Method for efficiently separating and enriching zinc and indium from lead-zinc liquid slag

Abstract

The invention discloses a method for efficiently separating and enriching zinc and indium from lead-zinc liquid slag, and belongs to the technical field of metallurgical solid waste resource utilization. The method comprises the following steps: directly carrying out neutral leaching on lead-zinc liquid slag by adopting a dilute sulfuric acid solution, carrying out reduction or replacement impurity removal on the obtained zinc sulfate leachate to obtain a purified zinc sulfate solution, and carrying out electrolysis to produce cathode zinc; carrying out acid leaching on the obtained indium-enriched leaching residues by adopting a sulfuric acid solution to obtain indium-containing leaching liquid and leaching residues; the indium-containing leachate is subjected to extraction-reverse extraction, and an indium-rich solution is obtained; and the indium-rich solution and elemental zinc are subjected to a replacement reaction, and sponge indium is obtained. According to the method, selective leaching and separation of zinc and indium can be achieved by precisely regulating and controlling reaction kinetics and thermodynamics conditions of two-stage leaching, and the technical problems that in traditional one-step leaching, zinc and indium are difficult to separate, the purification cost of leachate is high, and the indium recovery rate is low are solved.

Description

TECHNICAL FIELD

[0001] The application relates to a method for recovering valuable metals from lead-zinc liquid slag, in particular to a method for efficiently separating and enriching zinc and indium from lead-zinc liquid slag, and belongs to the technical field of metallurgical solid waste resource utilization. BACKGROUND

[0002] In the current high-speed economic development, rich rare and precious metals play a key role in many emerging industries such as electronics, new energy, aerospace, etc., and their demand is showing explosive growth. However, the independent mineralization of rich rare and precious metals is extremely rare in nature, and is mostly associated with other metals such as lead and zinc in complex ore systems. At present, with the long-term high-intensity mining of high-quality lead-zinc ore, resources are gradually exhausted, and the proportion of lean ore and refractory ore is increasing. Taking the common associated rare and precious metal indium in lead-zinc ore as an example, it is indispensable in flat panel displays, semiconductors, solar cells, etc., and with the rise of the 5G industry and the new energy vehicle industry, the demand for rare and precious metals has risen sharply. However, the content of rare and precious metals in lead-zinc ore is extremely low, and traditional mining and beneficiation techniques cannot achieve efficient enrichment and recovery.

[0003] The liquid slag produced in the lead-zinc smelting process is rich in zinc, indium and other elements. However, the existing methods for extracting zinc and indium from liquid slag have significant bottlenecks: on the one hand, the slag is water quenched or cooled into a solid state, and then ground and leached under normal pressure. This process is high in energy consumption, slow in reaction rate, and indium and other rare metals are often wrapped in a silicon-iron network structure due to their complex occurrence, resulting in an overall leaching efficiency of less than 80%. On the other hand, the existing technology generally uses one-step high-acid leaching process, which can improve the metal leaching rate, but leads to a large amount of impurities such as silicon, iron and aluminum entering the solution, resulting in a very long and complex solution purification process. It must rely on multiple neutralization precipitation and a large amount of activated carbon adsorption, which not only consumes a large amount of reagents and costs a lot, but also produces a large amount of intermediate sludge that is difficult to handle, and the regeneration and disposal of activated carbon itself is also a major environmental burden. Thirdly, zinc and indium enter the solution together at the beginning of leaching, and there is a lack of effective pre-enrichment and separation means, resulting in a large processing capacity, poor selectivity and high recovery cost in the subsequent extraction process. Therefore, the existing technology cannot achieve efficient and collaborative recovery of zinc and indium from liquid slag under the premise of economic and environmental protection, which is of great significance.

