Method for separating and recovering valuable metals from zinc leaching residues

Abstract

The invention discloses a method for separating and recovering valuable elements in zinc leaching residues, which comprises the following steps: mixing the zinc leaching residues with a sulfuric acid solution, adding a reducing agent, stirring to react under certain conditions, and performing solid-liquid separation after the reaction is finished to obtain filtrate 1 and filter residues 1; mixing the filter residue 1 with an oxalic acid solution, carrying out a stirring reaction under certain conditions, after the reaction is finished, filtering to obtain a filtrate 2 and a filter residue 2, carrying out a neutralization reaction on the filtrate 2 to hydrolyze and precipitate tin in the solution, and filtering to obtain a filtrate 3 and a stannous hydroxide precipitate; mixing and stirring the filtrate 1 and the filtrate 3, and filtering to obtain filtrate 4 and ferrous oxalate precipitate; according to the method, valuable elements in the zinc leaching residues are separated and recycled one by one, especially for Ga, Ge, In and Sn which are low in grade, effective enrichment and recycling can be achieved, maximum utilization of valuable metal resources is achieved, and generation of solid waste is reduced.

Description

Technical Field

[0001] The invention belongs to the field of resource recycling, and in particular relates to a method for separating and recovering valuable metals from zinc leaching residue. Background Art

[0002] At present, the hydrometallurgical zinc smelting process has become the mainstream zinc smelting method at home and abroad. The leaching slag produced by it contains not only valuable metals such as zinc, lead, indium, tin, silver, gallium, germanium, but also harmful elements such as arsenic and cadmium. Many domestic manufacturers combine the leaching slag with the iron removal slag for processing, resulting in a high iron content in the slag. For this kind of leaching slag, the more mature treatment methods are the volatilization kiln method and the fuming furnace method, but this method has high energy consumption and the tail gas SO 2 The tail gas absorption system is large in scale, with high operating costs, and the valuable metal elements, especially the rare metals, are seriously dispersed, with low recovery rates. How to separate and recycle the valuable elements has always been an important research topic for metallurgical researchers at home and abroad.

[0003] Patent CN106222429.A uses oxygen pressure leaching to treat zinc leaching residue, allowing valuable elements such as zinc and indium to enter the solution, and iron to enter the slag in the form of hematite, thus achieving the separation of zinc and iron, but there is still the problem of storage and treatment of iron slag. Patent CN116114935.A uses a side-blowing furnace and a fuming furnace for combined treatment, which can separate Cu, Ag, Au from Pb, Zn, and In, but there are also problems of high energy consumption and the inability to recycle iron. Patent CN104532295.B uses oxidation leaching to treat zinc leaching residue, and recovers zinc and cadmium from the leachate in the form of zinc sulfate and sponge cadmium, but the separation and treatment of other valuable elements is not considered.

[0004] The present invention provides a method for separating and recovering valuable elements in zinc leaching slag. The method can separate and recover valuable elements such as Zn, In, Sn, Ga, Ge, Pb, Ag, etc. in the zinc leaching slag, and in particular provides a method for recovering Fe, which can achieve almost zero solid waste generation. Summary of the invention

[0005] The present invention aims to provide a method for separating and recovering valuable elements in zinc leaching residue. Through the present invention, zinc, indium, gallium, germanium, tin, lead, silver and iron in zinc leaching residue can be recovered respectively, so as to maximize resource utilization and minimize solid waste production.

[0006] In order to achieve the above object, the present invention is implemented by the following technical scheme: a method for separating and recovering valuable elements in zinc leaching slag, the method specifically comprising the following steps:

[0007] S1) reduction leaching: mixing the zinc leaching residue with a sulfuric acid solution, adding a reducing agent, stirring and reacting under certain conditions, and performing solid-liquid separation after the reaction to obtain a filtrate 1 and a filter residue 1;

[0008] S2) combined leaching: the filter residue 1 is mixed with the oxalic acid solution, and stirred for reaction under certain conditions. After the reaction is completed, the mixture is filtered to obtain a filtrate 2 and a filter residue 2;

[0009] S3) neutralization and precipitation of tin: neutralization reaction is performed on filtrate 2 to hydrolyze and precipitate the tin in the solution, and filtrate is filtered to obtain filtrate 3 and stannous hydroxide precipitate;

[0010] S4) ferrous oxalate precipitation: the filtrate 1 obtained in S1) and the filtrate 3 obtained in S3) are mixed and stirred, and filtered to obtain a filtrate 4 and ferrous oxalate precipitation.

