阅读说明：本技术 一种湿法炼锌生产中含锌物料的处理方法 (Method for treating zinc-containing material in zinc hydrometallurgy production ) 是由 李存兄 张兆闫 魏昶 邓志敢 李兴彬 樊刚 李旻廷 于 2021-09-16 设计创作，主要内容包括：本发明涉及一种湿法炼锌生产中含锌物料的处理方法,属于湿法冶炼技术领域,本发明包括以下步骤：1)磨矿；2)常压中和浸出；3)I段加压浸出；4)Ⅱ段加压浸出；5)浸出渣浆化洗涤。本发明同时实现了湿法炼锌生产中的含锌物料中有价金属高效浸出和铁的高效同步沉淀。锌浸出率大于98%,获得铁含量低于2g/L的浸出液无需除铁可返回湿法炼锌主系统,巧妙地实现了湿法炼锌生产中的含锌物料中铁由危害杂质向火法炼铅原料的转变,并最终稳定固化于火法炼铅炉渣或窑渣中,省去了湿法炼锌过程溶液除铁的操作工序,简化了工艺流程,降低了过程操作成本,实现了湿法炼锌生产中的含锌物料危废铁渣的减排。(The invention relates to a method for treating zinc-containing materials in the production of zinc hydrometallurgy, which belongs to the technical field of hydrometallurgy and comprises the following steps: 1) grinding ore; 2) neutralizing and leaching under normal pressure; 3) i, pressurizing and leaching; 4) II, pressurizing and leaching; 5) and slurrying and washing the leached residues. The invention simultaneously realizes the high-efficiency leaching of valuable metals and the high-efficiency synchronous precipitation of iron in the zinc-containing materials in the zinc hydrometallurgy production. The zinc leaching rate is more than 98 percent, the obtained leaching solution with the iron content lower than 2g/L can return to a zinc hydrometallurgy main system without iron removal, the transformation of iron in zinc-containing materials in the zinc hydrometallurgy production from harmful impurities to pyrometallurgical lead smelting raw materials is skillfully realized, and the iron is finally stably solidified in pyrometallurgical lead smelting slag or kiln slag, the operation procedure of removing iron from the solution in the zinc hydrometallurgy process is omitted, the process flow is simplified, the process operation cost is reduced, and the emission reduction of hazardous waste iron slag of the zinc-containing materials in the zinc hydrometallurgy production is realized.)

1. A method for treating zinc-containing materials in the production of zinc hydrometallurgy is characterized in that: the zinc-containing material in the zinc hydrometallurgy production is subjected to two-stage pressure leaching, so that the high-efficiency leaching of valuable metals and the synchronous precipitation of iron in the zinc-containing material are realized, and the method specifically comprises the following steps:

(1) pressure leaching in the I section: the zinc-containing material, calcium lignosulfonate or lignin in the zinc hydrometallurgy productionMixing sodium sulfonate with acid liquor, and carrying out the mixing in a pressure kettle with a stirring devicePerforming sectional pressure leaching, controlling the pressure in the kettle to be 0.3-1.4 MPa, controlling the end-point reaction acidity to be 25-45 g/L, and performing liquid-solid separation after the reaction is finished to obtain an I-section pressure leaching liquid with iron content less than 2g/L and an I-section pressure leaching underflow;

（2）and (3) pressure leaching: mixing the I-stage pressure leaching underflow with calcium lignosulfonate or sodium lignosulfonate and zinc electrodeposition waste liquid, and performing the steps in a pressure kettle with a stirring devicePerforming pressure leaching in a section, controlling the pressure in the kettle to be 0.3-1.3 MPa, controlling the end-point reaction acidity to be 50-90 g/L, and performing liquid-solid separation on reaction ore pulp after the reaction is finished to obtainA stage-pressurized leach liquor andand (5) leaching slag under pressure.

2. The method for treating a zinc-containing material in the production of zinc hydrometallurgy according to claim 1, wherein the method comprises the following steps: the reaction temperature of the first stage pressure leaching and the second stage pressure leaching is higher than 100 ℃, and the reaction time is 60-180 min.

3. The method for treating a zinc-containing material in the production of zinc hydrometallurgy according to claim 1, wherein the method comprises the following steps: and the second stage of pressure leaching liquid returns to the first stage of pressure leaching.

4. The method for treating zinc-containing material in the production of zinc hydrometallurgy according to claim 1,the method is characterized in that: the I-section pressure leaching is finished in a vertical reaction kettle or a horizontal reaction kettle,the pressure leaching of the sections is completed in a vertical reaction kettle or a horizontal reaction kettle, and the two sections of reaction kettles are connected in series.

5. A method as claimed in any one of claims 1 to 4 for the treatment of a zinc bearing material in the production of zinc hydrometallurgy, and in which: before the pressure leaching of the section I, the zinc-containing material is ground or crushed.

6. A method as claimed in any one of claims 1 to 4 for the treatment of a zinc bearing material in the production of zinc hydrometallurgy, and in which: and pulping, washing and filtering the II-stage pressure leaching slag to obtain washing water and lead-silver-iron slag.

7. The method for treating a zinc-containing material in the production of zinc hydrometallurgy according to claim 6, wherein the method comprises the following steps: the obtained lead-silver-iron slag can adopt two processes to recover lead and silver in the lead-silver-iron slag, and realizes the harmless treatment of iron: the lead, silver and iron slag directly enters a pyrometallurgical lead smelting system to recover lead and silver in the lead, and iron is finally stably solidified in slag as a slagging agent to realize conversion from impurities to lead smelting raw materials; or firstly recovering silver in the lead-silver-iron slag by adopting a flotation technology to obtain lead-iron slag, then recovering lead in the lead-iron slag by utilizing a pyrogenic process lead smelting or lead enrichment technology, and finally stably solidifying iron serving as a slagging flux in furnace slag or kiln slag in a lead enrichment process to realize the conversion from impurities to lead smelting raw materials.

8. The method for treating a zinc-containing material in the production of zinc hydrometallurgy according to claim 1, wherein the method comprises the following steps: the zinc-containing material is a zinc-containing material accompanied with iron in a zinc hydrometallurgy process.

9. The method of claim 8, wherein the method comprises the steps of: the zinc-containing material is zinc roasted ore and neutral or acidic leaching residue in the zinc hydrometallurgy process.

10. A method according to claim 9 for the treatment of zinc bearing material in the production of zinc hydrometallurgy, including the steps of:

when the zinc-containing material is zinc roasted ore, grinding the zinc roasted ore, leaching at normal pressure, and performing two-stage pressure leaching on the normal-pressure leaching slag;

when the zinc-containing material is neutral or acidic leaching residue, the neutral or acidic leaching residue is subjected to ore grinding treatment and then two-stage pressure leaching.

Technical Field

The invention belongs to the technical field of hydrometallurgy, and particularly relates to a method for treating a zinc-containing material in the production of hydrometallurgy.

Background

The metal zinc is called as a protective agent of modern industry and is an indispensable strategic metal in the fields of aerospace, transportation, energy industry and the like. Sulphide ore is the main raw material for extracting zinc, and at present, more than 85 percent of the global sulphide ore is produced by adopting a wet main process flow of roasting, leaching, purifying and electrodepositing. According to different leaching modes and leaching slag treatment methods, the treatment method of the zinc-containing material is divided into a conventional treatment process and a hot acid leaching process.