[0004] In the prior art, there are documents ("New process for efficient extraction of indium from zinc oxide dust", Wang Dawei, etc., Non-ferrous metals (smelting part), 2015, 11:25-31) disclose the use of "neutral leaching-acid leaching-solvent extraction" process from indium-containing zinc oxide dust to extract indium, steps: 1) the temperature of neutral leaching is 60℃, the liquid-solid ratio is 10:1, the sulfuric acid acidity is 0.16mol / L, and the leaching time is 30min, the zinc leaching rate reaches 94.7%, and the indium is basically not leached. 2) The optimal conditions for acid leaching are: leaching temperature 70℃, liquid-solid ratio 6:1, sulfuric acid acidity 0.8mol / L, leaching time 2.0h, under the optimal conditions, the indium leaching rate reaches 91.6%. 3) The optimal conditions for solvent extraction of indium-containing leaching solution are: organic phase composition 20% P204+80% sulfonated kerosene, phase ratio A / O=4, initial acidity (H + concentration) 1.0~1.5mol / L, the single-stage extraction rate of indium can reach more than 90%. This method can realize high-efficiency leaching and extraction of indium in zinc oxide dust and separation from zinc. However, the liquid slag produced in the lead-zinc smelting process and the indium-containing zinc oxide dust are completely different in composition, the metal elements in the liquid slag are complexly embedded in the silicon-iron network structure, and the conventional leaching method leads to a leaching efficiency generally lower than 80%. SUMMARY

[0005] In view of the technical problems of difficult separation of zinc and indium and high purification cost of leaching solution in the prior art method for recovering zinc and indium from lead-zinc liquid slag, the purpose of the present application is to provide a method for efficiently separating and enriching zinc and indium from lead-zinc liquid slag. The method can selectively leach zinc and indium from lead-zinc liquid slag in two steps, that is, the efficient separation of zinc and indium is realized in the leaching process, which is beneficial to the recovery of high-purity zinc and indium, greatly simplifies (even eliminates) the deep purification steps such as activated carbon adsorption in the traditional process, significantly reduces the reagent consumption and solid waste production, achieves cost reduction and efficiency improvement, and fully utilizes the physical sensible heat and chemical activity of lead-zinc liquid slag to realize high leaching rate (both >95%) of zinc and indium, improve the resource recovery efficiency and economic benefit of the whole process. In summary, the method has the characteristics of short process, low energy consumption, environmental friendliness, good separation and recovery effect of zinc and indium, etc.

[0006] In order to achieve the above technical purpose, the present application provides a method for efficiently separating and enriching zinc and indium from lead-zinc liquid slag, which comprises the following steps:

[0007] 1) adding a sulfuric acid solution with a concentration of 100~200mL into a reaction kettle containing lead-zinc liquid slag, and carrying out leaching under the conditions of maintaining the temperature at 70~80℃ and the pressure at 0.5~1.0 MPa , the H2SO4 concentration of the sulfuric acid solution is controlled to be 90-110 g / L, and a zinc sulfate leaching solution and an indium-enriched leaching residue are obtained;

[0008] 2) the zinc sulfate leaching solution is mixed with elemental iron to carry out a reduction reaction and / or a displacement reaction to precipitate impurity metal ions, and a purified zinc sulfate solution is obtained, and the purified zinc sulfate solution is used to produce cathode zinc by electrolysis;

[0009] 3) the indium-enriched leaching residue is placed in a reaction kettle with a sulfuric acid solution, the concentration of the sulfuric acid solution is controlled to be 150-200 g / L, and leaching is carried out under the conditions of a temperature of 80-90 DEG C and a pressure of 0.5-1.0 MPa , and an indium-containing leaching solution and a leaching residue are obtained;

[0010] 4) the indium-containing leaching solution is neutralized to pH=2.0-3.0, and extraction is carried out on the indium-containing leaching solution with a P204-containing organic phase, and the indium-loaded organic phase is back-extracted with hydrochloric acid, and an indium-rich solution is obtained;

[0011] 5) the indium-rich solution is subjected to a displacement reaction with elemental zinc, and sponge indium is obtained.