[0011] Furthermore, the leaching reaction conditions in S1) are: temperature 50-100° C., reaction time 1-3 h, liquid-to-solid volume mass ratio 5-10 ml / g, initial sulfuric acid concentration 50-100 g / L, stirring speed 300-500 r / min, and the amount of reducing agent added is 10%-30% of the amount of zinc leaching slag.

[0012] Furthermore, the reducing agent is one or more of sodium sulfite, sulfur dioxide, and zinc sulfide.

[0013] Furthermore, the conditions of the complex leaching reaction in S2) are: liquid-to-solid volume mass ratio of 5-10 ml / g, initial oxalic acid concentration of the solution of 50-100 g / L, reaction temperature of 70-100° C., reaction time of 1-3 h, and stirring speed of 300-500 r / min.

[0014] Furthermore, in S2), the filter residue 2 is returned to the lead smelting system to recover lead and silver.

[0015] Furthermore, the neutralizing agent in S3) is one or a combination of sodium hydroxide, calcium hydroxide, calcium oxide, and sodium carbonate, and the neutralization reaction end point pH value is 8 to 10.

[0016] Furthermore, the stannous hydroxide in S3) is subjected to roasting treatment at a temperature of 100 to 300° C. for a time of 1 to 3 hours to obtain tin concentrate.

[0017] Furthermore, the specific process parameters in S4) are: filtrate 1 and 3 are mixed and reacted in a volume ratio of 1:1, the reaction temperature is controlled at 80-100°C, and filtrate 3 is slowly added to filtrate 1 at a fixed flow rate under stirring, and the time for adding all the filtrate 3 is controlled to be more than 1.0h.

[0018] Furthermore, when the concentration of Ga, Ge and In in the filtrate 4 in S4) reaches above 100 mg / L, the circuit is opened, and when it is below 100 mg / L, the filtrate is returned to S1) for recycling; and the solution of the filtrate 4 after recovering Ga, Ge and In is used to produce zinc sulfate heptahydrate.

[0019] Compared with the prior art, the beneficial effects and outstanding advantages of the present invention are:

[0020] 1) It realizes the resource recovery of iron elements and avoids the generation of hazardous waste of iron slag;

[0021] 2) For the low-content Ga, Ge, In, and Sn elements, better separation, enrichment, and recovery can be achieved, thus maximizing resource recycling;

[0022] 3) The entire process is simple and environmentally friendly, and can effectively reduce energy consumption and environmental pollution. BRIEF DESCRIPTION OF THE DRAWINGS

[0023] Figure 1 The present invention is a process flow chart of a method for separating and recovering valuable metals from zinc leaching residue. DETAILED DESCRIPTION

[0024] The present invention is further described in detail below by way of examples, but the scope of the present invention is not limited to these examples.

[0025] like Figure 1 As shown, the present invention provides a method for separating and recovering valuable elements in zinc leaching slag, which mainly comprises the following steps:

[0026] S1) reduction leaching: mixing the zinc leaching residue with a sulfuric acid solution, adding a reducing agent, stirring and reacting under certain conditions, and performing solid-liquid separation after the reaction to obtain a filtrate 1 and a filter residue 1, wherein valuable elements such as Zn, Fe, In, Ge, and Ga in the zinc leaching residue enter the filtrate 1, and Pb, Ag, and Sn remain in the filter residue 1;

[0027] S2) combined leaching: the filter residue 1 is mixed with the oxalic acid solution, and stirred under certain conditions for reaction. After the reaction is completed, the mixture is filtered to obtain a filtrate 2 and a filter residue 2. The element Sn enters the filtrate 2, and Pb and Ag remain in the filter residue 2. The filter residue 2 is sent to the lead smelting system to recover Pb and Ag.

[0028] S3) neutralization and precipitation of tin: neutralization reaction is performed on filtrate 2 to hydrolyze and precipitate the tin in the solution, and filtrate is filtered to obtain filtrate 3 and stannous hydroxide precipitate;

[0029] S4) ferrous oxalate precipitation: filtrate 1 and filtrate 3 are mixed and stirred, and filtered to obtain filtrate 4 and ferrous oxalate precipitate. The ferrous oxalate precipitate can be used for purification to prepare ferrous oxalate product for sale.

