WO2024251052A1 - Method for recovering valuable metals - Google Patents

Abstract

Provided in the present disclosure is a method for recovering valuable metals. The method comprises: a roasting procedure, involving roasting a material obtained by battery recycling and a lithium extraction agent under the protection of an inert gas, so as to obtain a roasted product; a first leaching separation procedure, involving subjecting the roasted product to first leaching separation, so as to obtain a lithium-containing leachate and leaching residues; a calcining procedure, involving mixing the leaching residues and a slag former, and then calcining same, so as to obtain a calcined product; and a matte smelting procedure, involving mixing the calcined product, nickel matte, pressure oxidation slag and a matte agent, and then subjecting the resulting mixture to matte smelting, so as to obtain nickel-containing matte. The recovery method provided in the embodiments of the present disclosure involves a simple process flow, has a high recovery rate and generates few pollutants, such that the recovery cost of valuable metals is low, and the recovery method is environmentally friendly.

Description

Methods for recovering valuable metals

Cross-references to related publications

The present disclosure claims priority to Chinese Patent Publication No. 202310664670.2, filed on June 6, 2023, entitled “Method for Recovering Valuable Metals”, the entire contents of which are incorporated herein by reference.

Technical Field

The present disclosure relates to the technical field of valuable metal recovery, and in particular to a valuable metal recovery method.

Background Art

Valuable metals generally refer to metals with recycling value among the raw materials for metal refining, such as nickel (Ni), cobalt (Co), manganese (Mn) and lithium (Li). These valuable metals can be used as manufacturing materials for batteries, giving batteries high energy density and high cycle performance, and are therefore widely used in electric vehicles, electronic equipment, aerospace and other fields.

At present, valuable metals in batteries can be recovered from nickel intermediates (such as nickel matte) or from batteries. However, in the relevant technology, the process of recovering valuable metals is complicated, the recovery rate is low, and it is easy to produce a large amount of pollutants, resulting in high recovery costs and environmental unfriendliness.

Summary of the invention

The present invention provides a method for recovering valuable metals, which has a simple process flow, a high recovery rate, and is environmentally friendly.

The present disclosure provides a method for recovering valuable metals, the method comprising:

A roasting process is to roast the battery recycling material and the lithium extraction agent under a protective atmosphere to obtain a roasting product;

A first leaching and separation step, performing a first leaching and separation on the roasted product to obtain a lithium-containing leaching solution and leaching residue;

a calcination step, mixing the leached slag and the slag-making agent and then calcining them to obtain a calcined product;

In the matte making and smelting process, the calcined product, nickel matte, oxygen pressed slag and matte making agent are mixed and then matte making and smelting is performed to obtain nickel-containing matte.

The recovery method provided by the embodiment of the present disclosure can enrich the valuable metal lithium in the battery recovery in the leachate in large quantities through the roasting process and the first leaching and separation process, so that the recovery rate of lithium is improved. The leached slag obtained by the first leaching and separation process is subjected to a slag-making reaction through a calcination process to remove some impurities in advance, such as Ca and Mg entering the slag-making agent, which can help shorten the time of matte smelting. The calcined product, nickel matte, oxygen-pressed slag and matte-making agent obtained through the calcination process are mixed and then smelted for matte making, which can separate the valuable metal nickel in the calcined product, nickel matte and oxygen-pressed slag from other impurities, that is, valuable metals such as valuable metal nickel enter the sulfur phase to obtain matte containing valuable metals such as nickel. In this way, valuable metals such as nickel in the calcined product, nickel matte and oxygen-pressed slag can be recovered at the same time, thereby simplifying the process of recovering valuable metals from battery recyclate, nickel matte and oxygen-pressed slag and increasing the recovery rate of valuable metal nickel. In addition, the large amount of oxygen contained in nickel matte, oxygen-pressed slag and calcined products can be used to reduce the amount of oxygen required for external introduction during the sulfur making process, which can further reduce the cost of recovering valuable metals while also reducing external The safety risks brought about by the introduction of a large amount of oxygen and the requirements for fire protection and recycling equipment in the factory. Moreover, impurities such as iron and manganese in the calcined products, nickel matte and oxygen-pressed slag will enter the slag phase, and the slag phase can be directly used as a by-product in fields such as construction. This can reduce the amount of impurities removed in subsequent production lines, thereby helping to reduce the demand for impurity removal equipment and the floor space of the recycling plant, thereby further helping to reduce the recovery cost of valuable metals in battery recyclates, nickel matte and oxygen-pressed slag. In addition, in the above-mentioned processes, few pollutants are generated, thereby making the method for recovering valuable metals environmentally friendly. Therefore, the recovery method provided in the embodiments of the present disclosure has a simple process flow, a high recovery rate, and few pollutants generated, thereby making the recovery cost of valuable metals low and environmentally friendly.

In some embodiments of the present disclosure, in the roasting process, the lithium-extracting agent includes one or more of a reducing agent, a sulfate, or concentrated sulfuric acid.

In some embodiments of the present disclosure, the reducing agent includes one or more of activated carbon, hydrogen, lignite, anthracite, and carbon monoxide.

In some embodiments of the present disclosure, the sulfate includes one or more of ammonium sulfate and sodium sulfite.

In some embodiments of the present disclosure, the mass ratio of battery recyclate to lithium extraction agent is 20:(1-30).

In some embodiments of the present disclosure, the calcination temperature is 300° C.-1000° C., the heating rate is 2° C./min-10° C./min, and the calcination time is 0.5 h-5 h.

In some embodiments of the present disclosure, in the first leaching and separation process, the process includes:

a first leaching step, mixing the roasted product and a first leaching agent and performing a first leaching treatment, so that lithium in the roasted product is dissolved into a solution to obtain a first leaching slurry containing lithium;

In the first separation step, the first leaching slurry is subjected to solid-liquid separation to obtain a lithium-containing leaching solution and leaching residue.

In some embodiments of the present disclosure, in the first leaching process, the pH of the first leaching treatment is 3-7.

In some embodiments of the present disclosure, in the first leaching process, the temperature of the first leaching treatment is 30° C.-90° C., and the time of the first leaching treatment is 0.5 h-4 h.

In some embodiments of the present disclosure, in the calcination process, the mass ratio of the leached slag to the slag-forming agent is (2-20):1.

In some embodiments of the present disclosure, the calcination temperature is 600-1000° C., and the calcination time is 2 h-8 h.

In some embodiments of the present disclosure, the slag-forming agent includes one or more of silicon dioxide, diatomaceous earth, bentonite, calcium oxide, and aluminum oxide.

In some embodiments of the present disclosure, in the matte smelting process, the mass ratio of the calcined product, nickel matte, oxygen pressed slag and matte forming agent is 10:(1-30):(1-25):(0.01-30).

In some embodiments of the present disclosure, the temperature of matte smelting is 1000° C.-1500° C., and the time of matte smelting is 1 h-10 h.

In some embodiments of the present disclosure, the matte-forming agent includes a sulfur-containing element and/or a sulfur-containing compound.

In some embodiments of the present disclosure, after the first leaching and separation process, it also includes: a drying process of drying the leached slag, nickel matte and oxygen-pressed slag respectively.

In some embodiments of the present disclosure, the drying temperature is 100° C.-300° C., the drying time is 0.5 h-4 h, and air and/or oxygen are introduced during the drying process.

In some embodiments of the present disclosure, the method for recovering valuable metals further comprises:

A first slurrying step, mixing the matte with a solvent and then performing a first slurrying treatment to obtain a first slurry liquid;

a second leaching and separation step, wherein the first slurry and the second leaching agent are mixed and then subjected to a second leaching treatment to dissolve the nickel in the first slurry into the solution to obtain a second leaching slurry containing nickel, and the second leaching slurry is subjected to a solid-liquid separation treatment to obtain a second filtrate and a second filter residue;

In the extraction step, the second filtrate is mixed with an extractant and then subjected to extraction treatment to obtain an extract and a raffinate.

In some embodiments of the present disclosure, in the first slurrying process, the matte is ground, and the ground matte is mixed with a solvent at a solid-liquid ratio of 1 g: (2-15) mL to perform a first slurrying process to obtain a first slurry liquid.

In some embodiments of the present disclosure, the particle size of the matte after grinding is less than or equal to 48 μm.

In some embodiments of the present disclosure, the temperature of the first slurrying treatment is 30° C.-90° C., and the time of the first slurrying treatment is 0.5 h-5 h.

In some embodiments of the present disclosure, in the second leaching and separation step, the second leaching agent includes concentrated sulfuric acid, wherein the amount of concentrated sulfuric acid added is 1 to 1.6 times the theoretical amount required to completely leach out the nickel in the first slurry.

In some embodiments of the present disclosure, the temperature of the second leaching treatment is 50° C.-100° C., and the time of the second leaching treatment is 2 h-8 h.

In some embodiments of the present disclosure, the method for recovering valuable metals further comprises:

a third leaching and separation step, mixing the second filter residue and the third leaching agent, performing a third leaching treatment to obtain a third leaching slurry, and performing a solid-liquid separation treatment on the third leaching slurry to obtain a third filtrate and a third filter residue;

a second slurrying step, mixing the third filter residue and the solvent and then performing a second slurrying treatment to obtain a second slurry liquid;

In the fourth leaching and separation step, oxygen is introduced into the second slurry to perform oxygen pressure leaching treatment to obtain a fourth leaching slurry, and the fourth leaching slurry is subjected to solid-liquid separation treatment to obtain a fourth filtrate and a fourth filter residue.

