Method for separating and extracting lead and zinc from lead-zinc sulfide ore powder

Abstract

The invention discloses a method for separating and extracting lead and zinc from lead-zinc sulfide ore powder, and belongs to the technical field of nonferrous metal extraction. The method for separating and extracting lead and zinc from lead-zinc sulfide ore powder comprises the following steps: adding a surfactant and a dilute acid solution into the lead-zinc sulfide ore powder, uniformly mixing, and pouring into a closed reaction kettle; introducing oxygen, heating to a constant temperature, starting stirring, carrying out oxygen pressure leaching, and carrying out liquid-solid separation after a preset time is reached, so as to obtain lead sulfate-based leaching residues and zinc-iron-containing leaching liquid; adding an oxidizing agent into the zinc-iron-containing leaching solution, adding a neutralizing agent for neutralization and hydrolysis after reaction, and removing impurities to obtain a zinc solution; adding the lead sulfate-based leaching residues into an alkaline desulfurizing agent to carry out desulfurization reaction, and leaching in an acid leaching agent to obtain lead-containing leaching liquid; and the lead-containing leaching agent is subjected to electro-deposition, metal lead is obtained, waste electrolyte is returned to the acid leaching agent, and recycling of the leaching agent is achieved. By adopting the method for separating and extracting the lead and the zinc from the lead-zinc sulfide ore powder, efficient and deep separation and recovery of the zinc and the lead in the lead-zinc sulfide ore are realized.

Description

Technical Field

[0001] The invention relates to the technical field of nonferrous metal extraction, in particular to a method for separating and extracting lead and zinc from lead-zinc sulfide ore powder. Background Art

[0002] Lead-zinc sulfide ore is an important polymetallic mineral resource, in which galena (PbS) and sphalerite (ZnS) usually exist in a closely associated form. At present, flotation is mainly used in industry to separate lead and zinc. According to the differences in the surface properties of the minerals, flotation technology can be divided into preferential flotation, mixed flotation, and other flotation processes. However, the existing flotation separation process generally has the following technical bottlenecks: (1) large consumption of reagents and high toxicity; (2) complex process flow; (3) low concentrate yield; (4) high lead-zinc mutual content, etc., which seriously restrict the efficient utilization of resources.

[0003] Lead smelting is primarily categorized into two main methods: pyrometallurgy and hydrometallurgy. Traditional pyrometallurgy utilizes a sintering roasting-blast furnace reduction process, which is subject to high energy consumption and environmental pollution from low-concentration SO2 emissions. While modern lead metallurgical methods have seen improvements, such as bottom-blown bath smelting, flash furnace smelting, and top-blown submerged smelting, smelting temperatures must still be maintained above 1100°C, resulting in high energy consumption and lead vapor pollution. Hydrometallurgical lead extraction technologies primarily include FeCl3 oxidation, galena-pyrolusite leaching, and silicic acid oxidative leaching. The FeCl3 oxidation method presents technical challenges such as the volatility of hydrochloric acid, severe equipment corrosion, and difficulty collecting high-temperature chlorine gas. The silicic acid oxidative leaching method, however, suffers from unstable fluoride, which easily decomposes to produce toxic hydrogen fluoride gas, leading to a harsh operating environment and posing a health risk to production personnel.

[0004] The existing patent CN202410140673.0 discloses a process for processing lead-zinc ores in which complex sulfide ores and carbonate ores coexist, including grinding the raw ore and then performing mixed flotation of sulfide ores, flotation of carbonate ores and self-oxidation pressure leaching, electrolysis, mineral separation and other steps. The above patent first floats the lead-zinc sulfide ore into concentrate and then adopts self-oxidation pressure leaching treatment. Instead of adding acid, the sulfur of the sulfide mineral is oxidized under pressure and oxygen to generate sulfuric acid to extract the metal. It has the advantages of low acidity and low impurities in the leachate. However, in order to obtain high-grade lead ore required for smelting in the above patent, the mixed lead-zinc ore needs to go through a flotation process to separate the sulfide ore and the carbonate ore. A variety of flotation aids such as frothers and collectors are added during the flotation process. There are disadvantages such as complicated and lengthy process flow, large amount of reagents used, and high cost of reagents. Therefore, it is necessary to innovate and upgrade the process and develop new green, low-carbon, clean and efficient new technologies for the separation and recovery of lead and zinc co-existing elements to maximize resource utilization and minimize environmental pollution. Summary of the Invention

[0005] The purpose of the present invention is to provide a method for separating and extracting lead and zinc from lead-zinc sulfide ore powder, thereby achieving efficient and deep separation and recovery of zinc and lead in the lead-zinc sulfide ore.

