Valorisation of a flotation tailing by bioleaching and brine leaching, fostering environmental protection and sustainable development

1 Valorisation of a flotation tailing by bioleaching and brine leaching, fostering environmental protection and sustainable development A. Romero-García, N. Iglesias-González, R. Romero, Juan Lorenzo-Tallafigo, A. Mazuelos and F. Carranza ABSTRACT Flotation tailing is a problematic mining waste, because contains sulphides that exposed to oxidising conditions generate acidic drainage and the subsequent metal mobilisation. In this study, a flotation tailing produced within an integral process for the treatment of polymetallic sulphide ores is cleaned and valorised seeking a better use of natural resources and a lower environmental impact. In this work, bacterial leaching followed by brine leaching is postulated as an alternative to the flotation tailing treatment. Bioleaching destroys pyritic matrix (99% Fe) producing biogenic ferric that can be used as oxidising agent and recycling to a hydrometallurgical process. 80-90% Cu and Zn are dissolved and critical raw materials as Sb, In and Co are recovered. Brine leaching achieves Pb and Ag extractions greater than 96% and generates a solid residue mainly composed of quartz, in which gold that initially was in flotation tailing is concentrated. A clean and easily treatable for gold recovery final residue is obtained. In conclusion, pyrite matrix has been destroyed avoiding the further acid generation and reducing dramatically waste volume, metals have been valorised, and tailing hazardousness has been removed. Therefore, the proposed process is a sustainable alternative for flotation tailing management, reducing the environmental impact, the management costs, and generating income from valorisation of metals and the production of leaching agent. Keywords: flotation tailing, bioleaching, waste management, biogenic ferric, brine leaching 2 1. INTRODUCTION Froth flotation is the most common mineral processing process for sulphide ores concentration, obtaining concentrates of a single metal, suitable for a subsequent pyrometallurgical processing (Fuerstenau et al., 2007; Peleka et al., 2018). Associated with this kind of concentration method is the accumulation of billions of tons of waste heavy metal containing all over the world for the decades of operation of mining plants (Hansen et al., 2005; Xie et al., 2013). This waste, called flotation tailing, is often stored in large piles and can cause environmental problems due to acid mine drainage (AMD), generated by the oxidation and dissolution of sulphide minerals contained. Acid water with dissolved metals can produce soil and water contamination by the seepage or run-off. Due to the water-borne transport of dissolved metals and the wind-borne dispersal of fine particles, tailing appears as great source of toxicity that can cause problems in ecosystems and residential areas (Navarro et al., 2008; Xie et al., 2013; Kefeni et al., 2017). In fact, the inadequate management of mining waste has already caused serious accidents, such as the one in Aznalcóllar (Spain, 1998), the catastrophe of Aberfan (Wales, 1966) or more recently the dam failure in Brumadinho (Brazil, January 2019). Bascetin et al. (2016), evaluated the costs of the traditional tailing dam method as well as alternative methods such as surface paste tailing disposal and geotextile or cement methods, which are gradually increasing in the world. The analysis indicated that the unit costs were 2.25 $/t, 2.29 $/t, 7.39 $/t, and 2.79 $/t respectively. In said study, the surface paste disposal method is mentioned as the most appropriate and advantageous. The danger that involve the storage of flotation tailing, increasingly rigorous environmental regulations