A novel hydrometallurgical treatment for the recovery of copper, zinc, lead and silver from bulk concentrates

1 A novel hydrometallurgical treatment for the recovery of copper, zinc, 1 lead and silver from bulk concentrates 2 Authors: Juan Lorenzo-Tallafigo*, Aurora Romero-García, Nieves Iglesias-González, 3 Alfonso Mazuelos, Rafael Romero, Francisco Carranza 4 Chemical Engineering Department, Universidad de Sevilla 5 Calle Profesor García González s/n 41012 Sevilla (Spain) 6 Telephone; +34 954 557 182 7 8 *Corresponding Author: [email protected] 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 2 Keywords 26 Bulk sulphide concentrate 27 Ferric leaching 28 Catalysed leaching 29 Hot brine leaching 30 Highlights 31 Copper, zinc, lead and silver can be benefited by a novel hydrometallurgical process 32 Direct catalytic leaching is not able to dissolve copper and zinc in a single step 33 Ferric leaching before catalytic leaching recovers more than 95% of copper and zinc 34 More than 95% of lead and 100% of catalyst are recovered by brine leaching 35 36 37 38 39 40 41 42 3 43 Abstract 44 Nowadays sulphide ores exploitation is undergoing some troubles, which are hindering 45 the treatment through traditional routes. Bulk flotation followed by a novel 46 hydrometallurgical process can dodge these difficulties. In this work, an integral 47 hydrometallurgical process consists of two ferric leaching steps, followed by a hot brine 48 leaching stage, is proposed to recover target metals from a bulk sulphide concentrate 49 (2.9% Cu, 7.4% Zn, 2.5% Pb, 67 ppm Ag and 37.2% Fe). In the first ferric leaching step, 50 sphalerite, galena and copper secondary sulphides are dissolved and, in the second 51 leaching step, a silver salt is added in order to catalyse chalcopyrite oxidation. If silver 52 salt is added at the beginning of the process, sphalerite passivation is observed, and 53 therefore zinc recovery is not possible. However, when catalytic leaching is performed 54 after a previous ferric leaching, copper and zinc recoveries higher than 95% are achieved. 55 The leached concentrate (0.3% Cu, 0.8% Zn, 3.3% Pb, 1438 ppm Ag, 38.0% Fe and 6.6% 56 S0), is treated by a hot brine leaching. When hot brine leaching is performed at high pulp 57 density, elemental sulphur removal is necessary to recover all silver added as a catalyst. 58 Extractions higher than 95% for Zn, Cu and Pb are achieved as well as the total recovery 59 of catalyst. The proposed process is silver surplus; therefore, this agent can be 60 recirculated. 61 62 63 64 65 66 67 68 69 70 71 4 1. Introduction 72 Currently, several changes are concerning to sulphide ores exploitation. Changes such as 73 the exhaustion of current ore deposits, the discovery of ore deposits that are more complex 74 to treat, the decrease in ore grade or more difficulties in environmental permits, which are 75 driving to a technological innovation necessity (Brierley, 2008; Norgate and Jahanshahi, 76 2010; Fomchenko and Muravyov, 2019). In this situation, (bio)-hydrometallurgical 77 technologies should play a key role to face these challenges and allow metals production 78 from polymetallic and low-grade ores, as well as mining waste. In this field, several 79 authors have studied different (bio)-hydrometallurgical processes to recover copper and 80 zinc from polymetallic sulphide ores, paying special attention to ferric leaching and 81 bioleaching (Lorenzo-Tallafigo et al., 2018; Fomchenko and Muravyov, 2019; 82 Fomchenko et al., 2019; Fomchenko and Muravyov, 2018; Carranza et al., 1997; Tipre 83 and Dave, 2004). Besides, other processes have been also proposed to recover target 84 metals from polymetallic ores, such as pressure leaching (with oxygen or sulphur 85 dioxide), roasting, chlorination, nitrate-sulphuric acid leaching, glycine leaching, sodium 86 meta-bisulphate leaching, or hydrogen peroxide leaching (Xu et al., 2011, Akcil and 87 Ciftci, 2003; Tarasov and Timoshenko, 2006; Cui et al., 2020; Sokic et al., 2017; Hara et 88 al., 2020; Sokic et al., 2019; Shin et al., 2019). 