Method for recovering valuable metals from lead-containing solid waste and wet zinc smelting waste residue

Abstract

The present application relates to a kind of methods for recycling valuable metals in lead-containing solid waste and wet zinc smelting slag, belong to metallurgy, industrial solid waste and secondary resource comprehensive recycling technical field, and lead-containing solid waste and the waste slag produced in smelting process are handled collaboratively.By collaborative disposal with industrial lead-containing solid waste, valuable metals such as Cu, Pb, Zn, Ag in iron pyrite residue and lead-silver slag can be efficiently extracted and recycled, avoiding environmental risks caused by long-term storage of the two residues;After collaborative disposal of lead-containing solid waste and zinc smelting slag, the direct recovery rate of lead and zinc is more than 95% and 90% respectively, more than 85% of the silver metal is enriched in the crude lead phase, and the sulfur solidification rate is more than 95%;The slag is mainly CaO-FeO-SiO2 ternary slag, which can be directly used for producing building materials after water quenching;Elemental sulfur in the material is solidified in the matte phase, thereby avoiding SO2 flue gas pollution problems in the smelting process.

Description

Technical Field

[0001] This invention belongs to the field of metallurgical, industrial solid waste and secondary resource comprehensive recycling technology, specifically relating to a method for recovering valuable metals from lead-containing solid waste and hydrometallurgical zinc smelting slag. Background Technology

[0002] Lead, as a basic metallic raw material, is widely used in industrial fields such as lead-acid batteries, chemical corrosion protection materials, and cable protection. my country, as a major producer and consumer of lead, uses approximately 85% of its refined lead in the manufacture of lead-acid batteries. Since lead-acid batteries have a lifespan of only 2-3 years, a large number of them become obsolete each year. If not disposed of and recycled in a timely manner, this will not only accelerate the depletion of primary lead mineral resources but also cause serious environmental pollution. Lead sulfate paste is a flocculent precipitate formed during the charging and discharging process of lead-acid batteries. Its chemical composition is complex and variable, mainly... Lead sulfate paste is composed of PbSO4, PbO2, PbO, and a small amount of Pb, making it difficult to process. Currently, lead sulfate paste is often processed using a combined wet and pyrometallurgical process. First, desulfurizing agents such as NaOH, NaHCO3, and Na2CO3 are used to convert PbSO4 in the lead paste into PbO. Then, pyrometallurgical smelting is used to reduce PbO to metallic Pb. This process converts high-melting-point lead sulfate into low-melting-point lead carbonate or lead oxide through a wet process, which greatly reduces smelting energy consumption. However, the process is relatively complex, consumes a large amount of desulfurizing agents, and the desulfurized products are difficult to sell.

[0003] Zinc hydrometallurgical systems generate large quantities of hazardous waste annually, including lead-silver slag and iron ore slag, which are listed as HW48 hazardous waste in the National Hazardous Waste List. These slags contain significant amounts of valuable metals such as lead, zinc, and silver. Long-term accumulation can cause severe heavy metal leaching pollution and a substantial waste of valuable metal resources. Current treatment processes for lead-silver slag and iron ore slag mainly include hydrometallurgy, pyrometallurgy, and combined hydrometallurgical and pyrometallurgical processes. Hydrometallurgical processes can be further divided into washing, conventional hydrometallurgical solvent leaching, hydrothermal leaching, and microwave-assisted leaching. The washing process recovers a large amount of soluble zinc entrained in the leaching residue through agitation and washing. After neutralization and impurity removal of the washing solution, solvent extraction and zinc electrowinning back-extraction are performed to obtain a zinc-rich solution (Zn). While the recovery rates of Ni, Co, and Cu reached 99.8%, 90%, and 99.8%, respectively, valuable metal elements such as In, Ga, Ge, and Ag in the leaching residue were not effectively recovered (68 g / L). Conventional wet solvent leaching typically uses acidic leaching agents such as hydrochloric acid, sulfuric acid, and nitric acid, or alkaline leaching agents such as concentrated ammonia and caustic soda to disrupt the crystal structure of the leaching residue, causing it to dissolve and leach out. Metal elements such as Cu, Pb, Zn, Sb, As, and Sn are enriched in the solution, and then valuable metal elements are recovered through processes such as iron powder replacement, solvent extraction, thiourea, or cyanide precipitation. Metal elements are separated from iron, and iron-containing solutions are purified and impurities removed to prepare high-value iron-based functional materials. Hydrothermal leaching and microwave-assisted leaching are both external field enhanced leaching methods. By enhancing the leaching temperature field, electric field, and magnetic field, the leaching rate can be improved. However, energy consumption and cost issues need to be considered comprehensively, and they have not yet been industrialized. Although wet treatment processes can recover valuable metals such as Cu, Pb, and Ag from the slag and prepare high-value iron-based functional materials, they cannot avoid the complex process routes, high treatment costs, and difficulty in operating on a scale that can meet the needs of the zinc-iron mixed slag produced by large zinc smelting systems.

