Method for resource utilization of zinc hydrometallurgy iron slag

Abstract

The invention provides a zinc hydrometallurgy iron slag resource utilization method which comprises the following steps: mixing zinc hydrometallurgy iron slag, a chlorinating agent and an inhibitor, roasting the obtained mixed material, recovering gas-phase zinc chloride in the roasting process, and obtaining roasted clinker after roasting; leaching the roasted clinker, and carrying out solid-liquid separation to obtain a leachate; and crystallizing the leachate to obtain ferrous chloride. The method provided by the invention is simple in process, short in flow and low in energy consumption, zinc and iron components in the zinc hydrometallurgy iron slag are converted into ferrous chloride and zinc chloride with high added values, resource utilization of the zinc hydrometallurgy iron slag is realized, the problems of selective extraction and high-value utilization of zinc and iron in the zinc hydrometallurgy iron slag are solved, and the method is suitable for industrial production. And the method has good economic, social and environmental benefits.

Description

Technical Field

[0001] The present invention belongs to the technical field of solid waste treatment, and relates to a method for treating hydrometallurgical zinc smelting slag, and in particular to a method for resource utilization of hydrometallurgical zinc smelting slag. Background Art

[0002] Hydrometallurgical zinc smelting is a key method in zinc smelting. It converts zinc from zinc ore into high-purity zinc metal through leaching, purification, extraction, and electrolysis. Compared to traditional pyrometallurgical zinc smelting, this process offers lower energy consumption and better environmental adaptability, making it widely used in the zinc smelting industry.

[0003] During the leaching stage of hydrometallurgical zinc smelting, iron from the raw materials enters the leachate. To prevent iron ions from depositing during the electrolysis process and affecting the quality of the cathode zinc, the leachate must be deironed. Common deironing methods include the goethite method and the jarosite method. This deironing process produces a large amount of iron slag, which contains incompletely recovered zinc and concentrated harmful impurities such as arsenic and antimony. The complex distribution of iron in the slag, particularly the coexistence of zinc and iron, makes it difficult to obtain high-quality iron concentrate using traditional magnetic separation methods. Improper handling can significantly impact resource recovery and environmental protection.

[0004] Iron slag from hydrometallurgical zinc smelting is typically recovered through either pyrometallurgical or hydrometallurgical processes. Traditional pyrometallurgical processes often use carbon or carbon monoxide as a reducing agent to reduce iron-containing components to metallic iron for recycling. While this process is shorter, it consumes a lot of energy and poses secondary environmental risks. Hydrometallurgical processes are highly selective but cannot completely destroy the slag structure. The high impurity concentration in the leachate leads to a sharp increase in purification costs, resulting in low overall recovery efficiency. As a result, slag is often stockpiled or used at a low value, resulting in both resource waste and environmental risks.

[0005] CN109796049A discloses a method for preparing iron red using iron precipitate slag from the hydrometallurgical zinc smelting process. Fine-grained iron red pigment is prepared through processes such as high-temperature leaching with dilute sulfuric acid, flotation separation of lead and silver, and hydrothermal treatment. The leachate is then recycled to the hydrometallurgical zinc smelting system. Although this process achieves waste slag resource utilization, it has problems such as complex procedures, low conversion efficiency, and high production costs. In addition, the lack of systematic assessment of the zinc and cadmium residues in the goethite slag may lead to excessive heavy metal residues in the iron red product, and there is still an environmental risk.

[0006] CN116287713A discloses a method for treating iron-removing slag from hydrometallurgical zinc smelting. This method involves leaching the iron-removing slag and zinc concentrate with sulfuric acid, synergistically leaching the zinc therein and separating the iron from the iron-removing slag. While this method allows for the recovery of zinc from the iron-removing slag, it still fails to address the resource utilization of the iron-removing slag, resulting in low resource utilization.

[0007] Existing technologies are generally limited by bottlenecks such as low zinc-iron separation efficiency, poor product quality, and difficulty in controlling secondary pollution. Therefore, there is an urgent need to develop short-process, high-value utilization methods. Summary of the Invention

[0008] The purpose of the present invention is to provide a method for resource utilization of hydrometallurgical zinc slag, which can separate and transform the iron and zinc metal components in the hydrometallurgical zinc slag to achieve resource utilization.

