Comprehensive recovery method for ferroalumen slag

Abstract

The invention relates to a comprehensive recovery method for ferroalumen slag, which comprises the following steps: (1) uniformly mixing ferroalumen slag with a reduction ammonia leaching solution, carrying out reduction ammonia leaching, and carrying out solid-liquid separation to obtain a nickel-cobalt-copper leaching solution and ferroalumen slag; wherein the iron-aluminum-vanadium slag comprises iron, aluminum, nickel, cobalt and copper elements; the reduction ammonia leaching solution comprises an ammonia source and a reducing agent; (2) extracting and separating the nickel-cobalt-copper leaching solution to obtain a nickel-cobalt-copper product; (3) the iron-aluminum slag and alkali liquor are mixed to be uniform, alkaline leaching is conducted, a sodium aluminate solution and iron slag are obtained through solid-liquid separation, the sodium aluminate solution is used for preparing an aluminum-containing product, and the iron slag serves as an iron-making raw material; wherein the step (2) and the step (3) do not have a sequence; the method realizes the recovery of nickel, cobalt and copper in the ferroalumen slag and the high-value utilization of aluminum resources, and can be used for solving the problems of large stockpiling amount, heavy pollution and difficult disposal of the ferroalumen slag at present.

Description

Technical Field

[0001] The present invention relates to the technical field of solid waste resource recovery, and in particular to a comprehensive recovery method for iron aluminum alum slag. Background Art

[0002] In recent years, with the rise of the new energy industry, lithium-ion batteries have experienced rapid development and are widely used in fields such as small electronic devices and electric vehicles. However, due to the limited lifespan of lithium-ion batteries, the number of scrapped batteries has been increasing year by year. Scraped lithium-ion batteries typically contain important resource elements such as nickel, cobalt, manganese, lithium, iron, aluminum, and copper. Therefore, efficient recycling of these batteries is not only a key path to sustainable resource utilization, but also an important means to reduce environmental pollution.

[0003] At present, the recycling processes of lithium-ion batteries mainly include pyrometallurgy and wet processing. Compared with the pyrometallurgical process, the wet process has significant advantages such as high metal recovery rate, milder operating conditions and lower environmental load. In the wet process, sulfuric acid leaching, neutralization precipitation to remove impurities such as iron and aluminum, extraction and separation are usually used to achieve the recovery of valuable metals. Among them, the neutralization precipitation process is the key step in impurity separation. Iron and aluminum impurities are usually removed in the form of goethite, hematite, iron and aluminum hydroxide precipitation or iron and aluminum alum slag. In particular, the iron and aluminum alum slag process has been widely used in the fields of zinc, nickel, cobalt metallurgy and secondary resource wet recovery due to its significant treatment effect on solutions with high impurity iron and aluminum content and its low precipitation pH value.

[0004] However, the widespread use of wet recovery technology has resulted in the production of large quantities of ferroalloy slag containing nickel, cobalt, and zinc. This solid waste is typically disposed of by stacking or landfill, leading to increasingly significant resource waste and environmental pollution. Therefore, the efficient recovery and utilization of the metal resources in ferroalloy slag is not only a key approach to improving resource utilization, but also a crucial direction for protecting the environment and promoting sustainable development.

[0005] Prior art CN113789447A discloses a method for separating iron and aluminum from ferroaluminum slag and recovering nickel. The method involves first dissolving the ferroaluminum slag with sulfuric acid, then adjusting the pH to 1.0-3.2, 3.2-5.5, and 7.0-8.8, respectively, using ammonia water to generate ferroaluminum hydroxide, aluminum hydroxide precipitate, and nickel complex. However, the method involves lengthy steps, significant nickel loss during the precipitation and separation process, and the aluminum resource cannot be reused.

[0006] Prior art CN115896462A discloses a method for recovering nickel-cobalt-containing iron-aluminum slag. The method sequentially washes the iron-aluminum slag with a copper removal solution and a sulfuric acid buffer solution to produce a nickel-cobalt washing solution and an iron-aluminum-washed slag. The nickel-cobalt washing solution is then precipitated to remove iron and aluminum before entering an extraction system. This method dissolves a large amount of iron and aluminum during the washing process, requiring chemical precipitation for removal. The overall nickel and cobalt recovery efficiency is low.

[0007] Prior art CN119237446A discloses a method for stabilizing nickel-cobalt-fluoride-iron-aluminum slag. The method involves mixing the ferro-aluminum slag with an alkaline substance to form a slurry, adding a calcium-containing substance and a phosphate ion solution, adjusting the pH to greater than 8, and performing solid-liquid separation to obtain a filtrate and stabilized ferro-aluminum slag. This method fails to recycle the valuable metals nickel, cobalt, and aluminum, resulting in a waste of resources.

[0008] In summary, based on the current situation that the ferroaluminum alum slag produced by the zinc-nickel-cobalt metallurgical industry has a large stockpile, heavy pollution and difficulty in disposal, the development of a new comprehensive recovery method for ferroaluminum alum slag has become an urgent problem to be solved. Summary of the Invention

[0009] In order to solve the above technical problems, the present invention provides a comprehensive recovery method for ferrous alum slag, which comprises the following steps: uniformly mixing the ferrous alum slag with a reducing ammonia leaching solution, performing reducing ammonia leaching, obtaining a nickel-cobalt-copper leachate and ferrous alum slag through solid-liquid separation, obtaining a nickel-cobalt-copper product through extraction and separation, obtaining a sodium aluminate solution and ferrous slag through alkaline leaching, and producing an aluminum-containing product through the sodium aluminate solution. The ferrous slag is used as an ironmaking raw material, thereby realizing comprehensive recovery of the ferrous alum slag, and solving the problems of heavy pollution, difficult disposal and low recovery efficiency of the ferrous alum slag.

[0010] To achieve this object, the present invention adopts the following technical solutions:

[0011] In a first aspect, the present invention provides a method for comprehensive recovery of aluminum sulfate slag, the method comprising the following steps:

[0012] (1) uniformly mixing the ferroaluminum slag with a reduction ammonia leaching solution, performing reduction ammonia leaching, and obtaining nickel, cobalt, and copper leaching solution and ferroaluminum slag through solid-liquid separation;

[0013] Wherein, the aluminum alum slag includes iron, aluminum, nickel, cobalt and copper elements; the reduced ammonia leaching solution includes an ammonia source and a reducing agent;

[0014] (2) extracting and separating the nickel-cobalt-copper leachate of step (1) to obtain a nickel-cobalt-copper product;

[0015] (3) uniformly mixing the iron-aluminum slag described in step (1) with an alkaline solution, performing alkaline leaching, and performing solid-liquid separation to obtain a sodium aluminate solution and iron slag, wherein the sodium aluminate solution is used to prepare an aluminum-containing product, and the iron slag is used as an ironmaking raw material;

[0016] There is no order between step (2) and step (3).

[0017] The comprehensive recovery method of ferrous alum slag provided by the present invention only involves uniformly mixing the ferrous alum slag with a reducing ammonia leaching solution, performing reducing ammonia leaching, obtaining a nickel-cobalt-copper leachate and ferrous alum slag through solid-liquid separation, obtaining a nickel-cobalt-copper product through extraction and separation, obtaining a sodium aluminate solution and ferrous slag through alkali leaching, and producing an aluminum-containing product through the sodium aluminate solution. The ferrous slag is used as an ironmaking raw material, thereby realizing comprehensive recovery of the ferrous alum slag. The reducing agent reduces some elements in the ferrous alum slag, such as trivalent cobalt ions, and the ammonia source provides ammonium ions for selective complexation with nickel-cobalt-copper, thereby realizing separation of nickel-cobalt-copper from the ferrous alum slag, and realizing separation and recovery of iron and aluminum through alkali leaching.

[0018] Preferably, the source of the iron-aluminum alum slag in step (1) includes the neutralization and iron-aluminum removal process of waste ternary battery leachate.

[0019] Preferably, the ferroaluminum alum slag in step (1) includes elements such as iron, aluminum, nickel, cobalt, copper, and manganese.