[0006] To achieve the above object, the present invention provides a method for separating and extracting lead and zinc from lead-zinc sulfide ore powder, comprising the following steps:

[0007] S1. Weigh a certain mass of lead-zinc sulfide ore powder, add a surfactant and a dilute acid solution, mix well, and pour into a closed reactor;

[0008] S2, after heating to a constant temperature through oxygen, stirring is started to perform oxygen pressure leaching, and after a predetermined time, liquid-solid separation is performed to obtain lead sulfate-based leaching residue containing pyrite and elemental sulfur and zinc-iron leachate;

[0009] S3, adding an oxidant to the zinc-iron leachate, adding a neutralizing agent after the reaction to neutralize and hydrolyze, and obtaining a zinc solution after removing impurities;

[0010] S4, adding the lead sulfate-based leaching residue to an alkaline desulfurizer to carry out a desulfurization reaction to generate a lead carbonate-based conversion residue, and leaching the lead carbonate-based conversion residue in an acidic leaching agent to obtain a lead-containing leachate;

[0011] S5. Electrodepositing the lead-containing leachate to obtain metallic lead, and returning the waste electrolyte to the acidic leaching agent to achieve recycling of the leaching agent;

[0012] S6. The insoluble leaching residue is subjected to flotation treatment to obtain pyrite and elemental sulfur.

[0013] Preferably, in S1, the particle size of the lead-zinc sulfide ore powder is not greater than 75 μm; the surface additive is one or more of lignin, lignin sulfonate, alkylbenzene sulfonate, anthracite, lignite, and o-phenylenediamine, and the amount of the surface additive is 0.2% to 2.0% of the mass of the lead-zinc sulfide ore powder; and the dilute acid is sulfuric acid, nitric acid, or methanesulfonic acid.

[0014] Preferably, in said S2, H is controlled during the oxygen pressure leaching process. + The concentration is 0.2mol / L-4.0mol / L, the temperature is 80℃-150℃, the time is 60min-300min, and the oxygen partial pressure is 0.4MPa-2.0MPa.

[0015] Preferably, in S3, the oxidant is hydrogen peroxide, oxygen or ozone, the neutralizer is CaO, ZnO, Ca(OH)2 or Zn(OH)2, and the hydrolysis endpoint pH is 2.5-5.5.

[0016] Preferably, in S4, the alkaline desulfurizer is one or more of Na2CO3, NaHCO3, (NH4)2CO3, NaOH, NH4HCO3 and ammonia water, and the liquid-solid ratio of lead sulfate-based leaching residue to the alkaline desulfurizer is 3mL / g-15mL / g.

[0017] Preferably, in S4, the desulfurization reaction temperature is 25°C or -95°C, and the desulfurization reaction time is 20 min to 120 min.

[0018] Preferably, in S4, the acidic leaching agent is nitric acid, methanesulfonic acid, HCl-NaCl or HCl-CaCl2 solution, and the liquid-solid ratio of the lead carbonate-based conversion slag liquid to the acidic leaching agent is 4mL:10g.

[0019] Preferably, in said S4, the temperature of acid leaching is 25°C-85°C, the time is 20min-60min, and the H + The molar ratio of Pb is n H+ :n Pb =1.0-3.5.

[0020] Preferably, in S5, the cathode current density during electrodeposition is 100 A / m 2 -300A / m 2 , the electrolyte temperature is 25℃-65℃, and the electrode distance is 3cm-10cm.

[0021] Preferably, in said S5, the electrodeposition anode plate adopts a graphite plate or a platinum-titanium anode, and the electrodeposition cathode plate adopts a lead starting plate, a titanium plate or a stainless steel plate.

[0022] The advantages and positive effects of the method for separating and extracting lead and zinc from lead-zinc sulfide ore powder of the present invention are:

[0023] 1. The present invention can realize the phase transformation of galena mineral into slag and leaching of sphalerite into liquid by controlling the oxygen pressure leaching conditions, thereby creating good conditions for the subsequent recovery of lead and zinc.