and the important value added that its content in metals supposes, require the proper search for technology which can successfully stabilise and recover metals from these materials. In addition, the concept of finite and non-renewable mineral resources should be integrated into future mining process. Thus, new alternatives to the current process in order to have both sound environmental protection and sustainable development are being sought. Bioleaching can be the solution for the treatment of flotation tailing where the shortage of high-grade minerals in valuables metals hinders the use of conventional methods and, therefore, requires being landfilled (Ahmadi et al., 2015; Muravyov and Fomchenko, 2018). This technique is based on the solubilisation of valuable metal cations from insoluble ores by biological complexation and oxidation processes (Rawlings, 2013; Johnson, 2014). There is a large variety of microorganisms in mining sites (Panda et al., 2015), nevertheless, Acidithiobacillus ferrooxidans is one of the most widely studied and used mesophilic, chemolithotrophic and acidophilic bacteria involved in metals solubilisation from mineral deposits (Valdés, 2008). Chemical reactions involved in leaching by action of these microorganisms can be generally 3 summarised in three basic steps (Jafari et al., 2018): oxidation–reduction reactions in the liquid phase catalysed by microorganisms, acid formation from organic or inorganic route, and extraction of metals from the sulphide matrix. The global bioleaching reaction of pyrite (FeS2), the main component in flotation tailing from polymetallic sulphides, is shown below (r1). 4 FeS2 + 15 O2 + 2 H2O → 2 Fe2(SO4)3 + 2 H2SO4 (r1) Many researches (Deveci et al., 2004; Antonijević et al., 2008; Mahmoud et al., 2017) have shown that during bioleaching, secondary copper sulphides, and zinc and lead sulphides are leached generating their corresponding sulphates and elemental sulphur. However, primary copper sulphides, such as chalcopyrite, are highly refractory and present challenges for bioleaching processes (Panda et al., 2013; Zhou et al., 2015). Lead bioleached from lead sulphide, precipitates in bioleaching conditions due to the low solubility of anglesite (PbSO4). On this matter, brine leaching is a suitable and extensively used method for its solubilisation (Frías et al., 2002; Ye et al., 2017). In the presence of chloride ions in aqueous medium, the formation of chlorocomplexes of lead gives ride to a decrease in the concentration of Pb2+ ions. The equilibrium of Pb2+ in contact with the solid phase, PbSO4, is displaced, and therefore, more solid passes into solution. In addition, this technique allows the solubilisation of silver. When the total concentration of chloride ions is greater than 10-2 M, the solubility of AgCl increases. The dissolution of AgC1 in chloride media is because of the progressive formation of soluble chlorocomplexes of silver (Guo et al., 2010; Bahram and Javad, 2011). This paper is focused on a flotation tailing produced in a process for the integral treatment of polymetallic sulphide ores (Fig. 1). In a first stage, a bulk flotation concentrate is obtained. Bulk flotation aims to recover jointly all the value metals in a sole concentrate for a later hydrometallurgical treatment. In this specific case, the treatment consists of a ferric leaching following by solvent extraction and electrodeposition to produce copper and zinc. Ferric leaching residue is subsequently subjected to a brine leaching for the solubilisation of silver and lead contained, which facilitates the recovery of both metals by cementation and precipitation respectively. This piece of work is focalised on flotation tailing valorisation through bioleaching and subsequent brine leaching, obtaining a non-polluting residue. In contrast with the usual flotation tailing management, storage, the proposed process contributes to a better use of natural resources, to the protection of the environment and to the circularity of these materials, which involves the sustainability sought in the mining industry of the future. 