Method for recovering valuable elements from waste battery material fumes

Abstract

The application discloses a method for recovering valuable elements from waste battery material smoke dust, comprising the following steps: uniformly mixing waste battery material smoke dust with concentrated sulfuric acid to obtain a mixture; performing roasting treatment on the mixture to obtain roasted powder; performing water or dilute acid leaching on the roasted powder, and then performing filtration to obtain a leaching solution and a leaching residue; performing extraction on the leaching solution to obtain a sulfate solution containing valuable metal elements; and performing water washing and drying on the leaching residue to obtain crude graphite. The method can improve the efficient leaching rate of valuable metals such as lithium, nickel, cobalt and manganese in the waste battery material smoke dust, can avoid the consumption of fluorine removal agents, can realize low acid consumption and low residual acid amount, can further reduce alkali consumption, and can realize the recovery of carbon elements in the waste battery material smoke dust.

Description

Technical Field

[0001] This application belongs to the field of resource recycling technology, specifically relating to a method for recovering valuable elements from waste battery material dust. Background Technology

[0002] With the rapid development of the new energy industry, the consumption of lithium-ion batteries has surged. However, the average lifespan of lithium-ion batteries is only 5-8 years, resulting in a large volume of waste batteries in the future. As the lithium battery market becomes increasingly segmented, battery types are becoming more diverse. Currently, there are two main recycling processes for waste lithium-ion batteries: wet and pyrometallurgical processes. Pyrometallurgical processes are suitable for processing lithium battery debris from various sources, and they offer high overall recovery rates and high production efficiency, making them a promising area for battery recycling. However, pyrometallurgical processes inevitably generate flue gas. Some of this flue gas is recycled and reused, while some, due to the enrichment of impurities, needs to be collected and processed separately. The flue gas also contains high levels of valuable metals such as lithium, nickel, and cobalt, necessitating effective technologies to recover these valuable elements.

[0003] The current process for treating flue dust mainly involves two-stage sulfuric acid leaching, pH adjustment to remove impurities such as iron and aluminum, fluoride removal, and extraction to separate nickel, cobalt, manganese, and lithium. However, this method has a low recovery rate of valuable metals, and the leaching residue still contains a significant amount of nickel and cobalt. It is difficult to achieve efficient leaching of flue dust and other impurities. Furthermore, the leachate contains high levels of fluoride impurities and residual acid, and there are also problems such as high consumption of defluorinating agents, acid, and alkali. Summary of the Invention

[0004] The purpose of this application is to overcome the shortcomings of the prior art and provide a method for recovering valuable elements from waste battery material dust, so as to solve the technical problems of low recovery rate of valuable metal elements, large consumption of defluorinating agent, and high consumption of acid and alkali in existing waste battery material dust.

[0005] To achieve the aforementioned objectives, this application provides a method for recovering valuable elements from waste battery material dust, comprising the following steps:

[0006] The waste battery material dust is mixed evenly with concentrated sulfuric acid to obtain a mixture.

[0007] The mixture is roasted to obtain roasted powder.

[0008] The roasted powder is extracted with water or dilute acid, filtered, and then a leachate and a leachate residue are obtained.

[0009] The leachate is extracted to obtain a sulfate solution containing valence metal elements. The leachate residue is washed with water and dried to obtain crude graphite.

[0010] Optionally, the concentrated sulfuric acid has a mass concentration greater than 80%; and / or the dilute acid has a mass concentration less than 20%.

[0011] Optionally, the mass ratio of the waste battery material dust to the concentrated sulfuric acid in the mixture is 1:(0.8-1.2).

[0012] Optionally, the calcination temperature is 200–300°C.

[0013] Optionally, the roasting process takes 1 to 2 hours.

[0014] Optionally, the liquid-to-solid ratio of the extraction is (3-5):1.

[0015] Optionally, the extraction temperature is 75–85°C.