[0024] 2. The present invention reduces the intermediate links, avoids the complex multi-stage flotation process in the flotation means, and avoids the environmental pollution caused by flotation reagents such as collectors and frothers. This not only reduces the equipment investment cost, but also reduces the labor intensity of operators and improves production efficiency. Moreover, the present invention has a good comprehensive utilization effect of valuable elements, not only realizing the separation of lead, zinc and iron, but also preparing metallic lead products and zinc-containing solutions (zinc can be recovered later). In addition, the present invention controls the oxygen pressure leaching conditions so that part of the elemental sulfur in the sulfide ore is oxidized to elemental sulfur, and part is overoxidized to SO4 2- , where elemental sulfur can be separated from insoluble pyrite by flotation.

[0025] 3. This invention utilizes a fully hydrometallurgical process to recover lead and zinc separately, offering the following advantages: ① It avoids pollution such as lead-containing fumes and SO2 generated during pyrometallurgical smelting; ② It achieves excellent lead-zinc separation, significantly improving metal recovery rates; and ③ It reduces energy consumption. Compared to pyrometallurgical smelting, which requires temperatures exceeding 1100°C, the hydrometallurgical process significantly reduces energy consumption, aligning with the development of green metallurgy.

[0026] The technical solution of the present invention is further described in detail below through the accompanying drawings and embodiments. BRIEF DESCRIPTION OF THE DRAWINGS

[0027] Figure 1 1 is a process flow chart of an embodiment of the present invention. DETAILED DESCRIPTION

[0028] In this application, unless otherwise defined, all technical and scientific terms used herein have the same meaning as commonly understood by those skilled in the art to which this application belongs. In the event of any inconsistency, the meaning described in this specification or the meaning derived from the contents recorded in this specification shall prevail. In addition, the terms used herein are only for the purpose of describing the embodiments of this application and are not intended to limit this application.

[0029] The embodiments of the present invention will be described in detail below with reference to the accompanying drawings.

[0030] like Figure 1 A method for separating and extracting lead and zinc from lead-zinc sulfide ore powder comprises the following steps:

[0031] S1. Weigh a certain mass of lead-zinc sulfide ore powder, add surfactant and dilute acid solution, mix well, and pour into a closed reactor for oxygen pressure leaching.

[0032] The particle size of the lead-zinc sulfide ore powder is not greater than 75 μm.

[0033] The surface additive is one or more of lignin, lignin sulfonate, alkylbenzene sulfonate, anthracite, lignite, and o-phenylenediamine, and the amount of the surface additive is 0.2% to 2.0% of the mass of the lead-zinc sulfide ore powder.

[0034] The dilute acid is sulfuric acid, nitric acid or methanesulfonic acid.

[0035] S2. After heating to a constant temperature through oxygenation, stirring is started, and after a predetermined time, liquid-solid separation is performed to obtain lead sulfate-based leaching residue and zinc-iron-containing leaching solution.

[0036] Control of H during oxygen pressure leaching + The concentration is 0.2mol / L-4.0mol / L.

[0037] The temperature is 80℃-150℃, the time is 60min-300min, and the oxygen partial pressure is 0.4MPa-2.0MPa.

[0038] S3. Add an oxidant to the zinc-iron leaching solution, add a neutralizer after the reaction to neutralize and hydrolyze, and obtain a zinc solution after removing impurities.

[0039] The oxidant is hydrogen peroxide, oxygen or ozone.

[0040] The neutralizing agent is CaO, ZnO, Ca(OH)2 or Zn(OH)2, and the hydrolysis endpoint pH is 2.5-5.5.

[0041] S4. Adding the lead sulfate-based leaching residue to an alkaline desulfurizer to carry out a desulfurization reaction to generate a lead carbonate-based conversion residue, and leaching the lead carbonate-based conversion residue in an acidic leaching agent to obtain a lead-containing leachate.

[0042] The alkaline desulfurization agent is one or more of Na2CO3, NaHCO3, (NH4)2CO3, NaOH, NH4HCO3 and ammonia water. The liquid-solid ratio of the lead sulfate-based leaching residue to the alkaline desulfurization agent is 3mL / g-15mL / g.

