The kinetics of high-temperature oxidation of copper of various chemical compositions by gaseous oxygen follows a parabolic rate law within the temperature range of 350–1050 °C. For specialists engaged in the theory and practice of fire refining, of particular interest is the kinetics of copper oxidation over a broader temperature interval of 350–1160 °C. Within this range, the processes include oxidation of solid copper and its melting, oxidation of liquid copper by oxygen introduced into the melt, oxygen solubility in copper, and slag formation. The duration of high-temperature interaction between copper and oxygen has a considerable effect on both the technical and economic indicators of anode smelting and the electrical properties of copper. Therefore, investigating the kinetics of high-temperature oxidation of copper and its influence on the electrical properties of the metal is essential for the optimal organization of fire refining. In the temperature range of 1100–1200 °C, copper oxidation occurs predominantly due to oxygen introduced into the melt with air. The copper(I) oxide formed migrates from the zone of direct contact with gaseous oxygen into the depth of the melt with lower oxygen concentration, where it dissociates into copper and oxygen, thus increasing oxygen concentration in the melt. Overoxidation of copper and excessive saturation of the anode metal with gases lead to its transfer into slag in the form of oxides, to excessive consumption of resources and refractory materials, and to a deterioration of the electrical properties of the metal. To identify optimal oxidation modes and to assess the influence of copper oxidation kinetics on the electrical properties of the melt, a comparative analysis was conducted of the kinetic patterns of oxidation of solid and liquid copper of various chemical compositions under identical experimental conditions, using differential thermogravimetric analysis (DTA) and by surface blowing of the copper melt with an air mixture. The results show that copper samples oxidize at nearly the same rate and that the presence of impurities does not affect the process. All oxygen intended for copper (I) oxide formation dissolves in copper up to the thermodynamic limit (up to 12 % Cu2O). It was established that oxygen concentrations in the melt above 0.06% adversely affect the electrical properties of copper. 

This study presents the results of oxidative leaching of chalcopyrite (CuFeS₂) and pyrite (FeS₂) in a sulfuric acid medium at low temperature in the presence of copper (Cu²⁺) and iron (Fe³⁺) ions. Using orthogonal experimental design, the optimal conditions were identified to maximize sulfide matrix decomposition and valuable metal recovery. Experiments were conducted at a constant temperature of 100 °C. The parameters investigated included partial oxygen pressure (0.2–0.75 MPa), concentrations of sulfuric acid (10–50 g/dm³), Fe³⁺ ions (2–10 g/dm³), Cu²⁺ ions (1–3 g/dm³), and leaching time (60–240 min). The composition of the feed minerals and leach products was analyzed by X-ray fluorescence (XRF) analysis, X-ray diffraction (XRD) analysis, and atomic absorption spectrometry (AAS). Maximum copper recovery from chalcopyrite (55 %) was achieved under the following conditions: O₂ partial pressure of 0.25 MPa, initial concentrations of H₂SO₄ – 50 g/dm³, Cu²⁺ – 1 g/dm³, Fe³⁺ – 2.5 g/dm³, and leaching time – 240 min. The maximum degree of pyrite oxidation (56 %) was obtained at an O₂ partial pressure of 0.75 MPa, initial concentrations of H₂SO₄ – 50 g/dm³, Cu²⁺ – 2 g/dm³, and Fe³⁺ – 10 g/dm³. The results showed that leaching time and oxygen pressure have the greatest effect on chalcopyrite and pyrite decomposition (p < 0.05). The interaction between Fe³⁺ and Cu²⁺ ions was also established: excess Fe³⁺ (>10 g/dm³) leads to hydrolysis and decreases chalcopyrite leaching efficiency, whereas Cu²⁺ promotes partial formation of secondary copper sulfides. Regression equations (R² = 0.98 for chalcopyrite and R² = 0.96 for pyrite) were derived, providing an adequate description of the process.

