Prolonged development of ore deposits, ore beneficiation and metallurgical smelting of concentrates result in the accumulation of wastes, forming large-scale dumps and sludge ponds negatively affecting the environment and human health. The creation and introduction of industrial waste treatment technologies will make it possible to dispose of them with the production of valuable commercial products and improve the environmental situation. In a research article on chemical beneficiation with the subsequent gravitational beneficiation of chrome-containing slurry tailings of Donskoy Ore-Mining and Beneficiation Plant (DOMBP), which processes chrome ore of the Kempirsay deposit in the Republic of Kazakhstan, Aktobe region, having considerable stocks of such waste is presented. The chromium-containing slurry tailings were heat-treated at 1100 °C, sintered with ammonium sulfate, water and sulfuric acid in a defined ratio at 300 °C, and then the resulting sinter was leached with water at 90 °C. The cake was washed to separate the liquid part from the solid part. The solid precipitate is the chromium-containing cake, which is upgraded by gravity concentration to a grade concentrate for ferroalloy production. When the solution is evaporated, magnesium sulfate with ammonium sulfate forms the Tutton’s salt ammoshenite. High-throughput chromium oxide extraction of 93.9% was achieved as a result of the research, and three products were obtained: (1) standard chromium concentrate with 49.48% Cr₂O₃ content, (2) ammoshenite used as a nitrogen–magnesium fertilizer and (3) forsterite, which is concentrated in gravity concentration tailings and can be used in the production of refractory minerals.

The results of research on the processing of ilmenite concentrate from the Obukhovskoye deposit are presented in this article. As the concentrate has a high chromium content, the study involved converting the iron into metal and the titanium into slag through the addition of soda. Positive results were obtained during the smelting of the ilmenite concentrate, and a one-stage smelting mode was established. This mode involved an increase in the temperature up to 1700 °C, with a heating step of 10 °C/min and using an argon supply. The holding time at this temperature was 30 min, followed by cooling to 700 °C in argon. The optimal parameters for sintering the non-magnetic fraction with soda and leaching the sinter with water and hydrochloric acid solution were also determined.

This study presents the results of the enlarged laboratory research on the melting of calcined composite pellets for ferrochrome obtained from fine-dispersed conditioned chrome concentrate containing 50.3% Cr₂O₃. This is a product of the gravitational beneficiation of waste sludge tailings from the Dubersay tailings dump at the Donskoy Mining and Processing Plant (DMPP) of JSC “TNC Kazchrome”. The composition of the charge for obtaining composite pellets consisted of 88.5% of chrome concentrate, 3% of mineral part of refined ferrochrome slag (RFC), 4% of ferruginous diatomite, 3% of coke and 1.5% of liquid glass. The initial charge was pelletized on a laboratory pelletizer to a size of 6–10 mm, dried at room temperature for 24 h and fired at 1200 °C for 60 min at a heating rate of 15 deg/min. On the basis of the developed composite annealed pellets, studies on the production of high-carbon ferrochrome at different melting temperatures were carried out. The results showed that with an increase in temperature from 1750 to 1850 °C, the iron–chromium phase in the composition of the alloy increases from 45.2 to 50.1%, the chromium carbide phase decreases from 23.7 to 11.3% and the chromium–iron phase increases from 7 to 11.2%. The carbon content in the alloys at temperatures from 1750 to 1850 °C varies from 7.2 to 8.94%, respectively. The maximum chromium content of the alloy is 64.82% with a melting point of 1850 °C; this alloy can be classified as FeCr60C90LP grade ferrochrome according to the international Chinese standard, which has no more than 0.03% phosphorus and no more than 0.1% sulfur.

