI

mineral surface. In this case leaching was mediated through the indirect mechanism with chemical Fe3+ oxidation. Bevilaqua et al. [188] showed that the pure culture of

SC

Acidithiobacillus thiooxidans (At.t) did not oxidise CuFeS2 while with added S0, the Cu dissolution rate was similar to that of chemical control, but without any new species formed

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even after 56 days in contact with the solution. It can be regarded that Acidithiobacillus thiooxidans is not suitable for bioleaching CuFeS2. But for experiments with

MA

Acidithiobacillus ferrooxidans, S0 and jarosite were found at the surface of chalcopyrite. Recently, Zeng et al. [163, 165] showed that a mixed culture of moderate thermophiles have

D

potential for application in the chalcopyrite bioleaching process. These authors found that at

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the end of bioleaching, the results of SEM/EDX of the chalcopyrite electrodes revealed the extracellular polymeric substances (EPS) and jarosite were the major products within the

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passivation layer to hinder the further leaching on the surface. Specifically, by using cyclic voltammetry analysis, the anodic and cathodic current signals were found to be reduced whilst the anodic peak moved gradually from low potential to high potential. Throughout the whole bioleaching process, the intermediate products of CuxS (1 < x < 2) could be found [165]. In a stirred tank reactor with a pulp density of 6 wt.%, the copper recovery was 84.7% after 24 days. Meanwhile, the leaching rate increased with increasing total Fe concentration in the solution [138].

Furthermore, Raman spectroscopy together with Fourier transform infrared (FT-IR) and XRD were used to analyse the secondary minerals produced in the process of bioleaching chalcopyrite with Acidithiobacillus ferrooxidants [189]. Potassium jarosite was first found prior to the formation of ammonio-jarosite, which accounted for the slow Cu leaching rate. Following the formation of potassium jarosite, CuS was detected as a second product of bioleaching, which was stable and remained even after over two months of leaching. Subsequent to CuS formation, elemental S was also detected in the leached solid. All three species were thought to contribute to the passivation of chalcopyrite bioleaching.

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ACCEPTED MANUSCRIPT It has been suggested that use of high temperature and moderate thermophile or thermophile bacteria can overcome the passivating effect of jarosite. D’Hugues et al [190] also did not find any evidence of jarosite hindrance to leaching chalcopyrite with extreme thermophiles at

PT

78 °C. Under anoxic conditions, moderate thermophiles are able to use Fe3+ in the jarosite surface precipitate as a terminal electron acceptor instead of oxygen [191]. However, in

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experiments conducted by Stott et al. [191], leaching of jarosite coated chalcopyrite samples with moderate thermophiles did not improve leaching rates. Leach residue analysis showed

SC

that although extensive bio-reduction of jarosite occurred, it was not completely removed. It was concluded that even a thin coating of jarosite was sufficient to hinder the dissolution

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process. The authors suggested that the best approach was to reduce or prevent jarosite formation rather than remediating it once it was formed during the leach. pH monitoring and

MA

control may also be required during leaching, as gangue minerals in the ore can consume acid and increase the pH to a level where significant jarosite precipitation would occur and lead to a significant decrease in soluble Fe concentration [192].

D

A popular bacteria, Sulfolobus metallicus, was tested in the bioleaching of pure chalcopyrite

TE

at 70 °C and pH 1.5 in shake flasks [193]. To prevent direct contact of bacteria with chalcopyrite the bacteria were restricted in a chamber packed by a 0.1 Millipore membrane.

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However, in this case, the Cu recovery was only half of that when there was bacterial contact with chalcopyrite. When there were no bacteria, the formation of bisulfite (HSO3-), bisulfate (HSO4-) and sulfate (SO42-) was detected. But thiosulfate (S2O32-) and sulfite (SO32-) were found on the surface when the bacteria were in contact with chalcopyrite. Furthermore, the intermediate products of HSO3- ,S2O32- ,SO32- can be oxidised by planktonic cells as part of the cooperative mechanism. Chalcocite and covellite have also been detected (using Leica phase contrast microscopy but not by XRD) by Ahmadi et al. [7] on the surface of bioleached chalcopyrite as have thiosulfate and polysulfide [5, 91]. 5.5 Reaction Kinetics The majority of papers published previously have mainly focussed on enhancing the bioleaching rate and obtaining improved recovery of metal. The following parameters and reagents have been shown to enhance bioleaching of chalcopyrite [5]: (i)

Higher temperature;

(ii)

Lower redox;

(iii)

Chloride improves the dissolution rate of chalcopyrite;

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ACCEPTED MANUSCRIPT (iv)

Silver ion catalysis improves chalcopyrite dissolution;

(v)

Moderate fine grinding particles with high surface area enhances rapid copper extraction.

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Temperature, pH, Eh or oxidation-reduction potential (ORP) and the activity of the bacteria are suggested to be the key factors controlling the copper extraction yield [194]. Vilcáz et al.

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[194] used Sulfolobus metallicus and Metallosphaera sedula to leach chalcopyrite. In the

SC

initial stage of bioleaching, the concentration of Fe3+ played the most important role for copper extraction, followed by temperature, but not pH. Then in the second stage, named the exponential stage, when the pH was small enough, the formation of jarosite was reduced and

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the temperature was pivotal to control of the leaching rate. When the protons were depleted with time, ORP became the most important factor as the concentration of Fe3+(aq) decreased

MA

[194]. Therefore, these three parameters can be controlled at different stages of bioleaching to enhance the Cu yield.

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The range of ORP for both optimised rate and yield of Cu dissolution is usually found to be

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around 600-650 mV (SHE) [150, 181, 195]. The redox potential can be controlled chemically by use of oxidising (e.g. H2O2) or reducing agents (e.g. sodium sulfite). It can also be

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controlled electrochemically by applying potential through an external source such as a working electrode [195]. Ahmadi et al. [7, 195] conducted a series of electrochemically controlled experiments using mesophiles and moderate thermophiles and found significant improvements (of up to 35 %) in the leach rate and yield, as compared to conventional bioleaching and chemical leaching approaches. An Eh range of 600 to 630 mV (SHE) was found to be the optimum range. Biological and analytical analysis of the leach residues (solution and solid) showed a four-fold increase in bacterial population and drastically reduced surface jarosite (as compared to conventional bioleaching and chemical leaching) as a result of an externally applied potential. Mineralogical analysis using a Leica phase contrast microscope (optical microscopy) also showed the presence of chalcocite and covellite on the surface. It is suggested that increased bacterial growth, decreased surface jarosite and electroreduction of the chalcopyrite surface to secondary sulfide minerals are the main reasons for enhancement of chalcopyrite leach rates and Cu yield. Using chemical redox control Gericke et al. [181] found that when using a thermophilic culture to leach chalcopyrite at 70 °C, more than 95 % of the Cu can be recovered without any redox control. However, redox control at 45 °C is needed for optimum Cu recovery at 620 mV (SHE). Studies conducted by Vilcáez et

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ACCEPTED MANUSCRIPT al. [194] also confirmed that redox control was needed for high leach rates with the optimum value close to 650 mV (SHE). A number of studies have also demonstrated that in addition to mixed cultures performing

PT

better than pure cultures, different mixes of microbes also yield better results than other mixes. Akcil et al. [196] showed that when leached under similar conditions, a pure culture of

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At. ferrooxidans performed better than pure cultures of Leptosppirillum ferrooxidans and At.

SC

thiooxidans. Mousavi et al. [197] found Sulfobacillus to perform better than At. ferrooxidans under their reaction conditions. Akcil et al. [196] suggest that for effective chalcopyrite oxidation, bacterial mixes consisting of Fe and S oxidising bacteria should be chosen to

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regenerate Fe3+ and to remove S0 passivating layers from the surfaces. In a series of experiments conducted by Akcil et al. [196], it was found that a mixture of At. ferrooxidans,

MA

L. ferrooxidans and At. thiooxidans performed better than a mixture of L. ferrooxidans and At. thiooxidans. It was further found that an increase in pulp density (1 – 5 % w/v) had detrimental effect on bioleaching activity of pure and mixed cultures. Wang et al. [198] also

D

showed pulp density to be one of the factors that can influence leach rates. Rubio and Frutos

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[199] showed that with 10 wt.% pulp density, it is possible to extract 94% of Cu within 10 days by using a mixed thermophilic culture. The yield decreases to about 80% in 14 days for

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the density of 20 wt.%. Using a similar approach, Dong et al. [138] found mixtures of moderate thermophiles, L. ferriphilum and At. Caldus, and Sulfobacillus sp. and Ferroplasma thermophilum perform better than all other mixtures they tested. Fu et al. [200] also found that a mixture of thermophilic bacteria, L. ferriphilum and At. Caldus, leached chalcopyrite more efficiently than the mesophilic mixture of At. ferrooxidans, and At. Thiooxidans. Hydraulic residence times (HRTs) when using a continuous column reactor can also have an effect on the chalcopyrite leach rates [201]. Liang et al. [201] showed that a residence time of 80 h was optimum for their setup. At other residence times there was competition between different microbes for Fe2+ oxidation which affected the leach rates. The moderate thermophiles in bioleaching systems with optimal living temperature of 45-52 °C are actively used by many researchers because compared with mesophilic conditions, increasing temperature increases reaction kinetics and activity [181]. On the other hand, compared with thermophiles, mesophiles can tolerate greater pulp densities and greater concentrations of catalysts such as silver [202]. Organics such as yeast extract have also been shown to positively influence bacterial growth and hence leach rates above 40 °C [7].

