RU2359046C1 - Processing method of copper sulphide materials on blister copper - Google Patents

Abstract

FIELD: metallurgy.

SUBSTANCE: method can be used for melting of sulphide copper materials on blister copper, including nickel-bearing. Method of continuous processing of copper including nickel-bearing sulphide materials includes oxidising fusion in furnace of sulphide material jointly with flux with receiving of blister copper, slag, concentrated by SO₂ gases at specific consumption of oxygen in the range 150-240 nm³ per 1 ton of processing dry sulphide material. After melting it is implemented depletion of received slag. Slag depletion is implemented uninterruptedly on reducing stage in reducing zone of furnace at usage for it mixture of oxygen-containing gas and hydrocarbon fuel at discharge coefficient of oxygen (α) in the range from 0.5 up to 0.9.

EFFECT: increasing of direct extraction of copper into blister.

5 cl, 2 dwg, 3 tbl, 1 ex

Description

The present invention relates to the field of non-ferrous metallurgy, in particular to a method for processing copper, including nickel-containing, sulfide materials into blister copper.

A known method (US 4802916, IPC С22В 15/00) for producing blister copper from copper concentrate from separation of Feinstein, according to which the concentrate in the first stage is subjected to weighted conversion to obtain blister copper, which contains 4-6% Ni and 2-3% S and then the semi-rough copper is converted to blister copper in a finish conversion converter using overhead purging with oxygen-containing gas and bottom stirring with an inert gas (nitrogen). The disadvantages of this method include the insufficiently high degree of sulfur utilization, since its utilization at the finishing conversion operations is not economically feasible, as well as the formation of rolled nickel slags, which are currently being converted into a liquid state by treatment with copper-nickel matte with the addition of expensive ferrosilicon.

A known method for the continuous melting of sulfide copper-containing concentrates (SU 1734389, IPC 6 C22B 15/00), which involves using a top supply of oxygen-containing blast to obtain a three-layer melt bath consisting of slag, matte and semi-copper. The height of the matte layer in the furnace is kept constant by maintaining a stable ratio of sulfide charge and oxygen, and the height of the slag and semi-copper layers is controlled by discharge from the furnace. The disadvantages of this method include the need to convert the resulting copper containing 1.5-2.5% sulfur and 5-6% nickel. When converting, rolled-up nickel copper-containing slag is formed, which is extremely inconvenient during further processing (not crushed, difficult to test) and with which a significant amount of mechanically entangled metallic copper is lost. In addition, the practical impracticability of sulfur utilization for finishing conversion operations should be attributed to shortcomings, and the technology of smelting in three-layer mode on semi-copper copper allows to utilize only 93-95% of sulfur contained in sulfide concentrate.

A known method of continuous conversion of liquid and solid sulfide materials (RU 2071982, IPC 6 C22B 15/06). The method includes loading sulfide materials into the furnace, supplying oxygen-containing blast to the matte-metal-slag emulsion layer through horizontal blowing devices located uniformly in the side walls of the furnace, and removing liquid conversion products from the furnace. The disadvantage of this method is the possibility of periodic formation between the layers of slag and copper of an intermediate layer of matte. The presence of an intermediate layer of matte inevitably leads to the formation of not rough, but semi-rough copper. Since periodic semi-copper production is allowed, this continuous conversion technology should also include the obligatory final conversion operation in this case. The disadvantages of this conversion are identical to the disadvantages of the two above methods: the inappropriateness of sulfur utilization for finishing operations and the formation of rolled nickel slags in the case of processing nickel-containing copper materials. In the case of obtaining not sinter, but blister copper in the furnace, the technology should be considered as a drawback of low direct copper extraction into blister copper, since the technology does not provide for the depletion of slag formed during oxidative smelting.

Closest to the invention according to the technical and technological essence is a method (RU 2169202, IPC 7 С22В 15/00) for processing copper concentrate into blister copper, including loading the charge, blowing the melt to form slag and blister copper, and releasing these smelting products. In this case, oxidative smelting of the concentrate is carried out at a ratio of concentrate loading and supply of oxygen-containing gas in the range of 1.0-1.3 from theoretically necessary for the oxidation of all sulfur and impurities (Fe, Ni, Co) to oxides, and before slag is released, which is carried out periodically , carry out the operation of depletion of slag, changing the ratio of the loaded concentrate and oxygen-containing blast in the range of 0.3-1.0 from theoretically necessary for the oxidation of all sulfur and impurities (Fe, Ni, Co) to oxides. At the same time, they achieve a reduction in the content of copper oxide in the slag from 35 to 22%.

