A method of selective pressure leaching of a zinc sulphide concentrate

Abstract

The application belongs to the technical field of hydrometallurgy, and particularly relates to a selective pressure leaching method for zinc sulfide concentrate. The method comprises the following steps: (1) one-stage leaching, (2) two-stage leaching, (3) neutralization and indium precipitation, and (4) neutralization and iron removal and zinc recovery. The method provided by the application has the advantages of short process, low cost, no three-waste emissions and full recovery of liquid slag, and can reduce environmental pollution and create economic benefits; in the two-stage leaching stage, high-grade sulfur, iron concentrate and other concentrates can be obtained by physical screening of leaching slag, without the need for a flotation process, so that the process flow is greatly shortened, and the cost is obviously reduced.

Description

Technical Field

[0001] This application belongs to the field of hydrometallurgical technology, specifically relating to a selective pressure leaching method for zinc sulfide concentrate. Background Technology

[0002] In the field of non-ferrous metal smelting, zinc extraction technology has always been a key focus of research and development. Zinc, as an extremely important non-ferrous metal, is used in automobile manufacturing for galvanizing to enhance the corrosion resistance of steel; in the battery industry, it is a key component of zinc-manganese batteries and zinc-air batteries; and in the construction sector, it is widely used in protective materials for roofs and walls. With the continuous development of various industries, the demand for zinc is steadily increasing, driving continuous innovation in zinc smelting technology. Traditionally, the extraction of metallic zinc has primarily used zinc sulfide concentrate as the main raw material. This raw material has a complex composition, typically containing 43-51% zinc, 7-15% iron, 28-32% sulfur, and other elements such as copper, cadmium, indium, tin, and lead, accounting for approximately 9-12%. Regarding the processing of zinc sulfide concentrate, two mainstream smelting methods have gradually emerged: pyrometallurgical + hydrometallurgical and hydrometallurgical + pyrometallurgical methods.

[0003] The first type of pyrometallurgical + hydrometallurgical process involves pyrometallurgical roasting of zinc sulfide concentrate, followed by the roasted ore entering the hydrometallurgical process. Through leaching, purification, electrowinning, and casting, zinc ingots are obtained. The leaching residue is then subjected to pyrometallurgical treatment. Waste heat from the roasting flue gas is recovered for hydrometallurgical applications, and the remaining flue gas is used to produce sulfuric acid. Other valuable metals are recovered during purification, and iron slag is either recycled or treated harmlessly. Its advantages include high zinc recovery, sufficient heat generated from sulfur to meet the requirements of the hydrometallurgical process, and the production of sulfuric acid as a byproduct while recovering various valuable metals. However, this process has significant drawbacks: a long process flow, complex procedures, low recovery rates of indium and other valuable metals, high requirements for the sulfur content of raw materials, high energy consumption, and difficulty in application in areas where sulfuric acid has no market.

[0004] The second process combines wet and pyrometallurgical methods. Zinc sulfide concentrate is finely ground and then subjected to two-stage pressure leaching. The first-stage leaching solution is purified, electrowinning, and cast into zinc ingots. The leaching residue enters the second-stage leaching, and the second-stage leaching solution is returned to the first stage. The second-stage leaching residue is used for sulfur recovery. During pressure leaching, most sulfur is converted to sulfur, with some forming sulfuric acid. Valuable metals are recovered during purification, and iron slag is further processed. The advantage of this process is that sulfur is converted into easily stored and transportable sulfur, making it suitable for regions with poor sulfuric acid sales. It also offers high leaching rates for indium and other valuable elements. However, its disadvantages are also significant: the conversion of sulfur to sulfur during pressure leaching leads to insufficient heat, requiring additional steam and resulting in high energy consumption; iron processing requires even more heat, necessitating additional iron removal or pyrometallurgical processes, further increasing energy consumption. Simultaneously, the iron slag generation process causes the loss of various valuable metals, leading to a reduced recovery rate. While some patents propose using pressure leaching to treat high-arsenic zinc sulfide concentrate, this method has the following shortcomings: it does not mention the recovery of valuable elements such as Cu, Fe, and In; sulfur is discarded in elemental form in the slag and is not recovered; and the reaction process requires high-purity oxygen, leading to high production costs. Ultimately, this results in low utilization value of zinc sulfide concentrate and unclear economic benefits.

[0005] In summary, existing processes have low overall utilization rates for zinc sulfide concentrate, resulting in problems such as insufficient processing space for zinc smelting enterprises, low smelting production capacity, excessive waste residue, high energy consumption, and environmental pollution. Therefore, there is an urgent need for an innovative technology and method to improve these problems. Summary of the Invention

[0006] Therefore, the technical problem to be solved by this application is to overcome the above-mentioned defects in the prior art, thereby providing a selective pressure leaching method for zinc sulfide concentrate.

[0007] Therefore, this application achieves the above objectives through the following technical solution:

[0008] This application provides a method for selective pressure leaching of complex zinc sulfide concentrates, the process flow of which is shown below. Figure 1 It includes the following steps:

[0009] S1, zinc sulfide concentrate is mixed with the first leaching solution and subjected to a first-stage leaching to obtain a first-stage filtrate and a first-stage leaching residue; the temperature of the first-stage leaching is 150-165℃, the time is 1-3h, and the oxygen partial pressure is 0.3-0.8MPa;

[0010] S2, the first-stage leaching residue is mixed with the second-stage leaching solution to perform a second-stage leaching, resulting in a second-stage filtrate and a second-stage leaching residue; the second-stage leaching temperature is 140-155℃, the time is 2-4h, and the oxygen partial pressure is 0.1-0.6MPa; the second-stage leaching residue is sorted to obtain iron concentrate, sulfur, and other valuable metal concentrates;

[0011] S3, mix a section of filtrate with the first neutralizing agent to neutralize and precipitate indium, and obtain indium residue and neutralized filtrate;

[0012] S4, the neutralized filtrate is mixed with the second neutralizing agent, and an oxidizing agent is added to neutralize and remove iron, resulting in iron slag and zinc-rich filtrate. The iron slag is returned to a first-stage leaching process, and the zinc-rich filtrate is electrolyzed to recover zinc products and waste electrolyte.