[0016] Optionally, the extraction time is 1 to 2 hours.

[0017] Optionally, before performing the extraction on the leachate, a step of removing impurities from the leachate is further included.

[0018] Optionally, the waste battery material dust contains the following components: Li 0.1–3.0 wt%, Ni 5.0–40 wt%, Co 3.0–15 wt%, Mn 3.0–15 wt%, F 0.2–3.0 wt%, Fe 0.5–2.0 wt%, Al 2.0–8.0 wt%, Cu 0.5–3.0 wt%, and C 10–40 wt%.

[0019] Compared with the prior art, this application has the following technical effects:

[0020] This application discloses a method for recovering valuable elements from waste battery material dust. The method involves leaching the dust after co-firing it with concentrated sulfuric acid. Firstly, concentrated sulfuric acid is more reactive; high-temperature firing can disrupt the ternary material structure in the waste battery material dust, allowing more valuable metals to participate in the high-temperature firing reaction. This, in turn, improves the leaching efficiency of valuable metals such as lithium, nickel, cobalt, and manganese in the subsequent extraction. Secondly, this application removes fluoride through high-temperature acidification firing, where sulfuric acid and fluorides generate hydrofluoric acid. The removal of acid gases prevents fluorine from entering the leachate, thus eliminating the need for defluorinating agents. Furthermore, in the high-temperature acid roasting process of this application, the waste battery material dust and concentrated sulfuric acid are pre-mixed evenly, and the water produced by the reaction is released in the form of water vapor. Therefore, there is no acidity decay, which can achieve low acid consumption and low residual acid, thereby reducing alkali consumption. Moreover, the leaching residue of this application can have a graphite content of over 96%, which can be directly recycled as crude graphite, realizing the recovery of carbon elements in waste battery material dust. Attached Figure Description

[0021] To more clearly illustrate the technical solutions in the specific embodiments of this application or the prior art, the drawings used in the description of the specific embodiments or the prior art will be briefly introduced below. Obviously, the drawings described below are some embodiments of this application. For those skilled in the art, other drawings can be obtained from these drawings without creative effort.

[0022] Figure 1 This is a process flow diagram of a method for recovering valuable elements from waste battery material dust according to an embodiment of this application. Detailed Implementation

[0023] To make the technical problems, technical solutions, and beneficial effects of this application clearer, the following detailed description is provided in conjunction with embodiments. It should be understood that the specific embodiments described herein are merely illustrative and not intended to limit the scope of this application.

[0024] In this application, the term "and / or" describes the relationship between related objects, indicating that three relationships can exist. For example, A and / or B can represent: A existing alone, A and B existing simultaneously, or B existing alone. A and B can be singular or plural. The character " / " generally indicates that the preceding and following related objects have an "or" relationship.

[0025] In this application, "at least one" means one or more, and "more than one" means two or more. "At least one of the following" or similar expressions refer to any combination of these items, including any combination of single or multiple items. For example, "at least one of a, b, or c", or "at least one of a, b, and c", can both mean: a, b, c, ab (i.e., a and b), ac, bc, or abc, where a, b, and c can be single or multiple.

[0026] It should be understood that in the various embodiments of this application, the order of the above processes does not imply the order of execution. Some or all steps may be executed in parallel or sequentially. The execution order of each process should be determined by its function and internal logic, and should not constitute any limitation on the implementation process of the embodiments of this application.

[0027] The terminology used in the embodiments of this application is for the purpose of describing particular embodiments only and is not intended to be limiting of this application. The singular forms “a,” “the,” and “the” used in the embodiments of this application and the appended claims are also intended to include the plural forms unless the context clearly indicates otherwise.

[0028] The weights of the relevant components mentioned in the embodiments of this application can refer not only to the specific content of each component, but also to the proportional relationship between the weights of the components. Therefore, any scaling up or down of the content of the relevant components according to the embodiments of this application is within the scope disclosed in the embodiments of this application. Specifically, the mass described in the embodiments of this application can be a mass unit known in the chemical industry, such as μg, mg, g, or kg.

