Process for intensified leaching of zinc leaching residues

Abstract

The invention relates to an enhanced leaching process for zinc leaching residues, and belongs to the technical field of zinc hydrometallurgy. According to the technical scheme, the method comprises the steps that reducing slag and zinc leaching slag are added into a first solvent for first leaching, and first leaching liquid and first underflow are obtained; the first underflow is added into a second solvent for second leaching, second leaching liquid and lead-silver-sulfur residues are obtained, and the second leaching liquid is returned for first leaching; and reducing the first leaching solution with zinc sulfide concentrate to obtain a reducing solution and reducing slag, and returning the reducing slag to the first leaching. Sulfuric acid is not generated in the leaching and reduction process, the leaching rate of zinc, iron and copper is high, and high-lead / high-silicon zinc concentrate which is difficult to treat by a roasting furnace can be adopted as a reducing agent.

Description

Technical Field

[0001] This invention belongs to the field of zinc hydrometallurgical technology, specifically relating to the leaching of zinc, iron, and copper from zinc leaching residue and zinc concentrate, and the leaching of Fe. 3+ The restoration. Background Technology

[0002] When zinc roasted ore is processed using a fully wet process, the zinc roasted ore is first subjected to neutral leaching and low-acid leaching to obtain zinc leaching residue. The zinc leaching residue is then subjected to enhanced leaching to dissolve elements such as zinc, iron, and copper in the residue, resulting in leachate and lead-silver slag. The leachate undergoes copper removal by displacement → pre-neutralization of acid → neutralization and indium precipitation → iron removal by hematite / goethite. The iron-removed liquid is then returned to the zinc roasted ore for neutral leaching and / or low-acid leaching, while the lead-silver slag is fed into a lead smelting furnace.

[0003] For enhanced leaching of zinc leaching residue, the following treatment methods exist in the existing technology.

[0004] The technical solution disclosed in Chinese patent CN105821208A is as follows: Zinc-containing materials are added to an acidic leaching agent, sulfur dioxide gas is introduced, and reduction leaching is carried out at 80-110℃. The total leaching rate of zinc is 97-99%, and the total leaching rate of iron is 91-97%. The ferric iron is reduced to ferrous iron for further copper removal. This method has high zinc and iron leaching rates and produces a small amount of lead-silver slag. However, the iron in the zinc leaching slag mainly exists in the form of zinc ferrite. During sulfur dioxide reduction leaching, 1.75 tons of sulfuric acid are generated for every 1 ton of iron leached, resulting in a large excess of sulfuric acid in the system. Calcium carbonate and / or calcium hydroxide are needed to neutralize the excess sulfuric acid, producing a huge amount of gypsum slag, which is extremely difficult to dispose of. Furthermore, during sulfur dioxide reduction leaching, some of the dissolved Cu... 2+ The copper leaching process results in the copper being reduced to CuS by sulfur dioxide and precipitated into the slag, leading to a Cu content in the lead-silver slag as high as 0.4–0.8 wt% and a low copper leaching rate. Therefore, this method is only used by one company in China and is difficult to promote within the zinc smelting industry.

[0005] The technical solution disclosed in Chinese patent CN104745810A is as follows: Weak acid slag and high-iron zinc sulfide concentrate are subjected to slag co-leaching at atmospheric pressure under conditions of 90℃ and 5 hours, producing co-leaching slag and co-leaching solution, wherein the co-leaching solution contains Fe. 3+ 3.54–3.59 g / L; the co-leaching residue was then subjected to atmospheric pressure high-acid leaching at 90℃ for 5 hours, resulting in high-acid residue containing 4.29–4.64 wt% Zn and 5.78–6.12 wt% Fe. The co-leaching solution obtained by this method contains Fe 3+ When iron / zinc powder replaces copper due to its high content, the iron / zinc powder must first remove the Fe. 3+ Reduced to Fe 2+ Talent and Cu 2+The reaction results in high consumption of iron / zinc powder, leading to high production costs. Furthermore, the ratio of weak acid slag to high-iron zinc sulfide concentrate is extremely critical. Excessive addition of high-iron zinc sulfide concentrate will result in a high zinc content in the acid slag, increasing zinc loss. Insufficient addition of high-iron zinc sulfide concentrate will lead to insufficient Fe content in the synergistic leachate. 3+ The content is higher; even if the ratio of the two is "just right," it is impossible to simultaneously achieve synergistic effects on Fe in the leachate. 3+ Low zinc content in high-acid slag is due to the fact that it is common knowledge in the metallurgical / chemical industry that to achieve high zinc content in the co-leaching solution... 3+ The zinc content is low, so an excess of zinc concentrate is required; conversely, to achieve a low zinc content in high-acid slag, an excess of weak-acid slag is necessary. Therefore, although this method has been publicly available for 10 years, it has not yet been industrialized.

