Process for recycling multi-metal in zinc smelting rotary kiln slag

Abstract

This invention discloses a multi-metal resource recovery process for zinc smelting rotary kiln slag. The process involves sequentially treating the slag with two-stage magnetic separation, stepped acid leaching, copper leaching with oxidation, iron powder replacement, and cooling crystallization. Magnetic separation pre-enriches valuable metals, followed by low-acid oxygen leaching of copper, with copper recovery and arsenic precipitation occurring simultaneously during replacement. A three-stage closed-loop mother liquor circulation system crystallizes to produce compliant ferrous sulfate heptahydrate, with gold-silver slag and copper-arsenic slag as byproducts. This invention achieves a comprehensive gold, silver, and copper recovery rate of ≥72% and an iron leaching rate of 96.1%, overcoming the shortcomings of existing pure mineral processing processes such as inability to recover gold and copper, low iron recovery rates, and arsenic pollution. The entire process operates under mild conditions with low reagent consumption, enabling the high-value and harmless co-processing of multi-metal solid waste.

Description

Technical Field

[0001] This invention patent relates to the field of comprehensive utilization technology of solid waste resources in non-ferrous metallurgy, specifically to a multi-metal resource recovery process for zinc smelting rotary kiln slag. Background Technology

[0002] In the hydrometallurgical zinc smelting process, after the zinc leaching residue is subjected to high-temperature volatilization roasting in a rotary kiln to recover lead and zinc, a large amount of rotary kiln slag is generated. A single production line can produce up to 300 tons of slag per day. This slag is enriched with various valuable metals such as gold, silver, copper, and iron, making it a high-value secondary solid waste resource. During the high-temperature roasting process, the slag is in a semi-molten state. After subsequent water quenching, its overall hardness increases significantly. The valuable components inside are mainly present in the form of metals, alloys, and sulfides, with minerals intertwined and embedded. Conventional beneficiation and single acid leaching processes are insufficient for achieving comprehensive and efficient recovery of multiple metals. Currently, the industry only targets the recovery of volatile lead and zinc components. The low-grade gold, silver, copper, and iron in the slag are not effectively developed and utilized. This not only results in the loss of a large amount of valuable metals and significant economic losses, but also causes long-term stockpiling of slag to occupy land resources and lead to a series of environmental problems such as heavy metal leakage and water and soil pollution.

[0003] Among the existing publicly available technologies, patent CN 108531740A discloses a process for recovering lead, zinc, carbon, silver, and iron from zinc smelting leaching slag. This process adopts a combined beneficiation and smelting route of high-temperature reduction roasting, water quenching and grinding, carbon flotation, silver flotation, and magnetic separation for iron: first, lead and zinc are recovered by volatilization through high-temperature roasting at 900~1200℃; the roasted slag is then water quenched and wet-grinded for classification, and then residual carbon, silver, and iron are recovered sequentially by flotation. While this process achieves the stepwise recovery of lead, zinc, carbon, silver, and iron from the leaching residue, and the treated tailings can be used as cement ingredients for harmless disposal, demonstrating a certain resource utilization effect, the patent only employs physical beneficiation methods such as flotation and magnetic separation throughout the entire process, without setting up a copper recovery process, nor involving the recovery and utilization of gold. Considering the characteristics of the raw materials in this application and actual production needs, it still has several shortcomings and limitations, and cannot adapt to the comprehensive treatment requirements of this type of zinc smelting rotary kiln slag. The specific defects are as follows: 1. Complete lack of gold and copper recovery processes: The entire process uses only physical beneficiation methods such as flotation and magnetic separation, without any copper recovery process. The kiln slag also contains copper, mainly in the form of copper sulfide and elemental copper. Physical flotation is insufficient to effectively separate and recover the copper components, resulting in the complete loss of copper resources. Furthermore, the process does not recover associated gold, leading to the complete loss of gold resources and extremely low solid waste utilization.

[0004] 2. Low iron recovery rate: This process relies solely on magnetic separation to recover iron minerals. Due to the influence of mineral embedding, metal symbiosis, and mineral encapsulation, the overall iron recovery rate is only 22.38% at most. Most of the iron remains in the tailings, resulting in a significant waste of iron resources.

