A method for extracting valuable metals from cyanide tailings

Abstract

The present invention relates to a method for extracting valuable metals from cyanide tailings, wherein the cyanide tailings to be treated, the first reducing agent, the chlorinating agent and the first conditioning agent are added to the first smelting furnace, an oxygen-rich gas is introduced, and oxidation smelting is performed to obtain CaO-SiO2-FeO molten slag and flue gas rich in metal chlorides; the CaO-SiO2-FeO molten slag in the first smelting furnace is transferred to the second smelting furnace, and the second reducing agent and the second conditioning agent are added to the second smelting furnace at the same time, and an oxygen-rich gas is introduced, and reduction smelting is performed at 1400-1600°C to obtain CaO-MgO-SiO2 molten slag and molten iron; dust is collected for the flue gas rich in metal chlorides to obtain smoke dust rich in metal chlorides. The present invention adopts the melting chlorination-reduction iron extraction process to comprehensively utilize the valuable elements such as gold, silver, copper, and iron in the cyanide tailings, which greatly reduces the energy consumption of extracting valuable resources from the cyanide tailings and realizes the maximum utilization of resources.

Description

Technical Field

[0001] The invention relates to a method for extracting valuable metals from cyanidation tailings, belonging to the technical field of comprehensive recovery of nonferrous metals. Background Art

[0002] The cyanidation tailings of gold concentrate are rich in rare and precious metals, and have few impurities, more than 95% of which are sulfur, iron, silicon and oxygen. The conventional treatment processes for cyanidation tailings include roasting-leaching, reduction roasting-magnetic separation, etc., but these processes have the disadvantages of long process, low recovery efficiency and high investment cost.

[0003] Chinese invention patent application CN201410422951.8 discloses a method for recovering valuable metals in tailings, wherein tailings, bentonite and chlorinating agent are mixed and then grinded to form green balls, the green balls are dried and then chlorinated and roasted in a rotary kiln, the roasted balls are cooled to obtain qualified pellets, the roasted flue gas is discharged through multi-stage washing, and the smoke containing valuable metals enters the solution to form a recovery solution; the filter cake obtained by filtering the recovery solution is gold, silver and lead slag, and the gold, silver and lead slag is treated to recover gold, silver and lead; the filtered filtrate is adsorbed on activated carbon and then neutralized with lime to obtain a neutralized filter residue as copper, lead and zinc slag, and the copper, lead and zinc slag is treated to recover copper, zinc and lead; the neutralized filtrate is evaporated and concentrated to recover the chlorinating agent. However, the method proposed in this patent application has problems such as high production cost, serious corrosion of high-temperature equipment, and ineffective recovery of iron.

[0004] Chinese invention patent application CN201510148846.4 discloses a method for extracting gold from cyanide tailings, which includes the following steps: mixing cyanide tailings, bentonite, and calcium chloride in proportion, pelletizing, drying, and roasting, wherein the metal gold and chlorine form a compound and volatilize in the form of smoke, the smoke is washed by spraying, concentrated by a thickener, and filtered by pulp pressure, and the obtained filter cake is leached by cyanide to extract gold. However, the method proposed in the patent application also has problems such as high production cost, severe corrosion of high-temperature equipment, and ineffective recovery of iron. Summary of the invention

[0005] In view of the shortcomings of the prior art, the object of the present invention is to provide a method for extracting valuable metals from cyanide tailings, so as to recover or enrich valuable elements such as gold, silver, copper, iron, etc. in the cyanide tailings.

[0006] In order to solve the above technical problems, the technical solution of the present invention is as follows:

[0007] A method for extracting valuable metals from cyanide tailings comprises the following steps:

[0008] S1, adding the cyanide tailings to be treated, the first reducing agent, the chlorinating agent and the first conditioning agent into the first smelting furnace, introducing oxygen-rich gas, and performing oxidation smelting to obtain CaO-SiO2-FeO molten slag (liquid) and flue gas rich in metal chlorides;

[0009] Wherein, the mass ratio of the cyanide tailings to be treated, the first reducing agent, and the chlorinating agent is 100:1-10:1-10, further 100:1-5:1-5, further 100:2-4:2-4, and further 100:2.5-3.5:2.5-3.5;