In the process of zinc hydrometallurgy, zinc-containing materials are often associated with iron, in the leaching process, iron and valuable metals such as zinc, germanium, copper, indium and the like are leached together and enter a leaching solution, and the leaching solution needs to precipitate the iron before the valuable metals are recovered. At present, the zinc hydrometallurgy enterprises mostly adopt an jarosite method or a goethite method to remove iron in the leachate, so that the iron is precipitated and separated out by dangerous waste jarosite slag or dangerous waste goethite slag, and the iron-containing waste slag has low iron content and is difficult to recycle. How to avoid the existence of valuable metals and iron in the leaching solution is a difficult problem in the industry.

Taking zinc roasted ore as an example, the conventional treatment process of the zinc roasted ore mainly realizes the high-efficiency volatilization of zinc, lead and other associated elements in the leaching slag of the zinc roasted ore and enriches the zinc, lead and other associated elements in zinc oxide smoke dust through a pyrogenic process of high-temperature carbon thermal reduction volatilization, iron in the zinc roasted ore is solidified at high temperature and enters kiln slag or water quenching slag, and secondary harmless treatment is not needed for the high-temperature solidified iron slag. But the conventional treatment process has the defects of large consumption of carbonaceous reducing agent, high energy consumption and low concentration of SO₂Serious smoke pollution or high tail absorption cost, low recovery rate of associated valuable metals such as copper, silver and the like, long process and the like.

The hot acid leaching process of the zinc roasted ore mainly realizes the high-efficiency leaching of valuable metals such as zinc, copper and the like in the leaching slag of the zinc roasted ore through a wet process of high-temperature high-acid leaching and enriches lead, silver and the like in the hot acid leaching slag. However, more than 80% of iron enters the solution in the hot acid leaching process, the iron in the hot acid leaching solution is removed by adopting jarosite method or goethite method in the current zinc hydrometallurgy enterprises, so that the iron is precipitated and separated out by dangerous waste jarosite slag or dangerous waste goethite slag, the iron-containing waste slag has low iron content and is difficult to recycle, and the dangerous waste iron slag needs to be treated in a harmless way according to the lead-zinc industry standard condition which is discharged from the national 2015. According to statistics, 1.2-1.4 t of dangerous waste jarosite slag or 0.8-1.0 t of dangerous waste goethite slag is generated when 1t of zinc roasting ore is treated, and the harmless treatment cost of each ton of dangerous waste iron slag is up to 600-800 yuan. In addition, the existing hot acid leaching process has the problems that valuable elements such as copper, indium, germanium and the like cannot be efficiently recovered and the like.

Disclosure of Invention

In order to overcome the problems in the prior art, the invention provides a method for treating zinc-containing materials in the production of zinc hydrometallurgy, which comprises the steps of carrying out two-stage pressure leaching on the zinc-containing materials, wherein the I-stage pressure leaching realizes the hydrolysis precipitation of iron and the leaching of partial valuable metals in a leaching system, and obtaining I-stage pressure leaching liquor with low acid and low iron and I-stage pressure leaching underflow; i-stage pressure leaching underflow channelStage pressure leachingRealizes the deep leaching of valuable metals and enriches iron, lead and silver inLeaching slag by stage pressurization;and pulping and washing the leaching slag, then sending the leached slag into a pyrometallurgical lead smelting system to comprehensively recover lead and silver in the leached slag, and taking iron as a slagging agent to be stably solidified in the pyrometallurgical lead smelting slag or kiln slag. Not only realizes the effective utilization of iron, but also has higher leaching rate of valuable metals.

In the present invention, unless otherwise specified, the percentages (%) are mass percentages.

In order to realize the purpose, the invention is realized by the following technical scheme:

the zinc-containing material in the zinc hydrometallurgy production is subjected to two-stage pressure leaching, so that the high-efficiency leaching of valuable metals and the synchronous precipitation of iron in the zinc-containing material are realized, and the method specifically comprises the following steps:

(1) pressure leaching in the I section: mixing zinc-containing material, calcium lignosulfonate or sodium lignosulfonate in zinc hydrometallurgy production with acid liquor, and stirring in a pressure kettle with a stirring devicePerforming sectional pressure leaching, controlling the pressure in the kettle to be 0.3-1.4 MPa, controlling the end-point reaction acidity to be 25-45 g/L, and performing liquid-solid separation after the reaction is finished to obtain an I-section pressure leaching liquid with iron content less than 2g/L and an I-section pressure leaching underflow;

（2）and (3) pressure leaching: mixing the I-stage pressure leaching underflow with calcium lignosulfonate or sodium lignosulfonate and zinc electrodeposition waste liquid, and performing the steps in a pressure kettle with a stirring deviceThe pressure in the kettle is controlled to be 0.3-1.3 MPa, and the end-point reaction acidity is controlled to be 50-90 g/L, and obtaining the reaction ore pulp through liquid-solid separation after the reaction is finishedA stage-pressurized leach liquor andand (5) leaching slag under pressure.

Further, the reaction temperature of the first stage pressure leaching and the reaction temperature of the second stage pressure leaching are both higher than 100 ℃, and the reaction time is 60-180 min.

And further, the pressure leaching liquid in the second stage returns to the pressure leaching in the first stage.

Further, the I-section pressure leaching is completed in a vertical reaction kettle or a horizontal reaction kettle,the pressure leaching of the sections is completed in a vertical reaction kettle or a horizontal reaction kettle, and the two sections of reaction kettles are connected in series.

Further, before the I-stage pressure leaching, the zinc-containing material is ground or crushed.

And further, pulping, washing and filtering the II-stage pressure leaching slag to obtain washing water and lead-silver-iron slag.

Further, the obtained lead-silver-iron slag can adopt two processes to recover lead and silver in the lead-silver-iron slag, and realize harmless treatment of iron: the lead, silver and iron slag directly enters a pyrometallurgical lead smelting system to recover lead and silver in the lead, and iron is finally stably solidified in slag as a slagging agent to realize conversion from impurities to lead smelting raw materials; or firstly recovering silver in the lead-silver-iron slag by adopting a flotation technology to obtain lead-iron slag, then recovering lead in the lead-iron slag by utilizing a pyrogenic process lead smelting or lead enrichment technology, and finally stably solidifying iron serving as a slagging flux in furnace slag or kiln slag in a lead enrichment process to realize the conversion from impurities to lead smelting raw materials.

Further, the zinc-containing material is a zinc-containing material accompanied with iron in a zinc hydrometallurgy process.

Further, the zinc-containing material is zinc roasting ore and neutral or acidic leaching residue in the zinc hydrometallurgy process.

Further, when the zinc-containing material is zinc roasted ore, after the zinc roasted ore is subjected to ore grinding treatment, normal-pressure leaching is carried out, and the normal-pressure leaching slag is subjected to two-stage pressure leaching; when the zinc-containing material is neutral or acidic leaching residue, the neutral or acidic leaching residue is subjected to ore grinding treatment and then two-stage pressure leaching.

The invention has the beneficial effects that:

the iron content of the I-stage pressurized leachate obtained by the method is less than 2g/L and is obviously lower than the level of 10-25 g/L of the iron content of the hot acid leachate obtained by the conventional industrial wet-process zinc smelting, and the leachate can return to a main system of the wet-process zinc smelting without iron removal; the obtained leaching final slag (namely the lead-silver-iron slag) contains less than 2% of zinc and is obviously lower than the zinc content of 5-8% of the hot acid leaching slag of zinc hydrometallurgy in the prior art, and the slag can directly enter a lead pyrometallurgical system for matching treatment. The invention greatly improves the comprehensive recovery rate of zinc in the zinc hydrometallurgy process and simplifies the zinc smelting process.

