Process for combined smelting of high-lead and low-lead complex zinc sulfide concentrate

Abstract

The invention relates to a combined smelting process for high-lead and low-lead complex zinc sulfide concentrate, and belongs to the technical field of metallurgy and chemical engineering. The method is characterized in that the low-lead zinc sulfide concentrate is treated through fluidized bed roasting, neutral leaching, low-acid leaching and hot acid leaching, and hot acid leaching liquid containing high-concentration ferric ions is obtained; and the hot acid leaching solution and the high-lead zinc sulfide concentrate are subjected to oxygen control reduction leaching to obtain a reduction leaching solution with the ferric ion concentration smaller than or equal to 2 g / L, the reduction leaching solution is subjected to copper precipitation, pre-neutralization and indium precipitation, obtained indium precipitation post-solution is subjected to high-temperature oxidation iron precipitation, an iron oxide by-product is produced, and the iron precipitation post-solution is returned to low-acid leaching. According to the method, the high-lead zinc sulfide concentrate and the low-lead zinc sulfide concentrate are subjected to combined treatment, and combined smelting and multi-metal comprehensive recovery of the high-lead and low-lead complex zinc sulfide concentrate are achieved. The method has the advantages of high raw material adaptability, low slag rate, high multi-metal comprehensive recovery rate, greenness, low carbon and the like.

Description

TECHNICAL FIELD

[0001] The application relates to a process for combined smelting of high-lead and low-lead complex zinc sulfide concentrates, and belongs to the technical field of metallurgy and chemical industry. BACKGROUND

[0002] Zinc is an important non-ferrous metal raw material and the third largest non-ferrous metal. Zinc metal has good rolling properties, wear resistance and corrosion resistance, and is mainly used as a surface coating of steel and steel structure, and is widely used in the fields of automobiles, buildings, ships, chemical industry, machinery, batteries and military industry. In recent years, with the acceleration of global industrialization, the demand for zinc continues to rise. Zinc smelting processes mainly include pyrometallurgical zinc smelting and hydrometallurgical zinc smelting, of which hydrometallurgical zinc smelting is the main zinc smelting method, and more than 85% of zinc in the world is produced by hydrometallurgical zinc smelting process. Zinc sulfide concentrate is the main raw material for zinc smelting. With the gradual depletion of high-grade zinc sulfide concentrate, multi-metal complex zinc sulfide concentrate has gradually become the main raw material for zinc smelting.

[0003] The conventional zinc sulfide concentrate hydrometallurgical zinc smelting process includes boiling roasting, neutral leaching and low-acid leaching. The leaching residue is treated by high-temperature reduction and volatilization in a rotary kiln or a fuming furnace to produce secondary zinc oxide dust, and the secondary zinc oxide is then comprehensively recovered. The conventional zinc sulfide concentrate hydrometallurgical zinc smelting process is suitable for treating zinc sulfide concentrate with low lead content. When the conventional hydrometallurgical zinc smelting process is used to treat high-lead zinc sulfide concentrate, a large amount of low-melting-point lead compounds are generated in the boiling roasting process, which causes the agglomeration of the roasted ore, and even causes sintering and shutdown in severe cases. In addition, lead sulfate or lead jarosite is formed in the leaching process, resulting in a high residue rate of the leaching residue, and many valuable metals such as copper and indium are difficult to be leached and remain in the leaching residue. The leaching residue needs to be further treated by pyrometallurgical method, which consumes a large amount of carbon energy such as coal coke, resulting in high energy consumption for residue treatment. In addition, the water-quenched slag produced by the pyrometallurgical treatment contains a high content of valuable metals such as copper, indium, silver, zinc and iron, which are difficult to be recovered and utilized, resulting in a low comprehensive recovery rate of valuable metals. In addition, the zinc sulfide concentrate usually contains a large amount of valuable metals such as iron, copper, indium and silver. In the boiling roasting process, iron and zinc react to form zinc ferrite and other insoluble compounds, and indium and other dispersed metals exist in the form of isomorphism in the zinc ferrite lattice. In the conventional neutral leaching and low-acid leaching process, zinc ferrite is difficult to be dissolved, resulting in a large amount of zinc, iron and indium remaining in the leaching residue, and a low comprehensive recovery rate of multi-metals. Therefore, how to efficiently and jointly treat high-lead zinc sulfide concentrate and low-lead zinc sulfide concentrate, how to improve the comprehensive recovery rate of multi-metals in the hydrometallurgical zinc smelting process, and how to reduce the residue rate of the leaching residue and the energy consumption of the hydrometallurgical zinc smelting process are common technical problems faced by the hydrometallurgical zinc smelting process. SUMMARY

[0004] The present application aims at the deficiency of the conventional process of treating high-lead zinc sulfide concentrate in the wet zinc smelting process, and provides a process of combined smelting of high-lead zinc sulfide concentrate and low-lead zinc sulfide concentrate. The method combines the treatment of high-lead zinc sulfide concentrate and low-lead zinc sulfide concentrate, realizes the combined smelting of high-lead and low-lead complex zinc sulfide concentrate, and the comprehensive high-efficiency recovery of zinc, copper, indium, silver, lead, iron and other metals. The present application uses boiling roasting, neutral leaching, low-acid leaching and hot-acid leaching to treat low-lead zinc sulfide concentrate, avoids the formation of a large amount of low-melting-point compounds such as low-melting-point lead silicate in the boiling roasting process, ensures the smoothness of the boiling roasting process, and uses the neutral leaching, low-acid leaching and hot-acid leaching process to maximize the destruction of the insoluble compounds such as zinc ferrite in the zinc calcine with low lead content, so as to maximize the dissolution of metal compounds such as zinc, copper, indium and iron into the solution, realize the high-efficiency leaching of various metals, greatly reduce the residue rate of the leaching residue, and at the same time, retain the insoluble metals such as silver and lead in the leaching residue, so as to realize the enrichment and recovery of valuable metals such as silver and lead. The hot-acid leaching solution and the high-lead zinc sulfide concentrate are subjected to controlled oxygen reduction leaching, the trivalent iron ions in the leaching solution are reduced to divalent iron ions, the hydrolysis precipitation of high-valence trivalent iron ions is avoided at the source, and the basic conditions are provided for the subsequent separation and recovery of copper, indium and iron; the reduction leaching solution with low trivalent iron ion concentration is subjected to copper precipitation, pre-neutralization and indium precipitation, and copper filter cake, pre-neutralization residue and indium enrichment are obtained, so as to realize the separation and recovery of copper and indium; the obtained post-indium precipitation solution is subjected to high-temperature oxidation and iron precipitation, and iron-containing oxide by-product with iron content ≥50% is obtained, so as to realize the comprehensive recovery of iron. The post-iron precipitation solution returns to the low-acid leaching; the low-acid leaching solution returns to the neutral leaching, and the neutral leaching solution is subjected to purification, electrodeposition and smelting to obtain zinc ingot; the waste electrolyte produced by electrodeposition is used as a leaching agent in the leaching process, so as to realize the acid balance and volume balance in the wet zinc smelting process.