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ACCEPTED MANUSCRIPT 5.6 Bioleaching Mechanisms Based on key intermediates and products Sand et al. [203] and Schippers and Sand [204], proposed two separate oxidation mechanisms for metal sulfides. The mechanisms are based

PT

on the chemical oxidation approach where key oxidants are Fe3+ and/or protons (H+). The role of bacteria is to regenerate these oxidants and concentrate them at the mineral/water or

RI

mineral /bacterial cell (EPS) interface. The mechanisms are based on the thiosulfate or the

SC

polysulfide pathway. Due to the crystal and electronic structure, sulfides like molybdenite, tungstenite and pyrite are considered to follow the thiosulfate pathway as these sulfides can only be oxidised through an oxidative attack by Fe3+. Dissolution of sulfides such as galena,

NU

sphalerite and chalcopyrite can also occur through an H+ attack, hence are considered to follow the polysulfide mechanism. According to Sand et al. [203], the valence band of

MA

minerals such as molybdenite, tungstenite and pyrite arise from the orbitals of metal atoms only, hence the valence band of these minerals do not contribute to the bonding between metal and sulfur moieties. The metal to sulfur bonds of these minerals can be broken through

D

a series of oxidation steps with attack from Fe(III) hexahydrate. The valence band of minerals

TE

such as galena, sphalerite and chalcopyrite result from a contribution from both metallic and S orbitals, and are prone to both Fe3+ and H+ attack. Thiosulfate and polythionates are key

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intermediates in the thiosulfate mechanism while S0 is the key product of the polysulfide pathway. The final product, mediated through bacterial oxidation, of both these pathways is sulfate. The two mechanisms are further explained by Equations 60 to 64 and Fig. 5. Thiosulfate mechanism:

  6    3      7    6

   8    5   2  8    10

60 61

Polysulfide mechanism:

C  2  2

   2  C     2  2

   ) 2

   2    % %  2    2



 % %



 1.5       2

62 63 64

The bioleaching of chalcopyrite may also follow the chemical mechanism suggested by Hiroyoshi et al. [45] for chalcopyrite oxidation in the presence of Fe3+ [185]. This mechanism is based on reduction where the chalcopyrite is converted to a chalcocite-like product which

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ACCEPTED MANUSCRIPT is more easily oxidised than chalcopyrite. Under bioleaching conditions, Cu2S would be more easily converted to S0 by bacterial action.

PT

5.7 Effect of Silver and Activated Carbon The bioleaching of chalcopyrite has also been shown to be enhanced in the presence of

RI

metallic cations apart from Fe3+. Silver is one such catalytic metallic species that has been actively studied for enhancing bioleaching of chalcopyrite. Gómez et al. [202] found a 3-fold

SC

increase in Cu extraction due to silver added to a leach at 45 °C. The authors used mesophiles and moderate thermophiles as silver is toxic to thermophiles at high temperatures.

NU

Experiments conducted at 50 °C by Gómez et al. [202], showed poor bacterial growth, possibly due to silver toxicity, and there was a significant jarosite precipitation which

MA

neutralised any benefits from the added silver. At greater temperatures, the authors found that the amount of jarosite precipitation increased with increasing silver. Using SEM and Auger electron spectroscopy (AES), Gómez et al. [182] established that Ag+ ions formed silver

D

sulfide (Ag2S) on chalcopyrite surfaces; surface S of which (Ag2S) is easily oxidised by

 
  4ST  2ST 
      

AC CE P

5.15.

TE

bacteria. The Ag2S species formation and its oxidation would occur through Equations 65 and

 2  

ST
  2    2ST  


65 66

Under bioleaching conditions Ag2S oxidation would proceed due to the presence of Fe3+ generated by bacterial action. Direct bacterial action would take place oxidising the S0 and Fe2+ produced through this process. While silver does have a positive influence on the leaching rate, one has to question the economic implications of using silver as a catalyst. Silver does promote chalcopyrite dissolution at lower temperatures, which may counteract the cost of using silver. Yuehua et al. [179] and Johnson et al. [205] have further shown that using argentite; silver concentrate and chalcopyrite ore that has significant amounts of silver may also provide similar benefits as using silver metal or ions (Ag+). Hence, any chalcopyrite ore with silver or silver bearing minerals would be able to be leached at lower temperatures than other chalcopyrite concentrates. Activated carbon has also been found to increase the bioleaching rate of chalcopyrite [201, 206]. Nakazawa et al. [206] leached a chalcopyrite concentrate using Thiobacillus ferrooxidans and found that the leaching rate and yield increased with increasing Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT concentration and decreasing particle size of activated carbon. Liang et al. [201] used an extreme thermophile, Acidianus manzaensis, to leach chalcopyrite with different concentrations of activated carbon. They found that a concentration of 2 g dm-3 activated

PT

carbon was optimum, increasing the yield from 64 % to 95 %. FT-IR, XRD and XANES measurements of the leach residues by Liang et al. [201] found jarosite, elemental S and

RI

chalcocite to be present on the surface. Activated carbon is a good electrical conductor and is thought to accelerate leaching by forming a galvanic couple with chalcopyrite [206].

on the activated carbon particle (Equation 68).

NU

Anodic reaction:

SC

Chalcopyrite undergoes anodic dissolution (Equation 67) while the cathodic reactions occur

Cathodic reaction:

68

TE

6. Galvanic Interactions

67

D


6  4  4    

MA

 
        2
  4 

Galvanic interactions are caused by electrochemical interactions as a result of the differing

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rest potentials of semi-conductive minerals [207, 208] and/or grinding media during electrochemical leaching/bioleaching [209], flotation [207, 210] and acid mine drainage [211, 212].

When two types of sulfide minerals are, particularly in the case of leaching, in electrical contact where electron transfer occurs from the smaller rest potential material (anode) to the material with the greater rest potential (cathode), they can form a galvanic cell. On the cathodic side where the minerals have the greater rest potential the following reaction may occur:   4  4   2 

69

For an anodic sulfide with the smaller rest potential the following reaction may occur: C
  C  
  2 

70

Examples of the rest potentials for sulfide minerals are summarised in Table 7.

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ACCEPTED MANUSCRIPT Galvanic interactions enhance the oxidation of the anodic mineral(s) leading to an increase of metal ions in solution though the cathodic mineral was protected from oxidation [213]. It has been reported that when chalcopyrite is in contact with pyrite, the leach rate of chalcopyrite is

PT

significantly increased. On the other hand, the dissolution rate of chalcopyrite is remarkably decreased in presence of sulfide minerals with smaller rest potentials such as sphalerite and

RI

galena [15].

Where chalcopyrite is in contact with pyrite, Fe3+(aq) if present in solution, will be reduced to

SC

Fe2+(aq) when in contact with the cathodic mineral, i.e. most commonly the pyrite surface [123, 214]:

NU

       

71

MA

The chalcopyrite acts as the anode to this reaction to produce the e-:  
        2
  4 

72

D

The S0 remains on the chalcopyrite surface. This galvanic interaction mechanism is shown

TE

schematically in Fig. 6.

As reviewed previously the passivation layer formed on chalcopyrite has variously been

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proposed to consist of one of, or a combination of, metal-deficient sulfides, S0, Sn2- and jarosites [27]. In the ferric sulfate leaching of chalcopyrite, Fe is selectively leached over Cu, in the absence of pyrite, with the consequent formation of a sulfur enriched copper sulfide. When pyrite was added, it is proposed that this provides a catalytic surface for the reduction of Fe3+(aq). When porous elemental sulfur forms on the chalcopyrite surface channels are available for the migration of Cu and Fe ions from the surface to the solution [126]. The galvanic current generated between pyrite and chalcopyrite has been measured by Chmielewski and Kaleta [215]. A sulfide electrode was prepared from rectangular sections (0.5 – 1 cm3) of chalcopyrite and pyrite joined by copper wire and conductive adhesive. The wired-sulfide was mounted in resin and polished. The measurement was carried out in solution (50 g H2SO4/dm3) which was either deoxygenated with argon (non-oxidative conditions) or saturated with O2 in the presence of Fe3+(aq) (oxidative leaching conditions) at temperatures of 25, 50, 70, and 90 °C. The addition of Fe3+(aq) to the solution dramatically increased the galvanic current between the sulfide electrodes. The increased potential

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ACCEPTED MANUSCRIPT difference was very stable due to the regeneration of Fe3+(aq) from Fe2+(aq) in the presence of dissolved O2. Galvanic interactions have been studied with respect to chalcopyrite leaching in relatively

PT

dilute acid solution such as pH 2.5 [216]. The dissolution rate of chalcopyrite was compared between chalcopyrite and chalcopyrite/pyrite systems. 0.3 g of solid (+45 -150 µm) was

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placed in 50 cm3 of solution maintained at pH 2.5 using 0.1 M HCl at 25 °C for a maximum of 3 h. The solution was constantly stirred. The chalcopyrite single mineral system required

SC

0.25 dm3 of 0.1 M HCl solution to maintain the pH at 2.5 while the mixed chalcopyrite/pyrite system required 0.40 dm3 of 0.1 M HCl indicating greater proton consumption due to the

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galvanic interaction. The dissolution rate of chalcopyrite was enhanced by factors of up to 18 times in the presence of pyrite. The leach period was terminated at 3 h, presumably prior to

MA

complete leaching and therefore the longer term effects of chalcopyrite-pyrite interactions were not examined.

D

When galena was mixed with either pyrite or chalcopyrite, the rest potential difference

TE

measured was greater than 0.3 V (SHE). Self-heating of these particular sulfide mixtures can occur in the presence of O2 and moisture due to galvanic interaction and can result in the

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ignition of fires due to the production of hydrogen sulfide by bacteria or intermediate reactions during the galvanic interaction [217]. The mixture of pyrite and pyrrhotite also has a high rest-potential difference causing self-heating. Self-hearint is particularly apparent for fine particle sizes [218].

When pyrite is present with arsenopyrite, pyrite is protected due to galvanic interaction. Arsenopyrite acts as the anode while pyrite acts as the cathode in the galvanic cell [219]. When galena (anode) forms a galvanic couple with sphalerite (cathode), sphalerite is also protected from oxidation [220]. 6.1 Galvanox Galvanox is the name given to a patented process based on the galvanic interaction between chalcopyrite and pyrite in ferric sulfate media [214]. The basis of the patent is described in [221]. 1.500 dm3 of solution (45 – 90 g H2SO4 dm-3) was stirred with a single 6 bladed impeller (1,200 rpm) in a reactor vessel. O2(g) was supplied through a gas flow controller and the temperature was varied from 60 to 80 °C. 50 g of chalcopyrite was leached with various masses (0 – 150 g) of pyrite all ground to -75 µm and the solution potential was varied from Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT 622 to 682 mV (SHE). The leach was completed within 24 h under the optimum conditions of: a) Pyrite-to-chalcopyrite ratio: 2:1 to 4:1;

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b) Acid concentration: stoichiometric + modest excess; c) Solution potential: > 637 mV (SHE) – preferably 667 mV (SHE) or higher;

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d) Temperature: > 70 °C – preferably 80 °C.