The invention skillfully realizes the conversion of iron in the zinc-containing material from harmful impurities to the pyrometallurgical lead smelting raw material, and finally stably solidifies the zinc-containing material in the form of pyrometallurgical lead smelting slag or kiln slag, thereby saving the operation procedure of removing iron from solution in the process of hydrometallurgy of zinc, simplifying the process flow, reducing the process operation cost and realizing the emission reduction of dangerous waste iron slag in the process of treating the zinc-containing material. Therefore, the conversion of the associated iron in the zinc-containing material from impurities to the resource of the raw material and the emission reduction effect of the hazardous waste iron slag generated by the conversion are important technical innovation and advantages of the invention.

The method overcomes the defects of high energy consumption, complex process flow, low recovery rate of associated metal, large amount of hazardous waste iron slag, high harmless treatment cost and the like in the conventional zinc-containing material treatment process, and simultaneously realizes efficient leaching of valuable metal and efficient synchronous precipitation of iron in the zinc-containing material.

Drawings

FIG. 1 is a schematic process flow diagram of the present invention.

Detailed Description

In order to make the objects, technical solutions and advantages of the present invention more apparent, preferred embodiments of the present invention will be described in detail below to facilitate understanding of the skilled person.

(1) Grinding: wet grinding or dry grinding is carried out on the zinc roasted ore to obtain ore pulp or ore powder with the particle size of 100-200 meshes.

(2) Leaching under normal pressure: mixing and size mixing the ore pulp or ore powder obtained in the step (1) with manganese-containing materials and acidic solution, and then performing series overflow neutralization leaching, wherein Mn is in a size mixing tank²⁺The concentration is 4-6 g/L, the pH value is 2.0-5.4, in the leaching process, the pH value in the leaching tank is sequentially increased along with the extension of the leaching time according to the overflow sequence, the leaching temperature is 60-85 ℃, and the reaction time is 90-180 min.

And (3) carrying out liquid-solid separation after atmospheric pressure leaching to obtain normal-pressure neutralized leaching residues and a leaching solution, and allowing the normal-pressure neutralized leaching solution to enter a purification-electrodeposition process of a zinc hydrometallurgy main system to produce electrodeposited zinc.

The main function of the step is to neutralize and leach the zinc roasted ore by using the I-stage pressurized leachate to realize leaching of most of zinc and Fe in the zinc roasted ore³⁺And neutralizing and hydrolyzing other impurity ions to purify and remove impurities to obtain solution meeting the purification procedure of the zinc hydrometallurgy main system, wherein zinc not leached in the zinc calcine is mainly zinc ferrite (ZnFe)₂O₄) The form exists in the neutralized leaching residue, and the main reactions occur as follows:

ZnO + 2H⁺ = Zn²⁺ + H₂O

2 Fe²⁺+MnO₂+ 4H⁺=2 Fe³⁺+Mn²⁺+ 2H₂O

Me³⁺+ 3OH^- = Me(OH)₃↓ (Me: Fe, Al, etc)

(3) Pressure leaching in the I section: mixing the normal pressure neutralization leaching residue produced in the step (2), calcium lignosulfonate or sodium lignosulfonate with the step (4)Mixing the leaching solution subjected to the stage pressurization with the washing water obtained in the step (5), and carrying out the stepStage pressure leaching, controlling the pressure in the kettle to be 0.3 toThe pressure leaching process comprises the following steps of 1.4MPa, leaching temperature higher than 100 ℃, end-point reaction acidity of 25-45 g/L, liquid-solid separation after reaction, obtaining I-section pressure leaching liquid and I-section pressure leaching underflow, wherein the iron content of the I-section pressure leaching liquid is lower than 2g/L, and the I-section pressure leaching liquid returns to the normal pressure neutralization leaching process. The concentration of sulfuric acid in the I-stage pressurized leaching solution is 25-45 g/L, the concentration of iron ions is 0.4-2 g/L, and the concentration of zinc ions is 70-90 g/L.

The main function of the I-stage pressure leaching process is to realize the Fe in the leaching system in the range of the operation conditions²⁺Oxidation and Fe³⁺And partial leaching of valuable elements in the leaching material, the main chemical reactions are as follows:

4Fe²⁺+O₂+ 4H⁺= 4Fe³⁺+ 2H₂O

2Fe³⁺+2SO₄ ^2-+2H₂O= 2FeOHSO₄↓+2H⁺

2M⁺+6Fe³⁺+4SO₄ ^2- +12H₂O= 2MFe₃(SO₄)₂(OH)₆↓+12H⁺(M: K⁺、Na⁺、H₃O⁺etc.)

ZnFe₂O₄+8H⁺= Zn²⁺ + 2Fe³⁺+ 4H₂O

（4）And (3) pressure leaching: mixing the I-stage pressure leaching underflow produced in the step (3) with calcium lignosulfonate or sodium lignosulfonate and zinc electrodeposition waste liquid, and carrying out the operation in a pressure kettle with a stirring devicePerforming pressure leaching in a section, controlling the pressure in the kettle to be 0.3-1.3 MPa, the reaction temperature to be more than 100 ℃, and the final reaction acidity to be 50-90 g/L, and performing liquid-solid separation on reaction ore pulp after the reaction is finished to obtainA stage-pressurized leach liquor andthe slag is leached out by pressure in a section,and returning the section pressure leaching liquid to the section I pressure leaching process.The concentration of sulfuric acid in the section pressure leaching solution is 50-90 g/L, the concentration of iron ions is 4-6 g/L, and the concentration of zinc ions is 50-70 g/L.

The main purpose of the section pressure leaching is to realize the high-efficiency leaching of valuable metal elements in materials within the range of the operation conditions, and the chemical reaction mainly generated in the working procedure is as follows:

4Fe²⁺+O₂+ 4H⁺= 4Fe³⁺+ 2H₂O

2MeS+O₂+ 4H⁺=2Me²⁺+ 2S⁰+2H₂o (Me: Zn, Fe, Cd, Cu, etc.)

ZnFe₂O₄ +8H⁺= Zn²⁺ + 2Fe³⁺+ 4H₂O

By the above-mentioned stage I pressure leaching andthe two-stage pressure countercurrent leaching process of the stage pressure leaching realizes the high-efficiency leaching of valuable metals and the high-efficiency precipitation of iron at the same time to the maximum extent.

(5) Pulping and washing leached residues: obtained in step (4)Pulping and washing the leaching residue and a weakly acidic solution with the pH = 1.5-3.5 in a normal-pressure stirring reaction tank at 30-80 ℃, pulping and washing,valuable metals in the stage pressure leaching slag are leached. And then carrying out liquid-solid separation to obtain washing water and lead-silver-iron slag. The washing water returns to the I-stage pressure leaching process, the iron in the obtained lead-silver-iron slag is mainly sulfate iron, and the lead-silver-iron slag contains zinc<2%, 18-30% of iron, 2-12% of lead and 0.02-0.06% of silver.

The lead and the silver in the lead-silver-iron slag can be recovered by adopting two processes, and the harmless treatment of iron is realized: the lead, silver and iron slag directly enters a pyrometallurgical lead smelting system to recover lead and silver in the lead, and iron is finally stably solidified in slag as a slagging agent to realize conversion from impurities to lead smelting raw materials; or firstly recovering silver in the lead-silver-iron slag by adopting a flotation technology to obtain lead-iron slag, then recovering lead in the lead-iron slag by utilizing a pyrogenic process lead smelting or lead enrichment technology, and finally stably solidifying iron serving as a slagging flux in furnace slag or kiln slag in a lead enrichment process to realize the conversion from impurities to lead smelting raw materials.

Wherein the manganese-containing material in the step (2) is one or a mixture of two of pyrolusite and anode mud generated in the zinc electrodeposition process, and the addition amount of the pyrolusite and the anode mud is obtained by neutralizing Mn in the leachate at normal pressure²⁺Determining the concentration, controlling the Mn in the normal pressure neutralization leaching solution²⁺The concentration is 4-6 g/L.

