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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.

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.

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.

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.

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.

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.

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.

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.

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:

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.

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.

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.

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.

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.

An essential operation in the mineral processing of copper ores into concentrates is blending, as it guarantees a constant feed for the flotation cells, increases metal recovery rate and reduces tailings. In this study, copper ores from Huelva province (Spain) were investigated by quantitative XRD (X-ray diffraction) methods to optimize blending and detect penalty minerals, which can affect flotation and concentrate quality. The Rietveld method in combination with cluster analysis, PLSR and more traditional chemical analysis provide a more complete and in-depth characterization of the ore and the whole process. The mineralogical monitoring can be fully automated to enable real-time decision making.

The mechanisms involved in the dissolution of chalcopyrite from a carbonatite concentrate in a ferric sulphate solution at pH 1.0, 1.5 and 1.8, and temperatures 25 °C and 50 °C were investigated. Contrary to expectations and thermodynamic predictions according to which low pH would favour high Cu dissolution, the opposite was observed. The dissolution was also highly correlated to the temperature. CuFeS2 phase dissolution produced intermediate Cu rich phases: CuS, Cu2S and Cu5FeS4, which appeared to envelop CuFeS2. Thermodynamic prediction revealed CuS to be refractory and could hinder dissolution. CuFeS2 phase solid-state dissolution process was further discussed. Free Fe3+ and its complexes (Fe(HSO4)2+, Fe(SO4)2– and FeSO4+ were responsible for Cu dissolution, which increased with increasing pH and temperature. The dissolution improved at pH 1.8 rather than 1.0 due to the increase of (Fe(HSO4)2+, Fe(SO4)2– and FeSO4+, which were also the predominating species at a higher temperature. The fast and linear first dissolution stage was attributed to the combined effect of Fe3+ and its complex (Fe(HSO4)2+, while Fe(SO4)2– was the main species for the second Cu dissolution stage characterised by a slow rate.

With the decrease in high-grade chalcopyrite resources, the copper extraction from low-grade chalcopyrite has attracted more and more attention. However, the kinetic rates of chalcopyrite leaching with traditional oxidants are usually very slow due to the formation of the passivation layer. In this study, a novel reagent of chlorinated oxidant, trichloroisocyanuric acid (TCCA), was used to leach chalcopyrite for the first time. The experimental results showed that when the initial oxidant concentration for TCCA was 0.054 mol·L−1, the leaching temperature was kept at 55 °C, and the pH of the pulp was controlled at 1, the oxidation efficiency of Cu can reach above 90% in less than 30 min. Various analyses of chalcopyrite mineral ore and its oxidized residues, such as chemical composition analysis, X-ray diffraction analysis, scanning electron microscopy analysis and X-ray photoelectron spectroscopy, were conducted, respectively. No obvious passivation layer was found on the chalcopyrite surface, though the sulfur product can also be generated during the leaching. Reaction kinetic analysis results showed that the different influence of surface reaction and diffusion process on the dissolution of chalcopyrite is little due to the fast leaching speed. After calculation, the activation energy of the whole leaching reaction is 9.06 kJ·mol−1, much lower than that in other reports. The mechanism was also proposed that TCCA was hydrolyzed in the solution to form hypochlorous acid, which is the strong oxidant, and cyanuric acid, which prevents the formation of a passivation layer. The processing in this study is expected to be applied as a novel method for atmospheric leaching of chalcopyrite.

The study of the depression effect of non-toxic depressants on the flotation separation of chalcopyrite from galena is of great importance for both industrial applications and theoretical research. The mixed depressant (DFinal) of four common inhibitors—sodium carboxymethyl cellulose, sodium silicate, sodium sulfite, and zinc sulfate—exhibited high selectivity during the separation of chalcopyrite from galena. Flotation tests on an industrial copper–lead bulk concentrate showed that using this depressant mixture can achieve highly efficient separation of chalcopyrite from galena at the natural pH of the pulp. Copper and lead concentrates were produced at grades of 21.88% (Cu) and 75.53% (Pb), with recoveries of 89.07% (Cu) and 98.26% (Pb). This showed a similar performance of DFinal with dichromate, which is a depressant that is widely used in industry, but without the environmental risks or the need for pH control. Zeta potential and Fourier transform infrared (FT-IR) results showed that interaction between the surface of the chalcopyrite and the mixed depressant was prevented by pre-treatment with a composite thiophosphate collector (CSU11), while the mixed depressant could expel/replace the composite thiophosphate on the surface of galena by chemical adsorption, depressing its flotation. This is the reason why this non-toxic depressant achieved the selective depression of galena from chalcopyrite, leading to efficient flotation separation.