[0010] The first conditioning agent contains one or more of an iron compound, a calcium compound, and a silicon compound, so that in the CaO-SiO2-FeO molten slag, the mass ratio of FeO to SiO2 is 1.0-2.2:1, further 1.2-2:1, and further 1.6-1.8:1; the mass ratio of CaO to SiO2 is 0.3-1.2:1, further 0.5-1.0:1, and further 0.6-0.8:1; specifically, the composition of the first conditioning agent and the amount of each component added are based on the composition of the CaO-SiO2-FeO molten slag. The mass ratio of FeO and SiO2 is 1.0-2.2:1, and the mass ratio of CaO and SiO2 is 0.3-1.2:1, which can be specifically calculated based on the composition and addition amount of the cyanide tailings;

[0011] The chlorinating agent includes a chloride salt;

[0012] The temperature of oxidation smelting is >1200℃;

[0013] The concentration of oxygen in the flue gas is 0.1 to 3 vol%;

[0014] S2, transferring the CaO-SiO2-FeO molten slag in the first smelting furnace to the second smelting furnace, adding the second reducing agent and the second tempering agent to the second smelting furnace, introducing the oxygen-rich gas, and performing reduction smelting at 1400-1600° C. to obtain CaO-MgO-SiO2 molten slag (liquid) and molten iron; preferably, performing reduction smelting at 1500-1550° C.; optionally, transferring the CaO-SiO2-FeO molten slag in the first smelting furnace to the second smelting furnace through a chute;

[0015] Collecting dust from the flue gas rich in metal chloride to obtain flue dust rich in metal chloride;

[0016] Wherein, the mass ratio of the CaO-SiO2-FeO molten slag to the second reducing agent is 100:10-30, further 100:15-25, further 100:18-23;

[0017] The second conditioning agent contains one or more of a calcium compound, a magnesium compound, and a silicon compound, so that the ratio of the total mass of CaO and MgO to the mass of SiO2 in the CaO-MgO-SiO2 molten slag is 0.6-1.2:1, and further 0.8-1.0:1. Specifically, the components of the second conditioning agent and the amount of each component added are such that the ratio of the total mass of CaO and MgO to the mass of SiO2 in the CaO-MgO-SiO2 molten slag is 0.6-1.2:1.

[0018] In this way, through the oxidation smelting process, on the one hand, the cyanide in the cyanide tailings can be harmlessly decomposed to achieve the harmlessness of the cyanide tailings; on the other hand, due to the addition of the chlorinating agent, the gold, silver, copper and other metals in the cyanide tailings are chlorinated to form chlorides that enter the flue gas and are thus separated from the molten slag; at the same time, due to the addition of the reducing agent and the control of the first smelting furnace to a weakly oxidizing atmosphere, the iron oxide in the cyanide tailings is reduced to form ferrous oxide, and a low-melting-point CaO-SiO2-FeO slag type is formed, which can effectively promote the chlorination reaction. Subsequently, the CaO-SiO2-FeO molten slag is transferred to the second smelting furnace, the second reducing agent and the second tempering agent are added, the oxygen-rich gas is introduced, and the reduction smelting is carried out at 1400-1600°C. On the one hand, the slag is transferred to the second smelting furnace in a molten state, which can effectively save energy consumption; on the other hand, the reducing agent is added, and the strong reducing atmosphere is controlled, so that the FeO in the CaO-SiO2-FeO molten slag can be further reduced to form molten iron; at the same time, the addition of the second tempering agent makes the CaO-SiO2 slag type formed after the CaO-SiO2-FeO is reduced to be a CaO-MgO-SiO2 slag type, thereby obtaining a lower melting temperature and viscosity, which is conducive to the ferrous reduction reaction and the separation of slag and iron. In summary, the present invention realizes the extraction of iron in cyanidation tailings through a shorter processing flow, and realizes the separation and enrichment of rare and precious metals such as gold, silver, and copper, and the cyanidation tailings are efficiently recycled, which can greatly reduce the recycling cost.

[0019] The cyanide tailings itself has a high calorific value and can reach the target temperature required for smelting without the addition of additional fuel.

[0020] Optionally, the first conditioning agent includes one or more of limestone, hematite, pyrite and quartz.

[0021] Optionally, the second conditioning agent includes one or more of limestone, dolomite, quartz stone and gypsum.

