A full-element recovery method for nickel-cobalt-manganese-oxide lithium waste

Abstract

The present invention discloses a method for recovering all elements of nickel-cobalt-manganese-oxide lithium waste, which belongs to the field of waste lithium battery recovery and solves the problems of large reagent consumption, complex process and large amount of sodium sulfate decahydrate that affects production in existing recovery methods. The method of the present invention comprises the following steps: mixing and calcining materials with carbon powder; slurry carbonization; pyrolysis; cyclic acid leaching, utilizing solubility characteristics, and controlling temperature to crystallize and precipitate different sulfates in steps. The present invention greatly reduces the generation of sodium sulfate and improves the fluency and continuity of production. Since there is no large amount of sodium sulfate decahydrate generated, the crystal water in the sodium sulfate decahydrate is prevented from taking away metallic lithium, and lithium loss is significantly reduced. The method of the present invention has a short process, is simple and easy to operate, does not use organic solvents, and saves production costs; does not use hydrogen peroxide, and does not have storage problems, safety problems, environmental problems, and equipment problems.

Description

Technical Field

[0001] The invention belongs to the field of waste lithium battery recycling, and specifically relates to a full-element recycling method for nickel-cobalt-manganese-oxide lithium waste. Background Art

[0002] At present, the recycling of nickel, cobalt and manganese waste generally adopts pyrolysis and hydrolysis. There are two main recycling technologies for hydrolysis of nickel, cobalt and manganese waste:

[0003] The first method: Use acid and hydrogen peroxide to dissolve the valuable metals nickel, cobalt, manganese and lithium in the nickel cobalt manganese waste into the solution, then add sodium hydroxide to adjust the pH to 12 to obtain a ternary precursor and a lithium-containing solution, and then add sodium carbonate to the lithium-containing solution to obtain lithium carbonate. Nickel, cobalt and manganese are recovered as a ternary precursor, and lithium is recovered as lithium carbonate. This method requires a large amount of acid, alkali and hydrogen peroxide, with a large amount of reagent consumption and a high storage risk factor. In addition, the biggest problem is the production of a large amount of sodium sulfate decahydrate. In the north where the temperature is cold in winter, the production of a large amount of sodium sulfate decahydrate will cause the entire production system pipeline to be blocked, all of which will slow down the solution of sodium sulfate decahydrate, causing production to stagnate. The crystals in the system are very hard and difficult to remove, and the only way is to cut the pipeline, which has a huge impact on production.

[0004] The second method is to use acid and hydrogen peroxide to dissolve the valuable metals nickel, cobalt, manganese and lithium in the nickel cobalt manganese waste into the solution, and then use a large amount of sodium hydroxide to saponify the extractant, extract the valuable metals in multiple stages, separate the organic phase and the inorganic phase, and extract the organic phase to make the corresponding sulfate, such as nickel sulfate, cobalt sulfate and manganese sulfate, and add sodium carbonate to the inorganic phase to make lithium carbonate. This method requires the use of a large amount of organic solvents, the acid and alkali process is complex, and the process flow is long. In addition, this method also produces a large amount of sodium sulfate decahydrate in the stripping stage and the lithium carbonate precipitation stage like the first method, and the above-mentioned problems also exist. Summary of the invention

[0005] The purpose of the present invention is to provide a full-element recovery method for nickel-cobalt-manganese-lithium oxide waste to solve the problems of large reagent consumption, complex process and large amount of sodium sulfate decahydrate that affects production in the existing recovery method.

[0006] The technical solution of the present invention is: a method for recovering all elements of nickel-cobalt-manganese-lithium oxide waste, comprising the following steps:

[0007] A. Mix the nickel cobalt lithium manganese oxide waste with carbon powder evenly, and then calcine at 600-650℃ for 2-3h under nitrogen flow;

[0008] B. Add water to the mixture calcined in step A to form a slurry, and then introduce carbon dioxide to perform carbonization for 2-3 hours. After the carbonization is completed, solid-liquid separation is performed to obtain carbonized slag and carbonized liquid;

[0009] C. Heating the carbonized liquid obtained in step B to a temperature greater than or equal to 90° C. for pyrolysis, reacting for 1-2 hours to obtain lithium carbonate and pyrolysis liquid;