[0013] As an example, the leaching temperature can be 150℃, 152℃, 154℃, 156℃, 158℃, 160℃, 162℃, 165℃, or within any range of the above values; the leaching time can be 1h, 1.5h, 2h, 2.5h, 3h, or within any range of the above values; the oxygen partial pressure during leaching can be 0.3MPa, 0.35MPa, 0.4MPa, 0.45MPa, 0.5MPa, 0.55MPa, 0.6MPa, 0.65MPa, 0.7MPa, 0.75MPa, 0.8MPa, or within any range of the above values.

[0014] In some alternative embodiments, the acidity of the first leachate is 40 to 70 g / L; as an example, the acidity of the first leachate may be 40 g / L, 45 g / L, 50 g / L, 55 g / L, 60 g / L, 65 g / L, 70 g / L, or within any of the above values.

[0015] It should be noted that in this application, the term "acidity" refers to the mass of sulfuric acid contained in 1L of leachate. For example, an acidity of 40g / L means that 1L of leachate contains 40g of sulfuric acid.

[0016] In some alternative embodiments, the first leachate includes at least one of the waste electrolyte generated from the electrolytic recovery step or the second-stage filtrate.

[0017] As an example, the temperature of the second-stage leaching can be 140℃, 142℃, 144℃, 146℃, 148℃, 150℃, 152℃, 155℃, or within any range of the above values; the leaching time can be 2h, 2.5h, 3h, 3.5h, 4h, or within any range of the above values; the oxygen partial pressure of the second-stage leaching can be 0.1MPa, 0.15MPa, 0.2MPa, 0.25MPa, 0.3MPa, 0.35MPa, 0.4MPa, 0.45MPa, 0.5MPa, 0.55MPa, 0.6MPa, or within any range of the above values.

[0018] In some alternative embodiments, the acidity of the second leachate is 50 to 70 g / L; as an example, the acidity of the second leachate is 50 g / L, 52 g / L, 54 g / L, 58 g / L, 60 g / L, 62 g / L, 64 g / L, 66 g / L, 68 g / L, 70 g / L, or within any range of the above values;

[0019] It should be noted that the control of acidity and oxygen partial pressure in the two-stage leaching is mainly due to the following considerations: a low oxygen partial pressure results in insufficient oxidation of S in the reaction system, low system acidity, and ultimately a low leaching rate of Zn, In, etc., which leads to a decrease in the grade of the concentrate in the subsequent process.

[0020] In some alternative embodiments, the second-stage leachate includes the waste electrolyte generated from the electrolytic recovery step.

[0021] As an example, the mass ratio of the first leachate to zinc sulfide concentrate can be 3:1, 4:1, 5:1, 6:1, 7:1, 8:1, 9:1, 10:1, or within any of the above values; the mass ratio of the second leachate to the first stage of leaching residue can be 3:1, 4:1, 5:1, 6:1, 7:1, 8:1, 9:1, 10:1, or within any of the above values.

[0022] It should be noted that the leaching residue obtained from one leaching stage is in a wet state. Its moisture content must be measured first, and then the mass of the dry residue of the leaching stage must be calculated.

[0023] As an example, the endpoint pH of the indium precipitation can be 4.0, 4.1, 4.2, 4.3, 4.4, 4.5, 4.6, 4.7, 4.8, or within any range of the above values; the reaction temperature can be 45℃, 48℃, 50℃, 55℃, 58℃, 60℃, 63℃, 65℃, or within any range of the above values; the reaction time can be 2h, 2.5h, 3h, 3.5h, 4h, 4.5h, 5h, or within any range of the above values.

[0024] In some alternative embodiments, the oxidant in the neutralization and iron removal reaction is at least one of air, oxygen, or hydrogen peroxide.

[0025] As an example, the endpoint pH for iron removal can be 5.1, 5.2, 5.3, 5.4, 5.5, or within any range of the above values; the reaction temperature can be 40℃, 41℃, 42℃, 43℃, 44℃, 45℃, 46℃, 47℃, 48℃, 49℃, 50℃, or within any range of the above values; the reaction time can be 1.2h, 1.4h, 1.6h, 1.8h, 2.0h, 2.2h, 2.4h, 2.6h, 2.8h, 3.0h, or within any range of the above values.

[0026] It should be noted that this application uses a method of indium precipitation followed by iron removal, because a large amount of Fe is present in the indium precipitation stage of this application system. 2+ The ions precipitate at a relatively high pH value, higher than Fe. 3+ The pH value for ion precipitation is important; therefore, in this application, indium is first neutralized and precipitated, followed by Fe precipitation. 2+ The ions are oxidized, neutralized, and precipitated. In the subsequent neutralization and iron removal process, an oxidant is added along with the neutralizing agent to remove iron.

[0027] In some alternative embodiments, the first neutralizing agent and the second neutralizing agent independently comprise at least one of quicklime, hydrated lime, or limestone.

[0028] In some optional embodiments, the iron concentrate obtained after the two-stage leaching residue is sorted contains ≥60 wt.% Fe and ≥90 wt.% S in the sulfur.

[0029] It should be noted that the Fe content in iron concentrate can be tested using methods and equipment known in the field, such as either atomic absorption spectrometry (AAS) or inductively coupled plasma atomic emission spectrometry (ICP); the S content in sulfur can be tested using methods and equipment known in the field, such as X-ray fluorescence spectrometry.

[0030] In some optional embodiments, the other valuable metal concentrate includes at least one of lead, tin, and silver, wherein the Sn content is ≥20 wt.% or the Pb content is ≥3 wt.% or the Ag content is ≥0.015 wt.%.

[0031] It should be noted that the types of metal elements contained in other valuable metal concentrates are determined by the composition of zinc sulfide concentrate. If zinc sulfide concentrate contains only one other valuable metal element, then one of lead concentrate, tin concentrate, or silver concentrate can be obtained; if zinc sulfide concentrate contains multiple other valuable metal elements, then the obtained other valuable metal concentrates contain multiple other valuable metals.

[0032] In some alternative embodiments, the zinc sulfide concentrate comprises, by mass percentage: 43-51% zinc, 7-15% iron, 28-32% sulfur, and 9-12% other valuable metals, including at least one of copper, cadmium, indium, tin, and lead.