[0005] The present application is realized by the following technical solutions:

[0006] A process of combined smelting of high-lead and low-lead complex zinc sulfide concentrate, characterized by the following:

[0007] Firstly, the low-lead zinc sulfide concentrate is treated by boiling roasting, neutral leaching, low-acid leaching and hot-acid leaching in sequence to obtain a hot-acid leaching solution containing high-concentration trivalent iron ions, and the high-lead zinc sulfide concentrate is subjected to controlled oxygen reduction leaching to obtain a reduction leaching solution with trivalent iron ion concentration ≤2 g / L, the lead content in the low-lead zinc sulfide concentrate is less than 2.0 wt%, and the lead content in the high-lead zinc sulfide concentrate is greater than or equal to 2.0 wt%.

[0008] After obtaining the reduction leaching solution, the reduction leaching solution is subjected to copper precipitation, pre-neutralization and indium precipitation, the post-indium precipitation solution is subjected to high-temperature oxidation and iron precipitation, and iron-containing oxide by-product is obtained, and the post-iron precipitation solution returns to the low-acid leaching.

[0009] The steps adopted are as follows:

[0010] Step (1): low-lead zinc sulfide concentrate roasting and leaching: low-lead zinc sulfide concentrate with an average lead content of less than 2.0wt% is treated by roasting, neutral leaching, low-acid leaching and hot-acid leaching to produce neutral leaching solution and hot-acid leaching solution. The neutral leaching solution is treated by conventional purification, electrodeposition and casting to produce zinc ingots.

[0011] Step (2): hot-acid leaching solution and high-lead zinc concentrate controlled oxygen reduction leaching: high-lead zinc sulfide concentrate with a lead content of ≥2.0wt%, preferably high-lead zinc sulfide concentrate with a lead content of ≥2.5wt%, is added to the hot-acid leaching solution produced in the above step (1), and controlled oxygen reduction leaching is carried out under the conditions of a temperature of 90-120℃, a reaction time of 60-360min, a liquid-solid ratio (mL / g) of 6-15:1, an oxygen addition amount of 0.5-1.2 times the theoretical amount, and a final sulfuric acid concentration of 10-50g / L, to obtain a reduction leaching solution with a ferric ion concentration of ≤2g / L and a reduction leaching residue.

[0012] Step (3): copper precipitation, pre-neutralization and indium precipitation in the reduction leaching solution: reduction iron powder is added to the reduction leaching solution produced in the above step (2), and copper precipitation is carried out under the conditions of a temperature of 75-85℃, a reaction time of 10-30min, an iron powder addition amount of 1.0-1.3 times the theoretical amount, and a final sulfuric acid concentration of 10-40g / L, to obtain a copper filter cake and a post-copper precipitation solution; a neutralizing agent is added to the post-copper precipitation solution, and the sulfuric acid concentration is neutralized to 5-15g / L under the conditions of a temperature of 70-85℃ and a reaction time of 60-120min, to obtain a pre-neutralization residue and a post-pre-neutralization solution; a precipitating agent is added to the post-pre-neutralization solution, and indium precipitation is carried out under the conditions of a temperature of 70-85℃ and a reaction time of 60-120min, to obtain an indium concentrate and a post-indium precipitation solution.

[0013] Step (4): high-temperature oxidation and iron precipitation in the post-indium precipitation solution: the post-indium precipitation solution produced in the above step (3) is heated to 150-185℃ and reacted for 120-240min, oxygen is introduced, and the oxygen partial pressure is maintained at 0.5-1.0MPa to carry out high-temperature oxidation and iron precipitation, and the reaction is separated into a liquid and a solid to obtain an iron oxide byproduct and a post-high-temperature oxidation and iron precipitation solution; the post-high-temperature oxidation and iron precipitation solution is returned to the low-acid leaching of the above step (1).

[0014] Further, the low-lead zinc sulfide concentrate comprises the following components in mass percentage: Pb 0.1-1.99%, Zn 45-50%, S 29-32%, Cu 0.2-0.8%, and In 50-200 g / t. The high-lead zinc sulfide concentrate comprises the following components in mass percentage: Pb 2.0-10%, preferably 2.5-10%, Zn 35-46%, S 28-32%, Cu 0.3-2.0%, and In 50-200 g / t.

[0015] Further, the process conditions for the boiling roasting of the low-lead zinc sulfide concentrate are as follows: roasting temperature 920-980℃, roasting time 30-60 min, and the sulfur content in the roasted product zinc calcine is less than 2.0%; the reaction temperature for the neutral leaching is 75-85℃, the leaching time is 2-3 hours, and the terminal iron ion concentration is less than 10 mg / L.

[0016] As preferred, the concentration of H2SO4 in the high-temperature oxidation and iron precipitation solution is 30-60 g / L, further preferably 45-55 g / L, and the concentration of Zn is 80-140 g / L, further preferably 100-130 g / L.