[0010] D. The carbonized slag obtained in step B is added with dilute sulfuric acid having an acid concentration of 100-120 g / L for acid leaching, the reaction temperature is not less than 95° C., the reaction time is 40-90 min, solid-liquid separation is performed to obtain a mixed solution I of lithium sulfate and manganese sulfate and acid leaching residue I, concentrated sulfuric acid is added to the mixed solution I to adjust the acid concentration to 100-120 g / L, and the carbonized slag is added, and the acid leaching operation is continued under the above conditions, and the leaching is carried out in a cycle so that the manganese concentration reaches 110 g / L, and then the mixture is leached at a temperature of 25-35° C. Circulating leaching is performed to make the manganese content in the leachate reach 180-190 g / L, then solid-liquid separation is performed at room temperature to obtain a mixed solution II and acid leaching residue II, the mixed solution II is heated to a temperature greater than or equal to 95° C., and the solid-liquid separation is performed while hot to obtain manganese sulfate crystals and a separated liquid, the temperature of the separated liquid is maintained at not less than 90° C. The separated liquid is continuously used for acid leaching operation, when the lithium content in the solution reaches 15-20 g / L, sodium hydroxide is added to the solution to adjust the pH to 11-12 to obtain manganese hydroxide and a lithium-containing liquid, and sodium carbonate is added to the lithium-containing liquid to precipitate lithium carbonate;

[0011] E. Wash the acid leaching residue I and the acid leaching residue II in step D with pure water at room temperature, and then separate the solid and liquid to obtain the washing residue and the washing liquid, and return the washing liquid to step D for further leaching;

[0012] F. Adding carbon powder to the washed residue obtained in step E, calcining at 650° C.-750° C. for 2-3 hours under nitrogen flow, and then calcining at 650° C.-700° C. for 2-3 hours under air flow to obtain a calcined material;

[0013] G. The calcined material obtained in step F is added with dilute sulfuric acid having an acid concentration of 180-200 g / L for acid leaching, the reaction temperature is not less than 95° C., the reaction time is 40-90 min, after solid-liquid separation, concentrated sulfuric acid is added to the obtained leachate to adjust the acid concentration to 180-200 g / L, and the calcined material is added, and the acid leaching operation is continued under the above conditions, and the cyclic leaching is performed so that the total concentration of nickel and cobalt in the leachate reaches 360-400 g / L, and then the temperature is maintained at not less than 90° C. for hot solid-liquid separation to obtain a mixed solution III and a leached residue III, and the mixed solution III is cooled to room temperature to precipitate crystals of cobalt sulfate and nickel sulfate;

[0014] H. The leached residue III obtained in step G is washed with pure water at a washing temperature of above 90° C., and the solid-liquid separation is performed while hot. The obtained washing liquid is returned to step G for further leaching.

[0015] As a further improvement of the present invention, in step A, the mass ratio of carbon powder is 10%-15% of the mixed powder.

[0016] As a further improvement of the present invention, in step B, when preparing the slurry, the ratio of the calcined material to water is prepared so that the lithium content in the carbonized liquid is 7.5-8 g / L.

[0017] As a further improvement of the present invention, in step D, during the acid leaching operation, the solid-liquid mass ratio of the carbonized slag is 1:5-6.

[0018] As a further improvement of the present invention, in step E, the washing solid-liquid mass ratio is 1:3-8, and the washing time is 30-50 minutes.

[0019] As a further improvement of the present invention, in step F, the carbon powder added accounts for 9%-12% by mass.

[0020] As a further improvement of the present invention, in step G, when the acid leaching operation is performed, the solid-liquid mass ratio of the calcined material is 1:5-6.

[0021] As a further improvement of the present invention, in step H, the washing solid-liquid mass ratio is 1:4-8, and the washing time is 30-50 minutes.

[0022] The present invention provides a recovery method with less alkali usage, no organic solvent, and little sodium sulfate decahydrate production. Compared with the prior art, the method has the following advantages:

[0023] 1. The present invention greatly reduces the generation of sodium sulfate, solves the problem that conventional methods easily generate a large amount of sodium sulfate, greatly reduces production costs, improves the smoothness and continuity of production, and will not cause the production system to collapse due to the problem of sodium sulfate crystallization.