[0033] The technical solution of this application has the following advantages:

[0034] 1. The selective pressure leaching method for complex zinc sulfide concentrate provided in this application is based on a two-stage oxygen pressure leaching process under low-acid conditions. By controlling the operating parameters of the two stages of leaching, the first stage of leaching efficiently solidifies iron ions in the mineral into the leaching residue, thereby significantly improving the grade of the iron concentrate. The second stage of leaching, through precise control of oxygen partial pressure, leaching time, and temperature, cleverly controls the form of elemental sulfur, ensuring that the sulfur in the leaching residue exists primarily as high-purity sulfur. This design brings multiple advantages: First, the iron ion concentration in the first-stage filtrate is extremely low, resulting in low iron content in the indium slag produced during the subsequent neutralization and indium precipitation process, thus improving the grade of the indium slag. Second, without the need for traditional flotation steps, high-purity, uniformly sized, and highly fluid sulfur products can be directly recovered from the second-stage leaching residue through physical sieving alone, greatly shortening the process flow and significantly reducing costs. Furthermore, this method utilizes the characteristics of different products (such as the phase transition behavior of sulfur at different temperatures) to effectively control the purity, morphology, and particle size of sulfur by adjusting key parameters such as oxygen partial pressure and temperature, thereby improving product quality and separation efficiency. This application, starting from a two-stage leaching process, maximizes the utilization of zinc sulfide concentrate, producing high-quality iron slag, copper slag, indium slag, sulfur, and zinc-rich solution. It effectively solves the problems of insufficient processing space, low smelting capacity, and high energy consumption due to excessive waste in zinc smelting enterprises, while avoiding environmental pollution and possessing significant social benefits. It is of great significance for the comprehensive utilization of zinc sulfide concentrate. The selective pressure leaching method for complex zinc sulfide concentrate provided in this application can be applied to other types of sulfide ores, such as nickel sulfide ores, sulfide gold ores, sulfide copper ores, and is easy to promote.

[0035] 2. The selective pressure leaching method for complex zinc sulfide concentrates provided in this application involves both leaching stages conducted under relatively low acid conditions. Maintaining this acidity effectively allows for the selective leaching of different elements in the mineral. For example, during the first leaching stage, the low acidity allows for the leaching of Zn, while Fe remains in the residue. This reduces overall acid consumption and also allows for a reduction in the amount of neutralizing agent used in the subsequent neutralization stage, thus lowering costs.

[0036] 3. The selective pressure leaching method for complex zinc sulfide concentrate provided in this application can effectively separate indium from iron by adjusting the parameters of the neutralization precipitation and neutralization iron removal steps, thereby improving the effective utilization value of the neutralization slag and reducing the difficulty of separating effective elements in the subsequent mixed slag.

[0037] 4. The selective pressure leaching method for complex zinc sulfide concentrate provided in this application allows the waste electrolyte to be directly reused in the first and second stage leaching steps, and the second stage filtrate to be directly reused in the first stage leaching step. This method combines the advantages of short process, low cost, no waste discharge, and full recovery of liquid residue. It directly solidifies iron and utilizes waste electrolyte without the need to purchase sulfuric acid, which reduces environmental pollution and creates economic benefits. Attached Figure Description

[0038] To more clearly illustrate the technical solutions in the specific embodiments of this application or the prior art, the drawings used in the description of the specific embodiments or the prior art will be briefly introduced below. Obviously, the drawings described below are some embodiments of this application. For those skilled in the art, other drawings can be obtained from these drawings without creative effort.

[0039] Figure 1 This is a process flow diagram of the selective pressure leaching method for zinc sulfide concentrate provided in this application. Detailed Implementation

[0040] The following embodiments are provided to better understand this application and are not limited to the preferred embodiments described herein. They do not constitute a limitation on the content and scope of protection of this application. Any product that is the same as or similar to this application, derived by any person based on the teachings of this application or by combining features of this application with other prior art, falls within the scope of protection of this application.

[0041] Unless otherwise defined, all technical and scientific terms used herein have the same meaning as commonly understood by one of ordinary skill in the art to which this application belongs; the terminology used herein is for the purpose of describing particular embodiments only and is not intended to limit the application; the terms “comprising” and “having” and any variations thereof in the text of this application are intended to cover non-exclusive inclusion.

[0042] In this document, the term "embodiment" means that a particular feature, structure, or characteristic described in connection with an embodiment may be included in at least one embodiment of this application. The appearance of this phrase in various places throughout the specification does not necessarily refer to the same embodiment, nor is it a separate or alternative embodiment mutually exclusive with other embodiments. It will be explicitly and implicitly understood by those skilled in the art that the embodiments described herein can be combined with other embodiments.

[0043] The "range" disclosed in this application is defined by a lower limit and an upper limit. A given range is defined by selecting a lower limit and an upper limit, which define the boundaries of the specific range. Ranges defined in this way can include or exclude endpoints and can be arbitrarily combined; that is, any lower limit can be combined with any upper limit to form a range. In this application, unless otherwise stated, the numerical range "ab" represents a shortened representation of any combination of real numbers from a to b, where a and b are real numbers. For example, the numerical range "0-5" indicates that all real numbers between "0-5" have been listed herein, and "0-5" is merely a shortened representation of these numerical combinations. Furthermore, when a parameter is described as an integer ≥ 2, it is equivalent to disclosing that the parameter can be, for example, integers 2, 3, 4, 5, 6, 7, 8, 9, 10, 11, 12, etc.

[0044] In the description of the embodiments of this application, the term "and / or" is merely a description of the relationship between related objects, indicating that there can be three relationships. For example, A and / or B can represent three situations: A exists alone, A and B exist simultaneously, and B exists alone.

[0045] In the description of the embodiments of this application, the term "at least one" refers to one or more (including two).

[0046] Unless otherwise specified, all steps in this application may be performed sequentially or randomly, preferably sequentially. For example, the method includes steps (a) and (b), indicating that the method may include steps (a) and (b) performed sequentially, or it may include steps (b) and (a) performed sequentially. For example, the mention that the method may also include step (c) indicates that step (c) may be added to the method in any order. For example, the method may include steps (a), (b), and (c), or it may include steps (a), (c), and (b), or it may include steps (c), (a), and (b), etc.

[0047] For experiments not specifically described in the examples, the procedures or conditions should be followed according to the conventional experimental procedures described in the literature in this field. Reagents or instruments whose manufacturers are not specified are all commercially available conventional reagent products.

[0048] The present application will now be described with reference to specific embodiments. It should be noted that these embodiments are merely descriptive and do not limit the present application in any way.

