A method for selectively recovering valuable metals from lithium-manganese slag and its application

Abstract

The present invention discloses a method and application for selectively recovering valuable metals from lithium-manganese slag, belonging to the field of resource recovery and utilization technology. The method comprises the following steps: subjecting the lithium-manganese slag to smelting, first-stage roasting, lithium extraction, second-stage roasting, and manganese extraction, in sequence, to achieve selective recovery of lithium and manganese, wherein a monovalent salt is added in the first-stage roasting process, a divalent salt is added in the second-stage roasting process, the volume proportion of oxygen in the first-stage roasting process and the second-stage roasting process is controlled to be less than 20%, and water immersion is adopted in the lithium extraction process and the manganese extraction process to obtain a lithium-containing solution and a manganese-containing solution, respectively. The method can be applied to selectively recover valuable metals in the recovery of lithium-ion batteries. The present invention can achieve step-by-step dissolution of lithium and manganese, and can also solve the problems of excessive dissolution of impurities, easy formation of silica gel, and difficulty in filtration, thereby shortening the subsequent purification and refining process of lithium.

Description

Technical Field

[0001] The present invention belongs to the technical field of resource recovery and utilization, and more specifically, relates to a method and application for selectively recovering valuable metals from lithium-manganese slag. Background Art

[0002] Lithium, known as "white oil," is a vital building block for modern industry, playing a key role in new energy sectors, such as lithium batteries. With the continued growth in global demand for renewable energy and electric vehicles, the demand for lithium batteries has also increased, directly driving the booming lithium market. Furthermore, lithium has a wide range of applications in fields such as glass and ceramics, aerospace, and more, further broadening its market prospects. Manganese metal is also of significant importance. Manganese is an indispensable alloying element in steel production, enhancing the steel's toughness, wear resistance, and corrosion resistance. With the continued development of global infrastructure construction and the real estate industry, demand for steel will continue to grow, thereby driving demand for manganese metal. Manganese also has a wide range of applications in batteries, magnetic materials, and other fields. In particular, in emerging battery technologies, where manganese is a key cathode material, demand is also increasing.

[0003] Lithium-ion batteries, due to their long cycle life, high capacity, and lightweight characteristics, are widely used in electric vehicles, communications, electronic equipment, and other fields. With the rapid increase in lithium-ion battery usage, it is expected that the number of used lithium-ion batteries will reach a peak by 2025. Used lithium-ion batteries contain toxic electrolytes and organic separators, which can cause irreversible damage to soil and water quality. In addition, they contain large amounts of strategic key metals such as Li (2-3%), Mn (2-20%), Co (5-30%), Ni (0.02-10%), and Cu (7-17%). If not recycled, they will result in a huge waste of resources.

[0004] The main recycling methods for used lithium-ion batteries include pyrometallurgy and hydrometallurgy. Pyrometallurgy is the most promising, short-process method due to its adaptability to raw materials and large processing scale. However, the current difficulty with pyrometallurgy is that conventional smelting slag contains a low lithium content and complex lithium-bearing mineral phases, making it difficult to selectively and efficiently extract lithium. Existing lithium extraction processes from battery slag typically utilize acid roasting, such as sulfuric acid or hydrochloric acid.

[0005] After searching, the patent application document with Chinese patent number 202211074571.0 and application publication date of January 31, 2023 discloses a method for selectively separating and recovering lithium and manganese from waste lithium manganate batteries. The method comprises: (1) discharging, disassembling, and classifying waste lithium manganate batteries to obtain positive electrode materials, and then subjecting them to high-temperature treatment to remove the binder; (2) roasting the lithium manganate positive electrode powder obtained by pretreatment with sulfates such as ferrous sulfate, ferric sulfate, ammonium sulfate, and ammonium bisulfate in two stages at a certain temperature to convert the lithium element of lithium manganate into the corresponding sulfate; (3) leaching the roasted product obtained in the second step with water to obtain lithium sulfate leachate and leaching residue rich in iron and manganese oxides. However, this process is for processing the lithium manganate positive electrode powder obtained after disassembly, the process is complicated, and the manganese element is insoluble. Subsequently, the filter residue of Fe2O3, Mn2O3 or FeMnO3 needs to be treated again to recover manganese.