[0030] Further, the reducing agent in S1) is one or more of sodium sulfite, sulfur dioxide, and zinc sulfide, and the leaching reaction conditions are a temperature of 50-100° C., a reaction time of 1-3 h, a liquid-to-solid volume mass ratio of 5-10 ml / g, an initial sulfuric acid concentration of 50-100 g / L, a stirring speed of 300-500 r / min, and an amount of the reducing agent added of 10%-30% of the amount of zinc leaching slag;

[0031] The conditions for the combined leaching reaction in S2) are as follows: a liquid-solid volume mass ratio of 5-10 ml / g, an initial oxalic acid concentration of 50-100 g / L, a reaction temperature of 70-100° C., a reaction time of 1-3 h, and a stirring speed of 300-500 r / min;

[0032] The neutralizing agent used in S3) is one or a combination of sodium hydroxide, calcium hydroxide, calcium oxide, and sodium carbonate, and the neutralization reaction end point pH value is 8 to 10;

[0033] S3) calcining the tin hydroxide at a temperature of 100 to 300° C. for a time of 1 to 3 hours;

[0034] The operation of S4) is: controlling the reaction temperature to 80-100°C, slowly adding the filtrate 3 to the filtrate 1 at a fixed flow rate under stirring, and controlling the time for the complete addition of the filtrate 3 to be more than 1.0h;

[0035] In S4), the filtrate 4 is selected to be open-circuited to recover Ga, Ge, In according to the concentration of Ga, Ge, In therein, or returned to S1) for recycling. When the concentration of Ga, Ge, In in the filtrate 4 reaches above 100 mg / L, the filtrate is open-circuited. When the concentration of Ga, Ge, In in the filtrate 4 is below 100 mg / L, the filtrate is returned to S1) for recycling. The solution after open-circuiting to recover Ga, Ge, In is used to produce zinc sulfate heptahydrate.

[0036] Example 1

[0037] Take 500 g of zinc leaching slag, the composition of which is shown in Table 1.

[0038] Table 1 Zinc leaching residue composition / %

[0039] Ga Ge In Ag Pb Zn Fe Sn 0.018 0.012 0.10 0.02 8.25 9.6 14.8 0.44

[0040] 1) Reduction leaching: Mix and stir with 3000ml sulfuric acid solution, the initial sulfuric acid concentration of the solution is 100g / L, heat to 80℃, add 100g sodium sulfite (added three times), stir and react for 3.0h, filter, and obtain 2950ml of filtrate, and the wet weight of the filter residue is 350.3g. The filtrate contains Ga 28mg / L, Ge 20mg / L, In 170mg / L, Zn 16.05g / L, Fe 24.61g / L;

[0041] 2) Combined leaching: Add the filter residue obtained in step (1) to 1300ml of 100g / L oxalic acid solution, heat to 80°C, stir and react for 3.0h, filter, and obtain 1250ml of filtrate, with a filter residue wet weight of 245g. The filtrate contains 1.7g / L Sn; the filter residue is sent to the lead smelting system to recover Pb and Ag.

[0042] 3) Neutralization of precipitated tin: Slowly add sodium hydroxide to the filtrate of step (2) while stirring until the end point pH value reaches about 9.0, stop stirring, let stand for 1.0h and filter. The obtained filter residue is roasted at 200℃ for 2.0h to obtain 4.58g of dry residue, containing 46.1% Sn, and the filtrate volume is 1180ml, containing 20mg / L Sn;

[0043] 4) Oxalic acid precipitation of iron: slowly add the filtrate of step (3) to the filtrate of step (1), stir and heat, when the temperature reaches 90°C, keep the temperature for 2.0h, filter. The obtained filtrate 3200ml is returned to step (1) for recycling, the Fe content of the filtrate is 0.75g / L, and the wet weight of the filter residue is 258g, which is used to prepare ferrous oxalate.

[0044] Example 2

[0045] Take 500 g of zinc leaching slag, the composition of which is shown in Table 2.

