Method for comprehensive recovery of lead slag to prepare high-purity lead and enrich silver

Abstract

This invention discloses a method for the comprehensive recovery of lead slag to prepare high-purity lead and enrich silver, relating to the field of resource utilization technology for non-ferrous metal smelting waste. The method includes: ball milling and mixing lead slag and coke, then adding water to granulate the mixture to obtain activated granules; sintering the activated granules under a reducing atmosphere to obtain sintered conversion slag; leaching the sintered conversion slag with acetic acid solution to obtain lead acetate leachate and acetic acid leachate residue; subjecting the lead acetate leachate to sequential oxidation flocculation and zinc powder replacement for impurity removal to obtain a qualified purified solution; electrolyzing the qualified purified solution to obtain high-purity electrolytic lead; and leaching the acetic acid leachate residue with sulfuric acid solution to recover indium, germanium, tin, and bismuth, obtaining sulfuric acid leachate residue. After treatment using this method, the total lead recovery rate is ≥96%, the purity of the high-purity electrolytic lead is ≥99.998%, the silver content in the sulfuric acid leachate residue is ≥1.5 kg / t, and the silver enrichment factor is ≥7 times.

Description

Technical Field

[0001] This invention relates to the field of resource utilization technology for non-ferrous metal smelting waste residue, specifically to a method for the comprehensive recovery of lead slag to prepare high-purity lead and enrich silver. Background Technology

[0002] Lead is a crucial raw material in the non-ferrous metals industry. Large quantities of lead slag are produced as a byproduct in lead-zinc smelting, hydrometallurgical zinc refining, and zinc oxide production. This lead slag typically contains 40%–50% lead, along with various valuable metals such as zinc, silver, germanium, indium, bismuth, and tin, as well as impurities like silicon, iron, arsenic, and antimony. It is a typical complex polymetallic secondary resource. Current technologies for treating lead slag are mostly traditional, commonly employing pyrometallurgical direct reduction to recover crude lead or hydrometallurgical sulfuric acid leaching to recover some lead and silver. Only lead and silver are recovered for valuation, while rare metals such as germanium, indium, tin, and bismuth are largely not efficiently recovered, resulting in significant resource waste and economic losses. Furthermore, traditional processes suffer from low lead recovery rates, silver dispersion in the tailings preventing accumulation, rare metal loss, and significant environmental impact.

[0003] Existing publicly available technologies for the resource utilization of lead slag still have significant shortcomings. For example, traditional pyrometallurgical smelting processes involve high temperatures, high energy consumption, and the production of arsenic and lead dust in the flue gas, resulting in high environmental treatment costs and significant losses due to the high-temperature volatilization of rare and dispersed metals. Conventional hydrometallurgical processes often use hydrochloric acid and sulfuric acid systems, which suffer from severe equipment corrosion, high acid consumption, large amounts of acidic wastewater, and large amounts of subsequent neutralization waste. Some technologies use sodium chloride systems to leach lead, which can improve the lead leaching rate, but chloride ions cause severe equipment corrosion, making it difficult to remove impurities from the leaching solution, thus hindering the improvement of lead product purity, and failing to simultaneously achieve the recovery of indium, germanium, tin, and bismuth, as well as silver enrichment.

[0004] Some studies have proposed using an acetic acid system to leach lead slag, but since lead in the slag mainly exists in the form of lead sulfate, direct acetic acid leaching kinetics are slow and the lead leaching rate is extremely low, making industrial application difficult. Some data show that sodium carbonate conversion followed by acetic acid leaching, without reduction sintering and granulation activation, results in poor material reactivity, incomplete conversion, and a lead recovery rate of less than 90%. Other processes employ a combined pyrometallurgical enrichment-wet extraction process, but this does not achieve simultaneous recovery of indium, germanium, tin, and bismuth; silver is only simply recovered as a by-product, failing to achieve efficient enrichment and resulting in limited economic benefits. Still other processes only achieve lead recovery, without involving rare and dispersed metal recovery and silver enrichment, leading to low resource utilization.

[0005] In summary, the existing lead slag treatment technologies generally have the following shortcomings: (1) The conversion efficiency of lead sulfate is low, the direct leaching effect of acetic acid is poor, and the total lead recovery rate is difficult to exceed 96%; (2) The impurities in the leaching solution are not completely removed, and it is impossible to prepare high-purity electrolytic lead with a purity of ≥99.998%; (3) Indium, germanium, tin and bismuth rare metals are not recovered in a coordinated manner, and the resource utilization rate is low; (4) Silver is highly dispersed in the tailings and cannot be enriched more than 7 times, and the value of silver is not fully explored; (5) The process flow is long, the energy consumption is high, the corrosion is serious, and the environmental risks are high.

[0006] Therefore, developing a new method for the comprehensive recycling of lead slag that features efficient lead recovery, synergistic extraction of rare and dispersed metals, deep silver enrichment, and a simple and environmentally friendly process has become a pressing technical challenge in this field. Summary of the Invention

[0007] This invention addresses the shortcomings of existing technologies by providing a method for the comprehensive recovery of lead slag to prepare high-purity lead and enrich silver. Using lead slag, a byproduct of zinc smelting, as raw material, high-purity lead is prepared through a combined process of granulation activation, reduction sintering, selective acetic acid leaching, segmented deep purification, electrolytic lead extraction, and sulfuric acid leaching. Simultaneously, rare metals such as indium, germanium, tin, and bismuth are recovered, and silver is efficiently enriched.

[0008] Therefore, in a first aspect, the present invention provides a method for the comprehensive recovery of lead slag to prepare high-purity lead and enrich silver, comprising: Lead slag and coke are ball-milled and mixed, and then water is added to granulate the mixture to obtain activated granules. The activated granular material was sintered under a reducing atmosphere to obtain sintered conversion slag. The sintering conversion slag was leached with acetic acid solution to obtain lead acetate leachate and acetic acid leachate residue; The lead acetate leachate was subjected to oxidative flocculation and zinc powder replacement for impurity removal in sequence to obtain a qualified purified solution. The qualified purified solution is electrolyzed to obtain high-purity electrolytic lead; The acetic acid leaching residue was leached with sulfuric acid solution to recover indium, germanium, tin and bismuth, and to obtain sulfuric acid leaching residue.

[0009] Furthermore, the amount of coke added is 5% to 15% of the mass of lead slag; after ball milling and mixing, the proportion of material with a particle size ≥ -200 mesh is ≥ 90%; the particle size of the activated granules is 1 mm to 5 mm, and the moisture content is 8% to 15%.

[0010] Furthermore, the sintering temperature is 700℃~850℃, and the sintering time is 3h~6h; the reducing atmosphere is a mixture of carbon monoxide and nitrogen, wherein the volume fraction of carbon monoxide is 8%~15%, and the balance is nitrogen.

[0011] Furthermore, the leaching of the sintered conversion slag with acetic acid solution includes: controlling the liquid-to-solid ratio to be (5~8) mL:1g, the final pH to be 3.5~5.5, the leaching temperature to be 55℃~75℃, the leaching time to be 2h~4h, and the initial concentration of the acetic acid solution to be 1.5mol / L~3.0mol / L; the lead leaching rate is ≥95%.

[0012] Further, the oxidative flocculation and impurity removal includes: adding hydrogen peroxide with a mass fraction of 20%~40% to the lead acetate leaching solution, adjusting the pH to 4.0~4.5, reacting at 55℃~65℃ for 40min~60min, and simultaneously adding an anionic flocculant during the reaction; the amount of hydrogen peroxide added is 1.0g / L~1.5g / L, and the amount of anionic flocculant added is 2.0mg / L~3.0mg / L.

[0013] Furthermore, the temperature for zinc powder replacement and impurity removal is 50℃~60℃, the pH is 3.5~4.5, and the reaction time is 50min~120min; the particle size of the zinc powder is -300~-200 mesh, and the amount added is 1.08~1.15 times the theoretical amount for impurity replacement.

[0014] Furthermore, the qualified purification solution contains iron ≤0.001g / L, arsenic ≤0.0005g / L, cadmium ≤0.0005g / L, tin ≤0.0005g / L, and bismuth ≤0.0003g / L.