This method has the following disadvantages.

A sufficiently high residual copper content in the slag after the depletion operation. This is due to the fact that upon reduction of the slag with a sulfide concentrate, iron, cobalt and part of the nickel concentrate pass into the slag by exchange reactions, which, against the background of a decrease in the concentration in copper slag, leads to a noticeable increase in the concentration of iron and nickel in the slag. When attempts are made to restore copper slag more deeply, the concentrations of iron and nickel in the slag increase even more and solid iron-nickel spinel precipitates out of the slag as a result of saturation of the homogeneous silicate melt with it. The consequence of the presence of significant amounts of hard spinel in the slag is, as you know, the inevitable foaming of slag and the creation of an emergency.

The variability of the levels of slag and blister copper in the furnace, which is associated with the frequency of alternating operations of oxidative smelting and depletion of slag in one furnace space. Since the blister copper zone in the furnace is lined with refractory bricks (the slag zone is coffered), the inconsistency of the slag and copper levels suggests periodic contact of the aggressive, due to the high content of oxide copper (copper concentration reaches 35 wt.% At the oxidation stage), slag melt with refractory lining with quick wear of the latter.

The combination of two processes in one furnace space (oxidative smelting and slag depletion) leads to inconsistency in the composition of the smelting products (copper, slag, exhaust gases) and makes automatic control of such technology very difficult.

The objective of the invention is to develop a method for the continuous processing of sulfide copper materials into blister copper. The technical result is to increase the direct extraction of copper into blister copper, maximize sulfur utilization, maintaining slag and copper levels in the furnace at a constant level.

The claimed technical result is achieved in that in a method for processing copper sulfide, including nickel-containing materials, including oxidative smelting of sulfide material together with fluxes to produce blister copper, slag, SO ₂ concentrated gases with specific oxygen consumption in the range of 150-240 nm ³ per 1 ton of dry sulphide material being processed and depletion of the oxidizing stage slags, according to the invention, slag depletion is carried out continuously in the reduction zone of the furnace using an oxygen mixture odosoderzhaschego gas and hydrocarbon fuels with a coefficient of oxygen consumption (α) in the range from 0.5 to 0.9.

It is possible to additionally supply coal to the oxidation stage of the process.

An additional supply of hydrocarbon fuel is possible at the oxidative stage of the process.

At the oxidizing stage of the process, together with a SiO ₂ -containing flux, CaO-containing flux can be supplied based on the production of slags with a SiO ₂ / CaO ratio of 3/1 to 1/1, and the total flux consumption for oxidative smelting is determined from the considerations of maintaining oxidative melting in the slag stage total concentration of iron, Nickel and cobalt is not more than 30 wt.%.

At the stage of depletion of slag along with hydrocarbon fuel can be fed hard coal with a flow rate of 0.1-5% by weight of the recovered slag.

In our proposed method, sulphide material (concentrate, granular or crushed matte) is melted in two stages, each of which is carried out in a separate zone of the furnace with separated gas spaces. At the first stage, the oxidative melting process is carried out, at the second stage, the reductive depletion of the slag melt is carried out. The process is carried out continuously, smelting products: slag and blister copper, are produced continuously. For example, a siphon device may be provided for the release of copper, the slag is continuously drained through a drain threshold located at a certain furnace height. The continuity of the oxidation and reduction stages creates the prerequisites for maintaining the constancy of the levels in the furnace of slag and blister copper, which is an important advantage of the process that allows you to reliably protect the refractory lining from destruction by aggressive slag of the oxidative stage (see the disadvantages of the prototype method RU 2169202).

Copper sulfide material, SiO ₂ and CaO-containing fluxes, hard coal (in case of heat deficiency during smelting), oxygen-containing blast (a mixture of process oxygen and air) and liquid or gaseous hydrocarbon fuel are fed to the oxidation stage. The latter is also served if necessary, i.e. with heat deficiency during the melting process. The specific oxygen consumption for the oxidation stage of the process is maintained within 150-240 m ³ / t of dry sulfide material and is calculated as the ratio of oxygen consumption to the consumption of sulfide material (per dry weight). In this case, the oxygen consumption in relation to the sulfide material is calculated as the difference in the total oxygen consumption minus the oxygen consumption that was used to burn solid and / or gaseous (liquid) fuel. The oxygen consumption for fuel combustion is calculated based on the complete combustion of the latter. Below is the equation for calculating the specific oxygen consumption:

- расход кислорода на окислительную зону печи, нм³/ч;

- oxygen consumption for the oxidation zone of the furnace, nm ³ / h;

% O ₂ - volume fraction of oxygen in the process oxygen;

V _air - air flow to the oxidation zone of the furnace, nm ³ / h;

Q _CxHy — consumption of hydrocarbon fuel, nm ³ / h;

To _CxHy - coefficient characterizing the type of fuel for conditions of its complete combustion (for example, for natural gas it is 2.1, for propane-butane - 5.8, etc.);

Q _C - consumption of solid carbonaceous reducing agent, t / h;

K _{C is a} coefficient characterizing the type of solid reducing agent (the content of carbon and volatiles in it) for the conditions of its complete combustion;

W _C - mass fraction of moisture in a solid carbonaceous reducing agent;

Q _sulf - consumption of sulfide material, t / h;

W _Sulf - the mass fraction of moisture in the sulfide material, t / h;

The amount of oxygen supplied to the oxidation of the sulfide material is calculated based on the need for the complete oxidation of sulfur, impurity metals (for example, iron, nickel, cobalt) and the partial oxidation of copper to copper oxide (I). Depending on the composition of the sulfide material being processed, the specific oxygen consumption for the implementation of the oxidative stage of the process may vary within the limits indicated above. The decrease in the specific oxygen consumption below 150 nm ³ / t leads to the incomplete oxidation of sulfur, impurity metals, and, therefore, the resulting copper is not rough, but semi-rough and requires an additional conversion operation. An increase in the specific oxygen consumption above 240 nm ³ / t leads to unjustified reoxidation of copper and, consequently, an increased consumption of hydrocarbon fuel in the reduction stage. In addition, it is possible to stratify the slag with the formation of an intermediate layer based on copper oxide (I) very aggressive for refractory lining between blister copper and slag layers.

At the reduction stage, the process of reduction of the slag melt coming from the oxidizing stage by a gas mixture consisting of oxygen, air, and hydrocarbon fuel is implemented. In this case, the ratio of fuel to air-oxygen mixture is selected so that the oxygen consumption coefficient (α) is in the range from 0.50 to 0.90. The coefficient (α) is calculated as follows:

Where

V _O2 — oxygen consumption in the reduction zone of the furnace, nm ³ / h;

% O ₂ - volume fraction of oxygen in the process oxygen;

V _air - air flow to the recovery zone of the furnace, nm ³ / h;

Q _CxHy — consumption of hydrocarbon fuel, nm ³ / h;

To _CxHy - coefficient characterizing the type of fuel for the conditions of its complete combustion.

Additionally, a solid carbonaceous reducing agent in an amount of 0.1 to 5% by weight of the slag can be loaded into the reduction zone. This is done in case of heat shortage. In this case, the coefficient α should be increased, and a situation may arise when the reduction potential is not enough to provide the required degree of copper recovery from slag. Then the necessary reduction potential is achieved by additional input to the stage of depletion of hard coal. In addition, the introduction of coal allows you to save part of the hydrocarbon fuel.

The purpose of the recovery stage is to provide partial recovery of oxide copper from the slag melt, while not allowing the saturation of the slag with iron-nickel spinel (magnetite). For this, firstly, carbon-containing reducing agents are used that effectively destroy spinels. Secondly, these reducing agents themselves, unlike sulfide concentrate, do not contain appreciable amounts of elements (iron, nickel, cobalt, etc.) that can oxidize and thereby increase the likelihood of formation of spinel in the slag. Thirdly, during melting, CaO-containing flux is added to the charge, which increases the solubility limit in spinel slag.

The inadmissibility of a very deep recovery of copper slag is dictated by the ultimate goal of the process - to obtain blister copper that is conditional on the content of metal impurities. In the case of deep reduction of slag in the reduction zone of the furnace, a part of the impurity metals (for example, iron, nickel and cobalt) will be reduced together with copper. The creation of a relatively low reduction potential in the reduction zone of the furnace provides the conditions under which practically only copper is reduced.