In some embodiments of the present disclosure, in the third leaching and separation step, the third leaching agent includes concentrated sulfuric acid, wherein the amount of concentrated sulfuric acid added is 1 to 1.3 times the theoretical amount required to completely leach out the nickel in the second filter residue.

In some embodiments of the present disclosure, in the third leaching and separation process, the temperature of the third leaching treatment is 40° C.-100° C., and the time of the third leaching treatment is 0.5 h-4 h;

In some embodiments of the present disclosure, in the second slurrying step, the third filter residue and the solvent have a solid-liquid ratio of 1 g: (3-15) mL.

In some embodiments of the present disclosure, in the fourth leaching and separation step, the oxygen partial pressure of oxygen pressure leaching is 0.6 MPa-1.2 MPa.

In some embodiments of the present disclosure, in the fourth leaching and separation step, the temperature of the oxygen pressure leaching is 150° C.-220° C., and the time of the oxygen pressure leaching is 2 h-8 h.

BRIEF DESCRIPTION OF THE DRAWINGS

In order to more clearly illustrate the specific embodiments of the present disclosure or the technical solutions in the prior art, the drawings required for use in the specific embodiments or the description of the prior art will be briefly introduced below. Obviously, the drawings described below are some embodiments of the present disclosure. For those skilled in the art, other drawings can be obtained based on these drawings without paying any creative work.

FIG1 is a flow chart of a method for recovering valuable metals provided in some embodiments of the present disclosure;

FIG2 is a flow chart of a method for recovering valuable metals provided in some embodiments of the present disclosure;

FIG3 is a flow chart of a method for recovering valuable metals provided in some embodiments of the present disclosure;

FIG4 is a flow chart of a method for recovering valuable metals provided in some embodiments of the present disclosure;

FIG5 is a flow chart of a method for recovering valuable metals provided in some embodiments of the present disclosure.

DETAILED DESCRIPTION

"Scope" disclosed in the present disclosure is limited in the form of lower limit and upper limit, and a given range is limited by selecting a lower limit and an upper limit, and the selected lower limit and upper limit define the boundary of a special range. The scope limited in this way can be including end values or not including end values, and can be arbitrarily combined, that is, any lower limit can form a scope with any upper limit combination. For example, if the scope of 60-120 and 80-110 is listed for a particular parameter, it is understood that the scope of 60-110 and 80-120 is also expected. In addition, if the minimum range values 1 and 2 listed, and if the maximum range values 3,4 and 5 are listed, the following scope can all be expected: 1-3, 1-4, 1-5, 2-3, 2-4 and 2-5. In the present disclosure, unless otherwise specified, the numerical range "a-b" represents the abbreviation of any real number combination between a and b, wherein a and b are real numbers. For example, the numerical range "0-5" means that all real numbers between "0-5" are listed in this document, and "0-5" is just an abbreviation of these numerical combinations. In addition, when a parameter is expressed as an integer ≥ 2, it is equivalent to disclosing that the parameter is, for example, an integer of 2, 3, 4, 5, 6, 7, 8, 9, 10, 11 and 12.

If not otherwise specified, all embodiments and optional embodiments of the present disclosure may be combined with each other to form a new technical solution.

Unless otherwise specified, all technical features and optional technical features of the present disclosure can be combined with each other to form a new technical solution.

If not otherwise specified, all steps of the present disclosure may be performed sequentially or randomly, preferably sequentially. For example, the method includes steps (a) and (b), which means that the method may include steps (a) and (b) performed sequentially, or may include steps (b) and (a) performed sequentially. For example, the method may further include step (c), which means that step (c) may be added to the method in any order, for example, the method may include steps (a), (b) and (c), or may include steps (a), (c) and (b), or may include steps (c), (a) and (b), etc.

If there is no special explanation, the "include" and "comprising" mentioned in the present disclosure represent open or closed forms. For example, the "include" and "comprising" may mean that other components not listed may also be included or only the listed components may be included or only the listed components may be included.

If not specifically stated, in the present disclosure, the term "or" is inclusive. For example, the phrase "A or B" means "A, B, or both A and B". More specifically, any of the following conditions satisfies the condition "A or B": A is true (or exists) and B is false (or does not exist); A is false (or does not exist) and B is true (or exists); or both A and B are true (or exist).

In the present disclosure, battery recyclate refers to a mixture of main battery components such as positive electrodes recovered from batteries, wherein the batteries may be nickel-sulfur secondary batteries, nickel-hydrogen secondary batteries, lithium-ion secondary batteries, etc. In some examples, the positive electrode in the lithium-ion secondary battery includes a positive electrode current collector and a positive electrode active material, and the positive electrode current collector may be a metal foil such as aluminum, copper, and nickel, and the positive electrode active material may include valuable metals such as nickel (Ni), cobalt (Co), manganese (Mn), and lithium (Li). In some specific examples, battery recyclate may be a mixture of valuable metals such as Ni, Co, Mn, and Li and carbon powder obtained through processes such as disassembly, crushing, screening, pyrolysis, and sorting.

Lithium extraction agent refers to a compound that can convert lithium metal in battery recycling into a compound that is easily soluble in water or a weakly acidic solution environment.

Nickel matte refers to a mixture of metal sulfides containing nickel, such as Ni3S2. In some examples, nickel matte can be an intermediate product in the nickel smelting process, such as low-grade nickel matte and high-grade nickel matte, wherein the nickel content in low-grade nickel matte is usually below 40%, and the nickel content in high-grade nickel matte is usually above 60%.

Oxygen-pressed slag refers to slag containing elements such as nickel, iron and oxygen. In some examples, oxygen-pressed slag may be slag produced by an oxygen-pressing process, the main component of which is iron oxide, and the total content of nickel and cobalt is about 2%-15%. The oxygen-pressing process usually uses oxygen or air (oxygen in the air) as an oxidant, which is fully mixed with the slurry under stirring to react with low-valent minerals such as reduced sulfide minerals to leach the target metal into the solution, thereby completing the efficient extraction of the target metal from the ore.

Referring to FIG. 1 , the present disclosure provides a method for recovering valuable metals, the method comprising:

S100, a roasting process, roasting the battery recycling material and the lithium extraction agent under a protective atmosphere to obtain a roasting product.

S200, a first leaching and separation step, performing a first leaching and separation on the roasted product to obtain a lithium-containing leaching solution and leaching residue.

S400, a calcination step, mixing the leached slag and the slag-making agent and calcining them to obtain a calcined product;

S500, matte making and smelting process, mixing the calcined product, nickel matte, oxygen pressed slag and matte making agent, and then performing matte making and smelting to obtain nickel-containing matte.

The recovery method provided by the embodiment of the present disclosure can enrich the valuable metal lithium in the battery recovery in the leachate in large quantities through the roasting process and the first leaching and separation process, so that the recovery rate of lithium is improved. The leached slag obtained by the first leaching and separation process is subjected to a slag-making reaction through a calcination process to remove some impurities in advance, such as Ca and Mg entering the slag-making agent, which can help shorten the time of matte smelting. The calcined product, nickel matte, oxygen-pressed slag and matte-making agent obtained through the calcination process are mixed and then smelted for matte making, which can separate the valuable metal nickel in the calcined product, nickel matte and oxygen-pressed slag from other impurities, that is, the valuable metal nickel enters the sulfur phase to obtain nickel-containing matte, so that valuable metals such as nickel in the calcined product, nickel matte and oxygen-pressed slag can be recovered at the same time, thereby simplifying the process of recovering valuable metals from battery recyclate, nickel matte and oxygen-pressed slag and increasing the recovery rate of valuable metal nickel. In addition, the large amount of oxygen contained in nickel matte, oxygen-pressed slag and calcined products can be used to reduce the amount of oxygen required for external introduction during the sulfur making process, which can further reduce the cost of recovering valuable metals while reducing the safety risks brought about by the large amount of oxygen introduced from the outside and the requirements for fire protection and recovery equipment in the factory. Moreover, impurities such as iron and manganese in the calcined products, nickel matte and oxygen-pressed slag will enter the slag phase, which can be directly used as a by-product in fields such as construction. This can reduce the amount of impurities removed in subsequent production lines, thereby helping to reduce the demand for impurity removal equipment and the floor space of recycling plants, thereby further helping to reduce the cost of recovering valuable metals in battery recyclates, nickel matte and oxygen-pressed slag. In addition, in the above-mentioned processes, few pollutants are generated, thereby making the method for recovering valuable metals environmentally friendly. Therefore, the recovery method provided in the embodiments of the present disclosure has a simple process flow, a high recovery rate, and few pollutants generated, thereby making the recovery cost of valuable metals low and environmentally friendly.

In the embodiments of the present disclosure, the lithium extraction agent used in the roasting process can convert the lithium metal in the battery recyclate into a compound that is easily soluble in water or a weakly acidic solution environment. In some embodiments, the lithium extraction agent may include one or more of a reducing agent, a sulfate, or concentrated sulfuric acid, wherein the reducing agent mainly undergoes a reduction reaction during the roasting process to form carbonate or bicarbonate, and the carbonate and bicarbonate can form soluble lithium carbonate or lithium bicarbonate with the lithium ions in the battery recyclate, thereby facilitating the leaching and recovery of lithium. Exemplarily, the reducing agent may include one or more of activated carbon, hydrogen, lignite, anthracite, and carbon monoxide, wherein the activated carbon may include coconut shell powdered activated carbon and/or wood activated carbon. The above-mentioned reducing agents are cost-effective, which can help reduce the cost of lithium extraction while also helping battery recycling. The lithium ions in the material are converted into soluble lithium carbonate or lithium bicarbonate to improve the lithium recovery rate.

