CN102534227B - Method for extracting indium from indium-rich smoke dust by using oxygen pressure technology - Google Patents

Abstract

The invention provides a method for extracting indium from indium-rich smoke dust by using an oxygen pressure technology. The method is characterized by comprising the following steps of: carrying out leaching indium extraction of the difficult-to-handle indium-rich smoke dust of a Pb-Sn reverberatory furnace by adopting a wet-process metallurgical oxygen pressure acid leaching technology, and highly enriching and recycling valuable metals in the raw material; and purifying to remove impurities of leachate, and carrying out extraction, replacement and electrolytic refining, thereby obtaining greater than 99.995% of electrical indium products. The technology method has the advantages that the indium leaching rate of the indium-rich smoke dust of the Pb-Sb reverberatory furnace and the enrichment rate of the valuable metals can be obviously increased, and the comprehensive recycling effect can be achieved, thereby waster water of indium smelting can be discharged up to the standard after being processed with low cost, the environmental pollution in the indium extraction process is eliminated. In the process, the leaching rate of the indium is more than 97%, leaching residues contain less than 0.01% of indium, the enrichment rate of plumbum, tin, bismuth and zinc is more than 98%, the quality of refined indium products is more than 99.995%, and plumbum enriching slag containing greater than 60% of Pd is obtained. The technical scheme can independently form a system, can also be used for improving and perfecting old technologies and has higher popularization and application value.

Description

A method for extracting indium from indium-rich dust by oxygen pressure

technical field

The invention belongs to the field of hydrometallurgy of non-ferrous metals, and relates to a process for extracting metal indium by oxygen pressure technology, in particular to a method for extracting indium from indium-rich dust by oxygen pressure. the

Background technique

Indium is a rare element, its content is rare in the earth's crust and its distribution is very scattered, and no independent mineralization has been found in nature. Usually indium is mainly associated with heavy non-ferrous metals such as copper, lead, zinc, tin and other minerals, especially sphalerite, chalcopyrite, kesterite and galena; therefore, it is extracted and recovered from the heavy non-ferrous metal smelting process Indium has become an important source of indium. Due to the special physical and chemical properties of indium, it is widely used in the electronics and semiconductor industries, especially the ITO target industry has accounted for more than 70% of indium consumption, while indium is also used in the field of advanced alloy materials and military, medical, chemical, electrical appliances , glass and other industries, indium has become an indispensable and important symbolic material for the development of modern science and technology. the

With the development of science and technology and the impact of resource shortages, although related industries at home and abroad have increased the intensity of secondary recycling of indium-containing waste, the raw materials required for indium metallurgy production mainly come from copper, lead, zinc, and tin. Among the indium-containing by-product materials produced in the smelting process of heavy non-ferrous metals, especially indium-rich dust. The main processes for the current industrialized indium production are: wet acid leaching solvent extraction method, enriched slag vacuum distillation purification method, and chlorination volatilization method; however, due to the characteristics of raw materials in China, most of the indium extraction processes in China still use wet acid leaching extraction. The method is the main method, that is: the indium-containing material is subjected to acid leaching-removal of impurities-extraction-reverse extraction-replacement-pressed alkali fusion-crude indium-electrolytic refining-refined indium. Although this conventional process is mature and reliable, due to acid leaching The leaching rate of indium in the process is generally about 70-80%, and the leaching rate is relatively low; the indium content in the acid leaching slag basically fluctuates at 0.05-0.1%, resulting in the direct recovery rate of indium in the whole process is only about 60-80%. It has the characteristics of long process, large reagent consumption, large metal occupation and turnover, serious loss of other valuable metals, special acid mist absorption equipment required for the acid leaching process, and difficult process wastewater discharge and treatment. the