[0022] Optionally, the chlorinating agent includes one or more of calcium chloride, sodium chloride, potassium chloride, and magnesium chloride, preferably calcium chloride.

[0023] Optionally, the first reducing agent includes one or more of carbon concentrate, waste activated carbon, coking coal, anthracite, and coke.

[0024] Optionally, the second reducing agent includes one or more of carbon concentrate, waste activated carbon, coking coal, anthracite, and coke.

[0025] The main reactions that may occur during oxidation smelting are:

[0026] 3FeS2+8O2(g)=Fe3O4+6SO2(g)

[0027] 2C+O2(g)=2CO(g)

[0028] Fe3O4+CO(g)=FeO+CO2(g)

[0029] 2CaCl2+O2(g)+2SiO2+4Au=4AuCl(g)+2CaSiO3

[0030] CaCl2+C+1.5O2(g)+SiO2+2Au=2AuCl(g)+CaSiO3+CO2(g)

[0031] 2CaCl2+O2(g)+2SiO2+4Ag=4AgCl(g)+2CaSiO3

[0032] CaCl2+C+1.5O2(g)+SiO2+2Ag=2AgCl(g)+CaSiO3+CO2(g)

[0033] CaCl2+CuS+1.5O2(g)+SiO2=CuCl2(g)+CaSiO3+SO2(g)

[0034] CaCl2+CuS+2.5O2(g)+C+SiO2=CuCl2(g)+CaSiO3+SO2(g)+CO2(g)

[0035] Main slag-forming reactions:

[0036] CaO+SiO2=CaO·SiO2

[0037] 2FeO+SiO2=2FeO*SiO2.

[0038] The main reactions that may occur during the reduction smelting process are:

[0039] 2C+O2(g)=2CO(g)

[0040] FeO+CO(g)=Fe+CO2(g)

[0041] 2FeSiO3+2CaO+C=2Fe+2CaSiO3+CO2(g)

[0042] 2FeSiO3+2MgO+C=2Fe+2MgSiO3+CO2(g)

[0043] MgO+SiO2=MgSiO3.

[0044] Optionally, in S1, the oxidation smelting temperature is 1250-1450°C.

[0045] Optionally, the oxygen concentration of the oxygen-rich gas is greater than 40 vol%, preferably 60-80 vol%, so that the concentration of sulfur dioxide in the flue gas can be increased and the acid production cost can be reduced.

[0046] Optionally, the flue gas rich in metal chlorides is subjected to waste heat utilization and dust collection in sequence to obtain residual flue gas and flue dust rich in metal chlorides;

[0047] The remaining flue gas is used for acid production and then discharged in compliance with the emission standards;

[0048] After acid leaching the metal chloride-rich smoke, solid-liquid separation is performed to obtain a leachate; water-soluble sulfide is added to the leachate, and after reaction, solid-liquid separation is performed to obtain metal sulfide.

[0049] Optionally, dilute sulfuric acid or dilute hydrochloric acid is used for pickling and leaching; optionally, the water-soluble sulfide is one or more of sodium sulfide and potassium sulfide.

[0050] Furthermore, in S2, the CO concentration in the outlet flue gas of the second smelting furnace is 15-45 vol%, and further 20-40 vol%, which can be specifically controlled according to the needs of ironmaking.

[0051] Optionally, after S2, the CaO-MgO-SiO2 molten slag is treated by method 1 or method 2;

[0052] Method 1: water quenching the CaO-MgO-SiO2 molten slag to obtain water-quenched slag;

[0053] Method 2:

[0054] (1) mixing the CaO-MgO-SiO2 molten slag with the third conditioning agent, heating and melting, and obtaining a mineral wool melt;

[0055] The third conditioning agent contains one or more of a calcium compound, an aluminum compound, a silicon compound, and a magnesium compound, so that the ratio of the total mass of SiO2 and Al2O3 to the total mass of CaO and MgO in the mineral wool melt is greater than 1.6:1, preferably 1.7-2.5:1;

[0056] Optionally, the CaO-MgO-SiO2 molten slag is transferred to an electric furnace through a chute, and at the same time, a third conditioning agent is added to the electric furnace, heated and melted to obtain a mineral wool melt;

[0057] (2) The mineral wool melt is filamentized to obtain mineral wool.