Concentrated sulfuric acid and sulfate mainly form soluble lithium sulfate with lithium ions in battery recyclate during the roasting process to facilitate lithium leaching and recovery. Exemplarily, the sulfate may include one or more of ammonium sulfate and sodium sulfite.

In this embodiment, the mass ratio of the lithium extraction agent to the battery recyclate is within a suitable range, which can also help improve the recovery rate of lithium. In some embodiments, the mass ratio of the lithium extraction agent to the battery recyclate is 20:(1-30). When the mass ratio of the lithium extraction agent to the battery recyclate is within the above-mentioned suitable range, the utilization rate of the lithium extraction agent can be increased, and the recovery rate of lithium can also be improved.

In other embodiments, the mass ratio of the lithium extraction agent to the battery recyclate can also be 20:(3-28), 20:(5-24), 20:(8-21), etc.

In addition, in the calcination process of S100, the calcination parameters are within a suitable range and can also facilitate the occurrence of a thermodynamic reaction of lithium ions forming soluble salts. In some embodiments of the present disclosure, the calcination temperature is 300°C-1000°C, the heating rate is 2°C/min-10°C/min, and the calcination time is 0.5h-5h.

Moreover, roasting is carried out under a protective atmosphere, which can facilitate the oxidation reaction of carbon in the battery recyclate, thereby helping to form soluble lithium carbonate or lithium bicarbonate with lithium ions, thereby improving the recovery rate of lithium. In this embodiment, the gas used in the protective atmosphere can be any protective gas known in the art. For example, an inert gas or nitrogen, wherein the inert gas can include helium (He), neon (Ne), argon (Ar), etc.

Please refer to FIG. 2 , in some embodiments, the first leaching and separation process of S200 includes:

S210, a first leaching step, mixing the roasted product with a first leaching agent and performing a first leaching treatment to dissolve lithium in the roasted product into a solution to form a first leaching slurry;

S220, a first separation step, performing solid-liquid separation on the first leaching slurry to obtain a lithium-containing leaching solution and leaching residue.

In the above embodiments, the first leaching process of S210 uses a solvent to mix and slurry the roasted product, which can provide a good leaching environment. The addition of the first leaching agent dissolves the lithium in the roasted product into the solution to form a first leaching slurry. Then, through the first separation process of S220, the first leaching slurry is subjected to solid-liquid separation to obtain a lithium-containing leaching solution and leaching residue.

In some embodiments, the solvent used in the first leaching step of S210 may be water, or an acidic washing liquid generated during the recovery of valuable metals, or a mixed solvent formed by water and the acidic washing liquid. The acidic washing liquid as a solvent can be beneficial to the reuse of waste liquid and can also achieve the effect of pre-leaching, thereby reducing the amount of the first leaching agent added.

In some embodiments, in the first leaching step of S210, the liquid-to-solid ratio of the solvent to the roasted product is (1-10) mL: 1 g. When the liquid-to-solid ratio of the solvent to the roasted product is within the above range, it can be beneficial to the leaching of lithium, and the solvent content in the leached residue obtained by subsequent separation can be within a reasonable range, thereby facilitating the subsequent reaction of the leached residue.

In this embodiment, the liquid-to-solid ratio generally refers to the ratio of the volume of the solvent to the mass of the calcined product.

In some embodiments, in the first leaching step of S210, the pH of the first leaching treatment is 3-7. When the pH of the first leaching treatment is within the above range, lithium can be easily formed into ions and exist in the aqueous solution, thereby reducing the recovery cost of lithium, thereby helping to reduce the recovery cost of valuable metals.

In the above embodiments, the addition of the first leaching agent can make the pH of the first leaching treatment within the range of 3-7, and the first leaching agent can be any compound that can make the pH of the first leaching treatment within the above range, such as concentrated sulfuric acid.

In addition, in the first leaching process of S210, the temperature of the first leaching treatment is 30°C-90°C, and the time of the first leaching treatment is 0.5h-4h. When the temperature of the first leaching treatment is within the above range, it is conducive to the occurrence of leaching reaction kinetics. When the time of the first leaching treatment is within the above range, it is helpful to improve the leaching reaction effect.

In the first separation step of S220, the solid-liquid separation treatment method can be any method of separating solids and liquids known in the art, for example, the solid-liquid separation treatment can be performed by centrifugation, tilting, filtration, and the like.

Furthermore, the leached slag obtained by the above solid-liquid separation treatment and the slag-forming agent are mixed and calcined to carry out a slag-forming reaction, so that some impurities (such as Ca and Mg, etc.) in the leached slag enter the slag-forming agent, thereby removing some impurities in the leached slag in advance. When the leached slag with reduced impurity content is used for matte smelting, the time required for matte smelting will be shortened, thereby improving the efficiency of matte smelting. In addition, the calcination process can also remove part of the carbon contained in the leached slag to reduce the influence of this part of carbon on the subsequent sulfur-forming reaction.

In some embodiments, in the calcination process of S400, the mass ratio of the leached slag to the slag-forming agent is (2-20): 1. When the mass ratio of the leached slag to the slag-forming agent is within the above range, it can help improve the slag-forming reaction efficiency to improve the nickel recovery rate, and can also further reduce the recovery cost of valuable metals.

In some embodiments, the slag-forming agent includes one or more of silicon dioxide, diatomaceous earth, bentonite, calcium oxide, and aluminum oxide. It is understandable that the slag-forming agent can be any one of silicon dioxide, diatomaceous earth, bentonite, calcium oxide, and aluminum oxide, or a mixture formed by mixing any two of silicon dioxide, diatomaceous earth, bentonite, calcium oxide, and aluminum oxide. Wherein, when the slag-forming agent includes a mixture, the components in the mixture cooperate with each other to selectively promote the impurities in the leached slag to enter the slag phase, thereby reducing the impurities contained in the sulfur phase, thereby improving the recovery rate of valuable metals.

In some specific embodiments, the calcination process may cause magnesium and part of the iron in the leached slag to undergo a slag-forming reaction with a slag-forming agent so that they enter the slag phase, wherein the following slag-forming reaction may occur:
10Fe ₂ O ₃ +FeS=7Fe ₃ O ₄ +SO ₂ ;
3Fe ₃ O ₄ +FeS+5SiO ₂ =5(2FeO·SiO ₂ )+SO ₂ ;
FeO+SiO ₂ =2FeO·SiO ₂ ;
CaO+SiO ₂ =CaO·SiO ₂ ;
MgO+SiO ₂ =MgO·SiO ₂ ;
2Mn ₃ O ₄ +S＝6MnO+SO ₂ ;
MnO+SiO ₂ =MnO·SiO ₂ .

In this embodiment, the matte smelting process S500 mixes the calcined product, nickel matte, oxygen-pressed slag and sulfur-forming agent and then performs matte smelting, which can make the impurities in the calcined product, nickel matte and oxygen-pressed slag enter the slag phase, and can also make the nickel in the calcined product, nickel matte and oxygen-pressed slag enter the sulfur phase, thereby simultaneously realizing the recovery of valuable metal nickel in battery recyclate, nickel matte and oxygen-pressed slag, and also has a high recovery rate of valuable metals. In addition, the oxygen element contained in the oxygen-pressed slag reduces the amount of oxygen introduced during the matte smelting process, so that the oxygen-pressed slag can be fully utilized to further reduce While reducing the cost of valuable metal recovery, it also reduces the safety risks brought about by the introduction of large amounts of oxygen from the outside and the demand for fire protection and recovery equipment in the plant.

In some embodiments, the matte-making agent includes sulfur-containing elements and/or sulfur-containing compounds. The reasonable combination of elemental sulfur and sulfur-containing compounds with the calcined product, nickel matte, and oxygen-pressed slag can enrich the valuable metal nickel in the calcined product, nickel matte, and oxygen-pressed slag in the sulfur phase through matte-making smelting, thereby helping to improve the recovery rate of valuable metals in the calcined product, nickel matte, and oxygen-pressed slag.

In the above embodiments, the sulfur-containing compound may be one or more of sulfur and sulfide ore.

In some embodiments, in the matte smelting process of S400, the mass ratio of the calcined product, nickel matte, oxygen slag and matte-forming agent is 10:(1-30):(1-25):(0.01-30). When the mass ratio of the calcined product, nickel matte, oxygen slag and matte-forming agent is within the above range, it can help the valuable metal nickel in the calcined product, nickel matte and oxygen slag to enter the sulfur phase to form nickel-containing matte, thereby further improving the recovery rate of the valuable metal.