[0020] In the present invention, there is no specific limitation on the aluminum sulfate slag described in step (1), and any aluminum sulfate slag containing valuable metals such as nickel, cobalt, manganese, and copper can be used in the present invention.

[0021] Preferably, the ammonia source in the reduced ammonia immersion solution in step (1) comprises any one or a combination of at least two of ammonia water, ammonium carbonate, ammonium bicarbonate or ammonium sulfate, wherein typical but non-limiting combinations are a combination of ammonia water and ammonium carbonate, a combination of ammonia water and ammonium bicarbonate, a combination of ammonia water and ammonium sulfate, a combination of ammonium carbonate and ammonium bicarbonate, a combination of ammonium carbonate and ammonium sulfate, a combination of ammonium bicarbonate and ammonium sulfate, a combination of ammonia water, ammonium carbonate and ammonium bicarbonate, a combination of ammonium carbonate, ammonium bicarbonate and ammonium sulfate, and preferably a combination of ammonium carbonate and ammonia water.

[0022] Preferably, the molar concentration of the ammonia source in the reduced ammonia immersion solution in step (1) is 1-15 mol / L, for example, it can be 1 mol / L, 2 mol / L, 3 mol / L, 5 mol / L, 7 mol / L, 8 mol / L, 9 mol / L, 11 mol / L, 13 mol / L or 15 mol / L, but is not limited to the listed values. Other values ​​not listed within this range are also applicable, preferably 2-8 mol / L.

[0023] Preferably, the ammonia source in the reduced ammonia immersion solution is a combination of ammonium carbonate and aqueous ammonia, and the molar concentration of aqueous ammonia in the reduced ammonia immersion solution is 1-10 mol / L, for example, 1 mol / L, 3 mol / L, 5 mol / L, 7 mol / L, 9 mol / L or 10 mol / L, etc., but is not limited to the listed values, and other values ​​not listed within this range are also applicable.

[0024] Preferably, the ammonia source in the reduced ammonia immersion solution is a combination of ammonium carbonate and aqueous ammonia, and the molar concentration of ammonium carbonate in the reduced ammonia immersion solution is 1-5 mol / L, for example, it can be 1 mol / L, 2 mol / L, 3 mol / L, 4 mol / L or 5 mol / L, but is not limited to the listed values, and other values ​​not listed within this range are also applicable.

[0025] In the present invention, the ammonia source provides ammonium ions, which selectively complex with nickel, cobalt and copper to achieve the leaching of nickel, cobalt and copper. When the concentration of the ammonia source increases, the concentration of ammonium ions in the solution increases, and the leaching rate of nickel, cobalt and copper increases.

[0026] Preferably, the reducing agent in the reduced ammonia leaching solution in step (1) comprises any one or a combination of at least two of ammonium sulfite, sodium sulfite or hydrogen peroxide, wherein typical but non-limiting combinations are a combination of ammonium sulfite and sodium sulfite, a combination of ammonium sulfite and hydrogen peroxide, and a combination of sodium sulfite and hydrogen peroxide, preferably ammonium sulfite.

[0027] Preferably, the molar concentration of the reducing agent in the reduced ammonia immersion solution in step (1) is 0.2-4 mol / L, for example, it can be 0.2 mol / L, 0.5 mol / L, 1.5 mol / L, 2 mol / L, 2.5 mol / L, 3 mol / L, 3.5 mol / L or 4 mol / L, but is not limited to the listed values. Other values ​​not listed within this range are also applicable, preferably 0.5-2 mol / L.

[0028] In the present invention, the concentration of the reducing agent has a significant effect on the leaching rate of cobalt. When the concentration of the reducing agent increases, the leaching rate of cobalt increases significantly. The reason is that part of the cobalt element is in the form of Co in the ferroaluminum slag. 3+ exists in the form of Co, which can be reduced to 2 + , thereby increasing the leaching rate.

[0029] Preferably, the liquid-to-solid ratio of the reduction ammonia leaching in step (1) is (2-20) mL: 1g, for example, it can be 2 mL: 1g, 5 mL: 1g, 7 mL: 1g, 10 mL: 1g, 12 mL: 1g, 15 mL: 1g, 17 mL: 1g or 20 mL: 1g, etc., but is not limited to the listed values. Other values ​​not listed within this range are also applicable, preferably (5-10) mL: 1g.

[0030] In the present invention, when the liquid-solid ratio of the reducing ammonia leaching is low, the concentration of nickel, cobalt and copper in the solution is high, which is not conducive to the equilibrium moving towards the leaching direction, resulting in a low nickel, cobalt and copper leaching rate; on the contrary, when the liquid-solid ratio of the reducing ammonia leaching is high, the concentration of nickel, cobalt and copper in the solution is low, which is more conducive to the equilibrium moving towards the leaching direction, thereby increasing the nickel, cobalt and copper leaching rate.

[0031] Preferably, the temperature of the reductive ammonia leaching in step (1) is 25-95°C, for example, it can be 25°C, 30°C, 35°C, 40°C, 45°C, 50°C, 55°C, 60°C, 65°C, 70°C, 75°C, 80°C, 85°C or 95°C, but is not limited to the listed values. Other values ​​not listed in this range are also applicable, preferably 50-80°C.

[0032] In the present invention, when the temperature of the reducing ammonia leaching is low, the decomposition efficiency of the iron aluminum alum slag structure in the reducing ammonia leaching solution is low, so that nickel, cobalt and copper cannot be effectively released from the iron aluminum alum slag structure, thereby resulting in a reduced nickel, cobalt and copper leaching rate. When the temperature of the reducing ammonia leaching is too high, the ammonia solution volatilizes and loses seriously, so that the ammonium ions in the solution are insufficient to complex the nickel, cobalt and copper, thereby resulting in a reduced nickel, cobalt and copper leaching rate.

[0033] Preferably, the time for the reductive ammonia leaching in step (1) is 30-600 min, for example, it can be 30 min, 60 min, 90 min, 120 min, 180 min, 210 min, 240 min, 270 min, 300 min, 330 min, 360 min, 390 min, 420 min, 450 min, 480 min, 510 min, 540 min, 570 min or 600 min, but is not limited to the listed values. Other values ​​not listed within this range are also applicable, preferably 240-480 min.

[0034] In the present invention, extending the time of reducing ammonia leaching can increase the nickel, cobalt and copper leaching rate.

[0035] Preferably, the reductive ammonia leaching in step (1) is carried out under stirring, and the stirring rate is 100-800r / min, for example, it can be 100r / min, 150r / min, 200r / min, 250r / min, 300r / min, 350r / min, 400r / min, 450r / min, 500r / min, 550r / min, 600r / min, 650r / min, 700r / min, 750r / min or 800r / min, etc., but is not limited to the listed values. Other values ​​not listed in this range are also applicable, preferably 300-600r / min.

[0036] Preferably, the alkali source in the alkali solution in step (3) comprises sodium hydroxide and / or potassium hydroxide, preferably sodium hydroxide.

[0037] Preferably, the mass concentration of the alkali source in the alkali solution of step (3) is 50-300 g / L, for example, it can be 50 g / L, 60 g / L, 70 g / L, 80 g / L, 90 g / L, 100 g / L, 110 g / L, 120 g / L, 130 g / L, 140 g / L, 150 g / L, 160 g / L, 170 g / L, 180 g / L, 190 g / L, 200 g / L, 210 g / L, 220 g / L, 230 g / L, 240 g / L, 250 g / L, 260 g / L, 270 g / L, 280 g / L, 290 g / L or 300 g / L, but is not limited to the enumerated values. Other unenumerated values ​​within this range are equally applicable, preferably 60-150 g / L.

[0038] In the present invention, when the alkali source concentration used in alkali leaching is low, the aluminum leaching rate is very low due to insufficient alkali. When the alkali source concentration is increased, the aluminum leaching efficiency can be significantly increased. However, if the alkali source concentration is too high, the equipment will be seriously corroded. Therefore, the alkali concentration should be controlled within an appropriate range.