[0029] The terms "first" and "second" are used for descriptive purposes only, to distinguish objects, such as substances, from one another, and should not be construed as indicating or implying relative importance or implicitly specifying the number of technical features indicated. For example, without departing from the scope of the embodiments of this application, "first XX" may also be referred to as "second XX," and similarly, "second XX" may also be referred to as "first XX." Thus, features defined with "first" and "second" may explicitly or implicitly include one or more of that feature.

[0030] This application provides a method for recovering valuable elements from waste battery material dust, and its process flow diagram is shown below. Figure 1 As shown, it includes the following steps:

[0031] (1) Mix waste battery material dust with concentrated sulfuric acid to obtain a mixture;

[0032] (2) The mixture is roasted to obtain roasted powder;

[0033] (3) After roasting, the powder is extracted with water or dilute acid, filtered, and then the leachate and leachate residue are obtained.

[0034] (4) After extracting the leachate, a sulfate solution containing valence metal elements is obtained. After washing and drying the leachate residue, crude graphite is obtained.

[0035] In step (1) above, the waste battery material dust used in this application embodiment contains the following components: Li 0.1-3.0 wt%, Ni 5.0-40 wt%, Co 3.0-15 wt%, Mn 3.0-15 wt%, F 0.2-3.0 wt%, Fe 0.5-2.0 wt%, Al 2.0-8.0 wt%, Cu 0.5-3.0 wt%, and C 10-40 wt%. For example, in a specific embodiment of this application, the elemental composition of the waste battery material dust (black powder dust) is shown in Table 1 below.

[0036] Table 1

[0037]

[0038] Current processes for recovering valuable elements from waste battery material dust typically employ conventional acid leaching, either high-acid or low-acid leaching. This involves placing the waste battery material dust in sulfuric acid with a concentration of 40% or 20%. However, this conventional acid leaching method is insufficient for efficiently recovering metal elements from waste battery material dust, especially nickel. Even with multi-stage high-acid leaching, high recovery rates are difficult to achieve. This is because the nickel in the ternary materials within the waste battery material dust exists as high-valence nickel (+3 valence), which is difficult to extract using conventional acid leaching, even with hydrochloric acid. Furthermore, conventional acid leaching requires high temperatures to enhance reactivity.

[0039] Although oxygen-pressure leaching is a commonly used method, it requires operating temperatures >200°C and oxygen pressure >2MPa to leach residual difficult-to-leach metals under these stringent conditions. However, oxygen pressure vessels are not resistant to fluorine; even ppm-level fluorine can corrode and permeate the vessel, posing an explosion risk. Since waste battery material dust mostly contains elemental fluorine, oxygen-pressure leaching cannot further improve the leaching rate. Unlike existing processes, this application uses concentrated sulfuric acid mixed with waste battery material dust for a reaction. Concentrated sulfuric acid has a higher acid concentration and stronger reactivity, effectively destroying the structure of ternary materials and thus improving the leaching rate of metal elements. The concentrated sulfuric acid in this application has a mass concentration greater than 80%; in a specific embodiment, 98% concentrated sulfuric acid is used.

[0040] In step (1) above, the mass ratio of waste battery material dust to concentrated sulfuric acid in the mixture is 1:(0.8~1.2), which ensures efficient recovery of valuable elements from the waste battery material dust. If the amount of concentrated sulfuric acid added is too small, the reactivity between the waste battery material dust and concentrated sulfuric acid will weaken, leading to a decrease in the recovery rate of valuable elements; while if the amount of concentrated sulfuric acid added is too large, it will not further improve the recovery rate of valuable elements from the waste battery material dust, and will also lead to an increase in acid consumption. In a specific embodiment, the mass ratio of waste battery material dust to concentrated sulfuric acid in the mixture is 1:1 as an example for illustration.