[0006] The "Design Handbook for Heavy Non-ferrous Metal Smelting - Lead, Zinc, Bismuth Volume" (Metallurgical Industry Press, 1st edition, May 1996, p. 304) discloses a hot acid leaching process for zinc leaching residue to goethite. The technical solution is as follows: the hot acid leaching solution is reduced with zinc concentrate at a temperature of 80-85℃ for 4-7 hours, with a zinc concentrate excess of 15-20%. After reduction, Fe... 3+ If the zinc content is below 2 g / L, the sulfur slag (i.e., reduction slag) is returned to the fluidized bed roaster. This method has a long reduction time, and because the reduction slag has a high zinc content, it needs to be returned to the roaster. This not only affects the roaster's ability to process zinc concentrate, but also limits the types of zinc concentrate that can be used for reduction. For example, high-lead, high-silicon zinc concentrate that is difficult for the roaster to process cannot be selected for reduction. Summary of the Invention

[0007] This invention aims to at least partially address the problems or deficiencies existing in the prior art. To this end, this invention proposes a method with high zinc, iron, and copper leaching rates and a reducing solution containing Fe. 3+ This method of enhanced zinc leaching using zinc leaching residue has low efficiency, short reduction time, and does not cause excess sulfuric acid during the reduction process, making it suitable for widespread application in the industry.

[0008] To achieve the above objectives, the present invention is implemented through the following technical solution: The zinc leaching residue enhanced leaching process includes the following steps: The reducing residue and zinc leaching residue are added to the first solvent to carry out the first leaching, resulting in the first leachate and the first underflow. The first underflow is added to the second solvent for a second leaching, resulting in a second leachate and lead-silver-sulfur slag. The second leachate is then returned to the first leaching. The first leaching solution is reduced with zinc sulfide concentrate to obtain a reducing solution and a reducing residue. The reducing residue is returned to the first leaching solution.

[0009] In addition, the zinc leaching residue enhanced leaching method according to embodiments of the present invention may also have the following additional / limiting features: According to one embodiment of the present invention, the amount of zinc sulfide concentrate added is equal to the Fe content in the first leaching solution. 3+ 1.5 to 2.0 times the theoretical reaction amount.

[0010] According to one embodiment of the present invention, the amount of zinc sulfide concentrate added is equal to the amount of Fe dissolved in the zinc leaching residue. 3+ 0.8 to 1.0 times the theoretical reaction amount.

[0011] According to one embodiment of the present invention, the first solvent is composed of a second leachate, zinc electrolysis waste liquid, and / or dilute sulfuric acid.

[0012] According to one embodiment of the present invention, the second solvent is one or both of zinc electrolytic waste liquid and dilute sulfuric acid.

[0013] According to one embodiment of the present invention, the first leaching is carried out at a temperature of 80-95°C for a time of 2.0-3.0 hours.

[0014] According to one embodiment of the present invention, the second leaching is carried out at a temperature of 80-95°C for a time of 2.0-4.0 hours.

[0015] According to one embodiment of the present invention, the first leachate reduction is carried out at a temperature of 80-90°C for a time of 1.0-3.0 h.

[0016] According to one embodiment of the present invention, the reducing solution contains Fe 3+ Not more than 1.0 g / L.

[0017] According to one embodiment of the present invention, the lead-silver-sulfur slag is fed into a top-blown furnace, side-blown furnace, or bottom-blown furnace for lead smelting to recover lead, silver, and sulfur. Attached Figure Description

[0018] Figure 1 This is a process flow diagram of the present invention. Detailed Implementation

[0019] To make the objectives, technical solutions, and technical effects of the embodiments of the present invention clearer, the technical solutions in the embodiments of the present invention will be clearly and completely described below. Obviously, the embodiments described below are only a part of the embodiments of the present invention, and not all of them. All other embodiments obtained by those skilled in the art in conjunction with the embodiments of the present invention without creative effort are within the protection scope of the present invention.