[0058] After the rare metals and iron are recovered, the CaO-MgO-SiO2 molten slag is basically silicon, calcium and magnesium oxides, which can be used as a good raw material for mineral wool. After further tempering, it can be used to prepare mineral wool, which can effectively increase the added value.

[0059] Optionally, the third conditioning agent includes one or more of basalt, quartz, dolomite, and bauxite, preferably quartz, bauxite, and dolomite.

[0060] Optionally, the mineral wool melt is spun using a centrifuge to obtain mineral wool.

[0061] Optionally, in the cyanide tailings to be treated, the Fe content is 20-50wt%, the Si content is 10-30wt%, the Cu content is 0-5wt%, the Pb content is 0-5wt%, the Au content is 1-5g / t, the Ag content is 10-300g / t, and the S content is 10-25wt%.

[0062] Optionally, in the cyanide tailings to be treated, the Fe content is 20.5-30wt%, the Si content is 12-28wt%, the Cu content is 0.5-4wt%, the Pb content is 0.5-4wt%, the Au content is 1.5-3.5g / t, the Ag content is 20-280g / t, and the S content is 15-20wt%.

[0063] Optionally, the flue gas generated by the second smelting furnace is subjected to waste heat utilization, surface cooling, bag dust collection, and flue gas purification in sequence to obtain clean coal gas.

[0064] Compared with the prior art, the present invention has the following beneficial effects:

[0065] 1. The complete comprehensive utilization of cyanide tailings can be achieved through two tempering and smelting. The molten chlorination-reduction iron extraction process is adopted to comprehensively utilize the valuable elements such as gold, silver, copper, iron, etc. in the cyanide tailings, thereby solving the problem that the existing technologies such as CN201510148846.4 fail to recover valuable metals such as iron; at the same time, by utilizing the large amount of heat released by the oxidation of sulfur in the original ore, the energy consumption of extracting valuable resources from the cyanide tailings is greatly reduced, thereby realizing the maximum utilization of resources.

[0066] 2. The molten chlorination process controls the oxygen potential to achieve desulfurization, chlorination volatilization, and selective reduction slag formation in one step. The addition of a small amount of the first reducing agent can not only achieve selective reduction of iron, but also promote the chlorination reaction of valuable metals.

[0067] 3. The reduction smelting process constructs a CaO-MgO-SiO2 slag type through tempering, which maximizes the use of the heat in the chlorination process. At the same time, since iron is reduced to ferrous iron in the previous process, the energy consumption and the amount of reducing agent required for this process are greatly reduced.

[0068] 4. By further tempering the CaO-MgO-SiO2 molten slag and utilizing the heat from the previous process, SiO2-Al2O3-CaO-MgO molten body can be efficiently formed by only keeping it warm, and used as a raw material for preparing mineral wool to obtain high value-added by-products.

[0069] 5. The present invention greatly reduces the amount of fuel such as coal in the reaction process by utilizing the heat of the raw materials in a cascade manner, recovers the rare and precious metals and iron in the materials, and reuses the slag, truly turning waste into treasure. BRIEF DESCRIPTION OF THE DRAWINGS

[0070] Figure 1 The invention discloses a flow chart for extracting valuable metals from cyanide tailings. DETAILED DESCRIPTION

[0071] The present invention will be described in detail below in conjunction with the embodiments. It should be noted that, in the absence of conflict, the embodiments and features in the embodiments of the present invention can be combined with each other. Unless otherwise specified, the relevant percentages refer to mass percentages.

[0072] Example 1

[0073] A method for extracting valuable metals from cyanide tailings comprises the following steps:

[0074] S1. Add the cyanide tailings to be treated, the first reducing agent, the chlorinating agent and the first conditioning agent to the first smelting furnace, introduce oxygen-enriched air, control the oxygen content of the outlet flue gas for oxidative smelting, and control the oxygen concentration of the outlet flue gas to be 1 vol%. The purpose is to ensure that there is a certain oxygen potential to completely burn the sulfur, but at the same time, the ferrous oxide will not be oxidized to ferric oxide due to excessive oxygen potential. After the smelting is completed, CaO-SiO2-FeO molten slag and flue gas rich in metal chlorides are obtained;

[0075] Wherein, the mass ratio of the cyanide tailings to be treated, the first reducing agent, and the chlorinating agent is 100:5:3;

[0076] The first conditioning agent is limestone and hematite, and the amount of the first conditioning agent added satisfies: in the CaO-SiO2-FeO molten slag, the mass ratio of FeO to SiO2 is 1.8:1, and the mass ratio of CaO to SiO2 is 0.5:1;

[0077] The chlorinating agent is calcium chloride; the temperature of oxidation smelting is 1350°C; the first reducing agent is anthracite, which is used for both reduction and as fuel to supplement heat for the smelting process;

[0078] The chemical composition of the cyanide tailings to be treated is shown in Table 1.