The objective of this work was to study the effect of muscovite and biotite on the flotation of chalcopyrite and molybdenite in seawater, and the use of sodium hexametaphosphate and sodium silicate to improve copper and molybdenum recoveries. The impact of the inorganic dispersants on the settling properties of the resulting flotation tailings was also studied. It was found that muscovite and biotite depress the flotation of chalcopyrite and molybdenite in seawater, with this depressing effect being stronger at pH 11 than at pH 9. Sodium hexametaphosphate and sodium silicate increased the recoveries of copper and molybdenum in seawater. These dispersants render the mineral particles more negatively charged and remove the hydroxy-complexes of magnesium and calcium from the mineral particles causing dispersion of the slimes. The settling rates of the flotation tailings slightly decrease when sodium hexametaphosphate and sodium silicate were added in the flotation stage. The presence of ultrafine particles dispersed by the action of the inorganic dispersants negatively impacted the flocculation and sedimentation processes leading to high flocculant consumption, low settling rates and high turbidity of the supernatant.

The leaching of chalcopyrite under different redox potentials was studied with a redox potential controlling device based on microcontroller technology. The leaching test was carried out in a 500 mL flask with 4 g pure chalcopyrite and 200 mL sulfuric acid. The additional total Fe concentration was 0.1 mol/kg, and the initial pH was 0.7. Chalcopyrite leaching tests with initial redox potential of 400 mV and 600 mV (vs. Ag/AgCl, the same below) and controlled redox potential of 350 mV, 400 mV, 450 mV, 500 mV, and 600 mV were carried out at a temperature of 60 °C and a stirring speed of 300 rpm. The results showed that the dissolution of chalcopyrite could not be hindered until the granular intermediates (S0 and Sn2−) accumulated to a certain amount on the chalcopyrite surface. The main passivating species in chalcopyrite dissolution may be iron-deficient polysulfides. At 400 mV, the chalcopyrite passivation by polysulfides was overcome by the galvanic effect between pyrite and chalcopyrite, and the chalcopyrite was rapidly oxidized with Cu leaching rate reaching up to 32% in 7 h, while it was only 15% in the other experimental groups. However, at 450–500 mV, pyrite was oxidized and the galvanic effect between chalcopyrite and pyrite was destroyed, leading to a significant decrease in Cu leaching rate. When the redox potential was 600 mV, the dissolution of metal-deficient polysulfides was promoted; thus, the passivation was weakened.

Overcoming the slow-leaching kinetics of refractory primary copper sulfides is crucial to secure future copper sources. Here, the effect of carbon was investigated as a catalyst for a bioleaching reaction. First, the mechanism of carbon-assisted bioleaching was elucidated using the model chalcopyrite mineral, under specified low-redox potentials, by considering the concept of Enormal. The carbon catalyst effectively controlled the Eh level in bioleaching liquors, which would otherwise exceed its optimal range (0 ≤ Enormal ≤ 1) due to active regeneration of Fe3+ by microbes. Additionally, Enormal of ~0.3 was shown to maximize the carbon-assisted bioleaching of the model chalcopyrite mineral. Secondly, carbon-assisted bioleaching was tested for three types of chalcopyrite/enargite-bearing complex concentrates. A trend was found that the optimal Eh level for a maximum Cu solubilization increases in response to the decreasing chalcopyrite/enargite ratio in the concentrate: When chalcopyrite dominates over enargite, the optimal Eh was found to satisfy 0 ≤ Enormal ≤ 1. As enargite becomes more abundant than chalcopyrite, the optimal Eh for the greatest Cu dissolution was shifted to higher values. Overall, modifying the Eh level by adjusting AC doses to maximize Cu solubilization from the concentrate of complex mineralogy was shown to be useful.

