The heterogeneity and local distribution of elements are established as a result of the study of nonferrous metals distribution and the composition of domestic wastes processing concentrate containing 20–40% copper, 40–50% brass, 20–28% zinc and up to 1% aluminium. Metals are mainly concentrated in granules of three types: zinc-based, copper-based and copper–zinc alloy, i.e., brass. The phase composition of these granule types and their structure are determined. Zinc granules are covered with a refractory oxide shell. A distillation processing method for such raw materials based on full state diagrams, including the phase transition of melting pairs of double systems of copper and tin with zinc and lead, is offered. The possibility of a rather complete zinc distillation (over 90%) with the accumulation of other metals in the copper-based alloy, containing more than 96 wt. % of the basic element, has been demonstrated by electric crucible melting in a highly reducing atmosphere. Copper-based alloys, after adjustment for the content of some metals (tin), can be realized as pressure-treated tin bronzes. Copper with the content of impurities corresponding to the standard for blister copper can be electrolytically processed (after dissolution in sulphuric acid) in copper production. Equipment for the implementation of the electro-thermal processing process that provides metal recycling of the metal concentrate is proposed.

This paper presents the results of a study of phase transformations occurring in copper-telluride by-products during its processing of oxidation-distillate roasting at low pressure. The results show that copper telluride is oxidized through intermediate compounds to the most stable tellurate (Cu₃TeO₆) at low temperatures. The increase in the roasting temperature above 900 °C and the presence of an oxidizer favor the copper orthotellurate decomposition. Thus, the tellurium extraction rate is 90–93% at a temperature of 1000 °C, the oxidant flow rate is 2.2 × 10⁻² m³/m²·s, and the roasting time is 60–90 min. One of the decomposition products is copper oxide alloy, which is the basis of the residue. The second product is tellurium in oxide form, which evaporates and then condenses in the cold zone of the condenser in crystalline form. The main constituent phase of the condensate is tellurium oxide (TeO₂), which can be further processed during one operation to elemental chalcogen by thermal reduction or electrolytic method.

The practicability of a pyrometallurgical scheme for raw material processing is established as a result of the analysis of methods intended to dearsenate and process gold-arsenic concentrates as well as equipment for the process execution. The conceptual design of vacuum equipment without forced movement of the dispersed material in the sublimator and of the reaction zone materials is proposed. In-process tests for the sublimation of arsenic sulfides from gravity and flotation concentrates received from the Bakyrchik deposit were executed at the pilot facility. As a result, it was found that more than 97–99% of arsenic passes into the gas phase and condenses in a sulfide form suitable for compaction by smelting. More than 99.5% of precious metals are concentrated in the sublimation residue. As a result of smelting residue from the sublimation of arsenic sulfides in a cyclone furnace, together with copper concentrates to copper matte, the gold recovery was 93.7–93.9% of the total amount loaded. Silver was 65.7–68% concentrated in copper matte, with a considerable amount in the dust. If the cyclone smelting dust is involved, the recovery rate of gold and silver can be increased to 97–99% and 94–95%, respectively. As a result of crucible smelting, the degree of recovery of gold in matte was 95.4%, with its content in slag being 3.6 g/t. The received matte according to the proposed scheme can be directed to the conversion process by obtaining blister copper, which is subjected to electrolytic refining with the recovery of gold from slimes.

This study investigates the use of preliminary thermochemical activation in a NaHCO₃ solution under elevated pressure and temperature to modify the chemically stable and hard-to-process phase composition of various mineral raw materials and improve the recovery of valuable components. The method was tested on various types of mineral raw materials, including slag from the reductive smelting of red mud from alumina production prior to acid leaching, ash before chemical beneficiation, gibbsite–kaolinite bauxite prior to gravity separation, and nephelines, for which the sintering process was replaced with chemical beneficiation. The slag from the reductive smelting of red mud was also tested before acid leaching. The activation of slag enhanced tricalcium silicate formation lead to leaching recoveries of ~96% for rare earth elements, ~92% for TiO₂, ~98% for CaO and Al₂O₃, and 50% for Fe₂O₃, compared to much lower values without activation. With ash, activation eliminated the sillimanite and hedenbergite phases, increased mullite and free silica, and formed calcite, resulting in a 15–20% higher silica recovery. With gibbsite–kaolinite bauxite, activation altered kaolinite, siderite, quartz, and hematite contents; eliminated calcium silicate; and improved the silicon modulus of the sand fraction by 35.9% during gravity beneficiation. For nepheline ore, activation promoted the formation of albite and hydrosodalite, eliminated corundum and andradite, and increased silica recovery from 33.58% to 59.6%. These results demonstrate that thermochemical activation effectively transforms mineral structures and significantly improves the efficiency of subsequent beneficiation processes.

