A process for the recovery of zinc and lead metals from a lead-rich slag produced in a zinc smelting process

Abstract

The application discloses a method for recovering zinc and lead metals from lead-rich slag generated in a zinc smelting process, which comprises low-temperature acid leaching, magnetic separation, oxidation heap leaching, stirring leaching and the like processes, so that the lead-rich slag is subjected to low-temperature acid leaching and magnetic separation first, zinc compounds which are easily soluble are leached out sufficiently, and the content of iron ions entering the liquid phase is effectively reduced, thereby reducing the interference of iron ions in the zinc liquid, then the processes of oxidation heap leaching and stirring leaching are combined, and zinc ferrite and zinc sulfide minerals which are difficult to be acid-dissolved can be further dissolved, the zinc-containing substances basically enter the liquid phase, the target of separating lead and zinc is achieved, and finally, combined with dilute hydrochloric acid washing, part of calcium and silicon compounds can be dissolved, so that the lead grade in the final slag is improved, and comprehensive utilization is realized.

Description

TECHNICAL FIELD

[0001] The present application relates to the technical field of comprehensive utilization of mineral resources, and particularly relates to a method for recovering zinc and lead from lead-rich slag generated in a zinc smelting process. BACKGROUND

[0002] Metal zinc is widely used in alloy, casting, electroplating, printing and dyeing industry, pharmaceutical industry, rubber industry and battery industry. Metal zinc has become an important metal material indispensable to the national economy and people's livelihood. Due to the wide range of uses and economy, the production and consumption of metal zinc are increasing. The increasing demand for metal zinc materials has brought unprecedented resource and environmental problems. On the one hand, the rapid increase in demand has led to a rapid decrease in rich ores and easy-to-select ores, and the zinc smelting industry is gradually facing the crisis of raw material supply shortage; on the other hand, the increasing production and consumption of metal zinc have led to an increasing amount of zinc-containing waste slag and waste materials, which not only wastes resources but also causes environmental pollution and ecological damage. Processing and utilization of low-grade ores, complex ores and secondary resources such as zinc-containing waste slag have become a problem to be solved for the sustainable and healthy development of the zinc metallurgy industry.

[0003] Due to the gradual enhancement of environmental awareness and the increasing strictness of national environmental protection policy, slag treatment has become a bottleneck restricting wet zinc smelting. There are many different processes for wet zinc smelting, and many kinds of slag are produced, and the chemical composition and chemical properties of these slags are different. The zinc that is difficult to leach in these zinc-containing waste slags mainly exists in the form of zinc ferrite (ZnFe2O4) and contains a small amount of zinc sulfide. Since zinc ferrite belongs to the spinel (AB2O4) type, the crystal lattice has considerable stability, so it is difficult to dissolve by ordinary acid leaching, and the high-temperature acid leaching process is limited in applicability due to high material requirements, high cost, and operation danger. SUMMARY

[0004] The present application aims at the above-mentioned problems, and provides a method for recovering zinc and lead from lead-rich slag generated in a zinc smelting process, so as to solve the deficiencies in the prior art.

[0005] The technical scheme adopted by the present application is as follows: a method for recovering zinc and lead from lead-rich slag generated in a zinc smelting process, comprising the following steps:

[0006] A, low-temperature acid leaching, the lead-rich slag is slurried with water, and then 98% concentrated sulfuric acid is added for stirring acid leaching to obtain an acid leaching slurry;

[0007] B, magnetic separation, the acid leaching slurry is subjected to magnetic separation to obtain a magnetic separation concentrate and a magnetic separation tailing;

[0008] C. Oxidation heap leaching, after the magnetic separation tailings are pressure filtered, the filter cake obtained is mixed with manganese dioxide and 98% concentrated sulfuric acid, and after stirring, oxidation heap leaching is carried out, and an oxidation heap leaching slurry is obtained;

[0009] D. Stirring leaching, the oxidation heap leaching slurry is subjected to stirring leaching, and after completion, water is added for washing and filtration, and a lead residue filter cake and a zinc-containing filtrate are obtained.