89 Pyro- and hydrometallurgical processes usually require a froth flotation step to produce 90 sulphide concentrates, which contain target metals. There are two flotation ways, 91 differential and bulk flotation. On the one hand, differential flotation obtains concentrates 92 of a single metal that later are usually processed by pyrometallurgical techniques 93 (traditional route). On the other hand, bulk flotation produces a unique concentrate which 94 contains all non-ferrous metals, being interesting new hydrometallurgical treatments to 95 benefit it. Bulk flotation has several advantages such as higher target metals recovery, 96 lower water and reagents consumption by far lower energetic supplies and minor particle 97 size requirements. Moreover, hydrometallurgy in contrast to pyrometallurgy can benefit 98 concentrates which contain impurities, such as arsenic or antimony (Carranza, 1985; 99 Carranza et al., 1997; Tipre et al., 1999). Bulk flotation followed by a hydrometallurgical 100 technology could be an attractive alternative compared to traditional routes, due to higher 101 recovery yields and to a greater flexibility regarding raw material composition could be 102 achieved (Frías et al., 2020). 103 5 Bulk sulphide concentrates are usually composed of pyrite (FeS2), sphalerite (ZnS), 104 galena (PbS) and chalcopyrite (CuFeS2). Chalcopyrite shows a refractory behaviour in 105 both bioleaching and leaching solutions, unlike sphalerite and galena (Lorenzo-Tallafigo 106 et al., 2018; Fomchenko and Muravyov, 2019; Fomchenko et al., 2019; Fomchenko and 107 Muravyov, 2018). Chalcopyrite is the most refractory of target sulphides in ferric sulphate 108 media, this is due to the formation of a passivating layer that hinders the diffusion of ferric 109 ion. For a time, the formation of a non-porous layer of elemental sulphur on chalcopyrite 110 surface was considered responsible of this passivation. Nevertheless, recent 111 electrochemical studies have observed that chalcopyrite passivation is because of the 112 formation of a non-stoichiometric deficient in metal sulphide film as a consequence of 113 the fact that iron is more reactive than copper, which remains on chalcopyrite surface 114 without reacting. This passivation process can be described according to r1 (Miller and 115 Portillo, 1979; Price and Warren, 1986; Hiroyoshi et al., 2002; Ghahreminezhad et al., 116 2015). 117 CuFeS2 → Cu1−xFe1−yS2+xCu2+ +yFe2+ +2(x+y)e−;y > x (r1) 118 Silver salts greatly improve copper dissolution from chalcopyrite (Carranza et al., 1997; 119 Cordoba et al., 2008; Hiroyoshi et al., 2002; Ghahreminezhad et al., 2015). The first 120 proposed model explains this catalytic activity by the formation of Ag2S on chalcopyrite 121 surface that avoids the refractoriness. Silver sulphide would be dissolved by ferric ions 122 producing silver ions (Ag+), starting the cycle again (r2-r3) (Miller and Portillo, 1979; 123 Price and Warren, 1986; Hiroyoshi et al., 2002). 124 CuFeS2+4Ag+ → Cu2+ +Fe2+ +2Ag2S (r2) 125 2Ag2S+4Fe3+ → 2S0+4Ag++4Fe2+ (r3) 126 According to Ghahreminezhad et al. (2015), silver ion is adsorbed onto chalcopyrite 127 surface forming silver sulphide, which in turn produces a sulphur vacancy in mineral 128 surface. The diffusion of silver and sulphur vacancies into passivated film improves the 129 diffusion of copper through it. The generated changes in the passivating layer enhances 130 the chalcopyrite dissolution kinetics. 131 The solid residue contains the silver, added as catalyst and contained in the concentrate, 132 and anglesite (PbSO4). As silver is an expensive metal, its recovery in a subsequent step 133 is a pivotal factor in the viability of the proposed process. Brine leaching is the most 134 6 widely recognised and extensively applied recovery method for silver and lead. In this 135 type of leaching, NaCl, MgCl2 or CaCl2 are used as source of chloride; FeCl3, H2O2, 136 NaClO or Ca(ClO)2 as oxidising agents; and HCl or H2SO4 to acidify leaching solutions 137 (Guo et al., 2010; Behnajady and Moghaddam, 2011). 