A method for coordinating treatment of cyanide tailings and copper-containing sludge

Abstract

The present invention relates to a method for coordinating cyanide tailings and copper-containing sludge, comprising the following steps: adding treated cyanide tailings, copper-containing sludge, a first reducing agent and a chlorinating agent to a first smelting furnace, introducing an oxygen-rich gas, performing chlorination smelting, and obtaining molten slag and flue gas rich in metal chlorides; transferring the molten slag in the first smelting furnace to a second smelting furnace, adding a second reducing agent and a flux to the second smelting furnace at the same time, introducing an oxygen-rich gas, performing reduction smelting, and obtaining CaO‑MgO‑SiO2 molten slag, molten iron and tail gas; collecting dust from the flue gas rich in metal chlorides to obtain flue dust rich in metal chlorides. The present invention realizes the complete comprehensive utilization of cyanide tailings and copper-containing sludge through two temperings, and adopts a combined process of melting chlorination-reduction iron extraction to comprehensively utilize valuable elements such as gold, silver, copper, iron, lead, and zinc in cyanide tailings and copper-containing sludge, respectively, solving the problem that the prior art fails to recycle valuable metals such as iron.

Description

Technical Field

[0001] The invention relates to a method for co-processing cyanidation tailings and copper-containing sludge, belonging to the technical field of comprehensive recovery of non-ferrous metals. Background Art

[0002] Cyanide tailings and copper-containing sludge are rich in valuable metals such as gold, silver, copper, and iron. However, the valuable metals have low grade and complex composition, which makes them difficult to handle. Conventional pyrometallurgical treatment processes for cyanide tailings usually use roasting-leaching, reduction roasting-magnetic separation and other processes for treatment, which have the disadvantages of long process flow, low recovery efficiency, and high investment cost. The valuable metals in copper-containing sludge have low grade and complex composition. Conventional copper-containing sludge treatment processes are usually combined with other multi-metal materials, such as circuit boards, for matte smelting, but this process has the disadvantages of difficult separation of valuable metals, poor adaptability of raw materials, and inability to recover iron resources.

[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.

[0005] Chinese invention patent application CN202110470389.6 discloses a method for treating copper-containing sludge, which comprises the following steps: drying the copper-containing sludge to obtain dried copper-containing sludge, wherein the moisture content of the dried copper-containing sludge is greater than or equal to a first preset value and less than or equal to a second preset value; and bottom blowing smelting of a preset proportion of dried copper-containing sludge, a slagging agent and a reducing agent. This method requires additional flux and can only recover copper in the sludge. Summary of the invention

[0006] In view of the deficiencies in the prior art, the object of the present invention is to provide a method for the coordinated treatment of cyanide tailings and copper-containing sludge, so as to recover or enrich valuable elements such as gold, silver, copper, iron, etc. in cyanide tailings and copper-containing sludge.

[0007] In order to solve the above technical problems, the technical solution of the present invention is as follows:

[0008] A method for co-processing cyanide tailings and copper-containing sludge comprises the following steps:

[0009] S1, adding the treated cyanide tailings, copper-containing sludge, the first reducing agent and the chlorinating agent into the first smelting furnace, introducing oxygen-rich gas, and performing chlorination smelting to obtain CaO-SiO2-FeO molten slag (liquid) and flue gas rich in metal chlorides;

[0010] The mass ratio of the total amount of the cyanide tailings and copper-containing sludge to be treated to the first reducing agent and the chlorinating agent is 100:5-20:5-20, further 100:5-15:5-15, further 100:8-15:5-12, and further 100:10-12:6-10;