Example 1

A method for treating a zinc-containing material in the production of zinc hydrometallurgy comprises the following specific steps:

(1) mixing zinc roasted ore (the mass percentage of main elements is 56.4 percent of zinc, 6.0 percent of iron, 3.0 percent of lead and 0.01 percent of silver) with water, and then mechanically activating the mixture in a wet ball mill to obtain finely ground ore pulp with the zinc roasted ore granularity of 100 meshes;

(2) mixing the finely ground ore pulp produced in the step (1) with pyrolusite, zinc electrodeposition waste liquid and I-section pressurized leachate in a No. 1 agitation leaching tank through a metering pump, carrying out four-stage normal pressure neutralization leaching in serial No. 2, No. 3, No. 4 and No. 5 agitation leaching tanks, overflowing the reaction ore pulp into the No. 2 to No. 5 agitation leaching tanks from the No. 1 agitation leaching tank in sequence through a chute, controlling the reaction temperature of the No. 1 to No. 5 agitation leaching tanks to be 60 ℃, 65 ℃ and 2 to No. 5 agitation leaching tanks in sequenceThe pH of the inner ore pulp is 2.0-3.5, 3.5-4.5, 4.5-5.0 and 5.2-5.4 in sequence, and the retention time of the reaction ore pulp in the stirring leaching tank from 1# to 5# is 120 min. After the reaction is finished, the ore pulp is subjected to liquid-solid separation by a thickener and a filter press to obtain the Mn-containing ore pulp²⁺The normal pressure neutralization leaching solution and the normal pressure neutralization leaching slag with the concentration of 4 g/L;

(3) neutralizing the leaching residue produced in the step (2) under normal pressure, calcium lignosulphonate and the acid content of 60g/LMixing the leaching solution and washing water under pressure in a pressure kettle with a stirring deviceAnd (3) performing stage pressure leaching, controlling the reaction temperature to be 170 ℃, the pressure in the kettle to be 1.4MPa, and the reaction time to be 120min, standing and layering the reaction ore pulp in a thickener after the reaction is finished, and obtaining the I-stage pressure leaching liquid and the I-stage pressure leaching underflow, wherein the concentration of iron ions is 2g/L, and the final acidity is 45 g/L.

(4) Mixing the I-stage pressure leaching underflow, calcium lignosulfonate and zinc electrodeposition waste liquid produced in the step (3), and carrying out the steps in a pressure kettle with a stirring devicePerforming pressure leaching in sections, controlling the leaching temperature in the process to be 120 ℃ and the pressure in the kettle to be 0.3MPa, reacting for 90min, and performing liquid-solid separation by using a thickener and a filter press to obtain the product with the iron ion concentration of 6g/L and the terminal acidity of 90g/LThe leaching solution is pressurized in a section way,and (5) leaching slag under pressure.

(5) Obtained in step (4)Stage pressure leaching slag and weak acidMixing the aqueous solutions, pulping and washing in a normal-pressure stirring reaction tank, controlling the temperature of ore pulp at 60 ℃, washing for 30min, and then carrying out liquid-solid separation to obtain washing water and lead-silver-iron slag (the mass percentage of main elements is 0.45 percent of zinc, 18 percent of iron, 12 percent of lead and 0.04 percent of silver). The washing water returns to the I-section pressure leaching process, the lead-silver-iron slag directly enters the modern pyrometallurgical lead smelting process to comprehensively recover lead and silver in the lead-silver-iron slag, and iron in the lead-silver-iron slag is used as a pyrometallurgical lead smelting slagging flux and is finally stably solidified in the slag, so that the conversion of impurities into lead smelting raw materials is realized.

After the zinc roasted ore is treated by the method, the leaching rate of zinc in the whole process is close to 99%, the concentration of iron ions in the I-stage pressurized leachate is only 2g/L, and the leachate can directly enter a zinc hydrometallurgy purification-electrodeposition system without further iron removal.

Example 2

A method for treating a zinc-containing material in the production of zinc hydrometallurgy comprises the following specific steps:

(1) the zinc roasted ore (the mass percentage of main elements are 45.0 percent of zinc, 9.2 percent of iron, 1.6 percent of lead and 0.005 percent of silver) and water are mixed into slurry and then are mechanically activated in a wet ball mill, and the finely ground ore slurry with the zinc roasted ore granularity of 150 meshes is obtained.

(2) And (2) mixing the finely ground ore pulp produced in the step (1) with pyrolusite, anode slime, zinc electrodeposition waste liquid and I-section pressurized leachate in a No. 1 stirring leaching tank through a metering pump, performing four-stage normal-pressure neutralization leaching in serial No. 2, No. 3, No. 4 and No. 5 stirring leaching tanks, overflowing the reaction ore pulp into the No. 2 to No. 5 stirring leaching tanks from the No. 1 stirring tank through a chute in sequence, controlling the reaction temperature of the No. 1 to No. 5 stirring leaching tanks to be 85 ℃, 80 ℃, and the pH of the ore pulp in the No. 2 to No. 5 stirring leaching tanks to be 3.5 to 4.5, 4.5 to 5.0, 5.0 to 5.2 and 5.2 to 5.4 in sequence, and keeping the reaction ore pulp in the No. 1 to No. 5 stirring leaching tanks for 90 min. After the reaction is finished, the ore pulp is subjected to liquid-solid separation by a thickener and a filter press to obtain the Mn-containing ore pulp²⁺The normal pressure neutralization leaching liquid with the concentration of 6g/L and the normal pressure neutralization leaching slag.

(3) Neutralizing the leaching residue produced in the step (2) under normal pressure, sodium lignosulfonate and acidity of 46g/LMixing the leaching solution and washing water under pressure in a pressure kettle with a stirring deviceAnd (3) performing stage pressure leaching, controlling the reaction temperature to be 155 ℃, the pressure in the kettle to be 1.2MPa, and the reaction time to be 180min, standing and layering the reaction ore pulp in a thickener after the reaction is finished, and obtaining the I-stage pressure leaching liquid and the I-stage pressure leaching underflow, wherein the iron ion concentration is 1.2g/L, and the end-point acidity is 38 g/L.

(4) Mixing the I-stage pressure leaching underflow, sodium lignosulfonate and zinc electrodeposition waste liquid produced in the step (3), and carrying out the steps in a pressure kettle with a stirring devicePerforming pressure leaching in a section, controlling the leaching temperature in the process to be 140 ℃ and the pressure in the kettle to be 1.15MPa, reacting for 60min, and performing liquid-solid separation on the reaction ore pulp by using a thickener and a filter press to obtain the product with the iron ion concentration of 5.5g/L and the end-point acidity of 70g/LA stage-pressurized leach liquor andand (5) leaching slag under pressure.

(5) Obtained in step (4)Mixing the leaching residue under pressure with a weakly acidic solution, pulping and washing in a normal-pressure stirring reaction tank, controlling the temperature of ore pulp at 30 ℃, and performing liquid-solid separation after washing for 40min to obtain washing water and lead-silver-iron residue (the mass percentage of main elements is 0.8% of zinc, 25.4% of iron, 7.1% of lead and 0.02% of silver). The washing water returns to the I-stage pressure leaching process, firstly, the flotation technology is adopted to recover the silver in the lead-silver-iron slag to obtain lead-iron slag, then, the pyrometallurgical lead smelting technology is adopted to recover the lead in the lead-iron slag, and the iron is used as a slagging flux to finally stabilize and solidifyThe lead-containing slag is converted into the lead smelting raw material from impurities.