[0015] Furthermore, the electrolysis includes controlling the current density to be 180 A / m. 2 ~260A / m 2 The cell voltage is 2.4V~3.0V, the electrolysis temperature is 35℃~45℃, the lead ion concentration in the electrolyte is 80g / L~120g / L, the free acetic acid concentration is 30g / L~60g / L, and the pH is 4.0~5.5. Electrolyte additives are also added. These additives include a colloidal leveling agent and / or a grain-refining agent. The colloidal leveling agent is selected from gelatin and bone glue, and the grain-refining agent is selected from resorcinol and β-naphthol.

[0016] Furthermore, the leaching of the acetic acid residue with sulfuric acid solution includes: controlling the liquid-to-solid ratio to be (4~6) L:1g, the leaching temperature to be 70℃~85℃, the leaching time to be 4.5h~8h, and the initial concentration of the sulfuric acid solution to be 100g / L~160g / L; the total recovery rate of the metals indium, germanium, tin and bismuth is ≥80%.

[0017] Furthermore, the purity of the high-purity electrolytic lead is ≥99.998%, and the total lead recovery rate is ≥96%; the silver content in the sulfuric acid leaching residue is ≥1.5kg / t, and the silver enrichment factor is ≥7 times.

[0018] Compared with the prior art, the present invention has at least the following beneficial effects: This invention provides a method for the comprehensive recovery of lead slag to prepare high-purity lead and enrich silver. Through granulation activation and reduction sintering, lead sulfate can be efficiently converted into lead oxide, significantly improving the acetic acid leaching effect, with a total lead recovery rate of over 96%. Segmented deep purification can completely remove various impurities such as iron, arsenic, silicon, antimony, cadmium, tin, and bismuth, producing high-purity lead with a purity of not less than 99.998% through electrolysis. Simultaneously, indium, germanium, tin, and bismuth rare metals are efficiently recovered, with a total rare metal recovery rate exceeding 80%, greatly improving resource utilization. It also allows for the effective enrichment of silver in the tailings, with an enrichment multiple of 7 times or more, and a silver content in the tailings exceeding 1.5 kg / t. This invention features a simple process flow, mild reaction conditions, recyclable reagents, and no large amount of acidic wastewater generated, resulting in significant environmental benefits and good industrial application value. Attached Figure Description

[0019] To more clearly illustrate the technical solutions in the embodiments of the present invention, the accompanying drawings used in the description of the embodiments will be briefly introduced below. Obviously, the accompanying drawings described below are only some embodiments recorded in the embodiments of the present invention. For those skilled in the art, other drawings can be obtained based on these drawings.

[0020] Figure 1 The flowchart illustrates the principle of the process for preparing high-purity lead and enriching silver through comprehensive recycling of lead slag, as provided in this embodiment of the invention. Detailed Implementation

[0021] To better understand the above technical solutions, the technical solutions of the embodiments of this application will be described in detail below with reference to the accompanying drawings and specific embodiments. It should be understood that the embodiments of this application and the specific features in the embodiments are detailed descriptions of the technical solutions of the embodiments of this application, rather than limitations on the technical solutions of this application. In the absence of conflict, the embodiments of this application and the technical features in the embodiments can be combined with each other.

[0022] A first aspect of this invention provides a method for the comprehensive recovery of lead slag to prepare high-purity lead and enrich silver, comprising: Lead slag and coke are ball-milled and mixed, and then water is added to granulate the mixture to obtain activated granules. The activated granular material was sintered under a reducing atmosphere to obtain sintered conversion slag; The sintering conversion residue was leached with acetic acid solution to obtain lead acetate leachate and acetic acid leachate residue; The lead acetate leachate was subjected to oxidative flocculation and zinc powder replacement for impurity removal in sequence to obtain a qualified purified solution. High-purity electrolytic lead is obtained by electrolyzing the qualified purified solution. Acetic acid leaching residue was leached with sulfuric acid solution to recover metallic indium, germanium, tin and bismuth, and to obtain sulfuric acid leaching residue.

[0023] Specifically, after selective leaching with acetic acid, the acetic acid leaching residue is enriched with rare metals such as indium, germanium, tin, and bismuth, as well as the precious metal silver. A secondary leaching of this residue using sulfuric acid solution offers the following multiple technical advantages and process benefits: Selective and efficient leaching of rare and dispersed metals: Acetic acid leaching preferentially dissolves only lead, while indium, germanium, tin, and bismuth remain in the slag; the sulfuric acid system has a strong dissolving ability for the above rare and dispersed metals, which can efficiently destroy the mineral lattice, allowing indium, germanium, tin, and bismuth to fully enter the leaching solution, achieving simultaneous recovery of rare and dispersed metals, with a total recovery rate of over 80%, greatly improving the comprehensive resource utilization rate of lead slag.

[0024] Achieving efficient enrichment and separation of silver: Silver is almost insoluble under sulfuric acid leaching conditions and remains stably retained in the sulfuric acid leaching residue; after removing metallic impurities such as indium, germanium, tin, and bismuth through sulfuric acid leaching, silver is further concentrated and enriched, with an enrichment factor of more than 7 times. The silver content in the tailings is ≥1.5kg / t, which can be directly sold as silver concentrate, significantly enhancing the economic value of silver.

[0025] The process is seamlessly integrated and resources are utilized in stages: lead is first leached with acetic acid, followed by rare metals with sulfuric acid. The graded selective leaching is clearly divided and does not interfere with each other. This avoids the separation difficulties caused by the simultaneous leaching of multiple metals, simplifies the subsequent extraction process of indium, germanium, tin, and bismuth, and reduces the separation and purification costs.

[0026] The process conditions are mild and the equipment is well-suited: the temperature, concentration, and time parameters of sulfuric acid leaching are controllable, the reaction conditions are mild, and high temperature and high pressure are not required. Conventional anti-corrosion stirring tanks can be used to achieve this. The leached slurry has good sedimentation and filtration performance, and solid-liquid separation is simple, making it suitable for continuous and stable industrial production.

[0027] Waste reduction and significant environmental benefits: After lead recovery, rare and dispersed metal recovery, and silver enrichment, the final sulfuric acid leaching residue is an inert tailings with low impurities and stable properties, which can be further disposed of harmlessly or utilized as building material resources. This achieves full-process utilization of smelting waste residue in terms of reduction, harmlessness, and high value, without the discharge of large amounts of waste acid and hazardous waste.

[0028] The reagent system is compatible and free from cross-contamination: the front-end acetic acid system and the back-end sulfuric acid system have clear division of labor, the reagents do not conflict or cross-contaminate, and the leachate is processed separately, which facilitates the separate purification and sale of lead, rare metals and silver. The process flow is logical and the production control is simple.

[0029] In some embodiments, the amount of coke added is 5% to 15% of the mass of lead slag; after ball milling and mixing, the proportion of material with a particle size ≥ -200 mesh is ≥ 90%; the particle size of the activated granules is 1 mm to 5 mm, and the moisture content is 8% to 15%.

[0030] Specifically, when coke is added at 5% to 15% of the lead slag mass as a reducing agent and skeletal support, it can create a weak CO reducing atmosphere in situ during sintering, ensuring the full reduction of lead sulfate to lead oxide. It also loosens the particle pores, facilitating the penetration of reducing gas, while minimizing the introduction of ash impurities, thus balancing reduction efficiency and product purity. If the addition amount is too small (<5%), it will lead to insufficient carbon source for reduction, low CO concentration in the furnace, weak reducing atmosphere, incomplete reduction of lead sulfate, and a significant decrease in subsequent acetic acid leaching rate and total lead recovery rate. The sintered granules will also have a dense structure and poor permeability, further hindering the reduction reaction. If the addition amount is too large (>15%), it will result in excessive coke redundancy, increasing raw material costs. Excessive coke ash will be introduced into the system, increasing the content of impurities such as iron and silicon, increasing the load on subsequent purification and impurity removal. Simultaneously, the particle pores will be too large, resulting in decreased strength and easy pulverization during sintering, affecting the permeability of the material layer and the subsequent solid-liquid separation effect.