[0039] Preferably, the liquid-to-solid ratio of the alkali leaching in step (3) is (1-15) mL:1g, for example, it can be 1 mL:1g, 1.5 mL:1g, 2 mL:1g, 2.5 mL:1g, 3 mL:1g, 3.5 mL:1g, 4 mL:1g, 5 mL:1g, 6 mL:1g, 7 mL:1g, 8 mL:1g, 9 mL:1g, 10 mL:1g, 11 mL:1g, 12.5 mL:1g, 13 mL:1g, 14 mL:1g or 15 mL:1g, etc., but is not limited to the listed values. Other values ​​not listed within this range are also applicable, preferably (2-4):1.

[0040] Preferably, the alkali leaching temperature of the alkali leaching in step (3) is 50-95°C, for example, it can be 50°C, 55°C, 60°C, 65°C, 70°C, 75°C, 80°C, 85°C, 90°C or 95°C, etc., but is not limited to the listed values. Other values ​​not listed in this range are also applicable, preferably 70-95°C.

[0041] In the present invention, the temperature of alkali leaching affects the leaching rate of aluminum. As the temperature increases, the reaction rate of alkali leaching increases, thereby increasing the leaching rate of aluminum.

[0042] Preferably, the alkali leaching time of the alkali leaching in step (3) is 30-300 min, for example, it can be 30 min, 45 min, 60 min, 75 min, 90 min, 120 min, 150 min, 180 min, 210 min, 240 min, 270 min or 300 min, but is not limited to the listed values. Other values ​​not listed within this range are also applicable, preferably 30-90 min.

[0043] Preferably, the sodium aluminate solution in step (3) contains impurities such as fluorine and phosphorus, which are removed by adding an alkaline earth metal compound to obtain a purified sodium aluminate solution, which is then used to prepare the aluminum-containing product.

[0044] Preferably, the alkaline earth metal compound comprises any one or a combination of at least two of calcium oxide, magnesium oxide, calcium carbonate, magnesium carbonate, calcium hydroxide, and magnesium hydroxide, wherein typical but non-limiting combinations are a combination of calcium oxide and magnesium oxide, a combination of calcium oxide and calcium carbonate, a combination of calcium oxide and magnesium carbonate, a combination of calcium oxide and calcium hydroxide, a combination of calcium oxide and magnesium hydroxide, a combination of magnesium oxide and calcium carbonate, a combination of magnesium oxide and magnesium carbonate, a combination of magnesium oxide and calcium hydroxide, a combination of magnesium oxide and magnesium hydroxide, a combination of calcium carbonate and magnesium carbonate, a combination of calcium carbonate and calcium hydroxide, a combination of calcium carbonate and magnesium hydroxide, a combination of magnesium carbonate and calcium hydroxide, or a combination of magnesium carbonate and magnesium hydroxide, etc., preferably a combination of calcium oxide and calcium hydroxide.

[0045] Preferably, the liquid-to-solid ratio of the sodium aluminate solution to the alkaline earth metal compound is 1L:(10-50)g, for example, 1L:10g, 1L:15g, 1L:20g, 1L:25g, 1L:30g, 1L:35g, 1L:40g, 1L:45g or 1L:50g, etc., but is not limited to the listed values. Other values ​​not listed within this range are also applicable, preferably 1L:(10-30)g.

[0046] Preferably, the impurity removal temperature of the impurity removal process is 50-95°C, for example, it can be 50°C, 55°C, 60°C, 65°C, 70°C, 75°C, 80°C, 85°C, 90°C or 95°C, but is not limited to the listed values. Other unlisted values ​​within this range are also applicable, preferably 70-95°C.

[0047] In the present invention, the liquid-to-solid ratio of the sodium aluminate solution to the alkaline earth metal compound and the impurity removal temperature affect the removal rate of the impurities fluorine and phosphorus. A decrease in the liquid-to-solid ratio and an increase in the impurity removal temperature are conducive to the precipitation and impurity removal reaction proceeding in a favorable direction, thereby making the removal of the impurities phosphorus and fluorine in the sodium aluminate solution more thorough.

[0048] Preferably, the impurity removal time is 60-300 min, for example, it can be 60 min, 70 min, 80 min, 90 min, 100 min, 110 min, 120 min, 130 min, 150 min, 180 min, 200 min, 220 min, 240 min, 260 min, 280 min or 300 min, but is not limited to the listed values. Other values ​​not listed in this range are also applicable, preferably 60-120 min.

[0049] Preferably, using the sodium aluminate solution in step (3) to prepare the aluminum-containing product comprises: adding an acid solution to the sodium aluminate solution to carry out a neutralization reaction to obtain the aluminum-containing product.

[0050] It is worth noting that if the sodium aluminate solution is impurity-removed, an acid solution is added to the purified sodium aluminate solution to carry out a neutralization reaction.

[0051] Preferably, the aluminum-containing product comprises a pseudo-boehmite product.

[0052] Preferably, the acid solution comprises sulfuric acid.

[0053] Preferably, the mass percentage concentration of the sulfuric acid is 10-50%, for example, it can be 10%, 15%, 20%, 25%, 30%, 35%, 40%, 45% or 50%, etc., but is not limited to the listed values. Other values ​​not listed in this range are also applicable, preferably 10-30%.

[0054] Preferably, the neutralization reaction is carried out under stirring, and the stirring rate is controlled to be 100-800 r / min, for example, it can be 100 r / min, 150 r / min, 200 r / min, 250 r / min, 300 r / min, 350 r / min, 400 r / min, 450 r / min, 500 r / min, 550 r / min, 600 r / min, 650 r / min, 700 r / min, 750 r / min or 800 r / min, etc., but is not limited to the listed values. Other values ​​not listed in this range are also applicable, preferably 300-600 r / min.

[0055] Preferably, the reaction time of the neutralization reaction is 5-240 min, for example, it can be 5 min, 10 min, 20 min, 30 min, 40 min, 50 min, 70 min, 90 min, 120 min, 140 min, 160 min, 180 min, 200 min, 220 min or 240 min, but is not limited to the listed values. Other values ​​not listed in this range are also applicable, preferably 5-120 min.

[0056] Preferably, the reaction temperature of the neutralization reaction is 50-90°C, for example, it can be 50°C, 55°C, 60°C, 65°C, 70°C, 75°C, 80°C, 85°C or 90°C, but is not limited to the listed values. Other values ​​not listed within this range are also applicable, preferably 70-90°C.

[0057] As a preferred technical solution of the present invention, the method comprises the following steps:

[0058] (1) mixing the ferroalite slag and the reduction ammonia leaching solution uniformly, according to the liquid-solid ratio of (2-20) mL: 1 g, reducing ammonia leaching at 25-95 ° C for 30-600 min, the reduction ammonia leaching is carried out under stirring, the stirring rate is 100-800 r / min, and solid-liquid separation is performed to obtain nickel, cobalt and copper leachate and ferroalite slag; wherein the ferroalite slag includes iron, aluminum, nickel, cobalt and copper elements, the source of the ferroalite slag includes the process of neutralizing the waste ternary battery leachate and removing iron and aluminum, the reducing ammonia The immersion solution includes an ammonia source and a reducing agent; the molar concentration of the ammonia source in the reduced ammonia immersion solution is 1-15 mol / L, the molar concentration of the reducing agent is 0.2-4 mol / L, the ammonia source in the reduced ammonia immersion solution is a combination of ammonium carbonate and ammonia water, the molar concentration of the ammonia water in the reduced ammonia immersion solution is 1-10 mol / L, and the molar concentration of ammonium carbonate in the reduced ammonia immersion solution is 1-5 mol / L; the reducing agent includes any one of ammonium sulfite, sodium sulfite or hydrogen peroxide, or a combination of at least two thereof;

[0059] (2) extracting and separating the nickel-cobalt-copper leachate of step (1) to obtain a nickel-cobalt-copper product;