The patent [221] proposed a wider range of application conditions: Solution potential: 547 mV – 717 mV;

•

Chalcopyrite-to-pyrite ratio: 1:1 – 1:10;

•

Temperature: 50 °C – melting temperature of sulfur (110 - 120 °C);

•

Atmosphere of oxygen-containing gas, such as air, oxygen-enriched air;

•

P80 particle size: 38 – 210 µm;

•

At least 2 moles of H2SO4 per 1 mole of Cu recovered from chalcopyrite.

MA

NU

SC

•

As noted in the back scattered electron images (Fig. 7), the chalcopyrite particles were coated

D

with porous S0 (non-conductive) which provide channels to the solution, hence the

TE

chalcopyrite was not directly in contact with pyrite. It is possible that the high sulfuric acid concentration might aid in ionic transfer of electrons between the pyrite and chalcopyrite that

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are not in contract. In the presence of pyrite, it has been found that the current density significantly decreases as copper is extracted, which occurs within a shorter period of time than in the absence of pyrite. Reduced reactivity of the chalcopyrite was observed when pyrite was absent [222].

Koleini et al. [150] found stirring speed and initial acid concentration do not affect the copper recovery. However the pyrite/chalcopyrite ratio, solution potential and temperature have significant influences on recovery. The optimum conditions found were a stirring speed of 1,150 rpm, pyrite/chalcopyrite ratio of 4, solution potential 607 mV (SHE), temperature 85 °C and initial H2SO4 concentration of 45 g/L achieved more than 95 % copper recovery within 24 h. A passivation layer was formed on the chalcopyrite surface in the absence of pyrite, while the layer had no effect on the leaching in the presence of pyrite. 6.2 Silver Addition When the Galvanox process was repeated with pyrite (pyrite-to-chalcopyrite ratio of 6:1) pretreated with silver nitrate (100 ppm Ag on pyrite), the leach was completed within 10 h [123].

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ACCEPTED MANUSCRIPT In this experiment, 1.500 dm3 of solution was prepared with H2SO4 corresponding to a 150 % stoichiometric concentration based on the equation: 73

PT

 
    2        2   2


When a 200 ppm of silver against pyrite was added, leaching is complete in about 5 h.

RI

In order to investigate the role of silver, galena coated with silver was added to the Galvanox

SC

process along with silver-enhanced pyrite [223]. The silver ions were more attracted to galena than pyrite, however the presence of galena was found not to be beneficial in this process. UV
  2ST  ST 
  UV 

NU

There are two possible reactions to be considered;

MA

2UV
  10ST  4   2UV  8ST
    8

74 75

The metallic Ag0 does not enhance the catalytic effect. Azizi et al. [224] reported that pyrite,

D

chalcopyrite and sphalerite improved gold and silver leaching due to the galvanic interaction whereas chalcocite and galena formed sulfide oxidation products and did not improve the

TE

leaching.

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Instead of pyrite, silver is an alternative chemical catalyst to enhance chalcopyrite leaching via redox interactions [6, 225]. When silver ions were deposited onto the chalcopyrite surface, they form silver sulfide resulting in the release of cupric and ferrous ions (Equations 76, 77 and 78).

 
  4ST        2ST 


 
  4ST     ST       

ST
  2    2ST  2    


2ST 
    4  4ST  2
  2 

76 76a 77 78

The activation energy was significantly reduced from 130.7 kJ mol-1 to 29.3 kJ mol-1 by adding silver [2]. It should be noted that this process is not analogous to the galvanic interaction of argentite (Ag2S) with chalcopyrite which would be predicted to retard chalcopyrite leaching due to the smaller rest potential of argentite (Table 7). Rather Equations 76 to 78 suggest a direct redox reaction of Ag+ with the chalcopyrite surface. Additionally it is unlikely that the bulk mineral phase argentite is actually formed, as may be suggested by Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT Equation 76. A more accurate representation of such a surface reaction is proposed in Equation 76a where the bulk of the chalcopyrite is left unaffected. When silver was applied to low grade chalcopyrite ore during bioleaching, it was

PT

preferentially attracted to galena and copper sulfide compared to pyrite and sphalerite [2]. Very small doses of silver were found to improve copper recovery during low grade

RI

chalcopyrite bioleaching [178].

SC

Further electrochemical tests on the chalcopyrite/pyrite galvanic cell have also been carried out with acid aqueous silver catalysed leaching at 35 °C for 2 days [2]. The pH was

NU

maintained at 1.8. The electrode was prepared by welding chalcopyrite and pyrite pieces to a copper wire. The electrochemical cell was equipped with three electrodes: reference electrode

MA

(Ag/AgCl), counter electrode (Pt) and working electrode (chalcopyrite or pyrite). In the absence of aqueous silver, pyrite behaved cathodically. When silver was added to the solution, the silver was attracted to the chalcopyrite surface and added anodic sites on the

TE

6.3 MnO2

D

surface to enhance the chalcopyrite dissolution.

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MnO2 has also been used as a cathode material [226] due to its very high rest potential (1.20 V SHE), [227, 228]. The dissolution of MnO2 is proposed to occur as: C
  4  2   C  2 

79

Unlike the chalcopyrite/pyrite system, Fe3+(aq) does not promote the dissolution rate because MnO2 lacks the capacity for Fe3+ to Fe2+ reduction. However the interaction with protons still accelerates the dissolution rate.

MnO2 (pyrolusite) was used as cathodic material to study the galvanic interactions with pyrite, chalcopyrite, sphalerite and galena [226, 229]. The anodic reactions observed are given in Equations 80 to 83. UM(WA: 
  4      2  16  14  '#YZ[\M(WA:  
        2
  4 

\'#Y(WA: ]
  ]  
  2  ^#Y#: UV
  UV   
  2 

Minerals and Materials Science & Technology Mawson Institute, University of South Australia

80 81 82 83

60

ACCEPTED MANUSCRIPT The rest potential of sulfide minerals measured in H2SO4 electrolyte did not show significant difference from HCl electrolyte. When the leaching experiments of CuFeS2-MnO2 and ZnSMnO2 couples were performed in HCl medium, Cl2 gas, which dissolves CuFeS2 and ZnS, was produced.

RI

C
  Y
9  CY  


PT

C
  4Y  CY  Y
9  2 

84 85

SC

As the rest potential of the four sulfide minerals decreases in the order pyrite > chalcopyrite > sphalerite > galena, their potential difference with MnO2 in the galvanic cells increases in this

NU

order, MnO2-pyrite < MnO2-chalcopyrite < MnO2-sphalerite < MnO2-galena. Though chalcopyrite does not dissolve well independently in diluted HCl leach media (3 or 4

MA

M), it dissolves in the presence of MnO2. The Cl2 gas formed from the MnO2-HCl reaction is the major leach reaction route compared to the galvanic reaction [226, 229].

D

When Gantayat et al. [227] leached chalcopyrite with MnO2 in sulfate leach media very

TE

strong galvanic interactions were observed. Because pyrite was present as an impurity in the chalcopyrite, the pyrite also dissolved to produce Fe2+(aq) , H+ and e- (see Equation 80) due to

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galvanic interaction with the MnO2 thus causing further MnO2 dissolution than predicted (Equation 79). The chalcopyrite surface developed a passivation film being during the anodic dissolution.

In the leaching with both ammoniacal and ammoniacal thiosulfate solutions, the dissolution rates of pyrite and chalcopyrite were increased in the presence of MnO2 [230]. 6.4 Activated Carbon

Because activated carbon is conductive, it forms galvanic couples with chalcopyrite and sphalerite. The activated carbon acts as a cathode while chalcopyrite and sphalerite act as anodes [206]. The anodic dissolution reactions are provided in Equation 81 for chalcopyrite and Equation 82 for sphalerite while the cathodic reaction is given in Equation 69. In bioleaching, activated carbon is beneficial at low redox potentials because the process goes through chalcocite oxidation [108]. Without activated carbon, the chalcopyrite is oxidised by Fe3+ and dissolved O2 at high redox potentials (Equations 86 and 87).  
  4       5    2


Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT  
  4          2
  2 

87

The S0 formed on the chalcopyrite surface may act as a passivation coating. In the presence of activated carbon, a chalcocite-like phase is formed by reduction of the chalcopyrite with

 
       
     


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Fe3+ and dissolved O2 to release Cu2+ (Equations 90 and 91).

PT

Cu2+ and Fe2+ at low redox potential (Equations 88 and 89). The chalcocite is oxidised by

SC

 
  3   3    2  
  3  

 2    
  4    2   


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  4    2   
  4  

88 89 90 91

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When 2 g/L of activated carbon was added to the chalcopyrite bioleaching systems, the Cu recovery was enhanced from 64% to 95% within 10 days of leaching [201]. The activated carbon was presumably acting as a cathode reducing Fe3+. Chalcocite was identified as an

6.5 Flotation

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intermediate species, though jarosite and S0 were also identified.

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Flotation requires grinding of the minerals, which often includes several semi-conductive minerals in contact. The grinding utilises media such as balls which are frequently metallic in nature. When minerals are ground for flotation, there may be galvanic interaction between: (1) two types of semi-conductive minerals; and/or (2) semi-conductive mineral and grinding media. Compared to leaching, both these types of galvanic interaction have to be considered as impacting on flotation recovery as well as other factors for instance the types of collectors, solution potential, particle size, etc. During grinding of chalcopyrite, a galvanic interaction also occurs between the grinding media (cast iron, mild steel, hyper steel and stainless steel) and chalcopyrite:


     2 

S[@WZ (#ZAW[

  2   4   4 

#A'[@WZ (#ZAW[ [   _ (`#Z_

92 93

Steels act as active metals while chalcopyrite acts as a noble material since the electrode potentials of steels are much lower than that of chalcopyrite as summarised in Table 8 [231].