[0031] When the proportion of material particles with a particle size of -200 mesh or larger is ≥90%, the material fineness is sufficient, the lead slag and coke are microscopically mixed evenly, the interface contact is sufficient, and the reduction sintering reaction activity is high. Fine particles have more lattice defects, making it easier for phase reconstruction to occur, which is conducive to the conversion of lead sulfate to lead oxide. The granulation formability is good and the particle strength is uniform. If the particle size is too coarse (the proportion of -200 mesh is <90%), the mineral particle size is large, the lead slag and coke are not mixed evenly, there are few contact interfaces, the reduction reaction only takes place on the surface of the particles, and the internal conversion is incomplete. The material reactivity is low, the acetic acid leaching kinetics are worse, the lead leaching rate is reduced, and the granulation is prone to cracking, low strength, and easy pulverization during sintering. If the particle size is too fine, the grinding energy consumption increases significantly, the production cost increases, and the material specific surface area is too large. The surface is prone to air adsorption oxidation, agglomeration and clumping, the granulation adhesion is poor, and too much fine powder is prone to caking and pore closing during sintering, which hinders the diffusion of the reducing atmosphere inward and reduces the conversion effect.

[0032] When the particle size of the activated granules is 1mm~5mm, the particle size is moderate, the material layer has good air permeability and low pressure drop, the reducing atmosphere can penetrate the material layer evenly, the sintering and conversion are uniform, and the mechanical strength is moderate. It does not pulverize during transportation and charging. During leaching, the mineral particles are well dispersed, the mass transfer contact is sufficient, and the leaching efficiency is high. If the particle size is too small (<1mm), the particles are too fine, the material layer is densely packed, the air permeability is poor, the flow of reducing atmosphere is blocked, and the reduction inside and outside is uneven. During leaching, it is easy to form a slurry that is too fine, the solid-liquid separation filtration resistance is large, the filter cake carries a lot of mother liquor, resulting in lead loss. If the particle size is too large (>5mm), the volume of a single particle is large, and the reducing atmosphere is difficult to penetrate to the core of the particle. The phenomenon of surface reduction and internal unconverted transformation occurs. During leaching, the internal lead minerals are difficult to contact with the acetic acid solution, the leaching reaction is incomplete, the lead recovery rate decreases, and the particles are easily broken and produce powder, affecting the stability of the working conditions.

[0033] When the moisture content of activated granules is 8%~5%, the moisture content is suitable, the adhesion between mineral powder and coke is good, the granulation rate is high, the particle strength is moderate, and it is not easy to lose powder or crack. The moisture vaporizes moderately in the early stage of sintering, which can form a microporous structure inside the particles, increase the porosity, and facilitate the diffusion of reducing gas and the penetration of subsequent leachate. If the moisture content is too low (<8%), the material is dry and loose, the adhesion is insufficient, the granulation is difficult, the particle strength is low, and it is easy to break into powder during the conveying and sintering process. The permeability of the material layer becomes poor, which affects the reduction sintering effect. If the moisture content is too high (>15%), the material is too wet and easily agglomerates, the particle size is uneven, and the adhesion is serious. The evaporation of water in the early stage of sintering is large and the energy consumption is high. Particle bursting and caking are likely to occur. Moreover, the high moisture content of the material entering the furnace prolongs the heating time and also causes the water vapor partial pressure in the furnace to be too high, which inhibits the reduction reaction.

[0034] The aforementioned lead slag raw material is a by-product of lead smelting, hydrometallurgical zinc refining, and zinc oxide production processes. The main phase of lead is lead sulfate, accompanied by valuable metals such as silver, indium, germanium, tin, and bismuth, and contains conventional impurities such as silicon, iron, arsenic, and antimony. The lead content of the lead slag is controlled at 40%~50% by mass, with stable grade, which is suitable for the reduction sintering-acetic acid selective leaching system of this process. Moreover, the lead slag is a dry waste residue without a large amount of organic matter or combustible impurities, which can prevent the atmosphere from getting out of control and the abnormal precipitation of impurities during sintering.

[0035] Preferably, the amount of coke added is 10% of the mass of lead slag; after ball milling and mixing, the proportion of material particles with a size of -240 mesh is 95%; the particle size of the activated granules is 3 mm and the moisture content is 10%.

[0036] In some embodiments, the activated granules are subjected to reduction sintering, with the sintering temperature controlled at 700℃~850℃ and the sintering time at 3h~6h, so that the lead sulfate in the lead slag is reduced to lead oxide; the reducing atmosphere is a mixture of carbon monoxide and nitrogen, wherein the volume fraction of carbon monoxide is 8%~15% and the balance is nitrogen.

[0037] Specifically, a sintering temperature of 700℃~850℃ provides sufficient thermodynamic activation energy to promote the lattice dissociation of lead sulfate in lead slag, smoothly converting it into lead oxide that is easily leached by acetic acid under a weak CO reducing atmosphere. This results in a high lead sulfate conversion rate and well-matched reaction kinetics. It ensures complete phase transformation while avoiding the high-temperature volatilization loss of rare and precious metals such as indium, germanium, tin, bismuth, and silver. Simultaneously, energy consumption is controllable, and the equipment's fire resistance and corrosion resistance loads are moderate. If the temperature is too low (<700℃), lattice activation is insufficient, and lead sulfate is difficult to leach. Ineffective decomposition and reduction, incomplete phase transformation, with lead sulfate phase remaining dominant due to its difficulty in leaching; subsequent acetic acid leaching rate drops significantly, making it difficult to achieve a total lead recovery rate of over 96%; if the temperature is too high (>850℃), the material is prone to overburning, agglomeration, and melting, sealing internal mineral pores and hindering the penetration of reducing atmosphere, thus reducing the lead sulfate conversion rate; at the same time, rare and dispersed metals and silver are severely lost through high-temperature volatilization, resulting in a decrease in rare and dispersed metal recovery rate and a deterioration in silver enrichment effect; energy consumption increases sharply, furnace refractory material wear accelerates, and production costs increase.

[0038] When the sintering time is between 3 and 6 hours, the reducing atmosphere can fully penetrate the interior of the granular material, allowing lead sulfate sufficient time to complete reduction and crystal phase reconstruction, transforming it into highly active lead oxide. This results in uniform material reaction, good conversion consistency, stable batch-to-batch process indicators, and suitability for continuous industrial production. If the time is too short (<3 hours), the reduction reaction is insufficient, leading to uneven conversion between the material's interior and exterior. The surface layer is converted, while a large amount of lead sulfate remains internally, resulting in a low lead leaching rate. If the time is too long (>6 hours), the phase conversion has reached equilibrium, offering no significant benefit. Prolonged high-temperature holding causes ineffective energy consumption and waste, exacerbates material densification and pore closure during sintering, increases the loss of valuable metals through volatilization, and reduces the overall resource recovery rate.

[0039] A weak reducing atmosphere with a CO volume fraction of 8%~15% and a balance of nitrogen provides moderate reducing power, selectively reducing lead sulfate to lead oxide. This achieves efficient phase conversion without excessive reduction to form metallic lead. Nitrogen acts as an air barrier, preventing secondary oxidation of materials and inhibiting the volatilization of valuable metals, ensuring the lead leaching rate and the simultaneous recovery of silver and rare metals. If the CO concentration is too low (<8%), the reducing atmosphere is insufficient, the reduction driving force is weak, the conversion of lead sulfate to lead oxide is incomplete, and there are many residual difficult-to-leach phases, resulting in a lead recovery rate that does not meet the standard. If the CO concentration is too high (>15%), the reducing power is too strong, and lead sulfate and lead oxide are easily excessively reduced to metallic lead. Metallic lead is difficult to leach with acetic acid, which significantly reduces the lead leaching rate. At the same time, it increases the safety risk of combustible gases in the furnace and increases energy and reagent consumption.

[0040] The process involves ball milling and mixing lead slag with coke, granulating, and then sintering. This results in well-developed pore structures in the particles, which facilitates the diffusion and penetration of the CO reducing atmosphere from the surface to the interior, leading to a more uniform and thorough reduction and conversion reaction. This avoids problems such as poor permeability, insufficient local reduction, or over-sintering of powder sintering, ensuring a high lead recovery rate and effective enrichment and recovery of silver and rare metals from the source.

[0041] Preferably, the sintering temperature is 800℃ and the sintering time is 5h.

[0042] The reducing atmosphere preferably uses carbon monoxide generated in situ from the incomplete combustion of coke in the sintering furnace, combined with nitrogen as a protective carrier gas to form a weak reducing protective atmosphere. More preferably, the volume fraction of carbon monoxide is 12%, and the volume fraction of nitrogen is 88%.

[0043] In some embodiments, leaching the sintering conversion slag with acetic acid solution includes: controlling the liquid-to-solid ratio to be (5~8) mL:1g, the final pH to be 3.5~5.5, the leaching temperature to be 55℃~75℃, the leaching time to be 2h~4h, and the initial concentration of the acetic acid solution to be 1.5mol / L~3.0mol / L; the lead leaching rate is ≥95%.