138 In previous studies, different ways of treating of sulphide concentrates have been 139 postulated, paying special attention to ferric leaching and bioleaching (Carranza et al., 140 1993; Carranza et al., 1997; Fomchenko and Muravyov, 2018, Fomchenko and 141 Muravyov, 2019; Fomchenko et al., 2019; Romero et al., 2019). Most of the works 142 (Carranza et al., 1993; Fomchenko and Muravyov, 2018, Fomchenko and Muravyov, 143 2019; Fomchenko et al., 2019) propose a hydrometallurgical step, where zinc is dissolved, 144 and an additional pyrometallurgical step to benefit copper. Otherwise, Carranza et al. 145 (1997) proposed two ferric leaching steps, a ferric leaching followed by a silver-catalysed 146 ferric leaching, where zinc and copper are extracted from zinc-copper concentrates. These 147 ways could present several troubles, such as an inadequate extractive strategy due to the 148 loss of other target metals (e.g., lead or silver) and to a part of chalcopyrite because of its 149 dissolution under (bio)-leaching conditions (Dutrizac, 1981). On the other hand, the sale 150 of copper concentrates obtained could have drawbacks such as very low grade, and 151 dangerous transport to other facilities, thereby increasing the carbon footprint. 152 In view of above mentioned, an integral hydrometallurgical treatment is proposed in this 153 work: two ferric leaching steps (ferric leaching followed by a silver catalysed leaching), 154 where zinc and copper are extracted, and a hot brine leaching to recover lead and silver. 155 In a continuous process, ferric ion can be regenerated in situ by oxygen supply or in a 156 bioreactor where extremophilic bacteria would catalyse this reaction (Mazuelos et al., 157 2012; Carranza et al., 1997). Reactions (r2-r8) take place in ferric leaching steps 158 (Carranza et al., 2004; Hiroyoshi et al., 2002; Da Silva, 2004). According to r2-r7, pyrite 159 produces the greatest generation of ferrous ion in these solutions. The dissolution of silver 160 and lead in chloride medium is due to the progressive formation of soluble chloro- 161 complexes, according to r9-r10 (Behnajady and Moghaddam, 2014). 162 FeS 2+ 14Fe3+ + 8H2O → 15Fe2+ + 16H+ + 2SO4 2− (r4) 163 ZnS+2Fe3+ → Zn2+ +2Fe2+ + S0 (r5) 164 Cu2S+4Fe3+ → 2Cu2+ +4Fe2+ + S0 (r6) 165 7 PbS +2Fe3+ → Pb2+ +2Fe2+ + S0 (r7) 166 Pb2+ + SO4 2− → PbSO4 (r8) 167 Ag++4Cl− → AgCl4 3− (r9) 168 Pb2+ +4Cl− → PbCl4 2− (r10) 169 This proposed route can be located within the mining facilities, changing the production 170 model: from trade in concentrate to the production and sale of metals (“from mine to 171 metal”). Fig. 1 summarises the three different options, including our proposal, to recover 172 non-ferrous metals from polymetallic sulphides. 173 174 175 176 177 178 179 180 Fig. 1. Summary of different options to recover target metals from complex sulphide ores (red arrows = 181 traditional route; blue arrows = previous studies; green arrows = proposal of this study). 182 The aim of this work is to demonstrate that the integral hydrometallurgical treatment of a 183 bulk concentrate, by ferric leaching plus brine leaching, can result in high recoveries of 184 Pb, Ag, Zn and Cu. This approach would avoid high costs and environmental risks 185 associated to traditional routes for this natural resource. 186 187 8 2. Material and methods 188 2.1 Bulk concentrate 189 Table 1 shows the chemical and mineralogical composition of the bulk sulphide 190 concentrate used in this work. This concentrate was obtained by bulk flotation of 191 polymetallic sulphide ores (Las Cruces Mine, Spain). Non-ferrous metal contentis low 192 regarding traditional differential concentrates (Cu + Zn + Pb = 12.8%). Low-grade 193 concentrates are usual in bulk flotation to ensure metal recoveries higher than 90% from 194 polymetallic sulphide ores (Frías et al., 2020). Soluble copper in cyanide solution can be 195 associated with secondary copper sulphides, such as covellite or chalcocite (Fisher, 1994). 