[0012] The method for efficiently separating and enriching zinc and indium from a lead-zinc liquid-state slag provided by the application is characterized in that hot and high-activity lead-zinc liquid-state slag is directly used as raw material, a two-step continuous leaching process is designed, and the lead-zinc liquid-state slag is sequentially subjected to high-selectivity leaching of zinc and indium, that is, efficient separation of zinc and indium is realized in the leaching process, which is beneficial to subsequent purification and recovery of zinc and indium. More specifically, under optimized leaching conditions, the hot and high-activity lead-zinc liquid-state slag can realize high-selectivity leaching of zinc to realize efficient separation of zinc and indium, wherein the leaching rate of zinc in the obtained zinc-containing leaching solution can reach more than 95%, and the enrichment rate of indium in the obtained indium-enriched leaching residue can reach more than 94%. Since the impurity content in the zinc-containing leaching solution is low, only simple displacement or reduction of impurities (high-valence impurity ions such as Cu 2+ , As 3+ / As 5+ , etc.) by elemental iron is needed, the traditional leaching solution needs to be neutralized to remove iron, activated carbon adsorption and other complex processes are avoided, and a pure zinc sulfate solution meeting the requirements of electrodeposition is obtained, which is beneficial to obtaining high-purity cathode zinc and realizing efficient recovery of zinc. The indium-enriched leaching residue is further subjected to high-acid pressure leaching, and under the optimized conditions, the indium in the indium-enriched leaching residue can be completely and efficiently leached, and the leaching rate of indium can reach more than 98%, and the final leaching residue mainly comprises harmless waste residues such as calcium silicate and calcium ferrite. The obtained indium-containing leaching solution is further enriched by extraction-back extraction, and high-purity sponge indium is obtained by displacement precipitation, and efficient recovery of indium is realized.

[0013] As a preferred scheme, the leaching The leaching time is 1 to 2 hours. Under the preferred leaching conditions, a high zinc leaching rate can be guaranteed.

[0014] As a preferred embodiment, the amount of elemental iron used is 1-3% of the mass of the zinc sulfate leaching solution. Elemental iron, such as iron powder, is mainly used to remove high-valence impurity metal ions (such as Cu) from the zinc sulfate leaching solution. 2+ As 3+ / As 5+ It can be removed by reduction precipitation or by displacement reaction. If the amount of elemental iron added is too low, the impurity removal effect will not be achieved; if the amount of elemental iron added is too high, too many iron ions will be introduced.

[0015] As a preferred embodiment, the reduction and / or displacement reactions are performed under the following conditions: a temperature of 50-60°C and a time of 30-60 minutes. These preferred reaction conditions improve the efficiency of the reduction and displacement reactions and accelerate the precipitation of impurity metal ions.

[0016] As a preferred embodiment, the leaching The leaching time is 2-3 hours. Under the preferred leaching conditions, a high indium leaching rate can be guaranteed.

[0017] As a preferred embodiment, the extraction employs a multi-stage extraction method, with the following extraction conditions: a phase volume ratio of O / A of 1~5:1~5, an extraction temperature of 20~30℃, a P2O4 volume concentration of 10~30% in the organic phase, and 2~5 extraction stages. The organic phase also contains a diluent, such as sulfonated kerosene. Using P2O4 results in high extraction efficiency for indium ions, which is beneficial for indium ion recovery.

[0018] As a preferred embodiment, the back-extraction employs a multi-stage back-extraction method. The extraction conditions are: a phase volume ratio of O / A of 1~5:1~5, an extraction temperature of 20~30℃, a back-extraction agent of 4~6 mol / L hydrochloric acid solution, and 1~3 back-extraction stages. This invention, through extraction and back-extraction, achieves highly efficient enrichment of indium-containing leachate. The indium concentration can be increased by more than 10 times through phase adjustment, while the purity is also significantly improved.