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ACCEPTED MANUSCRIPT The dissolved Fe is precipitated as oxy-hydroxide due to the generally alkaline pH of the system. Peng and Grano [232] reported grinding media acted as the active metal while chalcopyrite

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and galena acted as a noble metal. Iron hydroxide was deposited on chalcopyrite and galena affecting their floatability. Grinding media also galvanically interacted with pyrite and

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chalcopyrite and affected their floatability due to surface oxidation [233, 234].

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Bruckard et al. [235] compared the galvanic interactions in an O2 and N2 purged mill. The galvanic interaction was stronger in the presence of dissolved O2 because the oxygen can act

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as an electron acceptor to form hydroxyl ions. The galvanic interaction is weakened in the presence of N2 because nitrogen displaces the oxygen and thus eliminates the formation of

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hydroxyl ions that compete for adsorption sites with the collector. Ahn and Gebhardt [236] also reported that when chalcopyrite was ground with high-carbon steel media in a N2-purged mill, low grinding solution pH increased the galvanic interaction causing poorer floatability.

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The flotation recoveries of chalcopyrite are affected after contact with sphalerite or galena in

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the presence of O2. The floatability of the noble mineral, in this instance chalcopyrite, was significantly more affected than the floatability of the active minerals, sphalerite or galena

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[237]. Galvanic interaction between chalcopyrite and pyrite significantly affects the floatability of both minerals. Collectorless flotation of pyrite increases while chalcopyrite is depressed. On the pyrite surface, S0 or metal-deficient sulfide was formed. On the chalcopyrite surfaces, hydrophilic iron oxide/hydroxide and sulfur-oxygen species were formed [210].

6.6 Bioleaching

As stated previously galvanic interaction is also observed during bioleaching [238] particularly in the presence of pyrite [239]. Electrochemical leaching of chalcopyrite was enhanced by a factor of 4.6 at 30 °C due to galvanic interaction of chalcopyrite being in contact with pyrite. Adding bacteria further improved the leaching by an additional factor of 2.1. No significant improvement was observed when the temperature was increased to 55 °C [240]. The electrochemical behaviour of galena- and sphalerite-rich ores was also influenced by galvanic interaction [241]. During bioleaching sphalerite was passivated due to the formation of a galvanic cell with galena [242]. As the galena content was increased, the sphalerite leaching was further limited during bioleaching. However, after the majority of the Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT galena was oxidised, the Zn extraction rate increased most likely linearly, confirming the sphalerite leaching was retarded in presence of galena [242]. The galvanic interaction during bioleaching is very comparable to that during chemical leaching.

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Fe3+(aq) produced by bacterial sulfide oxidation is an important factor for bioleaching. High redox potential and low pH values (such that jarosite is not precipitated) are essential to

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achieve the fastest leaching for Cu and Co as the solubility increases while Ni and Zn are

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independent from pH and redox potential [243]. Particle size and ore mineralogy also strongly affect galvanic interactions during bioleaching [244]. Cruz et al. [245] used cyclic voltammetry to determine how much the galvanic reaction could occur in ore concentrates.

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during bioleaching of ore concentrates.

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They found a galvanic interaction could occur even with small amounts of sulfide minerals

7. Summary

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CuFeS2 is isostructural with ZnS with the size of unit cell of the former being approximately

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two times larger than the latter. One CuFeS2 unit cell contains four Cu, four Fe and eight S atoms. Every S atom in the bulk is coordinated to a tetrahedron of metal atoms while each

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metal atom is coordinated to a tetrahedron of S atoms. Most investigators regard CuFeS2 as having the formal oxidation states of Cu+Fe3+(S2-)2 while some also find the evidence for the presence of Cu2+ and Fe2+ although this has been largely discredited. However, the spectroscopic characterisation of chalcopyrite is clearly influenced by the considerable degree of covalency between S and both Fe and Cu indicating an actual valence state character intermediate between Cu+Fe3+(S2-)2 and Cu2+Fe2+(S2-)2. It has also been proposed that a partial Cu d9 configuration may be due to hybridisation of the occupied Cu 3d with empty Fe 3d states via the intervening S 3p. This possibility may suggest possible solid state electron flow routes on leaching and oxidation, which result in enhanced stabilisation and the observed slow reaction rates. Under normal atmospheric and processing conditions chalcopyrite is an antiferromagnetic narrow band gap (0.53 to 0.6 eV) semiconductor. The magnetic moment for Fe is found to be smaller than possible for any completely ionic Fe3+ oxidation state thus again suggesting a high degree of covalency between the Fe and S atoms within chalcopyrite. The poor cleavage of CuFeS2 results in conchoidal surfaces which are generally a mixture of different surface planes, e.g. (001), (100), (112). Reconstruction of the fractured surfaces to Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT form, from what was previously bulk S2-, a mixture of surface S2-, S22 and Sn2- takes place. These oxidation processes seem likely to be accompanied by the reduction of surface Fe3+ but clear experimental evidence for a reduced Fe3+ species seems to be lacking. Moreover it has

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been proposed that a pyrite FeS2 moiety is formed which would suggest the presence of low spin Fe2+. However, as far as the authors are aware this has not been observed

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experimentally. If this were to be the case this suggests that reaction rate between low spin Fe2+ and triplet O2 may be limited by the requirement for spin redistribution in order for    to take place prior to the dissociation of molecular oxygen.

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Oxidation of chalcopyrite in air (i.e. 0.21 atm of O2 equilibrated with water vapour) results in a Fe(III)-O-OH surface layer on top of a Cu rich sulfide layer overlying the bulk chalcopyrite. The formation of Cu(II) and Fe(III) sulfate, and Cu(I)-O have also been reported on

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prolonged oxidation, however the main Cu-bearing surface species appears to be Cu(I)-S. The surface speciation on aqueous leaching has been reported to predominantly consist of Fe(III)-O-OH initially, which on dissolution exposes a Fe-poor but Cu-rich sulfide layer.

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After prolonged leaching Fe(III)-S has been observed as have Cu(I)-Sn2- and S0. SO42- has

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also been observed in some instances but generally the surface concentration of this species is

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small due to its considerable solubility. Although considerable research has been carried in order to understand aqueous chalcopyrite leaching, the resulting surface chemistry is not been well understood. S22-, Sn2- (or metal deficient sulfide) and S0 are proposed to form on the chalcopyrite surface leached in both acidic and alkaline solutions. The latter two of these species along with a jarosite-like species are most frequently proposed to result in surface leaching passivation when chalcopyrite is leached in ferric sulfate medium. However, some investigators have reported the formation of S0 sufficiently porous to allow ion transportation to and from the chalcopyrite surface. Moreover, under some conditions, i.e. in the absence of the initial addition of Fe3+, both Sn2and S0 were observed to increase in surface concentration for the duration of the leach with no resulting passivation in HClO4, H2SO4 and HCl solutions. In contrast, in the presence of initially added Fe3+ the leach rate in H2SO4 was observed, overall to be slower, with greater surface components of SO42- and Sn2-. It seems probable, given these observations, that the rate of leaching is a function not only of surface speciation but also the porosity of the surface speciation, although how to control this is not yet clear.

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ACCEPTED MANUSCRIPT Interestingly, on a number of occasions the leach rate of chalcopyrite has been observed to reach a maximum on increase in agitation rate followed by a decrease thereafter on further increases in agitation rate. This has been proposed to result from reduced contact between the

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chalcopyrite particles and the solution borne oxidants. The leach rate of chalcopyrite is generally observed to increase with increasing acid concentration (i.e. decreased pH)

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however the assumption that increased [H+] is linked directly to leach rate may be incorrect as this also results in increased solubilisation of the oxidant Fe3+. It is noted that the acidic

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anion can also effect leach rates with various theories proposed including surface adsorption, e.g. in the case of Cl-, and solution complexation with Fe3+ thus reducing the concentration of

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free oxidant.

The effect of a number of oxidants. e.g. O2, H2O2, Cu2+, Cr6+ and Fe3+, has been examined.

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However, this is often accompanied by poor control of leach parameters so that the leaches with oxidant present may well have varying Eh, and possibly pH, as compared to the leaches in their absence. Hence, it is difficult to draw an overall conclusion regarding the real direct

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effect of the added chemical constituent. Where leaches have been carried out in the absence

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and presence of added Fe3+ with controlled pH and Eh the overall rate of leaching was found to be reduced in the presence of added Fe3+, however, this may not be the case for all

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oxidants or leaching conditions.

There is general agreement in the literature that chalcopyrite leaching is significantly affected by solution redox potential with an optimum Eh range above and below which the leach rate decreases. This suggests the participation of leach steps that involve both oxidation and reduction so that for high Eh reduction is rate determining while at low Eh the oxidation is rate determining. This conclusion is also supported by the observation that the addition of reductants, e.g. Fe2+ have resulted in increased leach rate for some conditions. There are essentially three kinetic models for CuFeS2 leaching: diffusion, surface reaction and a mixed kinetic model containing components of diffusion and surface reaction. It is not really possible to compare directly Ea derived due to the wide range of experimental conditions used, whether oxidant, i.e. Fe3+ was added or not and the extent of reaction examined, i.e. initial leach rates, or leach rate over a more extended period. Moreover, in many cases it appears that neither pH nor Eh were controlled and hence reaction conditions would have varied over the duration of the leach. However, it appears most likely the leach is surface reaction rate controlled with some initial period depending on leach conditions Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT where the rate is surface layer diffusion controlled. Further examination of changes in leach kinetics and Ea as a function of controlled comparative conditions would enable more in depth determination of what stages of leaching and for which specific conditions different

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rate controlling factors are at play. Although bioleaching of some copper ores has been adopted by industry, bioleaching has yet

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to be applied to predominantly chalcopyrite ores due to the slow resulting leach rates. It has

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been observed that mixed microbial strains usually yield higher leach rates, as compared to single strains, as different bacterial strains are able to adapt to the changing leach conditions throughout the leach process. The microbes used in bioleaching must be able to survive high

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temperature (> 30 °C), small pH and large metal concentrations. Such microbes are usually able to oxidise Fe and/or S and may therefore play a role in the breakdown of sulfidic

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passivating surface layers.