196 In this case, 37% of copper was dissolved in cyanide solution (10% NaCN, 30 min, 0.5% 197 p.d. and pH 11 at room temperature). From XRD analysis (Fig. 2) and microphotographs 198 (Fig. 3), different sulphide phases were identified. The most abundant mineral phase was 199 pyrite and the main non-ferrous sulphides found were sphalerite, galena, chalcopyrite and 200 chalcocite. Granulometric analysis determined a D80 of 32.1 µm, D50 of 11.7 µm and 201 D20 of 3.7 µm, being DX the maximum diameter for X% of particles. 202 203 204 205 206 207 208 209 210 211 212 213 214 215 216 217 218 219 9 Table 1: Chemical and mineralogical composition of bulk concentrate (Cu CN- corresponds to the copper 220 percentage extracted in cyanide media; Secondary copper is estimated as chalcocite; Identified phases are 221 marked with X; * mg/kg). 222 Element Cu Cu CN- Zn Pb Fe S Ag Ca As Sb Total Content (%) 2.9 1.07 (37%) 7.4 2.5 37.2 44.2 67* 0.6 0.67 0.14 95.61 Ore CuFeS2 ZnS PbS FeS2 Secondary copper CaSO4·2H2O XRD X X X X Microphotographs X X X X X Estimated (%) 5.6 11.0 2.9 76.2 1.3 2.6 99.6 223 224 225 226 227 228 229 16 3.2.2 Ferric leaching 370 In view of the above results, a ferric leaching as a previous step of catalytic leaching is 371 proposed. A set of ferric leaching tests was performed; time, pulp density and Fe2+/Fe3+ 372 ratio were the parameters tested. 373 Fig. 5 shows zinc and copper recoveries as a function of time. After 8h of ferric leaching, 374 zinc and copper dissolution are 95.5 and 47.2%, respectively. According to Carranza et 375 al. (2004), most of the dissolved copper must come from secondary sulphides, such as 376 chalcocite. On other hand, under these conditions chalcopyrite shows refractory 377 behaviour, reaching low dissolution levels (Mateos et al., 1987; Fomchenko and 378 Muravyov, 2018; Dutrizac, 1981; Córdoba et al., 2008). These results are in accordance 379 with copper mineralogical composition of the used concentrate (Table 1), where 37% of 380 copper can be directly related with secondary sulphides (Fisher, 1994). High dissolution 381 levels for zinc can be achieved in ferric sulphate media (Dutrizac, 2006; Souza et al., 382 2007b). 383 In both cases, at the beginning a fast leaching rate is observed followed by a decrease in 384 it. Chalcopyrite (Dutrizac, 1981; Dutrizac et al., 1969) and copper secondary sulphides 385 (Ruan et al., 2010; Chen and Lawson, 1991) leaching is controlled by diffusion 386 phaenomena. Sphalerite can show two kinetics regimes: a first stage where the chemical 387 reaction on the sphalerite surface is the rate limiting step and a second stage where 388 diffusion through a sulphur layer is the rate controlling step (Lorenzo-Tallafigo et al. 389 2018; Karimi et al., 2017; Souza et al., 2007; Weisener et al., 2003). 390 17 391 Fig. 5. Time effect on zinc and copper extraction (35 g/L ferric ion, 20 g/L sulphuric acid, 80 ºC, 1% pulp 392 density) (squares = zinc recovery; triangles = copper recovery; dashed lines = kinetic models). 393 In order to study the effect of pulp density from 1 to 25%, several tests were carried out 394 for 8 h. Fig. 6A presents the extraction of copper, zinc and iron as a function of pulp 395 density. Unlike zinc, whose recovery decreases notoriously when pulp density increases 396 above 10%, copper recovery does not show an important dependence with respect to pulp 397 density. Carranza et al. (2004) also observed similar results in copper recovery when pulp 398 density was increased in chalcopyrite and secondary copper sulphides 399 concentrates.Carranza et al. (1997) reported an important dependence of sphalerite 400 dissolution with respect pulp density (1-5% p.d.). However, in this work the last is only 401 observed when pulp density is drastically increased. This behaviour can be explained by 402 the low grade of non-ferrous metals in bulk concentrates. This low-grade allows working 403 at higher pulp densities with similar yields because ferric ion consumption per unit of 404 concentrate mass depends directly on the grade ofnon-ferrous sulphides. 