[0006] Chinese patent application number 201610320979.X, published on May 16, 2016, discloses a method for extracting lithium and manganese from lithium-containing manganese-rich slag. The method involves slurrying lithium-containing manganese-rich slag (a product obtained from the pyrometallurgical smelting of spent lithium-ion batteries) with a sulfuric acid solution to produce a mixed slurry; maintaining the mixed slurry at a temperature of no less than 50°C; calcining the product after calcination, controlling the calcination temperature to no less than 150°C for a period of no less than 30 minutes; leaching the calcined product, and then performing solid-liquid separation to produce a solution containing manganese sulfate and lithium sulfate. However, this method suffers from varying degrees of aluminum and silicon dissolution during the calcination process, making filtration difficult. The method also suffers from poor selectivity, complicating the subsequent separation of lithium, manganese, and aluminum, and causing varying degrees of corrosion to the equipment. Furthermore, the resulting leachate contains both manganese and lithium, rather than separate extractions of the two.

[0007] Therefore, developing a green and efficient recovery process for valuable elements in lithium-manganese slag will not only help solve the problem of environmental pollution, but also alleviate the global energy metal resource tension. Summary of the Invention

[0008] 1. Problem to be solved

[0009] To address the poor selectivity of existing roasting methods for recovering valuable metals, the present invention provides a method for selectively recovering valuable metals from lithium-manganese slag. By optimizing the roasting process, lithium and manganese can be dissolved in stages. This method also addresses issues such as excessive impurity dissolution and the formation of silica gel, which is difficult to filter, thereby shortening the subsequent lithium purification process.

[0010] 2. Technical solution

[0011] In order to solve the above problems, the technical solutions adopted by the present invention are as follows:

[0012] Based on the phenomenon of isomorphism in natural minerals, the present invention adds metal cation salts with the same valence and similar ionic radius to lithium-manganese slag, and achieves step-by-step separation and recovery of lithium and manganese through common ion replacement. The details are as follows:

[0013] A method for selectively recovering valuable metals from lithium-manganese slag, comprising the following steps:

[0014] S1. Melting process:

[0015] Smelting the lithium-manganese slag to obtain lithium-manganese smelting slag;

[0016] S2, first stage roasting process:

[0017] mixing the smelting slag in step S1 with a monovalent salt and then calcining the mixture, wherein the volume proportion of oxygen in the calcining atmosphere is controlled to be less than 20%, to obtain a calcined product comprising a soluble lithium salt, an insoluble aluminum silicate, and an insoluble manganese silicate;

[0018] S3, lithium extraction process:

[0019] The calcined product obtained in step S2 is water-soaked, and the filtrate obtained by filtration is a lithium-containing solution, and the filter residue is a manganese-rich slag, the main phases of which are aluminum silicate and manganese silicate; wherein the content of divalent Mn(II) in the manganese-rich slag is 3-25%, and the content of tetravalent Mn(IV) is 0.5-2.0%;

[0020] S4, second stage roasting process:

[0021] The manganese-rich slag in step S3 is dried and mixed with a divalent salt and then roasted, wherein the volume proportion of oxygen in the roasting environment atmosphere is controlled to be less than 20%, thereby obtaining a roasting product including soluble manganese salt and insoluble aluminosilicate. It should be noted that the oxygen content in the first-stage roasting determines the tetravalent manganese content in the manganese-rich slag. Although the oxygen content in the second-stage roasting does not determine the tetravalent manganese content in the manganese-rich slag, it affects the leaching rate of manganese in the second stage, because excessively high oxygen content in the second-stage roasting process will increase the tetravalent manganese content in the second-stage roasting product, thereby affecting the final manganese leaching rate.

[0022] S5, manganese extraction process:

[0023] The calcined product obtained in step S4 is soaked in water, and filtered to obtain a filtrate which is a manganese-containing solution and a filter residue which is aluminosilicate. The manganese-containing solution contains divalent manganese, and the filter residue contains a small amount of tetravalent manganese.

[0024] Furthermore, in step S1, the lithium-manganese-containing smelting slag includes but is not limited to lithium-manganese-containing slag obtained by pyrometallurgical high-temperature reduction smelting of one or more waste lithium-ion batteries (such as lithium manganate batteries, ternary power batteries, etc.), wherein the Li content is 0.5-7.5wt%, the mass fraction ratio of Al2O3 to SiO2 is (0.2-3.5):1, and the Mn content is 3-25wt%.

[0025] Furthermore, in step S1, in the smelting process, the smelting conditions are: temperature 1400-1700°C, time ≥20min, and the amount of slag-forming agent SiO2 added during the smelting process satisfies the mass fraction ratio of Al2O3 to SiO2 of (0.25-3.0):1.