[0044] Specifically, when the initial concentration of the acetic acid solution is between 1.5 mol / L and 3.0 mol / L, the concentration is moderate, allowing for sufficient complexation and dissolution with the lead oxide generated by reduction sintering. This results in favorable lead leaching kinetics, ensuring a lead leaching rate of ≥95%. Furthermore, the system has good acidity buffering capacity, preventing the large-scale dissolution of impurities and achieving selective lead leaching, thus reducing the pressure on subsequent purification and impurity removal. If the concentration is too low (<1.5 mol / L), the activities of hydrogen ions and acetate ions are insufficient, leading to weak dissolution capacity for lead oxide, slow lead leaching rate, incomplete leaching, and a significant decrease in lead recovery rate. The reaction reaches equilibrium time by a large margin, reducing production efficiency. If the concentration is too high (>3.0 mol / L), acid consumption increases significantly, production costs rise, and excessive acidity will dissolve large amounts of impurities such as iron, antimony, tin, and bismuth, causing a surge in the impurity load of the leaching solution. This increases the difficulty of subsequent oxidation flocculation and zinc powder replacement for impurity removal, while also exacerbating equipment corrosion and increasing the cost of environmental treatment of waste liquid.

[0045] When the liquid-to-solid ratio is (5~8) mL:1g, the slurry has good fluidity, the material is stirred evenly, the acetic acid and mineral particles are in full contact and mass transfer is sufficient, and the lead leaching is sufficient; the slurry concentration is moderate, the subsequent solid-liquid separation and filtration are smooth, the filter cake carries less mother liquor, and the lead loss is small; if the liquid-to-solid ratio is too small (<5:1), the slurry is too thick and viscous, it is easy to settle and clump, the material is not mixed evenly, mass transfer is hindered, and the lead leaching rate is low; the stirring load is large, filtration is difficult, and the loss of valuable metals is increased; if the liquid-to-solid ratio is too large (>8:1), the volume of the leaching liquid is too large, the lead ion concentration is diluted, the processing load of the subsequent purification and electrolysis processes increases, and the energy consumption increases; the acetic acid reagent is wasted seriously, and the overall economic efficiency deteriorates.

[0046] When the leaching temperature is between 55℃ and 75℃, moderately increasing the temperature accelerates the leaching reaction rate of acetic acid and lead oxide, shortens the leaching equilibrium time, and improves the lead leaching rate. A moderate temperature ensures leaching kinetics while inhibiting excessive dissolution of impurities, balancing leaching efficiency and solution purity. If the temperature is too low (<55℃), the reaction rate is slow, the leaching kinetics are poor, and lead dissolution is insufficient within the same time frame, resulting in a substandard leaching rate. This necessitates a significantly longer leaching time, reducing the production line's processing capacity. If the temperature is too high (>75℃), acetic acid volatilization results in severe losses, increased reagent consumption, and a high level of acid mist in the working environment. Water evaporates quickly, leading to large fluctuations in slurry concentration. It also easily promotes the hydrolysis of impurity ions to form colloids, increasing filtration difficulty and causing slight hydrolysis losses of lead salts.

[0047] A leaching time of 2-4 hours allows sufficient time for the acetic acid and mineral particles to react, enabling the lead leaching reaction to reach thermodynamic equilibrium and stably achieve a lead leaching rate of ≥95%. This process time is well-suited to the pace of continuous industrial production. If the time is too short (<2 hours), the leaching reaction will not reach equilibrium, the lead oxide will not dissolve sufficiently, and the lead leaching will be incomplete, directly reducing the overall lead recovery rate. If the time is too long (>4 hours), the increase in lead leaching is minimal, resulting in low marginal benefits. It also occupies equipment tank space, lengthens the production cycle, and reduces capacity. Prolonged acidic immersion will dissolve more impurities, increasing subsequent purification costs.

[0048] At a final pH of 3.5–5.5, lead oxide can be efficiently dissolved into lead acetate, while the hydrolysis and dissolution of impurities such as iron, silicon, and antimony can be inhibited, achieving selective and preferential leaching of lead. This avoids the hydrolysis of lead ions to form basic lead salt precipitates, reducing lead loss. If the pH is too low (<3.5), the acidity of the system is too high, and a large amount of impurities such as iron, arsenic, antimony, and bismuth will leach out simultaneously, resulting in severely excessive impurities in the leachate, which greatly increases the difficulty of subsequent purification and makes it difficult to prepare high-purity electrolytic lead. If the pH is too high (>5.5), the acidity of the system is insufficient, and lead ions are easily hydrolyzed to form basic lead acetate and lead hydroxide precipitates, which are suspended in the slurry and difficult to filter, causing a significant decrease in lead leaching rate and recovery rate.

[0049] Preferably, the initial concentration of the acetic acid solution is 2.5 mol / L, the liquid-to-solid ratio is 7 mL:1 g, the final pH is 4.5, the leaching temperature is 70 °C, and the leaching time is 3 h.

[0050] In some embodiments, oxidative flocculation for impurity removal includes: adding hydrogen peroxide with a mass fraction of 20% to 40% to lead acetate leaching solution, adjusting the pH to 4.0 to 4.5, reacting at 55°C to 65°C for 40 to 60 minutes, and simultaneously adding an anionic flocculant during the reaction; the amount of hydrogen peroxide added is 1.0 g / L to 1.5 g / L, and the amount of anionic flocculant added is 2.0 mg / L to 3.0 mg / L.

[0051] Specifically, the lead acetate leachate can be purified in stages by first oxidizing and flocculating to remove impurities such as iron, arsenic, silicon, and antimony.

[0052] When the hydrogen peroxide mass fraction is 20%~40%, the oxidizing capacity is moderate, which can fully oxidize low-valence impurities such as iron, arsenic, and antimony to high-valence states, facilitating subsequent hydrolysis, flocculation, and precipitation removal. The oxidation rate is also gradual, preventing localized over-oxidation side reactions and maintaining the stability of the lead acetate system. If the concentration is too low (<20%), the oxidizing capacity is insufficient, low-valence impurities are not completely oxidized, and iron, arsenic, and antimony residues are too high, making them difficult to remove completely and affecting the purity of high-purity lead. If the concentration is too high (>40%), the oxidizing power is too strong, easily leading to over-oxidation of system components, resulting in reagent waste, increased costs, and the generation of excessive bubbles, which disrupts the flocculation and sedimentation effect and exacerbates equipment corrosion.

[0053] When the hydrogen peroxide addition is 1.0 g / L to 1.5 g / L, the dosage is just enough to meet the quantitative oxidation requirements of iron, arsenic, silicon, and antimony impurities. The oxidation reaction is sufficient and the excess is controllable, with no excess hydrogen peroxide residue interfering with the subsequent zinc powder replacement process. If the dosage is too low (<1.0 g / L), the oxidant is insufficient, the impurities are not completely oxidized, the impurity removal rate after flocculation is low, and the impurities in the purified liquid exceed the standard. If the dosage is too high (>1.5 g / L), the excess hydrogen peroxide residue will consume the subsequent zinc powder, reduce the replacement and impurity removal effect, and at the same time decompose to generate a large number of bubbles, destroying the formation of flocculated flocs and worsening solid-liquid separation.

[0054] Within a pH range of 4.0 to 4.5, the system can efficiently hydrolyze oxidized iron, arsenic, silicon, and antimony ions to form hydroxide / silicate flocs, while ensuring that lead ions do not hydrolyze and precipitate, resulting in minimal lead loss. This optimal pH range is suitable for the best adsorption and bridging conditions of anionic flocculants, leading to the best flocculation and sedimentation effect. If the pH is too low (<4.0), the acidity is too high, making it difficult for impurity ions to hydrolyze into flocs, resulting in a significant decrease in flocculation removal rate and excessive iron, arsenic, and antimony residues. If the pH is too high (>4.5), it can easily cause slight hydrolysis of lead ions to form basic lead acetate precipitate, resulting in a decrease in lead recovery rate. At the same time, the flocs are too small and severely colloidal, making them difficult to filter and separate.