[0017] As preferred, the concentration of H2SO4 in the waste electrolyte is 160-175 g / L, and the concentration of Zn is 40-50 g / L. As further preferred, the concentration of H2SO4 in the waste electrolyte is 165-170 g / L, and the concentration of Zn is 40-45 g / L.

[0018] As preferred, in the low-acid leaching, the liquid-solid ratio is 9-11:1 in terms of mL / g, and the terminal sulfuric acid concentration is 5-10 g / L.

[0019] As preferred, in the hot-acid leaching, the liquid-solid ratio is 10-12:1 in terms of mL / g, and the terminal sulfuric acid concentration is 25-30 g / L.

[0020] As preferred, in step (2), in the controlled-oxygen reduction leaching, the liquid-solid ratio is 12-15:1 in terms of mL / g, the oxygen addition amount is 1.0-1.2 times the theoretical amount, and the terminal sulfuric acid concentration is 20-35 g / L.

[0021] In the present application, the concentration of H2SO4 in the high-temperature oxidation and iron precipitation solution, the concentration of Zn, the concentration of H2SO4 in the waste electrolyte, the liquid-solid ratio and the terminal sulfuric acid concentration in the low-acid leaching, the liquid-solid ratio and the terminal sulfuric acid concentration in the hot-acid leaching, and the liquid-solid ratio and the terminal sulfuric acid concentration in the controlled-oxygen reduction leaching are controlled in order to realize the smooth operation of the process and the efficient recovery of all metal resources in the raw materials.

[0022] Further, in step (3), the neutralizing agent is selected from at least one of zinc calcine, limestone powder; the precipitating agent is selected from at least one of limestone powder, zinc powder, and the amount of the precipitating agent is 5-15 g / L.

[0023] The zinc recovery rate is greater than or equal to 98.0%, the copper recovery rate is greater than or equal to 90.0%, the indium recovery rate is greater than or equal to 80%, the silver recovery rate is greater than or equal to 98.0%, and the lead recovery rate is greater than or equal to 98.0%.

[0024] The calculation method of the metal recovery rate is as follows:

[0025] Zinc recovery rate: η Zn =(ω 1Zn ×m1+ω 2Zn ×m2-ω aZn ×m a -ω bZn ×m b ) / (ω 1Zn ×m1+ω 2Zn ×m2) ×100%

[0026] In the formula, ω 1Zn is the mass percentage of zinc in the low-lead zinc sulfide concentrate, %; m1 is the mass of the low-lead zinc sulfide concentrate, g; ω 2Zn is the mass percentage of zinc in the high-lead zinc sulfide concentrate, %; m2 is the mass of the high-lead zinc sulfide concentrate, g; ω aZn is the mass percentage of zinc in the leaching final residue, %; m a is the mass of the leaching final residue, g; ω bZn is the mass percentage of zinc in the reduction leaching residue, %; m b is the mass of the reduction leaching residue, g.

[0027] Copper recovery rate: η Cu =(ω Cu ×m Cu ) / (ω 1Cu ×m1+ω 2Cu ×m2) ×100%

[0028] In the formula, ω Cu is the mass percentage of copper in the copper filter cake, %; m Cu is the mass of the copper filter cake, g; ω 1Cu is the mass percentage of copper in the low-lead zinc sulfide concentrate, %; m1 is the mass of the low-lead zinc sulfide concentrate, g; ω 2Cu is the mass percentage of copper in the high-lead zinc sulfide concentrate, %; m2 is the mass of the high-lead zinc sulfide concentrate, g.

[0029] Indium recovery rate: η In =(ω In ×mIn ) / (ω 1In ×m1+ω 2In ×m2) ×100%

[0030] Formula wherein ω In is the mass content of indium in the indium concentrate, g / t; m In is the mass of the indium concentrate, g; ω 1In is the mass content of indium in the low-lead zinc sulfide concentrate, g / t; m1 is the mass of the low-lead zinc sulfide concentrate, g; ω 2In is the mass content of indium in the high-lead zinc sulfide concentrate, g / t; m2 is the mass of the high-lead zinc sulfide concentrate, g.

[0031] Silver recovery rate: η Ag =(ω aAg ×m a -ω bAg ×m b ) / (ω 1Ag ×m1+ω 2Ag ×m2) ×100%

[0032] Formula wherein ω aAg is the mass content of silver in the leaching final residue, g / t; m a is the mass of the leaching final residue, g; ω bAg is the mass content of silver in the reduction leaching residue, g / t; m b is the mass of the reduction leaching residue, g; ω 1Ag is the mass content of silver in the low-lead zinc sulfide concentrate, g / t; m1 is the mass of the low-lead zinc sulfide concentrate, g; ω 2Ag is the mass content of silver in the high-lead zinc sulfide concentrate, g / t; m2 is the mass of the high-lead zinc sulfide concentrate, g.

[0033] Lead recovery rate: η Pb =(ω aPb ×m a -ω bPb ×m b ) / (ω 1Pb ×m1+ω 2Pb ×m2) ×100%

[0034] Formula wherein ω aPb is the mass percentage of lead in the leaching final residue, %; m a is the mass of the leaching final residue, g; ω bPb is the mass percentage of lead in the reduction leaching residue, %; m b is the mass of the reduction leaching residue, g; ω 1Pb is the mass percentage of lead in the low-lead zinc sulfide concentrate, %; m1 is the mass of the low-lead zinc sulfide concentrate, g; ω 2Pbwherein m1 is the mass percentage of lead in the high-lead zinc sulfide concentrate, m2 is the mass of the high-lead zinc sulfide concentrate, g.