4 Fig. 1: Proposal for the treatment of polymetallic sulphide ores by hydrometallurgical way. 2. MATERIALS & METHODS 2.1. Flotation tailing Flotation tailing coming from the bulk flotation of polymetallic sulphide ores from Las Cruces Mine, an important ore deposit in The Iberian Pyrite Belt (Southwest of Europe), is studied. 2.2. Microorganisms and culture Two types of inoculums are used. The first one, called FNN-9K, is originally isolated from Rio Tinto Mine drainage waters which consisted mainly of Acidithiobacillus ferrooxidans, in addition of Leptospirillum ferrooxidans and some heterotrophic bacteria (mainly related with Acidophilium genus). This culture is routinely maintained by successive subculturing (20% inoculum and 31oC) on 9K medium (Silverman and Lundgren, 1959). The second inoculum used, 5 named FNN-Py, is an adaptation of FNN-9K culture to the flotation tailing at 5% pulp density (Pd) (subcultured under the same conditions in addition to orbital agitation of 180 rpm). 2.3. Analytical methods Two types of analysis procedures are carried out for solid characterisation. One of them lies in an acid digestion with aqua regia dissolving completely the residue to establish the total extraction. The other analysis is based on the dissolution of the solid in two steps. First, an attack with 6M HCl (30 minutes at room temperature and 180 rpm) to establish the part that has been bioleached but has been precipitated along with the undissolved pyrite matrix. The solid obtained is subjected to a second attack with aqua regia. In this second digestion, the part of the mineral that is not bioleached is determined. The concentrations of Fe, Cu, Zn, Pb, Ag and Ca are measured by atomic absorption spectrometry (2380 spectrophotometer, Perkin Elmer) and the elements Bi, Co, Cr, Ni, In, As, W, Sb, Se y Sn are analysed by ICP-OES (715 ICP-OES, Agilent). Gold is determined by fire assay. Sulphur is measured by gravimetric analysis. Granulometric analysis is carried out through Beckman Coulter laser granulometer (LS 13 320 MW model). Crystalline phases are identified by X-ray diffraction (XRD) using Cukα1=1.5406 nm and scanning from 2θ=3o to 2θ=90o at CITIUS (Centro de Investigación Tecnología e Innovación de la Universidad de Sevilla) facilities. 2.4. Calculations Solids weight loss (WL) after bioleaching or brine leaching process is calculated according to Eq1. 𝑊𝐿 =(𝐼𝑛𝑖𝑡𝑖𝑎𝑙 𝑚𝑎𝑠𝑠−𝐹𝑖𝑛𝑎𝑙 𝑚𝑎𝑠𝑠) · 100 𝐼𝑛𝑖𝑡𝑖𝑎𝑙 𝑚𝑎𝑠𝑠 Eq1 Fe, Cu and Zn extractions by solid (XS) and by liquor (XL) in bioleaching process are calculated according to Eq2 and Eq3 respectively. 𝑋𝑆=100−(𝐹𝑖𝑛𝑎𝑙 𝑏𝑖𝑜𝑙𝑒𝑎𝑐ℎ𝑖𝑛𝑔 𝑠𝑜𝑙𝑖𝑑 𝑚𝑎𝑠𝑠 𝑎𝑓𝑡𝑒𝑟 𝑎𝑡𝑡𝑎𝑐𝑘 𝑤𝑖𝑡ℎ 𝐻𝐶𝑙 · % 𝑚𝑒𝑡𝑎𝑙 𝑓𝑖𝑛𝑎𝑙 𝑔𝑟𝑎𝑑𝑒 𝐼𝑛𝑖𝑡𝑖𝑎𝑙 𝑠𝑜𝑙𝑖𝑑 𝑚𝑎𝑠𝑠 · % 𝑚𝑒𝑡𝑎𝑙 𝑖𝑛𝑖𝑡𝑖𝑎𝑙 𝑔𝑟𝑎𝑑𝑒 ·100) Eq2 𝑋𝐿=𝑔 𝑚𝑒𝑡𝑎𝑙 𝑖𝑛 𝑙𝑖𝑞𝑢𝑜𝑟 + 𝑔 𝑚𝑒𝑡𝑎𝑙 𝑎𝑐𝑐𝑢𝑚𝑢𝑙𝑎𝑡𝑒𝑑 + 𝑔 𝑚𝑒𝑡𝑎𝑙 𝑖𝑛 𝑗𝑎𝑟𝑜𝑠𝑖𝑡𝑒 𝐼𝑛𝑖𝑡𝑖𝑎𝑙 𝑚𝑎𝑠𝑠 · % 𝑚𝑒𝑡𝑎𝑙 𝑖𝑛𝑖𝑡𝑖𝑎𝑙 𝑔𝑟𝑎𝑑𝑒 ·100 Eq3 Total extractions (bioleaching and brine leaching) for the rest of metals are calculated according to Eq4 (by the solid) and Eq5 (by the liquor). 𝑋𝑆=100−( 𝐹𝑖𝑛𝑎𝑙 𝑠𝑜𝑙𝑖𝑑 𝑚𝑎𝑠𝑠 · % 𝑚𝑒𝑡𝑎𝑙 𝑓𝑖𝑛𝑎𝑙 𝑔𝑟𝑎𝑑𝑒 𝐼𝑛𝑖𝑡𝑖𝑎𝑙 𝑠𝑜𝑙𝑖𝑑 𝑚𝑎𝑠𝑠 · % 𝑚𝑒𝑡𝑎𝑙 𝑖𝑛𝑖𝑡𝑖𝑎𝑙 𝑔𝑟𝑎𝑑𝑒 ·100) Eq4 𝑋𝐿=𝑝𝑝𝑚 𝑚𝑒𝑡𝑎𝑙 𝑖𝑛 𝑓𝑖𝑛𝑎𝑙 𝑙𝑖𝑞𝑢𝑜𝑟 · 10000 𝑝𝑝𝑚 𝑚𝑒𝑡𝑎𝑙 𝑖𝑛 𝑖𝑛𝑖𝑡𝑖𝑎𝑙 𝑠𝑜𝑙𝑖𝑑 · 𝑃𝑑 (%) 𝑖𝑛𝑖𝑡𝑖𝑎𝑙 Eq5 6 2.5. Acid-basic accounting (ABA) test Acid-basic accounting test (ABA) is used to predict acid drainage generation since provides a preliminary evaluation of acid generation potential (AP) and neutralisation potential (NP). This balance is expressed as neutralising potential ratio (NPR = NP/AP). AP is calculated according to the amount of CaCO3 needed to neutralise a ton of sample if all the sulphur is oxidised to sulphuric acid (under the assumption that all sulphur is present as FeS2). NP is determined by the Sobek method (Sobek et al.,1978). 