[0055] When the reaction temperature is between 55℃ and 65℃, the heating accelerates the oxidation reaction kinetics, promotes the hydrolysis and nucleation of impurities, and facilitates floc growth and dense sedimentation. The temperature range is moderate, balancing energy consumption and flocculation effect. If the temperature is below 55℃, the oxidation and hydrolysis reaction rates are slow, the flocs are formed slowly and the particles are small, the impurities are not completely removed, the sedimentation time is prolonged, and the production efficiency is low. If the temperature is above 65℃, the hydrogen peroxide decomposes too quickly, the effective oxidant is consumed in large quantities and the utilization rate is reduced. The solution evaporation is large, the acetic acid volatilization is aggravated, the system composition fluctuates greatly, and the flocs are easily broken and dissolved.

[0056] When the reaction time is 40-60 minutes, the entire process of oxidation, hydrolysis, adsorption, bridging, and flocculation of impurities is fully completed, the flocs grow stably, and iron, arsenic, silicon, and antimony are thoroughly removed, resulting in good process stability. If the reaction time is too short (<40 minutes), the oxidation and flocculation reaction is not complete, the flocs are not granulated, the impurities are not sufficiently retained, and the quality of the purified liquid does not meet the standards. If the reaction time is too long (>60 minutes), the formed flocs are prone to breakage and resuspension due to prolonged stirring, resulting in the re-dissolution of impurities. This also occupies equipment time and reduces processing capacity.

[0057] When the concentration of anionic flocculant is 2.0 mg / L to 3.0 mg / L, it can adsorb, bridge, and trap fine suspended colloidal impurities, accelerate floc aggregation and growth, result in fast sedimentation, clear filtrate, and efficient removal of fine suspended solids such as silicon, iron, arsenic, and antimony. If the concentration is too low (<2.0 mg / L), the flocculation and bridging effect is insufficient, the fine colloids cannot settle, the solution becomes turbid, and impurities enter subsequent processes with the filtrate. If the concentration is too high (>3.0 mg / L), excessive reagent will cause the colloidal particles to re-charge and disperse, resulting in restabilization, which makes sedimentation difficult. The solution viscosity will increase, the filtration resistance will rise, and organic impurities will be introduced, affecting the quality of subsequent electrolytic lead.

[0058] The preferred anionic flocculant is anionic polyacrylamide (PAM) flocculant, and the pH is preferably adjusted with sodium carbonate.

[0059] Preferably, the hydrogen peroxide has a mass fraction of 30%, an addition amount of 1.2 g / L, a pH of 4.2, and is reacted at 60°C for 50 min. The amount of anionic flocculant added is 2.5 mg / L.

[0060] In some embodiments, the temperature for zinc powder replacement to remove impurities is 50℃~60℃, the pH is 3.5~4.5, and the reaction time is 50min~120min; the particle size of the zinc powder is -300~-200 mesh, and the amount added is 1.08~1.15 times the theoretical amount for impurity replacement.

[0061] Specifically, the lead acetate leaching solution is purified in stages. After oxidation and flocculation to remove impurities, zinc powder reduction can remove impurities such as cadmium, tin, and bismuth.

[0062] When the reaction temperature is between 50℃ and 60℃, the kinetic rate of the zinc powder displacement reaction can be accelerated, improving the efficiency of cadmium, tin, and bismuth impurity displacement and precipitation. This ensures the displacement reaction proceeds completely while avoiding excessive self-dissolution of zinc powder and waste of reagents due to excessively high temperatures. It also ensures good sedimentation performance of impurity flocs and easy filtration. If the temperature is too low (<50℃), the reaction kinetics are slow, the displacement reaction is incomplete, the removal of cadmium, tin, and bismuth is insufficient, the impurities in the purified liquid exceed the standard, and subsequent electrolysis cannot produce 99.998% high-purity lead. The reaction time is forced to be longer, reducing production efficiency. If the temperature is too high (>60℃), the self-corrosion and self-dissolution of zinc powder are aggravated, the ineffective consumption increases significantly, and the reagent cost rises. Side reactions increase, and fine suspended colloidal particles are easily generated, making solid-liquid separation difficult. It also causes pH fluctuations in the system, affecting the displacement selectivity.

[0063] A pH range of 3.5–4.5 ensures the selective replacement of cadmium, tin, and bismuth by zinc powder without replacing the main metal lead. It also inhibits zinc powder hydrolysis, preventing the formation of zinc hydroxide flocs that coat the zinc powder surface, thus ensuring the replacement reaction proceeds continuously and fully. If the pH is too low (<3.5), the acidity is too high, resulting in a high hydrogen ion concentration. Zinc powder preferentially reacts with the acid to produce hydrogen, consuming a large amount of zinc powder and significantly reducing the effective utilization rate of impurities; the impurity removal rate is insufficient. If the pH is too high (>4.5), tin, bismuth, and cadmium ions are prone to premature hydrolysis, forming colloidal hydroxide precipitates that are difficult to filter out when suspended in the solution. Simultaneously, slight precipitation of basic lead salts is likely, causing lead loss and reducing the lead recovery rate.

[0064] A reaction time of 50-120 minutes allows for sufficient contact between zinc powder and impurity ions, achieving deep removal of cadmium, tin, and bismuth, meeting the internal control indicators for impurities in the purification solution. The reaction is stable, the process is easy to control, and it is suitable for continuous production. If the time is too short (<50 minutes), the displacement reaction will not reach equilibrium, impurities will not be completely removed, and the impurity content in the purification solution will exceed the standard, seriously affecting the quality of high-purity electrolytic lead. If the time is too long (>120 minutes), the increase in impurity displacement is minimal, resulting in low marginal returns. It also occupies equipment tank space, lengthens the production cycle, and reduces processing capacity. Prolonged residence time can cause back-dissolution, where a small amount of precipitated impurities dissolves back into the solution, worsening the purification effect.

[0065] Zinc powder with a particle size of -300 to -200 mesh is suitable, with sufficient specific surface area, allowing for full contact with impurity ions in the solution, resulting in high and fast displacement reaction activity; the particles have good settling properties, making it easy to filter and separate after the reaction, and preventing fine powder from penetrating the filter cloth and causing secondary pollution; if the particle size is too coarse (greater than -200 mesh), the specific surface area is small, the reaction activity is poor, the displacement rate is slow, and impurities are not completely removed; the utilization rate of zinc powder is low, and the amount used is increased in disguise; if the particle size is too fine (less than -300 mesh), the specific surface area is too large, the reaction is too fast, and local over-reaction is easy; the fine powder has extremely strong suspension properties, is difficult to settle and filter, and is easy to remain in the purification solution and be carried into the electrolysis process, causing impurities to be trapped in the electrode plates and affecting the purity of lead.

[0066] The zinc powder is added in a slightly excessive amount to compensate for the passivation of the zinc powder surface and the loss from local side reactions, ensuring the complete and deep replacement of cadmium, tin, and bismuth impurities. The excess amount is controllable, preventing reagent waste and excessive zinc ion enrichment in the solution. If the amount added is too small (<1.08 times the theoretical amount), there will be no surplus zinc powder to compensate for the loss, resulting in incomplete impurity replacement, excessive impurities in the purified solution, and failure to meet the electrolysis access requirements. If the amount added is too large (>1.15 times the theoretical amount), the zinc powder will be severely excessive, significantly increasing the reagent cost. Excess zinc powder residue in the solution will cause zinc ion enrichment in the electrolyte, interfering with electrolytic crystallization and increasing the load on subsequent wastewater treatment.

[0067] Preferably, the temperature for zinc powder replacement to remove impurities is 55℃, the pH is 4.0, and the reaction time is 90 min; the particle size of the zinc powder is -240 mesh, and the amount added is 1.13 times the theoretical amount for impurity replacement.

[0068] In some embodiments, the qualified purification solution contains iron ≤0.001g / L, arsenic ≤0.0005g / L, cadmium ≤0.0005g / L, tin ≤0.0005g / L, and bismuth ≤0.0003g / L.

[0069] In some embodiments, electrolysis includes: controlling the current density to 180 A / m 2 ~260A / m 2 The cell voltage is 2.4V~3.0V, the electrolysis temperature is 35℃~45℃, the lead ion concentration in the electrolyte is 80g / L~120g / L, the free acetic acid concentration is 30g / L~60g / L, and the pH is 4.0~5.5. Electrolyte additives are also added. The electrolyte additives include colloidal leveling agents and / or grain refiners. The colloidal leveling agent is selected from gelatin and bone glue, and the grain refiner is selected from resorcinol and β-naphthol.