Microbes may mediate bioleaching through contact, indirect and cooperative mechanisms.

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The contact mechanism may proceed through an enzymatic oxidation involving Fe3+ which is

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a constituent of the extracellular polymeric substance. The Fe3+ may extract electrons from the mineral surface through a complex redox chain within bacterial membranes enabling the reduction of which reduces molecular oxygen. The cooperative mechanism is established

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between the contacted and planktonic cells, when the contacted cells supply Fe2+ and S (or other S products) to the planktonic cells for their energy source. The net result is an efficient supply of oxidants (Fe3+) for oxidation and removal of surface S products. Bacterial activity can be beneficial for chalcopyrite leaching if leach conditions are optimised accordingly: low initial Fe3+, low redox potential (best if controlled), sufficient initial Fe2+, use of high temperature and moderate to extreme thermophiles and fine grinding. However, it remains unclear whether bioleaching can outperform optimised chemical leaching at the same temperature. Interestingly it has been observed for bioleaching, as for chemical leach experiments, that significant initial concentrations Fe3+ and high Eh were detrimental to leach rates. It has also been proposed that for cases where less acid (H2SO4) is needed to reduce downstream neutralisation costs, the bacteria should be prevented from contacting the mineral surface. Under this condition S oxidation produces elemental S and not H2SO4. As for chemical leaching, passivation is also observed on bioleaching with jarosite being likely to be the main contributor. Bacterial action tends to produce more Fe3+(aq) which on combination with high solution sulfate and monovalent alkali cations (needed as microbial Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT nutrients) can result in jarosite precipitation in the pH 1.5-2.5 range. Therefore, it may be advantageous to operate bioleaching at solution Eh conditions such that the solution Fe3+ concentration is reduced. S0 and various other intermediate Cu oxide and sulfide species have

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also been proposed to play a role in passivation. A number of factors have been identified as enhancing the rate of bioleaching: Higher

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temperature; lower redox; chloride; silver ions; and moderate fine grinding particles with high surface area. The addition of Ag+ is proposed to lead to the formation of Ag2S on the

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chalcopyrite surface which is relatively easily oxidised. However, the addition of Ag+ to solution can also lead to jarosite precipitation and is toxic to some bacteria. Furthermore, it is

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very unlikely that the addition of Ag+ to a chalcopyrite leach process would be an economic option. Activated carbon has also been shown to increase the leach rate of chalcopyrite,

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possibly due to the formation of an effective electronic couple. Galvanic interactions are caused by electrochemical interaction as a result of the differing rest

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potentials of semi-conductive minerals and can affect both chemical leaching and

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bioleaching. It is proposed that galvanic interactions enhance the oxidation of the anodic mineral(s) leading to an increase of metal ions in solution and protect the cathodic mineral

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from oxidation. This suggests that when chalcopyrite is in electrochemical contact with pyrite, chalcopyrite will dissolve prior to pyrite. However, when chalcopyrite is in electrochemical contact with many other metal sulfide minerals, e.g. sphalerite or galena the leaching of chalcopyrite will be delayed. The degree of galvanic interaction between minerals of contrasting rest potentials has been shown to be increased by increasing solution Fe3+ concentrations.

Interestingly, while S0 has often been observed to form thick surface coatings on leaching of chalcopyrite in the presence of pyrite these coatings are clearly porous and do not passivate the leach process. The presence of Ag is found to improve the leaching of chalcopyrite in contact with pyrite, but on introduction of further sulfide minerals the effect is less clear and may depend on which surfaces the Ag+ preferentially adsorb onto. However, overall the presence of Ag+ appears to be beneficial. Also, as for bioleaching, the addition of activated carbon also appears to be beneficial by acting as an electrical conduit aiding in the oxidation of the intermediate Cu2S product. In summary, while much has been observed at the macro-scale regarding the chalcopyrite leach process it is clear that interpretation of these phenomena is hampered by lack of Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT understanding at the molecular or atomic scale. Three primary questions that require elucidation, before the overall mechanism can be understood are: 1. How does the surface of chalcopyrite interact with solution or air borne oxidants?

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2. How does the nature of these oxidants affect the surface products formed? 3. What determines whether the surface formed will be passivating or not?

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These can only realistically be tackled by the application of near atomic-scale analytical

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approaches, which may include quantum chemical modelling, PEEM/SPEM, TEM, AFM etc.

8. Acknowledgement

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Financial support from Rio Tinto and Australian Research Council via the ARC-Linkage Project ‘Solution and surface speciation evolution during chalcopyrite leaching’

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(LP110200326) is gratefully acknowledged. Proof reading and add in interpretation by Dr. Paul Brown (Rio Tinto) is also gratefully acknowledged.

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9. References

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[1] Habashi F. Chalcopyrite. Its Chemistry and Metallurgy: McGraw-Hill; 1978.

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[2] Córdoba EM, Muñoz JA, Blázquez ML, González F, Ballester A. Leaching of chalcopyrite with ferric ion. Part I: General aspects. Hydrometallurgy. 2008;93:81-7. [3] Harmer SL, Thomas JE, Fornasiero D, Gerson AR. The evolution of surface layers formed during chalcopyrite leaching. Geochimica et Cosmochimica Acta. 2006;70:4392-402. [4] Pradhan N, Nathsarma KC, Srinivasa Rao K, Sukla LB, Mishra BK. Heap bioleaching of chalcopyrite: A review. Minerals Engineering. 2008;21:355-65. [5] Watling HR. The bioleaching of sulphide minerals with emphasis on copper sulphides -A review. Hydrometallurgy. 2006;84:81-108. [6] Hiroyoshi N, Arai M, Miki H, Tsunekawa M, Hirajima T. A new reaction model for the catalytic effect of silver ions on chalcopyrite leaching in sulfuric acid solutions. Hydrometallurgy. 2002;63:257-67. [7] Ahmadi A, Schaffie M, Petersen J, Schippers A, Ranjbar M. Conventional and electrochemical bioleaching of chalcopyrite concentrates by moderately thermophilic bacteria at high pulp density. Hydrometallurgy. 2011;106:84-92. [8] Hackl RP, Dreisinger DB, Peters E, King JA. Passivation of chalcopyrite during oxidative leaching in sulfate media. Hydrometallurgy. 1995;39:25-48. [9] Stott MB, Watling HR, Franzmann PD, Sutton D. The role of iron-hydroxy precipitates in the passivation of chalcopyrite during bioleaching. Minerals Engineering. 2000;13:1117-27.

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ACCEPTED MANUSCRIPT [10] Tshilombo AF. Mechanism and kinetics of chalcopyrite passivation and depassivation during ferric and microbial leaching: PhD Thesis, University of Brithish Columbia; 2004.

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[11] Córdoba EM, Muñoz JA, Blázquez ML, González F, Ballester A. Passivation of chalcopyrite during its chemical leaching with ferric ion at 68 °C. Minerals Engineering. 2009;22:229-35.

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[12] Viramontes-Gamboa G, Peña-Gomar MM, Dixon DG. Electrochemical hysteresis and bistability in chalcopyrite passivation. Hydrometallurgy. 2010;105:140-7.

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[13] Arce EM, González I. A comparative study of electrochemical behavior of chalcopyrite, chalcocite and bornite in sulfuric acid solution. International Journal of Mineral Processing. 2002;67:17-28.

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[14] Li J, Kawashima N, Kaplun K, Absolon VJ, Gerson AR. Chalcopyrite leaching: The rate controlling factors. Geochimica et Cosmochimica Acta. 2010;74:2881-93.

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[15] Dutrizac JE. Elemental sulphur formation during the ferric sulphate leaching of chalcopyrite. Canadian Metallurgical Quarterly. 1989;28:337-44.

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[16] Antonijevic MM, Jankovic Z, Dimitrijevic M. Investigation of the kinetics of chalcopyrite oxidation by potassium dichromate. Hydrometallurgy. 1994;35:187-201.

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[17] Harmer SL. Surface layer control for improved copper recovery for chalcopyrite leaching: University of South Australia; 2002.

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[18] Xian YJ, Wen SM, Deng JS, Liu J, Nie Q. Leaching chalcopyrite with sodium chlorate in hydrochloric acid solution. Canadian Metallurgical Quarterly. 2012;51:133-40. [19] Dutrizac JE. The leaching of sulphide minerals in chloride media. Hydrometallurgy. 1992;29:1-45. [20] Linge HG. Reactivity comparison of Australian chalcopyrite concentrates in acidified ferric solution. Hydrometallurgy. 1977;2:219-33. [21] Yin Q, Kelsall GH, Vaughan DJ, England KER. Atmospheric and electrochemical oxidation of the surface of chalcopyrite (CuFeS2). Geochimica et Cosmochimica Acta. 1995;59:1091-100. [22] Klauber C, Parker A, van Bronswijk W, Watling H. Sulphur speciation of leached chalcopyrite surfaces as determined by X-ray photoelectron spectroscopy. International Journal of Mineral Processing. 2001;62:65-94. [23] Parker AJ, Paul RL, Power GP. Electrochemistry of the oxidative leaching of copper from chalcopyrite. Journal of Electroanalytical Chemistry and Interfacial Electrochemistry. 1981;118:305-16. [24] Brion D. Photoelectron spectroscopic study of the surface degradation of pyrite (FeS 2), chalcopyrite (CuFeS 2), sphalerite (ZnS), and galena (PbS) in air and water. Appl Surf Sci. 1980;5:133–52.

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ACCEPTED MANUSCRIPT [25] Buckley AN, Woods R. An X-ray photoelectron spectroscopic study of the oxidation of chalcopyrite. Australian Journal of Chemistry. 1984;37:2403-13.