2.6. Bioleaching procedure 2.6.1. Previous bioleaching tests The performance of FNN-9K and FNN-Py inoculums are tested at 2, 5 and 10% of pulp density using the flotation tailing as substrate. All tests are carried out in sterile 250 mL- Erlenmeyer flasks at constant stirring rate (180 rpm) and temperature (31oC) in an orbital shaking incubator. Water is added daily to compensate for evaporation losses. Under the same conditions, control tests without bacteria at pH 1.5 are performed. 2.6.2. Bioleaching tests at constant pH Two bioleaching tests are carried out at constant pH. In bioleaching-1, the FNN-9K culture is used as inoculum while in bioleaching-2 the FNN-Py is utilised. In both cases, a cylindrical 7L- reactor, subjected to mechanical agitation and constant temperature fixed to 500 rpm and 31oC respectively, is used to perform the tests. A stirring rate of 500 rpm is sufficient to keep the solid in suspension. Air is injected through a 5 mm diameter spigot and pulp is kept aerated at constant flow rate of 3.5 L/min under standard conditions controlled by a rotameter. Pulp density is 5% and solution is composed of 80% of 9K medium without iron (II) sulphate, and of 20% of inoculum. pH is kept constant at 1.5 by the addition of 3 M sodium hydroxide solution (density = 1.098 g/mL) once or twice a day. Addition of NaOH solution is controlled by the mass. Loss of water by evaporation is replaced daily. At selected time intervals sample is taken, and both phases, solid and liquid, are analysed (section 2.3). Bioleaching is considered completed when the content of Fe in the liquor and in the soluble part in the attack with HCl of the solid from the sample taken, represent jointly around 95% of extraction of this metal. 2.7. Brine leaching procedure A study of the operating conditions for the solubilisation of silver and lead contained in the bioleaching-1 residue is performed. Tests are carried out in 250 mL-Erlenmeyer flasks at constant 7 temperature (30oC or 80oC), 5% pulp density, and stirring rate of 180 rpm in an orbital shaking incubator for two hours. Variables such as acidity, concentration of leaching agent, temperature (T) and Pd are studied. According to the best conditions for Ag and Pb extraction from the previous study, two tests at 12 and 20% pulp density are performed. 3. RESULTS AND DISCUSSION 3.1. Flotation tailing characterisation The qualitative analysis derived from the XRD pattern (Fig. 4A), indicates the presence of gypsum (CaSO4.2H2SO4), pyrite (FeS2), quartz (SiO2), chalcopyrite (CuFeS2). The chemical composition of flotation tailing is 46.32% S; 43.04% Fe; 0.81% Pb; 0,43% Zn; 0,37% Cu; 0,34% As; 0,19% Ca; 306 ppm In; 233 ppm Sb; 146 ppm Co; 106 ppm Bi; 24 ppm Ag; 21 ppm Sn; 18 ppm Cr; 17 ppm Ni; 0.9 ppm Au; less than 0.1 ppm Se and less than 0.05 ppm W. According to 46.3% of sulphur analysed, the AP value for the flotation tailing is 1432 kg CaCO3/t sample. The high value of AP and a NP=0 (absence of carbonates) indicates that the flotation tailing is a potential generator of acid drainage. The granulometric analysis indicates a D80 = 52 μm, that is, 80% by weight of material has a size equal to or lower than 52 μm. Moreover, D50 and D20 are 29 and 11 μm respectively. This fine granulometry, typical of flotation tailing, which gives it a high reactivity and, consequently, a high capacity to generate acid effluents. 