[0041] In step (2) above, the uniformly mixed concentrated sulfuric acid and waste battery material dust undergo a high-temperature acidification reaction through roasting heating. Under high-temperature roasting conditions, the reaction rate between the metal elements and concentrated sulfuric acid is significantly enhanced, and the metal elements are converted into water-soluble metal sulfate salts after reacting with concentrated sulfuric acid. In addition, the fluorine elements in the waste battery material dust react with concentrated sulfuric acid to generate hydrofluoric acid gas for removal. In this way, the fluorine elements will not enter the subsequent leachate, eliminating the need to use defluorinating agents to remove fluorine. In the embodiments of this application, the roasting temperature of the roasting treatment can be selected as 200-300℃, specifically 220℃, 250℃, 280℃, etc. If the roasting temperature is too low, the reaction between the metal elements and concentrated sulfuric acid will be insufficient, affecting the efficient recovery of metal elements; if the roasting temperature is too high, the volatility of concentrated sulfuric acid increases above 300℃, increasing acid consumption and energy consumption. By reasonably controlling the roasting temperature, not only can valuable metal elements in waste battery material dust be efficiently recovered, but acid consumption and energy consumption can also be reduced. Furthermore, roasting time is also a crucial factor affecting the recovery rate of valuable metal elements. Too short a roasting time will result in insufficient reaction between the metal elements and concentrated sulfuric acid, while too long a roasting time will increase energy consumption. The reasonable roasting time in this embodiment is 1-2 hours. In a specific embodiment, after roasting the waste battery material fume at 220°C for 2 hours under the action of concentrated sulfuric acid, more than 95% of the fluorine can be removed, directly obtaining material with qualified fluorine content.

[0042] In the embodiments of this application, the roasting process can be carried out in a steel strip furnace or a rotary kiln. In a specific embodiment, the roasting process of this application is carried out in a steel strip furnace, and the thickness of the mixture in the steel strip furnace is 3-6 cm. If the thickness of the mixture is too thin, it will increase energy consumption and reduce output; while if the thickness of the mixture is too thick, it will cause incomplete combustion of the mixture and affect the recovery rate of valuable metal elements.

[0043] Furthermore, in conventional acid leaching processes, the amount of acid in the solution is consumed as the reaction proceeds, leading to insufficient leaching power in the later stages of the reaction. To ensure the leaching rate, excessive acid is generally required, and more alkali is needed for neutralization in the later stages, resulting in low acid conversion rate, high acid consumption, and high alkali consumption. In contrast, in the sulfation roasting process of this application, the waste battery material dust and concentrated sulfuric acid are pre-mixed evenly, and the water produced in the reaction is released as steam. Therefore, there is no acidity decay, achieving low acid consumption and low residual acid. Existing processes have an acid utilization rate of approximately 60% in the low-acid leaching stage, and the H+ in the leachate... + The concentration is approximately 2 mol / L. However, the high-acid leaching section involves multiple back-leaching processes. Because residual metals are difficult to leach, a large amount of excess acid is required to form a high-concentration acid, resulting in even lower acid utilization and necessitating the consumption of large amounts of alkali for neutralization. In the embodiments of this application, the acid utilization rate is approximately 80%, and the H in the leachate... + The concentration is about 1 mol / L, so less alkali is needed for subsequent neutralization, reducing alkali consumption.

