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Copper leaching presents an environmentally friendly alternative to traditional sulfide ore processing methods. This study investigates an efficient leaching process for copper concentrate, utilizing a solution of sulfuric acid (H2SO4) and potassium iodide (KI) in a chloride medium (NaCl), enhanced by hydrogen peroxide (H2O2) at room temperature. A significant aspect of this research was optimizing the KI concentration to minimize iodide sublimation into iodine gas (I2). Through the experimental design, the optimal dosages of reagents were determined, leading to maximized copper extraction of approximately 27% in 45 min of testing at room temperature. The results showed that it is possible to obtain a suitable pregnant leach solution (PLS) (i.e., in the range of 3 to 8 g/L of Cu) for treatment in available copper solvent extraction (SX) plants with a cost of less than 4.5 USD/t Cu, according to the economic analysis carried out. The results of this study determine the most effective operational conditions for leaching and ensure a suitable PLS for SX plants in a cost-effective and environmentally friendly manner. This approach could significantly contribute to more sustainable practices in the mining and processing of copper ores.

Low-temperature bioleaching is relevant to the recovery of metals in alpine mines, but its development has been constrained by low bioleaching rates at high pulp concentrations. To this end, the bioleaching effect of the microbial community after the domestication of pulp concentration at 6 °C was studied. Domestication improved the bioleaching rate of copper. Environmental scanning electron microscopy (ESEM), X-ray diffraction (XRD), and electrochemical measurements revealed that the domestication process aggravated the corrosion of the chalcopyrite surface by accelerating its dissolution reaction. High-throughput sequencing technology indicated that Acidithiobacillus spp., Leptospirillum spp., and Acidiphilium spp. were the major lineages of the domesticated microbial community. The analysis of the microbial community revealed that domestication changed the microbial structure, enhancing the adaptability of the microbial community to pulp concentrations and acidic conditions. This study uncovered the mechanism by which domestication enhanced the bioleaching efficiency of the microbial community at low temperatures.

The demand for non-ferrous copper metals has increased dramatically with the development of the global economy; accordingly, some refractory copper sulfide ores with low grade and their associated minerals are beginning to be utilized, making the flotation separation of copper concentrates exceptionally difficult, especially the separation of chalcopyrite and pyrite. In this paper, S-allyl-O, O′-dibutyl phosphorodithioate (ADTP) was synthesized by a one-pot method and used as a chalcopyrite collector in the flotation separation of chalcopyrite and pyrite. Flotation experiments results have shown that ADTP exhibits better selectivity and greater collecting power for chalcopyrite under neutral or weak base conditions. The 95% recovery of chalcopyrite can be achieved at pH 8.0 and 8.0 mg/L ADTP. From the analysis results of the contact angle, the SEM-EDS spectrogram, and elemental mapping, it was found that ADTP adsorbed uniformly on a chalcopyrite surface and made a significant contribution to the hydrophobicity of the surface. Confirmed by FTIR and XPS analysis, ADTP was able to form P–S–Cu bonds on a chalcopyrite surface, proving that it was adsorbed on the chalcopyrite surface in the form of chemisorption.

In the flotation separation process of a Cu-Mo-W polymetallic ore, the wastewater from the scheelite cleaning flowsheet contains large numbers of residual flocculants and metal ions, and the separation of chalcopyrite and molybdenite requires a large number of environmentally harmful depressants. Therefore, it is necessary to find new methods to reduce the environmental and cost pressures of wastewater treatment and the use of depressants. In this work, the flotation wastewater from the scheelite cleaning flowsheet for the separation of chalcopyrite and molybdenite by selective surface passivation was investigated for the first time. Flotations of single minerals and artificially mixed minerals with or without immersion pretreatment in the presence and absence of aeration were performed. The results showed that pulp pH had no effect on the flotation of either mineral, and a molybdenite recovery of 93.22% with a chalcopyrite recovery of 10.77% was achieved under the conditions of 10 days of immersion pretreatment with aeration, 350 mg/L of kerosene, and 100 mg/L of MIBC. By combining the electrochemical cyclic voltammetry analysis and characterization by XRD and SEM, the selective surface passivation mechanism of chalcopyrite was discussed, which could be due to the coverage of the insoluble oxidation products, especially jarosite. This work has simultaneously achieved the depressant-free flotation separation of molybdenite and chalcopyrite and the reuse of scheelite flotation wastewater, which is of great significance for environmental protection and cost saving.

The Kodaro–Udokan region is a huge Cu metallogenic province in Southern Siberia, one of the largest on Earth. It contains world-class copper sandstone-hosted Udokan (Cu reserves of 26.7 Mt) and PGE-Ni-Cu Chineysky deposits related to gabbro–anorthosite pluton (Cu—10 Mt; Fe-Ti-V, 30 Gt of ore). Furthermore, there are many small deposits of sulfide ores in sedimentary and igneous rocks in this region as well. For many decades, their genesis has been hotly debated. We studied the mineral composition and the sulfur isotopes in several deposits located at different levels of the stratigraphic sequence and in gabbro intruded in sandstones of the Udokan complex. The differences in ore compositions were found. The Burpala and Skvoznoy deposits consisting of the chalcocite–bornite association are characterized only by negative δ34S. The δ34S values for the Udokan deposits are mostly 34S data characterize the ores of the Chineysky and Luktursky intrusions. Two Cu sandstone-hosted deposits are characterized by complex ore composition, i.e., the Krasny deposit, comprising chalcopyrite–pyrrhotite ores, is enriched in Co, Ni, Bi, Sb, Mo, Pb, Zn, Se, Te, and U and has a wide range of δ34S = −8.1–+13.5‰, and the Pravoingamakitsky deposit (Basaltovy section), consisting of quartz–chalcopyrite veins, has high PGE contents in ores with δ34S = +2.9–+4.0‰. These deposits are located near the gabbro massifs, and it is supposed that their ore compositions were influenced by magmatic fluids. The general regularities of the localization of the deposits in rift zones, and the proximity of mineral and isotopic composition allow us to conclude that the main source of copper could be rocks of basic composition because only they contain high Cu contents. Fluids from deep zones could penetrate to the surface and form Cu sandstone-hosted deposits.

Oxidative leaching, as a basic step of the hydrometallurgical process of pure copper production from chalcopyrite, is a slow process in which mineral acids with strong oxidants addition are usually used as a leaching medium. It was found experimentally that the copper leaching from chalcopyrite in the H2SO4–H2O2–H2O system, in the presence of isopropanol (IPA) and under other conditions (H2O2 concentration, rate of mixing and temperature), takes place with satisfactory rate and efficiency. To quantify how much the change of these crucial variables affects the rate of the process, experimentally obtained kinetic curves (conversion over time) were analyzed using a Shrinking Core Model (SCM). The determined values of the copper leaching rate constants (kobs) confirmed the positive influence of increasing IPA and H2O2 concentrations as well as the temperature on the kinetics and efficiency of the leaching. The kinetic studies were also supported by using X-ray diffraction (XRD), 57Fe Mössbauer spectroscopy, scanning electron microscopy (SEM), and adsorption measurements. The positive influence of IPA was explained by its stabilizing role for iron compounds (hematite, magnetite, and pyrite), which are catalysts during the Cu dissolution, as well as H2O2 protection from decomposition during free radical reactions. Finally, the optimal conditions for efficient leaching, the rate-limiting step as well as the mechanism suggestion of the copper dissolution, were given.

