WO2024037272A1 - 一种三元电池粉浸出方法 - Google Patents

一种三元电池粉浸出方法 Download PDF

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Publication number
WO2024037272A1
WO2024037272A1 PCT/CN2023/108096 CN2023108096W WO2024037272A1 WO 2024037272 A1 WO2024037272 A1 WO 2024037272A1 CN 2023108096 W CN2023108096 W CN 2023108096W WO 2024037272 A1 WO2024037272 A1 WO 2024037272A1
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Prior art keywords
leaching
slag
powder
ternary battery
roasting
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PCT/CN2023/108096
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English (en)
French (fr)
Inventor
刘勇奇
何然
程青云
郑宇�
巩勤学
李长东
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广东邦普循环科技有限公司
湖南邦普循环科技有限公司
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Priority to PCT/CN2023/108096 priority Critical patent/WO2024037272A1/zh
Priority to CN202380010131.5A priority patent/CN117223150A/zh
Publication of WO2024037272A1 publication Critical patent/WO2024037272A1/zh

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    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B26/00—Obtaining alkali, alkaline earth metals or magnesium
    • C22B26/10—Obtaining alkali metals
    • C22B26/12—Obtaining lithium
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
    • C22B3/06—Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic acid solutions, e.g. with acids generated in situ; in inorganic salt solutions other than ammonium salt solutions
    • C22B3/08—Sulfuric acid, other sulfurated acids or salts thereof
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
    • H—ELECTRICITY
    • H01—ELECTRIC ELEMENTS
    • H01M—PROCESSES OR MEANS, e.g. BATTERIES, FOR THE DIRECT CONVERSION OF CHEMICAL ENERGY INTO ELECTRICAL ENERGY
    • H01M10/00—Secondary cells; Manufacture thereof
    • H01M10/54—Reclaiming serviceable parts of waste accumulators
    • Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
    • Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
    • Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
    • Y02P10/00—Technologies related to metal processing
    • Y02P10/20—Recycling

Definitions

  • the present disclosure relates to the field of resource recovery, and in particular to a ternary battery powder leaching method.
  • the sulfuric acid atmospheric pressure leaching method uses a combination of multi-stage low-acid leaching and high-acid leaching to reduce and leach the nickel, cobalt, manganese, and lithium metal in the ternary battery powder into a metal sulfate solution. It takes a long time, consumes a huge amount of sulfuric acid, and Because its metal leaching rate is not high, a large amount of graphite slag mixed with nickel, cobalt, manganese, and lithium metal will be produced after leaching, causing environmental pollution and resource waste; the roasting-water leaching combined method is to extract nickel through wet-fire mixed metallurgy. Cobalt manganese lithium metal, but this method will produce a large amount of roasting slag mixed with nickel cobalt manganese lithium metal, which also has the problems of environmental pollution and low leaching rate.
  • the purpose of the present disclosure is to provide a ternary battery powder leaching method, which solves the problem of large leaching residue, environmental pollution and resource pollution caused by the existing ternary battery powder leaching method due to difficulty in leaching, long time consumption and low leaching rate. It solves the problem of waste and has the advantages of fast leaching and high leaching rate.
  • the ternary battery powder carries Cl in part of the solvent.
  • Cl has a great penetrating ability under high temperature and high pressure. It can directly react with the metal through the oxide film on the metal surface to generate the corresponding chloride and evaporate. The evaporated chloride will react with oxygen. The reaction produces chlorine gas and the corresponding oxides.
  • the oxide layer plays a role similar to that of catalysis: that is, chloride is oxidized close to the metal surface to generate chlorine gas, and the newly generated chlorine gas returns to the metal surface. Based on this, Cl's corrosion of metal can be relatively rapid. performed at a large rate.
  • This disclosure uses quicklime and sodium metaaluminate to remove chloride ions in the ternary battery powder slurry.
  • the principle is the calcium-aluminum precipitation method, which uses aluminum salts and calcium salts to combine with free chloride ions to form a Freund's salt precipitation that absorbs chloride ions. , and finally achieve the effect of chlorine removal simply and efficiently by filtering the sediment, which reduces the interference of flammable impurity gases and the risk of explosion for the subsequent high-temperature roasting process and high-pressure leaching process, and also protects the muffle furnace and reactor from The corrosive effect of chlorine is eliminated, and the sodium metaaluminate finally obtained can be reused, which is green and environmentally friendly.