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[26] Dutrizac J. The dissolution of chalcopyrite in ferric sulfate and ferric chloride media. Metallurgical and Materials Transactions B. 1981;12:371-8.

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[27] Klauber C. A critical review of the surface chemistry of acidic ferric sulphate dissolution of chalcopyrite with regards to hindered dissolution. International Journal of Mineral Processing. 2008;86:1-17.

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[28] Parker A, Klauber C, Kougianos A, Watling HR, van Bronswijk W. An X-ray photoelectron spectroscopy study of the mechanism of oxidative dissolution of chalcopyrite. Hydrometallurgy. 2003;71:265-76.

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[29] Dreisinger D, Abed N. A fundamental study of the reductive leaching of chalcopyrite using metallic iron part I: kinetic analysis. Hydrometallurgy. 2002;66:37-57.

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[30] Saxena NN, Mandre NR. Mixed control kinetics of copper dissolution for copper ore using ferric chloride. Hydrometallurgy. 1992;28:111-7.

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[31] Ikiz D, Gülfen M, AydIn AO. Dissolution kinetics of primary chalcopyrite ore in hypochlorite solution. Minerals Engineering. 2006;19:972-4.

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[32] Antonijevic MM, Bogdanovic GD. Investigation of the leaching of chalcopyritic ore in acidic solutions. Hydrometallurgy. 2004;73:245-56.

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[33] Hiroyoshi N, Kuroiwa S, Miki H, Tsunekawa M, Hirajima T. Synergistic effect of cupric and ferrous ions on active-passive behavior in anodic dissolution of chalcopyrite in sulfuric acid solutions. Hydrometallurgy. 2004;74:103-16. [34] Orth R, Liddell K. Rate law and mechanism for the oxidation of copper (I) by iron (III) in hydrochloric acid solutions. Industrial & Engineering Chemistry Research. 1990;29:117883. [35] Parker A, Klauber C, Kougianos A, Watling HR, van Bronswijk W. An X-ray photoelectron spectroscopy study of the mechanism of oxidative dissolution of chalcopyrite. Hydrometallurgy. 2003;71:265-76. [36] Zivkovic ZD, Mitevska N, Savovic V. Kinetics and mechanism of the chalcopyritepyrite concentrate oxidation process. Thermochimica Acta. 1996;282-283:121-30. [37] Al-Harahsheh M, Kingman S, Rutten F, Briggs D. ToF-SIMS and SEM study on the preferential oxidation of chalcopyrite. International Journal of Mineral Processing. 2006;80:205-14. [38] Yarar B, Spottiswood D. Interfacial phenomena in mineral processing. Engineering Foundation, New York, NY; 1982. [39] Dutrizac J, MacDonald R. The effect of some impurities on the rate of chalcopyrite dissolution. Canadian Metallurgical Quarterly. 1973;12:409-20.

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ACCEPTED MANUSCRIPT [40] Nicol M. Kinetics of the oxidation of copper (I) by oxygen in acidic chloride solutions. S Afr J Chem. 1984;37:77–80.

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[41] Padilla R, Pavez P, Ruiz MC. Kinetics of copper dissolution from sulfidized chalcopyrite at high pressures in H2SO4-O2. Hydrometallurgy. 2008;91:113-20. [42] Mahajan V, Misra M, Zhong K, Fuerstenau MC. Enhanced leaching of copper from chalcopyrite in hydrogen peroxide-glycol system. Minerals Engineering. 2007;20:670-4.

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[43] Bjorling G, Faldt I, Lindgren E, Toromanov I. NITRIC ACID ROUTE IN COMBINATION WITH SOLVENT EXTRACTION FOR HYDROMETALLURGICAL TREATMENT OF CHALCOPYRITE. 1976:725-37.

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[44] Prater JD, Queneau PB, Hudson TJ. NITRIC ACID ROUTE TO PROCESSING COPPER CONCENTRATES. Trans Soc Mining Eng AIME. 1973;254:117-22.

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[45] Hiroyoshi N, Miki H, Hirajima T, Tsunekawa M. A model for ferrous-promoted chalcopyrite leaching. Hydrometallurgy. 2000;57:31-8. [46] Aydogan S, Ucar G, Canbazoglu M. Dissolution kinetics of chalcopyrite in acidic potassium dichromate solution. Hydrometallurgy. 2006;81:45-51.

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[47] Burdick CL, Ellis JH. The crystal structure of chalcopyrite determined by X-rays. Journal of the American Chemical Society. 1917;39:2518-25.

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[48] Edelbro R, Sandström Å, Paul J. Full potential calculations on the electron bandstructures of Sphalerite, Pyrite and Chalcopyrite. Applied Surface Science. 2003;206:300-13. [49] Hall SR, Stewart JM. The crystal structure refinement of chalcopyrite, CuFeS2. Acta Crystallographica Section B. 1973;29:579-85. [50] Jones RT. Electronic structures of the sulfide minerals sphalerite, wurtzite, pyrite, marcasite, and chalcopyrite: PhD thesis,University of South Australia,Adelaide; 2006. [51] Petiau J, Sainctavit P, Calas G. K X-ray absorption spectra and electronic structure of chalcopyrite CuFeS2. Materials Science and Engineering: B. 1988;1:237-49. [52] Nikiforov KG. Magnetically ordered multinary semiconductors. Progress in Crystal Growth and Characterization of Materials. 1999;39:1-104. [53] Llanos J, Buljan A, Mujica C, Ramírez R. Electron transfer in the insertion of alkali metals in chalcopyrite. Materials Research Bulletin. 1995;30:43-8. [54] de Oliveira C, Duarte HA. Disulphide and metal sulphide formation on the reconstructed surface of chalcopyrite: A DFT study. Applied Surface Science. 2010;257:1319-24. [55] Von Oertzen G, Harmer S, Skinner WM. XPS and ab initio calculation of surface states of sulfide minerals: pyrite, chalcopyrite and molybdenite. Molecular Simulation. 2006;32:1207-12.

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ACCEPTED MANUSCRIPT [56] Raj D, Chandra K, Puri S. Mössbauer studies of chalcopyrite. Journal of the Physical Society of Japan. 1968;24:39-41.

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[57] Mikhlin Y, Tomashevich Y, Tauson V, Vyalikh D, Molodtsov S, Szargan R. A comparative X-ray absorption near-edge structure study of bornite, Cu5FeS4, and chalcopyrite, CuFeS2. Journal of Electron Spectroscopy and Related Phenomena. 2005;142:83-8.

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[58] Todd EC, Sherman DM. Surface oxidation of chalcocite (Cu2S) under aqueous (pH= 211) and ambient atmospheric conditions: Mineralogy from Cu L-and O K-edge X-ray absorption spectroscopy. American Mineralogist. 2003;88:1652.

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[59] Sainctavit P, Petiau J, Flank A, Ringeissen J, Lewonczuk S. XANES in chalcopyrites semiconductors: CuFeS2, CuGaS2, CuInSe2. Physica B: Condensed Matter. 1989;158:623-4. [60] Kono S, Okusawa M. X-ray photoelectron study of the valence bands in I-III-VI2 compounds. Journal of the Physical Society of Japan. 1974;37:1301-4.

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[61] Pearce CI, Pattrick RAD, Vaughan DJ, Henderson CMB, van der Laan G. Copper oxidation state in chalcopyrite: Mixed Cu d9 and d10 characteristics. Geochimica et Cosmochimica Acta. 2006;70:4635-42.

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ACCEPTED MANUSCRIPT [202] Gómez E, Ballester A, Blázquez ML, Gonzáez F. Silver-catalysed bioleaching of a chalcopyrite concentrate with mixed cultures of moderately thermophilic microorganisms. Hydrometallurgy. 1999;51:37-46.

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[203] Sand W, Gehrke T, Jozsa P-G, Schippers A. (Bio)chemistry of bacterial leaching— direct vs. indirect bioleaching. Hydrometallurgy. 2001;59:159-75.

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[204] Schippers A, Sand W. Bacterial leaching of metal sulfides proceeds by two indirect mechanisms via thiosulfate or via polysulfides and sulfur. Applied and Environmental Microbiology. 1999;65:319.

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[205] Johnson DB, Okibe N, Wakeman K, Yajie L. Effect of temperature on the bioleaching of chalcopyrite concentrates containing different concentrations of silver. Hydrometallurgy. 2008;94:42-7. [206] Nakazawa H, Fujisawa H, Sato H. Effect of activated carbon on the bioleaching of chalcopyrite concentrate. International Journal of Mineral Processing. 1998;55:87-94.

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[207] Rao SR, Finch JA. Galvanic Interaction Studies on Sulphide Minerals. Canadian Metallurgical Quarterly. 1988;27:253-9.

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[208] Subrahmanyam TV, Forssberg KSE. Mineral solution-interface chemistry in minerals engineering. Minerals Engineering. 1993;6:439-54.

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[209] Mehta AP, Murr LE. Fundamental studies of the contribution of galvanic interaction to acid-bacterial leaching of mixed metal sulfides. Hydrometallurgy. 1983;9:235-56.

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[210] Ekmekci Z, Demirel H. Effects of galvanic interaction on collectorless flotation behaviour of chalcopyrite and pyrite. International Journal of Mineral Processing. 1997;52:31-48. [211] Qing You L, Heping L, Li Z. Study of galvanic interactions between pyrite and chalcopyrite in a flowing system: implications for the environment. Environmental Geology. 2007;52:11-8. [212] Liu Q, Li H, Zhou L. Galvanic interactions between metal sulfide minerals in a flowing system: Implications for mines environmental restoration. Applied Geochemistry. 2008;23:2316-23. [213] Sui CC, Brienne SHR, Ramachandra Rao S, Xu Z, Finch JA. Metal ion production and transfer between sulphide minerals. Minerals Engineering. 1995;8:1523-39. [214] Dixon DG, Mayne DD, Baxter KG. GALVANOX(TM) - A novel galvanically-assisted atmospheric leaching technology for copper concentrates. Canadian Metallurgical Quarterly. 2008;47:327-36. [215] Chmielewski T, Kaleta R. Galvanic interactions of sulfide minerals in leaching of flotation concentrate from Lubin Concentrator. Physicochem Probl Miner Process. 2011;46:21-34.