[0024] 2. The lithium recovery rate is high. Since there is no large amount of sodium sulfate decahydrate produced, the crystal water in the sodium sulfate decahydrate is prevented from taking away the metallic lithium, and the lithium loss is significantly reduced.

[0025] 3. The method of the present invention has a short process, is simple and easy to operate, and only utilizes the different solubility characteristics of different sulfates as the temperature changes, and controls the temperature to crystallize different sulfates step by step.

[0026] 4. The method of the present invention does not use organic solvents, saving production costs. The method of the present invention does not use hydrogen peroxide, and there are no storage problems, safety problems, environmental problems, and equipment problems. DETAILED DESCRIPTION

[0027] The present invention is described in detail below in conjunction with specific implementation modes.

[0028] Embodiment 1,

[0029] A method for recovering all elements of nickel-cobalt-manganese-oxide lithium waste comprises the following steps:

[0030] A. Take 85g of ternary powder and 15g of carbon powder and mix them evenly, then calcine at 650℃ for 2h under nitrogen flow.

[0031] B. Take 80 g of the mixture calcined in step A, add 640 ml of pure water to make a slurry, and then introduce carbon dioxide for carbonization. The carbonization time is 2 hours. After the carbonization is completed, solid-liquid separation is performed to obtain carbonized slag and carbonized liquid. The lithium content of the carbonized liquid is 7.5 g / L.

[0032] C. The carbonized liquid obtained in step B is heated to above 90° C. for pyrolysis, and the reaction is carried out for 1 hour to obtain lithium carbonate and pyrolysis liquid.

[0033] D. The carbonized slag obtained in step B is added with dilute sulfuric acid with an acid concentration of 120 g / L for acid leaching. The solid-liquid ratio of the carbonized slag is 1:5, the reaction temperature is above 95° C., the reaction time is 80 min, and solid-liquid separation is performed to obtain a mixed solution I of lithium sulfate and manganese sulfate and acid leaching slag I. The pH value of the mixed solution I is about 5. Concentrated sulfuric acid is added to the mixed solution I to adjust the acid concentration to 120 g / L, and the carbonized slag is added. The acid leaching operation is continued under the above conditions. The leaching is performed in a cycle so that the manganese concentration reaches at least 110 g / L, and then the mixture is leached at a temperature of 30° C. The leaching is carried out under the condition of cyclic leaching, so that the manganese content in the leaching solution reaches 180g / L, and then the mixed solution II and the acid leaching residue II are obtained by solid-liquid separation at room temperature. The mixed solution II is heated to above 95°C, and the solid-liquid is separated into manganese sulfate crystals and separated liquid while hot. The temperature of the separated liquid is maintained at not less than 90°C, and the separated liquid is continuously used for the previous acid leaching operation to continuously precipitate manganese sulfate crystals. When the lithium content in the solution reaches 15g / L, sodium hydroxide is added to the solution to adjust the pH to 12 to obtain manganese hydroxide and lithium-containing solution. Sodium carbonate 9 times the mass of total lithium is added to the lithium-containing solution to precipitate lithium carbonate.

[0034] Leaching under low acidity of 100-120g / L and without hydrogen peroxide will only leach manganese in the carbided slag, without leaching nickel and cobalt. The solubility of manganese sulfate increases first and then decreases with increasing temperature.

[0035] E. Wash the acid leaching residue I and acid leaching residue II in step D with pure water, the washing solid-liquid mass ratio is 1:5, the washing time is 40 minutes, the washing temperature is room temperature, and then the solid-liquid separation is performed to obtain the washing residue and the washing liquid, and the washing liquid is returned to step D for further leaching.

[0036] F. Continue to add carbon powder to the washed residue obtained in step E, with the carbon powder adding accounting for 5% by mass. Calcine at 650° C. for 2 h under nitrogen flow, and then calcine at 800° C. for 2 h under air flow to obtain a calcined material.