In recent years, relevant scientific research institutions and technicians have continuously optimized and improved the process on how to increase the indium acid leaching rate, such as: leaching the indium-rich fume after concentrated sulfuric acid aging treatment, the application of secondary acid leaching strengthening technology, The promotion of mixed acid leaching technology, etc. Although these measures have improved the technical indicators of the indium acid leaching process to a certain extent, they have also increased the pressure and production costs of production operations, technical control, system adjustment, environmental protection and wastewater treatment, and weakened Without the effect of technical improvement, it is difficult to achieve satisfactory comprehensive benefits. With the rapid development of metallurgical technology and the continuous enhancement and improvement of human environmental protection awareness and environmental protection requirements, active research and development of efficient, green and environmentally friendly indium extraction technology has become the development trend of the industry. After the improvement and promotion of oxygen pressure technology in recent years, breakthrough progress has been made, the scope of application is wider, and the operation technology is becoming more and more mature, especially for the technological update and breakthrough in the field of non-ferrous hydrometallurgy. It not only strengthens the mass transfer effect of the metallurgical reaction process, improves the reaction efficiency, speeds up the reaction process, but also significantly improves the technical indicators and comprehensive processing capacity, enhances the recycling level of valuable metals, and eliminates process pollution Emissions reduce production costs and energy consumption, which is in line with the high-efficiency, environmental protection, and green development direction of nonferrous metallurgy. Therefore, the application of oxygen pressure acid leaching technology to the indium metallurgy process will have extremely far-reaching significance. the

Contents of the invention

In order to solve the problem of low yield of indium extraction from the indium-rich dust produced in the smelting process of heavy nonferrous metals, the present invention provides a method for extracting indium from the indium-rich dust by oxygen pressure, which uses oxygen pressure technology to extract indium by acid leaching , the process operation is simple, practical, efficient and clean, which not only significantly improves the leaching recovery rate of indium in the indium-rich fume, but also solves the problems of complex control of the leaching process, low enrichment rate of valuable metals, and large loss in conventional processes. The harm of acid mist to the environment is completely eliminated, and the goals of resource synthesis, recycling, high-efficiency utilization and green metallurgy are achieved. the

To achieve the above object, the present invention is achieved through the following technical solutions:

A method for extracting indium under oxygen pressure from indium-rich smoke dust, has the following specific steps:

A. Oxygen pressure acid leaching

Crush the indium-rich smoke and dust to a particle size of 100-120 mesh, and add it to the pre-configured dilute sulfuric acid solution with a mass percentage concentration of 10%-30% according to the liquid-solid ratio of 4-10:1, wherein the unit of the liquid-solid ratio is kg/ L; In the closed high-pressure reactor, introduce industrial oxygen; control the reaction leaching temperature to 120-180°C, the industrial oxygen pressure to 0.6-1.5MPa, and the reaction time to 2-5 hours; after the reaction is completed, the reactor is opened after cooling and pressure relief. Separating the reacted slurry from liquid to solid to obtain separation slag and oxygen pressure acid leaching filtrate; the separation slag is washed with hot water for 1 to 2 times and sent to lead smelting for comprehensive recovery of lead, tin and silver;

B. Reduction and removal of impurities

Put the oxygen pressure acid leaching filtrate into the reaction tank, add fresh finely crushed iron filings in an amount of 1.0-5.0g/L, stir and react at a temperature of 40-50°C for 1-3 hours, and then carry out clarification and separation; take the supernatant Calculated based on the total content of tin, bismuth, lead, iron and arsenic in the solution, add Na ₂ S or other sulfides 0.7 to 1.5 times the theoretical reaction amount, stir and react at a temperature of 50 to 60°C for 1 to 2 hours, then filter. Obtain impurity removal liquid and filter residue; the filter residue is washed and sent back to lead smelting for comprehensive recovery;