[0035] The recovery rate of zinc is ≥98.5% (zinc sulfide concentrate to leaching residue and reduction leaching residue), the recovery rate of copper is ≥90.0% (zinc sulfide concentrate to copper filter cake), the recovery rate of indium is ≥80% (zinc sulfide concentrate to indium concentrate), the recovery rate of silver is ≥99.5% (zinc sulfide concentrate to leaching final residue and reduction leaching residue), and the recovery rate of lead is ≥99.5% (zinc sulfide concentrate to leaching final residue and reduction leaching residue).

[0036] The present application first proposes to take a large amount of low-lead zinc sulfide concentrate as the processing object, to perform boiling roasting, neutral leaching, low-acid leaching and hot-acid leaching treatment with appropriate parameters, and then to introduce high-lead zinc concentrate at an appropriate node (after obtaining the hot-acid leaching solution) and to perform controlled oxygen reduction leaching, and to perform treatment on the reduction leaching solution to achieve effective separation and recovery of various metals.

[0037] The present application has the following advantages:

[0038] (1) The present method jointly processes high-lead zinc sulfide concentrate and low-lead zinc sulfide concentrate, realizes joint smelting of high-lead and low-lead complex zinc sulfide concentrate and comprehensive recovery of zinc, copper, indium, silver, lead, iron and other metals, and has the advantages of strong adaptability to raw materials, low slag rate, high comprehensive recovery rate of multiple metals, etc. In the present method, the low-lead zinc concentrate is treated by boiling roasting, neutral leaching, low-acid leaching and hot-acid leaching, which maximally destroys zinc ferrite and other compounds that are difficult to dissolve, maximally dissolves metal compounds such as zinc, copper, indium and iron into the solution, realizes efficient leaching of multiple metals, produces hot-acid leaching solution containing high-concentration ferric ions and leaching final residue rich in zinc, indium, copper and other metal ions, and leaching final residue rich in lead and silver is used as lead smelting raw material. After the treatment, the zinc ferrite component in the low-lead zinc sulfide concentrate roasting product is maximally destroyed, the slag rate of the leaching residue is reduced from 45-55% in the conventional process to 8-15%, the slag rate is greatly reduced, and the leaching rate of zinc, copper, indium, iron and other metals is greatly improved. At the same time, the high-lead zinc concentrate is mixed with the hot-acid leaching solution for controlled oxygen reduction leaching, the high-lead zinc concentrate reduces the ferric ions in the hot-acid leaching solution of the low-lead zinc concentrate roasting product to ferrous ions, the hydrolysis precipitation of ferric ions is avoided at the source, and the basis conditions for the recovery of copper, indium and iron are provided. Therefore, the present method jointly processes high-lead zinc sulfide concentrate and low-lead zinc sulfide concentrate, realizes comprehensive recovery of multiple metals, and has obvious advantages.

[0039] (2) The method uses boiling roasting, neutral leaching and hot acid leaching to process low-lead zinc sulfide concentrate, avoids the production of a large amount of low-melting-point lead-containing compounds in the roasting process of zinc concentrate, and maximizes the leaching of valuable elements such as zinc, iron, copper and indium into the solution, thereby reducing the slag rate of the leaching residue. The high-lead zinc sulfide concentrate and the hot acid leaching solution produced by roasting-leaching of low-lead zinc sulfide concentrate are directly subjected to controlled oxygen reduction leaching, realizing direct leaching of high-lead zinc sulfide concentrate and eliminating the boiling roasting step of high-lead zinc sulfide concentrate. The method can adapt to the treatment of complex zinc sulfide concentrate with different lead contents, has strong adaptability to raw materials, avoids the formation of a large amount of low-melting-point compounds such as lead silicate in the boiling roasting process, and ensures smooth boiling roasting process.

[0040] (3) The leaching residue produced by the method has a low slag rate and a high enrichment ratio of lead and silver in the leaching residue, and can be used as a lead smelting raw material for pyrometallurgical treatment. Compared with the method of using a rotary kiln or a fuming furnace for high-temperature pyrometallurgical treatment of the neutral leaching residue in conventional zinc hydrometallurgy, the amount of residue to be treated by the method is small, the use of carbon fuels such as coal is reduced, and the method has the advantages of low energy consumption, low carbon dioxide emissions and green low carbon. BRIEF DESCRIPTION OF DRAWINGS

[0041] Figure 1 The process flowchart of the present application. DETAILED DESCRIPTION

[0042] The present application will be further described below in conjunction with specific embodiments.

[0043] Example 1

[0044] A process for the combined smelting of high-lead and low-lead complex zinc sulfide concentrate, the specific steps are as follows: low-lead zinc sulfide concentrate boiling roasting and leaching: take 100 kg of low-lead zinc sulfide concentrate used by a certain zinc hydrometallurgy enterprise in Yunnan (its main chemical components are Zn 48.6%, Pb 1.1%, Cu 0.6%, Fe 8.5%, S 32.1%, SiO2 2.8%, In 200 g / t, Ag 95 g / t), the zinc sulfide concentrate is subjected to conventional boiling roasting at 950°C to produce roasted ore, and the roasted ore is leached with waste electrolyte (its main chemical components are H2SO4 170 g / L, Zn 45 g / L) and low-acid leaching solution as leaching agent at 80°C for 2 h, the reaction end point pH value is 5.2, and the liquid-solid separation after the reaction is carried out to obtain 60 kg of neutral leaching solution and neutral leaching residue. The neutral leaching solution is subjected to conventional three-stage purification process (zinc powder addition amount is 60 kg / t of zinc, first-stage purification temperature is 60-65°C, second-stage purification temperature is 85-90°C, and third-stage purification temperature is 55-65°C), electrodeposition (current density is 500 A / m2, and the cathode is made of copper), and zinc electrowinning (zinc powder addition amount is 60 kg / t of zinc), and the neutral leaching residue is subjected to pyrometallurgical treatment in a rotary kiln at 950-1050°C for 2-3 h, and then the pyrometallurgical treatment is carried out in a fuming furnace at 950-1050°C for 2-3 h, thereby obtaining 60 kg of zinc concentrate and 40 kg of slag. 2, slot voltage 3.4V, electrolytic cycle 36h, electrolyte temperature 40~42℃], melting and casting treatment (melting and casting temperature 460~480℃, melting and casting time 4~6 hours), to produce zinc ingot.