[0019] As a preferred embodiment, the amount of elemental zinc used is 110-130% of the molar amount of indium in the indium-rich solution. Using an appropriate excess of elemental zinc is beneficial for achieving deep conversion of indium ions and reducing the residual amount of indium ions in the indium-rich solution. The preferred time for replacing indium ions with elemental zinc is 1-2 hours.

[0020] The indium-containing leaching solution of the present invention uses an alkaline solution (such as lime milk or calcium carbonate) to neutralize to pH=2.0~3.0. On the one hand, it can hydrolyze and precipitate some impurities, and on the other hand, it can neutralize some acids, providing a favorable pH environment for the extraction of P204 (di-(2-ethylhexyl)phosphoric acid).

[0021] The sponge indium obtained by this invention is washed with deionized water, compressed into pellets, and then melted and cast into crude indium ingots at 250~350℃ under an inert atmosphere (such as argon).

[0022] Compared with existing technologies, the beneficial technical effects of the present invention are as follows:

[0023] This invention represents a significant technological advancement by greatly improving the recovery efficiency of zinc and indium in molten lead-zinc slag, while simplifying the process, reducing production costs, and lowering energy consumption. Specifically: 1) This invention significantly increases the recovery rates of zinc and indium in molten lead-zinc slag from 85% and 75% respectively (see Comparative Example 1 for the traditional method) to over 96% and 97%, achieving highly efficient recovery of valuable metals. 2) By designing a "two-stage pressure leaching" process, this invention achieves efficient separation of zinc and indium during the leaching process, solving the problem of high impurity content in the leaching solution at its source and avoiding the neutralization and activated carbon adsorption processes required in traditional methods. 3) This invention completely eliminates the use of activated carbon and reduces neutralizing agent consumption by over 70%, lowering reagent costs by approximately 60%. Simultaneously, due to the simplified purification process and an 80% reduction in solid waste generation, the overall operating cost is reduced by approximately 25-30% compared to traditional processes. 4) This invention directly utilizes the sensible heat of molten lead-zinc slag, eliminating the need for water quenching, crushing, and reheating of the slag in traditional processes. This reduces overall energy consumption by approximately 20-25%, equivalent to a reduction of 8,000-10,000 tons of carbon dioxide emissions annually. In summary, this invention not only improves the metal recovery rate from molten lead-zinc slag but also achieves a dual reduction in production costs and energy consumption. Detailed Implementation

[0024] The following specific embodiments are intended to further illustrate the content of the present invention, rather than to limit the scope of protection of the claims of the present invention.

[0025] Example 1

[0026] The molten slag produced by the oxygen-enriched side-blown smelting process of a lead-zinc smelter was used as the treatment object. Its chemical composition was analyzed as follows: Zn 18.5%, In 0.15%, Fe 22%, SiO 25%, Pb 3.5%, and the remainder being CaO, Al2O3, etc.

[0027] 1) Neutral sulfuric acid pressure leaching of zinc: Approximately 1000 kg of the high-temperature liquid slag (temperature approximately 1100℃) was weighed and directly introduced into a 2 m³ titanium pressure reactor via a chute. A pre-prepared 150 g / L dilute sulfuric acid solution was immediately injected, and cooling water and steam were introduced through the jacket to precisely control the reaction system temperature at 75 ± 2℃. The reactor was sealed, stirring was started, and compressed air was introduced to maintain the system pressure at 0.7 MPa. During the reaction, the pH was monitored online using an acidity meter, and the pH was controlled below 0.7. The rate of sulfuric acid solution addition was controlled to ensure that the final free sulfuric acid concentration was 100 g / L after 1.5 hours of reaction. After the reaction, hot pressure filtration was performed using a bottom filtration device. Approximately 750 L of clear neutral leachate was obtained. ICP analysis showed the following composition: Zn 48.5 g / L, In 0.008 g / L, Fe < 0.5 g / L. The calculated leaching rate of zinc reached 96.5%, and the inhibition enrichment rate of indium (the proportion remaining in the slag) was as high as 94.5%. Approximately 220 kg of neutral leaching residue was obtained (dry basis).