C. Extraction and refining of indium

①Extraction

After re-clarifying the impurity-removing liquid in step B for 8-10 hours, extract it at room temperature with a volume mixing ratio of 30% P204+70% sulfonated kerosene extraction system. The process control conditions are: four-stage extraction: compared to O/A =1: 10～15, mixing time 3～5min; three-stage washing: first two-stage acid washing with 80-100g/L H ₂ SO ₄ , and then one-stage water washing, compared to O/A=(2～ 5): 1, the separation time is 5 minutes; four-stage stripping: compared to O/A=(3-10): 1, wherein the stripping agent uses 6-8mol/L HCl solution, and the mixing time is 3-5 minutes;

② replacement

Use NaOH to adjust the pH value of the stripping solution under ventilated conditions, and control the pH of the solution to 1.5-2.0. After the solution is clarified and filtered, use hydrochloric acid to adjust the pH of the solution to 1.0, and then add fresh zinc that is equivalent to 5-8 times the weight of indium in the solution. The replacement is carried out with sheets, and the reaction time is 24-72 hours. When the indium content in the replacement solution is less than 0.001g/L; separate the sponge indium and the replacement solution; send the replacement solution to recover zinc products;

③Sponge indium alkali fusion

After washing and pressing the separated indium sponge, it is melted at 200-300°C under the cover of caustic soda, and the scum is removed to obtain crude indium with a content of more than 99%; the scum is returned to step A for oxygen pressure acid leaching Process cyclic acid leaching;

④ Electrolytic refining

Thick indium is melted and cast into a thick indium anode, and electrolysis is carried out under the condition of direct current. The electrolysis control conditions are: electrolysis temperature: 25-30°C, current density: 50-80A/m ² , cell voltage: 0.2-0.3V, electrolysis cycle : 5 to 7 days; the refined indium obtained by electrolysis is cast into an ingot after the glycerol iodination method removes impurities in a molten state, and more than 99.995% of the refined indium product is obtained;

D. Raffinate treatment

The raffinate produced in the indium extraction process is subjected to gas-phase oil separation and secondary clarification oil separation, and then is discharged by conventional sulfidation precipitation and lime-iron salt wastewater treatment, and the sulfide precipitation residue and waste water filter residue are sent to the reverberatory furnace for batching and recycling. the

The above-mentioned indium-rich dust is the enriched slag of high-indium-containing materials or other indium-containing products produced by volatilization furnaces and kilns in the metallurgical process of heavy metals copper, lead, zinc, and tin. the

The above-mentioned fresh and finely crushed iron filings are iron filings that have been produced, processed and stored for no more than 7 days. the

The above-mentioned fresh zinc flakes are 0# zinc flakes whose production, processing and storage time does not exceed 7 days, such as the Torch brand 0# zinc flakes produced by Zhuzhou Smelter. the

The H ₂ SO ₄ , NaOH, Na ₂ S, P204, sulfonated kerosene, industrial oxygen, and iron filings used in the above reaction process are all commercially available industrial grade products. Here P204 is the chemical name of the extractant: two (2-ethylhexyl) phosphoric acid; here the sulfonated kerosene is the diluent, also known as 200 ^#kerosene .

The high-pressure reaction equipment used in the above-mentioned oxygen pressure acid leaching process all meet the relevant national regulations on special pressure equipment, have acid and alkali resistance, anti-corrosion measures and meet the requirements of technical control conditions, and can meet the requirements of safety, standardization and skilled operation. the

Main reaction among the present invention is:

Oxygen pressure acid leaching:

In ₂ O ₃ +3H ₂ SO ₄ ＝In ₂ (SO ₄ ) ₃ +3H ₂ O (1)

2In+3H ₂ SO ₄ ＝In ₂ (SO ₄ ) ₃ +3H ₂ (2)

2In ₂ S ₃ +6H ₂ SO ₄ +3O ₂ ＝In ₂ (SO ₄ ) ₃ +6S+6H ₂ O (3)

SnO ₂ +2H ₂ SO ₄ ＝Sn(SO ₄ ) ₂ +4H ₂ O (4)

SnO+H ₂ SO ₄ ＝SnSO ₄ +H ₂ O (5)