[0037] G. The calcined material obtained in step F is added with dilute sulfuric acid having an acid concentration of 180 g / L for acid leaching, the reaction temperature is above 95°C, the reaction time is 80 min, after solid-liquid separation, concentrated sulfuric acid is added to the obtained leachate to adjust the acid concentration to 200 g / L, and the calcined material is added, and the acid leaching operation is continued according to the above conditions, and the cyclic leaching is performed so that the total concentration of nickel and cobalt in the leachate reaches 360 g / L, and then the solid-liquid separation is maintained at above 90°C while hot to obtain a mixed solution III and a leached residue III, and the mixed solution III is cooled to room temperature to precipitate crystals of cobalt sulfate and nickel sulfate. The liquid is returned to the leaching operation of this step, and mixed crystals of nickel sulfate and cobalt sulfate are continuously precipitated.

[0038] The calcined material is leached in an acid solution with an acidity of 180-200 g / L to obtain a sulfate solution containing nickel and cobalt. The solubility of nickel sulfate and cobalt sulfate increases first and then increases with increasing temperature.

[0039] H. The content of nickel and cobalt in the leached residue III obtained in step G is less than 4%. The leached residue III is washed with pure water, the washing solid-liquid mass ratio is 1:5, the washing time is 40min, the washing temperature is above 90°C, the solid-liquid separation is performed while hot, and the obtained washing liquid is returned to step G for further leaching.

[0040] Embodiment 2-4,

[0041] The difference between Example 2-4 and Example 1 is that the calcination temperature is different in step A. The effect of calcination temperature on the recovery rate of lithium is shown in Table 1.

[0042]

[0043] It can be seen from Table 1 that when the calcination temperature is 650°C, the recovery rate of lithium is high, and further increasing the temperature has little effect on the recovery rate of lithium. Therefore, the calcination temperature is preferably selected to be 650°C.

[0044] Embodiment 5-10,

[0045] The difference between Examples 5-10 and Example 1 is that the carbonization time is different in step B. The effect of carbonization time on the recovery rate of lithium is shown in Table 2.

[0046]

[0047] It can be seen from Table 2 that when the carbonization time is greater than or equal to 2h, the recovery rate of lithium is high, and further extending the carbonization time has little effect on the recovery rate of lithium. Therefore, the carbonization time is preferably selected as 2h.

[0048] Embodiment 11-13,

[0049] The difference between Examples 11-13 and Example 1 is that the pyrolysis time is different in step C. The effect of pyrolysis time on the recovery rate of lithium carbonate is shown in Table 3.

[0050]

[0051] It can be seen from Table 3 that when the carbonization time is greater than or equal to 1 h, the recovery rate of lithium carbonate is high, and further extending the pyrolysis time has little effect on the recovery rate of lithium carbonate. Therefore, the pyrolysis time is preferably selected as 1 h.

[0052] Embodiment 14-22,

[0053] The difference between Examples 14-22 and Example 1 is that the acid concentrations are different in step D. The effect of the acid concentration on the recovery rate of manganese is shown in Table 4.

[0054]

[0055] As shown in Table 4, when the acid concentration is 100-120 g / L, the recovery rate of manganese is high, and too high or too low acid concentration is not conducive to the leaching of manganese. Therefore, the concentration of leaching acid is preferably selected to be 100-120 g / L.

[0056] Embodiment 23-35,

[0057] The difference between Examples 23-35 and Example 1 is that in step D, the manganese concentrations are different during high temperature (not less than 95° C.) cyclic leaching. The effect of high temperature cyclic leaching on the crystallization of manganese sulfate is shown in Table 5.

[0058]

[0059] It can be seen from Table 5 that when the concentration of manganese ions is greater than 110g / L during high-temperature cyclic leaching, manganese sulfate will precipitate, which is not conducive to the subsequent low-temperature cyclic leaching to continue to increase the concentration of manganese. Therefore, it is preferred to control the manganese content of high-temperature cyclic leaching to be less than 110g / L, and when the manganese content reaches 110g / L, cyclic leaching at room temperature (25-35℃) is carried out.

[0060] Embodiment 36-48,

[0061] The difference between Examples 36-48 and Example 1 is that in step D, the manganese concentrations are different during the cyclic leaching at room temperature (25-35° C.). The effect of the cyclic leaching at room temperature on the crystallization of manganese sulfate is shown in Table 6.

[0062]

[0063] It can be seen from Table 6 that when the concentration of manganese ions is greater than 190 g / L during normal temperature circulation leaching, manganese sulfate will precipitate, which is not conducive to further increasing the concentration of manganese in subsequent normal temperature circulation leaching. Therefore, the content of nickel leached at normal temperature circulation should be controlled to be less than 190 g / L, and 180 g / L is preferably selected.