[0049] Example 1

[0050] This embodiment provides a selective pressure leaching method for zinc sulfide concentrate, and its process flow diagram is as follows: Figure 1As shown, the specific steps and operating parameters are as follows:

[0051] Zinc sulfide concentrate (composition: Zn 47%, Fe 12%, S 29%, In 0.06%, Sn 0.65%, and elements such as C, H, and O) was leached with a first leaching solution of 60 g / L. A first-stage leaching was performed under the conditions of a liquid-to-solid mass ratio of 5:1, a reaction temperature of 150℃, a reaction time of 1.5 h, and an oxygen partial pressure of 0.6 MPa. The resulting first-stage leaching residue was then leached with a second leaching solution of 60 g / L under the conditions of a liquid-to-solid mass ratio of 5:1, a temperature of 150℃, a reaction time of 3.5 h, and an oxygen partial pressure of 0.3 MPa. The second-stage filtrate returned from the first-stage leaching of the zinc sulfide concentrate had an acidity of 48 g / L. The resulting second-stage leaching residue, after sieving, yielded an iron concentrate content of 52 wt.%, a sulfur content of 96 wt.%, and a tin concentrate content of 21 wt.%.

[0052] Limestone was added to the first filtrate to adjust the pH to 4.5 for neutralization and indium precipitation reaction. The reaction temperature was 55℃ and the reaction time was 3.0h. The indium slag obtained from the indium removal and recovery process had an indium content of 0.12wt.%.

[0053] Air was introduced into the neutralized filtrate, and limestone was added to adjust the pH to 5.0 for neutralization and iron removal reaction. The reaction temperature was 50℃ and the reaction time was 3.0h. The resulting iron slag contained 46wt.% Fe and was returned to the first-stage leaching process for recovery. The zinc-rich filtrate contained 95g / L of Zn and was then electrolytically recovered. The waste electrolyte returned from the zinc electrolysis recovery process contained 43g / L of Zn and had an acidity of 60g / L.

[0054] Example 2

[0055] This embodiment provides a selective pressure leaching method for zinc sulfide concentrate, and its process flow diagram is as follows: Figure 1 As shown, the specific steps and operating parameters are as follows:

[0056] Zinc sulfide concentrate (composition: Zn 48%, Fe 11%, S 30%, In 0.05%, Pb 0.55%, and elements such as C, H, and O) was leached with a first leaching solution of 40 g / L. A first-stage leaching was performed under the conditions of a liquid-to-solid mass ratio of 3:1, a reaction temperature of 165℃, a reaction time of 1.0 h, and an oxygen partial pressure of 0.3 MPa. The resulting first-stage leaching residue was then leached with a second leaching solution of 40 g / L under the conditions of a liquid-to-solid mass ratio of 3:1, a temperature of 140℃, a reaction time of 4.0 h, and an oxygen partial pressure of 0.6 MPa. The second-stage filtrate returned from the first-stage leaching of the zinc sulfide concentrate had an acidity of 40 g / L. The resulting second-stage leaching residue, after sieving, yielded an iron concentrate content of 51 wt.%, a sulfur content of 90 wt.%, and a lead concentrate content of 3.2 wt.%.

[0057] Quicklime was added to the first filtrate to adjust the pH to 4.0 for neutralization and indium precipitation reaction. The reaction temperature was 45℃ and the reaction time was 2.0h. The indium slag obtained from the indium removal and recovery process had an indium content of 0.13wt.%.

[0058] Oxygen was introduced into the neutralized filtrate, and quicklime was added to adjust the pH to 5.2 for neutralization and iron removal reaction. The reaction temperature was 40℃ and the reaction time was 1.0h. The iron slag obtained contained 48wt.% Fe and was returned to the first-stage leaching process for recovery. The concentration of Zn in the obtained zinc-rich filtrate was 92g / L.

[0059] The waste electrolyte returned from the zinc recovery process has a Zn concentration of 44 g / L and an acidity of 40 g / L.

[0060] Example 3

[0061] This embodiment provides a selective pressure leaching method for zinc sulfide concentrate, and its process flow diagram is as follows: Figure 1 As shown, the specific steps and operating parameters are as follows:

[0062] Zinc sulfide concentrate (composition: Zn 49%, Fe 14%, S 31%, In 0.05%, Ag 0.0065%, and elements C, H, O, etc.) was leached with a first leaching solution of 70 g / L. A first-stage leaching was performed under the conditions of a liquid-to-solid mass ratio of 10:1, a reaction temperature of 155℃, a reaction time of 3.0 h, and an oxygen partial pressure of 0.8 MPa. The resulting first-stage leaching residue was then leached with a second leaching solution of 70 g / L under the conditions of a liquid-to-solid mass ratio of 10:1, a temperature of 155℃, a reaction time of 2.0 h, and an oxygen partial pressure of 0.1 MPa. The second-stage filtrate returned from the first-stage leaching of the zinc sulfide concentrate had an acidity of 60 g / L. The resulting second-stage leaching residue, after sieving, yielded an iron concentrate with a Fe content of 53 wt.% and a sulfur content of [missing information]. 94 wt.%, silver concentrate content was 156 g / t (Ag);

[0063] Quicklime was added to the first filtrate to adjust the pH to 5.0 for neutralization and indium precipitation reaction. The reaction temperature was 65℃ and the reaction time was 5.0h. The indium slag obtained from the indium recovery process had an indium content of 0.10wt.%.

[0064] Hydrogen peroxide was added to the neutralized filtrate, and quicklime was added to adjust the pH to 5.5 for neutralization and iron removal reaction. The reaction temperature was 45℃ and the reaction time was 2.0h. The resulting iron slag had an Fe content of 45wt.%, which was returned to the first-stage leaching process for recovery. The Zn concentration in the resulting zinc-rich filtrate was 94g / L. The waste electrolyte returned from the zinc recovery process had a Zn concentration of 42g / L and an acidity of 70g / L.