[0020] In the description of this invention, it should be understood that the weights of the relevant components mentioned in the embodiments of this invention can refer not only to the specific content of each component, but also to the proportional relationship between the weights of the components. Therefore, any scaling up or down of the content of the relevant components according to the embodiments of this invention is within the scope of this invention. Specifically, the weights mentioned in the embodiments of this invention can be well-known mass units in the metallurgical / chemical fields such as g, kg, and t, and the volume units can be mL, L, and m. 3 Units of volume commonly known in the metallurgical / chemical fields.

[0021] In this embodiment of the invention, the operation mode of the first leaching, the second leaching, and the reduction can be intermittent or continuous. That is, the material can be fed intermittently, and the material can be leached / reduced intermittently and then discharged, or the material can be fed continuously, and the leaching / reduction and discharge can be continuous.

[0022] In some embodiments, for the method of enhancing zinc leaching with leaching residue, please refer to... Figure 1 Specifically, it includes the following steps: In step S100, the reducing residue and zinc leaching residue are added to the first solvent for the first leaching. After the leaching slurry is clarified by conventional methods, the first leachate and the first underflow are obtained.

[0023] In the example, the first solvent is composed of a second leachate, zinc electrolysis waste liquid, and / or dilute sulfuric acid.

[0024] Specifically, zinc, iron, and copper in zinc leaching residue mainly exist in the form of zinc ferrite (ZnO.Fe2O3) and copper ferrite (CuO.Fe2O3). During the first leaching, zinc ferrite and copper ferrite in the zinc leaching residue are dissolved by the first solvent, generating zinc sulfate, copper sulfate, and ferric sulfate, which then enter the first leaching solution.

[0025] However, the dissolution of zinc ferrite and copper ferrite is affected not only by the reaction acidity and temperature, but also by the Fe content in the reaction slurry. 3+ The effect of concentration. With the increase of Fe in the reaction slurry... 3+ As the concentration increases, the dissolution potential of zinc ferrite and copper ferrite will rise, making dissolution more difficult, especially for those containing Fe. 3+ The dissolution of zinc ferrite and copper ferrite becomes more difficult when the second leaching solution is returned to the first leaching solution and when the sulfuric acid concentration in the slurry decreases in the later stages of the reaction.

[0026] Zinc, iron, and copper in the reduction residue mainly exist in the form of sulfides. After returning the reduction residue to the first leaching stage, the zinc, iron, and copper sulfides in the reduction residue will react with the Fe in the ore pulp. 3+ Reduced to Fe 2+ This process not only leached zinc, iron, and copper from the reducing slag, but also removed Fe from the first leaching pulp. 3+The reduced content stabilizes the dissolution potentials of zinc ferrite and copper ferrite, which is beneficial for the leaching of zinc, iron and copper from the zinc leaching residue.

[0027] In some embodiments, the first leaching is performed at a controlled temperature of 80–95°C for a duration of 2.0–3.0 h.

[0028] However, the first leaching is insufficient to completely remove zinc, iron, and copper from the zinc leaching residue and reduction residue. Therefore, the zinc, iron, and copper content in the underflow (i.e., solid residue) from the first leaching is too high and cannot be directly fed into the top-blown / side-blown / bottom-blown furnace of lead smelting for lead-silver-sulfur recovery. Therefore, a second leaching is required, producing a second leachate and lead-silver-sulfur residue.

[0029] In step S200, the first underflow is added to the second solvent for a second leaching. The leached slurry is clarified and filtered using conventional methods to obtain a second leaching solution and lead-silver-sulfur slag. The second leaching solution is returned to the first leaching, and the lead-silver-sulfur slag is used in a top-blown / side-blown / bottom-blown furnace for lead smelting to recover lead, silver, and sulfur.

[0030] In some embodiments, the second solvent is one or both of zinc electrolytic waste liquid and dilute sulfuric acid.