[0070] 180A / m 2 ~260A / m 2 The optimal current density ensures a stable electrolytic deposition rate, resulting in dense lead grains with smooth, burr-free surfaces, enabling the stable production of high-purity electrolytic lead exceeding 99.998%. Furthermore, the high current efficiency and balanced energy consumption make it suitable for continuous industrial production. However, excessively low current densities (<180 A / m) hinder this process. 2 Slow lead deposition rate significantly reduces production capacity; excessively loose grain growth results in soft electrode material that is prone to electrolyte impurities, affecting lead purity; low equipment utilization leads to increased production costs; and excessively high current density (>260A / m) 2 Severe polarization can lead to dendrite growth, nodule formation, and pinholes on the electrode plates; impurity ions are prone to co-deposition, resulting in substandard purity of high-purity lead; current efficiency decreases, cell voltage increases, and power consumption surges; in severe cases, it can cause short circuits, affecting the safe and stable operation of the electrolytic cell.

[0071] When the cell voltage is 2.4~3.0V, it can match the internal resistance of the acetic acid electrolyte and the electromotive force of the electrolysis reaction in this system. The electrolysis reaction is stable and controllable, the energy consumption is optimal, the electrode plate crystal morphology is good, and the high-purity lead deposition quality is guaranteed. If the cell voltage is too low (<2.4V), the electrolysis driving force is insufficient, the lead deposition is slow or even incomplete, the current efficiency is low, and the production capacity is low. If the cell voltage is too high (>3.0V), the ineffective energy consumption soars and the power consumption cost increases. The side reactions of the electrolyte are aggravated, a large number of hydrogen ions are precipitated to generate bubbles, the electrode plate surface is rough and the grains are coarse. Impurities are more likely to precipitate, which will damage the quality of high-purity lead.

[0072] When the electrolysis temperature is between 35℃ and 45℃, the electrolyte ion migration rate is moderate, and the conductivity is excellent, which helps to keep impurity ions in the solution and prevent them from precipitating. The lead crystals are uniform and dense, and the electrolyte additives have the best effect, balancing purity and energy consumption. If the temperature is too low (<35℃), the electrolyte viscosity is high, the ion diffusion is slow, the cell pressure increases, and the power consumption increases. The activity of the additives deteriorates, and the electrode plates are prone to peeling and loosening. If the temperature is too high (>45℃), the acetic acid volatilization is accelerated, the acid consumption increases, and the operating environment is poor. The electrolyte water evaporates quickly, and the component concentration fluctuates greatly. The additives are prone to decomposition and failure, the electrode plate grains are coarse, and they are prone to impurities, resulting in a decrease in purity. It will also accelerate equipment corrosion.

[0073] The electrolyte lead ion concentration is 80 g / L to 120 g / L, which provides good electrodeposition kinetics, uniform and stable crystallization, and is less prone to concentration polarization, ensuring continuous and stable precipitation of high-purity lead. If the concentration is too low (<80 g / L), concentration polarization is severe, which can easily lead to hydrogen evolution side reactions, resulting in loose, blackened, and poor-quality electrodes; current efficiency decreases, limiting production capacity. If the concentration is too high (>120 g / L), the electrolyte viscosity increases significantly, ion mass transfer is hindered, and a concentration layer is easily formed at the solid-liquid interface; large grains and nodules are likely to occur, while increasing the electrolyte circulation and filtration load.

[0074] When the free acetic acid concentration is 30g / L~60g / L, it can maintain the stable acidity and buffer system of the electrolyte, inhibit the hydrolysis and precipitation of lead ions, and ensure the conductivity of the electrolyte. It is suitable for the action environment of additives, effectively inhibits the precipitation of impurities, and stabilizes the quality of high-purity lead. If the concentration is too low (<30g / L), the buffer capacity of the system is insufficient, and lead ions are easily hydrolyzed to form basic salt precipitates, resulting in turbid electrolyte and component imbalance. Impurities are easily hydrolyzed and co-precipitated, and the purity of lead decreases. If the concentration is too high (>60g / L), the acidity of the electrolyte is too strong, which aggravates equipment corrosion. If the hydrogen ion concentration is too high, hydrogen evolution is easy, causing porosity and looseness of the electrode plates. Acid consumption increases, and the cost of subsequent waste liquid treatment increases.

[0075] An electrolysis pH of 4.0–5.5 is within the stable range for lead ions, preventing hydrolysis and precipitation while inhibiting the leaching of impurity ions such as iron, arsenic, and bismuth. Appropriate additives can level and refine the grains, ensuring ultra-high purity of electrolytic lead. If the pH is too low (<4.0), the acidity is too strong, leading to severe hydrogen evolution, porosity in the electrodes, and a loose texture; equipment corrosion worsens, and energy consumption increases. If the pH is too high (>5.5), lead ions easily hydrolyze to form lead hydroxide and basic lead acetate precipitates, causing electrolyte turbidity; impurity metal ions also simultaneously leach out through hydrolysis, severely reducing the purity of electrolytic lead.

[0076] Electrolyte additives can be adsorbed at the grain growth interface of the electrode plate, playing a role in leveling and refining the grains, inhibiting dendrites and burrs, and preventing the co-deposition of harmful impurities. This is the key to preparing 99.998% high-purity electrolytic lead. If the amount added is too small, the leveling and crystal inhibition effects will be insufficient, the electrode plate will be rough, prone to nodule growth, and impurities will be easily trapped, resulting in substandard purity. If the amount added is too large, the electrolyte viscosity will increase significantly, the conductivity will decrease, the cell voltage will increase, and the power consumption will increase. Organic additives are easy to be trapped inside the lead plate, affecting the physicochemical quality of high-purity lead.

[0077] Preferably, the current density is 220 A / m 2 The cell voltage is 2.6V, the electrolysis temperature is 40℃, the lead ion concentration in the electrolyte is 100g / L, the free acetic acid concentration is 45g / L, the pH is 4.8, and the electrolyte additive is a mixture of gelatin and resorcinol, with an addition amount of 0.3g / L.

[0078] In some embodiments, leaching the acetic acid leaching residue with sulfuric acid solution includes: controlling the liquid-to-solid ratio to be (4~6) L:1g, the leaching temperature to be 70℃~85℃, the leaching time to be 4.5h~8h, and the initial concentration of the sulfuric acid solution to be 100g / L~160g / L; the total recovery rate of indium, germanium, tin and bismuth is ≥80%.

[0079] Specifically, when the initial sulfuric acid concentration is 100 g / L to 160 g / L, sufficient hydrogen ion and sulfate ion activity is ensured, resulting in efficient leaching of valuable metals such as indium, germanium, tin, and bismuth from the acetic acid leaching residue. The leaching reaction kinetics are sufficient, and the total recovery rate of the four rare metals is consistently ≥80%. If the concentration is too low (<100 g / L), the acidity is insufficient, the driving force for the metal leaching reaction is weak, indium, germanium, tin, and bismuth are not completely leached, and the recovery rate of rare metals drops significantly. The leaching reaction rate is slow, and the production cycle is prolonged. If the concentration is too high (>160 g / L), acid consumption increases significantly, production costs rise, solution viscosity increases, and solid-liquid separation becomes difficult. Strong acidity exacerbates equipment corrosion and easily generates insoluble sulfates that coat mineral particles, which in turn inhibits metal leaching. It also generates a large amount of waste acid, greatly increasing the pressure on environmental treatment.

[0080] When the liquid-to-solid ratio is (4~6) L:1g, the slurry has good fluidity, sufficient stirring and mass transfer, uniform contact between mineral particles and sulfuric acid solution, and thorough leaching of valuable metals; it facilitates subsequent filtration and washing operations and ensures good process continuity; if the liquid-to-solid ratio is too small (<4:1), the slurry is too concentrated and viscous, stirring is difficult, the material is prone to settling and clumping, mass transfer is hindered, and the metal leaching rate is low; the solid-liquid separation resistance is high, the filter cake carries more mother liquor, and metal loss increases; if the liquid-to-solid ratio is too large (>6:1), the volume of the leachate is too large, the subsequent enrichment and concentration processes have high load, high energy consumption, and increased equipment footprint; the metal ion concentration is diluted, the subsequent extraction cost increases, and reagent consumption increases.