[0011] The mass ratio of the cyanide tailings to be treated and the copper-containing sludge is 1-5:1, so that in the CaO-SiO2-FeO molten slag, the mass ratio of FeO to SiO2 is 1.0-2.4:1, further 1.2-2.2:1, further 1.6-2.0:1, and the mass ratio of CaO to SiO2 is 0.1-0.7:1, further 0.2-0.5:1, further 0.2-0.4:1. More specifically, the mass ratio of the cyanide tailings and the copper-containing sludge is such that in the CaO-SiO2-FeO molten slag, the mass ratio of FeO to SiO2 is 1.0-2.4:1, and the mass ratio of CaO to SiO2 is 0.1-0.7:1, which can be calculated based on the composition of the cyanide tailings and the copper-containing sludge;

[0012] The chlorinating agent includes a chloride salt;

[0013] The temperature of chlorination smelting is 1100-1300℃;

[0014] The concentration of oxygen in the flue gas is 0.1 to 3 vol%;

[0015] S2, transferring the CaO-SiO2-FeO molten slag in the first smelting furnace to the second smelting furnace, adding a second reducing agent and a flux to the second smelting furnace, introducing an oxygen-rich gas, and performing reduction smelting at 1400-1600° C. to obtain CaO-MgO-SiO2 molten slag, molten iron and tail gas;

[0016] Collecting dust from the flue gas rich in metal chloride to obtain flue dust rich in metal chloride;

[0017] 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;

[0018] The flux contains one or more of calcium compounds, magnesium compounds, and silicon compounds, 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 0.8-1:1. Specifically, the composition of the flux and the amount of each component added are based on 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;

[0019] Furthermore, the concentration of CO in the tail gas is 15-45 vol%, and further 20-40 vol%, which can be specifically controlled according to the needs of ironmaking.

[0020] In this way, through the coordinated smelting of cyanide tailings and copper-containing sludge, on the one hand, the harmless disposal of cyanide tailings and copper-containing sludge can be achieved; on the other hand, the material properties of the raw materials are fully utilized, and the coordinated chlorination smelting is carried out to greatly reduce the disposal cost; due to the addition of the chlorinating agent, the gold, silver, copper and other metals in the cyanide tailings and copper-containing sludge are chlorinated to form chlorides that enter the flue gas and are thus separated from the molten slag; at the same time, through the coordinated slagging of cyanide tailings, copper-containing sludge and chloride salts, the characteristics of each component raw material are fully utilized, and the process does not require additional slagging agents; due to the addition of a 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 the Ca in the copper-containing sludge participates in slagging to form a low-melting-point CaO-SiO2-FeO slag type, 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 flux are added, and the oxygen-rich gas is introduced for reduction smelting. On the one hand, the slag is transferred to the second smelting furnace in a molten state, which can facilitate the transfer of materials between the first smelting furnace and the second smelting furnace and 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 flux converts the CaO-SiO2 slag type formed after the CaO-SiO2-FeO is reduced into the 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 and copper-containing sludge through a relatively short processing flow, and realizes the separation and enrichment of rare and precious metals such as gold, silver, and copper, and efficiently recycles cyanidation tailings and copper-containing sludge, which can greatly reduce the recycling cost.

[0021] The present invention adopts a molten chlorination-reduction iron extraction process to coordinately treat cyanide tailings and copper-containing sludge, fully utilizes the material properties of cyanide tailings and copper-containing sludge, and comprehensively recovers valuable metals such as gold, silver, copper, and iron in the materials. The present invention has the advantages of short process, high comprehensive recovery efficiency, harmless tailings, etc.

[0022] Furthermore, the flux includes one or more of limestone, dolomite, quartz stone, and gypsum. Optionally, the flux is preferably dolomite and limestone.

[0023] Furthermore, the chlorinating agent includes one or more of calcium chloride, sodium chloride, potassium chloride, and magnesium chloride, preferably calcium chloride and sodium chloride.

[0024] The chlorinating agent not only participates in the chlorination reaction as a chlorine source, but also participates in the slag-making reaction as a major calcium source, sodium source or magnesium source, thereby further tempering the slag composition.