[0026] Furthermore, in step S1, the lithium-manganese smelting slag is smelted according to the relevant requirements of YS / T 509.1-2008 (YS / T509.1-2008 Chemical analysis method of spodumene and lepidolite concentrates - Determination of lithium oxide, sodium oxide and potassium oxide content - Flame atomic absorption spectrometry).

[0027] Furthermore, the monovalent salt used in step S2 can be exchanged with monovalent lithium ions, and its monovalent cations are Na + , K + NH4 + The present invention includes one or more of Na2SO4, K2SO4, (NH4)2SO4, KCl, NaCl, and NH4Cl, preferably at least one of Na2SO4 and K2SO4. This is because chloride salts generally have lower melting points than sulfate salts, forming molten salts that separate from the slag during the roasting process, which is not conducive to the reaction kinetics. In addition, in the presence of a weak water vapor atmosphere, chloride salts easily decompose into monovalent metal oxides, forming a stable silicate phase with the SiO2 component in the lithium-rich manganese slag, and cannot achieve common ion exchange with lithium. Finally, it cannot be ignored that chloride salts easily generate hydrogen chloride gas.

[0028] Furthermore, in step S2, the mass ratio of smelting slag to monovalent salt is controlled to be (1-6):1. Within this range, the recovery rate of lithium increases with the increase of monovalent salt, but excessive monovalent salt will increase the cost and increase the Na and K content of the leachate, which is not conducive to subsequent lithium extraction.

[0029] Furthermore, the method for controlling the oxygen content in steps S2 and S4 is: adjusting the volume fraction of oxygen in the atmosphere by controlling the flow ratio of the introduced air and the inert gas.

[0030] Furthermore, the calcination temperature in step S2 is 600-1000° C., preferably 700-900° C. The calcination temperature not only affects the intensity of the common ion exchange reaction, but also affects the stability of the calcined product: if the temperature is too low, the material will not react fully; if the temperature is too high, the salt may decompose, which is not conducive to improving the lithium recovery rate.

[0031] Furthermore, the calcination time in step S2 is 0.5 to 3 hours, preferably 1.5 to 2 hours.

[0032] Furthermore, in step S4, the divalent salt can be exchanged with the divalent manganese ion, and the divalent metal cations are Ca 2+ Mg 2+ ; Including anhydrous CaSO4, MgSO4, CaCl2, MgCl2, and / or one or more of CaSO4, MgSO4, CaCl2, MgCl2 containing crystal water. The aqueous salt will increase the atmospheric water vapor during the roasting process, promote the decomposition of the chloride salt, and form a small amount of HCl gas, which will corrode the equipment on the one hand and reduce the salt conversion efficiency on the other hand. At least one of anhydrous CaCl2 and anhydrous MgCl2 is preferred.

[0033] Furthermore, in step S4, the mass ratio of manganese-rich slag to divalent salt is controlled to be (1-5):1. Within this range, the recovery rate of manganese increases with the increase of divalent salt, but excessive divalent salt will increase the cost and increase the calcium and magnesium content of the leachate, increase the viscosity of the leachate, and be unfavorable for subsequent manganese leaching.

[0034] Furthermore, in step S4, the calcination temperature is 700-1100° C., preferably 800-1000° C., and the calcination time is 0.5-3 h, preferably 1.5-2 h.

[0035] Furthermore, the water leaching conditions in steps S3 and S5 are: temperature of 10-80°C, leaching time of 2 minutes to 2 hours, and liquid-to-solid ratio of (3-50):1 mL / g. Preferably, the conditions are 40-60°C, 10-30 minutes, and (10-20):1. The lithium leaching rate is greater than 99%, and the impurity content in the lithium-containing solution is extremely low, with the concentrations of Ni, Co, Mn, Al, and Si all below 5 mg / L. The manganese leaching rate is greater than 90%, and the impurity metal content in the manganese-containing solution is extremely low, with the concentrations of Ni, Co, Li, Al, and Si all below 5 mg / L.

[0036] An application of the above method in recycling lithium-ion batteries, wherein the lithium-manganese slag includes but is not limited to one or more waste lithium-ion batteries.

[0037] In one possible embodiment of the present invention, the present invention relates to a process for extracting lithium and manganese from lithium-containing slag during the pyrometallurgical recovery of waste lithium-ion batteries, using slag from the pyrometallurgical smelting of waste lithium batteries as raw materials, and adopting a process containing K, Na, NH4 + The filter residue is subjected to a second-stage roasting with one or more salts containing Ca and Mg to obtain a Mn-containing solution with a low impurity content.