[0045] 65kg of neutral leaching residue was mixed with 450L of high-temperature oxidized iron precipitation solution (main chemical components: H2SO4 52g / L, Zn 115g / L) and 200L of waste electrolyte (main chemical components: H2SO4 170g / L, Zn 45g / L) to carry out low-acid leaching under the conditions of temperature 85℃, reaction time 120min, liquid-solid ratio (mL / g) 10:1, and final sulfuric acid concentration 15g / L, and then liquid-solid separation was carried out to obtain low-acid leaching solution and low-acid leaching residue, and the low-acid leaching solution was returned to the next round of neutral leaching. The low-acid leaching residue was mixed with the waste electrolyte to carry out hot-acid leaching under the conditions of temperature 85℃, reaction time 120min, liquid-solid ratio (mL / g) 10:1, and final sulfuric acid concentration 30g / L, and then liquid-solid separation was carried out to obtain hot-acid leaching solution and hot-acid leaching residue.

[0046] The hot acid leaching solution and high-lead zinc concentrate controlled oxygen reduction leaching: 30 kg of high-lead zinc sulfide concentrate with a lead content of 5.2% (the main chemical components of which are Zn 43.7%, Pb 5.2%, Cu 0.8%, Fe 9.5%, S 31.5%, SiO2 3.2%, In 260 g / t, and Ag 130 g / t) was added to the hot acid leaching solution produced in the above step, and controlled oxygen reduction leaching was carried out under the conditions of a temperature of 110°C, a reaction time of 90 min, a liquid-solid ratio (mL / g) of 12:1, an oxygen addition amount of 1.0 times the theoretical amount, and a final sulfuric acid concentration of 30 g / L, to obtain a reduction leaching solution with a ferric ion concentration of 1.5 g / L and a reduction leaching residue of 12 kg after liquid-solid separation at the end of the reaction. The main components of the reduction leaching residue are Zn 7.5%, Pb 12.9%, Cu 0.3%, Fe 4.2%, S 62%, SiO2 6.8%, In 60 g / t, and Ag 321 g / t. Copper precipitation, pre-neutralization, and indium precipitation were carried out on the reduction leaching solution: copper precipitation was carried out on the reduction leaching solution by adding reduction iron powder under the conditions of a temperature of 75°C, a reaction time of 20 min, an iron powder addition amount of 1.1 times the theoretical amount, and a final sulfuric acid concentration of 20 g / L, to obtain a copper filter cake with a copper content of 45.8% and a copper-precipitated solution after liquid-solid separation at the end of the reaction; zinc calcine was added as a neutralizing agent (the amount of the precipitant was 30 g / L) to the copper-precipitated solution, and the sulfuric acid concentration was neutralized to 7 g / L under the conditions of a temperature of 80°C and a reaction time of 60 min, to obtain a pre-neutralization residue and a pre-neutralization solution after liquid-solid separation at the end of the reaction; indium precipitation was carried out on the pre-neutralization solution by adding zinc powder as a precipitant under the conditions of a temperature of 75°C and a reaction time of 90 min, to obtain an indium concentrate of 9.0 kg and an indium-precipitated solution after liquid-solid separation at the end of the reaction. The indium content of the indium concentrate was 2533 g / t. High-temperature oxidation and iron precipitation were carried out on the indium-precipitated solution: the indium-precipitated solution was heated to 180°C and reacted for 180 min, oxygen was introduced, and the oxygen partial pressure was maintained at 0.8 MPa to carry out high-temperature oxidation and iron precipitation, to obtain an iron oxide byproduct of 17.9 kg and an iron-precipitated solution after high-temperature oxidation and iron precipitation, and the iron content of the iron oxide byproduct was 57.4%. The iron-precipitated solution was returned to the low-acid leaching step of the process.

[0047] In this embodiment example, the zinc recovery rate was 98.2%, the copper recovery rate was 90.0%, the indium recovery rate was 82.0%, the silver recovery rate was 98.6%, and the lead recovery rate was 99.1%.

[0048] Example 2

[0049] A process for combined smelting of high-lead and low-lead complex zinc sulfide concentrates, the specific steps being as follows: low-lead zinc sulfide concentrate boiling roasting and leaching: 100 kg of low-lead zinc sulfide concentrate (its main chemical components are Zn 47.5%, Pb 1.35%, Cu 0.75%, Fe 10.2%, S 32.3%, SiO2 2.6%, In 145 g / t, Ag 115 g / t) used by a zinc hydrometallurgy enterprise in Guangxi is taken, the zinc sulfide concentrate is produced into roasted ore at 950 DEG C by using conventional boiling roasting, the roasted ore is leached by using waste electrolyte (its main chemical components are H2SO4 175 g / L, Zn 47 g / L) and low-acid leaching solution as leaching agent at 82 DEG C for 2 h, the reaction end point pH value is 5.15, and neutral leaching solution and neutral leaching residue 63 kg are obtained by liquid-solid separation after reaction. The neutral leaching solution is subjected to conventional purification (zinc powder addition amount 60 kg / t of zinc, first-stage purification temperature 60-65 DEG C, second-stage purification temperature 85-90 DEG C, and third-stage purification temperature 55-65 DEG C), electrodeposition (current density 500 A / m 2 , tank voltage 3.4 V, electrolysis cycle 36 h, and electrolyte temperature 40-42 DEG C), and casting treatment (casting temperature 460-480 DEG C and casting time 4-6 h) to produce zinc ingot. The 63 kg of neutral leaching residue is mixed with 500 L of high-temperature iron-oxidation and precipitation solution (its main chemical components are H2SO4 48 g / L and Zn 108 g / L) and 130 L of waste electrolyte (its main chemical components are H2SO4 175 g / L and Zn 47 g / L) to carry out low-acid leaching under the conditions of temperature 85 DEG C, reaction time 150 min, liquid-solid ratio (mL / g) 10:1, and end point sulfuric acid concentration 20 g / L, and liquid-solid separation is carried out to obtain low-acid leaching solution and low-acid leaching residue, and the low-acid leaching solution is returned to the next round of neutral leaching. The low-acid leaching residue is mixed with waste electrolyte to carry out hot-acid leaching under the conditions of temperature 90 DEG C, reaction time 150 min, liquid-solid ratio (mL / g) 12:1, and end point sulfuric acid concentration 25 g / L, and liquid-solid separation is carried out to obtain hot-acid leaching solution and hot-acid leaching residue.