[0010] In the present application, the inventors have broken through the problem of leaching of zinc ferrite and zinc sulfide which is difficult to solve in the conventional acid leaching process. Through research, it has been found that by increasing the acid concentration and adding manganese dioxide oxidant, the structure of zinc ferrite and zinc sulfide can be effectively destroyed, and the influence of iron impurities can be reduced, so that part of the zinc enters the leaching solution, greatly reducing the zinc content in the residue, effectively separating lead and zinc, and improving the resource recycling efficiency. The key innovation of the method of the present application lies in magnetic separation and oxidation heap leaching. After low-temperature acid leaching, magnetic separation is performed. The purpose of magnetic separation is to remove the magnetic iron impurities in the slurry (zinc ferrite does not have magnetism and has no effect), thereby ensuring the recovery rate of lead and zinc while reducing the content of iron impurities and the influence of subsequent processes (for example, excessive consumption of oxidizing agent by magnetic iron impurities). After magnetic separation, oxidation heap leaching is performed. Oxidation heap leaching under high concentration liquid-solid ratio conditions is beneficial to the oxidation of zinc ferrite and zinc sulfide, thereby promoting their leaching from the slurry, improving the zinc recovery rate, and helping to realize the separation of lead and zinc.

[0011] Further, in step A, the liquid-solid mass ratio of the lead-rich residue to water is 3-4:1 (for example, it can be 3:1, 3.5:1, 4:1, etc.), the acid leaching temperature of low-temperature acid leaching is 65-75℃ (for example, it can be 65℃, 70℃, 75℃, etc.), the pH value of the acid leaching slurry is maintained at 0.5-1, and the acid leaching time is 3-4h.

[0012] Further, in step B, the magnetic field strength of the magnetic separation is 600-800mT, for example, it can be 600mT, 650mT, 700mT, 800mT, etc.

[0013] Further, in step C, the amount of manganese dioxide added is added in a ratio of 1:1.5-2 of zinc to manganese molar amount, and the ratio can be 1:1.5, 1:1.7, 1:2, etc.

[0014] Further, in step C, 98% concentrated sulfuric acid is added in a ratio of 1:1-1.25 of liquid to solid mass based on the dry mass of the filter cake, for example, it can be 1:1, 1:1.1, 1:1.25, etc.

[0015] Further, in step C, the oxidation heap leaching time is 16-24h.

[0016] Further, in step D, the temperature of the stirring leaching is 50-60 DEG C, and the leaching time is 3-3.5 h. Further, the amount of water used for washing is generally 2-3.5:1 by mass ratio of the lead-rich slag, and the specific ratio can be adjusted according to the actual situation.

[0017] Further, the obtained lead slag filter cake is washed with a dilute hydrochloric acid solution (the concentration can be 0.3-1 mol / L) to obtain a final lead slag, and the liquid-solid mass ratio can be 2-3:1, the washing temperature is generally normal temperature, and the washing time is 1-2 h. The washing liquid can be recycled as a dilute hydrochloric acid solution. The final lead slag contains Pb>35%, Zn<2%, and Fe<1.2% by mass fraction, and can be directly sold to the outside. If the lead content in the washed lead slag is less than 35%, the dilute hydrochloric acid solution can be used for washing again to obtain high-grade lead slag and a zinc-containing leaching solution.

[0018] Further, the lead-rich slag contains 11-14% of Pb, 15-18% of Zn, and 4-6% of Fe by mass percentage.

[0019] Further, the lead-rich slag is one or more of zinc-containing leaching slag and lead-zinc smelting slag, the lead phase is mainly lead sulfate, and the zinc phase is mainly zinc ferrite and zinc sulfide.

[0020] Experimental research shows that, according to the method of the present application, through the mutual coordination between the above steps, the lead and zinc in the lead-rich slag containing 11-14% of Pb, 15-18% of Zn, and 4-6% of Fe can be effectively separated, wherein the final lead slag product contains Pb>35%, Zn<2.5%, and Fe<1.2%, and about 90% of the lead metal enters the slag. At the same time, more than 92% of the zinc metal enters the leaching solution, and the SO4 2- content in the liquid phase reaches 70-120 g / L, which can be returned to the system as an acid leaching solution, so as to realize the separation and recycling of lead and zinc metals and increase economic benefits.