[0060] (3) The ferroaluminum slag of step (1) is mixed evenly with 50-300 g / L alkali solution at a liquid-solid ratio of (1-15) mL:1 g, and alkali-leached at 50-95 ° C for 30-300 min. After solid-liquid separation, a sodium aluminate solution and ferroaluminum slag are obtained. The sodium aluminate solution is used to prepare aluminum-containing products, and the ferroaluminum slag is used as an ironmaking raw material. The alkali source in the alkali solution includes sodium hydroxide and / or potassium hydroxide with a mass concentration of 50-300 g / L. In addition, the sodium aluminate solution contains impurities fluorine and phosphorus. Alkaline earth metal compounds are added to the sodium aluminate solution at a liquid-solid ratio of 1 L: (10-50) g. impurities at 50-95° C. for 60-300 minutes to obtain a purified sodium aluminate solution, which is then used to prepare an aluminum-containing product, wherein the alkaline earth metal compound includes any one or a combination of at least two of calcium oxide, magnesium oxide, calcium carbonate, magnesium carbonate, calcium hydroxide, or magnesium hydroxide; an acid solution is added to the sodium aluminate solution to carry out a neutralization reaction, the neutralization reaction being carried out under stirring at a temperature of 50-90° C., a reaction time of 5-240 minutes, and a stirring rate of 100-800 r / min to obtain a pseudo-boehmite product, wherein the acid solution includes sulfuric acid with a mass percentage concentration of 10-50%;

[0061] There is no order between step (2) and step (3).

[0062] Compared with the prior art, the present invention has at least the following beneficial effects:

[0063] The present invention provides a comprehensive recovery method for ferrous alum slag, which comprises the following steps: uniformly mixing the ferrous alum slag with a reducing ammonia leaching solution, performing reducing ammonia leaching, obtaining a nickel-cobalt-copper leachate and ferrous alum slag through solid-liquid separation, obtaining a nickel-cobalt-copper product through extraction and separation, obtaining a sodium aluminate solution and ferrous slag through alkali leaching, and producing an aluminum-containing product from the sodium aluminate solution. The ferrous slag is used as an ironmaking raw material, thereby realizing comprehensive recovery of the ferrous alum slag, and the overall process slag alkali quantification is greater than 60%, thereby solving the current problem of large stockpiles and difficult disposal of the ferrous alum slag, and having good economic and environmental benefits. BRIEF DESCRIPTION OF THE DRAWINGS

[0064] Figure 1 This is a process flow chart of the comprehensive recovery method of aluminum slag provided by the present invention;

[0065] Figure 2 This is a process flow chart of the comprehensive recovery method of ferroaluminum alum slag provided in Example 1 of the present invention. DETAILED DESCRIPTION

[0066] The technical solution of the present invention will be further described below with reference to the accompanying drawings and through specific embodiments. However, the following examples are merely simplified examples of the present invention and do not represent or limit the scope of protection of the present invention. The scope of protection of the present invention shall be subject to the claims.

[0067] See also Figure 1 The present invention provides a comprehensive recovery method for aluminum alum slag, which comprises the following steps:

[0068] (1) uniformly mixing the ferroaluminum slag with a reduction ammonia leaching solution, performing reduction ammonia leaching, and obtaining nickel, cobalt, and copper leaching solution and ferroaluminum slag through solid-liquid separation;

[0069] Wherein, the aluminum alum slag includes iron, aluminum, nickel, cobalt and copper elements; the reduced ammonia leaching solution includes an ammonia source and a reducing agent;

[0070] (2) extracting and separating the nickel-cobalt-copper leachate of step (1) to obtain a nickel-cobalt-copper product;

[0071] (3) uniformly mixing the iron-aluminum slag described in step (1) with an alkaline solution, performing alkaline leaching, and performing solid-liquid separation to obtain a sodium aluminate solution and iron slag, wherein the sodium aluminate solution is used to prepare an aluminum-containing product, and the iron slag is used as an ironmaking raw material;

[0072] There is no order between step (2) and step (3).

[0073] The source of the iron-aluminum slag used in the following examples includes the process of neutralizing the iron and aluminum removal process of the waste ternary battery leachate. The components of the iron-aluminum slag include: Fe 10.3%, Al 7.83%, Ni 2.48%, Co 0.67%, Mn 2.99%, Cu 0.29%, SO4 2- 45.00%.

[0074] Example 1

[0075] This embodiment provides a comprehensive recovery method for iron aluminum alum slag, the method flow chart is as follows Figure 2 As shown, the method includes the following steps:

[0076] (1) mixing the ferroalloy slag with 2 mol / L ammonium carbonate solution, 5 mol / L ammonia water and 1 mol / L ammonium sulfite, and subjecting the mixture to reduction ammonia leaching at 85°C for 300 min at a liquid-solid ratio of 5 mL:1 g. The reduction ammonia leaching is carried out under stirring, and solid-liquid separation is performed to obtain nickel, cobalt and copper leachate and ferroalloy slag;

[0077] (2) extracting and separating the nickel-cobalt-copper leachate of step (1) to obtain a nickel-cobalt-copper product;

[0078] (3) adding 150 g / L of sodium hydroxide solution to the ferroaluminum slag in step (1), alkali leaching at 90° C. for 90 min at a liquid-solid ratio of 3 mL:1 g, and obtaining a sodium aluminate solution containing impurities of fluorine and phosphorus and ferroaluminum slag through solid-liquid separation. The ferroaluminum slag is used as an ironmaking raw material. Calcium oxide is added to the sodium aluminate solution containing impurities of fluorine and phosphorus at a liquid-solid ratio of 1 L:20 g, and impurities are removed at 90° C. for 120 min to obtain a purified sodium aluminate solution. Sulfuric acid having a mass percentage concentration of 50% is added to the purified sodium aluminate to carry out a neutralization reaction. The neutralization reaction is carried out under stirring at a reaction temperature of 70° C. and a reaction time of 30 min to obtain a pseudo-boehmite product.

[0079] There is no order between step (2) and step (3).

[0080] Example 2

[0081] This embodiment provides a comprehensive recovery method for iron aluminum sulfate slag, which comprises the following steps:

[0082] (1) mixing the ferroalloy slag with 5 mol / L ammonium carbonate solution, 8 mol / L ammonia water and 2 mol / L sodium sulfite, and subjecting the mixture to reduction ammonia leaching at 55°C for 260 min at a liquid-solid ratio of 8 mL:1 g. The reduction ammonia leaching is carried out under stirring, and solid-liquid separation is performed to obtain nickel, cobalt and copper leachate and ferroalloy slag;

[0083] (2) extracting and separating the nickel-cobalt-copper leachate of step (1) to obtain a nickel-cobalt-copper product;

[0084] (3) adding 60 g / L potassium hydroxide solution to the ferroaluminum slag in step (1), alkali leaching at 50° C. for 60 min at a liquid-solid ratio of 2 mL:1 g, and obtaining a sodium aluminate solution containing impurities of fluorine and phosphorus and iron slag through solid-liquid separation. The iron slag is used as an ironmaking raw material. Magnesium carbonate is added to the sodium aluminate solution containing impurities of fluorine and phosphorus at a liquid-solid ratio of 1 L:30 g. The solution is subjected to impurity removal at 75° C. for 80 min to obtain a purified sodium aluminate solution. Sulfuric acid having a mass percentage concentration of 10% is added to the purified sodium aluminate to carry out a neutralization reaction. The neutralization reaction is carried out under stirring at a reaction temperature of 70° C. for a reaction time of 5 min to obtain a pseudo-boehmite product.

[0085] There is no order between step (2) and step (3).