Many studies have shown that the average iron (Fe) isotope compositions of mantle-derived rocks, mantle peridotite and model mantle are close to those of chondrites. Therefore, it is considered that chondrite values represent the bulk Earth Fe isotope composition. However, this is a brave assumption because nearly 90% of Fe of the Earth is in the core, where its Fe isotope composition is unknown, but it is required to construct bulk Earth Fe isotope composition. We approach the problem by assuming that the Earth’s core separation can be approximated in terms of the Sudbury-type Ni-Cu sulfide mineralization, where sulfide-saturated mafic magmas segregate into immiscible sulfide liquid and silicate liquid. Their density/buoyancy controlled stratification and solidification produced net-textured ores above massive ores and below disseminated ores. The coexisting sulfide minerals (pyrrhotite (Po) > pentlandite (Pn) > chalcopyrite (Cp)) and silicate minerals (olivine (Ol) > orthopyroxene (Opx) > clinopyroxene (Cpx)) are expected to hold messages on Fe isotope fractionation between the two liquids before their solidification. We studied the net-textured ores of the Sudbury-type Jinchuan Ni-Cu sulfide deposit. The sulfide minerals show varying δ56Fe values (−1.37–−0.74‰ (Po) < 0.09–0.56‰ (Cp) < 0.53–1.05‰ (Pn)), but silicate minerals (Ol, Opx, and Cpx) have δ56Fe values close to chondrites (δ56Fe = −0.01 ± 0.01‰). The heavy δ56Fe value (0.52–0.60‰) of serpentines may reflect Fe isotopes exchange with the coexisting pyrrhotite with light δ56Fe. We obtained an equilibrium fractionation factor of Δ56Fesilicate-sulfide ≈ 0.51‰ between reconstructed silicate liquid (δ56Fe ≈ 0.21‰) and sulfide liquid (δ56Fe ≈ −0.30‰), or Δ56Fesilicate-sulfide ≈ 0.36‰ between the weighted mean bulk-silicate minerals (δ56Fe[0.70ol,0.25opx,0.05cpx] = 0.06‰) with weighted mean bulk-sulfide minerals (δ56Fe ≈ −0.30‰). Our study indicates that significant Fe isotope fractionation does take place between silicate and sulfide liquids during the Sudbury-type sulfide mineralization. We hypothesize that significant iron isotope fractionation must have taken place during core–mantle segregation, and the bulk Earth may have lighter Fe isotope composition than chondrites although Fe isotope analysis on experimental sulfide-silicate liquids produced under the varying mantle depth conditions is needed to test our results. We advocate the importance of further research on the subject. Given the close Fe-Ni association in the magmatic mineralization and the majority of the Earth’s Ni is also in the core, we infer that Ni isotope fractionation must also have taken place during the core separation that needs attention.

Bacteria–mineral contact and noncontact leaching models coexist in the bioleaching process. In the present paper, dialysis bags were used to study the bioleaching process by separating the bacteria from the mineral, and the reasons for chalcopyrite surface passivation were discussed. The results show that the copper leaching efficiency of the bacteria–mineral contact model was higher than that of the bacteria–mineral noncontact model. Scanning electron microscopy (SEM), X-ray diffraction (XRD), and Fourier-transform infrared (FTIR) were used to discover that the leaching process led to the formation of a sulfur film to inhibit the diffusion of reactive ions. In addition, the deposited jarosite on chalcopyrite surface was crystallized by the hydrolysis of the excess Fe3+ ions. The depositions passivated the chalcopyrite leaching process. The crystallized jarosite in the bacteria EPS layer belonged to bacteria–mineral contact leaching system, while that in the sulfur films belonged to the bacteria–mineral noncontact system.

Bioleaching models to examine copper extraction from low-grade chalcopyrite ores were set up to identify the influence of pyrite on leaching efficacy. A combination of scanning electron microscopy and geochemical analysis showed that extraction was marginally enhanced by the addition of pyrite when using a combination of Leptospirillum ferrooxidans, an iron oxidiser, Acidithiobacillus thiooxidans, a sulphur oxidising species and Acidithiobacillus ferrooxidans, an iron and sulphur oxidiser. Extensive biofilms formed on the pyrite surfaces (>106 cells/mm2) but were severely limited on chalcopyrite, possessing approximately the same number of cells as quartz grains, an internal non-nutrient control “substrate” (with ca. 2 × 103 cells/mm2). The presence of dissolved copper did not inhibit the growth of this consortium. Indirect “bioleaching” of chalcopyrite appears to be limited by proton activity at the chalcopyrite surface.