2Sn ²⁺ +4H ₂ O+O ₂ →2H ₂ SnO ₃ ↓+4H ⁺ (6)

ZnO·Fe ₂ O ₃ +4H ₂ SO ₄ ＝ZnSO ₄ +Fe ₂ (SO ₄ ) ₃ +4H ₂ O (7)

MeO+H ₂ SO ₄ ＝MeSO ₄ +H ₂ O (8)

Restoration and removal of impurities:

Me ₁ ²⁺ +S ^2- ＝Me ₁ S↓ (9)

2In ³⁺ +3Zn＝2In+3Zn ²⁺ (10)

In ³⁺ +3e＝In (11)

The aforementioned Me—Zn, Pb, Cd, Cu, Bi, As, etc.; Me ₁ -Fe, Zn, Pb, Cd, Cu, Bi, Sn, As, etc.

Extraction reaction:

In ³⁺ (a)+3[H ₂ A ₂ ](o)＝[InA ₃ ·3HA](o)+3H ⁺ (a) (12)

Stripping reaction:

[InA ₃ ·3HA](o)+4HCl(a)=HInCl ₄ (a)+3[H ₂ A ₂ ](o) (13)

In the process of operation, the method of the present invention can not only solve the industry problems of low leaching rate of indium-containing smoke and dust from lead and tin reverberatory furnaces and extremely complicated material handling, but also significantly increase the value of heavy metals such as lead, tin and bismuth. The enrichment rate of elements into slag recycling reduces metal loss and environmental hazards, so that indium smelting wastewater can be treated at low cost and then discharged up to standard, eliminating people's concerns about the ineffective treatment of indium metallurgical process wastewater. In the process, the leaching rate of indium is over 96.5%, the leaching slag contains less than 0.01% indium, and the enrichment rate of lead, tin, and bismuth reaches over 97%. In the whole process, the recovery rate of indium reaches over 97%, and the raffinate contains less than 0.1mg/L, can obtain refined indium products containing more than 99.995% of In and lead-tin dross containing more than 55% of Pb and more than 30% of Sn. Therefore, the method of the present invention has the advantages of advanced technology, short flow process, strong adaptability to raw materials, high indium direct recovery rate, low energy consumption, good comprehensive recovery benefits, and process waste residue resources can be recycled and recycled, and waste water is easy to treat and discharge up to the standard. Advantages, it is very suitable for the technical transformation of old processes and process optimization, and has good promotion value. the

Description of drawings

Fig. 1 is a process flow diagram of the present invention. the

Detailed ways

The present invention will be further described below in conjunction with the accompanying drawings and individual embodiments. the

Example 1

Lead electrolysis scum enriched indium fume composition (wt, %): In 11.44, Zn 2.06, Pb24.86, Fe 0.33, As 0.89, Sb 1.45. the

Take 150Kg of the above-mentioned raw materials, prepare samples and pulverize them to a particle size of 120 mesh, add a solution of _H2SO4 with a concentration of 10% by mass percentage according to the reaction liquid-solid ratio of 7:1 to adjust the _slurry , and put it into a closed high-pressure reactor, and control the conditions It is: reaction temperature 160°C, reaction oxygen pressure 0.6MPa, reaction time 4.0h; after the reaction, the reactor is cooled and opened safely after pressure release, the slurry is separated from liquid to solid, and the filter residue is washed once with hot water at 80°C before returning to lead The recovery system is processed, and the filtrate is sent for impurity removal and extraction. The filter residue was sampled and analyzed: 0.014% indium, 68.18% lead, 2.63% tin; filtrate: 0.426g/L indium, 0.020g/L lead, 3.35g/L zinc. In the leaching reaction, the In leaching rate is 97.31%, the Sn leaching rate is 0.62%, the Pb leaching rate is 0.51%, and the Zn leaching rate is 99.2%.