3.2. Preliminary bioleaching tests Pyrite bioleaching (r1) produces sulphuric acid and Fe(III). In this sense, the variation of the concentration of H+ in the medium over time allows to describe qualitatively the evolution of cultures and the dissolution of pyrite matrix. Fig. 2 shows the evolution of pH for bioleaching tests of FNN-9K and FNN-Py cultures at 2, 5 and 10% pulp density. 8 Fig. 2: Kinetics comparison of bioleaching of FNN-9K and FNN-Py cultures at different pulp densities. The use of the inoculum FNN-Py involves to a decrease in the lag phase at 5 and 10% pulp density (Table 1) and, therefore, the beginning of the bioleaching in shorter times. In addition, this previous adaptation of the inoculum entails, unlike when the inoculum FNN-9K is used, that the concentration of substrate in the medium is not an influential factor in the inoculum lag time. The biofilm formed and adhered to the pyrite surface during this adaptation process protects the cells from the bioleaching medium, favouring the growth of the population (Bellenberg et al., 2015). Given that bacteria grow using pyrite as a substrate and that flotation tailing is mostly pyrite (Fig. 4A), the final Fe balance in the medium indicates the degree of attack suffered by the matrix of the bioleached mineral. However, high acidity conditions along with the presence of iron sulphides favour the formation of Fe (III) basic hydrosulphates (r2). 3 Fe3+ + X+ + 2 HSO4- + 6 H2O → X Fe3(SO4)2(OH)6 + 8 H+ (r2) In these compounds known as jarosites, with general formula XFe3(SO4)2(OH)6, where X = K+, Na+, Ag+, Pb2+, H3O+ or NH4+, a significant amount of values metals can be occluded (Ag or Pb) or immobilised by adsorption (Cu or Zn) (Daoud and Karamanev, 2006). The XRD analysis of bioleaching residue (Fig. 4B) shows potassium jarosite, as well as anglesite. Thus, to evaluate the fraction of tailing that has not been attacked by bacterial action, the previous dissolution of the jarosite by acid digestion is necessary. 9 Fe extractions obtained (Table 1) shows that the substrate depletion occurs at 2 and 5% pulp density. Bacterial growth is conditioned by the pH decrease in the medium. At pH values lower than 0.8-0.9, bacterial metabolic activity ceases. Thus, acidity achieved in the medium at 10% pulp density when inoculum is adapted, generates bacterial inhibition. On the other hand, there is no substrate depletion or bacterial inhibition due to the pH reached when operating at 10% pulp density with the inoculum FNN- 9K, in the time interval studied. Table 1. Lag time, final pH, weight loss, % of jarosite and total Fe extractions in bioleaching tests. Culture Pd (%) Lag time (h) WL (%) Final pH (%) Jarosite Total Fe XS (%) FNN-9K 2 47 69.59 1.17 54.65 97.52 FNN-9K 5 235 65.76 0.96 61.31 97.46 FNN-9K 10 235 39.25 1.13 23.18 72.07 FNN-Py 2 47 76.98 1.09 49.83 98.24 FNN-Py 5 47 68.09 0.93 72.82 99.36 FNN-Py 10 47 63.95 0.82 44.66 95.04 3.3. Bioleaching tests at constant pH The precipitation of Fe (III) depends on the acidity of the medium, therefore, an adequate pH control can be determinant to limit the extension of jarosite precipitation (Plumb et al., 2008; Kaksonen et al., 2014). In tests at constant pH, the neutralisation of the acid generated, besides the evolution of Fe in the medium (Fig. 3A), provides information about the bacterial growth and the degree of mineral ore attack. On the one hand, NaOH consumption is similar in both bioleaching tests at 5% pulp density, nevertheless, an earlier addition of NaOH evidences an earlier start of the attack to ore, producing shorter lag time when inoculum FNN-Py is used (Table 2). On the other hand, the more pronounced slope of the Fe extraction curve obtained in bioleaching-2 indicates a higher pyrite dissolution rate. Total Fe extractions close to 100% indicate exhaustion of the substrate, which, in turn, marks the end of the generation of acid and therefore cessation in the addition of NaOH solution. The generation of protons during the precipitation of jarosite (r2) and the complex chemistry of the system cause that the curves of NaOH added and % Fe extracted after the matrix dissolution (r1) are not coincident. 