[0009] In order to achieve the purpose of the invention, the present invention adopts the following technical solutions:

[0010] The present invention provides a method for resource utilization of hydrometallurgical zinc slag, the method comprising the following steps:

[0011] (1) mixing wet zinc smelting slag, a chlorinating agent, and an inhibitor, and roasting the resulting mixture; recovering vapor-phase zinc chloride during the roasting process; and obtaining roasted clinker after the roasting is completed;

[0012] (2) leaching the roasted clinker and performing solid-liquid separation to obtain a leachate;

[0013] (3) crystallizing the leachate to obtain ferrous chloride.

[0014] The composition of wet zinc smelting iron slag includes iron alum, goethite, yellow sodium alum, zinc ferrite, zinc sulfate, calcium sulfate, etc., and the physical phase composition is diverse. It also contains more impurities, especially iron-containing physical phases, and the types and structures are complex. The method provided by the present invention effectively destroys the iron slag structure by roasting. During the roasting process, under the synergistic effect of inhibitor and chlorinating agent, the iron-containing physical phase is converted into ferrous chloride and retained in the roasting material, and the zinc-containing physical phase is converted into volatile zinc chloride, so that the zinc component is recovered by gas phase separation, and effective selective separation of zinc is achieved. At the same time, the calcareous component (such as calcium sulfate) in the slag is stably present in the roasting material. Then, by controlling leaching, the iron (ferrous chloride) in the roasting material is separated from calcium (calcium sulfate) and other slag phases, and the iron component enters the leachate, so that the iron component is separated and finally obtained by crystallization.

[0015] The method provided by the present invention achieves efficient resource utilization of iron slag through roasting, condensation recovery, and leaching crystallization. This simple, short, and energy-efficient process converts the zinc and iron components in hydrometallurgical slag into high-value-added ferrous chloride and zinc chloride, achieving resource utilization of hydrometallurgical slag and resolving the challenges of selectively extracting and high-quality utilization of zinc and iron from hydrometallurgical slag. This method offers significant economic, social, and environmental benefits.

[0016] In the present invention, the composition of the hydrometallurgical zinc slag includes: 30-50wt% of iron, 5-25wt% of zinc, 1-8wt% of calcium, 5-15wt% of sulfur, and 10-30wt% of oxygen.

[0017] The iron content in the hydrometallurgical zinc slag is 30-50wt%, for example, it can be 30wt%, 32wt%, 34wt%, 35wt%, 36wt%, 38wt%, 40wt%, 42wt%, 44wt%, 45wt%, 46wt%, 48wt% or 50wt%, but is not limited to the listed values, and other values ​​not listed within the numerical range are also applicable.

[0018] The zinc content in the hydrometallurgical slag is 5-25wt%, for example, 5wt%, 10wt%, 15wt%, 20wt% or 25wt%, but is not limited to the listed values, and other values ​​not listed within the numerical range are also applicable.

[0019] The calcium content in the hydrometallurgical iron slag is 1-8wt%, for example, it can be 1wt%, 2wt%, 3wt%, 4wt%, 5wt%, 6wt%, 7wt% or 8wt%, but is not limited to the listed values. Other values ​​not listed within the numerical range are also applicable.

[0020] The sulfur content in the hydrometallurgical zinc slag is 5-15wt%, for example, 5wt%, 8wt%, 10wt%, 12wt% or 15wt%, but is not limited to the listed values, and other values ​​not listed within the numerical range are also applicable.

[0021] The oxygen content in the hydrometallurgical zinc slag is 10-30wt%, for example, it can be 10wt%, 15wt%, 20wt%, 25wt% or 30wt%, but is not limited to the listed values. Other values ​​not listed within the numerical range are also applicable.

[0022] Preferably, the chlorinating agent in step (1) comprises ammonium chloride.

[0023] Preferably, the amount of the chlorinating agent in step (1) is 20-50wt% of the mixed material, for example, it can be 20wt%, 25wt%, 30wt%, 35wt%, 40wt%, 45wt% or 50wt%, but is not limited to the listed values, and other values ​​not listed within the numerical range are also applicable.

[0024] Preferably, the inhibitor in step (1) comprises at least one of coke, anthracite or charcoal, and typical but non-limiting combinations include a combination of coke and anthracite, a combination of anthracite and charcoal, a combination of coke and charcoal, or a combination of coke, anthracite and charcoal.

[0025] Preferably, the amount of the inhibitor in step (1) is 1-5wt% of the mixed material, for example, it can be 1wt%, 2wt%, 3wt%, 4wt% or 5wt%, but is not limited to the listed values, and other values ​​not listed within the numerical range are also applicable.

[0026] Preferably, the calcination temperature in step (1) is 500-700°C, for example, it can be 500°C, 520°C, 540°C, 550°C, 560°C, 580°C, 600°C, 620°C, 640°C, 650°C, 660°C, 680°C or 700°C, but is not limited to the listed values, and other values ​​not listed within the numerical range are also applicable.