[0064] Embodiment 49-62,

[0065] The difference between Examples 49-62 and Example 1 is that in step F, different amounts of carbon powder are added. The effect of the amount of carbon powder added on the leaching rate of nickel and cobalt is shown in Table 7.

[0066]

[0067] It can be seen from Table 7 that when the usage of carbon powder is 8-12%, the leaching rate of nickel and cobalt reaches 99%. Therefore, the usage of carbon powder is 8%.

[0068] Embodiment 63-69,

[0069] The difference between Examples 63-69 and Example 1 is that in step F, the calcination temperature is different under the nitrogen gas flow condition. The effect of calcination temperature on nickel and cobalt leaching rates is shown in Table 8.

[0070]

[0071] It can be seen from Table 8 that when the nitrogen flow calcination temperature is 650-850°C, the leaching rate of nickel and cobalt reaches 99%. Therefore, the nitrogen flow calcination temperature is preferably selected to be 650°C.

[0072] Embodiment 70-76,

[0073] The difference between Example 70 and Example 1 is that in step D, when the acid leaching operation is performed, the solid-liquid mass ratio of the carburized slag is adjusted. The effect of the solid-liquid ratio on the leaching rate of nickel and cobalt is shown in Table 9.

[0074]

[0075] It can be seen from Table 9 that when the solid-liquid ratio is 1:5-1:8, the leaching rate of manganese reaches 89%. When the solid-liquid ratio is 1:5-1:6, the pH value of the leaching solution is 5-6, and when the solid-liquid ratio is 1:6-1:8, the pH value of the leaching solution is 2-4. Considering the utilization rate of the acid, the leaching solid-liquid ratio of 1:5-1:6 is selected as the optimal ratio.

[0076] Embodiment 77-83,

[0077] The difference between Examples 77-83 and Example 1 is that the washing solid-liquid mass ratio is different in step E. The effect of the washing solid-liquid ratio on the manganese content in the washing residue is shown in Table 10.

[0078]

[0079] It can be seen from Table 10 that when the washing solid-liquid ratio is 1:3-1:8, the manganese content in the washing residue is about 0.2%. Therefore, the washing solid-liquid ratio is preferably selected to be 1:3.

[0080] Embodiment 84-90,

[0081] The difference between Example 84 and Example 90 is that in step G, the solid-liquid mass ratio of the carburized slag is different during the acid leaching operation. The effect of the solid-liquid mass ratio on the leaching rate of nickel and cobalt is shown in Table 11.

[0082]

[0083] It can be seen from Table 11 that when the solid-liquid ratio is 1:5-1:8, the leaching rate of nickel and cobalt reaches 99%, the solid-liquid ratio is 1:5-1:6, the pH value of the leaching solution is 5-6, and the pH value of the leaching solution is 2-4 when the solid-liquid ratio is 1:6-1:8. Considering the utilization rate of the acid, the preferred leaching solid-liquid ratio is 1:5-1:6.

[0084] Embodiment 77-83,

[0085] The difference between Examples 77-83 and Example 1 is that the washing solid-liquid mass ratio is different in step E. The effect of the washing solid-liquid mass ratio on the manganese content in the washing residue is shown in Table 12.

[0086]

[0087] It can be seen from Table 12 that when the washing solid-liquid ratio is 1:4-1:8, the nickel content in the washing residue is 0.5%, and the cobalt content is about 0.4%. Therefore, the washing solid-liquid mass ratio is preferably selected to be 1:4.