[0065] Example 4

[0066] This embodiment provides a selective pressure leaching method for zinc sulfide concentrate, and its process flow diagram is as follows: Figure 1 As shown, the specific steps and operating parameters are as follows:

[0067] Zinc sulfide concentrate (composition: Zn 49%, Fe 14%, S 31%, In 0.05%, Ag 0.0065%, and elements C, H, O, etc.) was leached with a first leaching solution of 40 g / L. A first-stage leaching was performed under the conditions of a liquid-to-solid mass ratio of 6:1, a reaction temperature of 165℃, a reaction time of 2.0 h, and an oxygen partial pressure of 0.3 MPa. The resulting first-stage leaching residue was then leached with a second leaching solution of 60 g / L under the conditions of a liquid-to-solid mass ratio of 10:1, a temperature of 150℃, a reaction time of 2.0 h, and an oxygen partial pressure of 0.4 MPa. The second-stage filtrate returned from the first-stage leaching of the zinc sulfide concentrate had an acidity of 42 g / L. The resulting second-stage leaching residue, after sieving, yielded an iron concentrate with a Fe content of 50 wt.% and a sulfur content of [missing information]. 92 wt.%, silver concentrate content was 154 g / t (Ag);

[0068] Quicklime was added to the first filtrate to adjust the pH to 5.0 for neutralization and indium precipitation reaction. The reaction temperature was 55℃ and the reaction time was 2.0h. The indium slag obtained from the indium removal and recovery process had an indium content of 0.12wt.%.

[0069] Air was introduced into the neutralized filtrate, and quicklime was added to adjust the pH to 5.0 for neutralization and iron removal reaction. The reaction temperature was 50℃ and the reaction time was 3.0h. The resulting iron slag contained 47wt.% Fe and was returned to the first-stage leaching process for recovery. The zinc-rich filtrate contained 93g / L of Zn. The waste electrolyte returned from the zinc recovery process contained 41g / L of Zn and 40g / L of acidity.

[0070] Example 5

[0071] This embodiment provides a selective pressure leaching method for zinc sulfide concentrate, and its process flow diagram is as follows: Figure 1 As shown, the specific steps and operating parameters are as follows:

[0072] Zinc sulfide concentrate (composition: Zn 49%, Fe 14%, S 31%, In 0.05%, Ag 0.0065%, and elements C, H, O, etc.) was leached with a first leaching solution of 65 g / L. A first-stage leaching was performed under the conditions of a liquid-to-solid mass ratio of 5:1, a reaction temperature of 150℃, a reaction time of 2.5 h, and an oxygen partial pressure of 0.8 MPa. The resulting first-stage leaching residue was then leached with a second leaching solution of 55 g / L under the conditions of a liquid-to-solid mass ratio of 3:1, a temperature of 145℃, a reaction time of 4.0 h, and an oxygen partial pressure of 0.5 MPa. The second-stage filtrate returned from the first-stage leaching of the zinc sulfide concentrate had an acidity of 58 g / L. The resulting second-stage leaching residue, after sieving, yielded an iron concentrate with a Fe content of 54 wt.% and a sulfur content of [missing information - likely a percentage]. 91 wt.%, silver concentrate content was 165 g / t (Ag);

[0073] Quicklime was added to the first filtrate to adjust the pH to 4.0 for neutralization and indium precipitation reaction. The reaction temperature was 60℃ and the reaction time was 5.0h. The indium slag obtained from the indium recovery process had an indium content of 0.12wt.%.

[0074] Air was introduced into the neutralized filtrate, and quicklime was added to adjust the pH to 5.1 for neutralization and iron removal reaction. The reaction temperature was 45℃ and the reaction time was 4.0h. The resulting iron slag contained 45wt.% Fe and was returned to the first-stage leaching process for recovery. The Zn concentration in the resulting zinc-rich filtrate was 91g / L. The waste electrolyte returned from the zinc recovery process had a Zn concentration of 41g / L and an acidity of 65g / L.

[0075] Example 6

[0076] This embodiment provides a selective pressure leaching method for zinc sulfide concentrate, and its process flow diagram is as follows: Figure 1 As shown, the specific steps and operating parameters are as follows:

[0077] Zinc sulfide concentrate (composition: Zn 49%, Fe 14%, S 31%, In 0.05%, Ag 0.0065%, and elements C, H, O, etc.) was leached with a first leaching solution of 50 g / L. A first-stage leaching was performed under the conditions of a liquid-to-solid mass ratio of 10:1, a reaction temperature of 160℃, a reaction time of 1.0 h, and an oxygen partial pressure of 0.5 MPa. The resulting first-stage leaching residue was then leached with a second leaching solution of 70 g / L under the conditions of a liquid-to-solid mass ratio of 6:1, a temperature of 140℃, a reaction time of 3.0 h, and an oxygen partial pressure of 0.6 MPa. The second-stage filtrate returned from the first-stage leaching of the zinc sulfide concentrate had an acidity of 57 g / L. The resulting second-stage leaching residue, after sieving, yielded an iron concentrate with a Fe content of 51 wt.% and a sulfur content of [missing information]. 93 wt.%, silver concentrate content was 157 g / t (Ag);

[0078] Quicklime was added to the first filtrate to adjust the pH to 4.5 for neutralization and indium precipitation reaction. The reaction temperature was 45℃ and the reaction time was 4.0h. The indium slag obtained from the indium removal and recovery process had an indium content of 0.14wt.%.

[0079] Air was introduced into the neutralized filtrate, and quicklime was added to adjust the pH to 5.4 for neutralization and iron removal reaction. The reaction temperature was 50℃ and the reaction time was 2.0h. The resulting iron slag contained 46wt.% Fe and was returned to the first-stage leaching process for recovery. The Zn concentration in the resulting zinc-rich filtrate was 91g / L. The waste electrolyte returned from the zinc recovery process had a Zn concentration of 42g / L and an acidity of 50g / L.

[0080] Example 7

[0081] This embodiment provides a selective pressure leaching method for zinc sulfide concentrate, and its process flow diagram is as follows: Figure 1 As shown, the specific steps and operating parameters are as follows:

[0082] Zinc sulfide concentrate (composition: Zn 49%, Fe 14%, S 31%, In 0.05%, Ag 0.0065%, and elements C, H, O, etc.) was leached with a first leaching solution of 55 g / L. A first-stage leaching was performed under the conditions of a liquid-to-solid mass ratio of 3:1, a reaction temperature of 155℃, a reaction time of 3.0 h, and an oxygen partial pressure of 0.55 MPa. The resulting first-stage leaching residue was then leached with a second leaching solution of 40 g / L under the conditions of a liquid-to-solid mass ratio of 7:1, a temperature of 155℃, a reaction time of 3.5 h, and an oxygen partial pressure of 0.1 MPa. The second-stage filtrate returned from the first-stage leaching of the zinc sulfide concentrate had an acidity of 45 g / L. The resulting second-stage leaching residue, after sieving, yielded an iron concentrate with a Fe content of 53 wt.% and a sulfur content of [missing information]. 92 wt.%, silver concentrate content was 152 g / t (Ag);

[0083] Quicklime was added to the first filtrate to adjust the pH to 4.7 for neutralization and indium precipitation reaction. The reaction temperature was 65℃ and the reaction time was 3.5h. The indium slag obtained from the indium removal and recovery process had an indium content of 0.12wt.%.