[0005] 3. Failure to achieve high-value utilization of iron products: The process ultimately only produces ordinary iron concentrate as a primary product, without deep processing of iron elements, making it impossible to produce high-value-added ferrous sulfate products, which greatly limits the overall economic benefits of the process.

[0006] 4. Poor adaptability to complex minerals: The kiln slag after high-temperature roasting in rotary kilns is hard and has a dense mineral structure. Valuable metals are intertwined, and it is difficult to fully dissociate the target minerals by relying solely on physical beneficiation, which makes it impossible to further improve the recovery efficiency of various valuable metals.

[0007] 5. Failure to simultaneously treat arsenic impurities: Zinc smelting kiln slag is generally accompanied by toxic and harmful elements such as arsenic. This patent does not have a corresponding arsenic removal process, and arsenic components remain in the products or tailings for a long time, posing a risk of secondary heavy metal pollution. The process is not environmentally friendly.

[0008] Therefore, developing a method for the resource utilization of zinc smelting kiln slag that has stable operation, controllable production costs, and can simultaneously and efficiently recover gold, silver, copper, and iron, while also taking into account the harmless disposal of arsenic impurities and the high value of products, is a technical problem that the current hydrometallurgical zinc smelting industry urgently needs to solve. Summary of the Invention

[0009] In view of this, the purpose of this invention is to provide a multi-metal resource recovery process for zinc smelting rotary kiln slag. The comprehensive recovery rate of gold, silver and copper of this invention is ≥72%, and the iron leaching rate is 96.1%. This solves the defects of existing pure mineral processing processes, such as the inability to recover gold and copper, low iron recovery rate, and arsenic pollution. The whole process is mild and has low reagent consumption, which can realize the high-value and harmless co-treatment of solid waste multi-metals.

[0010] The present invention discloses a process for the resource recovery of polymetallic materials from the slag of a zinc smelting rotary kiln, which includes the following steps: (1) Two-stage magnetic separation: The rotary kiln slag is dry magnetically separated and classified with a magnetic separation intensity of 3300-3800 gs to obtain iron slag and carbon slag; then the iron slag is wet-grinded and then wet magnetically separated and classified with a magnetic separation intensity of 2300-2800 gs to obtain iron concentrate and iron tailings. (2) Low acid leaching: Add sulfuric acid with a mass concentration of 140-160 g / L to the iron concentrate in step (1), wherein the mass ratio of the iron concentrate to the sulfuric acid is 1:1.5-1.7, and react at 30-40℃ for 1.5-2.5 h. After filtration, low acid leaching residue and low acid leaching liquid are obtained. (3) High acid leaching: Add sulfuric acid with a mass concentration of 160-180 g / L to the low leaching residue in step (2), wherein the mass ratio of the low leaching residue to the sulfuric acid is 1:1.6-1.8, and react at 30-40℃ for 2-3 hours to obtain high leaching residue and high leaching liquid; The leaching reaction mechanisms of the first two stages are the same, as detailed below: 1. Amorphous / surface-active FeO (kiln slag is a product of a strong reducing atmosphere, and iron mainly exists in this form). FeO + H₂SO₄ → FeSO₄ + H₂O This portion of iron enters the solution as Fe 2+ exist.

[0011] 2. Surface oxide layer Fe2O3 (dissolves slowly in dilute sulfuric acid) Fe₂O₃ + 3H₂SO₄ → Fe₂(SO₄)₃ + 3H₂O This portion of iron enters the solution as Fe 3+ It exists and is subsequently reduced by residual FeS: FeS + H₂SO₄ → FeSO₄ + H₂S↑ FeS + Fe2(SO4)3 → 3FeSO4 + S↓ Fe 2+ After entering the low-leaching and high-leaching post-leaching solutions, the high-leaching residue contains unreacted solid metal impurities.

[0012] (4) Oxidative leaching: Add an oxidant to the high leaching residue in step (3). The initial acidity of the high leaching residue is controlled at 18~30g / L. After 2.5-3h of reaction, filter to obtain gold and silver residue and oxidative leaching liquid. The copper in the high-leaching residue mainly exists as cuprous sulfide (Cu2S) and copper sulfide (CuS).