[0004] As the drawbacks of hydrometallurgical processes become increasingly apparent, the resource-based and harmless treatment of iron ore slag and lead-silver slag through pyrometallurgical processes is receiving increasing attention. Currently, pyrometallurgical treatment processes for zinc hydrometallurgical leaching slag mainly include "high-temperature sintering," "reduction roasting-magnetic separation," and "oxygen-enriched bottom-blown smelting-fumigation volatilization." The high-temperature sintering process involves sintering iron ore slag and lead-silver slag with quartz sand, limestone, dolomite, granite, etc., at temperatures above 1300℃ to form iron-based building materials. This process fixes both toxic and valuable metal elements in the slag phase, significantly reducing the leaching toxicity of toxic elements and eliminating their environmental pollution. However, it results in economic losses due to the loss of valuable metals. The reduction roasting-magnetic separation method reduces iron in the leaching slag to magnetic iron(III) oxide or metallic iron under the action of a reducing agent. Subsequently, iron sand, containing Cu, Zn, Pb, and S, can be produced through magnetic separation. B and other components enter the non-magnetic material along with the gangue components, from which valuable metal elements are recovered. The oxygen-enriched bottom-blown smelting-fumigation volatilization process involves feeding iron ore slag and lead-silver slag into a high-temperature smelting furnace. Oxygen-enriched air is blown into the bottom of the furnace to enhance the stirring of the melt. The element S in the leaching slag mainly enters the flue gas in the form of SO2, while the valuable metal elements Zn, Pb, and Ag enter the slag phase as oxides. The hot slag enters the fumigation furnace through a chute for reduction and volatilization. Finally, Zn, Pb, Ag, In, etc., enter the flue dust, from which the various valuable metals are recovered. This process has advantages such as large production scale, strong processing capacity, and wide raw material adaptability. However, because lead and zinc are reduced to the metallic phase during the high-temperature smelting process and then re-oxidized to zinc oxide and lead oxide at the secondary tuyer, without generating metallic lead, the precious metals in the leaching slag are not effectively captured, resulting in a low overall recovery rate of precious metals. Summary of the Invention

[0005] The purpose of this invention is to provide a method for recovering valuable metals from lead-containing solid waste and hydrometallurgical zinc smelting slag in order to solve the above-mentioned problems.

[0006] The present invention achieves the above objectives through the following technical solutions:

[0007] This invention provides a method for recovering valuable metals from lead-containing solid waste and hydrometallurgical zinc smelting slag. The specific steps for clean and efficient recovery of valuable metals are as follows:

[0008] S1. Mix 10-20 parts lead-containing solid waste and 40-60 parts desulfurizing agent using a mixer for 30-40 minutes at a speed of 20-26 r / min until homogeneous. Then add desulfurizing activator, flux, and carbonaceous reducing agent. Grind and mix the mixture in a double planetary ball mill. Pour the mixture into a corundum crucible and cover it with a perforated lid. Place the crucible in a side-blown melting furnace and introduce an inert atmosphere quantitatively through a digital flow meter at a temperature of 1200℃~1300℃. Melt for 1 hour and let stand for 2-4 hours to obtain crude lead, slag, metallic zinc, iron matte, and sulfur-containing dust.

[0009] S2. The crude lead obtained above is refined by pyrometallurgical refining and then electrolytically refined to obtain metallic lead and anode mud. Metallic silver is recovered from the anode mud.