[0086] Example 3

[0087] This embodiment provides a comprehensive recovery method for iron aluminum sulfate slag, which comprises the following steps:

[0088] (1) mixing the ferroalloy slag with 1 mol / L ammonium carbonate solution, 2 mol / L ammonia water and 0.5 mol / L sodium sulfite, and subjecting the mixture to reduction ammonia leaching at 65°C for 150 min at a liquid-solid ratio of 10 mL:1 g. The reduction ammonia leaching is carried out under stirring, and solid-liquid separation is performed to obtain nickel, cobalt and copper leachate and ferroalloy slag;

[0089] (2) extracting and separating the nickel-cobalt-copper leachate of step (1) to obtain a nickel-cobalt-copper product;

[0090] (3) adding 130 g / L of sodium hydroxide solution to the ferroaluminum slag in step (1), alkali leaching at 80° C. for 150 min at a liquid-solid ratio of 4 mL:1 g, and obtaining a sodium aluminate solution containing impurities of fluorine and phosphorus and iron slag through solid-liquid separation. The iron slag is used as an ironmaking raw material. Calcium oxide is added to the sodium aluminate solution containing impurities of fluorine and phosphorus at a liquid-solid ratio of 1 L:10 g. The solution is subjected to impurity removal at 80° C. for 120 min to obtain a purified sodium aluminate solution. Sulfuric acid having a mass percentage concentration of 25% is added to the purified sodium aluminate to carry out a neutralization reaction. The neutralization reaction is carried out under stirring at a reaction temperature of 85° C. for 120 min to obtain a pseudo-boehmite product.

[0091] There is no order between step (2) and step (3).

[0092] Example 4

[0093] This embodiment provides a comprehensive recovery method for iron aluminum sulfate slag, which comprises the following steps:

[0094] (1) mixing the ferroalloy slag with 1 mol / L ammonium carbonate solution, 1 mol / L ammonia water and 1 mol / L ammonium sulfite, and subjecting the mixture to reduction ammonia leaching at 65°C for 200 min at a liquid-solid ratio of 10 mL:1 g. The reduction ammonia leaching is carried out under stirring, and solid-liquid separation is performed to obtain nickel, cobalt and copper leachate and ferroalloy slag;

[0095] (2) extracting and separating the nickel-cobalt-copper leachate of step (1) to obtain a nickel-cobalt-copper product;

[0096] (3) adding 200 g / L of sodium hydroxide solution to the ferroaluminum slag in step (1), alkali leaching at 80° C. for 30 min at a liquid-solid ratio of 5 mL:1 g, and obtaining a sodium aluminate solution containing impurities of fluorine and phosphorus and ferroaluminum slag through solid-liquid separation. The ferroaluminum slag is used as an ironmaking raw material. Calcium oxide is added to the sodium aluminate solution containing impurities of fluorine and phosphorus at a liquid-solid ratio of 1 L:10 g, and impurities are removed at 50° C. for 60 min to obtain a purified sodium aluminate solution. Sulfuric acid having a mass percentage concentration of 10% is added to the purified sodium aluminate to carry out a neutralization reaction. The neutralization reaction is carried out under stirring at a reaction temperature of 50° C. and a reaction time of 5 min to obtain a pseudo-boehmite product.

[0097] There is no order between step (2) and step (3).

[0098] Example 5

[0099] This embodiment provides a comprehensive recovery method for iron aluminum sulfate slag, which comprises the following steps:

[0100] (1) mixing the ferroalloy slag with 2 mol / L ammonium carbonate solution, 3 mol / L ammonia water and 1.5 mol / L ammonium sulfite, and subjecting the mixture to reduction ammonia leaching at 25°C for 60 min at a liquid-solid ratio of 3 mL:1 g. The reduction ammonia leaching is carried out under stirring, and solid-liquid separation is performed to obtain nickel, cobalt and copper leachate and ferroalloy slag;

[0101] (2) extracting and separating the nickel-cobalt-copper leachate of step (1) to obtain a nickel-cobalt-copper product;

[0102] (3) adding 100 g / L of sodium hydroxide solution to the ferroaluminum slag in step (1), alkali leaching at 50° C. for 300 min at a liquid-solid ratio of 5 mL:1 g, and obtaining a sodium aluminate solution containing impurities of fluorine and phosphorus and iron slag through solid-liquid separation. The iron slag is used as an ironmaking raw material. Calcium oxide is added to the sodium aluminate solution containing impurities of fluorine and phosphorus at a liquid-solid ratio of 1 L:20 g. 20 g / L of calcium oxide is added to remove impurities at 60° C. for 120 min to obtain a purified sodium aluminate solution. Sulfuric acid having a mass percentage concentration of 20% is added to the purified sodium aluminate to carry out a neutralization reaction. The neutralization reaction is carried out under stirring at a reaction temperature of 40° C. for 60 min to obtain a pseudo-boehmite product.

[0103] There is no order between step (2) and step (3).

[0104] Example 6

[0105] This embodiment provides a comprehensive recovery method for aluminum sulfate slag, which comprises the following steps:

[0106] (1) mixing the ferroalloy slag with 5 mol / L ammonium carbonate solution, 10 mol / L ammonia water and 2 mol / L sodium sulfite, and subjecting the mixture to reduction ammonia leaching at 70°C for 60 min at a liquid-solid ratio of 10 mL:1 g. The reduction ammonia leaching is carried out under stirring, and solid-liquid separation is performed to obtain nickel, cobalt and copper leachate and ferroalloy slag;

[0107] (2) extracting and separating the nickel-cobalt-copper leachate of step (1) to obtain a nickel-cobalt-copper product;

[0108] (3) adding 60 g / L of sodium hydroxide solution to the ferroaluminum slag in step (1), alkali leaching at 35° C. for 60 min at a liquid-solid ratio of 8 mL:1 g, and obtaining a sodium aluminate solution containing impurities fluorine and phosphorus and ferroaluminum slag through solid-liquid separation. The ferroaluminum slag is used as an ironmaking raw material. Magnesium oxide is added to the sodium aluminate solution containing impurities fluorine and phosphorus at a liquid-solid ratio of 1 L:5 g to obtain a purified sodium aluminate solution. Sulfuric acid having a mass percentage concentration of 40% is added to the purified sodium aluminate to carry out a neutralization reaction. The neutralization reaction is carried out under stirring at a reaction temperature of 50° C. and a reaction time of 20 min to obtain a pseudo-boehmite product.

[0109] There is no order between step (2) and step (3).

[0110] Example 7

[0111] This embodiment provides a comprehensive recovery method for iron aluminum sulfate slag, which comprises the following steps:

[0112] (1) mixing the ferroalloy residue with 1 mol / L ammonium carbonate solution, 6 mol / L ammonia water and 0.2 mol / L ammonium sulfite, and subjecting the mixture to reduction ammonia leaching at 25°C for 30 min at a liquid-solid ratio of 2 mL:1 g. The reduction ammonia leaching is carried out under stirring, and solid-liquid separation is performed to obtain nickel, cobalt and copper leachate and ferroalloy residue;

[0113] (2) extracting and separating the nickel-cobalt-copper leachate of step (1) to obtain a nickel-cobalt-copper product;

[0114] (3) adding 50 g / L of sodium hydroxide solution to the ferroaluminum slag in step (1), alkali leaching at 50° C. for 30 min at a liquid-solid ratio of 1 mL:1 g, and obtaining a sodium aluminate solution containing impurities of fluorine and phosphorus and iron slag through solid-liquid separation. The iron slag is used as an ironmaking raw material. Calcium oxide is added to the sodium aluminate solution containing impurities of fluorine and phosphorus at a liquid-solid ratio of 1 L:2 g, and impurities are removed at 50° C. for 60 min to obtain a purified sodium aluminate solution. Sulfuric acid having a mass percentage concentration of 10% is added to the purified sodium aluminate to carry out a neutralization reaction. The neutralization reaction is carried out under stirring at a reaction temperature of 50° C. and a reaction time of 5 min to obtain a pseudo-boehmite product.

[0115] There is no order between step (2) and step (3).

[0116] Example 8

[0117] This embodiment provides a comprehensive recovery method for iron aluminum alum slag, which is different from Example 1 in that, except for adjusting the concentration of the ammonium carbonate solution in step (1) to 7 mol / L, the rest is the same as Example 1.

[0118] Example 9

[0119] This embodiment provides a comprehensive recovery method for iron aluminum alum slag, which is different from Example 1 in that, except for adjusting the concentration of the ammonium carbonate solution in step (1) to 0.5 mol / L, the rest is the same as Example 1.