Understanding the oxidative mechanisms of chalcopyrite in various hydrothermal fluids is of great significance to improve copper extraction and to model the geochemical cycling of copper, iron, and sulfur. This paper investigated the effect of NaCl on the electrochemical oxidation of chalcopyrite at 340 °C and 21 MPa using polarization curves, electrochemical impedance spectroscopy (EIS), scanning electron microscopy (SEM), energy dispersive X-ray spectroscopy (EDS), and Raman spectroscopy. The experimental results showed that NaCl can promote chalcopyrite leaching. As NaCl concentration increases, the protective property of the oxidation layer degraded. In the absence of NaCl, the oxidation layer that consisted of CuSn, (n ≥ 1), probably with some Fe2O3 and Fe(OH)3 and also in the presence of NaCl, Fe2O3, is the principal oxidation product. More rapid ionic diffusion and further chemical reaction contributed to the improvement of chalcopyrite dissolution with increasing NaCl concentration. A dissolution mechanism is proposed in this paper to explain the chalcopyrite leaching processes which is dependent on NaCl concentration.

Mechanical activation as a means of accelerating the mineral dissolution may play an important role in chalcopyrite bioleaching. In the present work, the mechanical activation by ball-milling with 10 min, 30 min, 60 min, 90 min, 120 min and 180 min time periods of bioleaching of chalcopyrite was studied, and then evaluated by a Density Functional Theory (DFT) calculation. The results showed that the specific surface area increased sharply in the very beginning of mechanical activation and then increased slowly until the agglomeration of the particles occurred, while the chalcopyrite lattices increased with the mechanical activation. The reaction activity analyzed by cyclic voltammetry (CV) increased slowly in 30 min, increased quickly in the following 90 min, and then decreased, while the hydrophobicity analyzed by contact angles of the chalcopyrite after activation showed less of a change. The results showed that after 15 days of bioleaching, the Cu leaching by Sulfobacillus thermosulfidooxidans (S. thermosulfidooxidans) increased from 9.39% in the 0 min of mechanical activation to 87.41% in the 120 min of mechanical activation, and the copper leaching rate increased by about 78%. The DFT results provide solid proof that the activated chalcopyrite can be adsorbed more easily by cells with higher adsorption energies and stronger bonds.

The combined effect of dissolved ions and water temperature on the adsorption of a xanthate collector on chalcopyrite and pentlandite was investigated using multiple linear regression. Cationic species improved the adsorption of the collector on sulphide minerals through xanthate adsorption activation. Thiosulphate ions generally had a negative effect on collector adsorption, and the interaction of thiosulphate ions and cations effectively reduced collector adsorption on the sulphide minerals. With regards to temperature variation caused by seasonal variation, this study suggests that temperature can influence the adsorption of collectors in the flotation process and this should be approached on a case by case basis as it seems to differ with the type of mineral under investigation. These fundamental results prompt a discussion on how complex water matrices can affect interactions of reagents and sulphide minerals at the solid–liquid interface and the possible effect on flotation performance.

The Oktyabrsk PGE-Cu-Ni deposit is one of the largest resources in the Norilsk–Talnakh ore district, Russia, and it is viewed as an ore giant on a global scale. It contains three types of ores: massive, disseminated and veinlet-disseminated. The two former ore types were formed by a liquation process, whereas the latter was associated with fluid-induced selective metasomatic replacement of metamorphosed wall rocks. One of the major ore minerals in all ore types is chalcopyrite. In this study, we determined concentrations of trace elements in this mineral using laser ablation inductively coupled plasma mass spectrometry. It appeared that standard geochemical tools, such as plotting the data in the form of diagrams of normalized concentrations, binary and ternary plots, do not allow one to distinguish chalcopyrite from visually and genetically different ore types. In contrast, more advanced statistical methods such as cluster analysis show different groupings of elements for each ore type. Based on the element clustering, a classification tree was suggested, which allowed for the differentiation of massive, disseminated and veinlet-disseminated ore types of the Oktyabrsk deposit by Se, Te, Cd and Pb concentrations in chalcopyrite with a success rate of 86%. The general feature is that chalcopyrite of veinlet-disseminated ore is poorer in these elements compared to chalcopyrite of the two other ore types. Chalcopyrite of massive ore is poorer in Se and Te when compared to chalcopyrite of disseminated ore.