[0028] 2) Neutral Leachate Purification and Electrowinning: All the above-mentioned neutral leachate was transferred to a PPH material purification tank and heated to 55℃. Then, 2.0% of the total mass of high-quality reduced iron powder (200 mesh) was slowly added, and the mixture was mechanically stirred at 120 rpm for 45 minutes. After the reaction was complete, the solution was filtered to obtain a pure zinc sulfate solution. Atomic absorption spectroscopy analysis of this solution showed that the impurity contents of Cu, As, and Cd (Cu 0.1 mg / L, As 0.04 mg / L, Cd 0.2 mg / L) were all far below the requirements for electrowinning. Subsequently, this purified solution was directly sent to a standard electrowinning workshop, where, under a current density of 450 A / m², standard 0# grade cathode zinc was stably produced.

[0029] 2) High-acid pressure leaching enrichment of indium: The aforementioned 220 kg of neutral leaching residue was loaded into another pressure reactor. A pre-prepared sulfuric acid solution with a concentration of 180 g / L was added, controlling the liquid-to-solid ratio at 4 L: 1 kg. Leaching was carried out with stirring for 2.5 hours at 85℃ and 0.8 MPa. After the reaction was completed, the solution was filtered to obtain approximately 680 L of high-acid leachate, with the following composition: In 0.49 g / L, Zn 9.2 g / L, Fe 5.8 g / L. The leaching rate of indium in this step was calculated to be 98.5%.

[0030] 3) Indium Extraction and Back-Extraction: The high-acid leachate was introduced into a neutralization tank and slowly neutralized to pH 2.5 with lime slurry. After settling, the solution was filtered to remove small amounts of gypsum and ferric hydroxide. The clarified solution entered a three-stage countercurrent extraction system. The organic phase was a mixture of 20% (v / v) P2O4 and 80% sulfonated kerosene. The organic-to-aqueous phase ratio (O / A) was controlled at 1:1, and extraction was carried out at 25°C. The indium-loaded organic phase flowed into the back-extraction section and was back-extracted with 5 mol / L hydrochloric acid, controlling the O / A ratio at 3:1, to obtain a pure indium-rich solution with an indium concentration enriched to 14.8 g / L.

[0031] 4) Indium displacement and casting: The indium-rich liquid was transferred to a displacement tank, and 1.3 times the theoretical amount of high-purity zinc flakes (99.99%) were added. The mixture was mechanically stirred and reacted for 1.5 hours. After the reaction, the mixture was allowed to stand, the supernatant was discarded, and the deposited sponge indium was collected. The sponge indium was repeatedly washed with deionized water until neutral, and then pressed into dense indium lumps using a hydraulic press. Finally, the indium lumps were placed in a graphite crucible and cast at 300°C under high-purity argon protection to obtain crude indium ingots. The crude indium ingots weighed approximately 1.45 kg, and their purity was determined to be 99.35% by ICP-MS.

[0032] This embodiment starts with liquid slag and calculates the total metal recovery rate of the entire process: 96.0% for zinc and 96.9% for indium. Furthermore, no activated carbon was used in the entire hydrometallurgical stage.

[0033] Example 2

[0034] The liquid slag produced by the oxygen-enriched side-blown smelting process of a lead-zinc smelter was used as the treatment object. Its characteristics are that the indium content is relatively high: Zn 16.0%, In 0.28%, Fe 21%, SiO 26%, CaO 10%.

[0035] 1) Neutral sulfuric acid pressure leaching of zinc: Refer to Example 1, except that 800 kg of liquid slag is treated and the parameters are adjusted as follows: the leaching temperature is increased to 78°C, the pressure is maintained at 0.6 MPa, the reaction time is extended to 2 hours, the final free sulfuric acid concentration is strictly controlled at 95 g / L, and the sulfuric acid solution concentration is 150 g / L.