Add reduced iron powder (or fresh iron filings) at 3.5 g/L to the above filtrate at a temperature of 50°C, stir for 3 hours, then clarify and filter, and add Na ₂ S, stirring and reacting for 80 minutes, adding an appropriate amount of flocculant for filtration, and the filtrate was naturally clarified for 10 hours before being sent for extraction; the sulfide slag was washed and returned to lead smelting for recovery, and the slag contained less than 0.013% indium.

The indium-containing liquid after impurity removal is extracted with 30% P204+70% sulfonated kerosene extraction system, the control conditions are: extraction temperature is room temperature, 4-stage extraction O/A=1:15, 3-stage washing O/A=5 : 1, 4-stage back extraction O/A=7: 1, two-phase mixing time 5min, two-phase clarification separation time 5min; after the extraction, adjust the pH value of the back-extraction solution to 1.5 with NaOH, further separate impurities, and the filter residue returns to acid In the immersion process, add acid to the filtrate to adjust the pH to about 1.0, and then add fresh zinc flakes 6 times the theoretical amount for replacement. The reaction time is 24 hours at room temperature. After the replacement, the indium sponge is washed, compacted, alkali-fused, and electrolytically refined using conventional indium electrolysis conditions to obtain a product with an indium content of 99.998%. the

After indium replacement, the indium content in the liquid is less than 0.0008g/L, which can be sold or recycled to produce zinc products; the indium content in the raffinate in the indium extraction process is less than 0.09mg/L. , the conventional sulfide-lime wastewater treatment process can be used to meet the national discharge standards, and the wastewater treatment waste residue is returned to lead smelting ingredients for utilization. the

During the operation of this process, the indium leaching rate is greater than 97%, the direct recovery rate is greater than 93.2%, among which the recovery rate of purification and impurity removal is greater than 99.1%, the recovery rate of extraction is greater than 99.2%, the recovery rate of sponge indium replacement is greater than 97.9%, and the recovery rate of electrolysis is greater than 99.1%. Greater than 99.6%. And it can produce indium ingots that meet the national standard In99.995 or higher, and the enrichment rate of valuable metals is over 96.8%. the

Example 2

The raw material composition of indium-rich soot for lead-tin reverberatory furnace is (wt,%): In 0.287, Zn 1.36, Pb16.71, Fe 0.89, As 1.91, Sb 0.64, SiO ₂ 5.65, Cd 0.32, Sn 7.62.

Take 1Kg of the raw material that has been pulverized and prepared to a particle size of 120 mesh, add a solution of _H2SO4 with a mass percentage concentration of 30% according to the liquid-solid ratio of 4:1 to adjust the slurry, and put it into _a closed high-pressure reactor. The reaction conditions are controlled as follows: The temperature is 120°C, the reaction oxygen pressure is 0.8Mpa, and the reaction time is 2h; after the reaction is completed, the reactor is cooled and the pressure is released, and the slurry is opened safely to separate the liquid from the solid, and the filter residue is washed twice with hot water at 70°C, and weighed. analyze. The slag contains 0.008% indium, 56.38% lead, and 35.06% tin; the filtrate contains 0.48g/L indium and 0.012g/L lead. Based on slag, the indium leaching rate is 97.65%, the tin leaching rate is 0.86%, the lead leaching rate is 0.35%, and the zinc leaching rate is 98.7%.

Add reduced iron powder at 1.0g/L to the filtrate at a temperature of 40°C, stir for 1.0h, then clarify and filter, and add CaS at 55°C according to 1.5 times the reaction amount of heavy metal impurities in the solution, stir for 2h and then add Appropriate amount of flocculant for filtration. The filtrate is naturally clarified for 8 hours and then sent for extraction; the sulfide slag is washed and then returned to lead smelting for recovery, and the slag contains less than 0.015% indium. the