[0025] The main reactions occurring in the first smelting furnace are as follows:

[0026] 2CaCl2+O2(g)+2SiO2+4Au=4AuCl(g)+2CaSiO3

[0027] 2MgCl2+O2(g)+2SiO2+4Au=4AuCl(g)+2MgSiO3

[0028] 4NaCl+O2(g)+2SiO2+4Au=4AuCl(g)+2Na2SiO3

[0029] 2CaCl2+O2(g)+2SiO2+4Ag=4AgCl(g)+2CaSiO3

[0030] 2MgCl2+O2(g)+2SiO2+4Ag=4AgCl(g)+2MgSiO3

[0031] 4NaCl+O2(g)+2SiO2+4Ag=4AgCl(g)+2Na2SiO3

[0032] CaCl2+CuO+SiO2=CuCl2(g)+CaSiO3

[0033] MgCl2+CuO+SiO2=CuCl2(g)+MgSiO3

[0034] 2NaCl+CuO+SiO2=CuCl2(g)+Na2SiO3

[0035] Furthermore, the first reducing agent includes one or more of carbon essence, waste activated carbon, coking coal, anthracite, coke, natural gas, and organic materials, preferably anthracite and natural gas; the second reducing agent includes one or more of carbon essence, waste activated carbon, coking coal, anthracite, coke, natural gas, and organic materials, preferably anthracite and natural gas.

[0036] Optionally, before S1, the cyanide tailings and the copper-containing sludge are mixed and then dried to obtain a mixed material with a water content of less than 10 wt%.

[0037] Furthermore, the drying temperature is 300-500° C. The drying process can be carried out using clean coal gas obtained after the tail gas generated by the second smelting furnace is purified, making full use of the calorific value in the tail gas, while making the mixed material as dry as possible, thereby increasing the subsequent smelting efficiency.

[0038] Furthermore, in S1, the chlorination smelting temperature is 1150-1250°C.

[0039] Furthermore, the oxygen concentration of the oxygen-enriched gas is greater than 40 vol%, preferably 60-80 vol%, so that the amount of exhaust gas discharged from the furnace and the cost of exhaust gas treatment can be reduced.

[0040] Furthermore, 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;

[0041] The remaining flue gas is used for acid production and then discharged in compliance with the emission standards;

[0042] 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.

[0043] 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.

[0044] Optionally, the flue gas is first subjected to denitration treatment and then the waste heat is utilized. Optionally, dust is collected by a bag filter.

[0045] Furthermore, after S2, the CaO-MgO-SiO2 molten slag is water quenched to obtain water-quenched slag.

[0046] Furthermore, in the cyanide tailings to be treated, the Fe content is 30-60wt%, 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, and the Ag content is 10-300g / t.

[0047] Furthermore, in the copper-containing sludge, the content of Cu is 0.1-5wt%, the content of Au is 1-10g / t, the content of Ag is 10-200g / t, the content of Zn is 0-5wt%, and the content of CaO is 5-40wt%.

[0048] Optionally, the CaO-SiO2-FeO molten slag in the first smelting furnace is transferred to the second smelting furnace through a chute.

[0049] Optionally, the tail gas is subjected to waste heat utilization, cooling, dust collection and purification treatments in sequence to obtain clean coal gas.

[0050] Optionally, before S1, the cyanide tailings and the copper-containing sludge are input into a rotary kiln, and the materials are dried by the rotary kiln to obtain a mixed material; during this period, the clean coal gas is used as a heat source for the rotary kiln.

[0051] Furthermore, the tail 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. Optionally, the clean coal gas is used to dry the mixed material.

[0052] Optionally, before S1, the cyanide tailings and copper-containing sludge are input into a rotary kiln, and the materials are dried by the rotary kiln; during this period, the clean coal gas is used as a heat source for the rotary kiln. In this way, the concentration of CO in the tail gas is controlled to 15-45 vol%, which can not only meet the needs of ironmaking and obtain high-purity pig iron, but also obtain clean coal gas with a higher CO concentration to meet the needs of drying the mixed materials.