  • the process steps are simple and easy to industrialize.
  • the present disclosure uses alkali leaching to remove aluminum, and utilizes the acidity of aluminum ions to leach them into water-soluble metaaluminate ions, which can selectively remove impurities from insoluble matter such as lithium nickel cobalt manganate through solid-liquid separation.
  • roasting temperature exceeds 350°C
  • solid-solid reactions may occur between lithium nickel cobalt manganate, cobalt sulfate/nickel/manganese and carbon, and the conversion of lithium nickel cobalt manganate increases with the increase of temperature.
  • Strengthening can achieve complete destruction of the layered structure under milder conditions.
  • the ternary battery powder undergoes physical and chemical reactions to gradually generate NiSO 4 , CoSO 4 , MnSO 4 and Li 2 SO 4 , and due to the lithium salt of nickel cobalt manganese sulfate and graphite slag Due to the difference in water solubility, the sulfate solution can be leached through pure water at room temperature.
  • the present disclosure performs high-pressure leaching of roasting slag with a certain amount of ferrous sulfate, sulfuric acid, and graphite slag ingredients to deeply recover nickel, cobalt, manganese, and lithium metal.
  • SO 2 is used as a reducing agent, which increases the leaching rate of Ni/Co/Mn/Li, causing lithium nickel cobalt manganate to continuously leach out, and the reaction
  • the interface continues to shrink inward, and the participating solid substances form a loose and porous ash layer, and the graphite slag is transformed into loose and porous vermicular expanded graphite.
  • the high-pressure slag of the present disclosure is air separated to obtain vermicular expanded graphite and air separation slag that can be sold externally.
  • the air separation slag can be returned to roasting and reused to extract nickel cobalt manganese lithium, and the sodium metaaluminate solution obtained in the alkali leaching step can be It is reused in the chlorine removal and preliminary fluoride removal steps and has the characteristics of energy saving, high efficiency and environmental protection.
  • a ternary battery powder leaching method includes the following steps:
  • roasting residue After the alkali leaching residue and the roasting additive are mixed evenly, they are roasted to obtain roasting residue;
  • the ternary battery powder leaching method described in the present disclosure solves the problems of the existing ternary battery powder leaching method due to difficulty in leaching, long time consumption and low leaching rate, resulting in large leaching residue, environmental pollution and resource waste, and has the advantages The advantages of fast leaching and high leaching rate.
  • the method for preparing the reaction slurry is: adding pure water to the ternary battery powder according to preset pulping conditions, and stirring evenly; the preset pulping conditions include 3:1 to 10:1 liquid-to-solid ratio, a stirring speed of 200 to 400 rpm and a stirring time of 5 to 10 minutes.
  • a liquid-to-solid ratio of less than 3:1 will result in insufficient pulping, too rapid temperature rise of the reaction, and high slurry concentration, resulting in insufficient reaction; a liquid-to-solid ratio of greater than 10:1 will result in low slurry concentration, large volume, water consumption, etc.
  • the amount of the calcium oxide powder added before the first stirring reaction is determined based on the calcium to fluorine molar ratio of 1:1 to 2:1, and the first step is carried out.
  • the dosage of the sodium metaaluminate solution added after the stirring reaction is determined based on the mass ratio of aluminum to chloride being 1:10, and the dosage of the calcium oxide powder is determined based on the molar ratio of calcium to aluminum being 2:1. If the calcium to fluorine molar ratio is less than 1:1, the reaction will be incomplete. If the calcium to fluorine molar ratio is greater than 2:1, it will lead to material waste and increase production costs. If the aluminum to chlorine mass ratio is less than 1:10, the reaction will be incomplete.
  • the reaction will be incomplete. It will lead to material waste and increase production costs; if the calcium-aluminum molar ratio is less than 2:1, the reaction will be incomplete, and if the calcium-aluminum molar ratio is greater than 2:1, it will lead to material waste and increase production costs.
  • the first stirring reaction is performed at a stirring speed of 200 to 400 rpm for 30 minutes, and the second stirring reaction is performed at a stirring speed of 200 to 400 rpm. Stir for 60 minutes at a stirring speed of 400 rpm.