Galvanic interactions present alternative strategies to achieve a more efficient Cu dissolution from CuFeS2. The present work studied the interaction between chalcopyrite–magnetite (CuFeS2-Fe3O4) in acidified ferric sulphate Fe2(SO4)3-H2SO4 at a solution pH of 1.8 and a temperature of 25 or 50 °C. The addition of Fe3O4 to CuFeS2 forms a galvanic couple, which positively impacts the dissolution of Cu. The results showed that the presence of Fe3O4 led to high and fast Cu dissolution rates and decreased significantly the activation energy, from 83 to 57 kJ/mol. In addition to that, the solid residues revealed that CuFeS2 dissolution produced intermediate Cu-S-rich phases: CuS, Cu2S and Cu5FeS4, which appeared to envelop CuFeS2, had no observable intermediate phase while in the presence of Fe3O4. The results showed that 94% of Cu could be recovered after 5 h of leaching at 50 °C at a Fe3O4/CuFeS2 ratio of 4:1 and a 460 mV Ag/AgCl solution potential. The findings of this study present an option for efficient Cu dissolution from CuFeS2 in ferric sulphate at atmospheric pressure.

This study presents the results of a two-stage autoclave processing of a copper–arsenic concentrate. Copper concentrate is an important raw material to produce copper and other metals. However, in some cases, the concentrate may contain increased amounts of arsenic, which makes further processing difficult. Therefore, the development of modern hydrometallurgical methods for processing copper concentrate with a high arsenic content is an urgent task, which could lead to the optimization of the raw material processing process and the improvement of the quality of the concentrate. It has been established that the optimal conditions for the sequential two-stage autoclave processing of copper–arsenic concentrate are: t = 220–225 °C, τoxidation = 20 min, τtot = 90 min, Po2 = 0.4 MPa, and L:S = 10:1, [H2SO4]initial = 40 g/dm3; in this case, 85% of zinc, 44% of iron, and 78% of arsenic, respectively, are extracted into the solution during both stages and the loss of copper was about 0.01%. This is explained by the fact that at the first stage (oxidation) of the autoclave processing of the copper–arsenic concentrate, copper, together with iron, leaches into the solution, and at the second stage (reduction), copper precipitates out of the solution in the form of chalcocite. Copper in the residue after autoclave leaching is in the form of Cu2S, iron is in the form of pyrite (FeS2), and lead is in the form of anglesite (PbSO4), respectively. The obtained micrographs and EDX mappings clearly show no iron arsenates. This confirms that at the oxidative stage of the developed process, arsenic, removed by 78%, remains in the solution. The remaining arsenic is associated with tennantite, indicating the effectiveness of the treatment process in removing arsenic from the copper–arsenic concentrate. A second important observation is the presence of pronounced areas of copper sulfides in the microphotos without iron and arsenic impurities. This confirms that copper is deposited as chalcocite during the reduction phase of the process, which is the desired result.

In this study, GX2 was applied as a new high-efficiency chalcopyrite depressant to selectively separate molybdenite from Cu-Mo sulfide ores. The flotation performance and its interaction mechanisms with chalcopyrite and molybdenite were investigated using single-mineral and artificial-mixed-ore flotation, contact angle measurements, zeta-potential measurements, infrared spectroscopy, and X-ray optoelectronic spectrum analysis. The results indicated that molybdenite could be selectively separated from chalcopyrite under the optimal flotation conditions of pH 9.0, 80 mg/L GX2, 20 mg/L kerosene, 10 mg/L MIBC, and a flotation time of 3 min, while the molybdenite and chalcopyrite recoveries were around 90% and 5%, respectively. It was confirmed that GX2 could save ten times the depressant dosage compared to that of the Na2S baseline to achieve a similar separation efficiency. The contact angle test, zeta potential, infrared spectrum, and XPS results show that GX2 may be adsorbed on the surface of chalcopyrite via chemical adsorption, and the hydrophilic substances formed change its hydrophobicity, whereas the effect on the floatability of molybdenite is small, resulting in the excellent separation efficiency of chalcopyrite and molybdenite via flotation.

Nowadays, research on promoting the dissolution of chalcopyrite is important. As a natural symbiotic mineral of chalcopyrite, arsenopyrite will have an impact on the dissolution of chalcopyrite. This paper shows the influence of arsenopyrite on the dissolution of chalcopyrite in an acidic culture medium. The leaching results showed that adding arsenopyrite increased the leaching concentration of copper by 332 mg/L. The residues showed a decrease in sulfur through X-ray diffraction analysis (XRD) and an increase in dissolution degree through scanning electron microscope (SEM). Electrochemical experiments have shown that the rest potential of arsenopyrite is higher than that of chalcopyrite, so there is a galvanic interaction, and the impact on chalcopyrite is greater than that of arsenopyrite. The polarization curve also proves this. Under the interaction of galvanic couples, the reduction of S0 production and the enhancement of Cu2+ release can promote the dissolution of chalcopyrite. In addition, X-ray photoelectron spectrometer (XPS) analysis under the action of galvanic coupling indicates that more SO42− is generated on the surface of chalcopyrite, replacing Sn2−/S0, and SEM shows a stronger corrosion morphology. All results confirm that the electrochemical effect between arsenopyrite and chalcopyrite promotes the dissolution of chalcopyrite in the acidic culture medium.

To address the issue of mediocre separation efficiency of depressants in the copper-lead separation process, this article synthesized a macromolecular organic depressant, polymaleic anhydride-ethylenediaminetetraacetic acid (PMA–EDTC), using a polycarboxylic macromolecule as the backbone and also introducing –N–(C=S)–S– as the solidophilic group and employed as a galena depressant. The structure of PMA–EDTC was characterized using Fourier transform infrared (FT-IR). The effect of PMA–EDTC on the floatability of galena and chalcopyrite was investigated through micro-flotation and Contact angle measurements. The experimental results demonstrated that PMA–EDTC exhibited selectivity inhibition towards galena rather than chalcopyrite across a wide pH range. At a dose of 8 mg/L, there was effective separation between galena and chalcopyrite with a separation coefficient of 24.17, effectively altering the floatability of galena while having little impact on the floatability of chalcopyrite. The selective inhibition behavior and adsorption mechanism of PMA–EDTC on galena and chalcopyrite were investigated using FT-IR, Zeta potential, and X-ray photoelectron spectroscopy (XPS). FT-IR and Zeta potential studies indicated that PMA–EDTC formed chemical adsorption on the surface of galena. XPS confirmed the model of chemical adsorption of PMA–EDTC on lead atoms in the galena surface. The results indicate that PMA–EDTC adsorbs on the surface of galena via its –(C=S)–S–group, forming a hydrophilic complex and achieving selective depression of lead and the cleaning flotation of copper.

In the Carajás Mineral Province, gossan formation and lateritization have produced numerous supergene orebodies at the expense of IOCG deposits and host rocks. The Alvo 118 deposit comprises massive and disseminated hypogene copper sulfides associated with gossan and mineralized saprolites. The hypogene reserves are 170 Mt, with 1% Cu and 0.3 ppm Au, while the supergenes are 55 Mt, comprised of 30% gossan and 70% saprolite, with 0.92% Cu and 0.03 ppm Au. The gossan includes goethite, malachite, cuprite, and libethenite zones. The saprolite comprises kaolinite, vermiculite, smectite, and relics of chlorite. In the hypogene mineralization, Ag, Te, Pb, Se, Bi, Au, In, Y, Sn, and U are mainly hosted by chalcopyrite and petzite, altaite, galena, uraninite, stannite, and cassiterite. In the gossan, Ag, Te, Pb, Se, and Bi are hosted by Cu minerals, while Au, In, Y, Sn, and U are associated with iron oxyhydroxides, in addition to Zn, As, Be, Ga, Ga, Mo, Ni, and Sc. As supporting information, δ65Cu values indicate that the gossan is immature and, at least partly, not affected by leaching. In the saprolite, Ga, Sc, Sn, V, Mn, Co, and Cr are associated with the iron oxyhydroxides, partially derived from the host rock weathering. The δ56Fe values indicate that hypogene low contribution of the hypogene mineralization to the saprolite iron content. The association of Al2O3, Hf, Zr, Th, TiO2, Ce, La, Ba, and Sr represents the geochemical signature of the host rocks, with dominant contributions from chlorites, while In, Y, Te, Pb, Bi, and Se are the main pathfinders of Cu mineralization.