The suitability of microsilica as a raw material for the production of ceramic mullite–silicate proppants was assessed. The chemical and mineralogical compositions of the initial materials were studied. The mineral composition of bauxite is mainly represented by gibbsite Al(OH)₃ and, to a lesser extent, kaolinite Al₄[Si₄O₁₀](OH)₈, with impurities of hematite and quartz. It is established that, in order to obtain mullite–silicate proppants, compositions containing 10–20% microsilica are optimal. The sintering of these compositions occurs at 1350–1380 °C. A lightweight ceramic proppant was obtained with a bulk density of 1.21–1.41 g/cm³, breaking ratio at 51.7 MPa of 19.1–20.3%, and sphericity and roundness of 0.7–0.9, and the optimal roasting temperature was determined as 1370–1380 °C.

The article presents the results of large-scale laboratory tests of obtaining selenium concentrate from the slime of the sulfuric acid workshop (hereinafter: slime) of the Balkhash copper smelting plant and the extraction of selenium from it into solution. This slime, in contrast to similar slime from nonferrous metallurgy plants, significantly differs in chemical and material compositions, especially in the selenium content (4.6 - 12.5 wt.%). Involving it in the production sphere will increase the production of selenium by more than 20% from that obtained by the plant from copper electrolyte slime. The slime is washed with water to remove sulfuric acid, dried at 105 degrees C. Selenium concentrate was obtained by sequential leaching of slime with solutions of sodium carbonate and nitric acid, selenium from the concentrate was leached with solutions of sodium sulfite (atmospheric pressure) and sodium hydroxide (high pressure, autoclave leaching). Slime, concentrate, cakes, precipitates isolated from solutions, and solutions were analyzed using modern devices of a new generation: An Optima-8300 inductively coupled plasma spectrometer, an Axios X-ray fluorescence spectrometer, an Optima 2000 atomic emission spectroscope, a D8 Advance diffractometer, a D8 Advance infrared spectrometer Avatar 370. From the slime containing, wt. %: 51.2 Pb, 12.5 Se, 3.21 Hg and other elements, a selenium concentrate was obtained with the composition, wt. %: 0.41 Pb, 59.16 Se, 15.4 Hg. The technological indicators for obtaining a concentrate are given, %: concentrate yield - 20.74, Pb recovery - 0.81, Se - 98.23, Hg - 99.50. According to XRD and IR spectroscopy, the concentrate contains elemental selenium and mercury selenide from 10 selenium substances contained in the slime (given in the article). The extraction of selenium from the concentrate into the solution was, %: sodium sulfite - 76.84, sodium hydroxide - 89.65. The pulp from opening the concentrate with sodium hydroxide solution was filtered very poorly. The filtrates contained a colloidal suspension, which could not be filtered off either under vacuum or centrifugation. Therefore, to determine the qualitative and material composition of the filtrates, salts were obtained by evaporating a certain and then neutralized volume to dryness. It was found that only elemental selenium is leached from the concentrate by solutions of both reagents, while mercury selenide remains in the cakes. Moreover, over time, elemental selenium partially precipitated again from selenium-containing sodium sulfite solutions; the salts contain only elemental selenium. An explanation is given for the transition of soluble selenium salts to its elemental state. Despite the higher recovery of selenium from the concentrate by leaching in autoclaves, taking into account the equipment, the complexity of its maintenance, filtration of the pulp and the chemical composition of the filtrate, it is preferable to use the method of leaching with sodium sulfite.