[0027] Preferably, the calcination time in step (1) is 60-150 min, for example, it can be 60 min, 70 min, 80 min, 90 min, 100 min, 110 min, 120 min, 130 min, 140 min or 150 min, but is not limited to the listed values, and other values ​​not listed within the numerical range are also applicable.

[0028] Preferably, the method of recovering the gaseous zinc chloride in step (1) includes condensation.

[0029] Preferably, the liquid-to-solid ratio of the leaching in step (2) is 1-5 mL / g, for example, 1 mL / g, 2 mL / g, 3 mL / g, 4 mL / g or 5 mL / g, but is not limited to the listed values, and other values ​​not listed within the numerical range are also applicable.

[0030] Preferably, water is used for leaching in step (2).

[0031] Preferably, the crystallization method in step (3) includes evaporation crystallization and / or cooling crystallization;

[0032] Preferably, the cooling crystallization method includes temperature gradient crystallization.

[0033] Compared with the prior art, the present invention has the following beneficial effects:

[0034] The method provided by the present invention has a simple process, a short flow and low energy consumption, and can convert the zinc and iron components in the hydrometallurgical zinc slag into high-value-added ferrous chloride and zinc chloride, thereby realizing the resource utilization of the hydrometallurgical zinc slag, solving the difficult problem of selective extraction and high-value utilization of zinc and iron in the hydrometallurgical zinc slag, and having good economic, social and environmental benefits. BRIEF DESCRIPTION OF THE DRAWINGS

[0035] Figure 1 This is a process flow chart of the method for resource utilization of hydrometallurgical zinc slag provided in Example 1. DETAILED DESCRIPTION

[0036] The technical solution of the present invention is further described below by way of specific embodiments. It should be understood by those skilled in the art that the embodiments are merely to help understand the present invention and should not be regarded as specific limitations of the present invention.

[0037] In order to clearly illustrate the technical solution of the present invention, in a specific embodiment, the composition of the hydrometallurgical slag used includes: 40.6wt% of iron, 19.2wt% of zinc, 3.1wt% of calcium, 10.1wt% of sulfur, 18.2wt% of oxygen, and 8.22wt% of others.

[0038] Example 1

[0039] This embodiment provides a Figure 1 The method for resource utilization of hydrometallurgical zinc slag shown in the figure comprises the following steps:

[0040] (1) Wet zinc smelting slag, ammonium chloride, and coke are mixed to obtain a mixed material, wherein the ammonium chloride content is 30 wt% and the coke content is 3 wt%. The mixed material is roasted at 650° C. for 100 min. During the roasting process, volatile gas is collected and condensed to obtain zinc chloride. After the roasting is completed, roasted clinker is obtained.

[0041] (2) leaching the obtained roasted clinker with water, controlling the liquid-to-solid ratio of the leaching to 3 mL / g, and filtering to obtain a leachate;

[0042] (3) The obtained leachate is evaporated and crystallized to obtain ferrous chloride.

[0043] Example 2

[0044] This embodiment provides a method for resource utilization of hydrometallurgical zinc slag, the method comprising the following steps:

[0045] (1) Wet zinc smelting slag, ammonium chloride, and coke are mixed to obtain a mixed material, wherein the ammonium chloride content is 50 wt% and the coke content is 5 wt%. The mixed material is roasted at 500° C. for 150 min. During the roasting process, volatile gas is collected and condensed to obtain zinc chloride. After the roasting is completed, roasted clinker is obtained.

[0046] (2) leaching the obtained roasted clinker with water, controlling the liquid-to-solid ratio of the leaching to 5 mL / g, and filtering to obtain a leachate;

[0047] (3) The obtained leachate is evaporated and crystallized to obtain ferrous chloride.

[0048] Example 3

[0049] This embodiment provides a method for resource utilization of hydrometallurgical zinc slag, the method comprising the following steps:

[0050] (1) Wet zinc smelting slag, ammonium chloride, and coke are mixed to obtain a mixed material, wherein the content of ammonium chloride is 20 wt % and the content of coke is 1 wt %. The mixed material is roasted at 700° C. for 60 min. During the roasting process, volatile gas is collected and condensed to obtain zinc chloride. After the roasting is completed, roasted clinker is obtained.

[0051] (2) leaching the obtained roasted clinker with water, controlling the liquid-to-solid ratio of the leaching to 1 mL / g, and filtering to obtain a leachate;

[0052] (3) The obtained leachate is evaporated and crystallized to obtain ferrous chloride.