[0031] In some embodiments, the second leaching is performed at a temperature of 80–95°C for a time of 2.0–4.0 h.

[0032] Advantageously, through the first and second leaching processes, the total leaching rate of zinc in the zinc leaching residue and the reduced sulfur residue can be 96-97%, the total leaching rate of iron can be 92-94%, and the total leaching rate of copper can be 95-98%; the lead-silver sulfur residue contains 2-3 wt% Zn, 4-6 wt% Fe, and 0.05-0.10 wt% Cu.

[0033] Step S300: Reduce the first leachate with zinc sulfide concentrate, wherein the amount of zinc sulfide concentrate added is equal to the amount of Fe in the first leachate. 3+ The reaction rate is 1.5 to 2.0 times the theoretical reaction rate, the temperature is 80 to 90℃, and the time is 1.0 to 3.0 hours. After clarification and filtration, the reduced slurry produces a reducing solution and a reducing slag. The reducing slag is returned to the first leaching stage, and the reducing solution contains Fe. 3+ Not more than 1.0 g / L.

[0034] The main reaction formulas for the first leaching, second leaching, and reduction of the first leaching solution are as follows (expressed in terms of zinc ferrite in the zinc leaching residue and zinc sulfide in the zinc concentrate used for reduction): ZnFe2O4+ 4H2SO4= ZnSO4+ Fe2(SO4)3+ 4H2O (1) ZnS + Fe2(SO4)3= ZnSO4+ 2FeSO4+ S↓ (2) It should be noted that in this invention, zinc in zinc sulfide concentrate is assumed to exist entirely in the form of ZnS, and the amount of zinc sulfide concentrate added is determined based on the theoretical reaction amount of reaction formula (2).

[0035] Advantageously, during the reduction of the first leachate, reaction (2) mainly occurs, since the amount of zinc concentrate added is Fe 3+ The amount of ZnS is 1.5 to 2.0 times the theoretical reaction amount, meaning that ZnS is in excess, which allows the reaction to proceed more completely. The reducing solution contains Fe. 3+ The concentration is no greater than 1.0 g / L, and the reduction reaction time can be shortened to 1.0-3.0 h, which is much lower than the 4.0-7.0 h in the existing technology, thus solving the problem of long reduction time in the existing technology. The reduction slag is returned to the first leaching, which solves the problem that the reduction slag with high zinc content needs to be returned to the roasting furnace, which not only affects the roasting furnace's ability to process zinc concentrate, but also limits the type of zinc concentrate used for reduction. High-lead / high-silicon zinc concentrate that is difficult to process in the roasting furnace can be selected for reduction.

[0036] Advantageously, reactions (1) and (2) occur simultaneously during the first and second leaching processes. This is because the zinc sulfide concentrate has already reacted with the Fe in the first leaching solution. 3+ The reaction results in a decrease in zinc content in the reduction residue returned from the first leaching. At this point, Fe... 3+ The excess is excessive, while the deficiency is insufficient for ZnS; therefore, the zinc, iron, and copper sulfides and Fe in the reducing slag are excessive. 3+ The reaction is more thorough, which can reduce the zinc, iron, and copper content in lead-silver sulfur slag and improve the leaching rates of zinc, iron, and copper. This solves the problems of long leaching time and poor Fe content in the synergistic leaching solution of existing technologies (such as CN104745810A). 3+ Problems such as high content of zinc and high acidity in slag; reducing Fe with zinc sulfide concentrate. 3+ It does not generate sulfuric acid, which solves the problem of generating a huge amount of gypsum slag due to excessive sulfuric acid in existing technologies (such as CN105821208A), and also solves the problem of low copper leaching rate.

[0037] The amount of zinc sulfide concentrate added is the Fe dissolved in the zinc leaching residue. 3+ The theoretical reaction amount is 0.8 to 1.0 times. The amount of zinc sulfide concentrate added can be less than 1.0 times because lead, copper, and other sulfides in the zinc sulfide concentrate participate in the reduction of Fe. 3+ The reaction.

[0038] In some embodiments, Fe is dissolved in the zinc leaching residue 3+ The amount is greater than that of Fe in the first leachate. 3+ The reason for the reduced amount is that the reducing residue returned from the first leaching had already reduced some of the Fe during the first and second leaching processes. 3+ .