This study investigates the effect of preliminary autoclave oxidation with molecular oxygen (Т = 423 K, Р_О₂ = 0.6 MPa, τ = 2 h) on lignosulfonates differing in chemical composition and molecular weight distribution. Oxidation resulted in a reduction of hydroxyl groups and an increase in carbonyl groups, along with marked changes in solution properties such as redox potential, pH, specific conductivity, and surface tension at the liquid–gas interface. The functional activity of the initial and oxidized lignosulfonates was compared in terms of their ability to remove elemental sulfur films from the sphalerite surface under high-temperature oxidative pressure leaching conditions. The findings show that oxidative treatment decreases the effectiveness of lignosulfonates by diminishing their surface activity.

 The Ural region holds an estimated 1.5 million tons of nickel reserves, located in the industrially developed Chelyabinsk, Sverdlovsk, and Orenburg regions. At present, however, these deposits are not being exploited, and metallic nickel is not produced in the Urals, as metallurgical facilities have been completely shut down. The reserves are represented by oxidized nickel ores (ONO) – a complex raw material with low nickel and cobalt contents, whose processing by existing technologies is economically unfeasible. The challenge is compounded by the absence of a beneficiation method for ONO that yields a concentrate; therefore, all current technologies require processing the entire ore mass, which results in high reagent consumption and substantial energy costs. Research is ongoing to develop new technological approaches, including alternative methods for extracting nickel and cobalt from ONO in the Ural deposits. One promising option is the use of microwave (MW) energy to unlock nickel-bearing minerals and accelerate the dissolution of nickel and cobalt. This study evaluates the effect of microwave energy on nickel recovery from oxidized nickel ores of the Ural region. Comparative data are presented for conventional sulfuric acid leaching and for the process with microwave energy applied. A series of test studies was carried out to assess the feasibility of using microwave energy for ONO processing. The comparison of technological parameters demonstrated the advantage of atmospheric sulfuric acid leaching with microwave energy, which achieved nickel recovery of up to 95 % in a relatively short time. These results identify this approach as the most promising for practical implementation.

This study explores the nitric acid leaching of stibnite in the presence of tartaric acid, which acts as a complexing agent. The proposed approach is of considerable interest, as antimony is widely used across industries, from electronics to alloying applications. Thermodynamic analysis showed that nitric acid dissolution of stibnite inevitably leads to the formation of antimony oxides, which markedly reduces the extraction of the target metal into solution. To counteract these losses and enhance process efficiency, tartaric acid was introduced as an additive. The results demonstrated that tartaric acid promotes the formation of stable complexes with antimony ions, thereby retaining the metal in solution and minimizing the risk of oxide precipitation. Experimental design analysis revealed that the mass ratio of tartaric acid to antimony and the nitric acid concentration exert a stronger influence on leaching efficiency than temperature and leaching duration. Optimal conditions were established, achieving an antimony extraction of 87 %: temperature 35 °C, nitric acid concentration 5 mol/dm3 , leaching time 45 min, and a tartaric acid-to-antimony mass ratio of 4.5 : 1.0.

With the declining quality of feedstock in the copper industry, maintaining metal recovery rates and controlling production costs for non-ferrous and precious metals has become increasingly critical. A key research priority is therefore the development of processing strategies that not only concentrate target metals into flotation products but also recover valuable minor elements previously lost with slags and flue dust. One approach involves designing process flowsheets that integrate hydrometallurgical and beneficiation operations. Previous studies have demonstrated the effectiveness of combining autoclave leaching and flotation for decopperization of copper anode slimes and their concentration in gold, silver, and selenium. However, autoclave leaching requires significant capital and operating expenditures. For this reason, a series of tests was carried out on aeration leaching (decopperization) of copper anode slimes followed by flotation, yielding promising results. This study examined the influence of aeration leaching conditions (temperature, agitation, and specific oxidant consumption—air and oxygen), disintegration of the leached product, and flotation parameters on the selective separation of oxide and chalcogenide phases and the quality of the resulting concentrates. Based on the experimental results, process operations were developed that make it possible to concentrate precious metals in copper anode slimes two- to threefold without the use of autoclave leaching. Optimal conditions and equipment configurations were determined for deep decopperization of slimes (to less than 0.5–0.8 % residual copper). An acceptable degree of separation of precious-metal chalcogenides from lead and antimony oxides was achieved, enabling downstream recovery of marketable products from the respective concentrates. Analytical characterization of the products was performed using scanning electron microscopy (SEM) and energy-dispersive X-ray spectroscopy (EDS). The findings contribute to the development of an integrated hydrometallurgical technology for processing copper anode slimes.