[0038] 3. Beneficial effects

[0039] Compared with the prior art, the present invention has the following beneficial effects:

[0040] The present invention uses a two-step salt roasting process for lithium-manganese slag, which can be used to treat lithium-containing slag after pyrometallurgical smelting of waste lithium-ion batteries. The present invention has the following characteristics:

[0041] (1) Raw material characteristics

[0042] The raw material of the present invention can be lithium-containing slag obtained by high-temperature reduction smelting of waste lithium-ion batteries (such as lithium cobalt oxide batteries, lithium manganese oxide batteries, ternary power batteries, etc.). Compared with natural lithium-containing ores, the raw material of the present invention can be lithium-containing slag obtained by high-temperature smelting, which can achieve the self-transformation of lithium-containing mineral phase, high activity, simple phase and lower impurity ion content.

[0043] (2) One-stage calcination has strong selectivity for Li and high Li yield

[0044] The S2-S3 process is an in-situ ion replacement process, which utilizes the isomorphous substitution properties of metal cation salts with the same valence and similar ionic radius and the properties of large differences in chemical potential gradients. The monovalent cations in the salt open up ion channels in the lithium-rich manganese slag, and selectively replace Li under the appropriate conditions of time, temperature, and salt content to form soluble Li salts and insoluble potassium sodium silicates, thus achieving mineral phase reconstruction. The salt used in the first-stage roasting does not react with manganese and aluminum, and the concentrated Ni, Co, Mn, Al, and Si in the Li-containing solution are obtained. The concentrations are all below 5ppm. Since the salt roasting process is carried out at a low temperature, no chlorine- or sulfur-containing gases are volatilized, thus causing no corrosion to the equipment. At the same time, the filtration performance is excellent, and the leaching rates of manganese, aluminum, and silicon are all below 1%. Problems such as the difficulty in liquid-solid separation caused by silica gel will not occur. In addition, the type of salt used is simple, and only one or more potassium, sodium, and ammonium salts are used in the roasting process. The potassium-sodium salts contained in the lithium mother liquor obtained in the subsequent preparation of lithium carbonate can be recycled after evaporation and crystallization. At the same time, the recovery rate is higher than that of natural lithium-containing ores, and the lithium recovery rate is close to 100%.

[0045] (3) The second stage roasting has strong selectivity for Mn and high manganese yield

[0046] The manganese content in lithium-containing manganese slag is 3-25%. Conventional manganese extraction methods are mostly sulfuric acid roasting-water leaching methods, but there is a problem of poor selectivity, resulting in the presence of a large amount of impurity elements such as aluminum and silicon in the manganese-containing solution, and subsequent tedious impurity removal processes are still required. The key to the two-stage manganese extraction of the present invention is the use of a divalent salt roasting method, which can extract manganese with high selectivity. Therefore, the concentrations of Ni, Co, Li, Al, and Si in the Mn-containing solution are all lower than 5ppm; at the same time, controlling the oxygen content in the roasting process to reduce the formation of tetravalent manganese is also the key to improving the manganese recovery rate, because high-valent (>2-valent) manganese is more stable and difficult to achieve ion replacement. Specifically:

[0047] The present application is similar to the principles of S2-S3 in the S4-S5 process. It utilizes the isomorphous substitution properties and large chemical potential gradient differences of metal cation salts with the same valence and similar ionic radius. Divalent metal cations open up ion channels and selectively replace Mn under appropriate conditions of time, temperature, and salt content to form soluble manganese salts and insoluble calcium magnesium aluminum silicate. This not only ensures the integrity of the original lattice structure but also achieves mineral phase reconstruction. The salt used in the second-stage roasting does not react with aluminum. The concentrations of Ni, Co, Li, Al, and Si in the obtained Mn-containing solution are all less than 5ppm. Since the salt roasting process is at a low temperature, no chlorine- or sulfur-containing gas volatilization is generated, thus causing no corrosion to the equipment. The leaching rates of aluminum and silicon are both less than 1%, and the filtration performance is excellent. There is no problem of difficult liquid-solid separation caused by silica gel. The type of salt used is simple. Only one or two calcium salts or magnesium salts are used in the roasting process. The calcium- and magnesium-containing waste liquid generated by the subsequent production of manganese sulfate or other manganese products is evaporated and crystallized to obtain calcium-magnesium salts that can be recycled. At the same time, the recovery rate is higher than that of natural lithium-containing ores, and the Mn recovery rate is close to 90%.