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ACCEPTED MANUSCRIPT [216] Abraitis PK, Pattrick RAD, Kelsall GH, Vaughan DJ. Acid leaching and dissolution of major sulphide ore minerals: processes and galvanic effects in complex systems. Mineralogical Magazine. 2004;68:343-51.

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[217] Payant RA, Finch JA. The Effect of Sulphide Mixtures on Self-Heating. Canadian Metallurgical Quarterly. 2010;49:429-34.

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[218] Payant R, Rosenblum F, Nesset JE, Finch JA. The self-heating of sulfides: Galvanic effects. Minerals Engineering. 2012;26:57-63.

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[219] Urbano G, Reyes VE, Veloz MA, González I, Cruz Jn. Pyrite−Arsenopyrite Galvanic Interaction and Electrochemical Reactivity. The Journal of Physical Chemistry C. 2008;112:10453-61.

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[220] Urbano G, Meléndez AM, Reyes VE, Veloz MA, González I. Galvanic interactions between galena–sphalerite and their reactivity. International Journal of Mineral Processing. 2007;82:148-55.

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[221] Dixon DG, Tshilombo AF. Leaching process for copper concentrates. C22B 15/00, 3/08 ed2005. Chapter WO 2005/118894 A1.

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[222] Eghbalnia M, Dixon DG. Electrochemical study of leached chalcopyrite using solid paraffin-based carbon paste electrodes. Hydrometallurgy. 2011;110:1-12.

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[223] Nazari G, Dixon DG, Dreisinger DB. The role of galena associated with silverenhanced pyrite in the kinetics of chalcopyrite leaching during the Galvanox™ process. Hydrometallurgy. 2012;111–112:35-45. [224] Azizi A, Petre CF, Olsen C, Larachi F. Untangling galvanic and passivation phenomena induced by sulfide minerals on precious metal leaching using a new packed-bed electrochemical cyanidation reactor. Hydrometallurgy. 2011;107:101-11. [225] Muñoz JA, Dreisinger DB, Cooper WC, Young SK. Silver-catalyzed bioleaching of low-grade copper ores.: Part I: Shake flasks tests. Hydrometallurgy. 2007;88:3-18. [226] Devi NB, Madhuchhanda M, Rao KS, Rath PC, Paramguru RK. Oxidation of chalcopyrite in the presence of manganese dioxide in hydrochloric acid medium. Hydrometallurgy. 2000;57:57-76. [227] Gantayat B, Rath P, Paramguru R, Rao S. Galvanic interaction between chalcopyrite and manganese dioxide in sulfuric acid medium. Metallurgical and Materials Transactions B. 2000;31:55-61. [228] Miller JD, Wan R-Y. Reaction kinetics for the leaching of MnO2 by sulfur dioxide. Hydrometallurgy. 1983;10:219-42. [229] Devi N, Madhuchhanda M, Rath P, Rao K, Paramguru R. Simultaneous leaching of a deep-sea manganese nodule and chalcopyrite in hydrochloric acid. Metallurgical and Materials Transactions B. 2001;32:777-84. [230] Feng D, van Deventer JSJ. Interactions between sulphides and manganese dioxide in thiosulphate leaching of gold ores. Minerals Engineering. 2007;20:533-40. Minerals and Materials Science & Technology Mawson Institute, University of South Australia

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ACCEPTED MANUSCRIPT [231] Yelloji Rao MK, Natarajan KA. Influence of galvanic interaction between chalcopyrite and some metallic materials on flotation. Minerals Engineering. 1988;1:281-94.

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[232] Peng Y, Grano S. Inferring the distribution of iron oxidation species on mineral surfaces during grinding of base metal sulphides. Electrochimica Acta. 2010;55:5470-7.

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[233] Yuan XM, Palsson BI, Forssberg KSE. Flotation of a complex sulphide ore .2. Influence of grinding environments on Cu/Fe sulphide selectivity and pulp chemistry. International Journal of Mineral Processing. 1996;46:181-204.

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[234] Peng Y, Grano S, Fornasiero D, Ralston J. Control of grinding conditions in the flotation of chalcopyrite and its separation from pyrite. International Journal of Mineral Processing. 2003;69:87-100.

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[235] Bruckard WJ, Sparrow GJ, Woodcock JT. A review of the effects of the grinding environment on the flotation of copper sulphides. International Journal of Mineral Processing. 2011;100:1-13.

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[236] Ahn JH, Gebhardt JE. Effect of grinding media-chalcopyrite interaction on the selfinduced flotation of chalcopyrite. International Journal of Mineral Processing. 1991;33:24362.

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[237] Yelloji Rao MK, Natarajan KA. Electrochemical effects of mineral-mineral interactions on the flotation of chalcopyrite and sphalerite. International Journal of Mineral Processing. 1989;27:279-93.

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[238] Attia YA, El-Zeky M. Effects of galvanic interactions of sulfides on extraction of precious metals from refractory complex sulfides by bioleaching. International Journal of Mineral Processing. 1990;30:99-111. [239] Witne JY, Phillips CV. Bioleaching of Ok Tedi copper concentrate in oxygen- and carbon dioxide-enriched air. Minerals Engineering. 2001;14:25-48. [240] Mehta AP, Murr LE. Kinetic study of sulfide leaching by galvanic interaction between chalcopyrite, pyrite, and sphalerite in the presence of T. ferrooxidans(30°C) and a thermophilic microorganism (55°C). Biotechnology and Bioengineering. 1982;24:919-40. [241] Olubambi PA, Potgieter JH, Ndlovu S, Borode JO. Electrochemical studies on interplay of mineralogical variation and particle size on bioleaching low grade complex sulphide ores. Transactions of Nonferrous Metals Society of China. 2009;19:1312-25. [242] da Silva G, Lastra MR, Budden JR. Electrochemical passivation of sphalerite during bacterial oxidation in the presence of galena. Minerals Engineering. 2003;16:199-203. [243] Ahonen L, Tuovinen OH. Bacterial leaching of complex sulfide ore samples in benchscale column reactors. Hydrometallurgy. 1995;37:1-21. [244] Olubambi PA, Ndlovu S, Potgieter JH, Borode JO. Effects of ore mineralogy on the microbial leaching of low grade complex sulphide ores. Hydrometallurgy. 2007;86:96-104.

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ACCEPTED MANUSCRIPT [245] Cruz R, Luna-Sánchez RM, Lapidus GT, González I, Monroy M. An experimental strategy to determine galvanic interactions affecting the reactivity of sulfide mineral concentrates. Hydrometallurgy. 2005;78:198-208.

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[246] Mills P, Sullivan J. A study of the core level electrons in iron and its three oxides by means of X-ray photoelectron spectroscopy. Journal of Physics D: Applied Physics. 1983;16:723-32.

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[247] Baltrus JP, Diehl JR. Surface spectroscopic studies of factors influencing xanthate adsorption on coal pyrite surfaces. Surface and Interface Analysis. 1997;25:64-70.

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[248] Heuer J, Stubbins J. An XPS characterization of FeCO3 films from CO2 corrosion. Corrosion science. 1999;41:1231-43.

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[249] Descostes M, Mercier F, Thromat N, Beaucaire C, Gautier-Soyer M. Use of XPS in the determination of chemical environment and oxidation state of iron and sulfur samples: constitution of a data basis in binding energies for Fe and S reference compounds and applications to the evidence of surface species of an oxidized pyrite in a carbonate medium. Applied Surface Science. 2000;165:288-302.

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[250] Moulder JF, Stickle WF, Sobol PE, Bomben K. Handbook of X-ray photoelectron spectroscopy, Physical Electronics. Inc, Eden Prairie. 1995.

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[251] Harmer SL, Pratt AR, Nesbitt HW, Fleet ME. Reconstruction of fracture surfaces on bornite. The Canadian Mineralogist. 2005;43:1619-30.

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[252] Buckley AN, Skinner WM, Harmer SL, Pring A, Fan L-J. Electronic environments in carrollite, CuCo2S4, determined by soft X-ray photoelectron and absorption spectroscopy. Geochimica et Cosmochimica Acta. 2009;73:4452-67. [253] Deroubaix G, Marcus P. X-ray photoelectron spectroscopy analysis of copper and zinc oxides and sulphides. Surface and Interfaces Analysis. 1992;18:39-46. [254] McIntyre NS, Cook MG. X-ray photoelectron studies on some oxides and hydroxides of colbalt, nickel and copper. Analytical Chemistry. 1975;47:2208-13. [255] Chawla SK, Sankarraman N, Payer JH. Diagnostic spectra for XPS analysis of Cu-OSH compounds. Journal of Electron Spectroscopy and Related Phenomena. 1992;61:1-18. [256] van der Laan G, Westra C, C. H, Sawatzky GA. Satellite structure in photoelectron and Auger spectra of copper dihalides. Physical Review B. 1981;23:4369-80. [257] Schaufusz AG, Nesbitt HW, Kartio I, Laajalehto K, Bancroft GM, Szargan R. Incipient oxidation of fractured pyrite surfaces in air. Journal of Electron Spectroscopy and Related Phenomena. 1998;96:69-82. [258] Smart RSC, Skinner WM, Gerson AR. XPS of Sulphide Mineral Surfaces: Metaldeficient, Polysulphides, Defects and Elemental Sulphur. Surface and interface analysis. 1999;28:101-5.

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ACCEPTED MANUSCRIPT [259] Yin Q, Vaughan D, England K, Kelsall G, Brandon N. Surface Oxidation of Chalcopyrite (CuFeS) in Alkaline Solutions. Journal of the Electrochemical Society. 2000;147:2945.

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[260] Wagner C, Naumkin A, Kraut-Vass A, Allison J, Powell C, Rumble Jr J. NIST X-ray Photoelectron Spectroscopy Database, NIST Standard Reference Database 20, Version 3.4 (Web Version). U S Department of Commerce. 2003.