The objective of this study is to compare the reaction kinetics of copper leaching from chalcopyrite in acidic ferric sulfate media with (UAL) and without (non-UAL) ultrasound assistance. Four leaching parameters were evaluated and optimized. The parameter with the strongest effect was temperature, followed by ultrasonic power, the solid-to-liquid ratio (S/L), and acid concentration. Copper recovery showed an increase with rising temperatures in both systems. Ultrasonic power had a positive effect on copper leaching, but no significant difference was found among various power amplitudes. Copper extraction increased with decreasing S/L. At 0.1% S/L, the UAL leaching rate was double the non-UAL leaching rate. In both systems, acid concentration had little effect on copper extraction. Under optimized conditions, 20% amplitude power, 1% S/L, 0.5 M acid, and 80 °C leaching temperature, copper extraction was 50.4% and 57.5% in the non-UAL and UAL treatments, respectively. Ultrasonic waves enhanced the leaching rate, shortened the reaction time, and reduced acid consumption. Analysis of the rate-controlling step using a shrinking core model showed that leaching occurs after diffusion through the product layer but also chemical controlled in both non-UAL and UAL systems. The leaching mechanism was confirmed by characterizing the chalcopyrite and leached residue with X-ray diffraction and scanning electron microscopy/energy dispersive X-ray spectroscopy.

A multi-scale investigation was conducted to study the surface properties of xanthate-absorbed chalcopyrite at elevated temperature to understand the temperature effect on the xanthate’s performance during chalcopyrite flotation. Firstly, a macro-scale study was initiated to investigate the temperature effect on the hydrophobicity of mineral surface by means of contact angle measurement, Hallimond tube microflotation and lab flotation tests; secondly, a micro–scale study was conducted to clarify the temperature effect on the adsorption of chemicals on mineral surface employing an atomic force microscope (AFM) and Fourier transform infrared spectroscopy-attenuated total reflectance (FTIR-ATR). In the experiments, pure chalcopyrite samples were used for contact angle measurement, Hallimond tube microflotation, AFM and FTIR; and copper ore samples (1.51% Cu, 5.88% Fe 0.029% Mo, 5.23% S in weight percentage) were used for lab flotation tests. FTIR spectra and AFM images showed that, when potassium amyl xanthate (PAX) was used as the collector in this study, oily dixanthogen was the main hydrophobic species on the chalcopyrite surface. The morphological change of dixanthogen patches at elevated temperatures has a more significant impact than changes in the amount of adsorption species. Increasing temperature within a certain range is beneficial for the collector’s performance by increasing flotation recovery.

In flotation, the size of mineral particles is one of the most important parameters: when the size becomes fine, collision efficiency of the particles and air bubbles becomes low, causing low flotation recovery. To improve the collision efficiency and flotation kinetics, agglomeration using the emulsified oil of finely ground chalcopyrite (D50 = 3.5 μm) was carried out before flotation. In this study, the effects of agitation strength during agglomeration, kerosene dosage and potassium amyl xanthate (KAX) dosage on the flotation were investigated. Agglomeration using emulsified oil improved Cu recovery because the median diameter of agglomerate increased. With increasing agitation strength, KAX and kerosene dosages, Cu recovery was further increased. Agglomeration-flotation of a mixture containing chalcopyrite and quartz with 1:1 ratio (w/w, weight by weight) showed that Si recovery in froth was low and did not change with varying conditions (agitation strength, KAX and kerosene dosages); however, Cu recovery was significantly improved with increasing agitation strength, KAX and kerosene dosages, and thus the separation efficiency was improved.

The selective separation of chalcopyrite from galena (lead sulfide) through the flotation method is still a challenging task in the field of mineral engineering. Mercaptoacetic acid, though a common depressant of many gangue (commercially worthless) minerals, has shown to be less selective in the flotation separation of chalcopyrite and galena. This resent study therefore systematically investigated the selectivity of different mercapto acids (especially three types: 3-mercaptopropionic acid, 3-mercaptoisobutyric acid and mercaptoacetic acid) on the separation of chalcopyrite and galena by making the use of flotation experiments and first principle calculations. The calculation results demonstrated that the sulfhydryl and carboxyl groups existing on the molecular structure of three mercapto acids are the reactive and chelating centers to metal ions on sulfide mineral surfaces. Mercapto acids have higher binding energies to Cu2+ by 300−400 kJ/mol compared to Pb2+, indicating a higher affinity towards chalcopyrite. The order of reactivity and chelating ability noted was as follows: 3-mercaptopropionic acid > 3-mercaptoisobutyric acid > mercaptoacetic acid. Flotation results further showed that the selectivity of 3-mercaptopropionic acid or 3-mercaptoisobutyric acid was better than mercaptoacetic acid. The good agreement between the first principle calculations and the flotation results demonstrated that the former reagent could be served as a most selective depressant in the improved flotation separation of chalcopyrite and galena.