[0021] Multiple experimental researches show that the method of the present application is not only suitable for the separation and comprehensive recycling of lead and zinc resources in various existing lead-rich slags, but also has good adaptability to the slag samples generated in the zinc smelting process, which have high lead content and contain a large amount of zinc ferrite and zinc sulfide and other difficult-to-treat materials by conventional acid leaching. After the treatment by the process, part of the magnetic iron minerals can be magnetically separated, the content of iron ions in the subsequent acid leaching solution is reduced, zinc ferrite and zinc sulfide are decomposed, zinc metal basically enters the liquid phase, lead metal remains in the solid phase, and the effective separation and recycling of lead and zinc metals are realized.

[0022] As described above, due to the adoption of the technical solutions, the present application has the following beneficial effects:

[0023] 1、The present application can effectively separate lead and zinc in lead-rich slag, so that the final lead slag product contains Pb>35%, Zn<2.5%, Fe<1.2%, about 90% of the lead metal enters the slag, and more than 92% of the zinc metal enters the leaching solution, the obtained lead slag can be directly sold, the pricing coefficient of lead metal is greatly improved, the economic benefit is increased, meanwhile, the SO4 2- content in the liquid phase reaches 70-120g / L, which can be returned to the system as acid leaching solution, so that the separation and recycling of lead and zinc metals are realized;

[0024] 2、The present application can make water-soluble zinc salt and zinc oxide in the slag fully dissolve into the liquid phase in the low-temperature acid leaching process stage, which has low leaching cost, is easy to control and reduces environmental pollution;

[0025] 3、In the magnetic separation process stage, the magnetic iron in the slag is enriched to a certain extent through magnetic separation, compared with the conventional zinc leaching, the iron metal content in the liquid phase in the subsequent leaching process can be greatly reduced, so as to be beneficial to the recovery of zinc metal in the liquid phase;

[0026] 4、In the oxidation heap leaching-stirring leaching process stage, the insoluble zinc ferrite and zinc sulfide in the slag are decomposed at low temperature by the action of concentrated acid and manganese dioxide oxidant, so that zinc basically enters the liquid phase, high-concentration ore slurry is used for oxidation heap leaching reaction in this stage, the temperature is low, the equipment volume is small, the equipment cost and operation difficulty can be effectively reduced, the energy consumption cost is reduced, and when filtering and washing, the washing water amount can be adjusted according to the target value of SO4 2- in the circulating liquid, so as to save water as much as possible;

[0027] 5、When the zinc concentration reaches the requirement after the leaching zinc-containing liquid is recycled, the zinc can be recovered through purification into the electrolysis system, so as to realize the resource utilization of zinc metal;

[0028] 6、The iron content of the magnetic separation concentrate obtained by the present application is 26-30%, of which the lead content is 10-12% and the zinc content is 9-10%, the yield of this part of the magnetic concentrate is 5-7%, which can be stored, and the lead and zinc metals can be recovered by subsequent rotary kiln centralized treatment;

[0029] 7、The lead slag containing Pb<35% is washed by dilute hydrochloric acid at room temperature to remove a large amount of water-insoluble calcium oxide and part of silicon oxide in the slag, this process is simple and easy to operate as an independent section, so as to improve the lead content of the lead slag after washing, and the washing water can be recycled and treated in this process section and does not participate in the water circulation of the foregoing process. BRIEF DESCRIPTION OF DRAWINGS

[0030] Figure 1 is a flowchart of the method for recovering zinc and lead metals from lead-rich slag generated in the zinc smelting process;

[0031] Figure 2 This is a schematic diagram of the dilute hydrochloric acid washing process in the method of the present invention. Detailed Implementation

[0032] The present invention will now be described in detail with reference to the accompanying drawings.

[0033] To make the objectives, technical solutions, and advantages of this invention clearer, the invention will be further described in detail below with reference to the accompanying drawings and embodiments. It should be understood that the specific embodiments described herein are merely illustrative and not intended to limit the invention.

[0034] Example 1

[0035] like Figure 1 and Figure 2 As shown, according to the method for recovering zinc and lead metals from lead-rich slag produced in the zinc smelting process provided by the present invention, the lead-rich slag containing (mass fraction) Pb 12.1%, Zn 17.5%, and Fe 5% produced in the zinc smelting process of our company, Hongda Co., Ltd., is treated by the following steps for recovery:

[0036] S1. Low-temperature acid leaching: Add water to lead-rich slag at a liquid-to-solid ratio (mass ratio) of 4:1 to make a slurry. Then, add 98% concentrated sulfuric acid at 70℃ and stir to leach. Maintain pH = 0.5-1 and leach for 4 hours to obtain acid leaching slurry.