In some embodiments, the temperature of matte smelting is 1000°C-1500°C, and the time of matte smelting is 1h-10h. When the temperature and time of matte smelting are respectively within the above ranges, it is not only conducive to the thermodynamic chemical reaction, but also conducive to improving the efficiency of the reaction. Among them, the reactions that may occur in matte smelting are as follows:

Decomposition reaction of high-valent sulfides:
Fe ₇ S ₈ =7FeS+1/2S ₂ ;
2CuFeS ₂ =Cu ₂ S+2FeS+1/2S ₂ ;
(Ni, Fe) ₉ S ₈ =(x/3)Ni ₃ S ₂ +(9-x)FeS+(x/6-1/2)S ₂ ;
(Ni, Fe) ₉ S ₈ =(x/3)Ni ₃ S ₂ +(9-x)FeS+(x/6-1/2)S ₂ ;

Oxidation reaction of low-valent sulfides:
2FeS+3O ₂ =2FeO+2SO ₂ ;
Ni ₃ S ₂ +7/2O ₂ =3NiO+2SO ₂ ;
3FeS+3NiO=Ni ₃ S ₂ +3FeO+1/2S ₂ ;
3NiS＝Ni ₃ S ₂ +S;
Co ₃ S ₂ +7/2O ₂ =3CoO+2SO ₂ ;
3FeS+3CoO=Co ₃ S ₂ +3FeO+1/2S ₂ ;
3CoS＝Co ₃ S ₂ +S;

The chemical reactions that occur in battery recycling are mainly as follows:
NiO+S＝NiS+1/2O ₂ ;
CoO+S＝CoS+1/2O ₂ ;
Ni+S＝NiS;
Co+S＝CoS;
MnO+SiO ₂ =MnO·SiO ₂ .

Through the above reaction, the valuable metal nickel in the calcined product, nickel matte and oxygen-pressed slag can enter the sulfur phase to form matte, while manganese enters the slag phase, which can further remove impurities in the nickel phase, thereby further improving the recovery rate of the valuable metal nickel.

In the above embodiments, the battery may be a ternary battery, that is, the battery contains valuable metals such as Ni, Co, and Mn. The valuable metals such as Ni and Co can be recycled through the above recycling method.

Please refer to FIG. 3 , in some embodiments, after the first leaching and separation step in S200 , the method for recovering valuable metals further includes:

S300, a drying process, drying the leached slag, nickel matte and oxygen-pressed slag respectively.

In the above-mentioned embodiment, the drying process of drying the leached slag, nickel matte and oxygen-pressed slag can remove part of the moisture in the leached slag, nickel matte and oxygen-pressed slag.

In some embodiments, the drying temperature is 100° C.-300° C., and the drying time is 0.5 h-4 h. The drying parameters of the above drying process can further reduce the water content in the leached slag, nickel matte and oxygen-pressed slag, and reduce the adverse effects of water on the chemical reactions in the subsequent matte smelting process, such as the water vapor formed taking away part of the sulfiding agent.

In the above embodiments, introducing air and/or oxygen during the drying process can help accelerate the drying of the leached slag, nickel matte and oxygen-pressed slag.

As shown in FIG4 , after the matte smelting process, the nickel-containing matte obtained can be directly sold as a commodity, or can be further processed to separate nickel from other metal elements to prepare raw materials for producing batteries. Therefore, the present disclosure also provides a method for further processing the nickel-containing matte, including:

S600, a first slurrying step, mixing the matte with a solvent and then performing a first slurrying treatment to obtain a first slurry liquid.

S700, a second leaching and separation step, mixing the first slurry and the second leaching agent and performing a second leaching treatment to dissolve the nickel in the first slurry into the solution to obtain a second leaching slurry containing nickel, and performing a solid-liquid separation treatment on the second leaching slurry to obtain a second filtrate and a second filter residue;

S800, mixing the second filtrate and the extractant and performing extraction treatment to obtain an extract and a raffinate.

In the above embodiments, the first slurrying process can evenly disperse the matte in the solvent and form a first slurry liquid, so that the leaching reaction in the subsequent second leaching and separation process can be quickly completed. The second leaching and separation process can leach nickel and separate the second filtrate containing nickel and the second filter residue containing other impurities after solid-liquid separation. Then, the second filtrate is mixed with an extractant and then subjected to extraction treatment, so that part of the impurities are extracted into the extract, and nickel is enriched in the raffinate, thereby achieving separation of nickel from other metal impurities.

In some embodiments, in the first slurrying step of S600 , water may be used as a solvent, which may help reduce the recovery cost of valuable metals.

In some embodiments, matte is ground, and the ground matte is mixed with a solvent at a solid-liquid ratio of 1 g: (2-15) mL for a first slurry treatment to obtain a first slurry liquid. The grinding of matte can reduce the particle size of matte to help it be evenly dispersed in the solvent. The grinding process can also give the matte particles a larger reaction area, which is helpful for the subsequent matte leaching reaction. The ground matte is mixed with the solvent at the above solid-liquid ratio, which can further help improve the dispersion effect of matte to facilitate nickel leaching.

In this embodiment, the solid-liquid ratio generally refers to the ratio of the mass of the ground matte to the volume of the solvent.

In some embodiments, the particle size of the matte after grinding is less than or equal to 48 μm. This can make the matte have a suitable particle size, which can further improve its dispersion effect. Moreover, the particle size of the matte after grinding is within the above range, which can make it have a suitable reaction area, so as to facilitate the leaching of nickel.

In the above embodiment, the ground matte can be screened through a sieve with an appropriate mesh number so that the particle size of the ground matte meets the requirements, and the large particles that do not pass through the sieve can be further ground to reach the corresponding required particle size. Exemplarily, a sieve with a mesh size greater than or equal to 300 is used to screen out matte with a particle size less than or equal to 48 μm.

In some embodiments, the temperature of the first slurry treatment is 30° C.-90° C., and the time of the first slurry treatment is 0.5 h-5 h. When the temperature and time of the first slurry treatment are within the above range, it is beneficial to the thermal motion of molecules and helps to evenly disperse the ground matte into the solvent, which can help shorten the time of the subsequent leaching reaction.

In some embodiments, in the second leaching and separation step of S700, the second leaching agent may include concentrated sulfuric acid, and the amount of the second leaching agent added is 1 to 1.6 times the theoretical amount required to completely leach the nickel in the first slurry. By properly matching the leaching agent, the leaching rate of nickel can be improved while reducing the leaching cost.

In some embodiments, the temperature of the second leaching treatment is 50° C.-100° C., and the time of the second leaching treatment is 2 h-8 h. When the temperature and time of the second leaching treatment are within the above range, it is beneficial to the thermal motion of molecules and helps to improve the leaching efficiency.

In addition, in some embodiments, in the extraction process of S800, the extractant may include one or more of di(2-ethylhexyl) phosphate (P204), 2-ethylhexyl phosphoric acid (P507), and di(2,4,4-trimethylpentyl)phosphinic acid (C272). The specific type of the extractant can be selected according to the impurity situation of the solution.

In the above embodiments, after the second filtrate is subjected to extraction treatment, the obtained raffinate includes nickel sulfate, and the extract can also extract impurities from the extractant through the stripping solution (a mixed sulfate system formed by impurities such as Fe, Zn, Ca, Mg, Cd, etc.), and the extractant can be recovered after separation, thereby further reducing the recovery cost of valuable metals.

In addition, the second filter residue produced by the second leaching and separation step in S700 can be further processed to further recover part of the valuable metal nickel contained in the filter residue and form byproducts such as oxygen-pressed slag. Referring to FIG5 , in some embodiments, the method for recovering valuable metals further includes:

S900, a third leaching and separation step, mixing the second filter residue and the third leaching agent, performing a third leaching treatment to obtain a third leaching slurry, and performing a solid-liquid separation treatment on the third leaching slurry to obtain a third filtrate and a third filter residue;

S1000, a second slurrying step, mixing the third filter residue and the solvent and then performing a second slurrying treatment to obtain a second slurry liquid;

S2000, a fourth leaching and separation step, introducing oxygen into the second slurry to perform oxygen pressure leaching treatment to obtain a fourth leaching slurry, and performing solid-liquid separation treatment on the fourth leaching slurry to obtain a fourth filtrate and a fourth filter residue.

In the above embodiment, the nickel in the second filter residue can be further recovered to further improve the recovery rate of valuable metals.

In some embodiments, in the third leaching and separation process of S900, the third leaching agent includes concentrated sulfuric acid, and the amount of the third leaching agent added is 1 to 1.3 times the theoretical amount required to completely leach the nickel in the second filter residue. By properly matching the leaching agent, the leaching rate of nickel can be improved while reducing the leaching cost.

In some embodiments, the temperature of the third leaching treatment is 40° C.-100° C., and the time of the third leaching treatment is 0.5 h-4 h.

In some embodiments, in the second slurrying process of S1000, the solvent may be water, and the third filter residue and the solvent are subjected to a solid-liquid ratio of 1 g: (3-15) ml; the second slurrying process is performed. This is beneficial to the dispersion of the filter residue, and can also facilitate the occurrence of subsequent leaching reactions and reduce the occurrence of the destruction of the water system balance in the system.

In some embodiments, in the second slurrying process of S1000, the temperature of the second slurrying treatment is 30°C-90°C, and the time of the second slurrying treatment is 0.5h-5h. When the temperature and time of the second slurrying treatment are within the above range, it is beneficial to the thermal motion of molecules and helps to evenly disperse the filter residue into the third solvent, which can help to shorten the time of the subsequent leaching reaction.

In some embodiments, the oxygen partial pressure of oxygen pressure leaching is 0.6 MPa-1.2 MPa. When the oxygen partial pressure of oxygen pressure leaching is within the above range, it can facilitate the oxygen pressure leaching reaction while reducing the safety risks caused by oxygen.

In some embodiments, the temperature of oxygen pressure leaching is 150° C.-220° C., and the time of oxygen pressure leaching is 2 h-8 h, which can help remove the alloy phase in the material and reduce the safety risk caused by the reaction between the alloy and concentrated sulfuric acid in the oxygen pressure reaction.