An increase in the oxygen consumption coefficient (α) above 0.90 leads to the impossibility of reducing copper from the slag melt. Lowering the oxygen consumption coefficient below 0.50 leads to a deep reduction of copper, which in turn leads to the reduction of some of the metal impurities. Similarly, the selected flow rate range of the solid carbonaceous reducing agent is also explained. At a consumption of solid reducing agent below 0.1%, copper does not recover from slag, at a consumption of more than 5%, impurity metals are reduced in significant quantities. In this case, the quality of blister copper is deteriorating, and such copper must be subjected to conversion operations before refining. Maintaining the coefficient of oxygen consumption and consumption of solid reducing agent in the above ranges allows us to provide the required degree of copper recovery from slag melt with a slight reduction of impurity metals, which does not significantly affect the quality of blister copper.

Example

Tests of the new technology were carried out with copper nickel-containing concentrate from Feinstein separation at the Vanyukov pilot industrial dual-zone furnace (Fig. 1), built in 2005 at the Severonickel plant owned by MMC Norilsk Nickel.

The total hearth area of the Vanyukov dual-zone furnace is 11.4 m ² . The area of the oxidation zone is 4.7 m ² , the area of the reduction zone is 4.7 m ² , the siphon area is 2.0 m ² . The oxidation zone of the furnace is equipped with 6 tuyeres (4 main, 2 additional at the end of the furnace), the reduction zone is equipped with 4 tuyeres. The design of the tuyeres, their location, as well as the amount of air-oxygen blast and fuel supplied through them are identical to the industrial ones, which makes it possible to predict oxidation and reduction processes with a high degree of reliability according to the results of tests on a pilot industrial furnace. The working spaces of the furnace zones are separated by water-cooled partitions, the metal phase in the furnace is shared, the gas removal systems are separate. The gases of the oxidation zone containing SO ₂ are sent to sulfuric acid production, the gases of the reduction zone containing CO and H ₂ are burned with air through tuyeres specially arranged in the upper row of coffers and a system of organized suction, are cleaned of dust in the electrostatic precipitator and emitted into the atmosphere. The furnace siphon has an original design and provides for the possibility of producing blister copper both continuously and periodically. The level of the drain threshold for continuous removal of slag from the furnace is located at an altitude of 2200 mm from the hearth of the furnace, the level of copper in the unit was maintained at an altitude of 800-900 mm.

The composition of the processed copper concentrate, wt.%: Cu 67-67.5; Ni 3.8-3.9; C 0.23-0.25; Fe 4.4-4.6; S 21.2-21.5. The moisture content is 6.5%. Quartzite and limestone were used as fluxes. The content of silica flux SiO ₂ 73.0 wt.%, The content of limestone CaO 53.6 wt.%. Anthracite was used as solid fuel. The content in anthracite, wt.%: C 79.4; S 2.6. As a hydrocarbon fuel and a reducing agent used propane-butane mixture. Content in propane-butane mixture, wt.%: Propane and propylene 40-48; butane and butylenes 48-57; ethane and ethylenes 0.8-2.9. The oxygen-containing blast consisted of a mixture of process oxygen and compressor air.

The concentrate productivity was 4-5 t / h. During the tests, 1977 tons of copper concentrate were processed. The furnace worked in the technology development mode for 22 days.

Figure 1 - schematic representation of a dual-zone furnace Vanyukov: 1 - oxidizing zone; 2 - recovery zone; 3 - siphon; 4, 5 - pharmacies of the reduction and oxidizing zones; 6, 7 - flues of the reducing and oxidizing zones; 8 - burner; 9 - drain slag threshold; 10 - hole for the release of copper; 11 - emergency hole; 12, 13 - water-cooled partitions; 14 - tuyeres; 15 - tuyeres for afterburning of exhaust gases; 16 - copper caissons; 17 - embedded copper refrigerators.

Concentrate and a mixture of quartz and lime fluxes, anthracite, were continuously charged into the oxidation zone of the furnace. Quartz flux consumption was 8-15% of the concentrate, limestone consumption was 5-10% of the concentrate, anthracite consumption was 2.5-3.0% of the concentrate. An air-oxygen mixture was supplied through tuyeres in the side walls of the furnace with a total flow rate of 2100-2500 nm ³ / h. The oxygen blast enrichment ranged from 85 to 93%. A propane-butane mixture with a flow rate of 210-250 nm ³ / h was fed through a part of the tuyeres to compensate for the heat loss of a small-scale furnace.

