The search for effective reagents for the flotation of non-ferrous metals from mineral raw materials is an actual problem. The article discusses the results of research on improving the technology of processing copper–molybdenum ores using a combined collector emulsion. The object of research was copper–molybdenum ore containing 0.43% copper and 0.0089% molybdenum. Copper minerals are represented by chalcopyrite 1.2% and chalcocite 0.015%, while covellite is present in a much smaller amount. The main mineral of molybdenum is molybdenite. The cycle of collective copper–molybdenum flotation includes the main flotation, control flotation, and three re-cleanings of concentrate. In the basic mode, the collective Cu–Mo concentrate was produced with the copper content of 16.25% with an extraction of 77.79% and with the molybdenum content of 0.45% with an extraction of 79.38%. Reaflot, thionocarbamate, and butyl xanthate were used in a ratio of 1:3:15 to prepare a combined collector agent. In order to produce the combined collector emulsion, the best degree of dispersion, which amounted to 99.5% of particles with a size of −3.6 μm, was chosen. During the Cu–Mo ore flotation using an optimally dispersed microemulsion of the combined collector agent, a collective Cu–Mo concentrate was obtained with a copper content of 18.2% with an extraction of 83.58%, and with a molybdenum content of 0.49% with an extraction of 88.46%. The use of a combined collector agent increases the extraction of copper into the collective Cu–Mo concentrate by 5.79%, while the extraction of molybdenum increases by 9.08%.

This article presents the results of cyanide leaching of gold-containing concentrate using the trichlorocyanuric acid (TCCA) oxidizer. Gold-containing concentrate was obtained from a gold tailings sample from a gold recovery factory (GRF) in one of the deposits of Kazakhstan that have not previously been studied for concentrability. According to X-ray phase analysis and energy dispersive spectrometry (DSM) data, the main compounds in the tailings sample under study are pyrite FeS₂, quartz SiO₂, calcite CaCO₃, albite NaAlSi₃O₈, muscovite KAl₂Si₃AlO₁₀(OH)₈, dolomite CaMg(CO₃)₂, and oxidized iron compounds. Microscopic studies of the concentrate have established the presence of ultrafine gold with sizes from Au 0.9 to 10.2 μm in pyrite. Obtaining the gold-containing concentrate with a gold content of 15.95 g/t is possible according to the enrichment scheme, which includes centrifugal separation, classification according to the fineness class −0.05 mm, additional grinding of hydrocyclone sands to a fineness of 90.0–95.0% of the class finer than 0.050 mm, and control centrifugal separation. Since pyrite in technogenic raw materials is the main gold-containing mineral, this paper presents studies on the oxidizability of pyrite with the TCCA oxidizer. The results of studies on the oxidation of pyrite using the TCCA oxidizer show the products of its hydrolysis oxidize pyrite with the formation of various iron compounds on its surface. Pretreatment of gold-containing concentrate with oxidizer TCCA for 3 h before the cyanidation process (20 h) allows for an increase in the recovery of gold in the solution by 5.8%.

The results of experiments on manganese pellets based on beneficiated manganese-containing sludge and the melting of ferromanganese alloy with their use are presented. Via beneficiating manganese-containing sludge (16.32% Mn), a concentrate with a manganese content of 35.2% was obtained. The composition of the charge used to prepare manganese pellets is proposed, and the conditions affecting their strength are determined. It was established that manganese is present in the composition of calcined pellets in the form of jacobsite MnFe₂O₄ and hausmannite Mn₃O₄. The formation of a phase of ferrobustamite (Ca_0.79Fe_0.21)SiO₃, a ferrosilico-calcium binder that helps increase the strength of pellets, was found. Ferromanganese alloy was obtained as a result of the melting of calcined manganese pellets in a high-temperature Tamman-type unit. The Fe-Mn-Si alloy corresponds to the DIN 17 564 as-grade FeMn70Si in terms of manganese (63.76%) and silicon (17.21%) contents. The content of limiting impurity elements—carbon and phosphorus—in ferromanganese alloy is within acceptable limits. The structure of Fe-Mn-Si alloy and slag formed during their smelting process were studied.