[0120] Example 10

[0121] This embodiment provides a comprehensive recovery method for iron aluminum alum slag, which is different from Example 1 in that, except for adjusting the ammonia concentration in step (1) to 11 mol / L, the rest is the same as Example 1.

[0122] Example 11

[0123] This embodiment provides a comprehensive recovery method for iron aluminum alum slag, which is different from Example 1 in that, except for adjusting the ammonia concentration in step (1) to 0.5 mol / L, the rest is the same as Example 1.

[0124] Example 12

[0125] This embodiment provides a comprehensive recovery method for iron aluminum alum slag, which differs from Example 1 in that, except for adjusting the liquid-solid ratio of the reduction ammonia leaching in step (1) to 22mL:1g, the rest is the same as Example 1.

[0126] Example 13

[0127] This embodiment provides a comprehensive recovery method for iron aluminum alum slag, which differs from Example 1 in that, except for adjusting the liquid-solid ratio of the reduction ammonia leaching in step (1) to 1.5 mL:1 g, the rest is the same as Example 1.

[0128] Example 14

[0129] This embodiment provides a comprehensive recovery method for iron aluminum alum slag, which is different from Example 1 in that, except for adjusting the temperature of the reduction ammonia leaching in step (1) to 100°C, the rest is the same as Example 1.

[0130] Example 15

[0131] This embodiment provides a comprehensive recovery method for iron aluminum alum slag, which is different from Example 1 in that, except for adjusting the temperature of the reduction ammonia leaching in step (1) to 20°C, the rest is the same as Example 1.

[0132] Example 16

[0133] This embodiment provides a comprehensive recovery method for iron aluminum alum slag, which is different from Example 1 in that, except for adjusting the time of the reduction ammonia leaching in step (1) to 650 minutes, the rest is the same as Example 1.

[0134] Example 17

[0135] This embodiment provides a comprehensive recovery method for iron aluminum alum slag, which is different from Example 1 in that, except for adjusting the time of the reduction ammonia leaching in step (1) to 20 minutes, the rest is the same as Example 1.

[0136] Example 18

[0137] This embodiment provides a comprehensive recovery method for iron aluminum alum slag, which is different from Example 1 in that, except for adjusting the concentration of ammonium sulfite in step (1) to 4.5 mol / L, the rest is the same as Example 1.

[0138] Example 19

[0139] This embodiment provides a comprehensive recovery method for iron aluminum alum slag, which is different from Example 1 in that, except for adjusting the concentration of ammonium sulfite in step (1) to 0.1 mol / L, the rest is the same as Example 1.

[0140] Example 20

[0141] This embodiment provides a comprehensive recovery method for iron aluminum alum slag, which is different from Example 1 in that, except for adjusting the sodium hydroxide concentration in step (3) to 350 g / L, the rest is the same as Example 1.

[0142] Example 21

[0143] This embodiment provides a comprehensive recovery method for iron aluminum alum slag, which is different from Example 1 in that, except for adjusting the sodium hydroxide concentration in step (3) to 40 g / L, the rest is the same as Example 1.

[0144] Example 22

[0145] This embodiment provides a comprehensive recovery method for iron aluminum alum slag, which is different from Example 1 in that, except for adjusting the temperature of the alkaline leaching in step (3) to 45°C, the rest is the same as Example 1.

[0146] Example 23

[0147] This embodiment provides a comprehensive recovery method for iron aluminum alum slag, which is different from Example 1 in that, except for adjusting the temperature of the alkaline leaching in step (3) to 100°C, the rest is the same as Example 1.

[0148] Example 24

[0149] This embodiment provides a comprehensive recovery method for ferroaluminum alum slag, which differs from Example 1 in that, except for adding calcium oxide to the sodium aluminate solution at a liquid-to-solid ratio of 1L:60g in step (3), the rest is the same as Example 1.

[0150] Example 25

[0151] This embodiment provides a comprehensive recovery method for ferroaluminum alum slag, which differs from Example 1 in that, except for adding calcium oxide to the sodium aluminate solution at a liquid-to-solid ratio of 1L:5g in step (3), the rest is the same as Example 1.

[0152] Example 26

[0153] This embodiment provides a comprehensive recovery method for iron aluminum alum slag, which is different from Example 1 in that, except for adjusting the impurity removal temperature in step (3) to 100°C, the rest is the same as Example 1.

[0154] Example 27

[0155] This embodiment provides a comprehensive recovery method for iron aluminum alum slag, which is different from Example 1 in that, except for adjusting the impurity removal temperature in step (3) to 40°C, the rest is the same as Example 1.

[0156] Comparative Example 1

[0157] This comparative example provides a comprehensive recovery method for iron aluminum alum slag, which differs from Example 1 in that: in step (1), the addition of the reducing agent is omitted when the iron aluminum alum slag is subjected to reduction ammonia leaching with the reduction ammonia leaching solution.

[0158] Comparative Example 2

[0159] This comparative example provides a comprehensive recovery method for iron-aluminum-aluminum slag, which differs from Example 1 in that in step (3), iron and aluminum are not separated by alkali leaching, but are separated by magnetic separation and screening to obtain iron-enriched products and aluminum-enriched products.

[0160] Test method: Use inductively coupled plasma optical emission spectrometer (ICP-OES) to test the concentrations of nickel, cobalt, copper, aluminum, and phosphorus, use ion chromatography to measure the concentration of fluorine, and calculate the leaching rate and impurity removal rate according to the following formula: Nickel, cobalt, and copper leaching rate (%) = nickel, cobalt, and copper ion content in nickel-cobalt-copper leachate / nickel, cobalt, and copper ion content in ferroaluminum alum slag × 100%, aluminum leaching rate = aluminum ion content in sodium aluminate solution / aluminum ion content in ferroaluminum slag × 100%, fluorine and phosphorus removal rate = fluorine and phosphorus ion content in purified sodium aluminate solution / fluorine and phosphorus ion content in sodium aluminate solution containing impurities fluorine and phosphorus × 100%.

[0161] The test results are shown in Table 1.

[0162] Table 1

[0163]

[0164]

[0165] The test results show that:

[0166] (1) As can be seen from Table 1, the comprehensive recovery method of ferroaluminum alum slag provided by the present invention can achieve efficient recovery of nickel, cobalt, copper and aluminum, solve the current problem of large stockpiles and difficult disposal of ferroaluminum alum slag, and has good economic and environmental benefits.

[0167] (2) Combining Example 1 with Examples 8-11, it can be seen that in Example 1, 2 mol / L ammonium carbonate solution and 5 mol / L ammonia water are used as ammonia source solutions. Compared with the use of 7 mol / L ammonium carbonate solution in Example 8, the use of 0.5 mol / L ammonium carbonate solution in Example 9, the use of 11 mol / L ammonia water in Example 10, and the use of 0.5 mol / L ammonia water in Example 11, the leaching rates of nickel, cobalt, and copper in Example 1 are 92.5%, 89.8%, and 95.8%, respectively, and the leaching rates of nickel, cobalt, and copper in Example 8 are 93.4%, 89.9%, and 93.8%, respectively. The leaching rates of nickel, cobalt and copper in Example 9 are 75.5%, 73.23% and 91.3% respectively; the leaching rates of nickel, cobalt and copper in Example 10 are 93.2%, 90.8% and 96.3% respectively; the leaching rates of nickel, cobalt and copper in Example 11 are 71.5%, 68.2% and 87.6% respectively. This shows that when the concentrations of ammonium carbonate solution and ammonia water decrease, the leaching rate of nickel, cobalt and copper decreases; when the concentrations of ammonium carbonate solution and ammonia water increase, the concentration of ammonium ions in the solution increases, resulting in an increase in the leaching rate of nickel, cobalt and copper. When the concentration is too high, no obvious increase will occur, which will cause a waste of resources.