Chalcopyrite (CuFeS2) particles, exposing anisotropic crystal planes during the grinding process, possess comprehensive surface properties that affect their leaching behaviors. In order to investigate the influence of anisotropic crystal planes on the leaching mechanisms, CuFeS2 particles with anisotropic crystal planes were produced by employing three-head laboratory grinder mill (TM), rod mill (RM), and ball mill (BM) and were then leached in a sulfuric acid solution at pH = 1. Based on the XRD, SEM, XPS, and simulation results, (112), (102), (312), (110), (116), (100), and (001) planes were mainly exposed on the CuFeS2 surface during the crushing and grinding process. In addition, fewer (112), (102), (312), and (110) planes but more (116), (100), and (001) planes were exposed on the CuFeS2 samples in the TM method than in the RM and BM methods. Since the leaching rates were in the order of (001) > (100) > (116) > (110) > (312) > (102) > (112) planes, the Cu extractions followed an order of TM > RM > BM. This study, therefore, provides an excellent theoretical basis for the effect of anisotropic crystal planes on CuFeS2 leaching, further improving the understanding of CuFeS2 leaching mechanisms.

In the present work the effect of alkaline sulfide leaching (ASL) on the extraction of copper and zinc from low-grade copper concentrate containing chalcopyrite, tennantite, sphalerite, and pyrite during batch and continuous bioleaching experiments was studied. It was demonstrated that ASL and further bioleaching may be a promising approach for treatment of copper–zinc concentrates containing tennantite as this approach allows increasing copper extraction degree in comparison to one-stage bioleaching by 1.6–2.3 times. Thus, ASL was shown to be effective for pretreatment of tennantite containing concentrates to improve bioleaching for copper extraction. At the same time, ASL led to decrease in zinc extraction by 1.4–1.5 times. Therefore, the development of combined hydrometallurgical processes including ASL and bioleaching for effective metal leaching requires further studies to avoid negative effect on zinc extraction.

Research on chalcopyrite leaching represents a great challenge, given its importance as one of the most abundant copper minerals and its significant role in global copper extraction. This study aimed to evaluate the effects of different parameters on chalcopyrite leaching by hydrogen peroxide as a strong oxidizing reagent in sulfuric acid solution. A series of leaching tests were carried out to investigate the effect of the solid/liquid ratio, stirring speed, temperature, oxidant and acid concentrations, and lixiviant dosing method on copper extraction from chalcopyrite concentrate. The catalytic decomposition of hydrogen peroxide occurred in the investigated leaching system, as reflected in the obtained metal extraction values. Copper extraction was increased in the first 60 min of the reaction, after which it essentially ceased. The maximum final copper extraction of 64.5% was attained with 3.0 mol/L H2O2 in 3.0 mol/L H2SO4 at a temperature of 40 °C after 120 min of reaction. Due to the catalytic decomposition of hydrogen peroxide in the examined leaching system, the leaching experiment was performed with the periodic addition of lixiviant at specific time intervals as well. The dissolution process was described by the first-order kinetics equation with an apparent activation energy of ~39 kJ/mol. Finally, XRD and SEM-EDS analyses were used to characterize the leached residue, and the results showed that the formation of elemental sulfur on the chalcopyrite surface affected the dissolution process.

The mechanism of electrochemical interaction between chalcopyrite and hexagonal pyrrhotite was analyzed via electrochemical interaction, copper ion concentration testing, and X-ray photoelectron spectroscopy (XPS) characterization. Besides, the effect of electrochemical interaction between the two minerals on the flotation separation was investigated using the mineral flotation tests, adsorption capacity tests, and a microcalorimetric test. Our research results showed that chalcopyrite had higher electrochemical activity than hexagonal pyrrhotite, and when the former acted as an anode during the electrochemical interaction of the two, the corrosion current density was three times higher than that when it acted alone, and the surface oxidation corrosion was intensified. At the same time, the interaction between the two minerals was accompanied by a large number of copper ions dissolved and adsorbed on the surface of the hexagonal pyrrhotite, so that adsorption of butyl xanthate intensified, adsorption increased, and flotation recovery increased by 5%–20%. However, owing to the increase in metal defects and the generation of hydrophilic sulfate, the surface of chalcopyrite hindered the adsorption of butyl xanthate on its surface, and the flotation recovery decreased by nearly 10% compared with that before the occurrence of the electrochemical interaction. This action also significantly weakened the inhibition effect of lime on hexagonal pyrrhotite and increased the difficulty of the flotation separation of the two minerals. The research results of this study provide theoretical guidance for the flotation separation of copper–sulfur ores containing pyrrhotite.

In this study, SDS is used to enhance the sulfuric acid leaching of chalcopyrite in aqueous and isopropanol media. The presence of SDS increased copper extraction into the solution in both solvents. However, it was the “isopropanol–sulfuric acid–SDS” system that proved to be particularly effective for copper recovery from chalcopyrite. The positive effect of SDS can be attributed to the reduction in the solution’s surface tension and the enhancement of mineral wetting. Additionally, the presence of SDS as a surfactant induces changes in the adsorption patterns of formed sulfur species on the mineral surface. SDS competes with sulfur for occupancy on the surface binding sites. This competitive interaction has the potential to diminish the formation of a substantial sulfur layer on the mineral surface. Under optimal conditions (isopropanol media, 2 M H2SO4, 65 °C, 120 min, 0.6 g/L SDS), copper recovery into the solution was 83%, and this is a considerable achievement for chalcopyrite leaching at ambient pressure in the absence of strong oxidizers.

The complete reaction of chalcopyrite at ≥220 °C under pressure oxidation conditions (10 or 20% w/w pulp density, PO2 700 kPa) is a clean process producing a residue consisting of hematite and un-reacted gangue minerals. However, when the process water contains chloride ions, covellite intermediate formation is significant and subsequently generates elemental sulphur that can persist for up to 60 min. Increasing the temperature to 230 °C reduces this time, although the dissolution of copper and the oxidation of sulphur still follows non-parallel reaction pathways. At 245 °C, the production of elemental sulphur in the presence of moderate chloride levels, 15 g/L, is no longer significant. The effects of other chemical additions (including enhancement of aluminium content) are also examined. Particular emphasis is given to the mineralogy of the leach residues and the deportment of iron in these residues to various phases that include hematite, basic ferric sulphate and natrojarosite. The residues are found to also contain a number of other intermediate phases in addition to covellite and sulphur, such as antlerite and clinoatacamite, depending upon the leach conditions employed.