[0039] In some embodiments, the zinc leaching residue described in this invention is obtained by neutral leaching and low-acid leaching of zinc roasted ore. The zinc leaching residue contains 10–30 wt% zinc, 10–30 wt% iron, and 0.2–1.5 wt% copper. Neutral leaching and low-acid leaching of zinc roasted ore are prior art and are irrelevant to the problem solved by this invention; therefore, they will not be described in detail.

[0040] Similarly, after copper removal by displacement, indium and germanium precipitation by neutralization, and iron removal by goethite / hematite, the reducing solution is returned to the zinc roasted ore for neutral leaching and / or low-acid leaching. Since copper removal by displacement, indium and germanium precipitation by neutralization, and iron removal by goethite / hematite are existing technologies and are not related to the problem solved by this invention, they will not be described in detail.

[0041] Similarly, the recovery of lead, silver, and sulfur from lead-silver slag using top-blown / side-blown / bottom-blown furnaces in lead smelting is also existing technology and is unrelated to the problem solved by this invention, so it will not be described in detail.

[0042] Advantageously, the present invention is highly adaptable to the zinc concentrate added, and because the reducing slag is not returned to the roasting furnace, high-lead and high-silicon zinc concentrates that are difficult to process in the roasting furnace can be selected, thereby broadening the types of zinc concentrates.

[0043] To more clearly illustrate the present invention, the following specific embodiments will be described in detail.

[0044] Because a reduction residue needs to be prepared and added to the first leaching, unless otherwise specified, in specific embodiments, the first leaching, the second leaching, and the reduction of the first leaching solution are all the second round of leaching / reduction. Example 1

[0045] Zinc leaching residue: Zn 20.65wt%, Fe 29.06wt%, Cu 0.89wt%; Zinc concentrate: Zn 46.11wt%, Fe 10.02wt%, Cu 0.32wt%, Pb 1.36wt%, SiO 2 10.32wt%.

[0046] 200g of zinc leaching residue and the first round of reduction residue were used for the first leaching. The first solvent was dilute sulfuric acid and the second leaching solution from the first round. The first leaching temperature was 95℃ and the time was 2.0h. After the first leaching slurry was clarified, the first leaching solution and the first underflow were obtained.

[0047] A second solvent, zinc electrolysis waste liquid, was added to the first underflow for a second leaching process. The second leaching temperature was 95°C, and the time was 2.0 h. After filtration, the second leaching slurry yielded a second leachate and lead-silver-sulfur slag.

[0048] The first leachate was added to 70g of zinc concentrate for reduction. The added zinc concentrate was the Fe from the first leachate. 3+The theoretical reaction amount is 2.0 times that of Fe dissolved in the zinc leaching residue. 3+ The reaction rate was 1.0 times the theoretical reaction volume, at a temperature of 90℃, for 1.0 h. After filtration of the reduced slurry, a reducing solution and a reducing residue were obtained. Example 2

[0049] The zinc leaching residue contained 10.26 wt% Zn, 22.35 wt% Fe, and 1.36 wt% Cu; the zinc concentrate contained 44.03 wt% Zn, 3.95 wt% Fe, 0.62 wt% Cu, 9.33 wt% Pb, and 8.92 wt% SiO2.

[0050] 200g of zinc leaching residue and reduction residue were subjected to the first leaching. The first solvent was zinc electrolysis waste liquid and the second leaching solution from the first round. The first leaching temperature was 80℃ and the time was 3.0h. After the first leaching slurry was clarified, the first leaching solution and the first underflow were obtained.

[0051] A second solvent, zinc electrolysis waste liquid, was added to the first underflow for a second leaching process. The second leaching temperature was 80°C, and the time was 4.0 h. After filtration of the second leaching slurry, a second leachate and lead-silver-sulfur slag were obtained.

[0052] The first leachate was added to 58g of zinc concentrate for reduction. The added zinc concentrate was the Fe from the first leachate. 3+ The theoretical reaction amount is 1.5 times that of Fe dissolved in the zinc leaching residue. 3+ The reaction rate was 0.8 times the theoretical reaction rate, the temperature was 90℃, and the reaction time was 1.0 h. After filtration of the reduced slurry, a reducing solution and a reducing residue were obtained. Example 3

[0053] The zinc leaching residue contained 16.01 wt% Zn, 19.89 wt% Fe, and 0.66 wt% Cu; the zinc concentrate contained 49.11 wt% Zn, 6.78 wt% Fe, 0.46 wt% Cu, 5.42 wt% Pb, and 5.12 wt% SiO2.