  • the preset pH value is 12.
  • the method for performing the coagulation precipitation is to add a coagulant to the dechlorinated liquid, adjust the pH to 6-8, and stir the reaction.
  • the added amount of the coagulant is 1000 mg/L
  • the coagulant includes one or more of aluminum chloride, aluminum sulfate, ferric sulfate, and ferric chloride, and the pH is adjusted
  • the reagents are calcium oxide and sulfuric acid.
  • the method of performing the stirring reaction is to first stir at a stirring speed of 200 to 400 rpm for 5 min, and then stir at a stirring speed of 40 to 70 rpm for 60 min, and the standing time is 2 h.
  • the rotation speed is greater than 70 rpm, the large flocs will be broken and the flocculation effect will be reduced.
  • the rotation speed is less than 40 rpm, the solid particles will not be fully contacted, affecting the flocculation effect.
  • the mixing time is longer than 60 minutes, it will cause The originally large flocs are redispersed, affecting the flocculation effect.
  • the dosage of the sodium hydroxide is determined based on the molar ratio of sodium to aluminum being 5:1 to 7:1. If the molar ratio of sodium to aluminum is less than 5:1, the reaction will be incomplete. If the molar ratio of sodium to aluminum is greater than 7:1, it will lead to material waste and increase production costs.
  • the alkali leaching method is to stir at a stirring speed of 200 to 400 rpm for 1 to 2 hours. If the stirring time is less than 1 hour, the reaction will be incomplete. If the reaction time is greater than 2 hours, the production efficiency will be affected and the production energy consumption will be increased.
  • the roasting additives are graphite slag powder and sodium bisulfate dihydrate, and the weight ratio of the graphite slag, sodium bisulfate dihydrate and alkali leaching residue is 0.2:1:5.
  • the roasting method is roasting at 600°C for 2 to 3 hours. If the roasting time is less than 1 hour, the reaction will be incomplete. If the roasting time is greater than 2 hours, the production efficiency will be affected and the production energy consumption will be increased.
  • the solid-liquid ratio of the roasting residue and the pure water is 3:1 to 10:1, and the water immersion method is to stir for 45 minutes at a stirring speed of 200 to 400 rpm. If the preset solid-liquid ratio is less than 3:1, it will lead to insufficient leaching, resulting in a waste of resources; if the preset solid-liquid ratio is greater than 10:1, it will cause the metal sulfate solution to have low concentration, large volume, water consumption, etc., which will lead to uneconomical results. .
  • the leaching additive is ferrous sulfate and graphite powder, and the weight ratio of the ferrous sulfate, graphite slag and water leaching slag is 0.3:1:2
  • the solid-liquid ratio of the water-leached slag powder and the pure water is 6:1-8:1
  • the amount of sulfuric acid added is 250-400g/L
  • the concentration of the sulfuric acid is 98% .
  • the high-pressure leaching method is leaching at 160-180°C for 2-3 hours, and the maximum pressure of leaching is not greater than 2MPa.
  • the method further includes the following step: using the sodium metaaluminate solution obtained by alkali leaching to prepare the chlorine removal liquid.
  • the method further includes the following steps: removing iron, extracting and synthesizing the metal sulfate solution to obtain products containing nickel, cobalt, manganese and lithium.
  • the high-pressure slag is dried and air-separated to obtain expanded graphite and air-separated slag, and the air-separated slag is used as the alkali leaching slag to prepare the alkali leaching slag powder.
  • This disclosure uses quicklime to initially remove fluoride, and then uses aluminum sulfate as a coagulant to achieve deep fluoride removal.
  • the fluoride removal effect is obvious, which can eliminate the interference of hydrogen fluoride in the roasting process and avoid the corrosive effect of fluoride ions on the high-pressure reactor.
  • Calcium fluoride generated from fluorine and calcium has a large pKsp value, and the use of calcium salts is effective in removing fluoride.
  • Quicklime is not only widely available but also cheap and has a simple treatment process. It also has the advantages of turning waste into treasure and saving resources. When quicklime is added, it To neutralize acidic wastewater.
  • aluminum sulfate can be hydrolyzed to form alum flowers with adsorption effect, so that fluoride is condensed and precipitated.