[0050] The hot acid leaching solution and high-lead zinc concentrate controlled oxygen reduction leaching: 30 kg of high-lead zinc sulfide concentrate with a lead content of 4.8% (the main chemical components of which are Zn 45.2%, Pb 4.8%, Cu 0.68%, Fe 12.5%, S 32.1%, SiO2 2.9%, In 210 g / t, and Ag 173 g / t) was added to the hot acid leaching solution produced in the above step, and controlled oxygen reduction leaching was carried out under the conditions of a temperature of 100°C, a reaction time of 120 min, a liquid-solid ratio (mL / g) of 13:1, an oxygen addition amount of 1.1 times the theoretical amount, and a final sulfuric acid concentration of 20 g / L, and liquid-solid separation was performed after the reaction to obtain a reduction leaching solution with a ferric ion concentration of 1.2 g / L and a reduction leaching residue of 12.6 kg. The main components of the reduction leaching residue are Zn 6.7%, Pb 11.2%, Cu 0.21%, Fe 3.8%, S 67%, SiO2 6.5%, In 65 g / t, and Ag 395 g / t. Copper precipitation, pre-neutralization, and indium precipitation were performed on the reduction leaching solution: copper precipitation was performed on the reduction leaching solution by adding reduction iron powder under the conditions of a temperature of 80°C, a reaction time of 30 min, an iron powder addition amount of 1.1 times the theoretical amount, and a final sulfuric acid concentration of 20 g / L, and liquid-solid separation was performed after the reaction to obtain a copper filter cake with a copper content of 48.77% and a post-copper precipitation solution. Limestone powder was added as a neutralizing agent (the amount of the precipitating agent was 20 g / L) to the post-copper precipitation solution, the sulfuric acid concentration was neutralized to 5 g / L under the conditions of a temperature of 75°C and a reaction time of 90 min, and liquid-solid separation was performed after the reaction to obtain a pre-neutralization residue and a post-pre-neutralization solution. Limestone powder was added as a precipitating agent to the post-pre-neutralization solution under the conditions of a temperature of 75°C and a reaction time of 120 min, and indium precipitation was performed, and liquid-solid separation was performed after the reaction to obtain an indium concentrate of 8.6 kg and a post-indium precipitation solution. The indium content in the indium concentrate was 2007 g / t. High-temperature oxidation and iron precipitation were performed on the post-indium precipitation solution: the post-indium precipitation solution was heated to 185°C, oxygen was introduced, and iron precipitation was performed under the conditions of a reaction time of 180 min and an oxygen partial pressure of 0.6 MPa, and liquid-solid separation was performed after the reaction to obtain an iron oxide byproduct of 21.6 kg and a post-high-temperature oxidation and iron precipitation solution. The iron content in the iron oxide byproduct was 58.3%. The post-high-temperature oxidation and iron precipitation solution was returned to the low-acid leaching step of the process.

[0051] In this implementation example, the zinc recovery rate was 98.3%, the copper recovery rate was 90.2%, the indium recovery rate was 83.0%, the silver recovery rate was 98.4%, and the lead recovery rate was 98.1%.

[0052] Example 3

[0053] A process for combined smelting of high-lead and low-lead complex zinc sulfide concentrates, the specific steps are as follows: low-lead zinc sulfide concentrate boiling roasting and leaching: take 100 kg of low-lead zinc sulfide concentrate used by a certain zinc hydrometallurgy enterprise in Hunan (its main chemical components are Zn 49.3%, Pb 0.65%, Cu 0.45%, Fe 8.0%, S 31.6%, SiO2 2.9%, In 90 g / t, Ag 85 g / t), the zinc sulfide concentrate is roasted at 950 DEG C using conventional boiling roasting to produce roasted ore, the roasted ore uses waste electrolyte (its main chemical components are H2SO4 165 g / L, Zn 40 g / L) and low-acid leaching solution as leaching agent, and is reacted at 80 DEG C for 2 h, the reaction end point pH value is 5.2, and the reaction end liquid-solid separation obtains 55 kg of neutral leaching solution and neutral leaching residue. The neutral leaching solution is subjected to conventional purification (zinc powder addition amount 60 kg / t of zinc, first-stage purification temperature 60-65 DEG C, second-stage purification temperature 85-90 DEG C, and third-stage purification temperature 55-65 DEG C), electrodeposition (current density 500 A / m 2 , tank voltage 3.4 V, electrolysis cycle 36 h, and electrolyte temperature 40-42 DEG C), and casting treatment (casting temperature 460-480 DEG C and casting time 4-6 hours) to produce zinc ingots. The 55 kg of neutral leaching residue is mixed with 400 L of high-temperature iron-oxidation and precipitation liquid (its main chemical components are H2SO4 47 g / L and Zn 120 g / L) and 150 L of waste electrolyte (its main chemical components are H2SO4 165 g / L and Zn 40 g / L) to carry out low-acid leaching under the conditions of temperature 85 DEG C, reaction time 150 min, liquid-solid ratio (mL / g) 10:1, and end point sulfuric acid concentration 25 g / L, and liquid-solid separation obtains low-acid leaching solution and low-acid leaching residue, and the low-acid leaching solution is returned to the next round of neutral leaching. The low-acid leaching residue is mixed with waste electrolyte to carry out hot-acid leaching under the conditions of temperature 90 DEG C, reaction time 150 min, liquid-solid ratio (mL / g) 12:1, and end point sulfuric acid concentration 30 g / L, and liquid-solid separation obtains hot-acid leaching solution and hot-acid leaching residue.