[0054] 200g of zinc leaching residue and reduction residue were subjected to the first leaching. The first solvent consisted of dilute sulfuric acid, zinc electrolysis waste liquid, and the second leaching solution from the first round. The first leaching temperature was 88℃, and the time was 2.5h. After the first leaching slurry was clarified, the first leaching solution and the first underflow were obtained.

[0055] A second solvent, zinc electrolysis waste liquid, was added to the first underflow for a second leaching process. The second leaching temperature was 88°C, and the time was 2.5 hours. After filtration, the second leaching slurry yielded a second leachate and lead-silver-sulfur slag.

[0056] The first leachate was added to 40g of zinc concentrate for reduction. The added zinc concentrate was the Fe from the first leachate. 3+The amount of Fe dissolved in the zinc leaching residue is 1.7 times the theoretical reaction amount. 3+ The reaction rate was 0.9 times the theoretical reaction rate, the temperature was 85℃, and the reaction time was 2.0 h. After filtration of the reduced slurry, a reducing solution and a reducing residue were obtained.

[0057] Comparative Example 1 The zinc leaching residue from Example 1 was used. Following the technical solution disclosed in CN105821208A, a closed container was used, and sulfur dioxide reduction leaching was employed.

[0058] Take 200g of zinc leaching residue, add dilute sulfuric acid, and then pass sulfur dioxide gas through at 110℃ for 2.5 hours. After filtering the reduced leaching slurry, a reducing solution and lead-silver-sulfur residue are obtained.

[0059] Comparative Example 2

[0060] The zinc leaching residue and zinc concentrate from Example 1 were used. The ore-slag co-leaching was employed using the technical solution disclosed in CN104745810A.

[0061] Take 200g of zinc leaching residue and 70g of zinc concentrate. The amount of zinc concentrate is equal to the amount of Fe dissolved in the zinc leaching residue. 3+ The theoretical reaction volume was 1.0 times that of dilute sulfuric acid and the first round of high-acid leaching solution. The temperature was 90℃ and the time was 5.0h. After clarification and filtration of the co-leaching slurry, co-leaching solution (equivalent to reducing solution) and co-leaching residue were obtained.

[0062] The co-leaching residue was added to dilute sulfuric acid for high-acid leaching at 90℃ for 5.0 h. After clarification and filtration of the high-acid slurry, high-acid leachate and high-acid residue (equivalent to lead-silver-sulfur residue) were obtained.

[0063] Comparative Example 3

[0064] The zinc leaching residue and zinc concentrate from Example 1 are used. The reduction residue is not returned to the first leaching.

[0065] Take 200g of zinc leaching residue, add it to dilute acid solution, and leach at 90℃ for 3.0h. After the first leaching slurry clarifies, the first leachate and the first underflow are obtained.

[0066] Dilute sulfuric acid was added to the first underflow for a second leaching process at a temperature of 90°C for 3.0 hours. After filtration, the second leaching slurry yielded a second leachate and lead-silver-sulfur slag.

[0067] The first leachate was added to 70g of zinc concentrate for reduction. The added zinc concentrate was the Fe from the first leachate. 3+The reaction was carried out at 1.2 times the theoretical reaction rate, at a temperature of 90℃, for 3.0 hours. After filtration of the reduced slurry, a reducing solution and a reducing slag were obtained. The dry weight of the reducing slag was 43 g, with Zn 20.03 wt%, Pb 2.21 wt%, and SiO 2 16.76 wt%.

[0068] The results of the comparison of the effects of Examples 1-3 and Comparative Examples 1-3 are shown in Table 1.

[0069] Table 1 shows the results of the comparison between the examples and the comparative examples.