[0013] ① Leaching of cuprous sulfide (Cu2S) This is the most important reaction, where sulfur is typically oxidized to elemental sulfur (S), which is the ideal form for industrial sulfur recovery. 5Cu2S+8O2+8H2SO4→10CuSO4+8H2O+3S↓ Note: If the oxygen pressure is extremely high or the temperature is too high, sulfur may be further oxidized to sulfate (SO4). 2 ).

[0014] ② Leaching of copper sulfide (CuS) 3CuS+3O2+2H2SO4→3CuSO4+2H2O+2S↓ After CuSO4 enters the leaching solution, the remaining solids are gold and silver slag, which are sold directly as raw materials for the preparation of gold and silver metals.

[0015] (5) Displacement copper extraction: Add iron powder to the oxidative leaching solution in step (4), wherein the mass ratio of the oxidative leaching solution to the iron powder is 1:300-325, react at 50-60℃ for 1-1.5h, and then filter to obtain copper arsenic slag and displacement solution.

[0016] The specific reactions are shown in the table below: Table 1

[0017] Among them, Fe 2+ After entering the replacement liquid, the copper-arsenic slag that has been replaced is sold directly as a raw material for the preparation of copper metal.

[0018] Furthermore, in step (1), the mass percentage of each component in the rotary kiln slag is: Ag 60-70g / t, Au 0.4-0.6g / t, Cu 0.5-0.6%, Fe 22-26%, C 18-25%, SiO 2 14-16%, and the diameter of the rotary kiln slag is 1-3cm.

[0019] Furthermore, after wet grinding in step (1), the iron slag has a particle size of -200 mesh accounting for more than 80%.

[0020] Further, the liquid obtained after soaking in step (2) is cooled and crystallized. The stirring speed is controlled at 90-130 r / min, the cooling temperature is -7℃ to -2℃, and the cooling time is 5-6 h. Green vitriol seed crystals with an average particle size of 80-120 mesh are added, and the amount of seed crystals added is >3 g / L. After filtration, ferrous sulfate heptahydrate and crystallization mother liquor are obtained. Ferrous sulfate heptahydrate is dried in an environment of 45-50℃ for 2-4 h to obtain ferrous sulfate heptahydrate product.

[0021] Furthermore, the mother liquor of crystallization is heated to 30-40°C and then sent to step (2) to be mixed with the iron concentrate for closed-loop circulation. Because the iron content of the mother liquor of crystallization is still very high, it needs to be leached by return circulation.

[0022] Furthermore, the high-leaching liquid from step (3) is sent to step (2) to be mixed with the iron concentrate for closed-loop circulation. The acidity and iron content of the high-leaching liquid are still relatively high, so it needs to be recycled.

[0023] Furthermore, the oxidant in step (4) is oxygen, hydrogen peroxide, or ferric sulfate; wherein when the oxidant is oxygen, the mass ratio of the high-leaching residue to the oxygen is 1:10-15; when the oxidant is hydrogen peroxide, the mass ratio of the high-leaching residue to the hydrogen peroxide is 1:0.25-0.5; and when the oxidant is ferric sulfate, the mass ratio of the high-leaching residue to the ferric sulfate is 1:0.4-0.5.

[0024] Furthermore, in step (5), the initial acidity of the oxidative leaching solution is controlled to be <20 g / l.

[0025] Furthermore, in step (5), the iron powder has a particle size of 80 mesh to 100 mesh and a purity of >99.9%.

[0026] Furthermore, the replacement liquid from step (5) is sent to step (3) to be mixed with the low-leaching residue. The replacement liquid contains a certain amount of acid and iron and needs to be recycled.

[0027] Advantages of this invention: 1. This invention discloses a multi-metal resource recovery process for rotary kiln slag in zinc smelting, achieving high multi-metal recovery rates and strong mineral compatibility. Compared to the comparative patent CN108531740A, which relies solely on physical beneficiation and cannot recover gold and copper, this invention employs a combined beneficiation route of two-stage magnetic separation, stepped acid leaching, oxygen oxidation, and iron powder replacement. The comprehensive recovery rate of gold, silver, and copper is ≥72%, and the total iron leaching rate can reach 96.1%, improving iron recovery efficiency by 70 percentage points. For kiln slag where minerals are intertwined after high-temperature water quenching, the combination of magnetic pre-enrichment and wet dissociation effectively solves the problem of difficult dissociation of valuable metals by single beneficiation methods.