After the zinc roasted ore is treated by the method, the leaching rate of zinc in the whole process is close to 99%, the concentration of iron ions in the I-stage pressurized leachate is only 1.2g/L, and the solution can directly enter a zinc hydrometallurgy purification-electrodeposition system without further iron removal.

Example 3

A method for treating a zinc-containing material in the production of zinc hydrometallurgy comprises the following specific steps:

(1) the zinc roasted ore (the mass percentage of main elements are zinc 60.0%, iron 12.4%, lead 0.5% and silver 0.015%) and water are mixed and then are mechanically activated in a wet ball mill, and the finely ground ore pulp with the zinc roasted ore granularity of 200 meshes is obtained.

(2) And (2) mixing the finely ground ore pulp produced in the step (1) with anode mud, zinc electrodeposition waste liquid and I-section pressurized leachate in a No. 1 stirring leaching tank through a metering pump, carrying out four-stage normal-pressure neutralization leaching in serial No. 2, No. 3, No. 4 and No. 5 stirring leaching tanks, overflowing the reaction ore pulp into the No. 2 to No. 5 stirring leaching tanks from the No. 1 stirring tank through a chute in sequence, controlling the reaction temperature of the No. 1 to No. 5 stirring tanks to be 75 ℃, 75 ℃ and 75 ℃, controlling the pH of the ore pulp in the No. 2 to No. 5 stirring leaching tanks to be 4.5, 5.0, 5.2 to 5.4 and 5.2 to 5.4 in sequence, and keeping the reaction ore pulp in the No. 1 to No. 5 stirring tanks for 90 min. After the reaction is finished, the ore pulp is subjected to liquid-solid separation by a thickener and a filter press to obtain the Mn-containing ore pulp²⁺The normal pressure neutralization leaching solution with the concentration of 5g/L and the normal pressure neutralization leaching slag.

(3) Neutralizing the leaching residue produced in the step (2) under normal pressure, calcium lignosulfonate and the acid content of 30g/LMixing the leaching solution and washing water under pressure in a pressure kettle with a stirring devicePerforming pressure leaching in a section, controlling the reaction temperature at 110 ℃, the pressure in the kettle at 0.3MPa, and the reaction time at 150min, and after the reaction is finished, standing the reaction ore pulp in a thickenerLayering to obtain I-section pressure leaching liquid with iron ion concentration of 0.4g/L and end point acidity of 25g/L and I-section pressure leaching underflow.

(4) Mixing the I-stage pressure leaching underflow, calcium lignosulfonate and zinc electrodeposition waste liquid produced in the step (3), and carrying out the steps in a pressure kettle with a stirring devicePerforming pressure leaching in a section, controlling the leaching temperature in the process to be 160 ℃ and the pressure in the kettle to be 1.3MPa, reacting for 180min, and performing liquid-solid separation on the reaction ore pulp by using a thickener and a filter press to obtain the product with the iron ion concentration of 4g/L and the terminal acidity of 50g/LA stage-pressurized leach liquor andand (5) leaching slag under pressure.

(5) Obtained in step (4)The leaching residue under pressure is mixed with a weakly acidic solution, slurrying and washing are carried out in a normal-pressure stirring reaction tank, the temperature of ore pulp is controlled to be 80 ℃, liquid-solid separation is carried out after washing is carried out for 60min, and washing water and lead-silver-iron residue (the mass percentage of main elements (dry basis) is 1.5 percent of zinc, 28.2 percent of iron, 2.0 percent of lead and 0.06 percent of silver) are obtained. The washing water returns to the I-section pressure leaching process, firstly, the flotation technology is adopted to recover the silver in the lead-silver-iron slag to obtain the lead-iron slag, then, the lead enrichment technology is utilized to recover the lead in the lead-iron slag, and finally, the iron is stably solidified in the kiln slag to realize the conversion from impurities to lead smelting raw materials.

After the zinc roasted ore is treated by the method, the leaching rate of zinc in the whole process is close to 98%, the concentration of iron ions in the I-stage pressurized leachate is only 0.4g/L, and the solution can directly enter a zinc hydrometallurgy purification-electrodeposition system without further iron removal.

Comparative example 3 (same as example 3 except that no pressure was applied during leaching)

A method for treating a zinc-containing material in the production of zinc hydrometallurgy comprises the following specific steps:

(1) the zinc roasted ore (the mass percentage of main elements are zinc 60.0%, iron 12.4%, lead 0.5% and silver 0.015%) and water are mixed and then are mechanically activated in a wet ball mill, and the finely ground ore pulp with the zinc roasted ore granularity of 200 meshes is obtained.

(2) And (2) mixing the finely ground ore pulp produced in the step (1) with anode mud, zinc electrodeposition waste liquid and I-section pressurized leachate in a No. 1 stirring leaching tank through a metering pump, carrying out four-stage normal-pressure neutralization leaching in serial No. 2, No. 3, No. 4 and No. 5 stirring leaching tanks, overflowing the reaction ore pulp into the No. 2 to No. 5 stirring leaching tanks from the No. 1 stirring tank through a chute in sequence, controlling the reaction temperature of the No. 1 to No. 5 stirring tanks to be 75 ℃, 75 ℃ and 75 ℃, controlling the pH of the ore pulp in the No. 2 to No. 5 stirring leaching tanks to be 4.5, 5.0, 5.2 to 5.4 and 5.2 to 5.4 in sequence, and keeping the reaction ore pulp in the No. 1 to No. 5 stirring tanks for 90 min. After the reaction is finished, the ore pulp is subjected to liquid-solid separation by a thickener and a filter press to obtain the Mn-containing ore pulp²⁺The normal pressure neutralization leaching solution with the concentration of 5g/L and the normal pressure neutralization leaching slag.

(3) Neutralizing the leaching residue at normal pressure, calcium lignosulphonate and the like which are produced in the step (2),Mixing the normal pressure leaching liquid and washing water to obtain mixed acid liquid, and stirring in a stirrerAnd (3) leaching under normal pressure in the section, controlling the reaction temperature to be 90 ℃ and the reaction time to be 150min, standing and layering the reaction ore pulp in a thickener after the reaction is finished, and obtaining the I-section normal pressure leaching liquid and the I-section normal pressure leaching underflow, wherein the concentration of iron ions is 18g/L and the terminal acidity is 20 g/L.

(4) Mixing the I-stage atmospheric leaching underflow, calcium lignosulfonate and zinc electrodeposition waste liquid produced in the step (3), and carrying out the steps in equipment with a stirring deviceStaged atmospheric leachingControlling the leaching temperature in the process to be 90 ℃, reacting for 180min, and then carrying out liquid-solid separation on the reaction ore pulp by using a thickener and a filter press to obtain the product with the iron ion concentration of 12 g/L and the terminal acidity of 48 g/LLeaching solution under normal pressure andand leaching the slag at normal pressure.

(5) Obtained in step (4)Mixing the normal-pressure leaching residue with a weakly acidic solution, pulping and washing in a normal-pressure stirring reaction tank, controlling the temperature of ore pulp to be 80 ℃, washing for 60min, and then carrying out liquid-solid separation to obtain washing water and lead-silver residue (the mass percentage of main elements is (dry basis): zinc 4%, iron 3%, lead 2.0%, and silver 0.06%). The washing water returns to the stage I normal pressure leaching process, the lead-silver slag enters the modern pyrometallurgical lead smelting process to comprehensively recover lead and silver in the lead-silver slag, and materials such as pyrite and the like are added in the pyrometallurgical treatment process of the lead-silver slag to serve as slagging iron sources.

After the zinc roasted ore is treated by the method, the leaching rate of zinc in the whole process is close to 96%, the concentration of iron ions in the I-stage normal-pressure leaching liquid is 18g/L, and further neutralization and iron precipitation treatment are needed.