[0037] S2. Magnetic separation: The acid leaching slurry is subjected to wet magnetic separation under a magnetic field strength of 800mT to obtain magnetic concentrate and magnetic tailings (slurry concentration 20%, feed rate 30-50g / s);

[0038] S3. Oxidative heap leaching: After the magnetic separation tailings are filtered by pressure, the filtrate is treated as zinc-containing liquid A and can be recycled or enter the zinc electrolysis system after meeting the requirements. The filter cake is mixed with manganese dioxide oxidant at a zinc-to-manganese molar ratio of 1:1.5. After mixing, 98% concentrated sulfuric acid is added at a liquid-to-solid mass ratio of 1:1 (the solid amount is calculated based on the filter cake). After stirring and mixing, oxidative heap leaching is performed and the mixture is allowed to stand for 24 hours to obtain the oxidative heap leaching slurry.

[0039] S4. Stirred leaching: The slurry of the oxide heap is stirred and leached at a temperature of 60°C for 3.5 hours. After completion, water is added for washing and filtration to obtain lead slag filter cake and zinc-containing filtrate B.

[0040] S5. Washing with dilute hydrochloric acid: Add the lead slag filter cake to a 0.5 mol / L hydrochloric acid solution for washing. The liquid-to-solid mass ratio is 2:1. Stir at room temperature for 2 hours and then filter to obtain qualified lead slag filter cake and reusable dilute hydrochloric acid washing solution.

[0041] The lead grade of the obtained lead residue filter cake after washing is 36.8%, the zinc grade is 2.4%, and the iron grade is 0.8%. Relative to the entire process, the lead metal recovery rate entering the solid phase is 89.7%, and the zinc metal recovery rate entering the liquid phase is 92.3%.

[0042] Example 2

[0043] As shown in Figure 1 and Figure 2 A method for recovering zinc and lead metals from a lead-rich residue generated in a zinc smelting process is provided according to the present application. The lead-rich residue generated in the zinc smelting process of the company Hongda shares contains (mass fraction) Pb 12.1%, Zn 17.5%, and Fe 5%. The recovery treatment of the lead-rich residue is carried out by the following steps:

[0044] S1, low-temperature acid leaching: the lead-rich residue is slurried by adding water at a liquid-solid ratio (mass ratio) of 4:1, then 98% concentrated sulfuric acid is added for stirring leaching at a temperature of 75°C, the pH is maintained at 0.5-1, and the leaching time is 4h, to obtain an acid leaching slurry;

[0045] S2, magnetic separation: the acid leaching slurry is subjected to wet magnetic separation at a magnetic field strength of 800mT to obtain a magnetic separation concentrate and a magnetic separation tailing (slurry concentration 20%, feed speed 30-50g / s);

[0046] S3, oxidative heap leaching: the magnetic separation tailing is filtered, the filtrate is treated and then recycled or enters the zinc electrolysis system after meeting the requirements, and the filter cake is mixed with a manganese dioxide oxidant, the amount of which is added according to a zinc-manganese molar ratio of 1:1.75, then 98% concentrated sulfuric acid is added at a liquid-solid mass ratio of 1:1.25 (the solid amount is calculated based on the dry basis of the filter cake), the mixture is stirred and mixed, and then oxidative heap leaching is carried out, and the mixture is left to stand for 24h to obtain an oxidative heap leaching slurry;

[0047] S4, stirring leaching: the oxidative heap leaching slurry is subjected to stirring leaching at a temperature of 60°C for 3.5h, then water is added for washing and filtration to obtain a lead residue filter cake and a zinc-containing filtrate B;

[0048] S5, dilute hydrochloric acid washing: the lead residue filter cake is washed with a 0.5mol / L hydrochloric acid solution containing hydrochloric acid, the liquid-solid mass ratio is 2:1, and the mixture is stirred at room temperature for 2h before being filtered to obtain a qualified lead residue filter cake and a recycled dilute hydrochloric acid washing solution.