[0043] The desulfurization reaction temperature is 25°C and -95°C, and the desulfurization reaction time is 20min-120min.

[0044] The acidic leaching agent is nitric acid, methanesulfonic acid, HCl-NaCl or HCl-CaCl2 solution, and the liquid-solid ratio of the lead carbonate-based conversion slag liquid to the acidic leaching agent is 4mL:10g.

[0045] The acid leaching temperature is 25℃-85℃, the time is 20min-60min, and the H + The molar ratio of Pb is n H+ :n Pb =1.0-3.5.

[0046] S5. Electrodepositing the lead-containing leaching solution to obtain metallic lead, and returning the waste electrolyte to the acidic leaching agent to achieve recycling of the leaching agent.

[0047] The cathode current density during electrodeposition is 100 A / m 2 -300 A / m 2 , the electrolyte temperature is 25℃-65℃, and the electrode distance is 3cm-10cm.

[0048] The electrodeposition anode plate adopts graphite plate or platinum titanium anode, and the electrodeposition cathode plate adopts lead starting plate, titanium plate or stainless steel plate.

[0049] During the pressurized oxygen leaching process, by controlling the oxygen pressure leaching conditions, ZnS is converted into ZnSO4 and enters the leachate, and pyrite is partially dissolved into the leachate, which can be easily recovered after subsequent oxidation, neutralization and hydrolysis to remove impurities, while PbS is converted into PbSO4 and remains in the leaching residue, thus achieving lead and zinc separation. The elemental sulfur in the sulfide ore is oxidized to elemental sulfur, and some elemental sulfur is peroxidized to SO4 2- , and Pb in solution 2+ Combined to form PbSO4. The lead sulfate containing pyrite and elemental sulfur obtained in the oxygen pressure leaching stage (Ksp[PbSO4]=1.6×10 -8 )-based leaching residue can convert PbSO4 into acid-soluble lead carbonate (Ksp[PbCO3]=3.3×10 -14 ), which can then be dissolved in an acidic leaching agent and lead metal can be precipitated by electrodeposition. Insoluble pyrite and elemental sulfur can be separated by flotation to obtain pyrite and elemental sulfur.

[0050] The chemical reactions that mainly occur in the present invention are as follows:

[0051] Oxygen pressure leaching process:

[0052] 2FeS2+O2+4H + =2Fe 2+ +4S 0 +2H2O

[0053] PbS+8Fe 3+ +4H2O=PbSO4+8Fe 2+ +8H +

[0054] ZnS+2Fe 3+ =Zn 2+ +S 0 +2Fe 2+

[0055] 2S 0 +3O2+2H2O=2SO4 2- +4H +

[0056] Desulfurization conversion process:

[0057] PbSO4+Na2CO3=PbCO3+Na2SO4

[0058] The technical solution of the present invention is further described below through examples.

[0059] Example 1

[0060] A method for separating and extracting lead and zinc from lead-zinc sulfide ore powder comprises the following steps:

[0061] S1. Grind lead-zinc sulfide ore powder to a size of less than 75 μm. Weigh a certain amount of lead-zinc sulfide ore powder, mix it with calcium lignin sulfonate (0.7% by mass) and 1 mol / L dilute sulfuric acid, and pour the mixture into a sealed reactor.

[0062] S2. Adjust the oxygen partial pressure in the closed reactor to 0.4 MPa, start stirring after heating to 150° C. and react for 180 minutes, and perform liquid-solid separation while hot to obtain lead sulfate-based leaching residue and zinc-iron leaching solution.

[0063] S3. Add hydrogen peroxide to the zinc-iron leaching solution, and after sufficient reaction, add calcium oxide emulsion to neutralize to pH=3.5, remove impurities, and filter to obtain a zinc solution.

[0064] S4, the lead sulfate-based leaching residue is subjected to a desulfurization reaction in a NaHCO3 solution to generate a lead carbonate-based conversion residue, and then H+ :n Pb =3.2 Add methanesulfonic acid leaching agent and react for 60 minutes at a temperature of 50°C and a liquid-solid ratio of 8 mL / g.

[0065] S5, the lead-containing leachate obtained by filtering the above process is filtered at a cathode current density of 220A / m 2 Electrodeposition is carried out at 40°C to obtain metallic lead, and the waste electrolyte is returned to the leaching process to achieve the recycling of the leaching agent.

