Iron contained in coal fly ash of the Ekibastuz power station is distributed between magnetite and hematite. XRD data showed that ~80 wt % of iron is contained in magnetite and ~20 wt % in hematite. The leaching of iron from CFA by HCl was studied. It was determined that leaching efficiency increased with the increase in hydrochloric acid concentration and temperature. The maximum iron extraction efficiency was 52%. Aluminum is contained in the mullite and was practically not leached. The maximum aluminum extraction efficiency was 3.7%. The kinetics investigation showed that the process of iron leaching was controlled by chemical reaction and diffusion process steps, with an activation energy of 33.25 kJ·mol⁻¹. The aluminum leaching process is controlled by a diffusion process step with an activation energy of 19.89 kJ·mol⁻¹. The reaction order of hydrochloric acid is determined to be 0.9 and 0.23 for iron and aluminum, respectively.

The reduction roasting of the titanium–magnetite concentrate from the Gremyakha-Vyrmes deposit to produce granular cast iron and a titanium–vanadium slag has been studied. A preliminarily beneficiated titanium magnetite concentrate containing (%) 59.3 Fe_tot, 7.7 TiO₂, 1.5 SiO₂, 0.74 V₂O₅, etc. is used for an investigation. The reduction roasting of the concentrate is performed with coke in the temperature range 1450–1550°C on a coal substrate. The influence of temperature, the amount of a solid reducing agent, and the addition of CaCO₃ on the formation of metal and slag phases during reduction roasting and the distribution of iron and vanadium between these phases is studied. At 3% CaCO₃ and 12–14% coke in a charge, a good separation of a slag from metallic granules is found at roasting temperatures of 1475–1500°C, and the degree of iron recovery is ~99%. According to X-ray diffraction, chemical, and microprobe analyses, the slag produced during reduction at 12% coke and 3% CaCO₃ and 1500°C contains (%) 43.6 TiO₂, 19.8 Al₂O₃, 11.9 SiO₂, 7.8 MgO, 3.35 V₂O₅, and so on. Its phase composition is mainly represented by anosovite and Mg–Al spinels with a small amount of glass and Ca–Mg aluminosilicate (pyroxene).

This paper presents the results of benefication by wet magnetic separation of the black titanomagnetite concentrate of the Gremyakha-Vyrmes deposit, containing 9.1% TiO2, 55.1% Fe, 0.6% V2O5 and 6.0% SiO2. The main phases are titanomagnetite, ilmenite and a large number of silicate impurities, such as olivine (2(Mg, Fe)O·SiO2), albite (NaAlSi3O8) and tremolite (Ca2Mg5Si8O22(OH)2). Removal of non-metallic impurities will yield a concentrate suitable for further processing. According to the microscopic analysis it was found that the silicates are in the form of both free grains and clusters. In connection with this, a wet magnetic separation of the black titanomagnetite concentrate was carried out, which allowed separating both free grains of silicates and ilmenite and silicate from the splices after the preliminary grinding of the coarse fraction. As a result of the enrichment carried out, it was possible to increase the iron content in the titanomagnetite concentrate to 59.4% Fe2O3, and V2O5 to 0.74%, and to reduce the sulfur and phosphorus content by 25 and 30%, respectively. At the same time, up to 80% of silicates and practically all ilmenite passed into the non-magnetic fraction. The work was carried out under the state assignment No. 007-00129-18-00.

The results of studying the processing of the titanomagnetite concentrate from the Gremyakha-Vyrmes deposit intended to produce granulated iron and a titanium–vanadium slag are presented. The most suitable method for processing this concentrate is direct iron reduction in a rotary hearth furnace, a key feature of which is the concentration of vanadium along with titanium in a slag suitable for subsequent hydrometallurgical processing to recover vanadium and titanium. To perform the process in a rotary hearth furnace, a protective carbon bed should be used to prevent contact between the molten slag and the refractory lining. The influence of the reducing agent content in a charge and the reduction temperature on the composition of the resulting titanium–vanadium slag is studied under reduction roasting conditions on a protective bed. Reduction with the addition of 16% coke at 1550°C leads to the formation of a slag with the following composition (wt %): 6.2 FeO, 43.9 TiO₂, 20.6 Al₂O₃, 2.3 V₂O₅, 13.2 SiO₂, 7.5 MgO, 4.3 CaO, 0.6 MnO, 0.4 Cr₂O₃, and 0.9 (K,Na)₂O. The high content of titanium in the slag and its presence in the form of anosovite n[Al₂O₃⋅TiO₂]⋅m[(Mg,Fe)O⋅2TiO₂], which is easily decomposed in sulfuric acid, allows for further processing to recover vanadium and titanium. For vanadium recovery, the slag is proposed to be processed using the well-known lime–sulfuric acid method, which involves oxidizing roasting followed by sulfuric acid leaching of calcium vanadates. For titanium recovery, hydrometallurgical methods are suggested to synthesize products suitable for producing both metallic titanium and pigment-grade TiO₂.