[0010] As a further optimization of the present invention, the sulfur-containing dust in S1 is recovered through a dust collection system. The specific recovery steps are as follows: after the smelting temperature reaches 1200℃~1300℃, an alumina tube is inserted from the top of the furnace and extended to the bottom of the melt. Then, an inert atmosphere is quantitatively introduced through a digital flow meter, and the other end is connected to the alumina tube. The dust produced by smelting enters the cooler through the pipeline. When the dust is cooled to 320℃, cooling water is used to cool the flue gas. Then, the flue gas is desulfurized using the sodium alkali process. Finally, the flue gas produced by ionic liquid adsorption and desorption has a temperature not higher than 39℃ and an SO2 concentration not lower than 5.2%. The dust is collected from the pipeline after smelting is completed.

[0011] As a further optimization of the present invention, the lead-containing solid waste includes one or more of lead sulfate paste, CRT glass, lead-containing flue ash, lead anode mud, and lead-containing smelting slag.

[0012] As a further optimization of the present invention, the sulfur-fixing agent comprises 10-20 parts lead-silver slag and 10-20 parts iron alum slag by weight.

[0013] As a further optimization of the present invention, the atmosphere control in the side-blown melting furnace in S1 is achieved by adjusting the oxygen flow rate and the amount of carbonaceous reducing agent. The inert atmosphere is high-purity N2 or high-purity Ar, and the gas flow rate is 50 mL / min.

[0014] As a further optimization of the present invention, the sulfur-fixing activator is one or more of zinc oxide powder, sodium carbonate, and copper oxide, and the dosage is 8% to 12% of the total mass of the material; the carbonaceous reducing agent is metallurgical grade reduced coke, which has a fixed carbon content of 79.70% and a volatile matter content of 12.50%, and the dosage is 10% to 20% of the total mass of the material.

[0015] As a further optimization of the present invention, the flux mass is 10% to 20% of the total mass of lead-containing solid waste and zinc smelting slag.

[0016] The beneficial effects of this invention are as follows: By co-processing with industrial lead-containing solid waste, valuable metals such as Cu, Pb, Zn, and Ag can be efficiently extracted and recovered from iron ore slag and lead-silver slag, avoiding the environmental risks caused by long-term stockpiling of the two slags; after co-processing of lead-containing solid waste and zinc smelting slag, the direct recovery rates of metallic lead and zinc are over 95% and 90% respectively, over 85% of metallic silver is enriched in the crude lead phase, and the sulfur element solidification rate is over 95%; the slag is mainly composed of CaO-FeO-SiO2 ternary slag, which can be directly used for the production of building materials after water quenching; elemental sulfur in the material is solidified in the matte phase, thereby avoiding SO2 flue gas pollution problems during the smelting process. This invention has the advantages of a simple process, environmental friendliness, and high comprehensive recovery efficiency of valuable metals, and has good prospects for promotion and application in the field of secondary solid waste resource recycling. Attached Figure Description

[0017] Figure 1 This is a flow chart of the co-smelting process of lead-containing solid waste and hydrometallurgical zinc smelting slag according to the present invention. Detailed Implementation

[0018] The present application will now be described in further detail. It should be noted that the following specific embodiments are only used to further illustrate the present application and should not be construed as limiting the scope of protection of the present application. Those skilled in the art can make some non-essential improvements and adjustments to the present application based on the above application content.

[0019] I. Materials

[0020] Unless otherwise specified, all methods used in this invention are conventional methods known to those skilled in the art, and all reagents and materials used are commercially available products.

[0021] II. Methods

[0022] 2.1 Methods for recovering metals from lead-containing solid waste and hydrometallurgical zinc smelting slag:

[0023] S1. A mixture containing 15 parts lead-containing solid waste (specifically lead sulfate paste) and 50 parts desulfurizing agent was stirred for 35 minutes at a speed of 23 r / min using a mixer. After mixing evenly, a desulfurizing activator, flux, and carbonaceous reducing agent were added. The mixture was then ground and mixed in a double planetary ball mill and poured into a corundum crucible with a perforated lid. The crucible was then placed in a side-blown smelting furnace and smelted at 1300°C with an inert atmosphere introduced quantitatively through a digital flow meter. After smelting for 2 hours, the mixture was allowed to stand for 1 hour to obtain crude lead, slag, metallic zinc, iron matte, and sulfur-containing dust.