[0048] (3) The obtained Li-containing solution and Mn-containing solution are widely used in subsequent applications.

[0049] The impurity ion content of the Li-containing solution is low. After simple impurity removal, concentration and crystallization, high-purity Li2CO3 can be produced by introducing CO2 or adding potassium and sodium carbonate. The impurity ion content of the Mn-containing solution is low. The pH is adjusted to allow Ca and Mg ions to form precipitation. High-valent manganese can be obtained through concentration, crystallization and low-temperature calcination, which can be used in the production of positive electrode materials. BRIEF DESCRIPTION OF THE DRAWINGS

[0050] Figure 1 is a process flow chart of the present invention;

[0051] Figure 2 This is a SEM image of the lithium-manganese smelting slag in the embodiment;

[0052] Figure 3 This is the SEM image of the first stage calcined product in Example 1;

[0053] Figure 4 This is the SEM image of the second-stage calcined product in Example 1. DETAILED DESCRIPTION

[0054] The present invention will be further described below with reference to specific embodiments.

[0055] In order to better understand the present invention, the present invention is further described below through specific implementation methods, but the exemplified embodiments do not limit the protection scope of the present invention.

[0056] In a specific embodiment, lithium-manganese slag is used as an example to recover valuable metals, lithium and manganese, from waste lithium-ion batteries. The specific steps are as follows:

[0057] S1. Melting process:

[0058] The waste lithium-ion batteries were smelted by high-temperature reduction pyrolysis to obtain lithium-manganese smelting slag. The smelting conditions were as follows: temperature 1400-1700℃, time ≥20min, and the amount of slag-forming agent SiO2 added during the smelting process was such that the mass fraction ratio of Al2O3 to SiO2 was (0.25-3.0):1. The main components of the lithium-manganese smelting slag are shown in Table 1, and the SEM characterization results are shown in Table 1. Figure 2 shown.

[0059] S2, first stage roasting process:

[0060] According to Table 2, the lithium-manganese smelting slag in step S1 was mixed with a monovalent salt and then calcined. The calcination temperature was controlled to 800°C, the calcination time was controlled to 1.5h, and the oxygen volume fraction in the calcination environment was controlled to obtain a calcined product. The SEM characterization results of the calcined product in Example 1 are as follows: Figure 3 As shown, its main phases are soluble lithium salts, insoluble aluminum silicates and insoluble manganese silicates;

[0061] S3, lithium extraction process:

[0062] The calcined product obtained in step S2 is subjected to water leaching, wherein the leaching time of the water leaching process is 2 hours, the liquid-to-solid ratio is 20:1 mL / g, and the leaching temperature is 60° C.; the filtrate obtained by filtration is a lithium-containing solution, and the filter residue is a manganese-rich slag, the main phases of the manganese-rich slag are aluminum silicate and manganese silicate; wherein the impurity content in the lithium-containing solution is extremely low, and the concentrations of Ni, Co, Mn, Al, and Si are all less than 5 mg / L; the content of divalent Mn(II) in the manganese-rich slag is 3-25%, and the content of tetravalent Mn(IV) is 0.5-2.0%;

[0063] S4, second stage roasting process:

[0064] According to Table 3, the manganese-rich slag in step S3 was dried and mixed with a divalent salt and then calcined. The calcination temperature was controlled to be 900°C and the calcination time was 2h. The volume ratio of oxygen in the calcination atmosphere was controlled to obtain a calcined product. The SEM characterization results of the second-stage calcined product in Example 1 are as follows. Figure 4 As shown, its main phases include soluble manganese salts and insoluble aluminosilicates;

[0065] S5, manganese extraction process:

[0066] The roasted product obtained in step S4 was water-leached for 1.5 h, with a liquid-to-solid ratio of 20:1 mL / g and a leaching temperature of 50° C. The filtrate obtained by filtration was a manganese-containing solution, and the filter residue was aluminosilicate. The content of impurity metals in the manganese-containing solution was extremely low, wherein the concentrations of Ni, Co, Li, Al, and Si were all lower than 5 mg / L.