To enhance the leaching of chalcopyrite concentrates, this study evaluated a new process for extracting copper using iodized solutions and sulfuric acid diluted in seawater without pressure or high temperatures. The work involved a leaching test carried out under various conditions by varying the concentrations of chloride ions, H2SO4, and an evenly distributed oxygen supply in an aeration system. It was demonstrated that Cl− ion addition could promote the chalcopyrite-leaching process. The leaching efficiency of copper reached 70% after 96 h. However, a chloride ion dosage excess can have the opposite effect on extraction, reducing copper recovery. XRD and SEM-EDS results showed that cuprous chloride (CuCl) was formed at high dosages (>0.5 M); meanwhile, at a lower dosage, elemental sulfur (S) was formed in the presence of sulfuric acid solution and seawater medium. In contrast, in an aerated system, surface roughness markedly increased due to continuous oxidation on the surface of the ore. This change in morphology and the high value of the redox potential, given by the aerated system and the acidic environment, allowed copper recovery of up to 70% after 96 h. The results showed that an aerated system is the most effective factor in chalcopyrite concentrate leaching.

The complete reaction of chalcopyrite at ≥220 °C under pressure oxidation conditions (10 or 20% w/w pulp density, PO2 700 kPa) is a clean, near complete process, yielding high copper extractions (~99%) in an acidic leach liquor composed of dissolved metal sulphates, when high-quality process water is employed. However, when the process water contains chloride ions, here 3–100 g/L, although the copper extraction rate is enhanced, complete oxidation of sulphur under batch processing conditions is delayed. Chloride addition, therefore, appears to favour an oxidation mechanism that liberates cupric ions and preferentially forms elemental sulphur over sulphate. This provides evidence for the decoupling of the copper extraction and sulphate formation reactions. Increasing the reaction temperature, here to 245 °C, increases the rate of sulphuric acid formation and decreases the iron concentration in the leach liquor. The study also examines the effects of various upfront acid and other salt additions upon copper extraction. Added sulphuric acid was shown to slow the reaction, whereas salts such a cupric chloride and sodium sulphate had small effects on the rate and extent of copper extraction.

In this study, chemical leaching, XRD, SEM, and XPS analyses were conducted to investigate the dissolution and passivation mechanisms of chalcopyrite under pressurized oxidative conditions in water. The chemical leaching results showed that the chalcopyrite could be dissolved by pressurized leaching without any acid addition, i.e., in an O2–H2O system, and the copper leaching rate reached 96.4% under the optimal conditions of 180 °C, 1.5 MPa, 900 rpm, and 90 min. The XRD, SEM, and XPS data suggested that a large proportion of the chalcopyrite dissolved in solution via the mineral phase transformation of CuFeS2→Cu2+ and CuFeS2→CuS→Cu2+, i.e., some of the chalcopyrite directly leached into solution as Cu2+, and some of it was first converted to CuS and then to Cu2+. The primary passivation layers during the chalcopyrite pressurized water leaching were hematite, pyrite, and covellite; however, none of them covered the un-leached mineral surface or inhibited chalcopyrite dissolution, as long as the agitation speed and leaching time were maintained over 700 rpm and 60 min, respectively. Finally, a model of chalcopyrite’s dissolution and passivation mechanism during pressurized water leaching was proposed.

Carrier flotation is a technique that can recover fine particles by using coarse carrier particles during the flotation process. In heterogeneous carrier flotation, coarse mineral particles of different minerals are used as carriers to recover fine mineral particles. By using Cu2+-treated pyrite particles as carriers, fine chalcopyrite particles recovery could be improved. However, a disadvantage of this heterogeneous carrier flotation is that it requires a post-flotation separation process to improve the grade of the final Cu concentrate. This study tested mechanical and chemical treatments to detach finely ground chalcopyrite (D50~3.5 µm) particles from Cu2+-treated coarse pyrite particles (−125 + 106 µm) after flotation. The results showed that the ultrasonic treatment was not effective to detach chalcopyrite fines from Cu2+-treated pyrite particles. However, acid treatment was effective to detach chalcopyrite fines from coarse pyrite particles. At pH 2, approximately 96% of chalcopyrite fines were detached from coarse Cu2+-treated pyrite particles. The acid treatment of flotation froth (mixture of chalcopyrite fines and Cu2+-treated pyrite particles) decomposed the collector KAX (potassium amyl xanthate) and dissolved the Cu precipitates adsorbed on the pyrite surface. This weakened the hydrophobic attraction force between the chalcopyrite fines and coarse pyrite particles, thus promoting the detachment of chalcopyrite fines from Cu2+-treated coarse pyrite particles.

Disodium carboxymethyl trithiocarbonate (DCMT) is considered to have the potential to replace sulfide and cyanide as a new chalcopyrite inhibitor. However, the effect of its application in the industrial field is often not ideal, mainly because the flotation involves solid, liquid and gas three-phase flotation systems, leading to many influencing factors, especially the chemical changes in pulp caused by the liquid phase. In order to promote the industrial application DCMT, we studied the effect of water quality in the flotation liquid phase on the inhibition of DCMT on chalcopyrite. Water quality generally involves the physical, chemical and biological characteristics of water bodies. The water for beneficiation belongs to industrial water, and the main indicator of its water quality is the water hardness level. Flotation and contact angle studies showed that higher water hardness levels suppressed chalcopyrite inhibition by DCMT. Infrared and Raman spectra revealed that the free CO32− and Ca2+ in water coordinated with the residual organic chains on the surface of the pretreated chalcopyrite and was subsequently adsorbed onto the chalcopyrite surface. Moreover, the addition of DCMT dislodged the captured CO32− and Ca2+. X-ray photoelectron spectroscopy indicated that DCMT could adsorb on the chalcopyrite surface and compete with the Ca2+. When Ca2+ was trapped on the chalcopyrite surface, there were fewer adsorption sites available to the DCMT, resulting in a lower inhibition capacity. Simultaneously, the presence of DCMT promoted the release of Ca2+ from the chalcopyrite surface. Therefore, the influence of water quality must be considered when designing a flotation reagent system, and the water hardness level should be reduced to optimize the flotation process.

One of the main problems in processing chalcopyrite ore with hydrometallurgical methods is its refractoriness, which is due to the formation of a layer that inhibits the contact of the ore with the leaching solution, thus reducing the dissolution rate. The main objective of this paper is to evaluate the leaching potential of iodide ions in copper extraction from chalcopyrite concentrate in an acidic seawater medium. Leaching tests were carried out in glass reactors stirred at 45 °C. Parameters such as iodide salt concentration and acidity were evaluated in ranges of 0–5000 ppm and 0–1.0 M, respectively. According to the results obtained, adding iodide ions to a medium acid enhances the leaching kinetics in the chalcopyrite concentrate, observing that it improves copper extraction at low concentrations of 100 ppm KI compared to high concentrations of 5000 ppm KI. As a result, part of the iodide required to oxidize copper tends to sublimate or is associated with other ions producing iodinated compounds such as CuI. Copper extraction reached 45% within the first 96 h, while at 216 h, it reached an extraction of close to 70% copper. The recovery rate improves at potentials between 600 and 650 mV, while at lower potentials, the copper extraction decreases. The mineral surface was analyzed using SEM/EDS and XRD analyses for the identification of precipitates on the surface, finding porous elemental sulfur and precipitated jarosite. An increase in iodide ions improves the leaching kinetics in the chalcopyrite concentrate, observing that it improves copper extraction at low concentrations of 100 ppm KI compared to high concentrations of 5000 ppm KI. As a result, part of the iodide required to oxidize copper tends to sublimate or is associated with other ions producing iodinated compounds such as CuI. Copper extraction reached 45% within the first 96 h, while at 216 h, it reached an extraction of close to 70% copper. The recovery rate improves at potentials between 600 and 650 mV, while at lower potentials, the copper extraction decreases. The mineral surface was analyzed using SEM/EDS and XRD analyses for the identification of precipitates on the surface, finding porous elemental sulfur and precipitated jarosite.