[0081] A leaching temperature of 70℃~85℃ can improve the reaction kinetic rate, accelerate the destruction of mineral lattices and the dissolution of valuable metals, shorten the leaching time, and ensure high leaching rates of indium, germanium, tin, and bismuth. A moderate temperature balances energy consumption and equipment tolerance. If the temperature is too low (<70℃), the reaction rate is slow, the leaching kinetic conditions are poor, the metal dissolution is insufficient in the same time, and the recovery rate of rare and dispersed metals does not meet the standards; the leaching time needs to be significantly extended, reducing production capacity. If the temperature is too high (>85℃), the water evaporation is large, the acid mist volatilization is serious, the operating environment is poor, and the acid consumption increases; energy consumption increases significantly, the requirements for high temperature resistance and corrosion prevention of equipment are increased, and investment and maintenance costs increase; it is easy to cause some metal ions to hydrolyze and precipitate, which will reduce the recovery rate.

[0082] A leaching time of 4.5h to 8h allows for sufficient reaction of indium, germanium, tin, and bismuth minerals, achieving deep leaching, reaching thermodynamic leaching equilibrium, ensuring a total rare metal recovery rate of ≥80%, and strong process stability. If the time is too short (<4.5h), the reaction will not reach equilibrium, and valuable metals inside the minerals will not have enough time to dissolve, resulting in a low leaching rate and serious resource waste. If the time is too long (>8h), the leaching increment is minimal, resulting in low marginal benefits; it also occupies equipment capacity, lengthens the production cycle, and reduces the throughput; prolonged high-temperature acidic environment exacerbates equipment corrosion and increases operation and maintenance costs.

[0083] Preferably, the initial concentration of the sulfuric acid solution is 130 g / L, the liquid-to-solid ratio is 5 L: 1 g, the leaching temperature is 80 °C, and the leaching time is 6 h.

[0084] In some embodiments, the purity of high-purity electrolytic lead is ≥99.998%, the total lead recovery rate is ≥96%, the silver content in the sulfuric acid leaching residue is ≥1.5kg / t, and the silver enrichment factor is ≥7 times.

[0085] Example 1: A method for the comprehensive recycling of lead slag to prepare high-purity lead and enrich silver. 1. Preparation method 1) Granulation and activation: Add 100g of coke to 1000g of lead slag and mix by ball milling until the particle size of the material is 95% -240 mesh. Add water to granulate, control the particle size to 3mm and the moisture content to 10%, and obtain activated granules.

[0086] 2) Reduction sintering: The activated granular material is reduced and sintered at 800℃ for 5 hours in a reducing atmosphere with a carbon monoxide volume fraction of 12% and a nitrogen gas volume fraction of 88% to obtain sintered conversion slag.

[0087] 3) Acetic acid leaching: The sintering conversion slag was leached with an initial concentration of 2.5 mol / L acetic acid solution at a leaching temperature of 70℃, a liquid-to-solid ratio of 7:1 (mL / g), a leaching time of 3 h, and a final pH of 4.5. After filtration, lead acetate leaching solution and acetic acid leaching slag were obtained.

[0088] 4) Segmented purification: First, add 1.2 g / L of 30% hydrogen peroxide for oxidation, adjust the pH to 4.2 with sodium carbonate, heat to 60℃ and react for 50 min, while adding 2.5 mg / L of anionic PAM for flocculation to remove iron, arsenic, silicon, and antimony; after filtration, add zinc powder with a particle size of -240 mesh at 55℃ and pH 4.0, the amount added is 1.13 times the theoretical amount, and react for 90 min to selectively remove cadmium, tin, and bismuth; the purified solution meets the following requirements: Fe≤0.001 g / L, As≤0.0005 g / L, Cd≤0.0005 g / L, Sn≤0.0005 g / L, Bi≤0.0003 g / L, to obtain a qualified purified solution.

[0089] 5) Electrolytic purification: Electrolyze the qualified purified solution, controlling the current density at 220A / m³. 2 The cell voltage was 2.6V, the electrolysis temperature was 40℃, the electrolyte contained 100g / L of Pb, 45g / L of free acetic acid, and pH 4.8. A mixture of 0.3g / L of gelatin and resorcinol was added as an electrolyte additive to obtain high-purity electrolytic lead.

[0090] 6) Sulfuric acid leaching: The acetic acid leaching residue is leached with sulfuric acid at an initial concentration of 130 g / L, at a leaching temperature of 80℃, a liquid-to-solid ratio of 5:1 (L / g), and a leaching time of 6 hours. After filtration, sulfuric acid leachate and sulfuric acid leaching residue are obtained. The process flow is as follows: Figure 1 As shown.

[0091] 2. Results The total lead recovery rate was 96.8%, and the purity of electrolytic lead was 99.999%. The recovery rates of rare and dispersed metals were: indium 88%, germanium 81%, tin 92%, and bismuth 83%. The silver content in the sulfuric acid leaching residue was 1.52 kg / t, which is enriched by 7.2 times.

[0092] Example 2: A method for the comprehensive recovery of lead slag to prepare high-purity lead and enrich silver. 1. Preparation method 1) Granulation and activation: Add 50g of coke to 1000g of lead slag and mix by ball milling until the material particle size of -220 mesh is 90%. Add water to granulate, control the particle size to 1mm and the moisture content to 8%, and obtain activated granules.

[0093] 2) Reduction sintering: The activated granular material is reduced and sintered at 700°C for 3 hours in a reducing atmosphere with a carbon monoxide volume fraction of 12% and a nitrogen gas volume fraction of 88% to obtain sintered conversion slag.

[0094] 3) Acetic acid leaching: The sintering conversion slag was leached with an initial concentration of 1.5 mol / L acetic acid solution at a leaching temperature of 55℃, a liquid-to-solid ratio of 5:1 (mL / g), a leaching time of 2 h, and a final pH of 3.5. After filtration, lead acetate leaching solution and acetic acid leaching slag were obtained.

[0095] 4) The steps of segmented purification, electrolytic purification, and sulfuric acid leaching are the same as in Example 1.

[0096] 2. Results The total lead recovery rate was 96.1%, and the purity of electrolytic lead was 99.9984%; the silver content in the sulfuric acid leaching residue was 1.50 kg / t, which is 7.0 times enriched.

[0097] Example 3: A method for the comprehensive recovery of lead slag to prepare high-purity lead and enrich silver. 1. Preparation method 1) Granulation and activation: Add 150g of coke to 1000g of lead slag and mix by ball milling until the particle size of the material is 98% -280 mesh. Add water to granulate, control the particle size to 5mm and the moisture content to 15%, and obtain activated granules.

[0098] 2) Reduction sintering: The activated granular material is reduced and sintered at 850°C for 6 hours in a reducing atmosphere with a carbon monoxide volume fraction of 12% and a nitrogen gas volume fraction of 88% to obtain sintered conversion slag.

[0099] 3) Acetic acid leaching: The sintering conversion slag was leached with an initial concentration of 3.0 mol / L acetic acid solution at a leaching temperature of 75℃, a liquid-to-solid ratio of 8:1 (mL / g), a leaching time of 4 h, and a final pH of 5.5. After filtration, lead acetate leaching solution and acetic acid leaching residue were obtained.

[0100] 4) The steps of segmented purification, electrolytic purification, and sulfuric acid leaching are the same as in Example 1.

[0101] 2. Results The total lead recovery rate was 97.2%, and the purity of electrolytic lead was 99.999%. The silver content in the sulfuric acid leaching residue was 1.53 kg / t, which is 7.3 times enriched.

[0102] Example 4: A method for the comprehensive recycling of lead slag to prepare high-purity lead and enrich silver. 1. Preparation method 1) Granulation and activation: Add 80g of coke to 1000g of lead slag and mix by ball milling until the material particle size of -240 mesh is 95%. Add water to granulate, control the particle size to 2mm and the moisture content to 12%, and obtain activated granules.

[0103] 2) Reduction sintering: The activated granular material is reduced and sintered at 750°C for 4 hours in a reducing atmosphere with a carbon monoxide volume fraction of 12% and a nitrogen gas volume fraction of 88% to obtain sintered conversion slag.

[0104] 3) Acetic acid leaching: The sintering conversion slag was leached with an initial concentration of 2.0 mol / L acetic acid solution at a leaching temperature of 65℃, a liquid-to-solid ratio of 6:1 (mL / g), a leaching time of 2.5 h, and a final pH of 4.0. After filtration, lead acetate leaching solution and acetic acid leaching slag were obtained.

[0105] 4) The steps of segmented purification, electrolytic purification, and sulfuric acid leaching are the same as in Example 1.