[0046] Table 2 Zinc leaching residue composition / %

[0047] Ga Ge In Ag Pb Zn Fe Sn 0.018 0.012 0.10 0.02 8.25 9.6 14.8 0.44

[0048] 1) Reduction leaching: Use the previous batch of oxalic acid iron precipitation liquid and concentrated sulfuric acid to prepare 3000ml of initial solution with a sulfuric acid concentration of 100g / L, heat it to 80℃, add 500g of zinc leaching residue, pass SO 2 The gas was subjected to reduction leaching reaction, stirred for 3.0 hours, filtered, and 2800 ml of filtrate was obtained, and the wet weight of the filter residue was 351.6 g. The filtrate contained Ga 54 mg / L, Ge 39 mg / L, In 338 mg / L, Zn 32.1 g / L, and Fe 25.61 g / L;

[0049] 2) Combined leaching: Add the filter residue obtained in step (1) to 1300ml 100g / L oxalic acid solution, heat to 90°C, stir and react for 3.0h, filter, and obtain 1250ml filtrate, with a filter residue wet weight of 247g. The filtrate contains 1.8g / L Sn; the filter residue is sent to the lead smelting system to recover Pb and Ag.

[0050] 3) Neutralization of precipitated tin: Slowly add sodium hydroxide to the filtrate of step (2) while stirring until the end point pH value reaches about 9.0, stop stirring, let stand for 1.0h and filter. The obtained filter residue is roasted at 200℃ for 2.0h to obtain 4.45g dry residue, containing 47.1% Sn, and the filtrate volume is 1200ml, containing 22mg / L Sn;

[0051] 4) Oxalic acid precipitation of iron: slowly add the filtrate of step (3) to the filtrate of step (1), stir and heat, when the temperature reaches 90°C, keep the temperature for 2.0h, filter. The obtained filtrate 3100ml is returned to step (1) for recycling, the Fe content of the filtrate is 0.8g / L, and the wet weight of the filter residue is 248g, which is used to prepare ferrous oxalate.

[0052] Example 3

[0053] Take 500 g of zinc leaching slag, the composition of which is shown in Table 3.

[0054] Table 3 Zinc leaching residue composition / %

[0055] Ga Ge In Ag Pb Zn Fe Sn 0.10 0.11 0.20 0.05 9.45 10.26 13.8 0.55

[0056] 1) Reduction leaching: prepare 3000ml of initial solution with sulfuric acid concentration of 100g / L, add 500g of zinc leaching residue, heat to 90℃, add 150g of zinc sulfide (added three times), stir and react for 3.0h, filter, and obtain 2800ml of filtrate, with a wet weight of 361.6g of filter residue. The filtrate contains Ga 162mg / L, Ge 180mg / L, In 335mg / L, Zn 50.10g / L, Fe225.43g / L;

[0057] 2) Combined leaching: Add the filter residue obtained in step (1) to 1300ml of 100g / L oxalic acid solution, heat to 90°C, stir and react for 3.0h, filter, and obtain 1200ml of filtrate, with a filter residue wet weight of 236g. The filtrate contains 2.10g / L Sn; the filter residue is sent to the lead smelting system to recover Pb and Ag.

[0058] 3) Neutralization of precipitated tin: Slowly add sodium hydroxide to the filtrate of step (2) while stirring until the end point pH value reaches about 9.5, stop stirring, let stand for 1.0h and filter. The obtained filter residue is roasted at 200℃ for 2.0h to obtain 4.80g dry residue, containing 45.1% Sn, and the filtrate volume is 1150ml, containing 25mg / L Sn;

[0059] 4) Oxalic acid precipitation of iron: slowly add the filtrate of step (3) to the filtrate of step (1), stir and heat, when the temperature reaches 90°C, keep warm for 2.0h, filter. The obtained filtrate is returned to step (1) for recycling, 3000ml of filtrate, Fe content 0.7g / L, filter residue wet weight 265g, used to prepare ferrous oxalate, the filtrate is open circuit, and Ga, Ge, In and Zn therein are recovered.

[0060] The method for separating and recovering valuable metals from zinc leaching residue provided in the embodiment of the present application is described in detail. The description of the above embodiment is only used to help understand the method and its core idea of ​​the present application; at the same time, for those skilled in the art, according to the idea of ​​the present application, there will be changes in the specific implementation method and application scope. In summary, the content of this specification should not be understood as limiting the present application.

[0061] For example, certain words are used in the specification and claims to refer to specific components. Those skilled in the art should understand that hardware manufacturers may use different nouns to refer to the same component. This specification and claims do not use differences in names as a way to distinguish components, but use differences in the functions of components as the criteria for distinction. As mentioned throughout the specification and claims, "including" and "comprising" are open-ended terms, so they should be interpreted as "including / including but not limited to". "Approximately" means that within an acceptable error range, those skilled in the art can solve the technical problems within a preset error range and basically achieve the technical effects. The subsequent description of the specification is a preferred embodiment of the present application, but the description is for the purpose of illustrating the general principles of the present application, and is not used to limit the scope of the present application. The scope of protection of the present application shall be subject to that defined in the attached claims.