[0168] (3) Combining Example 1 with Examples 12 and 13, it can be seen that the liquid-solid ratio of the reduced ammonia leaching in Example 1 is 5 mL:1 g. Compared with the liquid-solid ratio of the reduced ammonia leaching in Example 12 of 22 mL:1 g and the liquid-solid ratio of the reduced ammonia leaching in Example 13 of 1.5 mL:1 g, the nickel, cobalt and copper leaching rates in Example 1 are 92.5%, 89.8% and 95.8%, respectively; the nickel, cobalt and copper leaching rates in Example 12 are 94.7%, 92.6% and 97.1%, respectively; and the nickel, cobalt and copper leaching rates in Example 13 are 82.4%, 78.5% and 94.4%, respectively. This shows that when the liquid-solid ratio of the reduced ammonia leaching is low, the higher the concentration of nickel, cobalt and copper in the solution, the less conducive it is for the equilibrium to move in the leaching direction, resulting in a lower nickel, cobalt and copper leaching rate; on the contrary, when the liquid-solid ratio of the reduced ammonia leaching is high, it is more conducive to the equilibrium to move in the leaching direction, resulting in an increased nickel, cobalt and copper leaching rate.

[0169] (4) Combining Example 1 with Examples 14-17, it can be seen that the temperature of the reduction ammonia leaching in Example 1 is 85°C and the time is 300 min. Compared with the temperature of the reduction ammonia leaching in Example 14 being 100°C, the temperature of the reduction ammonia leaching in Example 15 being 20°C, the reduction ammonia leaching time in Example 16 being 650 min, and the reduction ammonia leaching time in Example 17 being 20 min, the leaching rates of nickel, cobalt, and copper in Example 1 are 92.5%, 89.8%, and 95.8%, respectively; the leaching rates of nickel, cobalt, and copper in Example 14 are 88.8%, 85.1%, and 90.2%, respectively; and the leaching rates of nickel, cobalt, and copper in Example 15 are 69.3%, 55.6%, and 78.6%, respectively. In Example 16, the leaching rates of nickel, cobalt and copper were 93.5%, 91.8% and 96.3%, respectively. In Example 17, the leaching rates of nickel, cobalt and copper were 76.7%, 72.1% and 85.6%, respectively. This shows that when the reduction ammonia leaching temperature is lowered, the decomposition efficiency of the iron aluminum alum slag structure in the reduction ammonia leaching solution is reduced, so that nickel, cobalt and copper cannot be effectively released from the iron aluminum alum slag structure, resulting in a decrease in the leaching rate of nickel, cobalt and copper. However, when the reduction ammonia leaching temperature is too high, the ammonia solution volatilizes and loses seriously, so that the ammonium ions in the solution are insufficient to complex nickel, cobalt and copper, resulting in a decrease in the leaching rate of nickel, cobalt and copper. Therefore, it can be concluded that a suitable temperature is conducive to the reaction and reduces reagent consumption, and prolonging the reaction time can increase the leaching rate of nickel, cobalt and copper.

[0170] (5) Combining Example 1 with Examples 18 and 19, it can be seen that the concentration of the reducing agent solution ammonium sulfite solution in Example 1 is 1 mol / L. Compared with the ammonium sulfite solution concentration of 4.5 mol / L in Example 18 and the ammonium sulfite solution concentration of 0.1 mol / L in Example 19, the nickel, cobalt and copper leaching rates in Example 1 are 92.5%, 89.8% and 95.8%, respectively; the nickel, cobalt and copper leaching rates in Example 18 are 92.8%, 94.6% and 96.1%, respectively; and the nickel, cobalt and copper leaching rates in Example 19 are 93.5%, 80.4% and 93.3%, respectively. This shows that when the sodium sulfite concentration increases, the cobalt leaching rate increases significantly, while the effect on the nickel and copper leaching rates is not obvious. This is because part of the cobalt exists in the trivalent form in the iron aluminum alum slag, and sodium sulfite can reduce it to divalent form, thereby increasing the leaching rate.

[0171] (6) Combining Example 1 with Examples 20-23, it can be seen that the mass concentration of the alkali sodium hydroxide used in the alkali leaching in Example 1 is 150 g / L, and the alkali leaching temperature is 90°C. Compared with the mass concentration of sodium hydroxide in Example 20 being 350 g / L, the mass concentration of sodium hydroxide in Example 21 being 40 g / L, the alkali leaching temperature in Example 22 being 45°C, and the alkali leaching temperature in Example 23 being 100°C, the leaching rate of aluminum in Example 1 is 95.2%, the leaching rate of aluminum in Example 20 is 97.7%, and the leaching rate of aluminum in Example 23 is 100°C. The leaching rate of aluminum in Example 21 is 67.4%, the leaching rate of aluminum in Example 22 is 81.7%, and the leaching rate of aluminum in Example 23 is 96.1%. This shows that when the alkali concentration used in alkali leaching is low, the leaching rate of aluminum is very low due to insufficient alkali. Increasing the alkali concentration can significantly increase the leaching efficiency of aluminum, but too high an alkali concentration will seriously corrode the equipment. Therefore, the alkali concentration should be controlled within an appropriate range. The leaching efficiency of aluminum increases with the increase of the alkali leaching temperature. This is because increasing the temperature can accelerate the reaction rate, thereby increasing the leaching rate of aluminum.

[0172] (7) Combining Example 1 with Examples 24-27, it can be seen that the mass concentration of calcium oxide used for impurity removal in Example 1 is 20 g / L, and the impurity removal temperature is 90°C. Compared with the impurity removal in step (3) of Example 24 in which calcium oxide is added to the sodium aluminate solution at a liquid-solid ratio of 1 L:60 g, the impurity removal in step (3) of Example 25 in which calcium oxide is added to the sodium aluminate solution at a liquid-solid ratio of 1 L:5 g, the impurity removal temperature in Example 26 is 100°C, and the impurity removal temperature in Example 27 is 40°C, the removal rates of phosphorus and fluorine in Example 1 are 99.5%, 99.6%, 99.7%, 99.8%, 99.9%, 99.7%, 99.9%, 99.8%, 99.9 ... The removal rates of phosphorus and fluorine in Example 25 are 95.3% and 66.7% respectively. The removal rates of phosphorus and fluorine in Example 26 are 99.7% and 95.8% respectively. The removal rates of phosphorus and fluorine in Example 27 are 91.5% and 83.2% respectively. This shows that the reduction of the liquid-solid ratio of the sodium aluminate solution to the alkaline earth metal compound and the increase of the impurity removal temperature are beneficial to the precipitation and impurity removal reaction to proceed in a favorable direction, so that the removal of impurities phosphorus and fluorine in the sodium aluminate solution is more thorough.

[0173] (8) A comparison of Comparative Example 1 and Example 1 shows that the difference from Example 1 is that the addition of a reducing agent is omitted during the reduction ammonia leaching of the aluminum aluminum slag with the reduction ammonia leaching solution in step (1). In Comparative Example 1, the absence of a reducing agent during the reduction ammonia leaching has the greatest impact on the cobalt leaching rate, resulting in a significant decrease in the leaching rate.

[0174] (9) From the comparison between Comparative Example 2 and Example 1, it can be seen that the difference from Example 2 is that in step (3), iron and aluminum are separated by magnetic separation and screening, and the aluminum recovery rate is 82%. Compared with the alkaline leaching method in the present application, the aluminum recovery rate is significantly reduced.

[0175] In summary, the present invention uniformly mixes ferroalloy slag with a reducing ammonia leaching solution, performs reducing ammonia leaching, obtains nickel, cobalt and copper leachate and ferroalloy slag through solid-liquid separation, obtains nickel, cobalt and copper products through extraction and separation, obtains sodium aluminate solution and ferroalloy slag through alkali leaching, and produces aluminum-containing products from the sodium aluminate solution. The ferroalloy slag is used as an ironmaking raw material, thereby realizing comprehensive recovery of ferroalloy slag. The overall process slag alkali quantification is greater than 60%, which solves the problems of heavy pollution, difficult disposal and low recovery efficiency of ferroalloy slag, and has good economic and environmental benefits.