  • the introduced aluminum ions can be used to prepare sodium metaaluminate in the chlorine removal reagent.
  • Using quicklime and sodium aluminate as treatment agents to treat chloride ions can eliminate the interference caused by the conversion of chloride ions into combustible gases such as hydrogen chloride during back-end roasting and high-pressure leaching, ensuring the safety and reliability of the experiment and reducing the risk of explosion.
  • the two synergistic chlorine removal takes into account the advantages of excellent chloride ion treatment effect, environmental friendliness and low economic cost. The operation is simple and convenient for industrialization. This method has broad application prospects.
  • the layered structure of the battery powder can be completely destroyed.
  • the graphite slag and sodium bisulfate dihydrate roasting have a synergistic effect and can effectively improve the original roasting system. Thermodynamic conditions, lowering the temperature of the fired reaction, strengthening the leaching of valuable metals, changing the occurrence state of Li, Ni, Co, and Mn elements from high valence to divalent, all into water-soluble sulfuric acid complex Salt, through water leaching to achieve high leaching rates of nickel, cobalt, manganese and lithium (both greater than 99%), achieving efficient recovery of elements.
  • Adding a certain amount of graphite slag during the high-pressure leaching process can act as a reducing agent, which can effectively promote the high-pressure leaching reaction. At the same time, it is used as raw material through high-temperature and high-pressure reaction and then winnowed to obtain vermicular expanded graphite.
  • This expanded graphite is high-temperature resistant and loose. It is porous and has a large specific surface area, and can be used to prepare related refractory materials or adsorbent materials, thereby realizing the resource utilization of graphite slag.
  • ferrous sulfate as a high-pressure leaching preservative can not only protect the reactor from corrosion, but also avoid the introduction of new impurity elements, which is beneficial to back-end impurity removal.
  • the present disclosure achieves high pressure (not higher than 2Mpa) by controlling temperature rise, which is safe and controllable, allowing oxygen to participate in the leaching reaction at a higher partial pressure, and is energy-saving and environmentally friendly.
  • Figure 1 is a schematic flow chart of a ternary battery powder leaching method described in Embodiment 1;
  • Figure 2 is an SEM image of expanded graphite in Example 1;
  • Figure 3 shows the XRD patterns of raw materials, normal pressure low-acid leaching residue, and high-pressure leaching residue.
  • This embodiment provides a ternary battery powder leaching method. Please refer to Figure 1, which includes the following steps:
  • reaction slurry According to the preset pulping conditions, add 1000ml pure water to 200g ternary battery powder, stir at a stirring speed of 300rpm for 5 minutes to obtain a reaction slurry; detect the concentration of chloride ions and fluoride ions in the reaction slurry. After detection, The fluoride ion concentration in the reaction slurry is 458 mg/L, and the chloride ion concentration is 14.97 mg/L; in other embodiments, the preset pulping conditions can be a liquid-to-solid ratio of 3:1 to 10:1, 200 to 400 rpm The stirring speed and the stirring time are 5 to 10 minutes.
  • Dechlorination and preliminary fluoride removal Add calcium oxide powder to the reaction slurry according to the calcium to fluorine molar ratio of 1.5:1, stir at a stirring speed of 300 rpm for 30 minutes, and then press the aluminum to chlorine mass ratio of 1:10 and 2:1.
  • Calcium aluminum molar ratio Add calcium oxide powder and sodium metaaluminate solution to the reaction slurry, then stir for 60 minutes at a stirring speed of 300 rpm, perform solid-liquid separation to obtain the chlorine-removed liquid, and after testing, the concentration of chloride ions in the chlorine-removed liquid is 0.96 mg/L; in other embodiments, the amount of calcium oxide powder added before the first stirring reaction can be selected within the calcium to fluorine molar ratio range of 1:1 to 2:1; in In other embodiments, the dosage of the sodium metaaluminate solution added after the first stirring reaction can be 1:10 The amount of calcium oxide powder can be selected within the range of the mass ratio of aluminum to chlorine, and the amount of calcium oxide powder can be selected within the range of the molar ratio of calcium to aluminum of 2:1.