If normal pressure leaching is adopted, more than 85% of iron in the zinc roasted ore is dissolved and enters the I-stage normal pressure leaching liquid, the I-stage normal pressure leaching liquid needs further neutralization iron precipitation treatment, and the produced dangerous waste iron precipitation slag (jarosite slag or goethite slag) needs harmless treatment; and obtainedThe iron content in the normal-pressure leaching slag, namely the lead-silver slag, is low, and an additional iron source (usually pyrite) needs to be supplemented when the lead-silver slag enters the process of lead smelting by a pyrogenic process for slagging. When the two-stage pressure leaching process is adopted, the iron enters the lead-silver-iron slag, and when the two-stage atmospheric pressure leaching process is adopted, the iron enters the leaching solution, so that the two-stage pressure leaching process is obviously reducedThe treatment cost of the low-zinc roasted ore is favorable for comprehensively and efficiently recovering the associated valuable metals in the zinc roasted ore.

Example 4

A method for treating a zinc-containing material in the production of zinc hydrometallurgy comprises the following specific steps:

(1) mechanically activating zinc roasted ore (the mass percentage of main elements are 52.9 percent of zinc, 16.0 percent of iron, 2.2 percent of lead and 0.008 percent of silver) in a dry ball mill to obtain fine ground material with the zinc roasted ore granularity of 150 meshes.

(2) And (2) mixing the finely ground zinc roasted ore produced in the step (1) with pyrolusite, zinc electrodeposition waste liquid and I-section pressurized leachate in a No. 1 stirring leaching tank through a metering pump, carrying out four-stage normal-pressure neutralization leaching in serial No. 2, No. 3, No. 4 and No. 5 stirring leaching tanks, overflowing the reaction ore pulp into the No. 2 to No. 5 stirring leaching tanks from the No. 1 stirring tank through a chute in sequence, controlling the reaction temperature of the No. 1 to No. 5 stirring leaching tanks to be 72 ℃, 70 ℃, and the pH value of the ore pulp in the No. 2 to No. 5 stirring leaching tanks to be 5.0 to 5.2, 5.0 to 5.2 and 5.2 to 5.4 in sequence, and keeping the reaction ore pulp in the No. 1 to No. 5 stirring leaching tanks for 140 min. After the reaction is finished, the ore pulp is subjected to liquid-solid separation by a thickener and a filter press to obtain the Mn-containing ore pulp²⁺The normal pressure neutralization leaching solution with the concentration of 5g/L and the normal pressure neutralization leaching slag.

(3) Neutralizing the leaching residue produced in the step (2) under normal pressure, calcium lignosulphonate and the acid content of 40g/LMixing the leaching solution and washing water under pressure in a pressure kettle with a stirring deviceAnd (3) performing stage pressure leaching, controlling the reaction temperature to be 145 ℃, the pressure in the kettle to be 0.6MPa, and the reaction time to be 60min, standing and layering the reaction ore pulp in a thickener after the reaction is finished, and obtaining the I-stage pressure leaching liquid and the I-stage pressure leaching underflow, wherein the iron ion concentration is 0.8g/L, and the end-point acidity is 30 g/L.

(4) Pressurizing and leaching the section I produced in the step (3) to obtain a bottomMixing the waste liquid of calcium lignosulfonate and zinc electrodeposition in a pressure kettle with a stirring devicePerforming pressure leaching in a section, controlling the leaching temperature in the process to be 155 ℃ and the pressure in the kettle to be 0.85MPa, reacting for 150min, and performing liquid-solid separation on the reaction ore pulp by using a thickener and a filter press to obtain the product with the iron ion concentration of 5.0g/L and the end-point acidity of 56g/LA stage-pressurized leach liquor andand (5) leaching slag under pressure.

(5) Obtained in step (4)Mixing the leaching residue under pressure with a weakly acidic solution, pulping and washing in a normal-pressure stirring reaction tank, controlling the temperature of ore pulp at 50 ℃, and performing liquid-solid separation after washing for 50min to obtain washing water and lead-silver-iron residue (the mass percentage of main elements is 0.9% of zinc, 30.0% of iron, 9.6% of lead and 0.034% of silver). The washing water returns to the I-section pressure leaching process, firstly, the flotation technology is adopted to recover the silver in the lead-silver-iron slag to obtain the lead-iron slag, then, the pyrometallurgical lead smelting technology is adopted to recover the lead in the lead-iron slag, and the iron is used as a slagging flux and is finally stably solidified in the slag to realize the conversion from impurities to lead smelting raw materials.

After the zinc roasted ore is treated by the method, the leaching rate of zinc in the whole process is close to 99%, the concentration of iron ions in the I-stage pressurized leachate is only 0.8g/L, and the solution can directly enter a zinc hydrometallurgy purification-electrodeposition system without further iron removal.

Comparative example 4

A method for treating a zinc-containing material in the production of zinc hydrometallurgy comprises the following specific steps:

(1) mechanically activating zinc roasted ore (the mass percentage of main elements are 52.9 percent of zinc, 16.0 percent of iron, 2.2 percent of lead and 0.008 percent of silver) in a dry ball mill to obtain fine ground material with the zinc roasted ore granularity of 150 meshes.

(2) And (2) mixing the finely ground zinc roasted ore produced in the step (1) with pyrolusite, zinc electrodeposition waste liquid and I-section pressurized leachate in a No. 1 stirring leaching tank through a metering pump, carrying out four-stage normal-pressure neutralization leaching in serial No. 2, No. 3, No. 4 and No. 5 stirring leaching tanks, overflowing the reaction ore pulp into the No. 2 to No. 5 stirring leaching tanks from the No. 1 stirring tank through a chute in sequence, controlling the reaction temperature of the No. 1 to No. 5 stirring leaching tanks to be 72 ℃, 70 ℃, and the pH value of the ore pulp in the No. 2 to No. 5 stirring leaching tanks to be 5.0 to 5.2, 5.0 to 5.2 and 5.2 to 5.4 in sequence, and keeping the reaction ore pulp in the No. 1 to No. 5 stirring leaching tanks for 140 min. After the reaction is finished, the ore pulp is subjected to liquid-solid separation by a thickener and a filter press to obtain the Mn-containing ore pulp²⁺The normal pressure neutralization leaching solution with the concentration of 5g/L and the normal pressure neutralization leaching slag.

(3) And (3) mixing the normal-pressure neutralized leaching residue, calcium lignosulfonate and mixed acid liquor produced in the step (2) to obtain slurry, performing pressure leaching in a pressure kettle with a stirring device, controlling the reaction temperature to be 145 ℃, the pressure in the kettle to be 0.6MPa, and the reaction time to be 240min, standing and layering the reaction slurry in a thickener after the reaction is finished, and obtaining the pressure leaching solution and the pressure leaching residue with the iron ion concentration of 0.8g/L and the terminal acidity of 20 g/L.

(4) And (3) mixing the pressure leaching residue produced in the step (3) with a weakly acidic solution, pulping and washing in a normal-pressure stirring reaction tank, controlling the temperature of ore pulp to be 50 ℃, and performing liquid-solid separation after washing for 50min to obtain washing water and lead-silver-iron residue (the mass percentage of main elements is 12.5 percent of zinc, 19 percent of iron, 3.4 percent of lead and 0.012 percent of silver). The washing water returns to the I-section pressure leaching process, firstly, the flotation technology is adopted to recover the silver in the lead-silver-iron slag to obtain the lead-iron slag, then, the lead enrichment technology is utilized to recover the lead in the lead-iron slag, and finally, the iron is stably solidified in the kiln slag to realize the conversion from impurities to lead smelting raw materials.