16 Fig. 5: Treatment of Cu-Zn-Pb polymetallic complex sulphides in pyrite matrix proposal. Economic and mass balance of the process proposed in Fig. 5 are found in a spreadsheet as supplementary data. In mass balance can be observed the value added of Cu and Zn recirculated and Pb, Ag and Au recovered, just as savings of leaching agent in the hydrometallurgical process. The economic balance was used in this study for economic assessment of the incorporation of the new stage of residue treatment. In order to estimate the profitability of the project, the Fixed Asset Turnover Ratio is used. This coefficient equals annual sales divided by the value of fixed assets (FA). The annual sales consider, in addition to the income from the sale of metals such as Cu and Zn, the savings that the amount of Fe3+ recirculated with the bioleaching liquor would imply in the annual cost, as well as the savings in the annual management of the flotation tailing currently stored in piles. These sales forecasts do not take into account the potential profit from remaining metals (Ag, Pb, Au and CRMs) contained in the bioleaching residue due to the imprecision in predicting the extra cost that the addition of said metals to the initial integral process would entail. For calculations of the net present value (NPV) and the internal rate of return (IRR), it is estimated an interest of 15%, that the fixed asset represents 85% of total investment and a project lifetime of 15 years. The applicable tax rate is the general rate for the industry (25%), although it should be highlighted as a favourable argument from the tax point of view that tax incentives apply to mining companies have not been considered. On the other hand, production costs (PC) are 17 calculated according to the Eq6. In this process, raw material (RM) to be used is a polluting residue that would be recycled, assuming no cost. Auxiliary services (AS) include energy required in the bioleaching reactor, in addition to extra costs in electrodeposition and solvent extractions stages for Cu and Zn recovery incorporated in the process. Direct labour (DL) costs have been estimated according to Wessel method. PC = RM costs + AS costs + 2.5·DL costs + 0.25·FA Eq6 Table 7. Data for the economic balance. Fixed Asset Turnover Ratio 0.85 Total mass bioleached (t) 5,250,000.00 Net Annual sales (€/year) 28,531,790.00 Cu and Zn recovery (€/year) 6,815,204.96 Savings in tailing management (€/year) 721,350.00 Saving Fe3+ (€/year) 20,995,235.04 Fixed assets (€) 33,566,811.76 Investment (€) 39,490,366.78 Working capital (€) 5,923,555.02 Amortization (€/year) 2,237,787.45 Production costs 14,603,648.21 RM costs (€/year) 0.00 AS costs (€/year) 210,670.75 DL costs (€/year) 2,400,509.81 Table 8. Economic balance. Years/€ 0 1-14 15 Fixed assets 33,566,811.76 Working capital 5,923,555.02 Sales revenue 28,531,790.00 28,531,790.00 Production costs 14,603,648.21 14,603,648.21 Amortization 2,237,787.45 2,237,787.45 Tax rate 2,922,588.58 2,922,588.58 BB 11,690,354.34 11,690,354.34 NB 8,767,765.75 8,767,765.75 NCF -39,490,366.78 11,005,553.20 11,005,553.20 Based on the assumption of a credit granted at 15% interest to 15 years by a bank, an IRR of 27.11% and a NPV of 24,863,175.95 € are obtained, so it is a profitable project. 