  • Defluoridation of the dechlorinated liquid Use calcium oxide powder to adjust the pH of the dechlorinated liquid to 12, and add 1g of coagulant to the dechlorinated liquid at a dosage of 1000 mg/L (the coagulant in this embodiment is preferably 18 Aluminum sulfate hydrate), and adjust the pH to 6 to 8 through calcium oxide and sulfuric acid, stir quickly at 300 rpm for 5 minutes, and then stir slowly at 60 r/min for 60 minutes to perform coagulation and precipitation.
  • the coagulant in this embodiment is preferably 18 Aluminum sulfate hydrate
  • the coagulant includes one or more of aluminum chloride, aluminum sulfate, ferric sulfate, and ferric chloride.
  • Alkali leaching According to the molar ratio of sodium to aluminum of 6:1, add sodium hydroxide to the fluorochlorine rear liquid, stir at a speed of 300 rpm for 60 minutes, and perform solid-liquid separation to obtain sodium metaaluminate solution and alkali leaching residue, in which metaaluminum
  • the sodium acid solution can be used in the aforementioned chlorine removal and preliminary fluoride removal processes.
  • the alkali leaching residue is dried in an oven and then ball milled to obtain alkali leaching residue powder; in other embodiments, the molar ratio of sodium to aluminum can be 5:1. Select within the range of ⁇ 7:1.
  • roasting Add the alkali leaching residue powder and roasting additive (the roasting additive is dried graphite powder and sodium bisulfate dihydrate) at a weight ratio of 0.2:5:1. Mix evenly in the crucible. After mixing evenly, place the crucible in a muffle furnace and roast at 600°C for 2 hours to obtain roasting slag;
  • Water leaching Add pure water to the roasting slag at a preset solid-to-liquid ratio of 5:1, stir for 45 minutes at 300 rpm, and perform solid-liquid separation to obtain a metal sulfate solution and water leaching slag, of which the metal sulfate solution is recycled
  • the raw material of nickel, cobalt, manganese and lithium, the water leaching slag is dried in an oven and then ball milled to obtain water leaching slag powder; in other embodiments, the solid-liquid ratio of the roasting slag and pure water can be between 3:1 and 10:1. Make your selection.
  • High pressure leaching The water leaching slag powder and the leaching additive (the leaching additive is ferrous sulfate) are dried at a weight ratio of 0.3:1:2 of ferrous sulfate, graphite slag and water leaching slag.
  • Graphite powder is mixed evenly and added to the autoclave, then pure water is added to the autoclave at a solid-to-liquid ratio of 8:1, sulfuric acid is added at an amount of 350g/L and stirred evenly, followed by high-pressure leaching at 160°C for 2 hours. The maximum pressure during the period is 1.76MPa. After leaching is completed and cooled to below 50°C, solid-liquid separation is performed to obtain metal sulfate solution and high-pressure slag.
  • the metal sulfate solution is used as the raw material for recovering nickel cobalt manganese lithium, and the high-pressure slag is dried and then processed. 3.97g of expanded graphite and air separation slag were obtained through air separation. The air separation slag was used to mix with alkali leaching residue to prepare alkali leaching slag powder. After testing, the specific surface area of the expanded graphite was 41g/cm 2 .
  • the SEM picture of the expanded graphite is as follows As shown in Figure 2. In other embodiments, the solid-liquid ratio of the water-leached slag powder and pure water can be selected between 6:1 and 8:1.
  • the metal sulfate solution in this embodiment can be recycled to obtain products containing nickel, cobalt, manganese, and lithium through steps such as iron removal, extraction, and synthesis to prepare ternary precursors and lithium carbonate/lithium hydroxide products; the expanded graphite can be For preparing adsorbent materials or refractory materials.
  • This embodiment provides a ternary battery powder leaching method, which includes the following steps:
  • reaction slurry According to the preset pulping conditions, add 1000ml pure water to 200g ternary battery powder, stir at a stirring speed of 300rpm for 5 minutes to obtain a reaction slurry; detect the concentration of chloride ions and fluoride ions in the reaction slurry. After detection, The fluoride ion concentration in the reaction slurry is 376 mg/L, and the chloride ion concentration is 23.51 mg/L; in other embodiments, the preset pulping conditions can be a liquid-to-solid ratio of 3:1 to 10:1, 200 to 400 rpm The stirring speed and the stirring time are 5 to 10 minutes.