[0036] Results: The zinc leaching rate was 94.8%, and the indium inhibition enrichment rate reached 90.2%. This demonstrates that the selective leaching conditions remain effective for materials with high indium content.

[0037] 2) Neutral leachate purification and electrowinning: Refer to Example 1, except that the parameter is adjusted to: 50 minutes of reduction.

[0038] Result: The purified liquid successfully produced qualified electrolytic zinc.

[0039] 3) High acid pressure leaching enrichment of indium: Refer to Example 1, except that the parameters are adjusted as follows: the sulfuric acid solution concentration is 190 g / L, and leaching is carried out at 88°C for 2 hours.

[0040] Result: The leaching rate of indium reached 97.8%.

[0041] 4) Extraction and back-extraction of indium: Refer to Example 1.

[0042] Results: Approximately 2.2 kg of crude indium ingots were obtained, with a purity of 99.28%.

[0043] In this embodiment, the final calculation showed that the total recovery rate of zinc was 94.2% and the total recovery rate of indium was 96.5%. The results indicate that the present invention also has excellent adaptability and stability for processing high-indium liquid slag, and the key metal recovery indicators remain at a high level.

[0044] Comparative Example 1

[0045] 1) Neutral sulfuric acid pressure leaching of zinc: Compared with Example 1, the difference is that the concentration of free sulfuric acid at the endpoint is strictly controlled at 60 g / L.

[0046] Results: The zinc leaching rate was significantly reduced to 81.5%, while the indium inhibition enrichment rate was 92.0%. The results indicate that excessively low final acid concentration severely weakened the zinc leaching efficiency.

[0047] 2) Neutral leachate purification and electrowinning: See Example 1.

[0048] Result: Qualified electrolytic zinc can still be produced.

[0049] 3) High-acid pressure leaching enrichment of indium: Refer to Example 1, except that the parameters were adjusted to: sulfuric acid concentration of 190 g / L, 88°C for 2 hours. Due to the high residual zinc content in the neutral leaching residue, additional acid was consumed, and the indium leaching rate was affected, reaching 95.2%.

[0050] 4) Extraction and back-extraction of indium: Parameter Example 1

[0051] 5) Indium displacement and casting: See Example 1.

[0052] Results: In the final calculation of this comparative example, due to insufficient leaching in the first stage and subsequent chain reactions, the total recovery rate of zinc dropped to 80.1%, and the total recovery rate of indium was 90.3%. Although indium enrichment can still be partially achieved, the recovery rate of the main metal zinc has decreased significantly, resulting in slightly worse overall economic and technical indicators.

[0053] Comparative Example 2

[0054] 1) Zinc leaching under pressure with neutral sulfuric acid: The difference compared to Example 1 is that leaching is carried out under normal pressure.

[0055] Results: Due to the lack of pressure-enhanced mass transfer, the reaction rate was slow. After 1.5 hours, the final concentration of free sulfuric acid in the system was still as high as approximately 130 g / L, indicating slow acid consumption and incomplete reaction. Leaching effect: Analysis of the leachate showed that the zinc leaching rate was 87.2%, lower than the 96.5% under pressurized conditions. Meanwhile, because the actual acidity was higher than the target value, the leaching inhibition effect on indium was worse, with the indium concentration in the leachate reaching 0.12 g / L, and the indium inhibition enrichment rate decreasing to approximately 85%.

[0056] 2)~5) See Example 1.

[0057] Comparative Example 3

[0058] 1) Neutral sulfuric acid pressure leaching of zinc: The difference compared to Example 1 is that the leaching temperature is controlled at 65°C.

[0059] Results: Low temperature reduced the reaction rate. Although the final acidity could be controlled at around 105 g / L, the zinc leaching rate dropped significantly to 89.5%. Furthermore, the indium leaching rate was also low at low temperatures (In in the leachate < 0.005 g / L), indicating a high inhibition rate.