[0024] The sulfur-containing dust in S1 is recovered through a dust collection system. The specific recovery steps are as follows: after the smelting temperature reaches 1300℃, an alumina tube is inserted from the top of the furnace and extended to the bottom of the melt. Then, an inert atmosphere is quantitatively introduced through a digital flow meter. The other end is connected to the alumina tube. The dust produced by smelting enters the cooler through the pipeline. When the dust is cooled to 320℃, cooling water is used to cool the flue gas. Then, the flue gas is desulfurized using the sodium alkali process. Finally, the flue gas produced by ionic liquid adsorption and desorption has a temperature of 30℃ and an SO2 concentration of 7%. The dust is collected from the pipeline after smelting is completed.

[0025] The atmosphere control in the side-blown melting furnace in S1 is achieved by adjusting the oxygen flow rate and the amount of carbonaceous reducing agent. The inert atmosphere is high-purity N2 or high-purity Ar, and the gas flow rate is 50 mL / min.

[0026] The sulfur-fixing activator is sodium carbonate, and the dosage is 8% to 12% of the total mass of the material; the carbonaceous reducing agent is metallurgical grade reduced coke, which has a fixed carbon content of 79.70% and a volatile matter content of 12.50%, and the dosage is 10% to 20% of the total mass of the material; the fluxing agent is 15% of the total mass of lead-containing solid waste and zinc smelting slag.

[0027] S2. The crude lead obtained above is refined by pyrometallurgical refining and then electrolytically refined to obtain metallic lead and anode mud. Metallic silver is recovered from the anode mud.

[0028] The sulfur-fixing agent, by weight, comprises 15 parts lead-silver slag and 15 parts iron alum slag.

[0029] The method for recovering valuable metals from lead-containing solid waste and hydrometallurgical zinc smelting slag described above specifically includes the following chemical reaction formula:

[0030] PbS + FeO + C = Pb + FeS + CO↑;

[0031] PbS+Na2CO3+2C=Na2S+Pb+3CO↑;

[0032] PbS+2CuO+2C=Pb+Cu2S+2CO↑;

[0033] ZnS+Na2CO3+2C=Zn↑+Na2S+3CO↑;

[0034] ZnS+2CuO+2C=Zn↑+Cu2S+2CO↑.

[0035] Example 1

[0036] The main chemical components of No. 1 lead-silver slag and iron alum slag produced by the zinc hydrometallurgical system of a certain factory in Anhui Province, as well as the purchased lead sulfate paste, are listed in Table 1. The industrial composition and ash content test results of metallurgical grade coke are listed in Table 2.

[0037] Table 1. Main chemical composition (wt%) of lead sulfate paste, iron alum slag, and lead-silver slag.

[0038]

[0039] Table 2. Industrial composition and ash content analysis (wt%) of metallurgical grade reduced coke.

[0040]

[0041] According to the batching guidelines, the preset ratios are m(FeO) / m(SiO2) = 1.0 and m(CaO) / m(SiO2) = 0.6. Elemental sulfur is fixed according to theoretical values, requiring the addition of appropriate amounts of calcium oxide and ferrous oxide. The amount of metallurgical-grade reducing coke added is 14% of the total material mass, and the amount of sodium carbonate added is 10% of the total material mass. After thorough mixing and grinding, the materials are smelted at 1300℃ for 2 hours, kept at that temperature for 1 hour, then removed and poured into an iron mold for cooling and stratification. The flue dust is scraped from the corundum tube and collected. The obtained slag, matte, flue dust, and crude metal are weighed and analyzed. The final calculated direct recovery rates of lead, silver, and zinc are 95.23%, 88.26%, and 92.14%, respectively, with a sulfur fixation rate of 96.67%.

[0042] Example 2

[0043] The main chemical components of No. 2 lead-silver slag, iron alum slag, and purchased lead sulfate paste produced by the zinc hydrometallurgical system of a factory in Anhui are listed in Table 3.

[0044] Table 3. Main chemical composition (wt%) of lead sulfate paste, iron alum slag, and lead-silver slag.