[0044] In step (3) above, the calcined powder can be extracted with water or dilute acid. The water can be tap water or generated water, and the mass concentration of the dilute acid is less than 20%. The dilute acid can be dilute hydrochloric acid or dilute sulfuric acid, etc. In the specific embodiment of this application, dilute sulfuric acid is used as the dilute acid because the use of dilute sulfuric acid will not introduce additional impurity anions, which is beneficial to the subsequent separation operation of the system. Through the extraction treatment, the soluble metal sulfate salts in the calcined powder will dissolve in the water or dilute acid solution, and after filtration, a leachate and a leachate residue are formed. In the embodiment of this application, the liquid-solid ratio of the extraction treatment is (3-5):1. By controlling a reasonable liquid-solid ratio, soluble metal sulfate salts can be efficiently leached with reduced solvent usage. In the specific embodiment, the liquid-solid ratio of the extraction treatment is 4:1 as an example for explanation. The extraction temperature of the embodiment of this application can be selected as 75-85℃. At this leaching temperature, the leaching rate of metal sulfate salts can be accelerated. Furthermore, the leaching time in this embodiment can be selected as 1 to 2 hours. Within this leaching time, almost all of the metal sulfate salt can enter the leachate. Further increasing the leaching time will not further improve the leaching rate of the metal sulfate salt and will increase energy consumption. In this embodiment, the leaching rate of metallic nickel in the leachate can exceed 95%, and the leaching rate of lithium, cobalt, and manganese can exceed 98%.

[0045] The leaching residue of this application embodiment can contain more than 96% graphite, and the leaching residue can be directly recycled as crude graphite. However, the leaching residue of ordinary acid leaching recovery process contains up to 20% valuable metal elements and only about 70% carbon. Therefore, it is impossible to directly recover and utilize the carbon in the leaching residue.

[0046] In step (4) above, conventional extraction processes can be used to separate and efficiently recover different types of metal sulfates in the leachate, which will not be repeated in this embodiment. The leaching residue is washed with water and dried to obtain crude graphite. In order to reduce the extraction and recovery of valuable metals Li, Ni, Co and Mn by impurity elements in the leachate, the recovery method of this embodiment includes a step of removing impurities from the leachate before extraction. Specifically, the pH value of the leachate is adjusted to remove iron and aluminum elements from the leachate. The leachate enters the subsequent extraction and separation steps after iron and aluminum removal.

[0047] This application discloses a method for recovering valuable elements from waste battery material dust. The method involves leaching the waste battery material dust after co-firing it with concentrated sulfuric acid. Firstly, concentrated sulfuric acid is more reactive; high-temperature firing can disrupt the ternary material structure in the waste battery material dust, allowing more valuable metals to participate in the high-temperature firing reaction. This, in turn, improves the leaching efficiency of valuable metals such as lithium, nickel, cobalt, and manganese in the subsequent extraction. Secondly, this application removes fluoride through high-temperature acidification firing, where sulfuric acid and fluorides generate hydrogen. The removal of fluorine gas prevents fluorine from entering the leachate, thus eliminating the need for defluorinating agents. Furthermore, in the high-temperature acid roasting process of this application, the waste battery material dust and concentrated sulfuric acid are pre-mixed evenly, and the water produced by the reaction is released in the form of water vapor. Therefore, there is no acidity decay, which can achieve low acid consumption and low residual acid, thereby reducing alkali consumption. Moreover, the leaching residue of this application can have a graphite content of over 96%, which can be directly recycled as crude graphite, realizing the recovery of carbon elements in waste battery material dust.

[0048] The following examples illustrate a method for recovering valuable elements from waste battery material dust according to an embodiment of this application.

[0049] Example 1

[0050] Embodiment 1 of this application provides a method for recovering valuable elements from waste battery material dust, comprising the following steps:

[0051] (1) Weigh 10 kg of waste battery material dust and 10 kg of concentrated sulfuric acid. Slowly pour the concentrated sulfuric acid into the dust while stirring until no water vapor is emitted, and obtain the mixture.

[0052] (2) The mixture from step (1) is fed into a steel strip furnace. The temperature of the steel strip furnace is set to 220℃, the thickness of the material entering the furnace is 6cm, and the calcination time is 2h to obtain calcined powder. The perfluorine content of the calcined powder is reduced from 1.26% to 0.051%.