[0060] 2)~5) See Example 1.

[0061] Comparative Example 4

[0062] 1) Neutral sulfuric acid pressure leaching of zinc: Compared with Example 1, the difference is that 1000 kg of high-temperature liquid slag with the same composition is first water quenched and dried, then crushed and ball-milled to 200 mesh with more than 90% content to obtain solid slag powder, which is used to replace the hot molten slag in Comparative Document 1.

[0063] Results: This process completely lost the sensible heat of the slag and added high-energy-consuming steps such as water quenching, crushing, and grinding. Leaching effect: Even under the same chemical conditions as in Example 1, due to the dense crystal structure of the solid slag and its much lower reactivity than the amorphous liquid slag, the zinc leaching rate was 90.1%, and the indium concentration in the leachate (0.25 g / L) was significantly higher than in Example 1, resulting in a decrease in separation effect.

[0064] 2)~5) See Example 1.

[0065] Comparative Example 5

[0066] 1) Neutral sulfuric acid pressure leaching of zinc: Compared with Example 1, the difference is that the concentration of free sulfuric acid at the leaching endpoint is controlled at 150 g / L.

[0067] Results: The zinc concentration in the leachate reached 48.8 g / L (leaching rate 96.8%), but the indium concentration increased sharply to 0.68 g / L, and the iron concentration also rose to 9.2 g / L. This means that more than 85% of the indium and zinc were co-leached, and the "selective separation" effect of the first step was completely ineffective.

[0068] 2)~5) See Example 1.

[0069] However, due to the high concentration of iron and indium in the leachate, the simple "iron powder reduction purification method" is not applicable. It is necessary to reintroduce complex steps such as neutralization and iron removal and activated carbon adsorption in the traditional process, which increases costs.

[0070] Indium recovery process disorder: Indium is dispersed in neutral leaching solution and subsequent high acid leaching solution, making the recovery process complicated and uneconomical, and the overall recovery rate is reduced due to dispersion loss.

[0071] Comparative Example 6

[0072] The same lead-zinc molten slag as in Example 1 was used as the raw material, but the conventional "one-step high-acid atmospheric pressure leaching-deep purification" process was employed. The raw material composition was the same as in Example 1: Zn 18.5%, In 0.15%, Fe 22%, SiO 25%.

[0073] Slag pretreatment and grinding: First, 1000 kg of high-temperature liquid slag is quenched in water to obtain loose and porous solid water-quenched slag. Then, the water-quenched slag is crushed and ball-milled until more than 90% of the particles are smaller than 0.074 mm (-200 mesh) for leaching.

[0074] One-step high-acid atmospheric pressure leaching: All finely ground mineral powder was added to an atmospheric pressure leaching tank. A sulfuric acid solution with a concentration of 180 g / L was added, controlling the liquid-to-solid ratio at 5:1. Leaching was carried out at 90℃ with mechanical stirring for 3 hours. After the reaction, the mixture was filtered to obtain a complex high-acid leachate and leaching residue. Leachate analysis showed: Zn 42.5 g / L, In 0.14 g / L, Fe 25.8 g / L, and H₂SO₄ balance approximately 45 g / L. The calculated leaching rate of zinc was 92.0%, and the leaching rate of indium was 78.5%. Simultaneously, a large amount of impurities such as iron, silicon, and aluminum were co-leached.

[0075] Neutralization and purification: The resulting leachate has a high impurity content and must undergo multi-stage purification. First, lime milk is added to the hot leachate to neutralize it to pH 4.0-4.5, and air is bubbled in to remove Fe from the solution. 2+ Oxidized to Fe 3+The sludge is then hydrolyzed and precipitated into ferric hydroxide slag. This process generates a large amount of hazardous sludge containing iron and arsenic (approximately 35% of the total slag). After neutralization, the liquid is filtered, and the filtrate still contains some iron, copper, cadmium, cobalt, and organic impurities.