[0053] Optionally, the temperature of the reduction smelting is 1450-1500°C.

[0054] Compared with the prior art, the present invention has the following beneficial effects:

[0055] 1. Through two conditioning processes, the cyanide tailings and copper-containing sludge are fully and comprehensively utilized. The combined process of melt chlorination-reduction iron extraction is used to comprehensively utilize the valuable elements such as gold, silver, copper, iron, lead, and zinc in the cyanide tailings and copper-containing sludge, respectively, solving the problem that the existing technologies such as CN201510148846.4 fail to recycle valuable metals such as iron. This process greatly reduces the energy consumption of extracting valuable resources from cyanide tailings and copper-containing sludge, and helps to maximize the utilization of resources.

[0056] 2. The smelting process fully considers the characteristics of the two materials, makes full use of the similarities (multi-metal, high value, low grade) and differences (cyanide tailings high iron and low calcium, copper-containing sludge high calcium and low silicon) of their material characteristics, organically combines the two materials, and synergistically disposes of the two materials through molten chlorination.

[0057] 3. Make full use of the chlorination and slag-making functions of the chlorinating agent. While chlorinating and volatilizing low-grade valuable metals, metals such as Ca, Na, and Mg enter the slag for slag-making, which can further reduce the melting point of the slag system, thereby reducing energy consumption and processing costs.

[0058] 4. The reduction smelting process uses secondary tempering and flux to construct a CaO-MgO-SiO2 slag type. This process makes the most 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.

[0059] 5. This process greatly reduces the amount of coal used in the reaction process by utilizing the heat of the raw materials in a cascade manner. It not only recovers the rare and precious metals in the materials, but also reuses the slag, truly turning waste into treasure. BRIEF DESCRIPTION OF THE DRAWINGS

[0060] Figure 1 The present invention discloses a flow chart of the coordinated treatment of cyanide tailings and copper-containing sludge. DETAILED DESCRIPTION

[0061] 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.

[0062] Example 1

[0063] See also Figure 1 A method for co-processing cyanide tailings and copper-containing sludge comprises the following steps:

[0064] S1. The copper-containing sludge to be treated (the main components are shown in Table 2) and the cyanide tailings (the main components are shown in Table 1) are input into the rotary kiln drying system at a mass ratio of 1:3, and a mixed material is obtained at 300-500° C. The drying heat source comes from the tail gas generated in the reduction smelting process;

[0065] Then, the mixed material, the first reducing agent and the chlorinating agent are added to the first smelting furnace, and oxygen-rich gas is introduced to perform chlorination smelting to obtain CaO-SiO2-FeO molten slag and flue gas rich in metal chlorides;

[0066] Wherein, the mass ratio of the mixed material to the first reducing agent and the chlorinating agent is 100:12:8;

[0067] The mass ratio of the cyanide tailings to be treated to the copper-containing sludge is 3:1, so that 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 0.3:1;

[0068] The chlorinating agent is sodium chloride;

[0069] The temperature of chlorination smelting is 1150℃;

[0070] The concentration of oxygen in the flue gas is 1 vol%;

[0071] Table 1 Main chemical components of cyanide tailings (if no special instructions, the unit is wt%)

[0072]

[0073] Table 2 Main chemical components of copper-containing sludge (unless otherwise specified, in wt%)

[0074]

[0075] S2, transferring the CaO-SiO2-FeO molten slag in the first smelting furnace to the second smelting furnace, adding a second reducing agent and a flux to the second smelting furnace, introducing an oxygen-rich gas, and performing reduction smelting at 1450° C. to obtain CaO-MgO-SiO2 molten slag, molten iron and tail gas;

[0076] The molten iron is released, cast into ingots, and pig iron is obtained;

[0077] The mass ratio of the CaO-SiO2-FeO molten slag to the second reducing agent (anthracite) is 100:15;

[0078] The flux is limestone and dolomite, so that the ratio of the total mass of CaO and MgO to the mass of SiO2 in the Na2O-CaO-MgO-SiO2 molten slag is 0.6:1;

[0079] The concentration of CO in the tail gas is 35 vol%.