[0053] (3) Add the calcined powder obtained in step (2) to tap water at a liquid-to-solid ratio of 4:1 for extraction. Maintain the extraction temperature at 80°C and the extraction time at 2 hours. Filter, wash the leaching residue, and dry.

[0054] The leaching rates of Li, Ni, Co, and Mn were calculated to be 98.80%, 95.81%, 98.31%, and 98.84%, respectively. The F content in the leachate was 0.017 g / L, the fluorine removal rate was 96.6% (fluorine removal rate = 1 - (total fluorine in leachate / total fluorine in raw materials)), and the sulfuric acid conversion rate reached 80% (sulfuric acid conversion rate = 1 - (residual acid amount / total acid input)).

[0055] The leaching residue contains 96.6% carbon; the contents of nickel, cobalt, manganese and lithium in the leaching residue are 1.00%, 0.16%, 0.08% and 0.01% respectively (residue rate 30.86%).

[0056] Example 2

[0057] Embodiment 2 of this application provides a method for recovering valuable elements from waste battery material dust, including the following steps:

[0058] (1) Weigh 10 kg of waste battery material dust and 10 kg of concentrated sulfuric acid. Slowly pour the concentrated sulfuric acid into the dust while stirring until no water vapor is emitted, and obtain a mixture.

[0059] (2) The mixture from step (1) is fed into a steel strip furnace. The temperature of the steel strip furnace is set to 240℃, the thickness of the material entering the furnace is 6cm, and the calcination time is 2h to obtain calcined powder.

[0060] (3) Add the calcined powder obtained in step (2) to tap water at a liquid-to-solid ratio of 4:1 for leaching. Maintain the leaching temperature at 80°C and the leaching time at 2 hours. Filter, wash the leaching residue, and dry.

[0061] The calculated leaching rates of Li, Ni, Co, and Mn in the slag were 98.54%, 94.36%, 97.74%, and 98.47%, respectively. The F content in the leachate was 0.075 g / L, and the carbon content in the leaching residue was 94.3%.

[0062] Example 3

[0063] Embodiment 3 of this application provides a method for recovering valuable elements from waste battery material dust, comprising the following steps:

[0064] (1) Weigh 10 kg of waste battery material dust and 10 kg of concentrated sulfuric acid. Slowly pour the concentrated sulfuric acid into the dust while stirring until no water vapor is emitted, and obtain a mixture.

[0065] (2) The mixture from step (1) is fed into a steel strip furnace. The temperature of the steel strip furnace is set to 260℃, the thickness of the material entering the furnace is 6cm, and the calcination time is 2h to obtain calcined powder.

[0066] (3) Add the calcined powder obtained in step (2) to tap water at a liquid-to-solid ratio of 4:1 for leaching. Maintain the leaching temperature at 80°C and the leaching time at 2 hours. Filter, wash the leaching residue, and dry.

[0067] The calculated leaching rates of Li, Ni, Co, and Mn in the slag were 98.94%, 95.42%, 98.24%, and 98.98%, respectively. The F content in the leachate was 0.026 g / L, and the carbon content in the leaching residue was 96.5%.

[0068] Comparative Example 1

[0069] Comparative Example 1 of this application provides a method for recovering valuable elements from waste battery material dust, comprising the following steps:

[0070] (1) Weigh 10 kg of waste battery material dust and 12 kg of concentrated sulfuric acid. Slowly pour the concentrated sulfuric acid into the dust while stirring until no water vapor is emitted, and obtain a mixture.

[0071] (2) The mixture from step (1) is fed into a steel strip furnace. The temperature of the steel strip furnace is set to 140℃, the thickness of the material entering the furnace is 6cm, and the calcination time is 2h to obtain calcined powder.

[0072] (3) Add the calcined powder obtained in step (2) to tap water at a liquid-to-solid ratio of 4:1 for extraction. Maintain the extraction temperature at 80°C and the extraction time at 2 hours. Filter, wash the leaching residue, and dry.