[0045]

[0046]

[0047] According to the batching guidelines, the preset ratios of m(FeO) / m(SiO2) = 1.1 and m(CaO) / m(SiO2) = 0.7 are used. The elemental sulfur is fixed according to theoretical values. Appropriate amounts of calcium oxide and ferrous oxide need to be added respectively. The amount of metallurgical-grade reducing coke added is 16% of the total material mass, and the amount of sodium carbonate added is 8% of the total material mass. After the materials are thoroughly mixed and ground, they are placed at a constant temperature of 1300℃ for 2 hours and kept at that temperature for 1 hour. After that, they are taken out and poured into an iron mold for cooling and stratification. The soot is scraped out from the corundum tube and collected. The obtained slag, matte, soot and crude metal are weighed and analyzed. Finally, the direct recovery rates of lead, silver and zinc are calculated to be 96.12%, 86.38% and 93.32% respectively, and the sulfur fixation rate reaches 95.04%.

[0048] Example 3

[0049] The main chemical components of No. 2 lead-silver slag, iron alum slag, and purchased lead sulfate paste produced by the zinc hydrometallurgical system of a factory in Anhui are listed in Table 4.

[0050] Table 4. Main chemical composition (wt%) of lead sulfate paste, iron alum slag, and lead-silver slag.

[0051]

[0052] According to the batching guidelines, the preset ratios are m(FeO) / m(SiO2) = 1.0 and m(CaO) / m(SiO2) = 0.7. Elemental sulfur is fixed according to theoretical values, requiring the addition of appropriate amounts of calcium oxide and ferrous oxide. The amount of metallurgical-grade reducing coke added is 14% of the total material mass, and the amount of sodium carbonate added is 12% of the total material mass. After thorough mixing and grinding, the materials are smelted at 1300℃ for 2 hours, kept at that temperature for 1 hour, then removed and poured into an iron mold for cooling and stratification. The flue dust is scraped from the corundum tube and collected. The obtained slag, matte, flue dust, and crude metal are weighed and analyzed. The final calculated direct recovery rates of lead, silver, and zinc are 95.59%, 87.33%, and 91.76%, respectively, with a sulfur fixation rate of 96.47%.

[0053] Example 4

[0054] The main chemical components of No. 3 lead-silver slag, iron alum slag, and purchased lead sulfate paste produced by the zinc hydrometallurgical system of a factory in Anhui are listed in Table 5.

[0055] Table 5. Main chemical composition (wt%) of lead sulfate paste, iron alum slag, and lead-silver slag.

[0056]

[0057] According to the batching guidelines, the preset ratios are m(FeO) / m(SiO2) = 1.0 and m(CaO) / m(SiO2) = 0.7. Elemental sulfur is fixed according to theoretical values, requiring the addition of appropriate amounts of calcium oxide and ferrous oxide. The amount of metallurgical-grade reducing coke added is 16% of the total material mass, and the amount of sodium carbonate added is 10% of the total material mass. After thorough mixing and grinding, the materials are smelted at 1300℃ for 2 hours, kept at that temperature for 1 hour, then removed and poured into an iron mold for cooling and stratification. The flue dust is scraped from the corundum tube and collected. The obtained slag, matte, flue dust, and crude metal are weighed and analyzed. The final calculated direct recovery rates of lead, silver, and zinc are 95.59%, 87.33%, and 91.76%, respectively, with a sulfur fixation rate of 96.47%.

[0058] 2.2 Effect of sulfur-fixing agents on the direct recovery rate of metallic lead, metallic silver and metallic zinc

[0059] (1) The method steps for recovering metals from lead-containing solid waste and hydrometallurgical zinc smelting slag are referred to in 2.1, wherein the desulfurizing agent described in Example 5 is used instead of the desulfurizing agent in Example 1.

[0060] The raw material composition of the desulfurizing agent is shown in the table below:

[0061] Group Lead / silver residue / portion Iron sulfide residue / portion Example 1 15 15 Example 5 21 21

[0062] In Example 5, according to the batching criteria, the preset ratios of m(FeO) / m(SiO2) were 1.0 and m(CaO) / m(SiO2) were 0.71. The elemental sulfur was fixed according to theoretical values, requiring the addition of appropriate amounts of calcium oxide and ferrous oxide. The amount of metallurgical-grade reducing coke added was 14% of the total material mass, and the amount of sodium carbonate added was 10% of the total material mass. After the materials were thoroughly mixed and ground, they were smelted at a constant temperature of 1300℃ for 2 hours, kept at that temperature for 1 hour, and then removed and poured into an iron mold for cooling and stratification. The flue dust was scraped off from the corundum tube and collected. The obtained slag, matte, flue dust, and crude metal were weighed and analyzed. Finally, the direct recovery rates of lead, silver, and zinc were calculated to be 89.92%, 84.32%, and 90.13%, respectively, and the sulfur fixation rate reached 92.33%.