18 4. CONCLUSIONS Finally, it is concluded that bioleaching followed by brine leaching is an alternative for the waste management of the flotation process. Bioleaching destroys pyritic matrix (99% Fe) and dissolves 80-90% Cu and Zn and critical raw materials as Sb, In and Co. Brine leaching achieves Pb and Ag extractions greater than 96% and generates a rich in gold clean solid residue. The incorporation of bioleaching stage in the proposed hydrometallurgical process allows for obtaining a positive balance between the economic growth, the environmental care and the social welfare, due to it drastically reduces its environmental impact and generates income. 5. ACKNOGLEMENTS Thanks are given to AGQ Mining & Bioenergy for funding this project (2664/0124) and supply the flotation tailing. 6. REFERENCES Ahmadi, A., Khezri, M., Abdollahzadeh, A. A., and Askari, M. (2015). Bioleaching of copper, nickel and cobalt from the low grade sulfidic tailing of Golgohar Iron Mine, Iran. Hydrometallurgy, 154, 1-8. Antonijević, M. M., Dimitrijević, M. D., Stevanović, Z. O., Serbula, S. M., and Bogdanovic, G. D. (2008). Investigation of the possibility of copper recovery from the flotation tailings by acid leaching. Journal of Hazardous Materials, 158(1), 23-34. Bahram, B. and Javad, M. (2011). Chloride leaching of lead and silver from refractory zinc plant residue. Research Journal of Chemistry and Environment, 15(2). Bascetin, A., Tuylu, S., Adıguzel, D. and Ozdemir, O. (2016). New technologies on mine process tailing disposal. J Geol Resour Eng, 2(2016), 63-72. Bellenberg, S., Barthen, R., Boretska, M., Zhang, R., Sand, W., and Vera, M. (2015). Manipulation of pyrite colonization and leaching by iron-oxidizing Acidithiobacillus species. Applied microbiology and biotechnology, 99(3), 1435-1449. ‘Critical Raw Materials’ (2018). Available at: https://ec.europa.eu/growth/sectors/ rawmaterials/specific-interest/critical_en (Accessed: 16 June 2018). Daoud, J. and Karamanev, D. (2006). Formation of jarosite during Fe2+ oxidation by Acidithiobacillus ferrooxidans. Minerals Engineering. Pergamon, 19(9), pp. 960–967. Deveci, H., Akcil, A., and Alp, I. (2004). Bioleaching of complex zinc sulphides using mesophilic and thermophilic bacteria: comparative importance of pH and iron. Hydrometallurgy, 19 73(3-4), 293-303. Dutrizac, J. E. (1992). The leaching of sulphide minerals in chloride media. Hydrometallurgy, 29(1-3), 1-45. Frıas, C., Dıaz, G., Ocaña, N., and Lozano, J. I. (2002). Silver, gold and lead recovery from bioleaching residues using the PLINT process. Minerals Engineering, 15(11), 877-878. Fuerstenau, M. C., Jameson, G. J., and Yoon, R. H. (2007). Froth flotation: a century of innovation. Society for Mining, Metallurgy, and Exploration Inc. Guo, Z. H., Pan, F. K., Xiao, X. Y., Zhang, L., and Jiang, K. Q. (2010). Optimization of brine leaching of metals from hydrometallurgical residue. Transactions of Nonferrous Metals Society of China, 20(10), 2000-2005. Hansen, H. K., Yianatos, J. B., and Ottosen, L. M. (2005). Speciation and leachability of copper in mine tailings from porphyry copper mining: influence of particle size. Chemosphere, 60(10), 1497-1503. Jafari, M., Abdollahi, H., Shafaei, S. Z., Gharabaghi, M., Jafari, H., Akcil, A., and Panda, S. (2018). Acidophilic bioleaching: a review on the process and effect of organic–inorganic reagents and materials on its efficiency. Mineral Processing and Extractive Metallurgy Review, pp. 1-21. Johnson, D. B. (2014). Biomining—biotechnologies for extracting and recovering metals from ores and waste materials. Current opinion in biotechnology, 30, 24-31. Kefeni, K. K., Msagati, T. A., and Mamba, B. B. (2017). Acid mine drainage: prevention, treatment options, and resource recovery: a review. Journal of Cleaner Production, 151, 475-493. Kaksonen, A. H., Morris, C., Rea, S., Li, J., Usher, K. M., McDonald, R. G., Hilario, F., Hosken, T., Jackson, M. and du Plessis, C. A. (2014). Biohydrometallurgical iron oxidation and precipitation: Part II—Jarosite precipitate characterisation and