[0080] The first reducing agent is anthracite, which is used for both reduction and supplementation of heat required for smelting; the second reducing agent includes one or more of carbon concentrate, waste activated carbon, coking coal, anthracite, coke, natural gas, and organic materials.

[0081] The oxygen concentration of the oxygen-rich gas is 70 vol%.

[0082] 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;

[0083] The remaining flue gas is used for acid production and then discharged in compliance with the emission standards;

[0084] 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.

[0085] After S2, the CaO-MgO-SiO2 molten slag is water quenched to obtain water-quenched slag.

[0086] After chlorination smelting, the volatilization rates of gold, silver, copper, lead and zinc in copper-containing sludge and cyanide tailings were 98.3wt%, 97.2wt%, 99.2wt%, 99.5wt% and 99.7wt% respectively. The purity of pig iron was 97.2wt% and the recovery rate of iron was 98.4wt%.

[0087] Comparative Example 1

[0088] 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 was controlled to be 0%.

[0089] The results showed that the chlorination volatilization rate of each metal decreased, and the volatilization rates of gold, silver, copper, lead and zinc were 64.2wt%, 56.8wt%, 72.4wt%, 76.5wt% and 87.8wt% respectively. Due to the decrease in oxygen potential, their volatilization rates were significantly reduced.

[0090] Example 2

[0091] Example 1 was repeated, except that the oxygen concentration of the flue gas at the outlet of the first smelting furnace was controlled to be 0.1 vol%.

[0092] The results showed that the chlorination volatilization rate was slightly reduced, and the volatilization rates of gold, silver, copper, lead and zinc were 88.6wt%, 89.7wt%, 83.3wt%, 93.4wt% and 97.8wt% respectively. The mixed slag entered the next reduction process to obtain the product pig iron, with a purity of 94.4wt%, of which the gold grade was 1.7g / t, the silver grade was 16.5g / t, the copper grade was 0.5%, and the iron recovery rate was 97.4wt%. The gold, silver, copper, etc. that were not completely volatilized in the molten chlorination stage entered the pig iron, resulting in a decrease in the pig iron grade.

[0093] Example 3

[0094] Example 1 was repeated, except that the oxygen concentration of the flue gas at the outlet of the first smelting furnace was controlled to be 3 vol%.

[0095] The results showed that the chlorination volatilization rate was high, and the volatilization rates of gold, silver, copper, lead and zinc were 95.8wt%, 97.2wt%, 98.9wt%, 99.6wt% and 99.9wt% respectively. However, the viscosity of the slag increased, and the fluidity of the slag became poor. During the slag transfer process, it was necessary to manually poke the slag mouth frequently to avoid blockage. At this time, the ferrous iron in the slag was oxidized to ferric oxide. The mixed slag entered the next reduction process to obtain the product pig iron, with a purity of 97.2wt% and an iron recovery rate of 98.45wt%.

[0096] Comparative Example 2

[0097] Example 1 was repeated, except that the oxygen concentration of the flue gas at the outlet of the first smelting furnace 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 98.2 wt%, 99.1 wt%, 99.6 wt%, 99.7 wt%, and 99.9 wt%, 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.

[0098] Comparative Example 3

[0099] Example 1 was repeated, except that the mass ratio of copper-containing sludge to cyanide tailings was controlled to be 1:0.5.

[0100] The results showed that the slag viscosity was too high during the smelting process to achieve the purpose of subsequent reduction smelting. By adding 8% quartz stone into the mixed material, the slag was melted smoothly and discharged into the second smelting furnace.

[0101] Comparative Example 4

[0102] Example 1 was repeated, except that the mass ratio of copper-containing sludge to cyanide tailings was controlled to be 1:6.