[0063] The embodiments described above are merely examples of several implementations of the present invention, and while the descriptions are relatively specific and detailed, they should not be construed as limiting the scope of the present invention. It should be noted that those skilled in the art can make various modifications and improvements without departing from the concept of the present invention, and these modifications and improvements all fall within the scope of protection of the present invention.

Claims

1. A method for recovering valuable metals from lead-containing solid waste and hydrometallurgical zinc smelting slag, characterized in that, The specific steps for the clean recycling of valuable metals are as follows: S1. A mixture containing 10-20 parts lead-containing solid waste and 40-60 parts desulfurizing agent is stirred for 30-40 minutes at a speed of 20-26 r / min until homogeneous. A desulfurizing activator, flux, and carbonaceous reducing agent are then added. The mixture is ground and mixed in a double planetary ball mill, poured into a corundum crucible, and covered with a perforated lid. The crucible is then placed in a side-blown melting furnace. An inert atmosphere is introduced quantitatively through a digital flow meter at 1200°C-1300°C. After melting for 1 hour, the mixture is allowed to stand for 2-4 hours to obtain crude lead, slag, metallic zinc, iron matte, and sulfur-containing dust. The desulfurizing agent comprises 10-20 parts lead-silver slag and 10-20 parts iron alum slag by mass. The desulfurizing activator is one or more of zinc oxide powder, sodium carbonate, and copper oxide, used at 8%-12% of the total material mass. S2. The obtained crude lead is refined by pyrometallurgical refining and then electrolytically refined to obtain metallic lead and anode mud. Metallic silver is recovered from the anode mud.

2. The method for recovering valuable metals from lead-containing solid waste and hydrometallurgical zinc smelting slag according to claim 1, characterized in that, The sulfur-containing dust in S1 is recovered through a dust collection system. The specific recovery steps are as follows: After the smelting temperature reaches 1200°C~1300°C, an alumina tube is inserted from the top of the furnace and extends to the bottom of the melt. Then, an inert atmosphere is quantitatively introduced through a digital flow meter. The other end is connected to the alumina tube. The flue gas produced by smelting enters the cooler through the pipeline. When the flue gas is cooled to 320°C, cooling water is used to cool the flue gas. Then, the flue gas is desulfurized using the sodium alkali process. Finally, the flue gas temperature is not higher than 39°C and the SO2 concentration is not lower than 5.2% by ionic liquid adsorption and desorption. After smelting is completed, the flue gas is collected from the pipeline.

3. The method for recovering valuable metals from lead-containing solid waste and hydrometallurgical zinc smelting slag according to claim 1, characterized in that, The lead-containing solid waste includes one or more of the following: lead sulfate paste, CRT glass, lead-containing flue ash, lead anode mud, and lead-containing smelting slag.

4. The method for recovering valuable metals from lead-containing solid waste and hydrometallurgical zinc smelting slag according to claim 2, characterized in that, The atmosphere control in the side-blown melting furnace in S1 is achieved by adjusting the oxygen flow rate and the amount of carbonaceous reducing agent. The inert atmosphere is high-purity N2 or high-purity Ar, and the gas flow rate is 50 mL / min.

5. The method for recovering valuable metals from lead-containing solid waste and hydrometallurgical zinc smelting slag according to claim 1, characterized in that, The carbonaceous reducing agent is metallurgical grade reduced coke, which has a fixed carbon content of 79.70% and a volatile matter content of 12.50%, and is used at a rate of 10% to 20% of the total mass of the material.

6. The method for recovering valuable metals from lead-containing solid waste and hydrometallurgical zinc smelting slag according to claim 1, characterized in that, The flux is 10% to 20% of the total mass of lead-containing solid waste and zinc smelting slag.

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