[0028] 2. This invention discloses a multi-metal resource recovery process for zinc smelting rotary kiln slag, featuring comprehensive environmental protection facilities and an economical and efficient reagent and mother liquor recycling system. The process relies on iron powder replacement to simultaneously precipitate arsenic and enrich it in copper arsenic slag for centralized disposal, preventing secondary pollution from the discharge of arsenic-containing solids and liquids. Oxygen is used as the oxidant, achieving a copper leaching rate exceeding 94%, which is lower in cost and introduces no impurities compared to manganese powder, hydrogen peroxide, and ferric sulfate. The high-efficiency leaching, replacement, and crystallization mother liquor processes are all recycled in a closed loop, reducing sulfuric acid consumption, wastewater volume, and secondary recovery of valuable metals, significantly lowering water treatment and reagent costs.

[0029] 3. This invention discloses a multi-metal resource recovery process for zinc smelting rotary kiln slag, which co-produces multiple high-value-added products and has significant economic benefits. Unlike existing technologies that only produce ordinary iron concentrate, this process can simultaneously sell high-grade copper-arsenic slag, highly enriched gold and silver slag, and ferrous sulfate heptahydrate. The ferrous sulfate heptahydrate product meets the national standard Class II, broadening the value-added channels for solid waste and significantly improving resource recovery benefits.

[0030] 4. This invention discloses a multi-metal resource recovery process from the slag of a zinc smelting rotary kiln. The process parameters have been verified through orthogonal experiments, and the process is mild and easily industrialized. The optimal production range for cooling and crystallization is determined through four sets of gradient curve experiments involving temperature, stirring, time, and seed crystals, ensuring controllable crystallization yield and quality. The entire process operates under mild conditions from room temperature to 70°C, without the need for high temperature and high pressure, and ensures smooth material filtration. Existing zinc smelting production lines can achieve continuous large-scale production with minor modifications. Attached Figure Description

[0031] To more clearly illustrate the technical solutions in the embodiments of the present invention or the prior art, the drawings used in the description of the embodiments or the prior art will be briefly introduced below. Obviously, the drawings described below are only some embodiments of the present invention. For those skilled in the art, other drawings can be obtained based on these drawings without creative effort.

[0032] Figure 1 is a schematic diagram of the complete process flow for the multi-metal resource recovery of zinc smelting rotary kiln slag according to the present invention. Figure 2 shows the relationship between cooling temperature and crystallization yield. Figure 3 shows the relationship between stirring speed and crystallization yield; Figure 4 shows the relationship between cooling time and crystallization yield. Figure 5 shows the relationship between the amount of seed crystals added and the crystallization yield. Detailed Implementation

[0033] The present invention will be further described in detail below through embodiments.

[0034] Basic parameters of raw materials Zinc smelting rotary kiln slag, particle size 1-3cm, composition: Au 0.4-0.6g / t, Ag 60-70g / t, Cu 0.5-0.6%, Fe 22-26%, C 18-25%, SiO 214-16%; Each test group was fed 100 kg of material at a time, with the ambient temperature controlled between room temperature and 70°C throughout the process; the overall process flow is shown below. Figure 1 The entire process is designed with a three-stage closed-loop mother liquor circulation system: the high-acid leaching solution is returned to the low-acid leaching solution, the displacement solution is returned to the high-acid leaching solution, and the crystallization mother liquor is returned to the low-acid leaching solution. The cooling crystallization raw material is uniformly taken from the low-acid leaching solution of each embodiment, and the orthogonal experiment uses the dry weight of crystals as the evaluation index.

[0035] Example 1 (1) Two-stage magnetic separation: dry magnetic separation intensity of kiln slag is 3300-3800 gs, separating carbon slag and iron slag; iron slag is wet-ground to -200 mesh with a proportion of >80%, and wet magnetic separation is performed after adding water to adjust the slurry, with a magnetic separation intensity of 2300-2800 gs, to obtain iron concentrate and iron tailings.