[0084] Air was introduced into the neutralized filtrate, and quicklime was added to adjust the pH to 5.3 for neutralization and iron removal reaction. The reaction temperature was 45℃ and the reaction time was 3.0h. The resulting iron slag contained 49wt.% Fe and was returned to the first-stage leaching process for recovery. The zinc-rich filtrate contained 94g / L of Zn. The waste electrolyte returned from the zinc recovery process contained 43g / L of Zn and 55g / L of acidity.

[0085] Example 8

[0086] This embodiment provides a selective pressure leaching method for zinc sulfide concentrate, and its process flow diagram is as follows: Figure 1 As shown, the specific steps and operating parameters are as follows:

[0087] Zinc sulfide concentrate (composition: Zn 49%, Fe 14%, S 31%, In 0.05%, Ag 0.0065%, and elements C, H, O, etc.) was leached with a first leaching solution of 68 g / L. A first-stage leaching was performed under the conditions of a liquid-to-solid mass ratio of 6:1, a reaction temperature of 150℃, a reaction time of 1.5 h, and an oxygen partial pressure of 0.8 MPa. The resulting first-stage leaching residue was then leached with a second leaching solution of 65 g / L under the conditions of a liquid-to-solid mass ratio of 3:1, a temperature of 150℃, a reaction time of 4.0 h, and an oxygen partial pressure of 0.45 MPa. The second-stage filtrate returned from the first-stage leaching of the zinc sulfide concentrate had an acidity of 55 g / L. The resulting second-stage leaching residue, after sieving, yielded an iron concentrate with an Fe content of 51 wt.% and a sulfur content of [missing information]. 94 wt.%, silver concentrate content was 161 g / t (Ag);

[0088] Quicklime was added to the first filtrate to adjust the pH to 4.0 for neutralization and indium precipitation reaction. The reaction temperature was 55℃ and the reaction time was 5.0h. The indium slag obtained from the indium removal and recovery process had an indium content of 0.16wt.%.

[0089] Air was introduced into the neutralized filtrate, and quicklime was added to adjust the pH to 5.1 for neutralization and iron removal reaction. The reaction temperature was 45℃ and the reaction time was 3.0h. The resulting iron slag contained 49wt.% Fe and was returned to the first-stage leaching process for recovery. The zinc-rich filtrate had a Zn concentration of 94g / L. The waste electrolyte returned from the zinc recovery process had a Zn concentration of 43g / L and an acidity of 68g / L.

[0090] Example 9

[0091] This embodiment provides a selective pressure leaching method for zinc sulfide concentrate, and its process flow diagram is as follows: Figure 1 As shown, the specific steps and operating parameters are as follows:

[0092] Zinc sulfide concentrate (composition: Zn 49%, Fe 14%, S 31%, In 0.05%, Ag 0.0065%, and elements C, H, O, etc.) was leached with a first leaching solution of 45 g / L. A first-stage leaching was performed under the conditions of a liquid-to-solid mass ratio of 7:1, a reaction temperature of 165℃, a reaction time of 2.7 h, and an oxygen partial pressure of 0.3 MPa. The resulting first-stage leaching residue was then leached with a second leaching solution of 45 g / L under the conditions of a liquid-to-solid mass ratio of 10:1, a temperature of 147℃, a reaction time of 2.0 h, and an oxygen partial pressure of 0.35 MPa. The second-stage filtrate returned from the first-stage leaching of the zinc sulfide concentrate had an acidity of 41 g / L. The resulting second-stage leaching residue, after sieving, yielded an iron concentrate with a Fe content of 50 wt.% and a sulfur content of [missing information]. 90 wt.%, silver concentrate content is 158 g / t;

[0093] Quicklime was added to the first filtrate to adjust the pH to 5.0 for neutralization and indium precipitation reaction. The reaction temperature was 62℃ and the reaction time was 2.0h. The indium slag obtained from the indium removal and recovery process had an indium content of 0.13wt.%.

[0094] Air was introduced into the neutralized filtrate, and quicklime was added to adjust the pH to 5.5 for neutralization and iron removal reaction. The reaction temperature was 50℃ and the reaction time was 1.0 h. The resulting iron slag contained 45 wt.% Fe and was returned to the first-stage leaching process for recovery. The zinc-rich filtrate contained 90 g / L of Zn. The waste electrolyte returned from the zinc recovery process contained 41 g / L of Zn and 45 g / L of acidity.

[0095] Example 10

[0096] This embodiment provides a selective pressure leaching method for zinc sulfide concentrate, and its process flow diagram is as follows: Figure 1 As shown, the specific steps and operating parameters are as follows:

[0097] Zinc sulfide concentrate (composition: Zn 49%, Fe 14%, S 31%, In 0.05%, Ag 0.0065%, and elements C, H, O, etc.) was leached with a first leaching solution of 60 g / L. A first-stage leaching was performed under the conditions of a liquid-to-solid mass ratio of 10:1, a reaction temperature of 162℃, a reaction time of 1.0 h, and an oxygen partial pressure of 0.7 MPa. The resulting first-stage leaching residue was then leached with a second leaching solution of 70 g / L under the conditions of a liquid-to-solid mass ratio of 6:1, a temperature of 140℃, a reaction time of 2.5 h, and an oxygen partial pressure of 0.6 MPa. The second-stage filtrate returned from the first-stage leaching of the zinc sulfide concentrate had an acidity of 59 g / L. The resulting second-stage leaching residue, after sieving, yielded an iron concentrate with a Fe content of 53 wt.% and a sulfur content of [missing information]. 93 wt.%, silver concentrate content was 160 g / t (Ag);

[0098] Quicklime was added to the first filtrate to adjust the pH to 4.3 for neutralization and indium precipitation reaction. The reaction temperature was 45℃ and the reaction time was 2.5h. The indium slag obtained from the indium removal and recovery process had an indium content of 0.14wt.%.