[0062] It should also be noted that the terms "include", "comprises" or any other variations thereof are intended to cover non-exclusive inclusion, so that a product or system including a series of elements includes not only those elements, but also other elements not explicitly listed, or also includes elements inherent to such a product or system. In the absence of more restrictions, the elements defined by the sentence "comprises a ..." do not exclude the existence of other identical elements in the product or system including the elements.

[0063] It should be understood that the term "and / or" used in this article is only a description of the association relationship of associated objects, indicating that there can be three relationships. For example, A and / or B can represent: A exists alone, A and B exist at the same time, and B exists alone. In addition, the character " / " in this article generally indicates that the associated objects before and after are in an "or" relationship.

[0064] The above description shows and describes several preferred embodiments of the present application, but as mentioned above, it should be understood that the present application is not limited to the form disclosed herein, and should not be regarded as excluding other embodiments, but can be used in various other combinations, modifications and environments, and can be modified within the scope of the application concept described herein through the above teachings or the technology or knowledge of the relevant field. The changes and modifications made by those skilled in the art do not deviate from the spirit and scope of the present application, and should be within the scope of protection of the claims attached to the present application.

Claims

1. A method for separating and recovering valuable elements in zinc leaching slag, characterized in that: The method specifically comprises the following steps: S1) reduction leaching: mixing the zinc leaching residue with a sulfuric acid solution, adding a reducing agent, stirring and reacting under certain conditions, and performing solid-liquid separation after the reaction to obtain a filtrate 1 and a filter residue 1; S2) combined leaching: the filter residue 1 is mixed with the oxalic acid solution, and stirred for reaction under certain conditions. After the reaction is completed, the mixture is filtered to obtain a filtrate 2 and a filter residue 2; S3) neutralization and precipitation of tin: neutralizing the filtrate 2 to hydrolyze and precipitate the tin in the solution, filtering to obtain the filtrate 3 and the stannous hydroxide precipitate, and roasting the stannous hydroxide precipitate to obtain the tin concentrate; S4) ferrous oxalate precipitation: the filtrate 1 obtained in S1) and the filtrate 3 obtained in S3) are mixed and stirred, and filtered to obtain a filtrate 4 and ferrous oxalate precipitation.

2. The method according to claim 1, characterized in that The leaching reaction conditions in S1) are: temperature 50-100°C, reaction time 1-3h, liquid-to-solid volume mass ratio 5-10ml / g, initial sulfuric acid concentration 50-100g / L, stirring speed 300-500r / min, and the amount of reducing agent added is 10%-30% of the amount of zinc leaching slag.

3. The method according to claim 2, characterized in that The reducing agent is one or more of sodium sulfite, sulfur dioxide and zinc sulfide.

4. The method according to claim 1, characterized in that: The conditions of the complex leaching reaction in S2) are: liquid-to-solid volume mass ratio of 5-10 ml / g, initial oxalic acid concentration of the solution of 50-100 g / L, reaction temperature of 70-100° C., reaction time of 1-3 h, and stirring speed of 300-500 r / min.

5. The method according to claim 1, characterized in that The neutralizing agent in S3) is one or a combination of sodium hydroxide, calcium hydroxide, calcium oxide, and sodium carbonate, and the neutralization reaction end point pH value is 8 to 10; The calcination temperature is 100-300° C. and the calcination time is 1-3 hours.

6. The method according to claim 1, characterized in that The specific process parameters in S4) are: filtrate 1 and filtrate 3 are mixed and reacted in a volume ratio of 1:1, the reaction temperature is controlled at 80-100°C, and filtrate 3 is slowly added to filtrate 1 at a fixed flow rate under stirring, and the time for adding all the filtrate 3 is controlled to be more than 1.0h.

7. The method according to claim 1, characterized in that When the concentration of Ga, Ge, and In in the filtrate 4 in S4) reaches 100 mg / L or more, the circuit is opened; when the concentration is below 100 mg / L, the filtrate is returned to S1) for recycling.

8. The method according to claim 1, characterized in that The filter residue 2 in S2) is returned to the lead smelting system to recover lead and silver.

9. The method according to claim 1, characterized in that: According to the method of the present invention, the recovery rates of Zn, Fe and Sn in the zinc leaching slag reach 98%, 96% and 95% or more respectively.

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