The primary purpose of this study was to produce an ore-based high-capacity supercapacitor electrode. For this, chalcopyrite ore was first leached with nitric acid, and then metal oxide synthesis was carried out immediately on nickel foam using a hydrothermal technique from the solution. Cauliflower-patterned CuFe2O4 with a wall thickness of about 23 nm was synthesized on the Ni foam surface, characterized by XRD, FTIR, XPS, SEM, and TEM investigations. The produced electrode also displayed a feature of a battery-like charge storage mechanism with a specific capacity of 525 mF cm−2 at 2 mA cm−2 current density, energy of 8.9 mWh cm−2, and a power density of 233 mW cm−2. Additionally, even after 1350 cycles, this electrode still performed at 109% of its original capacity. The performance of this finding is 255% higher than that of the CuFe2O4 in our earlier investigation; despite being pure, it performs far better than some of its equivalents in the literature. Obtaining such performance from an electrode made from ore indicates that the use of ore has a lot of potential for supercapacitor production and property improvement.

The diversity and activity of sulfate-reducing bacteria (SRB) in the camel gut remains largely unexplored. An abundant SRB community has been previously revealed in the feces of Bactrian camels (Camelus bactrianus). This study aims to combine the 16S rRNA gene profiling, sulfate reduction rate (SRR) measurement with a radioactive tracer, and targeted cultivation to shed light on SRB activity in the camel gut. Fresh feces of 55 domestic Bactrian camels grazing freely on semi-arid mountain pastures in the Kosh-Agach district of the Russian Altai area were analyzed. Feces were sampled in early winter at an ambient temperature of −15 °C, which prevented possible contamination. SRR values measured with a radioactive tracer in feces were relatively high and ranged from 0.018 to 0.168 nmol S cm−3 day−1. The 16S rRNA gene profiles revealed the presence of Gram-negative Desulfovibrionaceae and spore-forming Desulfotomaculaceae. Targeted isolation allowed us to obtain four pure culture isolates belonging to Desulfovibrio and Desulforamulus. An active SRB community may affect the iron and copper availability in the camel intestine due to metal ions precipitation in the form of sparingly soluble sulfides. The copper-iron sulfide, chalcopyrite (CuFeS2), was detected by X-ray diffraction in 36 out of 55 analyzed camel feces. In semi-arid areas, gypsum, like other evaporite sulfates, can be used as a solid-phase electron acceptor for sulfate reduction in the camel gastrointestinal tract.

Environmentally achieving the flotation separation of chalcopyrite and galena is always a difficult problem due to the similar floatability of the two minerals. To conquer the problem, maleic acid–acrylic acid copolymer (MA/AA), an eco-friendly reagent, was applied as a potential depressant for flotation separation of chalcopyrite from galena for the first time. Single-mineral flotation tests exhibit that MA/AA has a much better depression ability and selectivity than those of traditional galena depressants (Na2S and K2Cr2O7), which can inhibit the floatability of galena well but barely affects the flotation behavior of chalcopyrite in a wide pH range (7–11). A satisfying flotation separation effect of artificially mixed galena and chalcopyrite was realized by using MA/AA as a depressant. Based on a series of measurements including zeta potential, XPS, and contact angle, it appears that MA/AA was much more inclined to be chemically adsorbed on the surface of galena than that of chalcopyrite, which restrains the further adsorption of collectors on galena. In contrast, for chalcopyrite, the low adsorption of MA/AA hardly affects the further adsorption of collectors. According to these findings, MA/AA is considered to be potentially applicable as an effective and eco-friendly depressant in the industrial flotation separation of chalcopyrite and galena.

Despite chalcopyrite (CuFeS2) being one of the oldest known copper ores, it exhibits various properties that are still the subject of debate. For example, the relative concentrations of the ionic states of Fe and Cu in CuFeS2 can vary significantly between different studies. The presence of a plasmon-like resonance in the visible absorption spectrum of CuFeS2 nanocrystals has driven a renewed interest in this material over recent years. The successful synthesis of CuAl1−pFepS2 nanocrystals that exhibit a similar optical resonance has recently been demonstrated in the literature. In this study, we use density functional theory to investigate Fe substitution in CuAlS2 and find that the formation energy of neutral [FeCu]2++[CuAl]2− defect complexes is comparable to [FeAl]0 antisites when p≥0.5. Analysis of electron density and density of states reveals that charge transfer within these defect complexes leads to the formation of local Cu2+/Fe2+ ionic states that have previously been associated with the optical resonance in the visible absorption of CuFeS2. Finally, we comment on the nature of the optical resonance in CuAl1−pFepS2 in light of our results and discuss the potential for tuning the optical properties of similar systems.

Chalcopyrite is well known as being refractory to conventional leaching approaches at atmospheric pressure. The current study investigated a hybrid approach using aqueous ammonia as a pH modifier for glycine-based lixiviant systems to leach copper from chalcopyrite while maintaining surface refreshment using ceramic media at room temperature. The glycine–ammonia system exhibited significantly better copper extraction than the traditional glycine–NaOH system. A copper extraction of 91.5% was achieved after 72 h of leaching by using 0.71 M ammonia, at a Gly:Cu molar ratio of 4:1, by using a solid content of 1%, with a ceramic media: solid ratio of 3:5 and at ambient temperature. Adding ceramic beads during leaching led to a breakup of particles and a refreshment of particles’ surfaces that significantly improved the copper extraction. At a solid content of 10%, oxygen is essential for leaching; a copper extraction of 95.4% was achieved with oxygen injection, while only 33.4% copper extraction was obtained without introducing oxygen. A kinetic analysis indicates that the leaching rate is limited by combined liquid film diffusion and diffusion through the product layer. A conceptual flowsheet is proposed, where chalcopyrite concentrate can be extracted by a leaching–grinding process and copper can be recovered by a solvent extraction–electrowinning circuit.

The nature of tennantite (Cu12As4S13), chalcopyrite (CuFeS2) and sphalerite (ZnS) particles’ mixture dissolution in nitric acid (HNO3) media was investigated in this study. The effects of temperature (323–368 K), HNO3 (1–8 mol/L) and Fe3+ (0.009–0.036 mol/L) concentrations, reaction time (0–60 min) and pyrite (FeS2) additive (0.5/1–2/1; FeS2/sulf.conc.) on the conversion of the minerals were evaluated. It has been experimentally shown that the dissolution of the mixture under optimal conditions (>353 K; 6 mol/L HNO3; FeS2/synt. conc = 1/1) allows Cu12As4S13, CuFeS2 and ZnS conversion to exceed 90%. The shrinking core model (SCM) was applied for describing the kinetics of the conversion processes. The values of Ea were calculated as 28.8, 33.7 and 53.7 kJ/mol, respectively, for Cu12As4S13, CuFeS2 and ZnS. Orders of the reactions with respect to each reactant were calculated and the kinetic equations were derived to describe the dissolution rate of the minerals. It was found that the interaction between HNO3 solution and Cu12As4S13, CuFeS2 and ZnS under the conditions investigated in this are of a diffusion-controlled nature. Additionally, the roles of Fe(III) in the initial solution and FeS2 in the initial pulp as catalysts were studied. The results indicated that the increase in Fe3+ concentration significantly accelerates the dissolution of the mixture, while the addition of FeS2 forms a galvanic coupling between FeS2, and Cu12As4S13 and CuFeS2, which also accelerates the reaction rate. The results of the study are considered useful in developing a hydrometallurgical process for polymetallic sulfide raw materials treatment.