[0067] Table 1 Detection results of main components of lithium-manganese slag (wt%)

[0068]

[0069] Table 2 Parameters of the first stage roasting process and the content of low-valent manganese in the manganese-rich slag in each embodiment

[0070]

[0071] Table 3 Parameters of the second stage roasting process of each embodiment

[0072]

[0073] It is worth noting that the reaction conditions in the above embodiments are all optimal reaction conditions for the reaction, and the complex of the present invention can be synthesized under the reaction conditions within the scope of the claims. In order to avoid redundancy, no further examples are given here, such as the temperature and time of the first and second stage roasting. Those skilled in the art can obtain better roasting effects by extending the roasting time at a low roasting temperature within the scope defined by the claims of the present application. In addition, those skilled in the art can reproduce the present invention based on the general formula of the complex of the present invention and the related solvents, catalysts and other reagents listed. Since there are too many examples, several representative examples are selected here to prove that the complex of the present invention has excellent luminous efficiency and low oxygen quenching rate, so the examples are not repeated here.

Claims

1. A method for selectively recovering valuable metals from lithium-manganese slag, characterized in that: The lithium-manganese slag is sequentially subjected to smelting, first-stage roasting, lithium extraction, second-stage roasting, and manganese extraction processes to achieve selective recovery of lithium and manganese, wherein a monovalent salt is added in the first-stage roasting process, and a divalent salt is added in the second-stage roasting process. The volume proportion of oxygen in the first and second-stage roasting processes is controlled to be less than 20%. Water immersion is used in the lithium extraction process and the manganese extraction process to obtain a lithium-containing solution and a manganese-containing solution, respectively. The lithium-manganese slag is one or more waste lithium-ion batteries.

2. The method for selectively recovering valuable metals from lithium-manganese slag according to claim 1, wherein: In the smelting process, the smelting conditions are: temperature 1400~1700℃, time ≥20 min.

3. The method for selectively recovering valuable metals from lithium-manganese slag according to claim 1, wherein: In the first roasting process, the monovalent salt is sulfate and / or chloride, wherein the cation of the monovalent salt is Na + , K + NH4 + At least one of the following: controlling the mass ratio of smelting slag to monovalent salt to be (1~6):

1.

4. The method for selectively recovering valuable metals from lithium-manganese slag according to claim 1, wherein: In the second stage roasting process, the divalent salt is sulfate and / or chloride, wherein the cation of the divalent salt is Ca 2+ Mg 2+ At least one of .

5. The method for selectively recovering valuable metals from lithium-manganese slag according to claim 4, wherein: The divalent salt is a salt without crystal water, a salt containing crystal water, or a mixture of salt without crystal water and salt containing crystal water; the mass ratio of manganese-rich slag to divalent salt is controlled to be (1-5):

1.

6. The method for selectively recovering valuable metals from lithium-manganese slag according to claim 1, wherein: In the first stage roasting process: the roasting temperature is 600~1000℃; the roasting time is 0.5~3h; In the second stage roasting process: the roasting temperature is 700~1100℃, and the roasting time is 0.5~3h.

7. The method for selectively recovering valuable metals from lithium-manganese slag according to claim 6, wherein: In the first stage of the roasting process: the roasting temperature is 700-900℃; the roasting time is 1.5-2h; In the second stage roasting process: the roasting temperature is 800~1000℃, and the roasting time is 1.5~2h.

8. The method for selectively recovering valuable metals from lithium-manganese slag according to claim 1, wherein: The lithium extraction process is as follows: the calcination product in the first calcination process is leached in water to obtain a filtrate which is a lithium-containing solution, and the filter residue is a manganese-rich slag, wherein the lithium leaching rate is greater than 99%, the concentrations of Ni, Co, Mn, Al, and Si in the lithium-containing solution are all lower than 5 mg / L, the content of Mn (II) in the manganese-rich slag is 3~25%, and the content of Mn (IV) is 0.5~2.0%.

9. The method for selectively recovering valuable metals from lithium-manganese slag according to claim 1, wherein: The manganese extraction process is as follows: the roasting product in the second-stage roasting process is leached in water to obtain a filtrate as a manganese-containing solution, and the filter residue is aluminosilicate, wherein the manganese leaching rate is greater than 90%, and the concentrations of Ni, Co, Li, Al, and Si in the manganese-containing solution are all lower than 5 mg / L.

10. The method for selectively recovering valuable metals from lithium-manganese slag according to any one of claims 1 to 9, characterized in that: The conditions for water immersion are: temperature of 10~80℃, leaching time of 2min~2h, and liquid-to-solid ratio of (3~50):1mL / g.

11. The method for selectively recovering valuable metals from lithium-manganese slag according to claim 10, wherein: The conditions for water immersion are: temperature of 40~60℃, leaching time of 10~30min, and liquid-to-solid ratio of (10~20):1 mL / g.

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