[0054] Hot acid leaching solution and high lead zinc concentrate controlled oxygen reduction leaching: 30 kg of high lead zinc sulfide concentrate with lead content of 7.5% (its main chemical components are Zn 43.7%, Pb 7.5%, Cu 1.1%, Fe 11.6%, S 31.2%, SiO2 2.5%, In 312 g / t, Ag 105 g / t) was added to the hot acid leaching solution produced in the above step, and controlled oxygen reduction leaching was carried out under the conditions of temperature 95°C, reaction time 240 min, liquid-solid ratio (mL / g) 10:1, oxygen addition amount 1.0 times of the theoretical amount, and final sulfuric acid concentration 30 g / L, to obtain a reduction leaching solution with ferric ion concentration of 1.0 g / L and a reduction leaching residue of 12.3 kg. The main components of the reduction leaching residue are Zn 7.3%, Pb 18.0%, Cu 0.32%, Fe 4.8%, S 65%, SiO2 6.0%, In 76 g / t, and Ag 252 g / t. Copper precipitation, pre-neutralization, and indium precipitation were carried out on the reduction leaching solution. Copper precipitation was carried out by adding reduction iron powder to the reduction leaching solution under the conditions of temperature 75°C, reaction time 25 min, iron powder addition amount 1.2 times of the theoretical amount, and final sulfuric acid concentration 25 g / L, to obtain a copper filter cake of 1.51 kg and a post-copper precipitation solution. The copper filter cake contains Cu 46.49%. Limestone powder was added to the post-copper precipitation solution as a neutralizing agent, and the sulfuric acid concentration was neutralized to 10 g / L under the conditions of temperature 75°C and reaction time 75 min, to obtain a pre-neutralization residue and a post-pre-neutralization solution. Zinc powder was added to the post-pre-neutralization solution as a precipitant (the amount of the precipitant was 10 g / L), and indium precipitation was carried out under the conditions of temperature 75°C and reaction time 90 min, to obtain an indium concentrate of 7.0 kg and a post-indium precipitation solution. The indium content in the indium concentrate is 2111.4 g / t. High-temperature oxidation and iron precipitation were carried out on the post-indium precipitation solution. The post-indium precipitation solution was heated to 170°C and reacted for 180 min, oxygen was introduced, and the oxygen partial pressure was maintained at 0.4 MPa, to carry out high-temperature oxidation and iron precipitation. The reaction was separated into a liquid and a solid to obtain an iron oxide byproduct of 18.8 kg and a post-high-temperature oxidation and iron precipitation solution. The iron content in the iron oxide byproduct is 55.3%. The post-high-temperature oxidation and iron precipitation solution was returned to the low-acid leaching step of the process.

[0055] In this implementation example, the zinc recovery rate is 98.3%, the copper recovery rate is 90.0%, the indium recovery rate is 80.5%, the silver recovery rate is 98.4%, and the lead recovery rate is 98.5%.

[0056] Comparative Example 1

[0057] The experimental raw materials and the implementation method are consistent with those of Example 1, except that the key step of reducing leaching of the hot acid leaching solution and the high-lead zinc concentrate under controlled oxygen is omitted, which leads to a 12-fold increase in the consumption of iron powder in the process of copper precipitation with iron powder, a sharp increase in the consumption of iron powder, an increase in the concentration of iron ions in the solution to 62 g / L, and difficulty in indium precipitation and high-temperature oxidation and iron precipitation, so that the process cannot run normally.

[0058] Comparative Example 2

[0059] The experimental raw materials and the implementation method are consistent with those of Example 2, except that the reaction temperature of the hot acid leaching solution and the high-lead zinc concentrate under controlled oxygen is 60°C. Under this condition, the concentration of trivalent iron ions in the reduction leaching solution obtained by liquid-solid separation of the reduction leaching reaction end liquid reaches 16 g / L, and the reduction leaching residue is 26.2 kg, and the zinc content in the reduction leaching residue reaches 41%, which leads to an increase in the concentration of trivalent iron ions in the reduction leaching solution, a significant increase in the leaching amount of the reduction residue, and a significant increase in the zinc content in the reduction leaching residue, which leads to difficulty in normal operation of the subsequent steps.

[0060] Comparative Example 3

[0061] The experimental raw materials and the implementation method are consistent with those of Example 3, except that the reduction leaching solution is first pre-neutralized, then indium is precipitated, and then copper is precipitated, which reverses the main steps, increases the mass of indium enrichment by 2.7 times, and reduces the indium content in the indium enrichment to 930 g / t. The copper filter cake contains 22% copper, and the yield of the copper filter cake is reduced by 80%. The overall process indium recovery rate is reduced by 30%, and the copper recovery is reduced by 40%.

[0062] The specific embodiments of the present application are described in detail above, but the present application is not limited to the above-described embodiments, and various changes can be made within the knowledge of those skilled in the art without departing from the spirit of the present application.

Claims

1. A process for the combined smelting of high lead and low lead complex zinc sulphide concentrates, characterised in that: The low-lead zinc sulfide concentrate is treated by boiling roasting, neutral leaching, low-acid leaching and hot-acid leaching to obtain a hot-acid leaching solution containing high-concentration ferric ions; the hot-acid leaching solution is subjected to controlled oxygen reduction leaching with high-lead zinc sulfide concentrate to obtain a reduction leaching solution with ferric ion concentration of ≤2 g / L, the low-lead zinc sulfide concentrate has a lead content of less than 2.0 wt%, and the high-lead zinc sulfide concentrate has a lead content of greater than or equal to 2.0 wt%.