[0073] The calculated leaching rates of Li, Ni, Co, and Mn in the slag were 96.24%, 88.59%, 94.51%, and 96.15%, respectively. The F content in the leachate was 0.456 g / L, and the carbon content in the leaching residue was 91.0%.

[0074] Comparative Example 2

[0075] Comparative Example 2 of this application provides a method for recovering valuable elements from waste battery material dust, comprising the following steps:

[0076] (1) Weigh 10 kg of waste battery material dust and 10 kg of concentrated sulfuric acid. Slowly pour the concentrated sulfuric acid into the dust while stirring until no water vapor is emitted, and obtain a mixture.

[0077] (2) The mixture from step (1) is fed into a steel strip furnace. The temperature of the steel strip furnace is set to 180℃, the thickness of the material entering the furnace is 6cm, and the calcination time is 2h to obtain calcined powder.

[0078] (3) Add the calcined powder obtained in step (2) to tap water at a liquid-to-solid ratio of 4:1 for extraction. Maintain the extraction temperature at 80°C and the extraction time at 2 hours. Filter, wash the leaching residue, and dry.

[0079] The calculated leaching rates of Li, Ni, Co, and Mn in the slag were 98.42%, 92.09%, 97.32%, and 98.60%, respectively. The F content in the leachate was 0.394 g / L, and the carbon content in the leaching residue was 95.5%.

[0080] Comparative Example 3

[0081] Comparative Example 3 of this application provides a method for recovering valuable elements from waste battery material dust, comprising the following steps:

[0082] (1) Weigh 10 kg of concentrated sulfuric acid, prepare a 40% sulfuric acid solution, and transfer it to the reaction vessel;

[0083] (2) Weigh 10kg of waste battery material dust and add the dust into the reactor while stirring.

[0084] (3) Heat the reactor to 80°C, maintain the stirring speed at 150 r / min, and leach for 2 hours. Filter, wash the leachate residue, and dry.

[0085] The calculated leaching rates of Li, Ni, Co, and Mn in the slag were 93.96%, 75.31%, 88.44%, and 93.40%, respectively. The F content in the leachate was 5.40 g / L, and the carbon content in the leaching residue was 69.8%.

[0086] After two stages of leaching, the contents of nickel, cobalt, manganese and lithium in the leaching residue were 4.90%, 2.48%, 2.04% and 0.22% respectively (residue rate 48.98%).

[0087] Comparative Example 4

[0088] Comparative Example 4 of this application provides a method for recovering valuable elements from waste battery material dust, including the following steps:

[0089] (1) Weigh 12 kg of concentrated sulfuric acid, prepare a 20% sulfuric acid solution, and transfer it to the reaction vessel;

[0090] (2) Weigh 10kg of waste battery material dust and add the dust into the reactor while stirring.

[0091] (3) Heat the reactor to 80°C, maintain the stirring speed at 150 r / min, and leach for 2 hours. Filter, wash the leachate residue, and dry.

[0092] The calculated leaching rates of Li, Ni, Co, and Mn in the slag were 86.90%, 59.23%, 64.11%, and 76.04%, respectively. The F content in the leachate was 1.65 g / L, and the carbon content in the leaching residue was 60.6%.

[0093] The leaching rate of metal M (M is one of Li, Ni, Co, and Mn) in Examples 1 to 3 and Comparative Examples 1 to 4 above is calculated as follows: Leaching residue mass × Mass fraction of M in the residue / (Mass of raw material × Mass fraction of M in the raw material) × 100%. The content of metal M in the solid is obtained by ICP determination of the solution composition after solid digestion and conversion.

[0094] The fluoride content in the leachate was determined using a Leici PXSJ-216 ion meter, while the fluoride content in the solid was determined by digesting the solid before using the ion meter.