The indium-containing liquid after impurity removal is extracted with 30% P204+70% sulfonated kerosene extraction system, the control conditions are: extraction temperature is room temperature, 4-stage extraction O/A=1:10, 3-stage washing O/A=5 : 1, 4-stage back extraction O/A=6: 1, two-phase mixing time 3min, two-phase clarification separation time 5min; after the extraction, adjust the pH value of the back-extraction solution to 2.0 with NaOH, further separate impurities, and the filter residue returns to acid In the immersion process, add acid to the filtrate to adjust the pH to about 1.0, and then add fresh zinc flakes 5 times the theoretical amount for replacement. The reaction time is 72 hours at room temperature. After the replacement, the sponge indium is washed, pressed, alkali-fused, and electrolytically refined using conventional indium electrolysis conditions to obtain a product containing more than 99.996% indium. The replacement fluid contains less than 1 mg/L indium. the

Example 3

Take 200Kg of the raw material used in Example 1, grind the sample preparation until _the particle size is 110 mesh, add a solution of _H2SO4 with a mass percentage concentration of 15% according to the reaction liquid-solid ratio of 5:1 to adjust the slurry, and put it in a closed titanium autoclave. The reaction control conditions are: temperature 180°C, oxygen pressure 1.5MPa, time 3h; after the reaction, the reactor is cooled and opened safely after pressure release, the slurry is separated from liquid to solid, the filter residue is washed twice with hot water at 80°C, and then returned to lead The recovery system is processed, and the filtrate is sent for impurity removal and extraction. The filter residue was sampled and analyzed: 0.012% indium, 55.67% lead, 34.86% tin; filtrate: 0.457g/L indium, 0.014g/L lead, 1.35g/L zinc. In the leaching reaction, the In leaching rate is 97.48%, the Sn leaching rate is 0.95%, the Pb leaching rate is 0.41%, and the Zn leaching rate is 98.3%.

Add the above filtrate to reduced iron powder (or fresh iron filings) at 3.0g/L at a temperature of 45°C, stir for 1.5h, then clarify and filter, and add 0.7 times the reaction amount of heavy metal impurities in the solution at 50°C Na ₂ S, stirred and reacted for 90 minutes, added an appropriate amount of flocculant to filter, and the filtrate was naturally clarified for 12 hours before being extracted; the sulfide slag was washed and then returned to lead smelting for recovery, and the slag contained less than 0.015% indium.

The indium-containing liquid after impurity removal is extracted with 30% P204+70% sulfonated kerosene extraction system, the control conditions are: extraction temperature is room temperature, 4-stage extraction O/A=1:15, 3-stage washing O/A=5 : 1, 4-stage back extraction O/A=6: 1, two-phase mixing time 4min, two-phase clarification separation time 5min; after the extraction, adjust the pH value of the back-extraction solution to 1.8 with NaOH, further separate impurities, and filter residue returns to acid In the immersion process, add acid to the filtrate to adjust the pH to about 1.0, then add fresh zinc flakes 6 times the theoretical amount for replacement, the reaction time is 60 hours, and the temperature is normal. After the replacement, the indium sponge is washed, pressed, alkali-fused, and electrolytically refined using conventional indium electrolysis conditions to obtain a product containing more than 99.997% indium. the

After indium replacement, the indium content in the liquid is less than 0.001g/L, which can be sold or recycled to produce zinc products; the indium content in the raffinate in the indium extraction process is less than 0.1mg/L. , the conventional sulfide-lime wastewater treatment process can meet the national discharge standards, and the wastewater treatment waste residue is returned to lead smelting ingredients for utilization. the

During the operation of the process, the indium leaching rate is greater than 97%, the direct recovery rate is greater than 92.5%, among which the purification and impurity removal recovery rate is greater than 98%, the extraction recovery rate is greater than 98.5%, the sponge indium replacement recovery rate is greater than 99%, and the electrolysis recovery rate is greater than 98%. Greater than 99.3%. And it can produce indium ingots that meet the national standard In99.995 or higher, and the enrichment rate of valuable metals is over 96%. the

Example 4

Lead electrolysis scum enriched indium fume composition (wt, %): In 10.29, Zn 3.08, Pb25.31, Fe 0.41, As 0.98, Sb 1.36. the