405 Unlike zinc and copper sulphides, pyrite oxidation (iron dissolution) decreases drastically 406 when pulp density increases. At high pulp densities, iron extraction remains constant 407 around 15%. Since pyrite has a higher rest potential than the other sulphides, at higher 408 ferrous ion concentrations pyrite is slightly oxidised. 409 18 When pulp density is increased, ferrous ion generation (g ferrous ion/ kg bulk 410 concentration) is reduced due to the decrease of pyrite dissolution. Low generation of 411 ferrous ion improves the process efficiency due to the increase of leaching selectivity 412 (Fig. 6B). Fig. 6B shows ferrous ion generation data, both experimental and 413 stoichiometric (considering the dissolution of sphalerite, galena and copper as chalcocite), 414 finding 3% of pulp density the most appropriate to perform the following tests. 415 416 Fig. 6. Effect of pulp density on Cu, Zn and Fe extraction (A) and on ferrous ion generation (g ferrous 417 ion/ kg of concentrate) (B) (80 ºC, 8h, 35 g/L Fe3+ and 20 g/L H2SO4) (squares = Zn extraction; triangles 418 = Cu extraction; diamonds = Fe extraction; circles = experimental data; crosses = stoichiometric data). 419 420 421 19 In commercial leaching processes, a considerable fraction of dissolved iron is ferrous ion 422 (Dutrizac, 2006). The presence of ferrous ion could affect the leaching process decreasing 423 the recovery of target metals. Ferrous sulphate effect on ferric leaching has not been 424 extensively studied; while Dutrizac (2006) observed a negative effect on sphalerite 425 dissolution, Lorenzo-Tallafigo et al. (2018) did not find any decrease of non-ferrous 426 sulphide dissolution when ferrous ion concentration increased. In order to study this 427 phenomenon, a set of ferric leaching tests with different Fe2+/Fe3+ ratio (total iron 428 concentration = 35 g/L) was performed. 429 Fig. 7 shows the effect of Fe2+/Fe3+ ratio on zinc and copper extraction and on ferrous ion 430 generation (g Fe2+/ kg bulk concentrate). Zn and Cu extractions remain almost constant 431 when the concentration of ferrous ion increases (Fig. 7A). However, ferrous ion 432 generation decreases nearly three times when initial ferrous ion concentration was 20 g/L 433 (Fig. 7B). In other words, the decrease of ferric ion concentration has an similar effect to 434 increasing pulp density by increasing the leaching selectivity of target sulphides. 435 20 436 Fig.7. Zn and copper extraction (A) and experimental and stoichiometric ferrous ion generation (B) as a 437 function of Fe (II)/ Fe (III) ratio (g Fe2+/ kg bulk concentrate) (B) (80 ºC, 8h, 3% pulp density, 20 g/L 438 H2SO4 and 35 g/L. FeT = Fe (II) + Fe (III)) (triangles = Cu extraction; squares = Zn extraction; circles = 439 experimental ferrous ion generation; crosses = stoichiometric ferrous ion generation). 440 3.2.3 Catalysed ferric leaching 441 Once direct catalysed leaching and ferric leaching have been tested separately, zinc and 442 copper recovery in two leaching steps in series (ferric leaching followed by catalysed 443 ferric leaching) was studied. After ferric leaching, sphalerite, galena and secondary 444 copper sulphides are oxidised, and the leached concentrate is mainly composed of pyrite, 445 anglesite, chalcopyrite and elemental sulphur. 446 First, in order to study the whole process, a set of leaching tests with different catalysts 447 amount, from 0 to 4000 mg Ag/ kg concentrate were performed as follow: a previous 448 21 ferric leaching followed by a catalysed ferric leaching with a filtration between both steps. 