In the process of pilot tests, both the oxidation stage of the process and the main two-stage technology variant were worked out: oxidative smelting and slag depletion. During the development of the oxidation stage in the reduction zone, the ratio of oxygen and propane-butane mixture was maintained, which ensured that the composition of the slag formed in the oxidation zone was unchanged (α = 0.91-0.95). Typical blister copper and slag compositions obtained at specific costs in the range from 198 to 206 nm ³ / t are presented in table 1.

Table 1
Compositions of blister copper and slag obtained as a result of pilot tests of the oxidative stage of the process U, nm ³ / t Content in blister copper, wt.% The content in the slag, wt.% Ni Cu With Fe S O Ni Cu With Fe SiO ₂ CaO 206 0.37 98.85 0.004 0.02 0.05 1.35 6.5 24.8 0.52 17.5 29.6 12.0 203 0.42 98.76 0.004 0.02 0.04 1.18 7.7 22.2 0.59 17.3 25.1 11.6 198 0.38 98.79 0.005 0.06 0.03 0.88 8.0 19.8 0.61 17.9 29.0 10.4 202 0.37 98.85 0.004 0.02 0.05 1.05 7.6 21.6 0.60 17.3 25.7 11.2 205 0.52 98.49 0.005 0.03 0.06 1.24 7.4 23.8 0.59 16.9 25.2 12.7 Wed value 0.41 98.75 0.005 0.03 0.05 1.14 7.4 22.4 0.58 17.4 26.9 11.6

When implementing a variant of the technology for producing blister copper in two stages in the reduction zone, the oxygen consumption coefficient was maintained within the limits providing partial recovery of copper from the slag melt. In the oxidation zone, the process was conducted without change, i.e. as in tests of the oxidative stage only. An oxygen-air mixture was supplied to the reduction zone with a total flow rate of 1100-1500 nm ³ / h. The oxygen blast enrichment ranged from 85 to 93%. The consumption of the propane-butane mixture was 240-260 nm ³ / hour. The oxygen consumption coefficient (a) ranged from 0.90 to 0.50.

The compositions obtained as a result of the implementation of the full version of the technology of slag and blister copper (combined in a furnace as a result of processes in the oxidation and reduction zones) are presented in table 2.

Figure 2 - the microstructure of the slag after depletion of copper concentrate according to the prototype method (a) and the proposed method (b).

1 - silicate glass; 2 - iron-nickel spinel (in the melt in the solid state); 3 - kings of metallic copper.

table 2
Compositions of blister copper and slag obtained as a result of testing a full (in two stages) technology version Coefficient t in rest.
zone Content in blister copper, wt.% The content of e in the slag, wt.% Ni Cu With Fe S O Ni Cu With Fe SiO ₂ CaO 0.79 0.47 98.5 0.005 0.03 0.05 0.53 7.7 8.0 0.72 21.1 31.3 17.9 0.78 0.42 99.0 0.005 0.02 0,03 0.88 8.5 8.0 0.71 20.8 32,5 14.9 0.78 0.55 98.8 0.005 0.02 0.05 0.72 8.4 9.6 0.72 21.5 33.7 15,5 Wed value 0.48 98.8 0.005 0.02 0.04 0.71 8.2 8.5 0.72 21.1 32,5 16.1

Note: the specific oxygen consumption in the oxidation zone was maintained in the range of 195–205 nm ³ / t.

Thus, the proposed oxidative smelting process of copper nickel-containing concentrate, combined with the process of continuous depletion of slag of the oxidizing stage in the reduction zone of the furnace, allows to reduce the concentration of copper in the slag by 2.5-3 times in comparison with the prototype method RU 2169202 (tested on the same sulphide material - copper concentrate from Feinstein separation) while maintaining the quality of blister copper in the nickel content in it at the same level. Such a significant decrease in copper concentration becomes possible as a result of the use of hydrocarbons and hard coal for slag recovery, i.e. reducing agents that effectively destroy spinels and do not themselves contain appreciable amounts of elements (iron, nickel, cobalt, etc.) that can oxidize and increase the likelihood of spinel formation. In addition, in the proposed method, CaO-containing flux is introduced into the charge, which increases the saturation limit of the silicate slag melt with iron-nickel spinel. All this, in contrast to the prototype method that uses depletion of sulfide concentrate, allows even with a very significant decrease in the concentration of copper in the slag to maintain the homogeneity of the slag melt and guarantee the stability of the process without foaming the melt.

As an example, figure 2 presents the microstructure of slag melts obtained by depletion of slag with the original copper concentrate in the prototype method and a gas mixture of oxygen and hydrocarbon fuel in the proposed method.