Separation of chalcopyrite from molybdenite is currently mainly carried out by flotation, but this process is costly because of the extensive use of inhibitors. This study briefly describes a 7.0T/100CGC low-temperature superconducting magnetic separator and discusses its separation principle as well as the effect of magnetic induction on chalcopyrite separation from molybdenite. A molybdenum (Mo) concentrate assaying 6.00% copper (Cu) and 19.01% Mo was magnetically sorted using a diamond-shaped steel rod medium mesh at a feed concentration of 20% and a pulp flow rate of 5 L/min from a Cu-Mo flotation concentrate with 88% of particles smaller than 23 μm using the separator. A Mo concentrate assaying 0.46% Cu and 16.28% Mo was finally obtained with a roughing (1.3 T)-cleaning (5 T) superconducting magnetic separation process. Similarly, the superconducting magnetic separator was performed to separate a Cu-Mo bulk flotation concentrate, and produced Cu concentrate assaying 19.64% Cu and 0.03% Mo from the bulk concentrate assaying 18.52% Cu and 0.39% Mo with a particle size of less than 0.074 mm. At a magnetic induction of 7 T, a pulp concentration of 20% and a feed velocity of 5 L/min, the grade and recovery of Cu in the magnetic product were 19.64% and 81.59%, respectively, whereas the grade and recovery of Mo in the non-magnetic product were 1.52% and 90.07%, respectively. Superconducting magnetic separation has potential applications for removing Cu from Mo concentrates, and separating Cu and Mo from Cu-Mo bulk flotation concentrates.

Bioleaching may be effectively used to extract nonferrous metals from sulfide ores and concentrates. At the same time, some minerals are refractory and their bioleaching rate is often comparatively low that does not allow the required metal extraction rate to be achieved. In the present work, we studied the two-stage process, which included stages of biological and chemical leaching, to improve copper extraction from low grade Cu–Zn sulfide concentrate containing chalcopyrite, tennantite, pyrite, and sphalerite. Bioleaching was conducted in the continuous mode in three laboratory scale reactors connected in series. The pulp density was 10% and the residence time was 7 days. The temperature was 40 °C in the 1st reactor and 50 °C in the 2nd and 3rd reactors. Bioleaching allowed the extraction of 29.5 and 78% of Cu and Zn, respectively. The solid bioleach residue obtained was then treated for additional Cu and Zn recovery using high temperature leaching at 90 °C for 25 h. The liquid phase of the bioleaching pulp contained Fe3+ ions, which is the strong oxidant, and the leach solution was supplemented with NaCl. In the presence of the maximal NaCl concentration (1 M), Cu and Zn extraction reached 48 and 84%. Thus, two-stage leaching may allow to increase bioleaching efficiency and may be used to improve the bioleaching rate of refractory minerals, such as chalcopyrite.

Clean energy infrastructure depends on chalcopyrite: the mineral that contains 70% of the world’s copper reserves, as well as a range of precious and critical metals. Smelting is the only commercially viable route to process chalcopyrite, where the oxygen-rich environment dictates the distribution of impurities and numerous upstream and downstream unit operations to manage noxious gases and by-products. However, unique opportunities to address urgent challenges faced by the copper industry arise by excluding oxygen and processing chalcopyrite in the native sulfide regime. Through electrochemical experiments and thermodynamic analysis, gaseous sulfur and electrochemical reduction in a molten sulfide electrolyte are shown to be effective levers to selectively extract the elements in chalcopyrite for the first time. We present a new process flow to supply the increasing demand for copper and byproduct metals using electricity and an inert anode, while decoupling metal production from fugitive gas emissions and oxidized by-products.

Porphyry copper deposits are important sources of copper and typically processed by flotation to produce copper concentrates. As mining areas become deeper, the amounts of impurities, such as sphalerite, can be increased in copper ores, so the appropriate depression of sphalerite floatability should be achieved to obtain saleable copper concentrates. In this study, the flotation behaviors of chalcopyrite and sphalerite in model samples mimicking copper ores with high Cu/Zn ratios (i.e., the ratio of chalcopyrite/sphalerite = 13:1) were investigated with zinc sulfate as a depressant for sphalerite. In addition, the effect of pyrite—a major gangue mineral in copper ores—on the depression of sphalerite floatability with zinc sulfate was examined. When sphalerite and chalcopyrite coexisted, the floatability of the former was effectively depressed by zinc sulfate (Zn recovery: 2+ due to the galvanic interaction with chalcopyrite, which resulted in the elimination of the effectiveness of zinc sulfate in depressing sphalerite floatability (Zn recovery: >90%). Despite the presence of much higher amounts of chalcopyrite and pyrite than sphalerite, the application of nitrogen (N2) gas limiting the galvanic interaction between pyrite and chalcopyrite by reducing the dissolved oxygen (DO) concentration in the system effectively depressed the floatability of sphalerite (Zn recovery:

This study aims to advance the knowledge of using carbon materials as catalysts in the oxidation of chalcopyrite. For this, two different materials (a commercial activated carbon (CC) and commercial biochar (BC)) were added to chalcopyrite ore (CPY) at three weight ratios (1:1, 1:0.5, and 1:0.25). Mixtures were treated with sulfuric/ferric solution for 96 h at 90 °C. Experimental results showed that extraction of copper from CPY was around 36%, increasing to higher than 90% with the addition of CC or BC at the proper ratio. The best result (99.1% Cu extraction) was obtained using a 1:1 ratio of CPY:CC. Analysis of solid residues shows that CC, with a high surface area, adsorbs sulfur onto its surface, limiting elemental sulfur formation. Additionally, the treatment of CPY in the CC’s presence transforms the chalcopyrite into CuS. Sulfur adsorption or CuS formation was not observed after the leaching of chalcopyrite with BC. However, the addition of BC to CPY at a ratio of 1:0.25 also increased the extraction of copper to 91.1%. Two carbon materials were oxidized after treatment with a sulfuric/ferric solution, and BC probably displayed catalytic properties in the leaching medium.

The surface oxidation of chalcopyrite is one of the most important factors affecting its flotation performance. In this study, a critical oxidation degree is proposed to define “slight” and “significant” oxidation in terms of surface species and chalcopyrite flotation recovery. Slight oxidation enhanced chalcopyrite hydrophobicity, but significant oxidation reduced its recovery apparently. Microthermokinetic measurements indicated that the apparent activation energy (Ea) of chalcopyrite oxidation was reduced from around 173 kJ·mol−1 to 163 kJ·mol−1 when the reaction changed from slight oxidation to significant oxidation when applying H2O2. The surface oxidation degree was defined as the ratio of hydrophilic species to hydrophobic species. The highest recovery (94.8%) and contact angle (93°) were achieved at a concentration of 0.1 vol.% H2O2, with the lowest oxidation degree of 0.388 being observed. The oxidation degree was correlated to the flotation recovery, with a quantitative relationship (y = −298.81x + 213.05, y and x represent flotation recovery and oxidation degree, respectively, 0.388 ≤ x ≤ 0.618) being established, thereby giving a guideline to better manage chalcopyrite flotation by controlling its surface oxidation and SBX adsorption on chalcopyrite surfaces.

The selective separation of molybdenite from copper sulfide concentrate in flotation process is realized using sodium hydrosulfide (NaHS) to depress the chalcopyrite and permit only the flotation of the molybdenite. However, this reagent is a highly toxic and flammable gas. The objective of this research was to study the feasible application of commercial lignosulfonates (LSs) in the separation by froth flotation process of molybdenite and chalcopyrite in seawater to present a novel application for LSs, as well as an alternative reagent to sodium hydrosulfide (NaHS). To achieve this, microflotation, absorbance tests and zeta potential measures were performed at pH 8 in seawater and 0.01 M NaCl. The results obtained in this study showed that it is possible to achieve selective separation of copper and molybdenum in both aqueous media due to high depressant effect of molybdenite and low depression of chalcopyrite in microflotation tests at 10 ppm of LSs, when the collector, PAX, is added prior to the addition of LSs. Absorbance study and zeta potential measurements showed that LSs adhere more to the molybdenite surface in seawater than in freshwater. This is related to the high ionic charge of the media that influences a greater interaction of LSs with the mineral surface.