In some specific embodiments, the method for recovering valuable metals may include:

In the roasting process, the battery recycling material and the lithium extraction agent are mixed evenly, and then roasted under a protective gas to obtain a roasted product;

a first leaching step, mixing the roasted product and a first leaching agent and performing a first leaching treatment to dissolve lithium in the roasted product into a solution to form a first leaching slurry;

A first separation step is to perform solid-liquid separation on the first leaching slurry to obtain a lithium-containing leaching solution and leaching residue, wherein the lithium-containing leaching solution can be used as a lithium source for producing positive electrode active materials in lithium-ion secondary batteries;

Drying process, drying the leached slag, nickel matte and oxygen pressed slag respectively;

Calcination step: In the calcination step, the leached slag and the slag-making agent are mixed and calcined to obtain a calcined product;

A matte smelting step, in which the calcined product, nickel matte, oxygen-pressed slag and a matte-forming agent are mixed and then smelted to obtain a nickel-containing matte, which can be used as a material for producing a ternary precursor;

In the first slurrying step, the matte is mixed with a solvent and then subjected to a first slurrying treatment to obtain a first slurry liquid.

a second leaching and separation step, wherein the first slurry and the second leaching agent are mixed and then subjected to a second leaching treatment to dissolve the nickel in the first slurry into the solution to obtain a second leaching slurry containing nickel, and the second leaching slurry is subjected to a solid-liquid separation treatment to obtain a second filtrate and a second filter residue;

An extraction step, mixing the second filtrate and the extractant and then performing an extraction treatment to obtain an extract and a raffinate containing nickel, wherein the raffinate can be used as a ternary precursor material;

a third leaching and separation step, mixing the second filter residue and the third leaching agent and performing a third leaching treatment to obtain a third leaching slurry, and performing a solid-liquid separation treatment on the third leaching slurry to obtain a third filtrate and a third filter residue, wherein the third filtrate can be recycled, for example, can be added to the leaching and separation step again;

a second slurrying step, mixing the third filter residue and the solvent and then performing a second slurrying treatment to obtain a second slurry liquid;

In the fourth leaching and separation step, oxygen is introduced into the second slurry to perform oxygen pressure leaching treatment to obtain a fourth leaching slurry. The slurry is subjected to solid-liquid separation treatment to obtain a fourth filtrate containing nickel and a fourth filter residue, wherein the fourth filtrate can be used as a ternary precursor material, and the fourth filter residue can be directly used in building materials.

The following examples describe the disclosure in more detail, and these examples are only for illustrative purposes, as it is obvious to those skilled in the art that various modifications and variations are made within the scope of the disclosure. Unless otherwise stated, all parts, percentages, and ratios reported in the following examples are by mass, and all reagents and raw materials used in the examples are commercially available or synthesized according to conventional methods, and the instruments used in the examples are commercially available.

Example 1

This embodiment provides a method for recovering valuable metals, comprising the following steps:

In the calcination process, 200 g of battery recyclate (the mass fractions of each component in the battery recyclate are shown in Table 1) and 10 g of coconut shell powdered activated carbon were mixed evenly and loaded into a crucible. The mixture was calcined at 400 °C for 0.5 h under inert gas protection to obtain a calcined product, with a heating rate of 2 °C/min.

In the first leaching and separation step, the roasted product is added into pure water at a liquid-solid ratio of 5 ml:1 g, concentrated sulfuric acid is added to adjust the pH of the initial solution to 4±0.1, and the reaction is stirred at 40°C for 2 hours. During the process, the pH is monitored in real time and acid is supplemented in time to ensure that the pH of the system is maintained at 4±0.1 for the first leaching treatment; after the leaching is completed, a lithium-containing leaching solution and a leaching residue are obtained by filtration; wherein, the Li leaching rate in the leaching solution is 96.18%, the Ni leaching rate is 4.98%, the Co leaching rate is 0.44%, and the Mn leaching rate is 0.5%.

In the drying process, the leached slag, low nickel matte and oxygen pressed slag are dried separately, and air is introduced during the drying process. The drying temperature is 100°C, the drying time is 4 hours, and the moisture content of the dried material is 0.5%.

In the calcination process, the dried leached slag is mixed with a slag-forming agent in a mass ratio of 8:1 and then calcined. The slag-forming agent is silicon dioxide and bentonite. The calcination temperature is 850°C and the calcination time is 6 hours.

In the matte smelting process, the calcined product, nickel matte, oxygen-pressed slag and sulfur are mixed in a mass ratio of 10:20:2:10 and then smelted at a temperature of 1450°C for 1.5 hours to obtain a matte containing nickel and cobalt. Among them, the compositions of low-grade nickel matte and oxygen-pressed slag are shown in Table 1. Calculation shows that the Ni recovery rate is 99.23% and the Co recovery rate is 99.85%.

The method for recovering valuable metals provided in this embodiment also includes:

In the first slurrying step, the obtained matte is ball-milled and then sieved through 300 mesh. The sieved matte is mixed with water at a solid-liquid ratio of 1g:9mL for the first slurrying treatment to obtain the first slurry liquid, wherein the temperature of the first slurrying treatment is 30°C and the time of the first slurrying treatment is 2h.

In the second leaching and separation step, concentrated sulfuric acid is added to the first slurry liquid for a second leaching treatment. The amount of sulfuric acid added is 1.1 times the theoretical amount. The temperature of the second leaching treatment is 70°C and the time of the second leaching treatment is 4 hours. After the second leaching treatment, filtering is performed to obtain a second filtrate and a second filter residue.

In the extraction step, the second filtrate is mixed with the P507 extractant and then subjected to extraction treatment to obtain an extract and a raffinate, wherein the raffinate is a nickel sulfate and cobalt sulfate solution.

The third leaching and separation step is to add concentrated sulfuric acid to the second filter residue for a third leaching treatment, the amount of sulfuric acid added is 1.1 times the theoretical amount, the temperature of the third leaching treatment is 80°C, and the time of the third leaching treatment is 2h; after the third leaching treatment, filtering is performed to obtain a third filtrate and a third filter residue, wherein the third filtrate contains nickel sulfate and cobalt sulfate.

In the second slurrying step, the third filter residue and water are subjected to a second slurrying treatment at a solid-liquid ratio of 1 g:10 mL to obtain a second slurry liquid.

In the fourth leaching and separation step, oxygen is introduced into the second slurry for oxygen pressure leaching, wherein the temperature of oxygen pressure leaching is 185°C, the oxygen partial pressure is 0.8 MPa, the oxygen pressure leaching time is 6 hours, and pressure filtration is performed after the oxygen pressure leaching is completed. The fourth filtrate obtained contains nickel sulfate and cobalt sulfate, and the fourth filter residue contains metals such as Ca, which can be directly sold as a by-product of building materials.

See Table 1 for the ingredients of each raw material in the embodiments of the present disclosure.

Table 1 Composition of raw materials

It should be noted that the components of battery recyclate, low-grade nickel matte and oxygen-pressed slag used in the embodiments of the present disclosure are merely exemplary and should not be understood as a limitation on the concept of the present disclosure, i.e., recycling valuable metals such as lithium, nickel and cobalt in battery recyclate, low-grade nickel matte and oxygen-pressed slag.

Example 2

The present embodiment is similar to the embodiment 1 in that: battery recycling, low-grade nickel matte and oxygen-pressed slag, and the difference lies in other process parameters, specifically including the following steps:

In the calcination process, 200 g of battery recycling and 30 g of wood activated carbon were mixed evenly and loaded into a crucible. The mixture was calcined at 800 °C for 2 h under inert gas protection to obtain a calcined product at a heating rate of 10 °C/min.

In the first leaching and separation step, the roasted product is added into pure water at a liquid-solid ratio of 2 ml:1 g, concentrated sulfuric acid is added to adjust the pH of the initial solution to 6.5±0.1, and the reaction is stirred at 60° C. for 2 hours. During the process, the pH is monitored in real time and acid is supplemented in time to ensure that the pH of the system is maintained at 6.5±0.1 for the first leaching treatment; after the leaching is completed, a lithium-containing leaching solution and a leaching residue are obtained by filtration; wherein, the Li leaching rate in the leaching solution is 97.36%, the Ni leaching rate is 3.49%, the Co leaching rate is 0.72%, and the Mn leaching rate is 0.81%.

In the drying process, the leached slag, low nickel matte and oxygen pressed slag are dried separately, and air is introduced during the drying process. The drying temperature is 150°C, the drying time is 2h, and the moisture content of the dried material is 0.25%.

In the calcination process, the dried leached slag is mixed with a slag-forming agent in a mass ratio of 10:1 and then calcined. The slag-forming agent is silicon dioxide, bentonite, and calcium oxide. The calcination temperature is 800°C and the calcination time is 2h.

In the matte smelting process, the calcined product, low-grade nickel matte, oxygen-pressed slag and liquid sulfur are mixed in a mass ratio of 2:3:2:4 and then smelted. The matte smelting temperature is 1300°C and the matte smelting time is 2 hours to obtain a matte containing nickel and cobalt. The Ni recovery rate is calculated to be 99.12% and the Co recovery rate is 99.29%.