From a comparison of the microstructures it is clearly seen that in the case of depletion of copper concentrate in the slag there is a significant amount of iron-nickel spinel, not observed in the slag, depleted in the proposed method.

Thus, as a result of the pilot industrial tests, it was found that the direct extraction of copper from the concentrate into blister copper is 95.5%, the recovery of sulfur in SO ₂ concentrated gases is 99.6%. The temperature of the slag was maintained at 1300-1400 ° C, the temperature of blister copper -1150-1200 ° C.

Table 3 presents other examples of the implementation of the smelting of copper concentrate in a two-zone Vanyukov furnace beyond the optimal ranges of the above parameters.

Table 3
The results of the processing of copper concentrate from the separation of Feinstein in a two-zone Vanyukov furnace with deviation of parameters from optimal values. Deviation of the parameter from the optimal The composition of blister copper, wt.% The composition of the slag, wt.% Comment Ni Cu With Fe S O Ni Cu With Fe SiO ₂ CaO U <150nm ³ / t
(50 nm ³ / t) 7.90 87.7 0.04 0,07 1.10 0.10 2,8 4.7 0.86 26.8 30,2 12.3 Obtained semi-copper with a significant content of sulfur and nickel. An operation is required to convert such copper before anode refining Between the layers of slag and semi-copper in the furnace, a matte layer is diagnosed containing, wt.%: Ni 3,4; Cu 76.6; With 0.12; Fe 0.35; S 19.5. U> 240 nm ³ / t (245 nm ³ / t) 0.12 99.3 <0.01 <0.01 <0.01 2.10 6.2 38.3 0.19 14.0 19.7 6.0 Between the layers of slag and blister copper, a layer of the oxide phase based on Cu ₂ O, which is aggressive with respect to the refractory lining, was diagnosed. The content in the oxide phase, wt.%: Ni 3,5; Cu 70.5; C 0.19; Fe 8.1; S 0.096. The extraction of copper into blister copper at the oxidative stage of the process decreased compared with oxidative smelting at optimal values of U from 89 to 75%. α> 0.9 (0.95) U = 206 nm ³ / t 0.38 98.8 <0.01 0.08 0,03 1.25 5,6 25,2 0.40 17.5 25.5 11.5 Copper slag depletion did not occur. The slag of the oxidizing zone is almost identical in composition to the slag of the reducing zone and contained, wt.%: Ni 5.5; Cu 25.7; Co 0.40; Fe 17.6; SiO ₂ 24.4; CaO 11.5. α <0.5 (0.45) U = 199 nm ³ / t 1,56 97.1 0.05 0.40 0.02 0.57 6.1 2,8 0.65 21.7 35,4 15,2 Received copper with a high nickel content. A conversion operation is required before anode refining to form coiled nickel slags.

Note: in the two upper examples, the compositions of slag and copper represent the operation of only the oxidation zone of the furnace, in the two lower examples, the compositions of slag and copper describe the operation of a dual-zone furnace, i.e. the composition of the slag after depletion, and the composition of copper refers to the combined metal phase of the furnace.

Claims (5)

1. A method for the continuous processing of copper and nickel-containing copper sulfide materials, including oxidative smelting in a furnace of sulfide material together with fluxes to produce blister copper, slag, SO ₂ concentrated gases with specific oxygen consumption in the range of 150-240 nm ³ per 1 ton of dry dry sulfide material and depletion of this slag, characterized in that the processing is carried out in a furnace with two zones - oxidative smelting is carried out in the oxidizing zone, and slag depletion is carried out continuously in the reduction zone of the furnace and when using a mixture of oxygen-containing gas and hydrocarbon fuel for this with an oxygen flow coefficient (α) in the range from 0.5 to 0.9. 2. The method according to claim 1, characterized in that the oxidizing melting is additionally fed with coal. 3. The method according to claim 1, characterized in that the oxidative smelting is additionally fed hydrocarbon fuel. 4. The method according to claim 1, characterized in that the CaO-containing flux is fed to the oxidative smelting together with the SiO ₂ -containing flux based on the production of slags with a ratio of SiO ₂ / CaO from 3/1 to 1/1, and the total flux consumption for oxidative smelting is determined from the conditions of maintaining in the slag the sum of the concentrations of iron, nickel and cobalt is not more than 30 wt.%. 5. The method according to claim 1, characterized in that for the depletion of slag together with hydrocarbon fuel serves hard coal.

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