This work introduces the CoarseAIR™, a novel system utilizing a three-phase fluidized bed and a system of inclined channels to facilitate coarse particle flotation and internal size classification. Internal classification in the CoarseAIR™ was investigated in a series of continuous steady-state experiments at different inclined channel spacings. For each experimental series, a low-grade chalcopyrite ore was milled to a top size of 0.53 mm and methodically prepared to generate a consistent feed. The air rate to the system was adjusted to determine the impact of the gas flux on coarse particle flotation and overall system performance, with a focus on maximizing both copper recovery and coarse gangue rejection. A new feed preparation protocol led to low variability in the state of the feed, and in turn strong closure in the material balance. Hence, clear conclusions were drawn due to the high-quality datasets. Inclined channel spacings of z = 6 and z = 9 mm were used. The z = 9 mm spacing produced more favourable copper recovery and gangue rejection. Higher gas fluxes of 0.30 to 0.45 cm/s had a measurable, adverse effect on the recovery of the coarser hydrophobic particles, while the gas flux of 0.15 cm/s delivered the best performance. Here, the cumulative recovery was 90%, and mass rejection was 60% at 0.50 mm, while the +0.090 mm recovery was 83% with a gangue rejection of 85%. The system displayed robust performance across all conditions investigated.

Dispersants under certain conditions favor the flotation of molybdenite in seawater; however, it is not clear if the entrainment of residues to the thickening stage can compromise the quality of the clarified water. In this work, the impact of small concentrations of sodium hexametaphosphate (SHMP) on the flocculation and sedimentation of synthetic tailings containing kaolinite, muscovite, and quartz in seawater is evaluated. The flocculant polymer is a high-molecular-weight polyacrylamide, and the pH is alkaline. The results are auspicious for mineral processing. On the one hand, the impact of SHMP is not entirely negative and can be lessened by limiting entrainment, which is good for copper and molybdenum ore processing. On the other hand, if the small increase in turbidity generated by the SHMP is tolerated, it is possible to expect improved settling speeds. Without SHMP, large but light agglomerates are formed. With SHMP, smaller but denser aggregates are formed, settling faster, and minute aggregates increase turbidity. The underlying mechanism derives from the competition between SHMP and polymer chains for the cations in solution; the result is a greater repulsion between the chains, which leads to greater repulsion and thus dispersion of smaller flocculant coils. The study shows that SHMP in concentrations of 1 to 3 kg/t is perfectly acceptable. The results represent an advance in the understanding of SHMP interactions with polymers and minerals in water clarification, which should be of interest to the industry whose sustainability in some regions depends on closing the water cycle.

The goal of this work was to study the bioleaching of arsenic-containing polymetallic concentrate that contained 6.2% Cu, 7.3% Zn and 1.7% As, depending on different temperatures and in the presence of CO2 and molasses in the medium, as well as the difference in the composition of microbial population formed under various conditions. A mixed population of moderately thermophilic and thermotolerant acidophilic microorganisms formed during the continuous bioleaching of copper concentrate was used as an inoculum. The experiments were carried out in a continuous mode in laboratory scale reactors, with a temperature range of 40 °C to 60 °C. To assess the effect of CO2 and molasses on metal leaching and microbial population composition, the experiments were carried out in three reactors: CO2 (~0.01 L/min) was supplied into the first reactor; 0.02% molasses were added to the pulp of the second reactor; and no additional carbon sources were supplied into the control reactor. The highest copper recovery (27%) was achieved at 50°C in the experiment with molasses, while the highest zinc recovery (82.1%) was reached at 45°C in the control experiment. Additional carbon sources affected the extraction of non-ferrous metals only at 60 °C and increased the extraction of copper and zinc by 12.6% and 24.2%, respectively. Both the temperature and carbon source used affected the microbial population composition. The main microbial genera revealed in the populations by next generation sequencing (NGS) were bacteria of the genera Sulfobacillus and Acidithiobacillus, as well as archaea of the genera Ferroplasma, Acidiplasma, and Cuniculiplasma. At low temperatures (40 and 45 °C), Acidithiobacillus, Sulfobacillus, and Ferroplasma predominated, while at temperatures 50–55 °C, the decrease in relative abundance of these genera occurred, and the predominance of Acidiplasma archaea was observed. The usage of both CO2 and molasses led to the increase in Sulfobacillus and Acidiplasma in relative abundance.

This paper reports on a study of column bioleaching of a low-grade chalcopyrite ore that is currently dump-leached under natural biological conditions without any control over microbial populations. The experimental methodology was focused on the effect of managing the bacterial populations in a raffinate solution sourced from a dump-leach operation. This study presents results from columns of two heights (0.45 and 1.0 m). We demonstrated that intermittent irrigation enhanced the chalcopyrite dissolution during column leaching, but excessively long rest periods negatively affected the chemical and bacterial activity due to the shortage of oxidizing agents and/or nutrients for microorganisms. The recovery of low-grade chalcopyrite ore was enhanced by increasing the microbial cell density. The addition of 1.5 × 108 cells/mL to the 0.45 m column and 5.0 × 107 cells/mL to the 1 m column resulted in increased extraction, with the copper dissolution increasing from 32% to 44% in the 0.45 m column and from 30% to 40% in the 1.0 m column over 70 days of leaching. Under these conditions, the pH level remained constant at ~1.8, and the redox potential was around 840 mV vs. the SHE throughout the experiment. These results provided useful insights for evaluating a sustainable controlled dump-based technology for mineral bioprocessing.

CuIn1-xZnxSe2 powders with various atomic percentages (x = 0, 0.05, 0.11, 0.16 and 0.21) were synthesized with the solvothermal method using metal chlorides and ethylendiamine as sources of precursors and a solvent, respectively. The experiment aims to investigate the effect of atomic percentages of Znx compounds on the structural and optical properties of CuIn1-xZnxSe2 in order to improve future technological applications based on this material. The powders’ chalcopyrite phases were identified by X-ray diffraction. Energy dispersive X-ray spectroscopy analysis revealed the presence of Cu, In, Zn and Se with the expected atomic ratio of Zn/(In + Zn). Scanning electron microscopy and transmission electron microscopy analysis showed that the powders have large-scale desert rose-like structures. The nanopowders’ optical study by UV-visible spectrophotometry showed that the CuIn1-xZnxSe2 energy gap values increase with the molar fraction of Znx. A change from 1.15 to 1.4 eV was observed.

This paper describes the nitric acid dissolution process of natural minerals such as tennantite, chalcopyrite and sphalerite, with the addition of Fe (III) ions and FeS2. These minerals are typical for the ores of the Ural deposits. The effect of temperature, nitric acid concentration, time, additions of Fe (III) ions and FeS2 was studied. The highest dissolution degree of sulfide minerals (more than 90%) was observed at a nitric acid concentration of 6 mol/dm3, an experiment time of 60 min, a temperature of 80 °C, a concentration of Fe (III) ions of 16.5 g/dm3, and an addition of FeS2 to the total mass minerals at 1.2:1 ratio. The most significant factors in the break-down of minerals were the nitric acid concentration, the concentration of Fe (III) ions and the amount of FeS2. Simultaneous addition of Fe (III) ions and FeS2 had the greatest effect on the leaching process. It was also established that FeS2 can be an alternative catalytic surface for copper sulfide minerals during nitric acid leaching. This helps to reduce the influence of the passivation layer of elemental sulfur due to the galvanic linkage formed between the minerals, which was confirmed by SEM-EDX.