[0036] (2) Low acid leaching: sulfuric acid mass concentration 150g / L, iron concentrate to sulfuric acid mass ratio 1:1.6, reaction at 35℃ for 2h; filter to obtain low leaching residue and low leaching liquid; the crystallization mother liquor is heated to 30-40℃ and then returned to this process for recycling.

[0037] (3) High acid leaching: sulfuric acid mass concentration 170g / L, low leaching residue: sulfuric acid = 1:1.7, reaction at 35℃ for 2.5h; filter to obtain high leaching residue and high leaching liquid; high leaching liquid is returned to the low acid leaching process for recycling, and the total iron leaching rate of the two acid stages is 96.1%.

[0038] (4) Oxygen oxidation leaching: High leaching residue is used to adjust the slurry, the initial acidity of the system is 25 g / L, oxygen is introduced as an oxidant, the residue-oxygen mass ratio is 1:12, and the reaction is carried out at a constant temperature for 3 hours; after filtration, the oxidized leaching liquid and gold and silver residue are obtained; the copper leaching rate is 94.2%, and the gold and silver residue is enriched with Au to 5 g / t and Ag to 1060 g / t, which can be directly sold to precious metal smelters.

[0039] (5) Copper extraction by iron powder replacement: The acidity of the leaching solution after oxidation is controlled at 18 g / L; 80-100 mesh coarse iron powder with a purity of 99.9% is selected, and the mass ratio of the leaching solution to the iron powder is 1:310. The reaction is carried out at 55℃ for 1.2 h; the copper replacement rate is 99.8% and the arsenic removal rate is 92.5%; the replacement solution is returned to the high acid leaching process for recycling, and copper arsenic slag with a copper content of 53.2% is produced.

[0040] (6) Cooling and crystallization (after leaching): Take the after-leaching liquid and carry out the crystallization operation. Control the stirring speed at 100 r / min, the cooling temperature at -5℃, and the crystallization at this temperature for 5.5 h. Add 3 g / L of 80-120 mesh ferrous sulfate seed crystals. Separate the solid and liquid, and dry it in an environment of 45-50℃ for 2-4 h to obtain the ferrous sulfate heptahydrate product. The test results of the ferrous sulfate heptahydrate product show that the indicators such as TiO2, heavy metals, and insoluble matter all meet the national standard Class II ferrous sulfate heptahydrate standard.

[0041] Comprehensive recovery indicators: gold recovery rate 72.1%, silver 78.6%, copper 75.3%, and total iron leaching rate 96.1%.

[0042] Example 2 Step 4 replaces the oxidant with hydrogen peroxide, with a slag:hydrogen peroxide ratio of 1:0.35. All other process parameters remain completely consistent.

[0043] Test results: Copper leaching rate was 86.7%, and the comprehensive recovery rate of gold, silver and copper was 69.3%; the procurement cost of hydrogen peroxide increased significantly, resulting in poor overall economic benefits.

[0044] Example 3 In step 4, the oxidant is ferric sulfate, and the ratio of slag to ferric sulfate is 1:0.45. The remaining steps remain unchanged.

[0045] Test results: 93.5% copper was leached out, but a large amount of iron impurities were introduced into the system, resulting in a decrease in the purity of the crystallized product, which could only meet the national standard Class II product.

[0046] Comparative Example 1 (i.e., the pure physical mineral processing technology compared to patent CN108531740) It only uses kiln slag grinding, carbon flotation, silver flotation, and a single magnetic separation process, without the complete set of wet process sections of acid leaching, oxidation, displacement, and crystallization. Test results: Copper recovery rate was 0%, total iron recovery rate was 21.7%, and the comprehensive gold and silver recovery rate was less than 50%; no ferrous sulfate heptahydrate by-product was produced, and the arsenic content in the tailings exceeded the standard, posing an environmental risk.

[0047] Comparative Example 2 Step 4: The oxidant is replaced with manganese powder at a dosage of 5%, and the initial acidity is 65 g / L. The remaining steps are exactly the same as in Implementation 1.

