The article presents the research results for the synthesis of inorganic sorbents based on manganese oxide compounds. It shows the results of the lithium sorption from brines with the use of synthesized sorbents. The effect of temperature, the molar ratio of Li/Mn, and the duration for obtaining a lithium-manganese precursor and its acid treatment was studied. The sorption characteristics of the synthesized sorbents were studied. The effect of the ratio of the sorbent mass to the brine volume and the duration of the process on the sorption of lithium from brine were studied. In this case, the sorbent recovery of lithium was ~86%. A kinetic model of the lithium sorption from brine on a synthesized sorbent was determined. The kinetics of the lithium sorption was described by a pseudo-second-order model, which implies limiting the speed of the process due to a chemical reaction.

This article presents studies on the ammonium fluoride processing of dusts from the reduction smelting of ilmenite concentrate with separation of silicon to obtain titanium dioxide. Optimal conditions for pyrohydrolysis of titanium fluorides were determined. The effects of temperature and duration on the process were studied. The optimal conditions for pyrohydrolysis of titanium fluorides were a temperature of 600 °C and duration of 240–300 min. The degree of titanium fluoride conversion to titanium oxide was 99.5% at these conditions. Titanium dioxide obtained by pyrohydrolysis of titanium fluorides was purified from iron, chromium, and manganese impurities. The effect of hydrochloric acid solution concentration, S:L ratio, and the process duration on the purification degree of titanium fluoride pyrohydrolysis was studied. The following optimum purification conditions were determined: hydrochloric acid solution concentration 12.5–15 wt%, temperature 25–30 °C, S:L = 1:6÷8, duration 20–30 min. The purified titanium dioxide consisted mainly of anatase. The pigmented titanium dioxide of rutile modification with 99.8 wt% TiO₂ was obtained after calcination at 900 °C for 120 min.

This paper presents the results of research on the development of a technology intended to process electric smelting dusts of ilmenite concentrate with the extraction of silicon and titanium and the production of products in the form of their dioxides. Dusts were processed for silicon separation using the ammonium fluoride method. The optimum conditions for the fluorination and sublimation process of silicon compounds from the electric smelting dust of the ilmenite concentrate were determined: a temperature of 260 °С, a 6 h duration, and mass ratio of dust to ammonium bifluoride of 1:0.5 ÷ 0.9. The sublimation degree of silicon compounds was ~84–91%. The sublimation of titanium fluorides from the remaining sinter was carried out at a temperature of 600 ± 10 °C for 2 h, the mass ratio titanium-containing residue: ammonium bifluoride of 1:0.5, and the degree of sublimation of titanium fluorides was 99%. Iron, manganese, and chromium impurities in the sublimation of titanium fluorides sublimate to a rather low degree. Pyrohydrolysis of titanium fluoride sublimes at 600 °C and allows for the conversion of fluorides into titanium dioxide by 99.5% in 4–5 h. Titanium dioxide of rutile modification with 99.8% TiO₂ was obtained after hydrochloric acid purification and calcination. A technological scheme for the complex processing of dust from the electric smelting of ilmenite concentrates with the production of silica and titanium dioxide is proposed.

This paper presents the results of studies for niobium sorption from chloride solutions with the use of anion-exchange organic sorbents: Amberlite IRA-67, Purolite A-100, AB-17-8, and AN-2FN. Niobium sorption was performed from model niobium-containing solutions. Data on comparative sorption characteristics of the studied sorbents were obtained, and the static exchange capacity of the sorbents, values of distribution coefficients, and extraction degree during the niobium sorption from chloride solutions were calculated. The Purolite A-100 anion-exchange resin exhibited the highest affinity for niobium ions under the conditions studied. Its distribution coefficient was 184 mL/g; the niobium extraction degree was 41.5%. To study the equilibrium sorption of niobium from solution on the Purolite A-100 anionite, three well-known models of isotherms were applied: Langmuir, Freundlich, and Dubinin–Radushkevich. The data obtained confirm the good agreement of the Langmuir model with the results of experiments and indicate that the process takes place in a monomolecular layer on the adsorbent having homogeneous adsorption centers. The optimum conditions of niobium sorption by the Purolite A-100 anion-exchange resin were determined as follows: hydrochloric acid concentration—5–10 wt.%, process temperature—35–40 °C, and duration—40–50 min. The calculated activation energy values for niobium sorption from hydrochloric acid solution in the temperature range of 20–50 °C were 25.32 kJ/mol, which corresponds to the intermediate region corresponding to the transition from the diffusion to the kinetic mode.

The increasing demand for rare refractory metals such as niobium in high-tech industries requires the development of efficient and sustainable methods to extract them from secondary sources, including industrial by-products. This study studied niobium extraction from a fluoride-sulfuric acid solution obtained by leaching a niobium-containing intermediate product. A comprehensive assessment of various organic extractants, including methyl isobutyl ketone, tributyl phosphate, trioctylamine, and Cyanex 923, was performed, focusing on the recoverability of niobium from this solution. Experimental results show that Cyanex 923 is significantly superior to other extractants in efficacy. The extraction of niobium at different concentrations of this extractant in toluene and the ratios of organic and aqueous phases were studied. It is established that an increase in the contact time of the phases does not contribute to additional niobium recovery. Studies show that applying concentrated Cyanex 923 in toluene ensures complete recovery of niobium into the organic phase. The three-stage counter-current extraction shows a slightly higher niobium recovery efficiency than the single-stage process.

The growing global demand for lithium, driven by its pivotal role in battery production, highlights the need for alternative technologies to recover this metal from low-grade and anthropogenic raw materials. This study investigates lithium extraction from aluminosilicate tailings of rare-metal production by sulfate roasting with concentrated sulfuric acid, followed by aqueous and hydrochloric acid leaching. Mineralogical analysis confirmed lithium mainly in muscovite and biotite (isomorphic substitutions) and partly as spodumene within the aluminosilicate matrix. The optimal parameters of thermochemical treatment were determined as 300 °C for 1 h at a liquid-to-solid ratio of 1:6. Subsequent aqueous leaching (90 °C, 1 h, L/S = 6:1) achieved a lithium recovery of 82.3%, while HCl proved less effective. Using response surface methodology (RSM) and a central composite design (CCD), a regression model was developed predicting up to 93.4% lithium extraction at 90 °C, a liquid-to-solid ratio of 10:1, and a leaching duration of 75 min. The calculated values showed good agreement with experimental data obtained at 90 °C, L/S = 10:1, and 30 min leaching, yielding 91.92% lithium recovery. These results confirm the efficiency of the proposed thermochemical approach and provide a scientific foundation for its further development and industrial scale-up.