[0066] In this embodiment, the zinc leaching rate is 94.76%, the lead recovery rate is 96.54%, and the purity of metallic lead is 99.99%.

[0067] Example 2

[0068] A method for separating and extracting lead and zinc from lead-zinc sulfide ore powder comprises the following steps:

[0069] S1. Grind lead-zinc sulfide ore powder to less than 75 μm. Weigh a certain amount of lead-zinc sulfide ore powder, mix it with sodium lignin sulfonate (0.5% by mass) and 1.5 mol / L methanesulfonic acid, and pour the mixture into a sealed reactor.

[0070] S2. Adjust the oxygen partial pressure in the closed reactor to 0.8 MPa, start stirring after heating to 150° C. and react for 180 minutes, and perform liquid-solid separation while hot to obtain lead sulfate-based leaching residue and zinc-iron leaching solution.

[0071] S3. Add hydrogen peroxide to the zinc-iron leachate, and after sufficient reaction, add calcium oxide emulsion to neutralize to pH=3, remove impurities, and filter to obtain a zinc solution.

[0072] S4, the lead sulfate-based leaching residue is subjected to a desulfurization reaction in a Na2CO3 solution to generate a lead carbonate-based conversion residue, and then H+ :n Pb =3.0, add methanesulfonic acid leaching agent, and react for 60 minutes at a temperature of 60°C and a liquid-solid ratio of 10 mL / g.

[0073] S5, the lead-containing leachate obtained by filtering the above process is heated at a cathode current density of 200A / m 2 Electrodeposition is carried out at 50°C to obtain metallic lead, and the waste electrolyte is returned to the leaching process to achieve the recycling of the leaching agent.

[0074] In this embodiment, the zinc leaching rate is 93.28%, the lead recovery rate is 96.78%, and the purity of metallic lead is 99.99%.

[0075] Example 3

[0076] A method for separating and extracting lead and zinc from lead-zinc sulfide ore powder comprises the following steps:

[0077] S1. Grind lead-zinc sulfide ore powder to a size of less than 75 μm. Weigh a certain amount of lead-zinc sulfide ore powder, mix it with calcium lignin sulfonate (0.6% by mass) and 0.8 mol / L dilute sulfuric acid, and pour the mixture into a sealed reactor.

[0078] S2. Adjust the oxygen partial pressure in the closed reactor to 1.0 MPa, start stirring after heating to 180° C. and react for 120 minutes, and perform liquid-solid separation while hot to obtain lead sulfate-based leaching residue and zinc-iron leaching solution.

[0079] S3. Add hydrogen peroxide to the zinc-iron leachate, and after sufficient reaction, add calcium oxide emulsion to neutralize to pH=4, remove impurities, and filter to obtain a zinc solution.

[0080] S4, the lead sulfate-based leaching residue is subjected to a desulfurization reaction in a (NH4)2CO3 solution to generate a lead carbonate-based conversion residue, and then H+ :n Pb =2.8 Add methanesulfonic acid leaching agent and react for 60 minutes at a temperature of 80°C and a liquid-solid ratio of 9 mL / g.

[0081] S5, the lead-containing leachate obtained by filtering the above process is heated at a cathode current density of 200A / m 2 Electrodeposition is carried out at 50°C to obtain metallic lead, and the waste electrolyte is returned to the leaching process to achieve the recycling of the leaching agent.

[0082] In this embodiment, the zinc leaching rate is 93%, the lead recovery rate is 96%, and the purity of metallic lead is 99.99%.

[0083] Therefore, the method for separating and extracting lead and zinc from lead-zinc sulfide ore powder of the present invention is adopted to achieve efficient and deep separation and recovery of zinc and lead in the lead-zinc sulfide ore.

[0084] Finally, it should be noted that the above embodiments are only used to illustrate the technical solutions of the present invention rather than to limit the same. Although the present invention has been described in detail with reference to the preferred embodiments, those skilled in the art should understand that they can still modify or replace the technical solutions of the present invention with equivalents, and these modifications or equivalent replacements cannot cause the modified technical solutions to deviate from the spirit and scope of the technical solutions of the present invention.