[0079] Table 1 Main chemical components of cyanide tailings (if no special instructions, the unit is wt%)

[0080]

[0081] S2, transferring the CaO-SiO2-FeO molten slag in the first smelting furnace to the second smelting furnace, adding the second reducing agent and the second tempering agent to the second smelting furnace, introducing oxygen-rich gas so that the CO concentration of the outlet flue gas is 35 vol%, the purpose of which is to completely reduce the ferrous iron to elemental iron, and performing reduction smelting at 1550° C. to obtain CaO-MgO-SiO2 molten slag and molten iron;

[0082] The molten iron is released and cast into ingots to obtain pig iron;

[0083] Collecting dust from the flue gas rich in metal chloride to obtain flue dust rich in metal chloride;

[0084] Wherein, the mass ratio of CaO-SiO2-FeO molten slag to the second reducing agent is 100:15;

[0085] The second conditioning agent is limestone and dolomite. The amount of the second conditioning agent added is such that the ratio of the total mass of CaO and MgO to the mass of SiO2 in the CaO-MgO-SiO2 molten slag is 0.8:1.

[0086] The second reducing agent is anthracite.

[0087] The oxygen concentration of the oxygen-rich gas is 70 vol%.

[0088] The flue gas rich in metal chlorides is subjected to waste heat utilization and dust collection in sequence to obtain residual flue gas and flue dust rich in metal chlorides;

[0089] The remaining flue gas is used for acid production and then discharged in compliance with the emission standards;

[0090] After acid leaching the metal chloride-rich smoke, solid-liquid separation is performed to obtain a leachate; water-soluble sulfide is added to the leachate, and after reaction, solid-liquid separation is performed to obtain metal sulfide.

[0091] After S2, the CaO-MgO-SiO2 molten slag is mixed with the third conditioning agent, heated and melted to obtain a mineral wool melt;

[0092] The third conditioning agent is basalt, dolomite, quartz stone, and bauxite, and the amount of the third conditioning agent added is such that the ratio of the total mass of SiO2 and Al2O3 to the total mass of CaO and MgO in the mineral wool melt is greater than 1.6:1;

[0093] The mineral wool melt is spun through a centrifuge to produce silk, thereby obtaining mineral wool.

[0094] After testing, the volatility of gold, silver, copper, lead and zinc in the cyanide tailings after oxidation smelting was 97wt%, 98wt%, 98.6wt%, 99wt% and 99.5wt% respectively, and the SO2 concentration in the flue gas rich in metal chloride was 48vol%. The purity of pig iron was 97.2wt%, the recovery rate of iron was 98.4wt%, and the sulfur content was 0.026wt%.

[0095] Comparative Example 1

[0096] Example 1 was repeated, with the only difference being that the oxygen concentration of the flue gas at the outlet of the first smelting furnace in S1 was controlled to be 0%.

[0097] The results showed that the sulfur in the slag was not completely removed, and the discharged slag was still a mixture of iron sulfide and part of the slag. The slag could not be used for ironmaking and needed further treatment. At the same time, the volatility of gold, silver, copper, lead and zinc were 58wt%, 62wt%, 71wt%, 53wt% and 78wt% respectively. Due to the decrease in oxygen potential, their volatility was significantly reduced.

[0098] Example 2

[0099] Example 1 was repeated, with the only difference being that the oxygen concentration of the flue gas at the outlet of the first smelting furnace in S1 was controlled to be 0.1 vol%.

[0100] The results showed that the sulfur in the slag was almost removed, the slag viscosity was acceptable, the fluidity was good, and the volatilization rates of gold, silver, copper, lead and zinc were 89wt%, 85wt%, 78wt%, 91wt% and 96wt% respectively. The mixed slag entered the next reduction process to obtain the product pig iron, the purity of the pig iron was 96.4wt%, the iron recovery rate was 97.4wt%, and the sulfur content was 0.055%.

