The accumulated amount of lead-zinc ore flotation tailings in the dumps of concentrating plants today can be considered as independent man-made deposits. In addition to their resource value as sources of lead and zinc, as well as associated gold, silver, cadmium, selenium and other metals, tailings are an environmentally hazardous source of heavy metal pollution of ground and surface waters. The environmental hazard of stale tailings is exacerbated by the fact that they occupy large areas that cannot be used for agricultural or other purposes of the national economy. Wastes of flotation enrichment of lead-zinc ores significantly differ from the source material not only in the content of minerals, but also in the degree of oxidation of their surface, fractional composition, and the presence of a significant amount of mineral intergrowths. In view of this, the use of existing flotation technologies is ineffective for obtaining standard lead and zinc concentrates from enrichment tailings. This paper describes the technology that has been developed for processing of zinc and lead-bearing enrichment wastes by sulfidizing roasting followed by magnetic and flotation concentration of cinders. It was found that, as a result of the sulfidizing-pyrrhotizing roasting process, the flotation ability increases for lead compounds and decreases for iron compounds, while the magnetic susceptibility of lower iron sulfides formed during roasting increases. It has been established that sulfidizing takes place with sufficient completeness, and during subsequent flotation, it is possible to extract up to 95% of zinc and up to 80% of lead into sulfide concentrate. These results have a technological advantage in contrast to the other methods that have been used. It was found that at roasting temperatures of 700-800 °C, pyrrhotites have a maximum magnetic susceptibility of 3.75, 5.43 and 2.18 SI units for Fe 0.855 S, Fe 0.888 S and Fe 0.909 S, respectively. Technological recommendations are acceptable for similar raw materials.

Zinc and lead are one of the most used metals in the world. The average annual growth rate of the zinc market is about 3.5%. Half of the zinc consumed in the world is used as electroplating, more than 30% is spent on the production of zinc alloys, including for the production of brass and bronze. At present, polymetallic ores are the main raw material for the production of lead and zinc. The production technology includes flotation enrichment of the initial ore raw material with subsequent pyrometallurgical or hydrometallurgical processing of the resulting concentrates. Unfortunately, the reserves of well-enriched sulfide ores are declining, which leads to the need to involve mixed and oxidized ores in the production. Obtaining zinc is also a relatively expensive process, one of the stages of which is the roasting of zinc sulfide concentrates in a fluidized bed furnace using oxygen-enriched air blast. In this regard, technologies aimed at processing hard-to-cut oxidized lead-zinc ores, as well as improving the process of roasting in a fluidized bed, are topical and in demand today. The article presents the technology and method for processing oxidized lead-zinc ore, including high-temperature sulfidizing roasting of oxidized compounds of lead and zinc, the results of roasting carried out in the presence of a high-sulfur sulfidizing agent in the form of pyrite (sulfur content is 45.15%), at molar ratios of metal oxide to pyrite NZnO/NFeS 2 and NPbO/NFeS 2 equal to 0.1-0.14 for sulfidizing in an air-blown fluidized bed furnace at a flow rate of 10 to 20 l/min, at a temperature of 750-800 °C, with a roasting time of 30-45 minutes. As a result of sulfidizing roasting, the degree of sulfidization reaches 88%, and the subsequent extraction of zinc from the non-magnetic fraction into a froth product in an open flotation cycle is 90% with a content of 23.4% zinc.

The SX-EW technology is more effective for the production of copper from refractory oxidized ores for the present day. A wide range of modern extractants is currently offered on the market for the extraction of copper from leaching solutions and its choice is a very important issue in the production of copper using the SX-EW technology. The aim of this work was to study the extraction properties of modified extractants of the Acorga series (5747, 5910, 5640) and unmodified extractant Lix 984N. It has been established that extractants Acorga 5640 and unmodified Lix984 have high selectivity to copper. During the extraction of copper from the productive solution with the use of all extractants, the formation of a third phase, cruda, is observed. Physical and chemical studies have established that the steak contains a large amount of silica, magnetite, hematite and anglesite. The distribution of iron, silica, and copper ions during extraction was studied and it was found that the extractant Lix 984 N (10%), then Acorga5640 (10%), has a high selectivity to copper/iron and copper/silica. Acorga 5640 (10%) is an effective extractant of copper from the productive solution of the Almaly deposit. It is observed that the amount of crud formed during extraction also depends on the rate of phase mixing, the number of revolutions of the stirrer, at a speed of 350-450 rpm, the formation of crud is 0.73%. The addition of the Acorga CR60 reagent in the amount of 5-10 ppm leads to a 3-3.2-fold decrease in the volume of the crud.

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The paper presents the results of the development of a technology for processing of pyrite concentrates with the extraction of nonferrous metals, iron and sulphur using the technology of two-stage dissociating roasting with hydrometallurgical processing of cinder. Hydrometallurgical processing of cinder was studied. It comprised of two steps - cinder leaching and extraction for processing multi-component solutions. Based on the results of laboratory tests, balance experiments were conducted on the chemical enrichment of nickel-poor and cobaltpyrite concentrates. The experiments included (i) Thermal decomposition of pyrite concentrate at a temperature of 660 - 700° C and a duration of 60 minutes, an oxygen flow rate of 30-50% of the stoichiometric; (ii) Leaching of the cinder with a HCl solution at a temperature of 90°C, liquid: solid = 8:1, pH 2,7 and a duration of 60 minutes; (iii) Filtration of pulp and precipitation of nickel and cobalt from solutions with iron sulphide. During leaching, mainly iron is transferred to the solution; nickel and cobalt remain in a small amount of cake (the cake weight is reduced by almost ten times compared to the thermal decomposition product). As a result, up to 99% of non-ferrous metals are extracted into the concentrate, and the solution becomes more concentrated in iron.