[0048] Experimental results: Copper leaching rate was only 57.8%, the theoretical amount of iron powder added increased to 4.5 times, the reagent cost increased by 33%, and the comprehensive recovery rate of gold, silver and copper dropped to 61%.

[0049] Comparative Example 3 The high-temperature leaching solution, the displacement solution, and the crystallization mother liquor are all not returned to the front-end leaching and are directly discharged.

[0050] Experimental results: Sulfuric acid consumption increased by 23%, iron, copper and silver recovery rates decreased by 10.2%, 8.6% and 7.9% respectively, the discharged wastewater contained iron and arsenic, and the wastewater treatment cost increased significantly.

[0051] test The experimental raw material was uniformly the low-impregnation liquid produced in Example 1. Only one parameter was adjusted for each group, while the other three parameters remained at their optimal values ​​(stirring 100 r / min, crystallization 6.5 h, seed crystals 3 g / L). The dry weight of the crystals was used as the evaluation index. The yield trends of each factor are shown below. Figure 2 (temperature), Figure 3 (Stirring) Figure 4 (time), Figure 5 As shown in the (seed) diagram.

[0052] 1. Cooling temperature gradient test Option 1: -7℃, crystallization yield 55.1g; Option 2: -2℃, crystallization yield 50.9g; Comparison with Scheme 1: 9℃, crystallization yield 0.6g; Pattern: Combination Figure 2As can be seen from the temperature change curve, the lower the temperature, the higher the crystallization yield. The yield difference between -7℃ and -2℃ is minimal. The yield drops sharply when the temperature rises to 4℃, and almost no crystals precipitate at 9℃. The solution freezes completely when the temperature is below -10℃, making industrial production impossible.

[0053] 2. Stirring speed gradient test Option 1: 100 r / min, crystallization yield 55.1 g (optimal); Option 2: 150 r / min, crystallization yield 52.4 g; Option 3: 200 r / min, crystallization yield 53.9 g; Comparison scheme 2: 0 r / min (no stirring), crystallization yield is only 1.5g.

[0054] Pattern: From Figure 3 The stirring speed curve shows that without stirring, the seed crystals agglomerate and very few crystals precipitate. When stirring exceeds 100 r / min, the continuous impact of the fluid disrupts the crystal growth interface, and the crystal yield gradually decreases. 100 r / min is the optimal stirring range.

[0055] 3. Cooling time gradient test Option 1: 53.2g of crystals were obtained after 6 hours; Option 2: 7 hours, 53.9g of crystals were obtained; Comparison scheme 3: 2.5h, crystallization 1.5g; Comparison scheme 4: 4h, crystallization 18.2g.

[0056] Pattern: such as Figure 4 As shown in the time curve, the crystallization yield increases slowly over time, but the increase is minimal after the crystallization time exceeds 5.5 hours. Considering the overall energy consumption of production, 5 to 6 hours is the optimal time.

[0057] 4. Seed crystal addition gradient test Option 1: Seed crystals 3g / L, 55.1g; Option 2: Seed crystals 2g / L, 42.2g; Option 3: Seed crystals 1g / L, 18.2g; Comparison Option 5: No seed crystals added, only 17.4g.

[0058] Pattern: Combination Figure 5 The seed crystal addition curve shows that the crystallization yield is linearly positively correlated with the seed crystal addition amount. The seed crystal addition amount for industrial production needs to be ≥3g / L.

[0059] The above are preferred embodiments of the present invention. For those skilled in the art, several improvements and modifications can be made without departing from the principle of the present invention, and these improvements and modifications should also be considered within the scope of protection of the present invention.