[0099] Air was introduced into the neutralized filtrate, and quicklime was added to adjust the pH to 5.2 for neutralization and iron removal reaction. The reaction temperature was 42℃ and the reaction time was 3.0h. The resulting iron slag contained 46wt.% Fe and was returned to the first-stage leaching process for recovery. The zinc-rich filtrate had a Zn concentration of 94g / L. The waste electrolyte returned from the zinc recovery process had a Zn concentration of 42g / L and an acidity of 60g / L.

[0100] Comparative Example 1

[0101] The main difference between Comparative Example 1 and Example 1 is that the partial pressure of oxygen leaching in one stage is 0.1 MPa.

[0102] The results of the selective pressure leaching method for zinc sulfide concentrate in Comparative Example 1 are as follows:

[0103] The acidity of the second-stage leaching filtrate was 23 g / L; the indium slag produced by the neutralization and indium precipitation reaction had a content of 0.06 wt.%; the iron slag obtained from the neutralization and iron removal reaction had a content of 34 wt.%; and the Zn concentration in the resulting zinc-rich filtrate was 70 g / L. The waste electrolyte produced from the zinc-rich filtrate during zinc recovery had an acidity of 42 g / L and a Zn concentration of 38 g / L. The second-stage leaching residue obtained from the second-stage leaching reaction, after screening, showed an iron concentrate content of 43 wt.% Fe, a sulfur content of 84 wt.% S, and a tin concentrate content of 11 wt.% Sn.

[0104] Comparative Example 2

[0105] The main difference between Comparative Example 2 and Example 1 is that the leaching reaction temperature is 200°C.

[0106] The results of the selective pressure leaching method for zinc sulfide concentrate in Comparative Example 2 are as follows:

[0107] The acidity of the second-stage leaching filtrate was 30 g / L; the indium slag produced by the neutralization and indium precipitation reaction had a content of 0.10 wt.%; the iron slag obtained from the neutralization and iron removal reaction had a content of 24 wt.%; and the Zn concentration in the resulting zinc-rich filtrate was 93 g / L. The waste electrolyte produced from the zinc-rich filtrate during zinc recovery had an acidity of 44 g / L and a Zn concentration of 45 g / L. The second-stage leaching residue obtained from the second-stage leaching reaction, after screening, showed an iron concentrate content of 45 wt.% Fe, a sulfur content of 71 wt.% S, and a tin concentrate content of 13 wt.% Sn.

[0108] Comparative Example 3

[0109] The main difference between Comparative Example 3 and Example 1 is that the leaching reaction time is 4.0 h.

[0110] The results of the selective pressure leaching method for zinc sulfide concentrate in Comparative Example 3 are as follows:

[0111] The acidity of the second-stage leaching filtrate was 37 g / L; the indium slag produced by the neutralization and indium precipitation reaction had a content of 0.11 wt.%; the iron slag obtained from the neutralization and iron removal reaction had a content of 37 wt.%, and the Zn concentration in the resulting zinc-rich filtrate was 92 g / L; the waste electrolyte produced from the zinc-rich filtrate during zinc recovery had an acidity of 41 g / L and a Zn concentration of 43 g / L; the second-stage leaching residue obtained from the second-stage leaching reaction, after screening, showed an iron concentrate content of 51 wt.%, a sulfur content of 90 wt.%, and a tin concentrate content of 12 wt.%.

[0112] Comparative Example 4

[0113] The main difference between Comparative Example 4 and Example 1 is that the oxygen partial pressure during the two-stage leaching is 0.8 MPa.

[0114] The results of the selective pressure leaching method for zinc sulfide concentrate in Comparative Example 4 are as follows:

[0115] The acidity of the second-stage leaching filtrate was 62 g / L; the indium slag produced by the neutralization and indium precipitation reaction had a content of 0.13 wt.%; the iron slag obtained from the neutralization and iron removal reaction had a content of 46 wt.%; and the Zn concentration in the resulting zinc-rich filtrate was 93 g / L. The waste electrolyte produced from the zinc-rich filtrate during zinc recovery had an acidity of 43 g / L and a Zn concentration of 41 g / L. The second-stage leaching residue obtained from the second-stage leaching reaction, after screening, showed an iron concentrate content of 38 wt.% Fe, a sulfur content of 76 wt.% S, and a tin concentrate content of 22 wt.% Sn.

[0116] Comparative Example 5

[0117] The main difference between Comparative Example 5 and Example 1 is that the two-stage leaching reaction time is 1.0 h.

[0118] The results of the selective pressure leaching method for zinc sulfide concentrate in Comparative Example 5 are as follows:

[0119] The acidity of the second-stage leaching filtrate was 26 g / L; the indium slag produced by the neutralization and indium precipitation reaction had a content of 0.11 wt.%; the iron slag obtained from the neutralization and iron removal reaction had a content of 47 wt.%; and the Zn concentration in the resulting zinc-rich filtrate was 91 g / L. The waste electrolyte produced from the zinc-rich filtrate during zinc recovery had an acidity of 41 g / L and a Zn concentration of 40 g / L. The second-stage leaching residue obtained from the second-stage leaching reaction, after screening, showed an iron concentrate content of 35 wt.% Fe, a sulfur content of 72 wt.% S, and a tin concentrate content of 16 wt.% Sn.

[0120] Comparative Example 6

[0121] The main difference between Comparative Example 6 and Example 1 is that the two-stage leaching reaction temperature is 130°C.

[0122] The results of the selective pressure leaching method for zinc sulfide concentrate in Comparative Example 6 are as follows:

[0123] The acidity of the second-stage leaching filtrate was 23 g / L; the indium slag produced by the neutralization and indium precipitation reaction had a content of 0.12 wt.%; the iron slag obtained from the neutralization and iron removal reaction had a content of 46 wt.%, and the Zn concentration in the resulting zinc-rich filtrate was 92 g / L; the waste electrolyte produced from the zinc-rich filtrate during zinc recovery had an acidity of 46 g / L and a Zn concentration of 39 g / L; the second-stage leaching residue obtained from the second-stage leaching reaction, after screening, showed an iron concentrate content of 31 wt.%, a sulfur content of 67 wt.%, and a tin concentrate content of 13 wt.%.

[0124] Comparative Example 7

[0125] The main difference between Comparative Example 7 and Example 1 is that the partial pressure of oxygen leaching in one stage is 1.1 MPa.