After the zinc roasted ore is treated by the method, the leaching rate of zinc in the whole process is close to 84%, and the iron ion concentration of the I-stage pressurized leaching liquid is 0.8/L.

If only one-stage pressure leaching is adopted, the leaching rate of valuable metals such as zinc and the like is obviously reduced, and the step I is added in the inventionThe main function of pressure leaching is that the iron precipitation and part of valuable metals are leached,the function of the section pressure leaching is to realize the deep leaching of valuable metals by increasing the temperature and the acidity of a leaching system.

Example 5

A method for treating a zinc-containing material in the production of zinc hydrometallurgy comprises the following specific steps:

(1) 200kg of zinc roasted ore (the mass percentage of main elements is 53.0 percent of zinc, 7.5 percent of iron, 1.2 percent of lead and 0.009 percent of silver) and 300L of water are mixed according to the solid-to-liquid ratio of 1.5: 1 (kg/L) of size mixing and then carrying out mechanical activation in a wet ball mill to obtain the finely ground ore pulp with the zinc roasted ore granularity of 150 meshes.

(2) And (2) mixing the finely ground ore pulp produced in the step (1) with pyrolusite, anode mud, zinc electrodeposition waste liquid and I-section pressurized leachate in a No. 1 stirring leaching tank through a metering pump, performing four-stage normal-pressure neutralization leaching in serial No. 2, No. 3, No. 4 and No. 5 stirring leaching tanks, overflowing the reaction ore pulp into the No. 2 to No. 5 stirring leaching tanks from the No. 1 stirring tank through a chute in sequence, controlling the reaction temperature of the No. 1 to No. 5 stirring leaching tanks to be 75 ℃, and pH of the ore pulp in the No. 2 to No. 5 stirring leaching tanks to be 4.0 to 4.5, 4.5 to 5.2, 5.2 to 5.4 and 5.2 to 5.4 in sequence, and staying the reaction ore pulp in the No. 1 to No. 5 stirring leaching tanks for 120 min. After the reaction is finished, the ore pulp is subjected to liquid-solid separation by a thickener and a filter press to obtain the Mn-containing ore pulp²⁺The normal pressure neutralization leaching solution with the concentration of 5g/L and the normal pressure neutralization leaching slag.

(3) Neutralizing the leaching residue produced in the step (2) under normal pressure, calcium lignosulphonate and the acid content of 40g/LMixing the leaching solution and washing water under pressure in a pressure kettle with a stirring devicePressure leaching in section, controlling reaction temperature at 145 deg.C and pressure in kettle at 0.9MPa, and reactingThe reaction time is 90min, and after the reaction is finished, the reaction ore pulp is kept stand and layered in a thickener to obtain I-section pressurized leach liquor and I-section pressurized leach underflow, wherein the concentration of iron ions is 1.0g/L, and the final acidity is 35 g/L.

(4) Mixing the I-stage pressure leaching underflow, calcium lignosulfonate and zinc electrodeposition waste liquid produced in the step (3), and carrying out the steps in a pressure kettle with a stirring devicePerforming pressure leaching in a section, controlling the leaching temperature in the process to be 155 ℃ and the pressure in the kettle to be 0.75MPa, reacting for 120min, and performing liquid-solid separation on the reaction ore pulp by using a thickener and a filter press to obtain the product with the iron ion concentration of 5.0g/L and the end-point acidity of 75g/LA stage-pressurized leach liquor andand (5) leaching slag under pressure.

(5) Obtained in step (4)Mixing the leaching residue under pressure with a weakly acidic solution, pulping and washing in a normal-pressure stirring reaction tank, controlling the temperature of ore pulp to 65 ℃, washing for 30min, and then carrying out liquid-solid separation to obtain washing water and lead-silver-iron residue (the mass percentage of main elements is 0.8% of zinc, 25.6% of iron, 4.4% of lead and 0.03% of silver). The washing water returns to the I-section pressure leaching process, firstly, the flotation technology is adopted to recover the silver in the lead-silver-iron slag to obtain the lead-iron slag, then, the lead enrichment technology is utilized to recover the lead in the lead-iron slag, and finally, the iron is stably solidified in the kiln slag to realize the conversion from impurities to lead smelting raw materials.

After the zinc roasted ore is treated by the method, the leaching rate of zinc in the whole process is close to 99%, the concentration of iron ions in the I-stage pressurized leachate is only 1.0g/L, and the solution can directly enter a zinc hydrometallurgy purification-electrodeposition system without further iron removal.

Comparative example 5

The same zinc roasted ore as in example 5 was used and compared using the existing industrial treatment process, namely zinc roasted ore-neutral leaching-weak acid leaching-hot acid leaching-jarosite process iron precipitation.

Examples of the process of neutral leaching of zinc roasted ore, weak acid leaching, hot acid leaching and iron precipitation by jarosite method

(1) The zinc roasted ore (the mass percentage of main elements are 53.0 percent of zinc, 7.5 percent of iron, 1.2 percent of lead and 0.009 percent of silver) and water are mixed into slurry and then are mechanically activated in a wet ball mill, and the finely ground ore slurry with the zinc roasted ore granularity of 150 meshes is obtained.

(2) Mixing the fine ground ore pulp produced in the step (1) with pyrolusite, anode slime, a high-iron solution (sulfuric acid concentration is 80g/L and iron content is 4 g/L), zinc electrodeposition waste liquid and 366L of a weak acid leaching solution (sulfuric acid concentration is 0.5 g/L) in a No. 1 stirring leaching tank, carrying out four-stage normal-pressure neutralization leaching in serial No. 2, No. 3, No. 4 and No. 5 stirring leaching tanks, overflowing the reaction ore pulp into the No. 2 to No. 5 stirring leaching tanks from the No. 1 stirring tank in sequence through a chute, controlling the reaction temperature of the No. 1 to No. 5 stirring leaching tanks to be 75 ℃, and 5.2 to 5.5, pH of the ore pulp in the No. 2 to No. 5 stirring leaching tanks to be 4.0 to 4.5, 4.5 to 5.2 to 5.4 in sequence, and staying for 120min in the No. 1 to No. 5 stirring leaching tanks. After the reaction is finished, the ore pulp is clarified and separated by a thickener to obtain Mn²⁺The concentration of the atmospheric neutralization leaching liquid is 5g/L, and the bottom flow of the atmospheric neutralization leaching liquid.

(3) Uniformly adding the weak acid leaching underflow and zinc electrodeposition waste liquid produced in the step (2) into a No. 1 stirring weak acid leaching tank, performing four-stage weak acid leaching in the No. 1, No. 2, No. 3 and No. 4 stirring weak acid leaching tanks which are connected in series, overflowing reaction ore pulp into the No. 2 to No. 4 stirring leaching tanks from the No. 1 stirring tank in sequence through a chute, controlling the reaction temperature of the No. 1 to No. 4 stirring leaching tanks to be 80 ℃, controlling the pH of the ore pulp in the No. 1 to No. 4 stirring leaching tanks to be 2.5, and staying the reaction ore pulp in the No. 1 to No. 4 stirring leaching tanks for 160 min. After the reaction is finished, the ore pulp is subjected to liquid-solid separation by a thickener and a filter press to obtain weak acid leaching underflow and weak acid leaching liquid with the iron content of 2 g/L. The weak acid leaching liquid returns to normal pressure for neutralization leaching, and the bottom flow of the weak acid leaching enters a hot acid leaching process.