[0070]

[0071] Note: The amount of zinc sulfide concentrate added is the Fe dissolved in the zinc leaching residue. 3+ Multiples of theoretical reaction amount As can be seen from the comparative results provided in Table 1 for Examples 1-3 and Comparative Examples 1-3, compared with Comparative Example 1, no sulfuric acid is generated during the leaching and reduction process of the specific Examples 1-3 of the technical solution provided by the present invention, and the Fe in the reducing solution is... 3+ It was basically equivalent to Comparative Example 1, although the iron leaching rate was slightly lower than that of Comparative Example 1, the copper leaching rate was better than that of Comparative Example 1; compared with Comparative Example 2, the Fe in the reducing solution was... 3+ The leaching rates of zinc, iron, and copper were significantly better than those of Comparative Example 2; compared to Comparative Example 3, the Fe content in the reducing solution was also significantly higher. 3+ The reduction time of Comparative Example 3 is better than that of Comparative Example 3, indicating that the reduction time of Comparative Example 3 needs to be greater than 3 hours to allow Fe to be reduced. 3+ The leaching rate of zinc and iron is less than 2 g / L, which is better than that of Comparative Example 3. Moreover, the reduction slag of Comparative Example 3 contains 20.03 wt% Zn and 16.76 wt% SiO2, which cannot be used in the lead smelting furnace or returned to the roasting furnace. In other words, the technical solution adopted by Comparative Example 3 cannot use high-lead / high-silicon zinc concentrate as a reducing agent.

[0072] In the description of this invention, it should be understood that the terms "first" and "second" are used for descriptive purposes only and should not be construed as indicating or implying relative importance or implicitly specifying the number of features indicated. Thus, a feature defined as "first" or "second" may explicitly or implicitly include at least one of that feature.

[0073] In the description of this specification, references to terms such as "one embodiment," "some embodiments," or "example" indicate that a specific feature or characteristic described in connection with that embodiment or example is included in at least one embodiment or example of the present invention. In this specification, the illustrative expressions of the above terms do not necessarily refer to the same embodiment or example. Furthermore, the specific features or characteristics described may be combined in any suitable manner in one or more embodiments or examples. Moreover, without contradiction, those skilled in the art can combine and / or combine the different embodiments or examples described in this specification, as well as the features of different embodiments or examples.

[0074] Although embodiments of the present invention have been shown and described above, it is understood that the above embodiments are exemplary and should not be construed as limiting the present invention. Those skilled in the art can make changes, modifications, substitutions and variations to the above embodiments within the scope of the present invention.

Claims

1. A process for enhancing zinc leaching residue leaching, characterized in that, Includes the following steps: The reducing residue and zinc leaching residue are added to the first solvent to carry out the first leaching, resulting in the first leachate and the first underflow. The first underflow is added to the second solvent for a second leaching, resulting in a second leachate and lead-silver-sulfur slag. The second leachate is then returned to the first leaching. The first leaching solution is reduced with zinc sulfide concentrate to obtain a reducing solution and a reducing residue. The reducing residue is returned to the first leaching solution.

2. The process according to claim 1, characterized in that, The amount of zinc sulfide concentrate added is equal to the Fe content in the first leaching solution. 3+ 1.5 to 2.0 times the theoretical reaction amount.

3. The process according to claim 1 or 2, characterized in that, The amount of zinc sulfide concentrate added is the Fe dissolved in the zinc leaching residue. 3+ 0.8 to 1.0 times the theoretical reaction amount.

4. The process according to claim 1, characterized in that, The first solvent is composed of a second leachate, zinc electrolysis waste liquid, and / or dilute sulfuric acid.

5. The process according to claim 1, characterized in that, The second solvent is one or both of zinc electrolysis waste liquid and dilute sulfuric acid.

6. The process according to claim 1, characterized in that, The first leaching is performed at a temperature of 80–95°C for 2.0–3.0 hours.

7. The process according to claim 1, characterized in that, The second leaching is carried out at a temperature of 80–95°C for 2.0–4.0 hours.

8. The process according to claim 1, characterized in that, The first leachate reduction is carried out at a temperature of 80–90°C for a time of 1.0–3.0 h.

9. The process according to claim 1 or 8, characterized in that, The reducing solution contains Fe 3+ Not more than 1.0 g / L.

10. The process according to claim 1, characterized in that, The lead-silver sulfur slag is then processed in a lead smelting furnace.

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