Take 200Kg of the above-mentioned raw material, grind the sample to a particle size of 120 mesh, add a solution of _H2SO4 with a mass percentage concentration of 250% according to the reaction liquid-solid ratio of 10:1 to adjust the _slurry , and put it into a closed high-pressure reactor, and control the conditions It is: reaction temperature 160°C, reaction oxygen pressure 1.0MPa, reaction time 5.0h; after the reaction, the reactor is cooled and opened safely after pressure release, the slurry is separated from liquid to solid, and the filter residue is washed once with hot water at 80°C before returning to lead The recovery system is processed, and the filtrate is sent for impurity removal and extraction. The filter residue was sampled and analyzed: 0.014% indium, 68.18% lead, 2.63% tin; filtrate: 0.426g/L indium, 0.020g/L lead, 3.35g/L zinc. In the leaching reaction, the In leaching rate is 97.31%, the Sn leaching rate is 0.62%, the Pb leaching rate is 0.51%, and the Zn leaching rate is 99.2%.

Add reduced iron powder (or fresh iron filings) to the above filtrate at a temperature of 50°C at a rate of 5g/L, stir for 2 hours, then clarify and filter, and add _Na2 S, stirring and reacting for 1 hour, adding an appropriate amount of flocculant for filtration, and the filtrate was naturally clarified for 9 hours before being sent for extraction; the sulfide slag was washed and then returned to lead smelting for recovery, and the slag contained less than 0.013% indium.

The indium-containing liquid after impurity removal is extracted with 30% P204+70% sulfonated kerosene extraction system, the control conditions are: extraction temperature is room temperature, 4-stage extraction O/A=1:15, 3-stage washing O/A=5 : 1, 4-stage back extraction O/A=7: 1, two-phase mixing time 5min, two-phase clarification separation time 5min; after the extraction is completed, adjust the pH value of the stripping solution to 1.9 with NaOH, further separate impurities, and filter residue returns to acid In the immersion process, add acid to the filtrate to adjust the pH to about 1.0, then add fresh zinc flakes 8 times the theoretical amount for replacement, the reaction time is 36 hours, and the temperature is normal. After the replacement, the sponge indium is washed, pressed, alkali fused, and electrolytically refined using conventional indium electrolysis conditions to obtain a product containing 99.995% indium. the

After indium replacement, the indium content in the liquid is less than 0.0008g/L, which can be sold or recycled to produce zinc products; the indium content in the raffinate in the indium extraction process is less than 0.09mg/L. , the conventional sulfide-lime wastewater treatment process can be used to meet the national discharge standards, and the wastewater treatment waste residue is returned to lead smelting ingredients for utilization. the

During the operation of this process, the indium leaching rate is greater than 97%, the direct recovery rate is greater than 92.3%, among which the purification and impurity removal recovery rate is greater than 98.2%, the extraction recovery rate is greater than 98.7%, the sponge indium replacement recovery rate is greater than 99.2%, and the electrolysis recovery rate is greater than 98.2%. Greater than 99.5%. And it can produce indium ingots that meet the national standard In99.995 or higher, and the enrichment rate of valuable metals is over 96.3%. the

Claims (5)

