An overview of the composition of wolframite ores of the Akchatau deposit and the technologies for processing concentrates using NaOH and Na₂CO₃ by hydro- and pyrometallurgical methods is given, and the disadvantages associated with both the technology and the equipment are noted. To develop a technology for processing Akchatau wolframite concentrates, samples of ore materials were taken, the chemical and mineralogical composition of the samples was studied, and enrichment was carried out to obtain rich concentrates. The kinetics of the sintering of the wolframite concentrate with soda was investigated, the dependences of the degree of transformation of the tungsten minerals into sodium tungstate were obtained, and the rate constants, the order of the reaction, and the values of the apparent activation energy were calculated. The results of sintering an enlarged sample of wolframite concentrate with soda in a muffle furnace are presented. After the subsequent leaching, studies were carried out to purify the obtained solutions of sodium tungstate from the impurities while eliminating the operations of the neutralizing solutions through the use of electrodialysis with an MK-40 cation-exchange membrane. The scheme of processing the wolframite ores of Akchatau is proposed.

The article theoretically justified and experimentally confirmed the possibility of implementing the process of the electric melting of Satbayevskiy ilmenite concentrates with new fluxing additives based on oxides and nitrides of aluminium, calcium, and boron. They also include boron carbonitride CNV that will expand the raw material base of Kazakhstan titanium production by involving local substandard material in the process, as well as the technical and economic performance of electric melting. In order to conduct experiments in the area of extremum, the ilmenite concentrate from the Satbayev deposit was taken. Furthermore, the optimum conditions of the electric melting of Satbayevski ilmenite concentrates (such as the process temperature of 1550–1600 °C, the reducing agent consumption of 8–10% of the concentrate mass, and the duration of 90 min), using new fluxing additives, were selected. As a result of the experiment (performed at a temperature of 1600 °C), it has been found that the introduction of 3 to 6% of fluxes in the charge of electric melting promotes the reduction of iron oxides from 45 to 80% and achievement of the extraction of titanium oxide in slag of up to 83.5–90.1%. The addition of 6% boron oxide and carbonitride in the charge of electric melting reduces the melting temperature of the charge to ~1400–1450 °C and the melting time to 90 min. It also creates conditions for a quieter electric melting mode.

This work presents the results of thermodynamic and technological studies of the reducing electric smelting of low-quality and difficult-to-recover Satpaevsk ilmenite concentrates in the presence of sodium carbonate (Na₂CO₃) as a flux. In the course of thermodynamic studies, according to literature data and using the HSC Chemistry 8 program, the state diagrams of the FeO-TiO₂, FeO-Fe₂O₃-TiO₂, Na₂O-TiO₂, Na₂O-SiO₂, Ti-C-O and Fe-C-O systems and the possibility of reactions in the temperature range of 500–1600 °C (without smelting of titanium, 1678 °C), taking into account phase transitions, were studied. The article also presents the results of technological studies of the above process. In general, the research results showed the possibility of using sodium carbonate (Na₂CO₃) as a flux in the charge of the electric smelting of low-quality Satpaevsk ilmenite concentrates. It has been established that the addition of 4.5–5% of Na₂CO₃ allows for reducing the temperature and the duration of electric smelting, the viscosity of the melt, and the complete separation of the smelting products (cast iron from slag).

Developing efficient technologies for vanadium recovery from refractory raw materials requires an understanding of the thermodynamic and electrochemical behavior of vanadium during acid leaching. This study investigates hydrogen peroxide-assisted sulfuric acid leaching of roasted vanadium-bearing ores from the Greater Karatau region using standard Gibbs free-energy calculations, Pourbaix diagrams, and electrode-potential calculations in HSC Chemistry 8.1.5. The reactions of the model vanadates NaVO₃, Na₃VO₄, and Na₄V₂O₇ with H₂SO₄ exhibited negative ΔG⁰ values at 25 and 65 °C, indicating thermodynamically favorable acid dissolution. The more negative ΔG⁰ values in the presence of H₂O₂ largely reflect the thermodynamic contribution of its decomposition and do not directly quantify its effect on vanadium extraction. Pourbaix diagrams indicated VO₂⁺ as the thermodynamically stable dissolved V(V) species under strongly acidic oxidizing conditions. The higher equilibrium potential of the H₂O₂/H₂O couple relative to VO₂⁺/VO²⁺ indicates the thermodynamic feasibility of oxidizing residual V(IV) to V(V). Competing reactions of Na₂CO₃ with mineral–matrix components may consume the sodium reagent during roasting. The calculations were qualitatively consistent with previously obtained vanadium extraction of 78.70–80.15% from roasted Balasauskandyk and Kurumsak ores at 65 °C. An integrated five-stage physicochemical model is proposed, requiring direct experimental validation of individual stages.

Lead-bearing dusts from metallurgical processes are hazardous secondary resources due to their complex composition and toxicity. At the same time, their high lead content makes them a promising feedstock for resource recovery. This study proposes an energy-efficient electrosmelting approach based on the partial substitution of sodium carbonate with calcium-rich industrial waste (sugar-industry defecate). Thermodynamic analysis and equilibrium modeling of the Pb–Sb–Fe–Na–Ca–Si–S–Cl–As system were performed in the temperature range of 200–1200 °C using Outotec HSC Chemistry. The results indicate that under equilibrium conditions approximately 90–95% of lead is concentrated in the metallic phase (~56 kg from ~60 kg in the feed), while antimony is co-recovered (~1.9–2.0 kg). The slag is dominated by calcium silicates, primarily Ca₂SiO₄, confirming the important role of CaO in slag formation and impurity fixation. Chlorine is predominantly bound as NaCl and partially as CaCl₂, while sulfur is distributed between Na₂S and Na₂SO₄. A significant portion of arsenic is predicted to be retained in the slag as calcium and sodium arsenates (Ca₃(AsO₄)₂ and Na₃AsO₄), whereas its volatilization is thermodynamically negligible under equilibrium conditions. Preliminary experimental results are generally consistent with the thermodynamic predictions and confirm the feasibility of partially replacing Na₂CO₃ with sugar-industry defecate. The proposed approach contributes to reducing the consumption of conventional fluxes and promotes the utilization of industrial waste within a circular-economy framework.

The article presents experimental data on the electrochemical oxidation of vanadium-bearing ore with the aim of increasing the efficiency of vanadium extraction during subsequent hydrometallurgical processing. Three different charge compositions were studied during the preliminary oxidative roasting stage, differing in the type of oxidizers used: calcined soda, sodium chloride, and their mixture in a mass ratio of 9:1. Electrochemical oxidation was carried out in a sulfuric acid medium using a membrane electrolysis cell equipped with an MK-40 type diaphragm. Experimental studies were conducted by varying key technological parameters: H₂SO₄ concentration (5–15%), solid-to-liquid phase ratio (1:2–1:5), temperature (25–85 °C), process duration (0.5–3 h), and current density (100–1000 A/m²). It was found that preliminary roasting promotes the conversion of vanadium into higher oxidation states, predominantly V⁵⁺, which significantly increases its solubility during subsequent electrochemical treatment. For the first time, the kinetic patterns of electrochemical vanadium leaching were identified, as well as the limiting mechanism of the process, associated with the formation of poorly soluble oxide films on the ore surface. Optimization of the electrochemical oxidation parameters allowed us to achieve vanadium extraction into solution up to 92%.