The article presents the results of process research for the concentration of a previously unstudied sulfide gold-containing ore from one of the deposits of Kazakhstan. In an X-ray phase analysis, the following minerals were identified in the sample: quartz, chlorite, feldspar, muscovite, and sulfides. The material composition studies demonstrate that up to 52.22 % of gold in the ore sample is free or occurring in intergrowths with pyrite. The results of scanning electron microscopy and X-ray spectral microanalysis indicate that gold is represented by inclusions of petcite and electrum grains in pyrite, with the dimensions of thousandths of a millimeter. It is a specific mineralogical and processing feature of the test ore sample that the gold and pyrite in the sample are mainly represented by finely disseminated elements with the size of thousandths and first tenths of a millimeter, which requires fine grinding of the feed material. The influence of the ore grinding fineness on gold recovery into the flotation concentrate was studied. It was found that, with the grinding size of gravity tailings of 90–95 % class –0.044+0 mm, the gold content in the flotation tailings was reduced to 0.5 g/t. The flotation concentration results obtained confirm the presence of fine and ultrafine gold in the ore and the requirement for applying a staged flotation technology. Under the optimal conditions, the combined gold-bearing concentrate with the gold content of 44.0 g/t was obtained using a gravity/flotation concentration process, with the recovery of 92.86 %.The research was completed with the financial support of the Science Committee of the Ministry of Education and Science of the Republic of Kazakhstan under program No. BR0523640.

The Donskoy Mining and Concentration Complex slime tailings concentration studies were conducted. By means of 0.2+0.071 mm slime size fraction concentration on concentrating table in open cycle, chromite concentrate with 48.8 % Cr2O3 mass fraction, yield being 29.42 %, was produced. Chromium oxide losses in tailings were 18.62 %, with 6.5 % Cr2O3 mass fraction. By means of –0.071+0 mm slime size fraction concentration on Knelson centrifugal separator in open cycle, chromite concentrate with 42.8 % Cr2O3 mass fraction, yield being 48.5 %, was produced. Chromium oxide losses in centrifugal separator tailings were 19.16 %, with 15.75 % Cr2O3 mass fraction. By means of gravity concentration circuit, total chromite concentrate, containing 43.58 % Cr2O3, was produced. Total concentrate yield was 37.0 %, with Cr2O3 recovery into concentrate — 62.45 %. Chromium total losses in tailings and in +0.2 mm size fraction will amount to 24.98 %, with its mass fraction of ~12.0 %. A technology of slime tailings activation in soda solution containing 120 g NaHCO3 per 1 dm3, at temperature +120 °С and duration of 60 min, prior to gravity concentration, resulting in slime phase composition change, has been developed. Slimes concentration on concentrating table and centrifugal separator, following preliminary activation in soda solution, permitted to produce joined salable concentrate with 51.3 % Cr2O3 mass fraction, which is 8.42 % higher, if compared to processing without activation.

The article presents the results of laboratory studies on ore flotation of copper-molybdenum deposits in the East Kazakhstan region of Kazakhstan using a combined reagent. It is required to use new beneficiation methods and flotation reagents to increase the efficiency of flotation of ore minerals to improve the processing of finely disseminated ores. The problem can be solved by using a combination of different collectors. The objective of the research is to increase the recovery of copper and molybdenum during flotation of copper-molybdenum ore using a combined reagent. The studied ore sample contains 0.42% copper; 0.009% molybdenum. A combination of sodium butyl xanthate, TS-100 thionocarbamate, and reaflot in a ratio, in %: 15: 3: 1, was used as a combined flotation reagent. The combined flotation reagent was preliminarily passed through an ultrasonic homogenizer JY96-IIN to obtain a reagent microemulsion before flotation. The optimum dispersion time for the combined flotation reagent is 60 seconds. At the same time, 99.4% are microemulsion particles with a particle size of fewer than 3.7 microns. The use of a combined reagent microemulsion increases the extraction of copper into the copper-molybdenum concentrate by 3.69%, and the extraction of molybdenum by 6.05%. The copper content in the copper-molybdenum concentrate increases by 1.26%. The copper content in the flotation tailings decreases from 0.07 to 0.056%. The consumption of the combined reagent is reduced by 15% in comparison with the basic butyl xanthate.