Claims

1. A method for recovering all elements of nickel-cobalt-manganese-oxide lithium waste, characterized in that The following steps are involved: A. Mix the nickel cobalt lithium manganese oxide waste with carbon powder evenly, and then calcine at 600-650℃ for 2-3h under nitrogen flow; B. Add water to the mixture calcined in step A to form a slurry, and then introduce carbon dioxide to perform carbonization for 2-3 hours. After the carbonization is completed, solid-liquid separation is performed to obtain carbonized slag and carbonized liquid; C. Heating the carbonized liquid obtained in step B to a temperature greater than or equal to 90° C. for pyrolysis, and reacting for 1-2 hours to obtain lithium carbonate and pyrolysis liquid; D. The carbonized slag obtained in step B is added with dilute sulfuric acid having an acid concentration of 100-120 g / L for acid leaching, the reaction temperature is not less than 95° C., the reaction time is 40-90 min, solid-liquid separation is performed to obtain a mixed solution I of lithium sulfate and manganese sulfate and acid leaching residue I, concentrated sulfuric acid is added to the mixed solution I to adjust the acid concentration to 100-120 g / L, and the carbonized slag is added, and the acid leaching operation is continued under the above conditions, and the leaching is carried out in a cycle so that the manganese concentration reaches 110 g / L, and then the mixture is leached at a temperature of 25-35° C. Circulating leaching is performed to make the manganese content in the leachate reach 180-190 g / L, then solid-liquid separation is performed at room temperature to obtain a mixed solution II and acid leaching residue II, the mixed solution II is heated to a temperature greater than or equal to 95° C., and the solid-liquid separation is performed while hot to obtain manganese sulfate crystals and a separated liquid, the temperature of the separated liquid is maintained at not less than 90° C. The separated liquid is continuously used for acid leaching operation, when the lithium content in the solution reaches 15-20 g / L, sodium hydroxide is added to the solution to adjust the pH to 11-12 to obtain manganese hydroxide and a lithium-containing liquid, and sodium carbonate is added to the lithium-containing liquid to precipitate lithium carbonate; E. Wash the acid leaching residue I and the acid leaching residue II in step D with pure water at room temperature, and then separate the solid and liquid to obtain the washing residue and the washing liquid, and return the washing liquid to step D for further leaching; F. Adding carbon powder to the washed residue obtained in step E, calcining at 650° C.-750° C. for 2-3 hours under nitrogen flow, and then calcining at 650° C.-700° C. for 2-3 hours under air flow to obtain a calcined material; G. The calcined material obtained in step F is added with dilute sulfuric acid having an acid concentration of 180-200 g / L for acid leaching, the reaction temperature is not less than 95° C., the reaction time is 40-90 min, after solid-liquid separation, concentrated sulfuric acid is added to the obtained leachate to adjust the acid concentration to 180-200 g / L, and the calcined material is added, and the acid leaching operation is continued under the above conditions, and the cyclic leaching is performed so that the total concentration of nickel and cobalt in the leachate reaches 360-400 g / L, and then the temperature is maintained at not less than 90° C. for hot solid-liquid separation to obtain a mixed solution III and a leached residue III, and the mixed solution III is cooled to room temperature to precipitate crystals of cobalt sulfate and nickel sulfate; H. The leached residue III obtained in step G is washed with pure water at a washing temperature of above 90° C., and the solid-liquid separation is performed while hot. The obtained washing liquid is returned to step G for further leaching.

2. The method for recovering all elements of nickel-cobalt-manganese-oxide waste according to claim 1, characterized in that: In step A, the mass ratio of carbon powder is 10%-15% of the mixed powder.

3. A method for recovering all elements of nickel-cobalt-lithium manganese oxide waste according to claim 1 or 2, characterized in that: In step B, when preparing the slurry, the ratio of the calcined material to water is prepared so that the lithium content in the carbonized liquid is 7.5-8 g / L.

4. The method for recovering all elements of nickel-cobalt-lithium manganese oxide waste according to claim 3, characterized in that: In step D, during the acid leaching operation, the solid-liquid mass ratio of the carbonized slag is 1:5-6.

5. The method for recovering all elements of nickel-cobalt-lithium manganese oxide waste according to claim 4, characterized in that: In step E, the washing solid-liquid mass ratio is 1:3-8, and the washing time is 30-50 minutes.

6. The method for full element recovery of nickel-cobalt-lithium manganese oxide waste according to claim 5, characterized in that: In step F, the carbon powder replenishment accounts for 9%-12% by mass.

7. The method for recovering all elements of nickel-cobalt-lithium manganese oxide waste according to claim 6, characterized in that: In step G, during the acid leaching operation, the solid-liquid mass ratio of the calcined material is 1:5-6.

8. The method for recovering all elements of nickel-cobalt-lithium manganese oxide waste according to claim 7, characterized in that: In step H, the washing solid-liquid mass ratio is 1:4-8, and the washing time is 30-50 min.

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