1. A method for extracting indium from oxygen pressure in indium-rich dust, characterized in that it has the following specific steps: A. Oxygen pressure acid leaching Crush the indium-rich dust to a particle size of 100-120 mesh, and add it to the pre-configured dilute sulfuric acid solution with a mass percentage concentration of 10%-30% according to the liquid-solid ratio of 4-10:1, wherein the unit of the liquid-solid ratio is L/ kg; In the closed high-pressure reactor, feed industrial oxygen; control the reaction leaching temperature to 120-180°C, the industrial oxygen pressure to 0.6-1.5MPa, and the reaction time to 2-5h; after the reaction is completed, the reactor is opened after cooling and pressure relief. The reacted slurry is subjected to liquid-solid separation to obtain the separated slag and oxygen pressure acid leaching filtrate; the separated slag is washed with hot water for 1 to 2 times and sent to lead smelting for comprehensive recovery of lead, tin and silver; B. Reduction and removal of impurities Put the oxygen pressure acid leaching filtrate into the reaction tank, add fresh fine iron filings in an amount of 3.0g/L, stir and react at a temperature of 45°C for 1.5h, and then carry out clarification and separation; take the supernatant according to the amount of tin, Calculate the total content of bismuth, lead, iron, and arsenic, add Na ₂ S or other sulfides of 0.7 times the theoretical reaction amount, stir and react at a temperature of 50°C for 90 minutes, and then filter to obtain impurity removal liquid and filter residue; the filter residue is washed, Return to lead smelting for comprehensive recycling; C. Extraction and refining of indium ①Extraction After clarification of the impurity-removing solution in step B for 8-10 hours, extract at room temperature using a volume mixing ratio of 30% P204+70% sulfonated kerosene extraction system. The process control conditions are: four-stage extraction: compared to O/A = 1: 10~15, mixing time 3~5min; three-stage washing: use 80~100g/L H ₂ SO ₄ to perform two-stage acid washing first, and then one-stage water washing, compared to O/A=(2~ 5): 1, the separation time is 5 minutes; four-stage back extraction: compared to O/A=(3~10): 1, where the back extraction agent uses 6~8mol/L HCl solution, and the mixing time is 3~5 minutes; ② replacement Use NaOH to adjust the pH value of the stripping solution under ventilated conditions, and control the pH of the solution to 1.5-2.0. After the solution is clarified and filtered, use hydrochloric acid to adjust the pH of the solution to 1.0, and then add fresh zinc that is equivalent to 5-8 times the weight of indium in the solution. Sponge replacement, the reaction time is 24-72 hours, when the replacement liquid contains less than 0.001g/L indium; separate the sponge indium and the replacement liquid; send the replacement liquid to recover zinc products; ③Sponge indium alkali fusion After washing and pressing the separated indium sponge, it is melted at 200-300°C under the cover of caustic soda, and the scum is removed to obtain crude indium with a content of more than 99%; the scum is returned to step A for oxygen pressure acid leaching Process cyclic acid leaching; ④ Electrolytic refining Thick indium is melted and cast into ^a thick indium anode, and electrolysis is carried out under the condition of direct current. Period: 5 to 7 days; the refined indium obtained by electrolysis is cast into an ingot after the glycerol iodination method removes impurities in a molten state, and more than 99.995% of refined indium products are obtained; D. Raffinate treatment The raffinate produced in the indium extraction process undergoes gas-phase oil separation and secondary clarification oil separation, and then adopts conventional sulfide precipitation and lime-iron salt wastewater treatment to reach the standard discharge, and the sulfide precipitation residue and waste water filter residue are sent to the reverberatory furnace for batching and recycling. 2. A method for extracting indium under oxygen pressure from indium-rich fumes according to claim 1, further characterized in that: the indium-rich fumes are produced by volatilization furnaces and kilns in the metallurgical process of heavy metals copper, lead, zinc and tin. Produce high-indium-containing materials or enriched slag of other indium-containing products. 3. A method for extracting indium from indium-rich fumes under oxygen pressure according to claim 1, further characterized in that: said fresh finely crushed iron filings are iron filings that have been produced, processed and stored for no more than 7 days. 4. A method for extracting indium from indium-rich fumes under oxygen pressure according to claim 1, further characterized in that: said fresh zinc flakes are 0# zinc flakes whose production, processing and storage time does not exceed 7 days. 5. A method for extracting indium under oxygen pressure from indium-rich dust according to claim 1, further characterized in that: _H2SO4 , _NaOH , _Na2S , P204, Sulfonated kerosene, industrial oxygen, and iron filings are commercially available industrial grade products.

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