[0101] Example 3

[0102] Example 1 was repeated, with the only difference being that the oxygen concentration of the flue gas at the outlet of the first smelting furnace in S1 was controlled to be 3 vol%.

[0103] The results showed that the sulfur in the slag was almost removed, the viscosity of the slag increased, and the fluidity became worse, but it could still be discharged into the reduction furnace. At this time, the volatilization rates of gold, silver, copper, lead, and zinc were 96wt%, 98wt%, 99.1wt%, 99.4wt%, and 99.9wt%, respectively, and the sulfur dioxide concentration in the flue gas was reduced to 38%. The mixed slag entered the next reduction process to obtain the product pig iron, with a purity of 96.7wt%, an iron recovery rate of 98.2wt%, and a sulfur content of 0.021%.

[0104] Comparative Example 2

[0105] Example 1 was repeated, except that the oxygen concentration of the flue gas at the outlet of the first smelting furnace in S1 was controlled to be 5 vol%, iron oxide was generated during the smelting process, the viscosity of the slag increased, and the slag was difficult to be discharged smoothly into the next process. At the same time, the volatility of gold, silver, copper, lead, and zinc were 97.6wt%, 98.9wt%, 99.2wt%, 99.5wt%, and 99.9wt%, respectively. Due to the increase in oxygen potential, the iron oxide content increased, resulting in an increase in slag viscosity, and the slag was difficult to be discharged smoothly. At the same time, the increase in oxygen potential is conducive to the improvement of chlorination index.

[0106] Comparative Example 3

[0107] Example 1 was repeated, with the only difference being that the CO concentration of the flue gas at the outlet of the second reduction furnace in S1 was controlled to be 8 vol%.

[0108] The results showed that a large amount of iron in the reduction furnace was not reduced to elemental iron, but still existed in the form of ferrous oxide. The direct iron recovery rate dropped to 48.4%. At the same time, since the iron grade in the reduction slag was 18.5%, which was much greater than the iron content required for the preparation of mineral wool, it was no longer suitable for entering the third electric furnace to prepare mineral wool.

[0109] Comparative Example 4

[0110] Example 1 was repeated, with the only difference being that the CO concentration of the flue gas at the outlet of the second reduction furnace in S1 was controlled to be 12 vol%.

[0111] The results showed that the iron in the reduction furnace was not carburized enough and the iron was not completely reduced into the molten iron. The direct recovery rate of iron dropped to 78.4%. The iron content in the reduction slag was 10.5%. After entering the third electric furnace for tempering, the quality of the rock wool produced was poor due to the high iron content.

[0112] Example 4

[0113] Example 1 was repeated, with the only difference being that the CO concentration of the flue gas at the outlet of the second reduction furnace in S1 was controlled to be 15 vol%.

[0114] The results showed that the iron in the reduction furnace was not carburized enough and the iron was not completely reduced into the molten iron. The direct recovery rate of iron was 88.4%. Part of the iron was lost into the slag and was brought into the third electric furnace. After tempering, it settled in the electric furnace and needed to be cleaned regularly.

[0115] Example 5

[0116] Example 1 was repeated, with the only difference being that the CO concentration of the flue gas at the outlet of the second reduction furnace in S1 was controlled to be 45 vol%.

[0117] The results showed that the direct iron recovery rate was 98.45%, with no significant increase, but coal consumption increased. At the same time, the iron grade in pig iron dropped to 92.6%. Therefore, the CO concentration was too high and there was no obvious further benefit. It was no longer suitable to further increase the CO concentration.

[0118] The contents explained in the above embodiments should be understood as these embodiments are only used to more clearly illustrate the present invention, and are not used to limit the scope of the present invention. After reading the present invention, various equivalent forms of modifications to the present invention by those skilled in the art all fall within the scope defined by the claims attached to this application.