In this work, we applied the technology of magnetic treatment to the flotation of chalcopyrite. The mechanism of the effect of pre-magnetized butyl xanthate on chalcopyrite flotation was studied using monomineral flotation tests, adsorption tests, conductivity tests, Fourier transform infrared spectroscopy (FTIR), etc. The monomineral flotation test results showed that, after the magnetization pretreatment of butyl xanthate solution, the chalcopyrite flotation recovery was increased by nearly two percentage points, and the dosage was reduced by 4–10 mg/L at the same recovery. The adsorption, FTIR, dissolved oxygen, and conductivity test results all showed that the magnetization pretreatment increased the dissolved oxygen content and promoted the oxidation of butyl xanthate to double xanthate with better selectivity to chalcopyrite. An electrochemical analysis showed that the magnetization pretreatment of butyl xanthate reduced the corrosion potential and corrosion current density of chalcopyrite surface and inhibited the self-corrosion process of chalcopyrite surface. The flotation test results of actual copper sulfide ore showed that pre-magnetized butyl xanthate could increase the copper recovery of copper concentrate by 3.06 percentage points and the sulfur recovery of sulfur concentrate by nearly 3 percentage points, and effectively reduce the mutual content of copper and sulfur concentrates.

Currently, the high demand for copper is in direct contrast with the decrease in the mineral grade and, more significantly, the concerns regarding the environmental impact that arise as a result of processing such low-grade materials. Consequently, new mineral processing concepts are needed. This work explores the chemical dissolution of chalcopyrite concentrate at ambient pressure and moderate temperatures in a deep eutectic solvent. Copper and iron are dissolved without changing their oxidation state, without solvent pH change, and stabilized as a chloride complex with no evidence of passivation. Chemical equilibria of the metallic chloride complexes limit the dissolution, and the step that is rate-controlling of the kinetics is the interdiffusion of species in the solvent. The chemical mechanism may involve initial chloride adsorption at positive sites of the solid surface, pointing out the importance of surfaces states on chalcopyrite particles. A model based on a shrinking particle coupled with pseudo-second-order increase in the liquid concentration of copper describes the dissolution kinetics and demonstrates the importance of the liquid to solid ratio. Iron and copper can be recovered separately from the solvent, which highlights that this concept is an interesting alternative to both redox-hydrometallurgy and pyrometallurgy to obtain copper by the processing of chalcopyrite concentrate.

Sodium metabisulfite (MBS) was used in this study for selective flotation of chalcopyrite and molybdenite. Microflotation tests of single and mixed minerals were performed to assess the floatability of chalcopyrite and molybdenite. The results of microflotation of single minerals showed that MBS treatment significantly depressed the floatability of chalcopyrite and slightly reduced the floatability of molybdenite. The results of microflotation of mixed minerals demonstrated that the MBS treatment could be used as a selective chalcopyrite depressant in the selective flotation of chalcopyrite and molybdenite. Furthermore, the addition of diesel oil or kerosene could significantly improve the separation efficiency of selective flotation of chalcopyrite and molybdenite using MBS treatment. A mechanism based on X-ray photoelectron spectroscopy analysis results is proposed in this study to explain the selective depressing effect of MBS on the flotation of chalcopyrite and molybdenite.

Seawater has been increasingly used as an alternative to freshwater in mineral flotation. Although previous studies suggest that Mg2+ ions in seawater have the primary negative roles in chalcopyrite flotation, insufficient work has been conducted to understand the effects of kerosene as a collector in chalcopyrite flotation. In this study, the influence of kerosene emulsion on chalcopyrite floatability in a solution containing Mg2+ was systematically investigated. The results indicated that the addition of kerosene significantly reduced the adsorption of hydrophilic Mg-precipitates onto the chalcopyrite’s surface. In addition to contact angle, zeta potential, optical microscopy, and Fourier-transform infrared spectroscopy analyses, extended Derjguin–Landau–Verwey–Overbeek (EDLVO) theory and density functional theory (DFT) calculations were conducted to understand the influencing mechanisms of kerosene on chalcopyrite flotation. The adsorption energies showed an order of kerosene and Mg(OH)2 > kerosene and chalcopyrite > chalcopyrite and Mg(OH)2, indicating kerosene was preferentially adsorbed on the Mg(OH)2 surface, forming agglomerates and therefore reducing the adsorption of Mg(OH)2 precipitates onto the chalcopyrite’s surface. In addition, hydrophobic agglomerates were also formed due to the attachment of kerosene to the chalcopyrite’s surface when additional kerosene was added, further enhancing chalcopyrite floatability.

The galvanic interaction between chalcopyrite and monoclinic pyrrhotite and its effect on flotation separation were studied using monomineral flotation tests, adsorption capacity tests, X-ray photoelectron spectroscopy (XPS) characterization, and scanning electron microscopy (SEM) test. These results showed that the interaction promoted the reduction of O2 on the cathodic chalcopyrite surface and accelerated the generation of Fe(OH)3, which was not conducive to collector adsorption; hence, the flotation recovery decreased by 10–16%. On the other hand, galvanic interaction accelerated the oxidation of S on the anodic monoclinic pyrrhotite surface to S0 or SO42− and produced a large amount of H+, thus preventing the formation of Fe(OH)3. Meanwhile, the Cu2+ eluted from chalcopyrite surface activated monoclinic pyrrhotite; hence, the flotation recovery increased by 3–10%. Galvanic interaction reduced the floatability difference between the two minerals, and the separation difficulty was significantly increased. Even with an increase in the amount of lime, the separation could not be improved.

This paper proposes selective leaching of molybdenum from Mo/Cu complex bulk concentrates in a 5 M NaCl solution using the electro-oxidation method. Here, the effects of several factors such as pH, pulp density, current density, and temperatures were investigated. A higher leaching yield of Mo increased with increasing pH from 5 to 9 and decreased with increasing pulp density from 1 to 10%. A rise in current density did not help enhance Mo, and the elevating temperature did not always result in a higher leaching yield. Application of ultrasonic led to higher leaching yield of Mo. Ninety-two percent of leaching yield was obtained upon leaching of Mo in 5 M NaCl at 25 °C, pulp density of 5%, and the current density of 0.292 A/g under ultrasonic irradiation with a power of 27 kW. The resultant residue mainly consisted of chalcopyrite.

The requirement of a ceramic product with high technical and aesthetic performance makes it necessary to select and control raw materials to avoid losses caused by low-quality products. Many defects have their origin in impurities present in clay minerals such as sulfides and sulfates. It is important to study the oxidation, decomposition, and pyrolysis reactions that affect these minerals and their dependence on environment conditions (humidity and temperature) during the extraction and beneficiation of clay minerals in an open pit mine. The development of hyphenated techniques coupling mass spectrometry with a thermal analysis instrument provides information that is not available from either technique alone, such as decomposition behavior and the determination of emissions with a lower limit of detection. The evolution of sulfur dioxide from the oxidation of different sulfides provides information on the in situ oxidation and decomposition reactions that happen when a thermal treatment is applied. The results obtained show important differences in the reactions that take place in sulfides when they are stored under different environmental conditions. Specifically, the general tendency is that the sulfurs stored under high relative humidity show a decrease in the intensity of the emission as well as changes in the onset of the peak emission.