449 Each leaching stage was carried out with fresh leaching liquor. Fig. 8 shows zinc and 450 copper recoveries obtained with different catalyst amounts. An improvement in copper 451 extraction can be observed when low silver amounts were added as a consequence of the 452 catalytic dissolution of chalcopyrite. Copper and zinc recoveries higher than 95% can be 453 reached when 1000 mg Ag/ kg concentrate is added. These results are in accordance with 454 Carranza et al. (1997), Mateos et al. (1987) and Cordoba et al. (2008a), where 455 chalcopyrite, contained in sphalerite-chalcopyrite and chalcopyrite concentrates, was 456 dissolved using silver as a catalyst. The previous ferric leaching dissolves non-ferrous 457 sulphides, except chalcopyrite, avoiding the passivation of them by catalyst addition. 458 459 460 Fig. 8. Zn and Cu extraction in two leaching steps as a function of silver added (3% pulp density, 8h ferric 461 leaching, 12h catalytic leaching, 35 g/L ferric ion, 20 g/L H2SO4 and 80 ºC) (triangles = copper 462 extraction; squares = zinc extraction). 463 464 3.2.3.1 Ferric leaching time effect 465 In order to test the effect of ferric leaching time in catalytic step, a set of leaching tests, 466 where the ferric leaching time was varied, was conducted. In these tests, both stages took 467 place in the same flask, once ferric leaching was finished, silver was added, starting the 468 catalysed leaching. Fig. 9 shows copper and zinc recovery achieved in these assays. Ferric 469 leaching time does not modify copper recovery, and the mean copper extraction is 93%. 470 22 In contrast, Zn extraction progressively enhances with reaction time in the first step, from 471 84 to 96%. Four hours of ferric leaching are enough to leach more than 90% of copper 472 and zinc. These results are in accordance with Carranza et al. (1997) and Carranza et al. 473 (2004), where the time of ferric leaching previous to catalysed leaching was also studied 474 in sphalerite-chalcopyrite and copper concentrates. Nevertheless, in the present study 475 lower ferric leaching times were required to reach high metal recoveries, as this bulk 476 concentrate contains a minor grade of target sulphides. In addition, Carranza et al. (1997) 477 and Carranza et al. (2004) carried out these tests with low ferric ion concentrations. 478 In these assays, catalytic step starts with a part of ferric ion reduced to ferrous ion, from 479 5.4 to 13.5 g/L Fe2+ (leaching liquor was not regenerated between both steps). This fact 480 does not affect the process efficiency, allowing the copper and zinc recovery in a single 481 stream. 482 483 484 Fig 9. Zn and Cu recovery in two steps as a function of ferric leaching time (12h catalytic leaching, 35g/L 485 ferric ion, 20 g/L H2SO4, 3% pulp density, 1000 mg Ag/ kg bulk concentrate, 80 ºC) (triangles = Cu 486 recovery; squares = Zn recovery). 487 488 489 490 491 492 23 3.2.3.2 Pulp density effect 493 A set of catalysed ferric leaching tests was conducted in order to study pulp density from 494 1 to 25 %. In order to observe a possible effect of pulp density, catalyst amount was 495 decreased to 500 mg Ag/kg concentrate. Several ferric leaching tests were performed in 496 a stirred tank reactor (STR) with a volume of 6L to generate enough amount of leached 497 concentrate to study the catalytic stage, the results are shown in supplementary data (Fig. 498 S1-3). In these tests (STR1-3), mean copper and zinc extractions of 44.4 and 92.2, 499 respectively, were achieved. 