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[261] Antonijevic MM, Jankovic ZD, Dimitrijevic MD. Kinetics of chalcopyrite dissolution by hydrogen peroxide in sulphuric acid. Hydrometallurgy. 2004;71:329-34.

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[262] Rath P, Paramguru R, Jena P. Kinetics of dissolution of sulphide minerals in ferric chloride solution. I. Dissolution of galena, sphalerite and chalcopyrite. Trans Inst Min Metall C. 1988;97.

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[263] Al-Harahsheh M, Kingman S, Hankins N, Somerfield C, Bradshaw S, Louw W. The influence of microwaves on the leaching kinetics of chalcopyrite. Minerals Engineering. 2005;18:1259-68.

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ACCEPTED MANUSCRIPT Table 1 Binding energy for Fe and Fe containing materials (eV) Formal oxidation state

Fe 2p3/2

Fe 2p1/2

Fe 3p

O 1s

Fe

0

706.50

719.70

52.90

FeS2

2

706.50

719.70

--

FeO

2

710.70

724.40

56.00

FeCO3

2

710.20

723.70

FeSO4

2

711.25

Fe3O4

8/3

710.20

723.70

FeOOH

3

710.60

723.80

Fe2O3

3

Fe2(SO4)3

3

[246]

284.60

[247]

284.90

[246]

531.90

289.40

[248]

532.50

285.00

[24]

284.60

[249]

284.60

[249]

--

[250]

PT

284.90

530.00

55.60 --

58.15

529.60530.20 532.40

[24]

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713.50

D

710.90

530.10

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MA 724.20

Reference

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56.00

710.80-

C 1s

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Compound

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ACCEPTED MANUSCRIPT Table 2 Cu 2p3/2 XPS binding energy (eV) of Cu compounds. Chemical formula

Formal oxidation state

Peak position

Bornite

Cu5FeS4

1

932.2

[251]

Chalcopyrite

CuFeS2

1

932.3

[71]

Carrollite

CuCo2S4

1

932.5

[252]

Chalcocite

Cu2S

1

Cuprite

Cu2O

1

Covellite

CuS

Copper hydroxide

Cu(OH)2

Tenorite

CuO

Copper bromide

CuBr2

Copper sulfate

CuSO4

Copper chloride Copper fluoride

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Mineral

Reference

[253]

932.4

[79]

2?

932.0-932.4

[79]

2

934.4

[254]

2

933.8

[255]

2

933.1

[256]

2

934.5

[3]

CuCl2

2

934.6

[256]

CuF2

2

936.6

[256]

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932.9

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ACCEPTED MANUSCRIPT Table 3 S 2p3/2 binding energies (BE) for S species. BE (eV)

FWHM (eV)

Reference

S2-

161.1-161.8

0.5-0.7

[95, 96, 124, 251]

S22-

162.3-162.4

0.5-0.7

[3, 25, 96]

Sn2-

163.0-163.9

1.1-1.3

S0

163.05-164.7

0.7-1.4

SO42-

168.0-169.0

0.9

S2O32- (attached S) 161.7-163.2

—

[249, 257, 260]

S2O32(central S)

167.4-167.8

—

[249, 257]

SO32-

166.4-166.5

Energy loss

163.9-164.5

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[3, 25, 96]

[25, 124, 141, 257, 258] [3, 25, 69, 249, 259]

—

[249, 257]

2.3-2.6

[3, 69, 71, 95, 96]

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S species

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ACCEPTED MANUSCRIPT Table 4 Electrochemical reactions at the chalcopyrite surface on progressive potential increase. Potential

Reaction

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(mV, SHE)



 
  
 

 
 2 

 



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< 740


  
 
 
  

 1 


 2   1 

 

740-940



  8
 

   2
  16  17  and/or

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  1   

 2  2 1      and



  3
 

 
 

  6  8  2

  13


 0.75
  1.25

 


  0.25
  26

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1060-1340

 28 

3


 , ,  
  12


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>1340

D


  4
 

 
  8  8 and

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    6


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Chalcopyrite leaching rate and mechanisms. Note: X is the oxidised fraction. Rate dependence

Ea (kJ mol-1)

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Leach media

Rate constant (k)

Mechanism

Reference

Diffusion through product layer

[31]

24

Diffusion through a porous product layer by shrinking core model

[46]

42

Diffusion

[26]

15-28

Mixed control model (chemical reaction control; diffusion of ions through product layer)

[30]

83

Mixed control model (surface reaction control; lixiviant diffusion through sulfur layer)

[132]

Surface reaction control

[135]

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T (°C)

1-17×10-4 min-1 HClO + HCl

Particle size


  


  1  1  

0.30-0.93×10-3 min-1

FeCl3+ H2SO4

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45 - 100

2
  1    1 3
  

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50 - 97

K2Cr2O7 + H2SO4

19.88

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10 - 40

[Fe3+]0.12

2.63-6.93×10-6 s-1 2.3-15.9×10-7 s-1

30 - 90

FeCl3 + HCl

70 - 90

NaNO3 + H2SO4

40 - 95

Fe3+ + HCl (H2SO4)

[H+]0.8[Fe3+]0.42

48±10

23 - 40

FeCl3 + HCl

[Cu-][Fe3+] [Cl-]2/[H+]

86.4

25 - 50

H2O2 + H2SO4

[H+]0.3

30 - 80

K2Cr2O7 +

[H+]0.8-0.92

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Particle size


   ln1  

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[34]

60

Surface reaction control by shrinking core model

[261]

48 - 54

Chemical reaction control

[16]

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FeCl3 + HCl

25 - 70

HCl at pH 3

[Fe3+]0.38

93 32±5 1  31     21  
" 22,423   exp    

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0.386-3.321×10-3 min-1 

125 - 150

O2 + H2SO4

 మ 1

25-75

H2O2 + H2SO4

25-75

Cu2+ + HCl




 1  1  [H2O2]1

Eh 750 mV SHE, pH 1




 1  1 


 1

H2SO4 55-85


 

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Fe2(SO4)3 + H2SO4


  1  1  0.45

3+

Fe

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Surface reaction control

[153]

Diffusion of hydrogen ions through a product layer by shrinking core model

[29]

93.5

Surface chemical reaction by shrinking core model

[41]

79.5

Surface reaction control by shrinking core model

[263]


 

30

Surface reaction control by shrinking core model

[42]

72

Surface chemical reaction and electrochemical control

[139]

ECu<10h = 21±5

Mixed control model by shrinking core model (diffusion or transport control; chemical reaction control)

   


 1  1

[262]


 

AC

50-92

22.4

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[H+]1

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Fe + HCl

PT ED

25 - 85

Surface chemical reaction

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70 - 100

PT

H2SO4


  

ECu>10h = 83±10

93

[131]

H2O2+ H2SO4

50-90

[bmim*]HSO4

Without Fe3+




Without Fe3+




[H2O2]1.45




+ 0.77

[H ]


 

EFe = 76 ± 10

ECu = 80 ± 10

 1  1  

EFe = 84±10

 1  1  





 

 1  1  


 

Chemical reaction control

39

Surface chemical reaction by shrinking core model

[105]

69.4

Electrochemical surface control

[138]

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* 1-butyl-3-methyl-imidazolium hydrogen sulfate

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Chemical reaction control

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Chemical reaction control

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 1  1  

MA

30-80




PT ED

Fe3+

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Table 6 Kinetic models and equations.
  1  1  

ଵൗ ଷ

Surface chemical reaction


  1  1  

ଵൗ ଷ

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Fluid film diffusion

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  1  2 3  1  

Diffusion through the product layer

ଵൗ ଶ ଷ

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ଶൗ ଷ

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Table 7

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Rest potential for sulfide minerals at pH 4 [208].

Chemical Formula

Rest Potential (V, SHE)

Pyrite

FeS2

0.66

Marcasite

FeS2

Chalcopyrite

CuFeS2

Sphalerite

(Zn,Fe)S

Covellite

CuS

Bornite

Cu5FeS4

Galena

PbS

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0.63

0.56

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0.46

Argentite

0.42 0.40

Ag2S

0.28

Sb2S3

0.12

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Stibnite

0.45

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ACCEPTED MANUSCRIPT Table 8 Steady-state electrode potentials (mV vs SHE) in 0.5 M sodium chloride (pH 10.5) for various electrodes [231].

Chalcopyrite

79

Stainless steel

-96

Hyper steel

-356

Mild steel

-396

Cast iron

-436

PT

259

354

139

RI

Platinum

O2 purged

-26

SC

N2 purged

-306

-326 -416

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Figure Captions

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Fig. 1 Schematic representation of the chalcopyrite unit cell structure, S - large yellow

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sphere, Cu – red sphere, Fe – blue sphere [50].

Fig. 2 Unit cell of CuFeS2 to show metal spins, Cu – red sphere, Fe – blue sphere. Redrawn

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from [86].

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Fig. 3. Cu concentration and % extraction in H2SO4 solutions at controlled pH 1, Eh 750 mV

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(SHE), 75 °C with and without the initial addition of 4 mM Fe3+.Redrawn from [131]. Fig. 4. (a) Bacterial attachment onto the mineral surface. Redrawn from [164]. (b) Two mode bacterial attachment onto the mineral surface. Redrawn from [160]. (c) Schematic of the

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Fig. 5. Schematics of the sulfide bio-oxidation mechanism. Redrawn from [204]. Fig. 6. Diagram of cathodic and anodic reactions for a chalcopyrite/pyrite galvanic cell. Fig. 7. Back scattered electron images of solids sampled during the Galvanox leaching experiments [214]. A: in progress. B: complete.

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Figure 1

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Figure 2

3.740

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3.714

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Figure 3

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Figure 4

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(b)

(c)

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Figure 5

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Figure 6

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Figure 7

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Graphical Abstract

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Highlights

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• The crystalline and magnetic structure of chalcopyrite is reviewed; • Surface speciation upon atmospheric and aqueous oxidation are compared; • Proposed mechanism of bio and chemical leaching are critically examined; • The effects of galvanic interactions are reviewed. • The remaining unknowns regarding leach mechanisms are summarised.

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