The method for recovering valuable metals provided in this embodiment also includes:

In the first slurrying step, the obtained matte is ball-milled and then sieved through 300 mesh. The sieved matte is mixed with water at a solid-liquid ratio of 1g:6mL for the first slurrying treatment to obtain the first slurry liquid, wherein the temperature of the first slurrying treatment is 60°C and the time of the first slurrying treatment is 1.5h.

In the second leaching and separation step, concentrated sulfuric acid is added to the first slurry for the second leaching treatment. The amount of sulfuric acid added is 1.2 of the theoretical amount. times, the temperature of the second leaching treatment is 70° C., and the time of the second leaching treatment is 6 hours; after the second leaching treatment, filtering is performed to obtain a second filtrate and a second filter residue.

In the extraction step, the second filtrate is mixed with the P507 extractant and then subjected to extraction treatment to obtain an extract and a raffinate, wherein the raffinate is a nickel sulfate and cobalt sulfate solution.

The third leaching and separation step is to add concentrated sulfuric acid to the second filter residue for a third leaching treatment, the amount of sulfuric acid added is 1.2 times the theoretical amount, the temperature of the third leaching treatment is 60°C, and the time of the third leaching treatment is 2 hours; after the third leaching treatment, filtering is performed to obtain a third filtrate and a third filter residue, wherein the third filtrate contains nickel sulfate and cobalt sulfate.

In the second slurrying step, the third filter residue and water are subjected to a second slurrying treatment at a solid-liquid ratio of 1 g:8 mL to obtain a second slurry liquid.

In the fourth leaching and separation step, oxygen is introduced into the second slurry liquid for oxygen pressure leaching treatment, wherein the temperature of the oxygen pressure leaching is 185° C., the oxygen partial pressure is 0.85 MPa, and the oxygen pressure leaching time is 4 hours. After the oxygen pressure leaching is completed, filter pressing is performed to obtain a fourth filtrate containing nickel sulfate and cobalt sulfate and a fourth filter residue containing metals such as Ca.

Example 3

The present embodiment is similar to the embodiment 1 in that: battery recycling, low-grade nickel matte and oxygen-pressed slag, and the difference lies in other process parameters, specifically including the following steps:

In the roasting process, 200 g of battery recycling, 180 g of ammonium sulfate and concentrated sulfuric acid were mixed evenly and loaded into a crucible. The mixture was roasted at 750°C for 2 h under inert gas protection to obtain a roasted product at a heating rate of 5°C/min.

In the first leaching and separation step, the roasted product is added into pure water at a liquid-solid ratio of 2ml:1g, concentrated sulfuric acid is added to adjust the pH of the initial solution to 6.5±0.1, and the reaction is stirred at 60°C for 2h. During the process, the pH is monitored in real time and acid is supplemented in time to ensure that the pH of the system is maintained at 6.5±0.1 for the first leaching treatment; after the leaching is completed, a lithium-containing leaching solution and a leaching residue are obtained by filtration; wherein, the Li leaching rate in the leaching solution is 99.28%, the Ni leaching rate is 3.86%, the Co leaching rate is 1.2%, and the Mn leaching rate is 0.36%.

In the drying process, the leached slag, low nickel matte and oxygen pressed slag are dried separately, and air is introduced during the drying process. The drying temperature is 200°C, the drying time is 1.5 hours, and the moisture content of the dried material is 0.15%.

In the calcination process, the dried leached slag is mixed with a slag-forming agent in a mass ratio of 12:1 and then calcined. The slag-forming agent is a mixture of silicon dioxide and calcium oxide. The calcination temperature is 900°C and the calcination time is 3 hours.

In the matte smelting process, the calcined product, low-grade nickel matte, oxygen-pressed slag and liquid sulfur are mixed in a mass ratio of 10:2:2:5 and then smelted. The matte smelting temperature is 1250°C and the matte smelting time is 2 hours to obtain a matte containing nickel and cobalt. Calculation shows that the Ni recovery rate is 99.14% and the Co recovery rate is 98.99%.

The method for recovering valuable metals provided in this embodiment also includes:

In the first slurrying step, the obtained matte is ball-milled and then sieved through 300 mesh. The sieved matte is mixed with water at a solid-liquid ratio of 1g:7mL for the first slurrying treatment to obtain the first slurry liquid, wherein the temperature of the first slurrying treatment is 70°C and the time of the first slurrying treatment is 2h.

In the second leaching and separation step, concentrated sulfuric acid is added to the first slurry for a second leaching treatment. The amount of sulfuric acid added is 1.15 of the theoretical amount. times, the temperature of the second leaching treatment is 60° C., and the time of the second leaching treatment is 5 hours; after the second leaching treatment, filtering is performed to obtain a second filtrate and a second filter residue.

In the extraction step, the second filtrate is mixed with the P507 extractant and then subjected to extraction treatment to obtain an extract and a raffinate, wherein the raffinate is a nickel sulfate and cobalt sulfate solution.

The third leaching and separation step is to add concentrated sulfuric acid to the second filter residue for a third leaching treatment, the amount of sulfuric acid added is 1.3 times the theoretical amount, the temperature of the third leaching treatment is 60°C, and the time of the third leaching treatment is 2 hours; after the third leaching treatment, filtering is performed to obtain a third filtrate and a third filter residue, wherein the third filtrate contains nickel sulfate and cobalt sulfate.

In the second slurrying step, the third filter residue and water are subjected to a second slurrying treatment at a solid-liquid ratio of 1 g:12 mL to obtain a second slurry liquid.

In the fourth leaching and separation step, oxygen is introduced into the second slurry liquid for oxygen pressure leaching treatment, wherein the temperature of the oxygen pressure leaching is 185° C., the oxygen partial pressure is 1 MPa, and the oxygen pressure leaching time is 3 hours. After the oxygen pressure leaching is completed, filter pressing is performed to obtain a fourth filtrate containing nickel sulfate and cobalt sulfate and a fourth filter residue containing metals such as Ca.

Example 4

The present embodiment is similar to the embodiment 1 in that: battery recycling, low-grade nickel matte and oxygen-pressed slag, and the difference lies in other process parameters, specifically including the following steps:

In the roasting process, 200 g of battery recycling and 300 g of concentrated sulfuric acid were mixed evenly and loaded into a crucible. The mixture was roasted at 1000° C. for 5 h under inert gas protection to obtain a roasted product at a heating rate of 8° C./min.

In the first leaching and separation step, the roasted product is added into pure water at a liquid-solid ratio of 10 ml: 1 g, concentrated sulfuric acid is added to adjust the pH of the initial solution to 7±0.1, and the reaction is stirred at 90° C. for 4 hours. During the process, the pH is monitored in real time and acid is added in time to ensure that the pH of the system is maintained at 7±0.1 for the first leaching treatment; after the leaching is completed, a lithium-containing leaching solution and a leaching residue are obtained by filtration; wherein, the Li leaching rate in the leaching solution is 99.89%, the Ni leaching rate is 4.50%, the Co leaching rate is 2.1%, and the Mn leaching rate is 1.34%.

In the drying process, the leached residue, low nickel matte and oxygen pressed residue are dried separately. Oxygen is introduced during the drying process. The drying temperature is 300°C, the drying time is 0.5h, and the moisture content of the dried material is 0.05%.

In the calcination process, the dried leached slag is mixed with a slag-forming agent in a mass ratio of 20:1 and then calcined. The slag-forming agent is a mixture of silicon dioxide, calcium oxide and aluminum oxide. The calcination temperature is 1000°C and the calcination time is 8 hours.

In the matte smelting process, the calcined product, low-grade nickel matte, oxygen-pressed slag and sulfide ore are mixed in a mass ratio of 10:10:7:6 and then smelted. The matte smelting temperature is 1500°C and the matte smelting time is 5 hours to obtain a matte containing nickel and cobalt. The Ni recovery rate is calculated to be 98.22% and the Co recovery rate is 98.89%.

The method for recovering valuable metals provided in this embodiment also includes:

In the first slurrying step, the obtained matte is ball-milled and then sieved through 300 mesh. The sieved matte is mixed with water at a solid-liquid ratio of 1g:15mL for the first slurrying treatment to obtain the first slurry liquid, wherein the temperature of the first slurrying treatment is 90°C and the time of the first slurrying treatment is 5h.

In the second leaching and separation step, concentrated sulfuric acid is added to the first slurry for the second leaching treatment. The amount of sulfuric acid added is 1.6 of the theoretical amount. times, the temperature of the second leaching treatment is 100° C., and the time of the second leaching treatment is 4 hours; after the second leaching treatment, filtering is performed to obtain a second filtrate and a second filter residue.

In the extraction step, the second filtrate is mixed with the P507 extractant and then subjected to extraction treatment to obtain an extract and a raffinate, wherein the raffinate is a nickel sulfate and cobalt sulfate solution.

The third leaching and separation step is to add concentrated sulfuric acid to the second filter residue for a third leaching treatment, the amount of sulfuric acid added is 1.3 times the theoretical amount, the temperature of the third leaching treatment is 100°C, and the time of the third leaching treatment is 2h; after the third leaching treatment, filtering is performed to obtain a third filtrate and a third filter residue, wherein the third filtrate contains nickel sulfate and cobalt sulfate.

In the second slurrying step, the third filter residue and water are subjected to a second slurrying treatment at a solid-liquid ratio of 1 g:15 mL to obtain a second slurry liquid.