500 Pulp density was studied from 3 to 25%. The tests for 3 to 12% were conducted with solid 501 obtained in STR 1 and those between 12 and 25% were performed with the solid obtained 502 in STR 2. Fig. 10 shows the results obtained after two ferric leaching steps (ferrous ion 503 generation corresponds with the catalysed step). An increase in pulp density progressively 504 enhances copper recovery (Fig. 10A), nonetheless, when pulp density exceeds 15% 505 copper dissolution is reduced due to ferric ion exhaustion. The optimum pulp density to 506 dissolve chalcopyrite from this concentrate is in the range of 9-15%. This improvement 507 could be attributed to higher amounts of silver when pulp density increases, which could 508 be mobilised between chalcopyrite and pyrite surfaces to dissolve copper faster. The 509 increase of pulp density in catalytic leaching would allow decreasing the reactor volume 510 or increasing the leaching productivity. In addition, higher copper concentrations in PLS 511 would ease the subsequent concentration and purification steps. Zn recovery varies 512 slightly, from 91 to 97%. Fig. 10B shows ferrous ion generation, which decreases rapidly 513 as pulp density increases. At 15% of pulp density, the required ferric ion is reduced almost 514 3.5 times to reach better yields than with a pulp density of 3%. Like in ferric leaching, 515 increasing the pulp density decreases the generation of ferrous ion, bringing experimental 516 closer to stoichiometric ferrous ion generation (considering only the chalcopyrite 517 dissolution). 518 519 24 520 Fig. 10. Results obtained at different pulp densities in catalysed ferric leaching: Cu extraction (A) and 521 ferrous ion generation (g Fe2+/ kg bulk concentrate) (B) (500 mg Ag/ kg concentrate, 80 ºC, 35 g/L Fe3+, 522 20 g/L H2SO4, 12h of catalytic leaching) (triangles = copper extraction; circles = experimental ferrous ion 523 generation; crosses = stoichiometric ferrous ion generation). 524 3.2.3.3 Effect of catalyst amount 525 A third ferric leaching assay in a stirred tank reactor (test STR 3) was carried out to study 526 the effect of catalyst amount; results are presented in Fig. S3 and Table S1. This leached 527 concentrate was used to perform three catalysed ferric leaching tests in order to study the 528 effect of catalyst amount, from 125 to 500 mg Ag/ kg concentrate. The other conditions 529 were chosen according to the above results: 12% of pulp density, 80 ºC, 12h, 35 g/L ferric 530 ion, 20 g/L sulphuric acid and 180 rpm of orbital shaking. 531 25 Fig. 11 shows copper and zinc extractions with different catalyst amounts. Copper 532 extraction falls when silver amount is below 250 mg Ag/ kg concentrate, which decreases 533 drastically to 70% when only 125 mg Ag/ kg concentrate is added. Nevertheless, zinc 534 extraction remains almost constant. 535 Carranza et al. (1997) reported that 2000 mg Ag/ kg concentrate would be necessary to 536 dissolve chalcopyrite from a copper-zinc concentrate (43.7% chalcopyrite). Carranza et 537 al. (2004) studied the catalysed leaching in several copper concentrates, observing that 538 the catalyst requirement decreased when the chalcopyrite content fell in the used 539 concentrates; 1000 mg Ag/ kg concentrate was necessary to dissolve copper in a low- 540 grade copper concentrate (1.40% chalcopyrite). In both studies, pulp density was low, 541 between 1-3%. In this case, the chalcopyrite content in the leached concentrate is 542 estimated at 6.94%, obtaining copper recoveries higher than 90% with 250 mg Ag/ kg 543 concentrate and 12% of pulp density. This disagreement could be due to silver amount 544 effect also depends on the leaching conditions (e.g. pulp density) and the mineral 545 composition. 546 547 Fig. 11. Zn and Cu extraction as a function of catalyst amount (80 ºC, 12h catalysed leaching, 12% of 548 pulp density, 20 g/L H2SO4 and 35 g/L Fe3+) (triangles = Cu extraction; squares = Zn extraction). 549 550 551 552 553 32 In addition, since silver is a metal of great economic value, a 1.82% difference in its 678 extraction can be economically significant. In spite of Fe(III) did not result a significant 679 variable for silver extraction at 1% of pulp density, it becomes important at higher pulp 680 densities, as the low silver extractions in tests A25-A28 indicate. 