[0106] 2. Results The total lead recovery rate was 96.5%, and the purity of electrolytic lead was 99.9986%; the silver content in the sulfuric acid leaching residue was 1.51 kg / t, which is 7.1 times enriched.

[0107] Example 5: A method for the comprehensive recovery of lead slag to prepare high-purity lead and enrich silver. 1. Preparation method Unlike Example 1, in the acetic acid leaching step, an acetic acid solution with an initial concentration of 1.8 mol / L was used to leach the sintering conversion slag. The leaching temperature was 65°C, the liquid-to-solid ratio was controlled at 6:1 (mL / g), the leaching time was 3 hours, and the final pH was controlled at 4.2. After filtration, lead acetate leaching solution and acetic acid leaching slag were obtained.

[0108] 2. Results The total lead recovery rate was 96.3%, and the purity of electrolytic lead was 99.999%. The silver content in the sulfuric acid leaching residue was 1.50 kg / t, which is 7.0 times enriched.

[0109] Comparative Example 1 1. Preparation method Other conditions are the same as in Example 1, except that the granulation and activation step is omitted, and the lead slag is directly sintered.

[0110] 2. Results Uneven sintering resulted in a 72% conversion rate of lead sulfate to lead oxide, a lead leaching rate of 81.3%, and a total lead recovery rate of 88.6%; silver enrichment was 3.1 times.

[0111] Comparative Example 2 1. Preparation method Other conditions are the same as in Example 1, except that the reduction sintering step is omitted and the activated granules are directly leached with acetic acid.

[0112] 2. Results The lead leaching rate was 12.7%, and the total lead recovery rate was 11.2%, making industrialization impossible.

[0113] Comparative Example 3 1. Preparation method Other conditions are the same as in Example 1, except that zinc powder is not added for displacement and impurity removal.

[0114] 2. Results Electrolytic lead purity was 99.52%, but impurities exceeded the standard; silver enrichment was 2.8 times.

[0115] By comparing the above comparative examples with the embodiments of the present invention, it can be seen that the present invention significantly overcomes the bottlenecks of traditional processes, such as low lead leaching rate, incomplete impurity removal, poor purification effect, inability to recover rare and dispersed metals, and difficulty in silver enrichment, through granulation activation, reduction sintering modification, selective acetic acid leaching, segmented deep purification, and electrolytic refining of lead slag raw materials. Even in complex multi-metal lead slag systems, the present invention can still achieve a total lead recovery rate of over 96%, greatly improving the utilization rate of raw materials. Simultaneously, it adopts a green route of reduction sintering-acetic acid leaching instead of traditional methods. Replacing the traditional high-acid, high-corrosion, and high-energy-consumption process, the conversion rate of lead sulfate to lead oxide is ≥97%, and the lead leaching rate is ≥95%. While improving the purity of high-purity electrolytic lead to ≥99.998% and ensuring stable electrowinning quality, it significantly reduces production costs and energy consumption, and achieves high-value, harmless, and resource-based utilization of smelting waste. Indium, germanium, tin, and bismuth are simultaneously and efficiently recovered, with a total recovery rate of indium, germanium, tin, and bismuth ≥80%. The silver enrichment factor is ≥7 times, and the silver content reaches 1.5 kg / t, resulting in a significant improvement in silver efficiency.

[0116] The entire process generates no large amounts of waste acid, uses no highly toxic gases, and discharges no hazardous waste. The process is green, safe, and can be stably scaled up. Compared with existing pyrometallurgical reduction processes, sodium chloride leaching processes, direct acetic acid leaching processes, and traditional water washing purification processes, this invention has significant synergistic advantages and outstanding inventiveness in terms of raw material compatibility, resource utilization rate, preparation cost, product quality, environmental safety, synergistic recovery of rare and dispersed metals, and efficient silver enrichment.

[0117] It will be readily understood by those skilled in the art that the above-described advantageous methods can be freely combined and superimposed without conflict. The above are merely preferred embodiments of this application and are not intended to limit this application. Any modifications, equivalent substitutions, and improvements made within the spirit and principles of this application should be included within the protection scope of this application. The above are merely preferred embodiments of this application. It should be noted that for those skilled in the art, several improvements and modifications can be made without departing from the technical principles of this application, and these improvements and modifications should also be considered within the protection scope of this application.

Claims

1. A method for the comprehensive recovery of lead slag to prepare high-purity lead and enrich silver, characterized in that, include: Lead slag and coke are ball-milled and mixed, and then water is added to granulate the mixture to obtain activated granules. The activated granular material was sintered under a reducing atmosphere to obtain sintered conversion slag. The sintering conversion slag was leached with acetic acid solution to obtain lead acetate leachate and acetic acid leachate residue; The lead acetate leachate was subjected to oxidative flocculation and zinc powder replacement for impurity removal in sequence to obtain a qualified purified solution. The qualified purified solution is electrolyzed to obtain high-purity electrolytic lead; The acetic acid leaching residue was leached with sulfuric acid solution to recover indium, germanium, tin and bismuth, and to obtain sulfuric acid leaching residue.

2. The method according to claim 1, characterized in that, The amount of coke added is 5% to 15% of the mass of lead slag; after ball milling and mixing, the proportion of material with a particle size of ≥-200 mesh is ≥90%; the particle size of the activated granules is 1mm to 5mm, and the moisture content is 8% to 15%.

3. The method according to claim 1, characterized in that, The sintering temperature is 700℃~850℃, and the sintering time is 3h~6h; the reducing atmosphere is a mixture of carbon monoxide and nitrogen, wherein the volume fraction of carbon monoxide is 8%~15%, and the balance is nitrogen.

4. The method according to claim 1, characterized in that, The leaching of the sintered conversion slag with acetic acid solution includes: controlling the liquid-to-solid ratio to be (5~8) mL:1g, the final pH to be 3.5~5.5, the leaching temperature to be 55℃~75℃, the leaching time to be 2h~4h, and the initial concentration of the acetic acid solution to be 1.5mol / L~3.0mol / L; the lead leaching rate is ≥95%.

5. The method according to claim 1, characterized in that, The oxidation flocculation and impurity removal process includes: adding hydrogen peroxide with a mass fraction of 20% to 40% to the lead acetate leachate, adjusting the pH to 4.0 to 4.5, reacting at 55℃ to 65℃ for 40 to 60 minutes, and simultaneously adding an anionic flocculant during the reaction; the amount of hydrogen peroxide added is 1.0 g / L to 1.5 g / L, and the amount of anionic flocculant added is 2.0 mg / L to 3.0 mg / L.

6. The method according to claim 1, characterized in that, The zinc powder is used for impurity removal at a temperature of 50℃~60℃, a pH of 3.5~4.5, and a reaction time of 50min~120min. The zinc powder has a particle size of -300~-200 mesh and is added at 1.08~1.15 times the theoretical amount for impurity removal.

7. The method according to claim 1, characterized in that, The qualified purification solution contains iron ≤0.001g / L, arsenic ≤0.0005g / L, cadmium ≤0.0005g / L, tin ≤0.0005g / L, and bismuth ≤0.0003g / L.

8. The method according to claim 1, characterized in that, The electrolysis includes: controlling the current density to 180 A / m³. 2 ~260A / m 2 The cell voltage is 2.4V~3.0V, the electrolysis temperature is 35℃~45℃, the lead ion concentration in the electrolyte is 80g / L~120g / L, the free acetic acid concentration is 30g / L~60g / L, and the pH is 4.0~5.

5. Electrolyte additives are also added. These additives include a colloidal leveling agent and / or a grain-refining agent. The colloidal leveling agent is selected from gelatin and bone glue, and the grain-refining agent is selected from resorcinol and β-naphthol.

9. The method according to claim 1, characterized in that, The leaching of the acetic acid residue with sulfuric acid solution includes: controlling the liquid-to-solid ratio to be (4~6) L:1g, the leaching temperature to be 70℃~85℃, the leaching time to be 4.5h~8h, and the initial concentration of the sulfuric acid solution to be 100g / L~160g / L; the total recovery rate of the metals indium, germanium, tin and bismuth is ≥80%.

10. The method according to any one of claims 1-9, characterized in that, The high-purity electrolytic lead has a purity of ≥99.998% and a total lead recovery rate of ≥96%; the silver content in the sulfuric acid leaching residue is ≥1.5kg / t, and the silver enrichment factor is ≥7 times.

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