In the fourth leaching and separation step, oxygen is introduced into the second slurry liquid for oxygen pressure leaching treatment, wherein the temperature of the oxygen pressure leaching is 220° C., the oxygen partial pressure is 1 MPa, and the oxygen pressure leaching time is 8 hours. After the oxygen pressure leaching is completed, filter pressing is performed to obtain a fourth filtrate containing nickel sulfate and cobalt sulfate and a fourth filter residue containing metals such as Ca.

Example 5

The present embodiment is similar to the embodiment 1 in that: battery recycling, low-grade nickel matte and oxygen-pressed slag, and the difference lies in other process parameters, specifically including the following steps:

In the calcination process, 200 g of battery recycling material and 200 g of concentrated sulfuric acid were mixed evenly and loaded into a crucible. The mixture was calcined at 750°C for 2 h under inert gas protection to obtain a calcined product at a heating rate of 7°C/min.

In the first leaching and separation step, the roasted product is added into pure water at a liquid-solid ratio of 5 ml:1 g, concentrated sulfuric acid is added to adjust the pH of the initial solution to 6.5±0.1, and the reaction is stirred at 80° C. for 2 hours. During the process, the pH is monitored in real time and acid is supplemented in time to ensure that the pH of the system is maintained at 6.5±0.1 for the first leaching treatment; after the leaching is completed, a lithium-containing leaching solution and a leaching residue are obtained by filtration; wherein the Li leaching rate in the leaching solution is 99.78%, the Ni leaching rate is 2.50%, the Co leaching rate is 1.69%, and the Mn leaching rate is 0.65%.

In the drying process, the leached slag, low nickel matte and oxygen pressed slag are dried separately. Compressed air is introduced during the drying process. The drying temperature is 200°C, the drying time is 2h, and the moisture content of the dried material is 0.1%.

In the calcination process, the dried leached slag is mixed with a slag-forming agent in a mass ratio of 5:1 and then calcined. The slag-forming agent is a mixture of silicon dioxide, calcium oxide and bentonite. The calcination temperature is 750°C and the calcination time is 3 hours.

In the matte smelting process, the calcined product, low-grade nickel matte, oxygen-pressed slag and sulfide ore are mixed in a mass ratio of 10:20:20:25 and then smelted at a temperature of 1250°C for 3 hours to obtain a matte containing nickel and cobalt. Calculated Ni recovery rates are 99.85% and Co recovery rates are 99.56%.

The method for recovering valuable metals provided in this embodiment also includes:

In the first slurrying step, the obtained matte is ball-milled and then sieved through 300 mesh. The sieved matte is mixed with water at a solid-liquid ratio of 1g:5mL for the first slurrying treatment to obtain the first slurry liquid, wherein the temperature of the first slurrying treatment is 90°C and the time of the first slurrying treatment is 2h.

In the second leaching and separation step, concentrated sulfuric acid is added to the first slurry for the second leaching treatment. The amount of sulfuric acid added is 1.2 of the theoretical amount. times, the temperature of the second leaching treatment is 80° C., and the time of the second leaching treatment is 8 hours; after the second leaching treatment, filtering is performed to obtain a second filtrate and a second filter residue.

In the extraction step, the second filtrate is mixed with the P507 extractant and then subjected to extraction treatment to obtain an extract and a raffinate, wherein the raffinate is a nickel sulfate and cobalt sulfate solution.

The third leaching and separation step is to add concentrated sulfuric acid to the second filter residue for a third leaching treatment, the amount of sulfuric acid added is 1.2 times the theoretical amount, the temperature of the third leaching treatment is 80°C, and the time of the third leaching treatment is 2 hours; after the third leaching treatment, filtering is performed to obtain a third filtrate and a third filter residue, wherein the third filtrate contains nickel sulfate and cobalt sulfate.

In the second slurrying step, the third filter residue and water are subjected to a second slurrying treatment at a solid-liquid ratio of 1 g:5 mL to obtain a second slurry liquid.

In the fourth leaching and separation step, oxygen is introduced into the second slurry liquid for oxygen pressure leaching treatment, wherein the temperature of the oxygen pressure leaching is 200° C., the oxygen partial pressure is 0.9 MPa, and the oxygen pressure leaching time is 8 hours. After the oxygen pressure leaching is completed, filter pressing is performed to obtain a fourth filtrate containing nickel sulfate and cobalt sulfate and a fourth filter residue containing metals such as Ca.

Example 6

The present embodiment is similar to the embodiment 1 in that: battery recycling, low-grade nickel matte and oxygen-pressed slag, and the difference lies in other process parameters, specifically including the following steps:

In the calcination process, 200 g of battery recycling material and 300 g of concentrated sulfuric acid were mixed evenly and loaded into a crucible. The mixture was calcined at 600° C. for 2 h under inert gas protection to obtain a calcined product at a heating rate of 6° C./min.

In the first leaching and separation step, the roasted product is added into pure water at a liquid-solid ratio of 5 ml:1 g, concentrated sulfuric acid is added to adjust the pH of the initial solution to 5±0.1, and the reaction is stirred at 80° C. for 1 hour. During the process, the pH is monitored in real time and acid is supplemented in time to ensure that the pH of the system is maintained at 5±0.1 for the first leaching treatment; after the leaching is completed, a lithium-containing leaching solution and a leaching residue are obtained by filtration; wherein the Li leaching rate in the leaching solution is 99.18%, the Ni leaching rate is 3.50%, the Co leaching rate is 3.25%, and the Mn leaching rate is 4.78%.

In the drying process, the leached slag, low nickel matte and oxygen pressed slag are dried separately, oxygen is introduced during the drying process, the drying temperature is 150°C, the drying time is 3h, and the moisture content of the dried material is 0.13%.

In the calcination process, the dried leached slag is mixed with a slag-forming agent in a mass ratio of 7:1 and then calcined. The slag-forming agent is a mixture of silicon dioxide, aluminum oxide, calcium oxide and bentonite. The calcination temperature is 750°C and the calcination time is 3 hours.

In the matte smelting process, the calcined product, low-grade nickel matte, oxygen-pressed slag and liquid sulfur are mixed in a mass ratio of 10:30:25:30 and then smelted at a temperature of 1250°C for 8 hours to obtain a matte containing nickel and cobalt. Calculation shows that the Ni recovery rate is 98.23% and the Co recovery rate is 98.65%.

The method for recovering valuable metals provided in this embodiment also includes:

In the first slurrying step, the obtained matte is ball-milled and then sieved through 300 mesh. The sieved matte is mixed with water at a solid-liquid ratio of 1g:10mL for the first slurrying treatment to obtain the first slurry liquid, wherein the temperature of the first slurrying treatment is 60°C and the time of the first slurrying treatment is 2h.

In the second leaching and separation step, concentrated sulfuric acid is added to the first slurry for the second leaching treatment. The amount of sulfuric acid added is 1.2 of the theoretical amount. times, the temperature of the second leaching treatment is 80° C., and the time of the second leaching treatment is 2 h; after the second leaching treatment, filtering is performed to obtain a second filtrate and a second filter residue.

In the extraction step, the second filtrate is mixed with the P507 extractant and then subjected to extraction treatment to obtain an extract and a raffinate, wherein the raffinate is a nickel sulfate and cobalt sulfate solution.

The third leaching and separation step is to add concentrated sulfuric acid to the second filter residue for a third leaching treatment, the amount of sulfuric acid added is 1.2 times the theoretical amount, the temperature of the third leaching treatment is 80°C, and the time of the third leaching treatment is 2 hours; after the third leaching treatment, filtering is performed to obtain a third filtrate and a third filter residue, wherein the third filtrate contains nickel sulfate and cobalt sulfate.

In the second slurrying step, the third filter residue and water are subjected to a second slurrying treatment at a solid-liquid ratio of 1 g:7 mL to obtain a second slurry liquid.

In the fourth leaching and separation step, oxygen is introduced into the second slurry liquid for oxygen pressure leaching treatment, wherein the temperature of the oxygen pressure leaching is 200° C., the oxygen partial pressure is 1.2 MPa, and the oxygen pressure leaching time is 8 hours. After the oxygen pressure leaching is completed, filter pressing is performed to obtain a fourth filtrate containing nickel sulfate and cobalt sulfate and a fourth filter residue containing metals such as Ca.

Comparative Example 1

On the basis of Example 1, oxygen was additionally introduced only during the matte smelting process, with an oxygen flow rate of 50 Nm ³ /h. The nickel yield was 98.02%, and the Co yield was 98.67%.

Comparative Example 2

On the basis of Example 1, only the matte smelting process was changed, and oxygen was not added to slag pressing during this process. The nickel yield was 97.28% and the Co yield was 97.99%.

Comparative Example 3

On the basis of Example 1, only the matte smelting process was changed. In this process, oxygen was not added to slag pressing and oxygen was additionally introduced. The oxygen flow rate was 50 Nm ³ /h. The nickel yield was 98.12% and the Co yield was 98.93%.

Test Section

(1) Valuable metal leaching rate test

Take 5 mL of the leaching solution, add 45 mL of concentrated sulfuric acid to dilute it, and obtain the test sample solution;

Take 10mL of sample solution and put it into Inductive Coupled Plasma Emission Spectrometer (ICP) for testing. The concentration of each valuable metal in the sample solution is obtained through testing. The concentration of the valuable metal in the leaching solution is calculated based on the concentration, and then the leaching rate of the valuable metal (such as Li, Ni, Co, Mn, etc.) is calculated according to the following formula:

Leaching rate = (volume of leachate * concentration of valuable metals in leachate) / (mass of battery recyclate * mass content of valuable metals in battery recyclate).