  • Dechlorination and preliminary fluoride removal Add calcium oxide powder to the reaction slurry according to the molar ratio of calcium to fluorine of 1.5:1, stir at a stirring speed of 300 rpm for 30 minutes, and then use the mass ratio of aluminum to chlorine of 1:10 and calcium of 2:1.
  • Aluminum molar ratio Add calcium oxide powder and sodium metaaluminate solution to the reaction slurry, and then stir for 60 minutes at a stirring rate of 300 rpm. Perform solid-liquid separation to obtain the chlorine-removed liquid.
  • the concentration of chloride ions in the chlorine-removed liquid is: 0.87mg/L; in other embodiments, the amount of calcium oxide powder added before the first stirring reaction can be selected within the calcium to fluorine molar ratio range of 1:1 to 2:1; in other embodiments, the dosage of the sodium metaaluminate solution added after the first stirring reaction can be selected within the aluminum-chlorine mass ratio range of 1:10, and the dosage of the calcium oxide powder can be selected within the range of 2: Choose within a calcium to aluminum molar ratio range of 1.
  • Defluoridation of the dechlorinated liquid Use calcium oxide powder to adjust the pH of the dechlorinated liquid to 12, and add 1g of coagulant to the dechlorinated liquid at a dosage of 1000 mg/L (the coagulant in this embodiment is preferably 18 Aluminum sulfate hydrate), and adjust the pH to 6 to 8 through calcium oxide and sulfuric acid, stir quickly at 300 rpm for 5 minutes, and then stir slowly at 60 r/min for 60 minutes to perform coagulation and precipitation.
  • the coagulant in this embodiment is preferably 18 Aluminum sulfate hydrate
  • the coagulant includes one or more of aluminum chloride, aluminum sulfate, ferric sulfate, and ferric chloride.
  • Alkali leaching According to the molar ratio of sodium to aluminum of 6:1, add sodium hydroxide to the fluorochlorine rear liquid, stir at 300 rpm for 60 minutes, and perform solid-liquid separation to obtain sodium metaaluminate solution and alkali leaching residue, in which metaaluminate
  • the sodium solution is used in the aforementioned chlorine removal and preliminary fluoride removal process, and the alkali leaching residue is dried in an oven and then ball milled to obtain alkali leaching residue powder; in other embodiments, the molar ratio of sodium to aluminum can be between 5:1 and 7 Select within the range of :1.
  • roasting Add the alkali leaching residue powder and roasting additive (the roasting additive is dried graphite powder and sodium bisulfate dihydrate) at a weight ratio of 0.2:5:1. Mix evenly in the crucible. After mixing evenly, place the crucible in a muffle furnace and roast at 600°C for 3 hours to obtain roasting slag;
  • Water leaching Add pure water to the roasting slag at a solid-to-liquid ratio of 6:1, stir at 300 rpm for 45 minutes, and perform solid-liquid separation to obtain a metal sulfate solution and water leaching slag.
  • the metal sulfate solution is used to recover nickel and cobalt.
  • the water-leached slag is dried in an oven and then ball-milled to obtain water-leached slag powder; in other embodiments, the solid-liquid ratio of roasting slag and pure water can be selected between 3:1 and 10:1. .
  • High pressure leaching The water leaching slag powder and the leaching additive (the leaching additive is ferrous sulfate) are dried at a weight ratio of 0.3:1:2 of ferrous sulfate, graphite slag and water leaching slag.
  • Graphite powder is mixed evenly and added to the autoclave, then pure water is added to the autoclave at a solid-to-liquid ratio of 6:1, sulfuric acid is added at an amount of 350g/L and stirred evenly, followed by high-pressure leaching at 160°C for 3 hours. The maximum pressure during the period is 1.76MPa. After leaching is completed and cooled to below 50°C, solid-liquid separation is performed to obtain metal sulfate solution and high-pressure slag.
  • the metal sulfate solution is used as the raw material for recovering nickel cobalt manganese lithium, and the high-pressure slag is dried and then processed. 4.05g of expanded graphite and air separation slag were obtained through air separation. The air separation slag was used to mix with alkali leaching residue to prepare alkali leaching residue powder. After detection, the specific surface area of the expanded graphite was 45g/cm 2 ; in other embodiments, The solid-liquid ratio of water-leached slag powder and pure water can be selected between 6:1 and 8:1.