[0176] 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 comprehensive recovery method for aluminum sulfate slag, characterized in that: The method comprises the following steps: (1) uniformly mixing the ferroaluminum slag with a reduction ammonia leaching solution, performing reduction ammonia leaching, and obtaining nickel, cobalt, and copper leaching solution and ferroaluminum slag through solid-liquid separation; Wherein, the aluminum alum slag includes iron, aluminum, nickel, cobalt and copper elements; the reduced ammonia leaching solution includes an ammonia source and a reducing agent; (2) extracting and separating the nickel-cobalt-copper leachate of step (1) to obtain a nickel-cobalt-copper product; (3) uniformly mixing the iron-aluminum slag described in step (1) with an alkaline solution, performing alkaline leaching, and performing solid-liquid separation to obtain a sodium aluminate solution and iron slag, wherein the sodium aluminate solution is used to prepare an aluminum-containing product, and the iron slag is used as an ironmaking raw material; There is no order between step (2) and step (3).

2. The method according to claim 1, characterized in that The source of the iron-aluminum alum slag in step (1) includes the neutralization and iron-aluminum removal process of the waste ternary battery leachate.

3. The method according to claim 1 or 2, characterized in that The ammonia source in the reduced ammonia immersion solution in step (1) comprises any one or a combination of at least two of ammonia water, ammonium carbonate, ammonium bicarbonate or ammonium sulfate, preferably a combination of ammonium carbonate and ammonia water; Preferably, the molar concentration of the ammonia source in the reduced ammonia immersion solution in step (1) is 1-15 mol / L, preferably 2-8 mol / L; Preferably, the ammonia source in the reduced ammonia immersion solution is a combination of ammonium carbonate and aqueous ammonia, the molar concentration of aqueous ammonia in the reduced ammonia immersion solution is 1-10 mol / L, and the molar concentration of ammonium carbonate in the reduced ammonia immersion solution is 1-5 mol / L; Preferably, the reducing agent in the reduced ammonia leaching solution in step (1) comprises any one or a combination of at least two of ammonium sulfite, sodium sulfite or hydrogen peroxide, preferably ammonium sulfite; Preferably, the molar concentration of the reducing agent in the reducing ammonia immersion solution in step (1) is 0.2-4 mol / L, preferably 0.5-2 mol / L.

4. The method according to any one of claims 1 to 3, characterized in that The liquid-to-solid ratio of the reduction ammonia leaching in step (1) is (2-20) mL:1 g, preferably (5-10) mL:1 g; Preferably, the temperature of the reduction ammonia leaching in step (1) is 25-95°C, preferably 50-80°C; Preferably, the time of the reduction ammonia leaching in step (1) is 30-600 min, preferably 240-480 min; Preferably, the reducing ammonia leaching in step (1) is carried out under stirring, with a stirring rate of 100-800 r / min, preferably 300-600 r / min.

5. The method according to any one of claims 1 to 4, characterized in that The alkali source in the alkali solution of step (3) comprises: sodium hydroxide and / or potassium hydroxide, preferably sodium hydroxide; Preferably, the mass concentration of the alkali source in the alkali solution of step (3) is 50-300 g / L, preferably 60-150 g / L; Preferably, the liquid-to-solid ratio of the alkaline leaching in step (3) is (1-15) mL:1 g, preferably (2-4) mL:1 g; Preferably, the alkali leaching temperature of step (3) is 50-95°C, preferably 70-95°C; Preferably, the alkali leaching time of step (3) is 30-300 min, preferably 30-90 min.

6. The method according to any one of claims 1 to 5, characterized in that In step (3), the sodium aluminate solution contains impurities such as fluorine and phosphorus, which are removed by adding an alkaline earth metal compound to obtain a purified sodium aluminate solution, which is then used to prepare the aluminum-containing product; Preferably, the alkaline earth metal compound comprises any one or a combination of at least two of calcium oxide, magnesium oxide, calcium carbonate, magnesium carbonate, calcium hydroxide or magnesium hydroxide, preferably a combination of calcium oxide and calcium hydroxide; Preferably, the liquid-to-solid ratio of the sodium aluminate solution to the alkaline earth metal compound is 1 L:(10-50) g, preferably 1 L:(10-30) g.

7. The method according to claim 6, characterized in that The impurity removal temperature is 50-95°C, preferably 70-95°C; Preferably, the impurity removal time is 60-300 min, preferably 60-120 min.

8. The method according to any one of claims 1 to 7, characterized in that The sodium aluminate solution in step (3) is used to prepare the aluminum-containing product, comprising: adding an acid solution to the sodium aluminate solution to carry out a neutralization reaction to obtain the aluminum-containing product; Preferably, the aluminum-containing product comprises a pseudo-boehmite product; Preferably, the acid solution comprises sulfuric acid; Preferably, the mass percentage concentration of the sulfuric acid is 10-50%, preferably 10-30%.

9. The method according to claim 8, characterized in that The neutralization reaction is carried out under stirring, and the stirring rate is controlled to be 100-800 r / min, preferably 300-600 r / min; Preferably, the reaction time of the neutralization reaction is 5-240 min, preferably 5-120 min; Preferably, the reaction temperature of the neutralization reaction is 50-90°C, preferably 70-90°C.

10. The method according to any one of claims 1 to 9, characterized in that The method comprises the following steps: (1) mixing the ferroalite slag and the reduction ammonia leaching solution uniformly, according to the liquid-solid ratio of (2-20) mL: 1 g, reducing ammonia leaching at 25-95 ° C for 30-600 min, the reduction ammonia leaching is carried out under stirring, the stirring rate is 100-800 r / min, and solid-liquid separation is performed to obtain nickel, cobalt and copper leachate and ferroalite slag; wherein the ferroalite slag includes iron, aluminum, nickel, cobalt and copper elements, the source of the ferroalite slag includes the process of neutralizing the waste ternary battery leachate and removing iron and aluminum, the reducing ammonia The immersion solution includes an ammonia source and a reducing agent; the molar concentration of the ammonia source in the reduced ammonia immersion solution is 1-15 mol / L, the molar concentration of the reducing agent is 0.2-4 mol / L, the ammonia source in the reduced ammonia immersion solution is a combination of ammonium carbonate and ammonia water, the molar concentration of the ammonia water in the reduced ammonia immersion solution is 1-10 mol / L, and the molar concentration of ammonium carbonate in the reduced ammonia immersion solution is 1-5 mol / L; the reducing agent includes any one of ammonium sulfite, sodium sulfite or hydrogen peroxide, or a combination of at least two thereof; (2) extracting and separating the nickel-cobalt-copper leachate of step (1) to obtain a nickel-cobalt-copper product; (3) The ferroaluminum slag of step (1) is mixed evenly with 50-300 g / L alkali solution at a liquid-solid ratio of (1-15) mL:1 g, and alkali-leached at 50-95 ° C for 30-300 min. After solid-liquid separation, a sodium aluminate solution and ferroaluminum slag are obtained. The sodium aluminate solution is used to prepare aluminum-containing products, and the ferroaluminum slag is used as an ironmaking raw material. The alkali source in the alkali solution includes sodium hydroxide and / or potassium hydroxide with a mass concentration of 50-300 g / L. In addition, the sodium aluminate solution contains impurities fluorine and phosphorus. Alkaline earth metal compounds are added to the sodium aluminate solution at a liquid-solid ratio of 1 L: (10-50) g. impurities at 50-95° C. for 60-300 minutes to obtain a purified sodium aluminate solution, which is then used to prepare an aluminum-containing product, wherein the alkaline earth metal compound includes any one or a combination of at least two of calcium oxide, magnesium oxide, calcium carbonate, magnesium carbonate, calcium hydroxide, or magnesium hydroxide; an acid solution is added to the sodium aluminate solution to carry out a neutralization reaction, the neutralization reaction being carried out under stirring at a temperature of 50-90° C., a reaction time of 5-240 minutes, and a stirring rate of 100-800 r / min to obtain a pseudo-boehmite product, wherein the acid solution includes sulfuric acid with a mass percentage concentration of 10-50%; There is no order between step (2) and step (3).

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