(2) Valuable metal recovery test

Using strong acid to leach valuable metals in the sulfur phase material to obtain a leachate;

Take 5 mL of the leaching solution, add 45 mL of concentrated sulfuric acid to dilute it, and obtain the test sample solution;

Take 10mL of sample solution and put it into ICP for testing. The concentration of each valuable metal in the sample solution is obtained through testing. The total amount of valuable metals in the leaching solution is calculated according to the following formula to obtain the recovery rate of valuable metals (such as Ni and Co):

The mass fraction of Ni/Co in matte = (Ni/Co concentration of matte leaching solution * leachate volume) / mass of matte material (Note: matte leaching solution is obtained by completely dissolving matte material with concentrated sulfuric acid.)

Ni/Co recovery rate = (mass of matte material * mass fraction of Ni/Co in matte)/((mass content of Ni/Co in nickel matte * mass of nickel matte) + (mass content of Ni/Co in oxygen-pressed slag * mass of oxygen-pressed slag) + (mass content of Ni/Co in battery recyclate after lithium extraction * mass of battery recyclate after lithium extraction)).

Table 2 Recovery rate of valuable metals in each embodiment and comparative example

Referring to Table 2, it can be seen from Examples 1 to 6 that by adopting the recycling method provided by the present invention, the recovery rate of lithium in battery recyclate can be as high as over 92%, which is significantly improved compared to the traditional wet lithium extraction rate of only about 85%. At the same time, the recovery rates of nickel and cobalt are also over 92%, which can achieve the purpose of recovering valuable metals such as lithium, nickel and cobalt from battery recyclate, nickel matte and oxygen-pressed slag, and the recovery rate of valuable metals is high and the recovery cost is relatively low.

Comparing the experimental data of Comparative Example 1 with that of Example 1, although additional oxygen was introduced during the matte smelting process in Comparative Example 1, the recovery rates of nickel and cobalt were not improved. Therefore, it can be seen that there is no need to additionally introduce oxygen during the matte smelting process.

Comparing the experimental data of Comparative Example 2 with that of Example 1, since no oxygen slag was added to the matte smelting process in Comparative Example 2, the recovery rate of nickel and cobalt was significantly reduced. Combining the experimental data of Comparative Example 3 with that of Example 1, it can be seen that in the matte smelting process in Comparative Example 3, no oxygen slag was added, and additional oxygen was introduced. The recovery rate of nickel and cobalt in Comparative Example 3 was still lower than that in Example 1. It can be seen that in the matte smelting process, since oxygen slag was added, no additional oxygen was needed, which can improve the utilization value of oxygen slag to reduce the cost of valuable metal recovery, and can also improve the recovery rate of valuable metal nickel and cobalt, reduce the safety risks caused by the introduction of a large amount of oxygen from the outside, and reduce the requirements for fire protection and recovery equipment in the factory.

The various technical features described above can be combined arbitrarily. Although all possible combinations of these technical features are not described, any combination of these technical features should be considered to be covered by this specification as long as there is no contradiction in such combination.

Finally, it should be noted that the above embodiments are only used to illustrate the technical solutions of the present disclosure, rather than to limit them. Although the present disclosure has been described in detail with reference to the aforementioned embodiments, those skilled in the art should understand that they can still modify the technical solutions described in the aforementioned embodiments, or replace some or all of the technical features therein with equivalents. However, these modifications or replacements do not make the essence of the corresponding technical solutions deviate from the scope of the technical solutions of the embodiments of the present disclosure.

Industrial Applicability

The recovery method provided by the present disclosure has a simple process flow, a high recovery rate, and generates few pollutants, thereby reducing the recovery cost of valuable metals and being environmentally friendly.

Claims (10)

A method for recovering valuable metals, the method comprising: A roasting process is to roast the battery recycling material and the lithium extraction agent under a protective atmosphere to obtain a roasting product; A first leaching and separation step, performing a first leaching and separation on the roasted product to obtain a lithium-containing leaching solution and leaching residue; a calcination step, mixing the leached slag and the slag-making agent and then calcining to obtain a calcined product; In the matte making and smelting process, the calcined product, nickel matte, oxygen pressed slag and matte making agent are mixed and then matte making and smelting is performed to obtain nickel-containing matte. The recovery method according to claim 1, wherein in the roasting step, the lithium extraction agent includes one or more of a reducing agent, a sulfate or concentrated sulfuric acid; Optionally, the reducing agent includes one or more of activated carbon, hydrogen, lignite, anthracite, and carbon monoxide; Optionally, the sulfate includes one or more of ammonium sulfate and sodium sulfite; Optionally, the mass ratio of the battery recyclate to the lithium extraction agent is 20:(1-30); Optionally, the calcination temperature is 300°C-1000°C, the heating rate is 2°C/min-10°C/min, and the calcination time is 0.5h-5h. The recovery method according to claim 1 or 2, wherein, in the first leaching and separation step, it comprises: a first leaching step, mixing the roasted product and a first leaching agent to perform a first leaching treatment, so that the lithium in the roasted product is dissolved into a solution to obtain a first leaching slurry containing lithium; A first separation step, performing solid-liquid separation on the first leaching slurry to obtain the lithium-containing leaching solution and the leaching residue; Optionally, in the first leaching step, the pH of the first leaching treatment is 3-7; Optionally, in the first leaching process, the temperature of the first leaching treatment is 30°C-90°C, and the time of the first leaching treatment is 0.5h-4h. The recovery method according to any one of claims 1 to 3, wherein in the calcination step, the mass ratio of the leached slag to the slag-forming agent is (2-20):1; Optionally, the calcination temperature is 600°C-1000°C, and the calcination time is 2h-8h; Optionally, the slag-forming agent includes one or more of silicon dioxide, diatomaceous earth, bentonite, calcium oxide, and aluminum oxide. The recovery method according to any one of claims 1 to 4, wherein, in the matte smelting process, the mass ratio of the calcined product, the nickel matte, the oxygen pressed slag and the matte making agent is 10:(1-30):(1-25):(0.01-30); Optionally, the matte-forming agent includes a sulfur-containing element and/or a sulfur-containing compound; Optionally, the temperature of the matte smelting is 1000°C-1500°C, and the time of the matte smelting is 1h-10h. The recovery method according to any one of claims 1 to 5, wherein after the first leaching and separation step, it further comprises: a drying step of drying the leached residue, nickel matte and oxygen-pressed slag respectively; Optionally, the drying temperature is 100° C.-300° C., the drying time is 0.5 h-4 h, and air and/or oxygen are introduced during the drying process. The recycling method according to any one of claims 1 to 6, further comprising: A first slurrying step, mixing the matte with a solvent and then performing a first slurrying treatment to obtain a first slurry liquid; a second leaching and separation step, mixing the first slurry and the second leaching agent and performing a second leaching treatment to dissolve the nickel in the first slurry into the solution to obtain a second leaching slurry containing nickel, and performing a solid-liquid separation treatment on the second leaching slurry to obtain a second filtrate and a second filter residue; In the extraction step, the second filtrate is mixed with an extractant and then subjected to extraction treatment to obtain an extract and a raffinate. The recovery method according to claim 7, wherein in the first slurrying step, the matte is ground, and the ground matte is mixed with the solvent at a solid-liquid ratio of 1 g: (2-15) mL to perform the first slurrying treatment to obtain the first slurry liquid; Optionally, the particle size of the matte after grinding is less than or equal to 48 μm; Optionally, the temperature of the first slurry treatment is 30°C-90°C, and the time of the first slurry treatment is 0.5h-5h; Optionally, in the second leaching and separation step, the second leaching agent includes concentrated sulfuric acid, wherein the amount of the concentrated sulfuric acid added is 1 to 1.6 times the theoretical amount required to completely leach the nickel in the first slurry; Optionally, the temperature of the second leaching treatment is 50°C-100°C, and the time of the second leaching treatment is 2h-8h. The recycling method according to claim 7, further comprising: a third leaching and separation step, mixing the second filter residue and a third leaching agent, performing a third leaching treatment to obtain a third leaching slurry, and performing a solid-liquid separation treatment on the third leaching slurry to obtain a third filtrate and a third filter residue; A second slurrying step, mixing the third filter residue and the solvent and then performing a second slurrying treatment to obtain a second slurry liquid; The fourth leaching and separation step is to introduce oxygen into the second slurry to perform oxygen pressure leaching treatment to obtain a fourth leaching slurry, and to perform solid-liquid separation treatment on the fourth leaching slurry to obtain a fourth filtrate and a fourth filter residue. The recovery method according to claim 9, wherein, in the third leaching and separation step, the third leaching agent comprises concentrated sulfuric acid, wherein the amount of concentrated sulfuric acid added is 1 to 1.3 times the theoretical amount required to leach all the nickel in the second filter residue; Optionally, in the third leaching and separation step, the temperature of the third leaching treatment is 40° C.-100° C., and the time of the third leaching treatment is 0.5 h-4 h; Optionally, in the second slurrying step, the third filter residue and the solvent are in a solid-liquid ratio of 1 g: (3-15) ml; Optionally, in the fourth leaching and separation step, the oxygen partial pressure of the oxygen pressure leaching is 0.6MPa-1.2MPa; Optionally, in the fourth leaching and separation step, the temperature of the oxygen pressure leaching is 150° C.-220° C., and the time of the oxygen pressure leaching is 2 h-8 h.