[0103] The results showed that the viscosity of the slag was too high during the smelting process to achieve the purpose of subsequent reduction smelting. By adding 8% limestone to the mixed material, the slag was melted smoothly and discharged into the reduction furnace.

[0104] Example 4

[0105] Example 1 was repeated, except that the mass ratio of the copper-containing sludge to the cyanide tailings was controlled to be 1:1.

[0106] The results show that the mixed material can be smelted and deslagging smoothly, and the volatilization rates of gold, silver, copper, lead and zinc are 96.2wt%, 95.4wt%, 97.2wt%, 98.3wt% and 99.7wt% respectively. The CaO-SiO2-FeO molten slag enters the next reduction process to obtain the product pig iron, the purity of the pig iron is 96.2wt%, and the iron recovery rate is 97.32wt%.

[0107] Example 5

[0108] Example 1 was repeated, except that the mass ratio of the copper-containing sludge to the cyanide tailings was controlled to be 1:5.

[0109] The results show that the mixed material can be smelted and deslagging smoothly, and the volatilization rates of gold, silver, copper, lead and zinc are 94.3wt%, 92.7wt%, 95.7wt%, 97.8wt% and 98.9wt% respectively. The CaO-SiO2-FeO molten slag mixed slag enters the next reduction process to obtain the product pig iron, the purity of the pig iron is 94.3wt%, and the iron recovery rate is 96.78wt%.

[0110] Comparative Example 5

[0111] Example 1 was repeated, except that the mass ratio of the mixture to sodium chloride was controlled to be 100:3:.

[0112] The results showed that the slag viscosity was too high during the smelting process and the smelting purpose could not be achieved. At the same time, the chlorination volatilization rate of each metal was reduced. The volatilization rates of gold, silver, copper, lead and zinc were 38.2wt%, 42.3wt%, 48.4wt%, 69.5wt% and 74.8wt%, respectively. Due to the reduction in the amount of chlorinating agent, the slag viscosity was too high and its volatility was significantly reduced.

[0113] Comparative Example 6

[0114] Example 1 was repeated, except that the mass ratio of the mixture to the sodium chloride was controlled to be 100:22.

[0115] The results showed that the furnace condition deteriorated sharply, and the excess sodium chloride volatilized into the smoke before participating in the smelting reaction, resulting in a sharp increase in the amount of smoke, which increased by 60%. The volatility of gold, silver, copper, lead and zinc were 98.2wt%, 97.4wt%, 96.4wt%, 99.5wt% and 99.8wt% respectively. The CaO-SiO2-FeO molten slag entered the next reduction process to obtain the product pig iron, with a purity of 96.5wt% and an iron recovery rate of 96.7wt%.

[0116] Example 6

[0117] Example 1 was repeated, except that the mass ratio of the mixture to sodium chloride was controlled to be 100:5.

[0118] The results showed that the viscosity of the slag was high, but it could still be discharged smoothly. The volatilization rates of gold, silver, copper, lead and zinc were 88.4wt%, 78.6wt%, 86.5wt%, 92.6wt% and 97.4wt% respectively. The CaO-SiO2-FeO molten slag entered the next reduction process to obtain pig iron with a purity of 96.8wt% and an iron recovery rate of 97.4wt%.

[0119] Example 7

[0120] Example 1 was repeated, except that the mass ratio of the mixture to sodium chloride was controlled to be 100:18.

[0121] The results showed that the mixed material could be smelted and deslagging smoothly, but some sodium chloride volatilized directly into the flue gas, resulting in a large smoke rate and a 36% increase in smoke volume. The volatilization rates of gold, silver, copper, lead and zinc were 95.7wt%, 98.8wt%, 98.2wt%, 98.7wt% and 99.8wt% respectively.

[0122] Example 8 repeats Example 1, except that the added chlorinating agent is changed to magnesium chloride.