Claims

1. A process for the resource recovery of multiple metals from the slag of a zinc smelting rotary kiln, characterized in that, It includes the following steps: (1) Two-stage magnetic separation: The rotary kiln slag is dry magnetically separated and classified with a magnetic separation intensity of 3300-3800 gs to obtain iron slag and carbon slag; then the iron slag is wet-grinded and then wet magnetically separated and classified with a magnetic separation intensity of 2300-2800 gs to obtain iron concentrate and iron tailings. (2) Low acid leaching: Add sulfuric acid with a mass concentration of 140-160 g / L to the iron concentrate in step (1), wherein the mass ratio of the iron concentrate to sulfuric acid is 1:1.5-1.7, and react at 30-40℃ for 1.5-2.5 h. After filtration, low acid leaching residue and low acid leaching liquid are obtained. (3) High acid leaching: Add sulfuric acid with a mass concentration of 160-180 g / L to the low leaching residue in step (2), wherein the mass ratio of the low leaching residue to sulfuric acid is 1:1.6-1.8, and react at 30-40℃ for 2-3 hours to obtain high leaching residue and high leaching liquid. (4) Oxidative leaching: Add an oxidant to the high leaching residue in step (3). The initial acidity of the high leaching residue is controlled at 18~30g / L. After 2.5-3h of reaction, filter to obtain gold and silver residue and oxidative leaching liquid. (5) Displacement copper extraction: Add iron powder to the oxidative leaching solution in step (4), wherein the mass ratio of the oxidative leaching solution to the iron powder is 1:300-325, react at 50-60℃ for 1-1.5h, and then filter to obtain copper arsenic slag and displacement solution.

2. The process for recovering polymetallic resources from the slag of a zinc smelting rotary kiln according to claim 1, characterized in that, In step (1), the mass percentage of each component in the rotary kiln slag is: Ag 60-70g / t, Au 0.4-0.6g / t, Cu 0.5-0.6%, Fe 22-26%, C 18-25%, SiO 2 14-16%, and the diameter of the rotary kiln slag is 1-3cm.

3. The process for the resource recovery of polymetallic materials from the slag of a zinc smelting rotary kiln according to claim 1, characterized in that, After wet grinding in step (1), the iron slag has a particle size of -200 mesh and accounts for more than 80%.

4. The process for the resource recovery of polymetallic materials from the slag of a zinc smelting rotary kiln according to claim 1, characterized in that, The liquid obtained after soaking in step (2) is cooled and crystallized. The stirring speed is controlled at 90-130 r / min, the cooling temperature is -7℃ to -2℃, and the cooling time is 5-6 h. Green vitriol seed crystals with an average particle size of 80-120 mesh are added, and the amount of seed crystals added is >3 g / L. After filtration, ferrous sulfate heptahydrate and crystallization mother liquor are obtained. Ferrous sulfate heptahydrate is dried in an environment of 45-50℃ for 2-4 h to obtain ferrous sulfate heptahydrate product.

5. The process for the resource recovery of polymetallic materials from the slag of a zinc smelting rotary kiln according to claim 4, characterized in that, After heating the crystallization mother liquor to 30-40℃, it is sent to step (2) to be mixed with the iron concentrate for closed-loop circulation.

6. The process for the resource recovery of polymetallic materials from the slag of a zinc smelting rotary kiln according to claim 1, characterized in that, The high-leaching liquid from step (3) is sent to step (2) to be mixed with the iron concentrate for closed-loop circulation.

7. The process for the resource recovery of polymetallic materials from the slag of a zinc smelting rotary kiln according to claim 1, characterized in that, The oxidant in step (4) is oxygen, hydrogen peroxide, or ferric sulfate; wherein when the oxidant is oxygen, the mass ratio of the high-efficiency leaching residue to the oxygen is 1:10-15; when the oxidant is hydrogen peroxide, the mass ratio of the high-efficiency leaching residue to the hydrogen peroxide is 1:0.25-0.5; and when the oxidant is ferric sulfate, the mass ratio of the high-efficiency leaching residue to the ferric sulfate is 1:0.4-0.

5.

8. The process for the resource recovery of polymetallic materials from the slag of a zinc smelting rotary kiln according to claim 1, characterized in that, In step (5), the initial acidity of the oxidative leaching solution is controlled to be <20 g / l.

9. The process for the resource recovery of polymetallic materials from the slag of a zinc smelting rotary kiln according to claim 1, characterized in that, In step (5), the iron powder has a particle size of 80-100 mesh and a purity of >99.9%.

10. The process for the resource recovery of polymetallic materials from the slag of a zinc smelting rotary kiln according to claim 1, characterized in that, The replacement liquid from step (5) is sent to step (3) and mixed with the low-impregnation residue.

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