[0053] Example 4

[0054] This embodiment provides a method for resource utilization of hydrometallurgical zinc slag. Compared with Example 1, the roasting temperature in step (1) is controlled to be 450° C., and the rest is the same as Example 1.

[0055] Example 5

[0056] This embodiment provides a method for resource utilization of hydrometallurgical zinc slag. Compared with Example 1, the roasting temperature in step (1) is controlled to be 850° C., and the rest is the same as Example 1.

[0057] Comparative Example 1

[0058] This comparative example provides a method for resource utilization of hydrometallurgical zinc smelting slag. Compared with Example 1, ammonium chloride is not added in step (1), that is, the mass of ammonium chloride is replaced by hydrometallurgical zinc smelting slag. The rest is the same as Example 1.

[0059] Comparative Example 2

[0060] This comparative example provides a method for resource utilization of hydrometallurgical slag. Compared with Example 1, no coke is added in step (1), that is, the mass of coke is replaced by hydrometallurgical slag, and the rest is the same as Example 1.

[0061] The purity and element recovery results of the products obtained in the examples and comparative examples are summarized in Table 1.

[0062] Table 1

[0063]

[0064]

[0065] “ / ” in the table indicates no data.

[0066] As can be seen from Table 1, the method provided by the present invention can obtain high value-added zinc chloride and ferrous chloride using wet zinc smelting slag as raw materials. Compared with Example 1, in Example 4, when the roasting temperature is too low, zinc cannot be completely converted into vapor-phase zinc chloride, resulting in a reduced zinc recovery rate; in Example 5, when the roasting temperature is too high, most of the ferrous chloride enters the vapor phase, causing the zinc chloride purity to decrease and the iron recovery rate to decrease. In Comparative Example 1, no chlorinating agent is added, zinc oxide does not volatilize at a relatively low temperature, and zinc and iron elements do not participate in the reaction during the subsequent water immersion process, making it impossible to recover zinc and iron elements; in Comparative Example 2, no inhibitor is added, causing the iron element to be partially converted into ferric chloride and enter the vapor phase, resulting in a decrease in zinc chloride purity and a low iron recovery rate.

[0067] In summary, the method provided by the present invention can convert the zinc and iron components in the hydrometallurgical slag into high-value-added ferrous chloride and zinc chloride, thereby realizing the resource utilization of the hydrometallurgical slag, solving the difficult problem of selective extraction and high-value utilization of zinc and iron in the hydrometallurgical slag, and having good economic, social and environmental benefits.

[0068] The applicant declares that the above is only a specific embodiment of the present invention, but the scope of protection of the present invention is not limited thereto. Those skilled in the art should understand that any changes or substitutions that can be easily thought of by those skilled in the art within the technical scope disclosed by the present invention fall within the scope of protection and disclosure of the present invention.

Claims

1. A method for resource utilization of hydrometallurgical zinc slag, characterized in that: The method comprises the following steps: (1) mixing wet zinc smelting slag, a chlorinating agent, and an inhibitor, and roasting the resulting mixture; recovering vapor-phase zinc chloride during the roasting process; and obtaining roasted clinker after the roasting is completed; (2) leaching the roasted clinker and performing solid-liquid separation to obtain a leachate; (3) crystallizing the leachate to obtain ferrous chloride.

2. The method according to claim 1, characterized in that The chlorinating agent in step (1) includes ammonium chloride.

3. The method according to claim 1 or 2, characterized in that The amount of the chlorinating agent in step (1) is 20-50wt% of the mixed material.

4. The method according to any one of claims 1 to 3, characterized in that The inhibitor in step (1) comprises at least one of coke, anthracite or charcoal.

5. The method according to any one of claims 1 to 4, characterized in that The amount of the inhibitor in step (1) is 1-5 wt% of the mixed material.

6. The method according to any one of claims 1 to 5, characterized in that The calcination temperature in step (1) is 500-700°C.

7. The method according to any one of claims 1 to 6, characterized in that The calcination time in step (1) is 60-150 min.

8. The method according to any one of claims 1 to 7, characterized in that The method for recovering the gaseous zinc chloride in step (1) includes condensation.

9. The method according to any one of claims 1 to 8, characterized in that The liquid-to-solid ratio of the leaching in step (2) is 1-5 mL / g; Preferably, water is used for leaching in step (2).

10. The method according to any one of claims 1 to 9, characterized in that The crystallization method in step (3) includes evaporation crystallization and / or cooling crystallization; Preferably, the cooling crystallization method includes temperature gradient crystallization.

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