[0123] The results show that the mixed material can be smelted and deslagging smoothly. The volatilization rates of gold, silver, copper, lead and zinc are 93.4wt%, 96.2wt%, 97.9wt%, 98.1wt% and 99.4wt% respectively. The CaO-SiO2-FeO molten slag enters the reduction process to obtain the product pig iron, the purity of the pig iron is 95.7wt%, and the recovery rate of iron is 96.2wt%. The water slag is tested after water quenching (CaO-MgO): SiO2=1.1.

[0124] Example 9

[0125] Example 1 was repeated, except that the added chlorinating agent was changed to calcium chloride.

[0126] The results show that the mixed material can be smelted and deslagging smoothly. The volatilization rates of gold, silver, copper, lead and zinc are 93.4wt%, 96.2wt%, 97.9wt%, 98.1wt% and 99.4wt% respectively. The CaO-SiO2-FeO molten slag enters the reduction process to obtain the product pig iron, the purity of the pig iron is 95.7wt%, and the iron recovery rate is 96.2wt%. The water slag is tested after water quenching (CaO-MgO): SiO2=1.2.

[0127] 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 co-processing cyanide tailings and copper-containing sludge, characterized in that: The steps include: S1, adding the treated cyanide tailings, copper-containing sludge, the first reducing agent and the chlorinating agent into the first smelting furnace, introducing oxygen-rich gas, and performing chlorination smelting to obtain CaO-SiO2-FeO molten slag and flue gas rich in metal chlorides; The mass ratio of the total amount of the cyanide tailings and copper-containing sludge to be treated to the first reducing agent and the chlorinating agent is 100:5-20:5-20; The mass ratio of the cyanide tailings to be treated to the copper-containing sludge is 1-5:1, so that in the CaO-SiO2-FeO molten slag, the mass ratio of FeO to SiO2 is 1.0-2.4:1, and the mass ratio of CaO to SiO2 is 0.1-0.7:1; The chlorinating agent includes a chloride salt; The temperature of chlorination smelting is 1100-1300℃; The concentration of oxygen in the flue gas is 0.1~3vol%; In the cyanide tailings to be treated, the content of Fe is 30-60wt%, 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, and the content of Ag is 10-300g / t; in the copper-containing sludge, the content of Cu is 0.1-5wt%, the content of Au is 1-10g / t, the content of Ag is 10-200g / t, the content of Zn is 0-5wt%, and the content of CaO is 5-40wt%; S2, transferring the CaO-SiO2-FeO molten slag in the first smelting furnace to the second smelting furnace, adding a second reducing agent and a flux to the second smelting furnace, introducing an oxygen-rich gas, and performing reduction smelting at 1400-1600° C. to obtain CaO-MgO-SiO2 molten slag, molten iron and tail gas; 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 flux 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; The concentration of CO in the tail gas is 15-45 vol%.

2. The method according to claim 1, characterized in that The flux includes one or more of limestone, dolomite, quartz stone and gypsum; the chlorinating agent includes one or more of calcium chloride, sodium chloride, potassium chloride and magnesium chloride.

3. 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, coke, and natural gas; the second reducing agent includes one or more of carbon essence, waste activated carbon, coking coal, anthracite, coke, and natural gas.

4. The method according to claim 1, characterized in that: In S1, the chlorination smelting temperature is 1150-1250°C.

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

6. The method according to claim 5, characterized in that The oxygen concentration of the oxygen-rich gas is 60-80 vol%.

7. The method according to any one of claims 1 to 6, 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.

8. The method according to any one of claims 1 to 6, characterized in that: After S2, the CaO-MgO-SiO2 molten slag is water quenched to obtain water-quenched slag.

9. The method according to any one of claims 1 to 6, characterized in that: The tail gas is processed in sequence through waste heat utilization, cooling, dust collection and purification to obtain clean coal gas.

10. The method according to claim 9, characterized in that Before S1, the cyanide tailings and copper-containing sludge are input into a rotary kiln, and the materials are dried by the rotary kiln to obtain a mixed material; during this period, the clean coal gas is used as a heat source for the rotary kiln.

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