[0049] The lead grade of the obtained lead residue filter cake after washing is 35.8%, the zinc grade is 2.0%, and the iron grade is 0.9%. Relative to the entire process, the lead metal recovery rate entering the solid phase is 89.1%, and the zinc metal recovery rate entering the liquid phase is 92.8%.

[0050] Example 3

[0051] As Figure 1 and Figure 2 shown, a method for recovering zinc and lead metals from a lead-rich slag generated in a zinc smelting process is provided according to the present application. The lead-rich slag generated in the zinc smelting process of the present company contains (mass fraction) Pb 13.7%, Zn 15.8%, and Fe 5.8%. The recovery treatment is carried out by the following steps:

[0052] S1, low-temperature acid leaching: the lead-rich slag is slurried with water at a liquid-solid ratio (mass ratio) of 4:1, then 98% concentrated sulfuric acid is added for stirring leaching at a temperature of 70°C, the pH is maintained at 0.5-1, and the leaching time is 4h, to obtain an acid leaching slurry;

[0053] S2, magnetic separation: the acid leaching slurry is subjected to wet magnetic separation at a magnetic field strength of 800mT, to obtain a magnetic separation concentrate and a magnetic separation tailing (slurry concentration 20%, feed speed 30-50g / s);

[0054] S3, oxidative heap leaching: the magnetic separation tailing is filtered, the filtrate is treated and then recycled or enters the zinc electrolysis system after meeting the requirements, the filter cake is mixed with manganese dioxide oxidant, the mixing amount is 1:1.2 according to the mole ratio of zinc to manganese, then 98% concentrated sulfuric acid is added at a liquid-solid mass ratio of 1:1 (the solid amount is calculated on a dry basis), the mixture is stirred and mixed, and then oxidative heap leaching is carried out after standing for 24h, to obtain an oxidative heap leaching slurry;

[0055] S4, stirring leaching: the oxidative heap leaching slurry is subjected to stirring leaching at a temperature of 60°C, the time is 3.5h, then water is added for washing and filtration, to obtain a lead slag filter cake and a zinc-containing filtrate B.

[0056] It is detected that the lead grade of the obtained lead slag filter cake is 36.2%, the zinc grade is 1.7%, and the iron grade is 0.9%. Relative to the entire process, the recovery rate of lead metal entering the solid phase is 90.7%, and the recovery rate of zinc metal entering the liquid phase is 93.3%.

[0057] Comparative Example 1

[0058] Comparative Example 1 is the same as Example 1, except that Comparative Example 1 does not perform magnetic separation operation, but directly performs oxidative heap leaching after filtration, and the subsequent steps are the same as Example 1.

[0059] The test results show that the lead grade of the lead residue filter cake obtained after washing is 32.4%, the zinc grade is 3.4%, and the iron grade is 1.4%. Compared with the whole process, the lead metal recovery rate into the solid phase is 89.2%, the zinc metal recovery rate into the liquid phase is 93.5%, and the amount of iron metal into the liquid phase reaches 90%. Therefore, without magnetic separation, the lead grade in the leaching residue is reduced, which is not conducive to the recovery of lead, and the zinc metal grade cannot be reduced to the target value of 2% or less. Without the magnetic separation process, the magnetic iron minerals will enter the leaching process, which will consume part of the acid and oxidant, resulting in a large amount of iron elements entering the liquid phase, and the iron element content entering the liquid phase is increased by about 40% compared with the magnetic separation process.

[0060] Comparative Example 2

[0061] Comparative Example 2 is the same as Example 1, except that the magnetic separation strength of Comparative Example 2 is 1000 mT, and the subsequent steps are the same as Example 1.

[0062] The test results show that the lead grade of the lead residue filter cake obtained after washing is 34.8%, the zinc grade is 2.8%, and the iron grade is 0.9%. Compared with the whole process, the lead metal recovery rate into the solid phase is 89.1%, and the zinc metal recovery rate into the liquid phase is 91.84%. Therefore, increasing the magnetic separation strength beyond a certain range does not significantly improve the metal recovery effect of the whole process, and cannot effectively reduce the iron grade in the final leaching residue. It will also cause part of the zinc elements to enter the magnetic concentrate during magnetic separation, thereby reducing the amount of metal entering the liquid phase.