[0126] The results of the selective pressure leaching method for zinc sulfide concentrate in Comparative Example 7 are as follows:

[0127] The acidity of the second-stage leaching filtrate was 70 g / L; the indium slag produced by the neutralization and indium precipitation reaction had a content of 0.08 wt.%; the iron slag obtained from the neutralization and iron removal reaction had a content of 43 wt.%; and the Zn concentration in the resulting zinc-rich filtrate was 94 g / L. The waste electrolyte produced from the zinc-rich filtrate during zinc recovery had an acidity of 50 g / L and a Zn concentration of 43 g / L. The second-stage leaching residue obtained from the second-stage leaching reaction, after screening, showed an iron concentrate content of 34 wt.% Fe, a sulfur content of 53 wt.% S, and a tin concentrate content of 23 wt.% Sn.

[0128] Comparative Example 8

[0129] The main difference between Comparative Example 8 and Example 1 is the addition of a first leachate with an acidity of 20 g / L.

[0130] The results of the selective pressure leaching method for zinc sulfide concentrate in Comparative Example 8 are as follows:

[0131] The acidity of the second-stage leaching filtrate was 21 g / L; the indium slag produced by the neutralization and indium precipitation reaction had a content of 0.07 wt.%; the iron slag obtained from the neutralization and iron removal reaction had a content of 41 wt.%, and the Zn concentration in the resulting zinc-rich filtrate was 75 g / L; the waste electrolyte produced from the zinc-rich filtrate during zinc recovery had an acidity of 37 g / L and a Zn concentration of 41 g / L; the second-stage leaching residue obtained from the second-stage leaching reaction, after screening, showed an iron concentrate content of 36 wt.%, a sulfur content of 63 wt.%, and a tin concentrate content of 14 wt.%.

[0132] The results from the examples and comparative examples show that the temperature, oxygen partial pressure, and reaction time of the first-stage leaching significantly affect the concentrations of Zn, Fe, and In, as well as the acidity, in the leachate, thus determining the grade of the neutralized indium slag and the neutralized iron slag. Simultaneously, these parameters also determine the content of each element in the leaching residue, thereby affecting the grade of the Sn concentrate obtained from the second-stage leaching and the content of Fe and S in the leaching residue. Similarly, the temperature, oxygen partial pressure, and reaction time of the second-stage leaching mainly determine the final grade of the iron concentrate, sulfur, and valuable metal concentrate, but have a relatively small impact on the grade of the neutralized indium slag, the neutralized iron slag, and the Zn concentration in the zinc-rich filtrate.

[0133] Obviously, the above embodiments are merely illustrative examples for clear explanation and are not intended to limit the implementation. Those skilled in the art will recognize that other variations or modifications can be made based on the above description. It is neither necessary nor possible to exhaustively list all possible implementations here. However, obvious variations or modifications derived therefrom are still within the scope of protection of this invention.

Claims

1. A method for selective pressure leaching of zinc sulfide concentrate, characterized in that, Includes the following steps: S1, zinc sulfide concentrate is mixed with the first leaching solution and subjected to a first-stage leaching to obtain a first-stage filtrate and a first-stage leaching residue; the temperature of the first-stage leaching is 150~165℃, the time is 1~3h, and the oxygen partial pressure is 0.3~0.8MPa; the acidity of the first leaching solution is 40~70g / L. S2, the first-stage leaching residue is mixed with the second-stage leaching solution for a second-stage leaching process, yielding a second-stage filtrate and a second-stage leaching residue; the second-stage leaching temperature is 140~155℃, the time is 2~4h, and the oxygen partial pressure is 0.1~0.6MPa; the second-stage leaching residue is sieved to obtain iron concentrate, sulfur, and other valuable metal concentrates; the acidity of the second-stage leaching solution is 50~70g / L; S3, mix a section of filtrate with the first neutralizing agent to neutralize and precipitate indium, and obtain indium residue and neutralized filtrate; S4, the neutralized filtrate is mixed with the second neutralizing agent and the oxidizing agent to neutralize and remove iron, resulting in iron slag and zinc-rich filtrate. The iron slag is returned to a first-stage leaching process, and the zinc-rich filtrate is electrolyzed to recover zinc products and waste electrolyte.

2. The selective pressure leaching method for zinc sulfide concentrate according to claim 1, characterized in that, In S1, the first leachate includes at least one of the waste electrolyte generated by the electrolytic recovery or the second-stage filtrate.

3. The selective pressure leaching method for zinc sulfide concentrate according to claim 1, characterized in that, In S2, the second-stage leachate includes the waste electrolyte generated from the electrolytic recovery.

4. The selective pressure leaching method for zinc sulfide concentrate according to claim 1, characterized in that, In S1, the mass ratio of the first leachate to zinc sulfide concentrate is 3~10:

1.

5. The selective pressure leaching method for zinc sulfide concentrate according to claim 1, characterized in that, In S2, the mass ratio of the second leachate to the dry leaching residue of the first stage is 3~10:

1.

6. The selective pressure leaching method for zinc sulfide concentrate according to claim 1, characterized in that, In S3, during the neutralization and indium precipitation reaction, the final pH of the indium precipitation is 4.0~4.8, the reaction temperature is 45~65℃, and the reaction time is 2~5h.

7. The selective pressure leaching method for zinc sulfide concentrate according to claim 1, characterized in that, In S4, during the neutralization and iron removal reaction, the final pH of the iron removal endpoint is 5.0~5.5, the reaction temperature is 40~50℃, and the reaction time is 1~3h. And / or, the oxidant includes at least one of air, oxygen or hydrogen peroxide.

8. The selective pressure leaching method for zinc sulfide concentrate according to claim 1, characterized in that, The first neutralizing agent and the second neutralizing agent independently include at least one of quicklime, hydrated lime, or limestone.

9. The selective pressure leaching method for zinc sulfide concentrate according to any one of claims 1-8, characterized in that, In S2, the other valuable metal concentrate includes at least one of lead, tin, and silver.

10. The selective pressure leaching method for zinc sulfide concentrate according to claim 9, characterized in that, In S1, the zinc sulfide concentrate comprises, by mass percentage: 43-51% zinc, 7-15% iron, 28-32% sulfur, and 9-12% other valuable metals, wherein the other valuable metals include at least one of copper, cadmium, indium, tin, and lead.

Images