(4) And (4) uniformly adding the weak acid leaching underflow and the zinc electrodeposition waste liquid produced in the step (3) into a hot acid leaching and stirring weak acid leaching tank for hot acid leaching, controlling the reaction temperature in the process to be 85 ℃, and keeping the reaction ore pulp in the stirring leaching tank for 180 min. After the reaction is finished, the ore pulp is subjected to liquid-solid separation by a thickener and a filter press to obtain hot acid leaching slag (the mass percentage of main elements is 6.5 percent of zinc, 6.4 percent of iron, 6.8 percent of lead and 0.05 percent of silver) and hot acid leaching liquid (the main components are that the concentration of iron ions is 19.6g/L and the concentration of zinc ions is 110 g/L). The hot acid leaching slag enters a pyrogenic process lead smelting system to recover lead and silver or directly carry out harmless treatment, and the hot acid leaching liquid enters a jarosite process iron removal process.

(5) Continuously adding zinc roasted ore into the hot acid leaching solution produced in the step (4), maintaining the pH value of the process to be 1.5-2.0, carrying out jarosite method iron precipitation, controlling the reaction temperature of the process to be 90 ℃, reacting for 120-180 min, carrying out liquid-solid separation on the ore pulp when the concentration of iron ions in the solution is less than 2g/L to obtain jarosite slag (the main components are 6% of zinc, 29% of iron, 1.5% of lead and 0.016% of silver) and iron-removing liquid with the concentration of iron ions of 1.2g/L, carrying out high-temperature curing and harmless treatment on the jarosite slag, and returning the iron-removing liquid to the neutral leaching process.

The leaching rate of zinc in the whole process after the zinc-germanium-containing leaching residue is treated by adopting a zinc roasted ore neutral leaching-weak acid leaching-hot acid leaching-jarosite method iron precipitation process is 96%, the concentration of iron ions in a hot acid leaching solution is 19.6g/L, 370kg jarosite residue is generated when the iron is precipitated by adopting the jarosite method, the dangerous waste iron residue needs to be further subjected to harmless treatment, and the hazardous waste iron residue is rich in zinc, lead, silver and the like and is difficult to recover.

Example 6

A method for treating a zinc-containing material in the production of zinc hydrometallurgy comprises the following specific steps:

(1) acid leaching residues (the mass percentage of main elements is 18 percent of zinc, 15.5 percent of iron, 3.9 percent of lead and 0.02 percent of silver) are mixed with water to be pulped and then are mechanically activated in a wet ball mill, and fine grinding ore pulp with the granularity of 150 meshes is obtained;

(2) mixing the fine ground ore pulp and the lignin sulfonic acid produced in the step (1)Calcium carbonate, acid of 75g/LMixing the leaching solution and washing water under pressure in a pressure kettle with a stirring deviceAnd (3) performing stage pressure leaching, controlling the reaction temperature to be 155 ℃, the pressure in the kettle to be 0.8MPa, and the reaction time to be 90min, standing and layering the reaction ore pulp in a thickener after the reaction is finished, and obtaining the I-stage pressure leaching liquid and the I-stage pressure leaching underflow, wherein the concentration of iron ions is 1.3g/L, and the final acidity is 42 g/L.

(4) Mixing the I-stage pressure leaching underflow, calcium lignosulfonate and zinc electrodeposition waste liquid produced in the step (2), and carrying out the steps in a pressure kettle with a stirring devicePerforming pressure leaching in a section, controlling the leaching temperature in the process to be 145 ℃ and the pressure in the kettle to be 0.8MPa, performing liquid-solid separation on reaction ore pulp by a thickener and a filter press after reacting for 90min to obtain the product with the iron ion concentration of 7.5 g/L and the end-point acidity of 70g/LThe leaching solution is pressurized in a section way,and (5) leaching slag under pressure.

(5) Obtained in step (4)Mixing the leaching residue under pressure with a weakly acidic solution, pulping and washing in a normal-pressure stirring reaction tank, controlling the temperature of ore pulp to 65 ℃, washing for 30min, and then carrying out liquid-solid separation to obtain washing water and lead-silver-iron residue (the mass percentage of main elements is 0.8% of zinc, 18% of iron, 7.1% of lead and 0.05% of silver). The washing water returns to the I-section pressure leaching process, the lead-silver-iron slag firstly adopts the flotation technology to recover the silver in the lead-silver-iron slag to obtain the lead-iron slag, and then the lead-iron slag is beneficial toLead in the lead-iron slag is recovered by a lead enrichment technology, and iron is finally stably solidified in the kiln slag to realize the conversion from impurities to lead smelting raw materials.

After the acid leaching residue is treated by the method, the leaching rate of zinc in the whole process is close to 98%, the concentration of iron ions in the I-stage pressurized leaching solution is only 1.3g/L, and the solution can directly enter a zinc hydrometallurgy purification-electrodeposition system without further iron removal.

Example 7

A method for treating a zinc-containing material in the production of zinc hydrometallurgy comprises the following specific steps:

(1) mixing neutral leaching residues (the mass percentage of main elements is 22 percent of zinc, 13 percent of iron, 3.5 percent of lead and 0.02 percent of silver) with water, and then mechanically activating the neutral leaching residues in a wet ball mill to obtain fine ground ore pulp with the granularity of the acidic leaching residues being 200 meshes;

(2) mixing the fine ground ore pulp produced in the step (1), calcium lignosulfonate and the mixture with acidity of 80g/LMixing the leaching solution and washing water under pressure in a pressure kettle with a stirring deviceAnd (3) performing stage pressure leaching, controlling the reaction temperature to be 160 ℃, the pressure in the kettle to be 1.0MPa, and the reaction time to be 120min, standing and layering the reaction ore pulp in a thickener after the reaction is finished, and obtaining the I-stage pressure leaching liquid and the I-stage pressure leaching underflow, wherein the concentration of iron ions is 1.8g/L, and the final acidity is 39 g/L.

(4) Mixing the I-stage pressure leaching underflow, calcium lignosulfonate and zinc electrodeposition waste liquid produced in the step (2), and carrying out the steps in a pressure kettle with a stirring devicePerforming pressure leaching in a section, controlling the leaching temperature in the process to be 155 ℃ and the pressure in the kettle to be 0.9MPa, performing liquid-solid separation on reaction ore pulp after reacting for 120min by using a thickener and a filter press to obtain the product with the iron ion concentration of 8.0 g/L and the end-point acidity of 75g/LThe leaching solution is pressurized in a section way,and (5) leaching slag under pressure.

(5) Obtained in step (4)Mixing the leaching residue under pressure with a weakly acidic solution, pulping and washing in a normal-pressure stirring reaction tank, controlling the temperature of ore pulp at 60 ℃, washing for 30min, and then carrying out liquid-solid separation to obtain washing water and lead-silver-iron residue (the mass percentage of main elements is zinc 1.2%, iron 20%, lead 7.5%, and silver 0.04%). The washing water returns to the I-section pressure leaching process, the lead-silver-iron slag firstly adopts the flotation technology to recover the silver in the lead-silver-iron slag to obtain the lead-iron slag, then the lead enrichment technology is utilized to recover the lead in the lead-iron slag, and the iron is finally stably solidified in the kiln slag to realize the conversion from impurities to lead smelting raw materials.

After the acid leaching residue is treated by the method, the leaching rate of zinc in the whole process is close to 97%, the concentration of iron ions in the I-stage pressurized leaching solution is only 1.8g/L, and the solution can directly enter a zinc hydrometallurgy purification-electrodeposition system without further iron removal.

Finally, it is noted that the above-mentioned preferred embodiments illustrate rather than limit the invention, and that, although the invention has been described in detail with reference to the above-mentioned preferred embodiments, it will be understood by those skilled in the art that various changes in form and detail may be made therein without departing from the scope of the invention as defined by the appended claims.