Due to the specific features of hydrometallurgical processing, where cyanide solutions are used, the composition of the leaching solution undergoes periodic changes referred to in the literature as “fatigue.” This adversely affects the rate of gold recovery and cementation, and therefore the overall efficiency of cyanide leaching technology. One of the most important markers determining the “fatigue” of the solution is copper. This study examined the possibility of applying a combined scheme for the purification of circulating cyanide solutions with high concentrations of copper (1196 mg/dm³), iron (111 mg/dm³), arsenic (19 mg/dm³), and sodium cyanide (0.94 g/dm³). A two-stage technology (reverse osmosis + chemical precipitation) was developed for the reduction of treated solution volumes and the removal of impurities. At the first stage, the solution was separated in a reverse osmosis unit equipped with LP22-8040 membranes, producing permeate and concentrate in a 1 : 1 ratio. The permeate (12 mg/dm³ Cu and 0.01 mg/dm³ Fe, pH = 11.13) was returned to the process cycle. At the second stage, the concentrate, which contained 99 % of the initial impurities, was further purified by the stepwise addition of a CuSO₄ solution (70 g/dm³ Cu) in 1–11 cm³ doses under stirring (500 rpm, 10 min). The results showed that the optimal CuSO₄ dose (11 cm³) provided removal of more than 86 % of Cu from the initial solution, as well as 100 % of Fe and more than 96 % of As. The precipitate obtained in the process consisted of 68.3 % copper, with CuCN and Cu(OH)₂ as the main components.

Global gold consumption has steadily increased in recent decades, driven by expanding industrial applications and reserve accumulation by many countries. In parallel, depletion of high-grade deposits has shifted processing toward low-grade and refractory ores. These trends—together with tighter environmental regulations—highlight the need for alternative lixiviants for gold extraction. Although cyanide remains the industry standard, it is highly toxic and often ineffective for refractory sulfide ores. Other systems—thiosulfate (including ammoniacal thiosulfate), thiourea, and bromide/iodide lixiviants—are used far less frequently due to significant disadvantages. Among acidic chloride lixiviants, sodium dichloroisocyanurate (NaDCC) was investigated as a promising candidate. Use of NaDCC requires strongly acidic solutions (pH < 1.0) and an excess of Cl^–, i.e., conditions consistent with the stability domain of the Au(III) chloride complex (AuCl₄^–). Using the rotating-disk technique, we examined the effects of temperature, disk rotation rate, and HCl concentration on the specific dissolution rate of the reagent (NaDCC), as well as on solution pH and redox potential (Eh). NaDCC hydrolyzes in water to form hypochlorous acid (HClO)—the primary source of active chlorine—while the concurrent pH decrease arises from formation of weak acids (hypochlorous and cyanuric). Adding HCl to NaDCC solutions generates molecular chlorine (Cl₂), which evolves once its solubility limit is exceeded. Gold-dissolution tests across NaDCC and HCl concentrations identified an optimum at [HCI] = 14.4 g/dm³ and [NaDCC] = 3.0 g/dm³, yielding a maximum gold dissolution rate of υ_Au = 0.118 mg/(cm²·min).