[0063] Comparative Example 3

[0064] Comparative Example 3 is the same as Example 1, except that during the oxidation heap leaching, the liquid-solid mass ratio is 2:1 (the same amount of acid is used, and the amount of water is increased), and the rest is the same as Example 1.

[0065] The test results show that the lead grade of the lead residue filter cake obtained after washing is 34.8%, the zinc grade is 5.3%, and the iron grade is 0.7%. Compared with the whole process, the lead metal recovery rate into the solid phase is 88.8%, and the zinc metal recovery rate into the liquid phase is 87.7%. Therefore, increasing the liquid-solid ratio and reducing the sulfuric acid concentration of the reaction has a greater impact on zinc leaching. The zinc grade in the leaching residue is too high, which is not conducive to the leaching of zinc into the liquid phase.

[0066] Comparative Example 4

[0067] Comparative Example 4 is the same as Example 1, except that when the manganese dioxide oxidant is added, the zinc-manganese molar ratio is 1:1.

[0068] The test results show that the lead grade of the lead residue filter cake obtained after washing is 34.1%, the zinc grade is 3.2%, and the iron grade is 1.2%. Relative to the whole process, the lead metal recovery rate into the solid phase is 87.9%, and the zinc metal recovery rate into the liquid phase is 91.1%. It can be seen that when the amount of manganese dioxide is reduced, the zinc grade in the leaching residue cannot be effectively reduced, and the increase in the lead grade is limited.

[0069] The above merely describes preferred embodiments of the present application, but should not be used to limit the present application, and any modification, equivalent replacement, and improvement within the spirit and principle of the present application should be included in the protection scope of the present application.

Claims

1. A method for recovering zinc and lead metals from lead-rich slag produced during zinc smelting, characterized in that, Includes the following steps: A. Low-temperature acid leaching: Add water to lead-rich slag to make a slurry, then add 98% concentrated sulfuric acid and stir for acid leaching to obtain an acid leaching slurry; B. Magnetic separation: The acid leaching slurry is subjected to magnetic separation to obtain magnetic concentrate and magnetic tailings. C. Oxidative heap leaching: After the magnetic separation tailings are filtered by pressure, the resulting filter cake is mixed with manganese dioxide and 98% concentrated sulfuric acid, stirred and mixed evenly, and then subjected to oxidative heap leaching to obtain oxidative heap leaching slurry. D. Stirred leaching: The oxide heap leaching slurry is stirred and leached. After completion, water is added for washing and filtration to obtain lead slag filter cake and zinc-containing filtrate.

2. The method as described in claim 1, characterized in that, In step A, the liquid-solid mass ratio of lead-rich slag to water is 3-4:1, the acid leaching temperature is 65-75℃, the pH value of the acid leaching slurry is maintained at 0.5-1, and the acid leaching time is 3-4 hours.

3. The method as described in claim 1, characterized in that, In step B, the magnetic field strength of the magnetic separation is 600-800 mT.

4. The method as described in claim 1, characterized in that, In step C, manganese dioxide is added in a ratio of zinc to manganese molars of 1:1.5-2.

5. The method as described in claim 1, characterized in that, In step C, based on the mass of the filter cake base, 98% concentrated sulfuric acid is added at a liquid-to-solid mass ratio of 1:1-1.

25.

6. The method as described in claim 1, characterized in that, In step C, the oxidation heap leaching time is 16-24 hours.

7. The method as described in claim 1, characterized in that, In step D, the temperature for stirring and leaching is 50-60℃, and the leaching time is 3-3.5h.

8. The method according to any one of claims 1-7, characterized in that, The obtained lead slag filter cake was washed with dilute hydrochloric acid solution to obtain the final lead slag. By mass fraction, the final lead slag contained >35% lead, <2% zinc, and <1.2% iron.

9. The method as described in claim 8, characterized in that, By mass percentage, lead-rich slag contains 11-14% lead, 15-18% zinc, and 4-6% iron.

10. The method as described in claim 9, characterized in that, The lead-rich slag is one or more of zinc-containing leaching slag and lead-zinc smelting slag, with lead phase mainly being lead sulfate and zinc phase mainly being zinc ferrite and zinc sulfide.

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