Claims

1. A method for extracting valuable metals from cyanide tailings, characterized in that: The steps include: S1, adding the cyanide tailings to be treated, the first reducing agent, the chlorinating agent and the first conditioning agent into the first smelting furnace, introducing oxygen-rich gas, and performing oxidation smelting to obtain CaO-SiO2-FeO molten slag and flue gas rich in metal chlorides; Wherein, the mass ratio of the cyanide tailings to be treated, the first reducing agent, and the chlorinating agent is 100:1-10:1-10; The first conditioning agent contains one or more of an iron compound, a calcium compound, and a silicon compound, so that in the CaO-SiO2-FeO molten slag, the mass ratio of FeO to SiO2 is 1.0-2.2:1, and the mass ratio of CaO to SiO2 is 0.3-1.2:1; The chlorinating agent includes a chloride salt; The temperature of oxidation smelting is >1200℃; The concentration of oxygen in the flue gas is 0.1~3vol%; S2, transferring the CaO-SiO2-FeO molten slag in the first smelting furnace to the second smelting furnace, adding the second reducing agent and the second tempering agent to the second smelting furnace, introducing oxygen-rich gas, and performing reduction smelting at 1400-1600° C. to obtain CaO-MgO-SiO2 molten slag and molten iron; Collecting dust from the flue gas rich in metal chloride to obtain flue dust rich in metal chloride; Wherein, the mass ratio of CaO-SiO2-FeO molten slag to the second reducing agent is 100:10-30; The second tempering agent contains one or more of a calcium compound, a magnesium compound, and a silicon compound, so that in the CaO-MgO-SiO2 molten slag, the ratio of the total mass of CaO and MgO to the mass of SiO2 is 0.6~1.2:1; and the concentration of CO in the outlet flue gas of the second smelting furnace is 15~45vol%.

2. The method according to claim 1, characterized in that The first conditioning agent includes one or more of limestone, hematite, pyrite, and quartz; the second conditioning agent includes one or more of limestone, dolomite, quartz, and gypsum.

3. The method according to claim 1, characterized in that The chlorinating agent includes one or more of calcium chloride, sodium chloride, potassium chloride and magnesium chloride.

4. The method according to claim 1, characterized in that: The first reducing agent includes one or more of carbon essence, waste activated carbon, coking coal, anthracite, and coke; the second reducing agent includes one or more of carbon essence, waste activated carbon, coking coal, anthracite, and coke.

5. The method according to claim 1, characterized in that In S1, the oxidation smelting temperature is 1250-1450°C.

6. The method according to claim 1, characterized in that The oxygen concentration of the oxygen-rich gas is greater than 40 vol%.

7. The method according to claim 6, characterized in that The oxygen concentration of the oxygen-rich gas is 60-80 vol%.

8. The method according to claim 1, characterized in that: The flue gas rich in metal chlorides is subjected to waste heat utilization and dust collection in sequence to obtain residual flue gas and flue dust rich in metal chlorides; The remaining flue gas is used for acid production and then discharged in compliance with the emission standards; After acid leaching the metal chloride-rich smoke, solid-liquid separation is performed to obtain a leachate; water-soluble sulfide is added to the leachate, and after reaction, solid-liquid separation is performed to obtain metal sulfide.

9. The method according to claim 1, characterized in that: After S2, the CaO-MgO-SiO2 molten slag is treated by method 1 or method 2; Method 1: water quenching the CaO-MgO-SiO2 molten slag to obtain water-quenched slag; Method 2: (1) mixing the CaO-MgO-SiO2 molten slag with the third conditioning agent, heating and melting, and obtaining a mineral wool melt; The third conditioning agent contains one or more of a calcium compound, an aluminum compound, a silicon compound, and a magnesium compound, so that the ratio of the total mass of SiO2 and Al2O3 to the total mass of CaO and MgO in the mineral wool melt is greater than 1.6:1; (2) The mineral wool melt is filamented to obtain mineral wool.

10. The method according to claim 9, characterized in that In the mineral wool melt, the ratio of the total mass of SiO2 and Al2O3 to the total mass of CaO and MgO is 1.7-2.5:

1.

11. The method according to claim 9, characterized in that The third conditioning agent includes one or more of basalt, quartz, dolomite and bauxite.

12. The method according to claim 1, characterized in that In the cyanide tailings to be treated, the content of Fe is 20-50wt%, the content of Si is 10-30wt%, the content of Cu is 0-5wt%, the content of Pb is 0-5wt%, the content of Au is 1-5g / t, the content of Ag is 10-300g / t, and the content of S is 10-25wt%.

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