[0095] Carbon content was determined using the oxidative roasting weight loss method: the leaching residue was roasted in air at 900℃ for 1 hour, and the burn loss was recorded. The remaining material was then re-stirred and roasted again for 1 hour, and the burn loss was recorded. This process was repeated until the burn losses were identical for the next two roasting cycles, ensuring sufficient contact between the carbon and air. This burn loss was considered the carbon content. The static single-phase roasting volatile matter in this material is negligible, and the other substances are metal oxides that do not react with oxygen; therefore, this method can be used to determine the carbon content.

[0096] As can be seen from the comparison between Comparative Examples 1 and 2 and Example 1 of this application, since the calcination temperatures in Comparative Examples 1 and 2 are 140°C and 180°C respectively, that is, the calcination temperatures are both less than 200°C, the fluorine removal rate is low, the fluorine content in the solution is still high, and the nickel leaching rate is significantly reduced.

[0097] As can be seen from the comparison of Comparative Examples 3 and 4 with Example 1, Comparative Examples 3 and 4 do not use the sulfation roasting process, but directly use 40% and 20% sulfuric acid solutions for extraction. The metal leaching rate is low, and the fluorine content in the solution is very high. The leaching residue still contains a high content of valuable metals, and the carbon content in the residue is below 70%, making it impossible to directly recover carbon.

[0098] As seen in Comparative Example 3, leaching the flue dust with 40% sulfuric acid at 80°C resulted in a leaching rate of approximately 90% for lithium, cobalt, and manganese, and approximately 80% for nickel. This demonstrates that direct sulfuric acid leaching yields low metal leaching rates. Furthermore, the leachate contains a large amount of fluoride, necessitating the consumption of significant amounts of defluorinating agents. Additionally, the leaching residue exhibits high nickel, cobalt, manganese, and lithium content, rendering sulfuric acid leaching alone insufficient for efficient recovery of valuable metals from the flue dust. Moreover, the leaching residue contains approximately 70% graphite, failing to meet the standards for crude graphite and hindering its recovery. In contrast, the leaching residue in Example 1 has lower nickel, cobalt, manganese, and lithium content, and a graphite content exceeding 96%, making it suitable for recycling as crude graphite.

[0099] The above embodiments merely illustrate several implementation methods of this application, and while the descriptions are relatively specific and detailed, they should not be construed as limiting the scope of this patent application. It should be noted that those skilled in the art can make various modifications and improvements without departing from the concept of this application, and these all fall within the protection scope of this application. Therefore, the protection scope of this patent application should be determined by the appended claims.

Claims

1. A method for recovering valuable elements from waste battery material dust, characterized in that, Includes the following steps: Waste battery material dust is mixed with concentrated sulfuric acid to obtain a mixture, wherein the mass ratio of the waste battery material dust to the concentrated sulfuric acid in the mixture is 1:(0.8~1.2). The waste battery material dust contains the following components: Li 0.1~3.0 wt%, Ni 5.0~40 wt%, Co 3.0~15 wt%, Mn 3.0~15 wt%, F 0.2~3.0 wt%, Fe 0.5~2.0 wt%, Al 2.0~8.0 wt%, Cu 0.5~3.0 wt%, C 10~40 wt%. The mixture is calcined to obtain calcined powder; wherein the calcination temperature is 200~300℃ and the calcination time is 1~2 h. The roasted powder is extracted with water or dilute sulfuric acid, filtered, and then a leachate and a leachate residue are obtained; wherein the liquid-to-solid ratio of the extraction is (3~5):1, the extraction temperature is 75~85 ℃, and the extraction time is 1~2 h. The leachate is extracted to obtain a sulfate solution containing valence metal elements. The leachate residue is washed with water and dried to obtain crude graphite.

2. The method as described in claim 1, characterized in that, The concentrated sulfuric acid has a mass concentration greater than 80%; the dilute sulfuric acid has a mass concentration less than 20%.

3. The method according to any one of claims 1 to 2, characterized in that, Before the extraction of the leachate, the step of removing impurities from the leachate is also included.

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