681 On the other hand, the effect of Fe(III) on the extraction of Cu and Zn is observed. Copper 682 extractions for tests A17-A24 are on the order of 20% while Zn extractions of 50%. It 683 could be postulated that in chloride medium, Fe(III) is more active than in sulphate 684 medium, and pyrite is partially attacked resulting in the release of occluded copper and 685 zinc sulphides that are oxidised in these conditions. 686 3.3.2.3. Leaching time effect 687 The influence of time on chloride leaching was studied at 20% of pulp density under the 688 best conditions, that is, elemental sulphur layer removed, 5 g/L of Fe(III), 0.5M HCl and 689 250 g/L of NaCl. Table 8 shows the extractions of silver and lead and residue composition 690 at 15, 30, 60, 90, 120 and 150 minutes. 691 Table 8. Ag and Pb extractions at different times (20% dp, 5 g/L Fe(III), 0.5 M HCl and 250 g/L NaCl). 692 Extractions Residue composition Test Time (min) Ag (%) Pb (%) Ag (ppm) Pb (%) A29 15 70.2 98.8 509 0,04 A30 30 76.8 98.2 397 0.07 A31 60 82.6 98.8 302 0.05 A20 120 94.3 99.1 96 0.03 A32 150 94.3 98.7 98 0.05 693 In the studied sample, lead is found as anglesite as a result of the previous ferric leaching, 694 which is soluble in chloride medium. Lead extractions higher than 98% are obtained in 695 just 30 minutes. Lead remaining in solid must be occluded because lead extraction 696 remains almost constant. On the other hand, 120 minutes of leaching are necessary to 697 reach silver extraction of 94.3%; afterwards the extraction is not significantly increased. 698 It is worthy to note that an Ag extraction of 94.3% means both the total extraction of 699 added silver as a catalyst, and the 22% of silver contained initially in the bulk concentrate. 700 701 702 33 4. Conclusions 703 In order to recover copper and zinc, a leaching process consisting of a ferric leaching 704 followed by a catalysed ferric leaching is proposed. High recoveries of both metals are 705 not feasible in a single catalysed step. Sphalerite dissolution is passivated when silver is 706 added, reaching recoveries lower than 60% with 1000 mg Ag/kg concentrate. When a 707 previous ferric leaching is performed, higher zinc and copper recoveries can be achieved 708 (>95%). A minimum time of 4h of ferric leaching is necessary to achieve recoveries 709 higher than 90% of both metals. In the catalysed leaching, copper recoveries greater than 710 90% are achieved with low catalyst amounts (250 mg/kg concentrate) when pulp density 711 is increased. 712 Chloride leaching is a successful method to recover Pb and Ag from the catalysed ferric 713 leaching residue. In chloride leaching tests, lead extractions are between 95 and 99% in 714 the studied experimental conditions. The best conditions for silver extraction in chloride 715 medium at 1% of pulp density and 80°C are the absence of elemental sulphur, 0.5 M HCl 716 and 50 g/L NaCl. In these conditions, silver extraction is 96.62%. Within the limits of the 717 studied range, the best conditions for silver extraction at high pulp density and 80°C are 718 20% dp, 120 minutes as reaction time, removal of So, 5 g/L Fe (III), 0.5 M HCl and 250 719 g/L NaCl. Under these conditions, the extractions of Pb and Ag are 98.9% and 94.3% 720 respectively, which means the total extraction of silver added as a catalyst and the 22% 721 of silver contained in the bulk concentrate. 722 It can be concluded that silver-catalysed ferric leaching followed by brine leaching with 723 prior removal of elemental sulphur is a viable hydrometallurgical alternative for the 724 treatment of a Cu-Zn-Pb bulk concentrate. Extractions higher than 95% for Zn, Cu and 725 Pb are achieved as well as the total recovery of catalyst. The proposed process is silver 726 surplus, therefore this agent can be recirculated. 727 5. Acknowledgements 728 Thanks are given to CLC (project 2665/0124) and VI PPIT-US for funding this work. 729 6. References 730 Akcil, A., & Ciftci, H. (2003). Metals recovery from multimetal sulphide concentrates 731 (CuFeS2–PbS–ZnS): combination of thermal process and pressure leaching. 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