2. A process for the combined smelting of high lead and low lead complex zinc sulphide concentrates according to claim 1, characterized in that: The reduction leaching solution is subjected to copper precipitation, pre-neutralization and indium precipitation, and the post-indium precipitation solution is subjected to high-temperature oxidation and iron precipitation to produce an iron oxide byproduct, and the post-iron precipitation solution is returned to the low-acid leaching.

3. The process for combined smelting of high lead and low lead complex zinc sulphide concentrates according to claim 1, characterized in that; The method comprises the following steps: Step (1): boiling roasting and leaching of low-lead zinc sulfide concentrate: a low-lead zinc sulfide concentrate with an average lead content of less than 2.0 wt% is treated by boiling roasting, neutral leaching, low-acid leaching and hot-acid leaching to produce a neutral leaching solution and a hot-acid leaching solution. The neutral leaching solution is subjected to conventional purification, electrodeposition and smelting treatment to produce zinc ingots; Step (2): controlled oxygen reduction leaching of hot-acid leaching solution with high-lead zinc concentrate: the hot-acid leaching solution produced in the above step (1) is added with high-lead zinc sulfide concentrate with a lead content of greater than or equal to 2.0 wt%, and subjected to controlled oxygen reduction leaching under the conditions of a temperature of 85-120 ℃, a reaction time of greater than or equal to 60 min, preferably 60-360 min, a liquid-solid ratio of 6-15:1 (mL / g), an oxygen addition amount of 0.5-1.2 times the theoretical amount, and a final sulfuric acid concentration of 10-50 g / L, to obtain a reduction leaching solution with ferric ion concentration of ≤2 g / L after liquid-solid separation; Step (3): copper precipitation, pre-neutralization and indium precipitation of reduction leaching solution: the reduction leaching solution produced in the above step (2) is added with reduced iron powder, and subjected to copper precipitation under the conditions of a temperature of 75-85 ℃, a reaction time of greater than or equal to 10 min, preferably 10-30 min, an iron powder addition amount of 1.0-1.3 times the theoretical amount, and a final sulfuric acid concentration of 10-40 g / L, to obtain a copper filter cake and a post-copper precipitation solution after liquid-solid separation; the post-copper precipitation solution is added with a neutralizing agent, and subjected to pre-neutralization under the conditions of a temperature of 70-85 ℃, a reaction time of greater than or equal to 120 min, preferably 120-240 min, to neutralize the sulfuric acid concentration to 5-15 g / L, and obtain a pre-neutralization residue and a post-pre-neutralization solution after liquid-solid separation; the post-pre-neutralization solution is added with a precipitating agent, and subjected to indium precipitation under the conditions of a temperature of 70-85 ℃, a reaction time of greater than or equal to 60 min, preferably 60-120 min, to obtain an indium concentrate and a post-indium precipitation solution after liquid-solid separation; Step (4): high-temperature oxidation and iron precipitation of post-indium precipitation solution: the post-indium precipitation solution produced in the above step (3) is heated to 150-185 ℃, and subjected to high-temperature oxidation and iron precipitation under the conditions of a reaction time of greater than or equal to 150 min, preferably 180-240 min, oxygen is introduced to maintain an oxygen partial pressure of 0.5-1.0 MPa, to obtain an iron oxide byproduct and a post-iron precipitation solution after liquid-solid separation; the post-iron precipitation solution is returned to the low-acid leaching of the above step (1).

4. The process for combined smelting of high lead and low lead zinc sulphide concentrates as claimed in claim 1, wherein: The low-lead zinc sulfide concentrate comprises the following components in mass percentage: Pb 0.1-1.99%, Zn 45-50%, S 29-32%, Cu 0.2-0.8%, and In 50-200 g / t.

5. The process according to claim 1, wherein the high-lead low-lead zinc sulfide concentrate is subjected to combined smelting. The high-lead zinc sulfide concentrate comprises the following components in mass percentage: Pb 2.0-10%, Zn 35-46%, S 28-32%, Cu 0.3-2.0%, and In 50-200 g / t.

6. The process for combined smelting of high lead and low lead complex zinc sulphide concentrates as claimed in claim 1, wherein: The process conditions for the fluidized roasting of the low-lead zinc sulfide concentrate are as follows: roasting temperature 920-980 ℃, roasting time 30-60 min, and sulfur content in the roasted zinc calcine less than 2.0 wt%; reaction temperature for neutral leaching 75-85 ℃, leaching time 2-3 hours, and terminal iron ion concentration in the neutral leaching solution less than 10 mg / L.

7. The process according to claim 1, wherein the high-lead low-lead complex zinc sulfide concentrate is subjected to combined smelting. When the low-acid leaching is performed, the liquid-solid ratio is 9-12:1 (mL / g), and the terminal sulfuric acid concentration is 5-20 g / L. When the hot-acid leaching is performed, the liquid-solid ratio is 8-12:1 (mL / g), and the terminal sulfuric acid concentration is 30-60 g / L.

8. The process for combined smelting of high lead and low lead complex zinc sulphide concentrates as claimed in claim 1, wherein: In step (2), the controlled-oxygen reduction leaching is performed under the following conditions: liquid-solid ratio 10-15:1 (mL / g), oxygen addition amount 1.0-1.2 times the theoretical amount, and terminal sulfuric acid concentration 15-35 g / L.

9. A process for the combined smelting of high lead and low lead complex zinc sulphide concentrates according to claim 2, characterized in that: In step (3), the neutralizing agent is at least one of zinc calcine and limestone powder, and the precipitating agent is at least one of limestone powder, zinc calcine, and zinc powder, and the amount of the precipitating agent is 5-15 g / L.

10. The process for combined smelting of high lead and low lead complex zinc sulphide concentrates as claimed in claim 1, wherein: The zinc recovery rate is ≥98.0%, the copper recovery rate is ≥90.0%, the indium recovery rate is ≥80%, the silver recovery rate is ≥98.0%, and the lead recovery rate is ≥98.0%.

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