acid recovery by conversion to hematite. Hydrometallurgy, 147, 264-272. Lazer, M. J., Southwood, A. J. and Southwood, M. J. (1986). The release of refractory gold from sulphide minerals during bacterial leaching. Proceedings of International Conference on Gold, Vol. 2, Johannesburg: South Africa Institute of Mining and Metallurgy, pp. 287–297. Lorenzo-Tallafigo, J., Iglesias-González, N., Mazuelos, A., Romero, R. and Carranza, F. (2019). An alternative approach to recover lead, silver and gold from black gossan (polymetallic ore). Study of biological oxidation and lead recovery stages. Journal of Cleaner Production, 207, 510-521. 20 Mahmoud, A., Cézac, P., Hoadley, A. F. A. and Contamine, F. (2017). A review of sulfide minerals microbially assisted leaching in stirred tank reactors. International Biodeterioration and Biodegradation, 119, pp. 118–146. Muravyov, M. I. and Fomchenko, N. V. (2018). Biohydrometallurgical treatment of old flotation tailings of sulfide ores containing non-nonferrous metals and gold. Minerals Engineering, 122, 267-276. Navarro, M. C., Pérez-Sirvent, C., Martínez-Sánchez, M. J., Vidal, J., Tovar, P. J. and Bech, J. (2008). Abandoned mine sites as a source of contamination by heavy metals: a case study in a semi-arid zone. Journal of Geochemical exploration, 96(2-3), 183-193. Panda, S., Parhi, P. K., Nayak, B. D., Pradhan, N., Mohapatra, U. B. and Sukla, L. B. (2013). Two step meso-acidophilic bioleaching of chalcopyrite containing ball mill spillage and removal of the surface passivation layer. Bioresource technology, 130, 332-338. Panda, S., Akcil, A., Pradhan, N., and Deveci, H. (2015). Current scenario of chalcopyrite bioleaching: a review on the recent advances to its heap-leach technology. Bioresource technology, 196, pp. 694-706. Peleka, E. N., Gallios, G. P. and Matis, K. A. (2018). A perspective on flotation: a review. Journal of Chemical Technology & Biotechnology, 93(3), 615-623. Plumb, J. J., Muddle, R. and Franzmann, P. D. (2008). Effect of pH on rates of iron and sulfur oxidation by bioleaching organisms. Minerals engineering, 21(1), 76-82. Rawlings, D. E. (2013). Biomining: theory, microbes and industrial processes. Springer Science & Business Media. Sandberg, R. G. and Huiatt, J. L. (1986). Ferric chloride, thiourea and brine leach recovery of Ag, Au and Pb from complex sulfides. Journal of the Mineral, metals and Materials Society, 38(6), 18-22. Silverman, M. P. and Lundgren, D. G. (1959). Studies on the chemoautotrophic iron bacterium Ferrobacillus ferrooxidans. I. An improved medium and a harvesting procedure for securing high cell yields. Journal of bacteriology, 77(5), pp. 642–647. Sobek, A.A., Schuller, W.A., Freeman, J.R., Smith, R.M. (1978). Field laboratory methods applicable to overburdens and mine soils. US EPA 60012-78-054. Valdés, J., Pedroso, I., Quatrini, R., Dodson, R. J., Tettelin, H., Blake, R., A Eisen, J. and Holmes, D. S. (2008). Acidithiobacillus ferrooxidans metabolism: from genome sequence to 21 industrial applications. BioMed Central genomics, 9(1), 597. Xie, X. D., Min, X. B., Chai, L. Y., Tang, C. J., Liang, Y. J., Li, M., Ke, Y., Chen, J. and Wang, Y. (2013). Quantitative evaluation of environmental risks of flotation tailings from hydrothermal sulfidation–flotation process. Environmental Science and Pollution Research, 20(9), 6050-6058. Ye, M., Yan, P., Sun, S., Han, D., Xiao, X., Zheng, L., Huang, S., Chen, Y. and Zhuang, S. (2017). Bioleaching combined brine leaching of heavy metals from lead-zinc mine tailings: transformations during the leaching process. Chemosphere, 168, 1115-1125. Zhou, S., Gan, M., Zhu, J., Li, Q., Jie, S., Yang, B. and Liu, X. (2015). Catalytic effect of light illumination on bioleaching of chalcopyrite. Bioresource technology, 182, 345-352.