[0076] Deep purification by activated carbon adsorption: To obtain a qualified pre-electrowinning solution, the neutralized solution is heated to 70°C, and 0.2% (by weight) of powdered activated carbon is added. The mixture is stirred and adsorbed for 1 hour to remove organic impurities and some residual metal ions. The activated carbon is then separated by filtration.

[0077] Zinc electrowinning and indium recovery: The deeply purified solution is fed into the electrowinning system to produce cathode zinc, but due to the long process and metal dispersion, the total zinc recovery rate drops to 85.1%. For indium recovery, it is re-enriched from the complex neutralized liquid or waste liquid. Initial enrichment is carried out by sulfide precipitation, followed by acid dissolution and extraction (P204) processes, with a final total indium recovery rate of 74.6%.

Claims

1. A method for efficiently separating and enriching zinc and indium from lead-zinc molten slag, characterized in that: Includes the following steps: 1) Add a sulfuric acid solution with a concentration of 100~200g / L to a reactor containing molten lead-zinc slag, and carry out leaching under conditions of maintaining a temperature of 70~80℃ and a pressure of 0.5~1.0MPa. The addition rate of sulfuric acid solution was controlled to maintain the H2SO4 concentration at the leaching endpoint at 90~110 g / L, resulting in zinc sulfate leaching solution and indium-enriched leaching residue. 2) The zinc sulfate leaching solution is mixed with elemental iron to carry out a reduction reaction and / or a displacement reaction to precipitate impurity metal ions, thereby obtaining a purified zinc sulfate solution, which is then electrolyzed to produce cathode zinc. 3) Place the indium enrichment leaching residue and sulfuric acid solution in a reaction vessel, control the concentration of the sulfuric acid solution to be 150~200 g / L, and carry out leaching under the conditions of 80~90℃ and 0.5~1.0MPa. The indium-containing leachate and leachate residue were obtained. 4) Neutralize the indium-containing leachate to pH 2.0-3.0, extract the indium-containing leachate with an organic phase containing P2O4, and back-extract the indium-loaded organic phase with hydrochloric acid to obtain an indium-rich solution; 5) Displace indium-rich solution with elemental zinc to obtain sponge indium.

2. The method for efficiently separating and enriching zinc and indium from lead-zinc molten slag according to claim 1, characterized in that: The leaching The time is 1 to 2 hours.

3. The method for efficiently separating and enriching zinc and indium from lead-zinc molten slag according to claim 1, characterized in that: The amount of elemental iron used is 1-3% of the mass of the zinc sulfate leaching solution.

4. A method for efficiently separating and enriching zinc and indium from lead-zinc molten slag according to claim 1 or 3, characterized in that: The conditions for the reduction and / or displacement reactions are: a temperature of 50-60°C and a time of 30-60 minutes.

5. A method for efficiently separating and enriching zinc and indium from lead-zinc molten slag according to claim 1 or 3, characterized in that: The leaching The time is 2 to 3 hours.

6. The method for efficiently separating and enriching zinc and indium from lead-zinc molten slag according to claim 1, characterized in that: The extraction adopts a multi-stage extraction method, and the extraction conditions are: phase volume ratio O / A is 1~5:1~5, extraction temperature is 20~30℃, P2O4 volume concentration in organic phase is 10~30%, and the number of extraction stages is 2~5.

7. The method for efficiently separating and enriching zinc and indium from lead-zinc molten slag according to claim 1, characterized in that: The back-extraction adopts a multi-stage back-extraction method, and the extraction conditions are: phase volume ratio O / A is 1~5:1~5, extraction temperature is 20~30℃, back-extraction agent is 4~6mol / L hydrochloric acid solution, and the number of back-extraction stages is 1~3 stages.

8. The method for efficiently separating and enriching zinc and indium from lead-zinc molten slag according to claim 1, characterized in that: The amount of elemental zinc used is 110-130% of the molar amount of indium in the indium-rich solution.