Claims

1. A method for separating and extracting lead and zinc from lead-zinc sulfide ore powder, characterized in that: The following steps are involved: S1. Weigh a certain mass of lead-zinc sulfide ore powder, add a surfactant and a dilute acid solution, mix well, and pour into a closed reactor; S2, after heating to a constant temperature through oxygen, stirring is started to perform oxygen pressure leaching, and after a predetermined time, liquid-solid separation is performed to obtain lead sulfate-based leaching residue containing pyrite and elemental sulfur and zinc-iron leachate; S3, adding an oxidant to the zinc-iron leachate, adding a neutralizing agent after the reaction to neutralize and hydrolyze, and obtaining a zinc solution after removing impurities; S4, adding the lead sulfate-based leaching residue to an alkaline desulfurizer to carry out a desulfurization reaction to generate a lead carbonate-based conversion residue, and leaching the lead carbonate-based conversion residue in an acidic leaching agent to obtain a lead-containing leachate; S5. Electrodepositing the lead-containing leachate to obtain metallic lead, and returning the waste electrolyte to the acidic leaching agent to achieve recycling of the leaching agent; S6. The insoluble leaching residue is subjected to flotation treatment to obtain pyrite and elemental sulfur.

2. The method for separating and extracting lead and zinc from lead-zinc sulfide ore powder according to claim 1, characterized in that: In S1, the particle size of the lead-zinc sulfide ore powder is not greater than 75 μm; the surface additive is one or more of lignin, lignin sulfonate, alkylbenzene sulfonate, anthracite, lignite, and o-phenylenediamine, and the amount of the surface additive is 0.2% to 2.0% by mass of the lead-zinc sulfide ore powder; and the dilute acid is sulfuric acid, nitric acid, or methanesulfonic acid.

3. The method for separating and extracting lead and zinc from lead-zinc sulfide ore powder according to claim 1, characterized in that: In the S2, H is controlled during the oxygen pressure leaching process. + The concentration is 0.2mol / L-4.0mol / L, the temperature is 80℃-150℃, the time is 60min-300min, and the oxygen partial pressure is 0.4MPa-2.0MPa.

4. The method for separating and extracting lead and zinc from lead-zinc sulfide ore powder according to claim 1, characterized in that: In the S3, the oxidant is hydrogen peroxide, oxygen or ozone, the neutralizer is CaO, ZnO, Ca(OH)2 or Zn(OH)2, and the hydrolysis endpoint pH is 2.5-5.

5.

5. The method for separating and extracting lead and zinc from lead-zinc sulfide ore powder according to claim 1, characterized in that: In the S4, the alkaline desulfurizer is one or more of Na2CO3, NaHCO3, (NH4)2CO3, NaOH, NH4HCO3 and ammonia water, and the liquid-solid ratio of the lead sulfate-based leaching residue to the alkaline desulfurizer is 3mL / g-15mL / g.

6. The method for separating and extracting lead and zinc from lead-zinc sulfide ore powder according to claim 1, characterized in that: In the step S4, the desulfurization reaction temperature is 25°C or -95°C, and the desulfurization reaction time is 20 min to 120 min.

7. The method for separating and extracting lead and zinc from lead-zinc sulfide ore powder according to claim 1, characterized in that: In the S4, the acidic leaching agent is nitric acid, methanesulfonic acid, HCl-NaCl or HCl-CaCl2 solution, and the liquid-to-solid ratio of the lead carbonate-based conversion slag liquid to the acidic leaching agent is 4mL / g-10mL / g.

8. The method for separating and extracting lead and zinc from lead-zinc sulfide ore powder according to claim 1, characterized in that: In S4, the acid leaching temperature is 25°C-85°C, the time is 20min-60min, and the H + The molar ratio of Pb is n H+ :n Pb =1.0-3.

5.

9. The method for separating and extracting lead and zinc from lead-zinc sulfide ore powder according to claim 1, characterized in that: In the S5, the cathode current density during the electrodeposition is 100 A / m 2 -300 A / m 2 , the electrolyte temperature is 25℃-65℃, and the electrode distance is 3cm-10cm.

10. The method for separating and extracting lead and zinc from lead-zinc sulfide ore powder according to claim 1, characterized in that: In the above S5, the electrodeposition anode plate is a graphite plate or a platinum-titanium anode, and the electrodeposition cathode plate is a lead starting plate, a titanium plate or a stainless steel plate.

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