In the wastes of the joint stock Aktobe Ferro-alloy plant, there are over 12 mln ton of refined ferrochrome slags that contain significant amounts of the metal dorebead. Dust-like mineral part of the slagof refined ferrochrome mainly consists of di-calcium silicate contaminated with poisonous compounds of hexatomic chrome. The metal can be separated from the mineral part by air separation and screen sizing, while the mineral part of the slag can be very efficiently mixed with Kazakhstani ferri-ferrous diatomite(opoka) and fluid glass and thermally processed to obtainglass-like heat insulating material. Being heated up to t ≥ 1000оС Di-calcium silicate interacts with finely divided ferrumdioxyde and hydrated silica to formporous glass phase of ferro-silica-calcium compounds that strengthen the heat-processed pellets against any mechanical influence. It was found that the optimum conditions for pellet production are: the ratio of diatomite ore to the mineral part of the RFX slag 2: 1, because at other ratios, the formation of a rolling glass-phase pellet is observed, the processing temperature is 1050 ° C, because at lower temperatures there is practically no formation of a glass phase. The compounds of hexatomic chrome het destroyed at heating and dissolve in the glass mass where they become insoluble. It is shown that the mineral part of RFX slags due to the relatively high content of calcium and silicon oxides is a valuable raw material for the production of granular porous heat-insulating material. The proposed technology makes it possible to extract a high-value metal component of RFX slags in the form of an expensive metal concentrate of refined ferrochromium and to obtain porous pellet fillers for reinforced concrete products.

A combination technology for high-silica raw material treatment by means of fluoride method is presented, and the process conditions are determined for production of ammonium hexafluorosilicate, amorphous silica, aluminium fluorine compounds and ferrous concentrate. The process of raw material and ammonium bifluoride sintering, providing for sinter mix separation into gas-phase as ammonium hexafluorosilicate and aluminium nonvolatile fluorine compounds and iron, concentrated in sinter solid phase, was investigated. The optimal sintering process conditions were established: the ratio of high-silica raw material and ammonium bifluoride, step-up process temperature increase, duration. It was determined, that with interaction of ammonium hexafluorosilicate and ammonium solutions, amorphous silica is formed. Mother solution, following precipitation of amorphous silica, is subjected to isothermal crystallization through concentration by evaporation with a view to regeneration of ammonium bifluoride. In the pyro-hydrolysis process, with heating of sinter, upon water steam and temperature effect, metal fluorides pass into metal-oxides and sublimation separation of fluoride with formation of fluorohydrogen acid is accomplished. In order to separate aluminium and iron oxides in concentrate, autoclave-alkali-solution leaching is applied, producing alkali-aluminate solution and ferrous concentrate. Alkalialuminate solution interaction with fluorohydrogen acid solution produces cryolite, used in metallic aluminum electrolysis.

Nitric–sulfuric acid leaching of substandard molybdenite concentrate from the Aktogay deposit (Kazakhstan) transfers only part of the molybdenum into solution: under the baseline process conditions, approximately 30% of the sulfide is oxidized, while roughly 70% of the molybdenum remains in the leach cake as unoxidized molybdenite. Molybdenum is thus distributed between two technologically dissimilar streams—the solution and the solid cake—and complete recovery requires processing of both. In the present work, both branches are combined into a single integrated scheme applied to a common feedstock. In the first (sorption) branch, molybdenum is concentrated from the solution. Among four anionites tested (Purolite A-100, Amberlite IRA-95, AV-17-8, and Lewatit M-800), Amberlite IRA-95 exhibited the best performance (sorption 84.21%, desorption with ammonia solution 82.15%, and selectivity coefficients Mo/Cu = 304.7 and Mo/Fe = 192.2). To remove iron, copper, and silicon impurities prior to concentration, a two-stage scheme is proposed incorporating preliminary impurity sorption on KU2-8 cationite (separation coefficients Mo/Cu = 5.7, Mo/Fe = 3.58, Mo/SiO₂ = 2.17). In the second (roasting) branch, the cake is processed: the cake, granulated with molasses, is subjected to oxidative roasting at 600 °C for 150 min under an oxygen flow (2 dm³/min) until complete desulfurization is achieved, after which the calcine is leached with ammonia. Both streams converge into a single product: the ammoniacal desorbate from the sorption branch and the alkaline solution from the calcine are acidified with nitric acid to pH 2–3, ammonium paramolybdate is precipitated by evaporation, and the precipitate is calcined to molybdenum trioxide. The resulting MoO₃, according to XRD data, is represented predominantly by orthorhombic α-MoO₃ (molybdite). Overall, the proposed two-branch scheme closes the molybdenum material balance for the substandard concentrate from both sides and outlines a technically straightforward pathway for bringing stockpiled technogenic raw material into processing.