  • the metal sulfate solution in this embodiment can be recycled to obtain products containing nickel, cobalt, manganese, and lithium through steps such as iron removal, extraction, and synthesis to prepare ternary precursors and lithium carbonate/lithium hydroxide products; the expanded graphite can be For preparing adsorbent materials or refractory materials.
  • This embodiment provides a ternary battery powder leaching method, which includes the following steps:
  • reaction slurry According to the preset pulping conditions, add 1000ml pure water to 200g ternary battery powder, stir at a stirring speed of 300rpm for 5 minutes to obtain a reaction slurry; detect the concentration of chloride ions and fluoride ions in the reaction slurry. After detection, The fluoride ion concentration in the reaction slurry is 610 mg/L, and the chloride ion concentration is 14.97 mg/L; in other embodiments, the preset pulping conditions can be a liquid-to-solid ratio of 3:1 to 10:1, 200 to 400 rpm The stirring speed and the stirring time are 5 to 10 minutes.
  • Dechlorination and preliminary fluoride removal Add calcium oxide powder to the reaction slurry according to the molar ratio of calcium to fluorine of 1.5:1, stir according to the first preset stirring method at a stirring speed of 300 rpm for 30 minutes, and then press the aluminum chloride ratio of 1:10 Mass ratio, add calcium oxide powder and sodium metaaluminate solution to the reaction slurry according to the calcium to aluminum molar ratio of 2:1, then stir at a stirring rate of 300 rpm for 60 minutes, perform solid-liquid separation to obtain the chlorine-free liquid. After testing, the solution is removed.
  • the concentration of chloride ions in the post-chlorine solution is 0.92 mg/L; in other embodiments, the amount of calcium oxide powder added before the first stirring reaction can be between 1:1 and 2:1 of calcium fluoride. Select within the molar ratio range; in other embodiments, the amount of the sodium metaaluminate solution added after the first stirring reaction can be selected within the aluminum to chlorine mass ratio range of 1:10. The dosage of calcium oxide powder can be selected within the calcium to aluminum molar ratio range of 2:1.
  • Defluoridation of the dechlorinated liquid Use calcium oxide powder to adjust the pH of the dechlorinated liquid to 12, and add 1g of coagulant to the dechlorinated liquid at a dosage of 1000 mg/L (the coagulant in this embodiment is preferably 18 Aluminum sulfate hydrate), and adjust the pH to 6 to 8 through calcium oxide and sulfuric acid, stir quickly at 300 rpm for 5 minutes, and then stir slowly at 60 r/min for 60 minutes to perform coagulation and precipitation.
  • the coagulant in this embodiment is preferably 18 Aluminum sulfate hydrate
  • the coagulant includes one or more of aluminum chloride, aluminum sulfate, ferric sulfate, and ferric chloride.
  • Alkali leaching According to the molar ratio of sodium to aluminum of 6:1, add sodium hydroxide to the fluorochlorine rear liquid, stir at 300 rpm for 60 minutes, and perform solid-liquid separation to obtain sodium metaaluminate solution and alkali leaching residue, in which metaaluminate
  • the sodium solution can be used in the aforementioned chlorine removal and preliminary fluoride removal processes.
  • the alkali leaching residue is dried in an oven and then ball milled to obtain alkali leaching residue powder; in other embodiments, the molar ratio of sodium to aluminum can be between 5:1 and Choose from a range of 7:1.
  • roasting Add the alkali leaching residue powder and roasting additive (the roasting additive is dried graphite powder and sodium bisulfate dihydrate) at a weight ratio of 0.2:5:1. Mix evenly in the crucible. After mixing evenly, place the crucible in the muffle furnace and roast at 600°C for 2 hours according to the preset roasting method to obtain roasting slag;
  • the roasting additive is dried graphite powder and sodium bisulfate dihydrate
  • Water leaching Add pure water to the roasting slag at a preset solid-to-liquid ratio of 5:1, stir for 45 minutes at 300 rpm, and perform solid-liquid separation to obtain a metal sulfate solution and water leaching slag, of which the metal sulfate solution is recycled
  • the raw material of nickel, cobalt, manganese and lithium, the water leaching slag is dried in an oven and then ball milled to obtain water leaching slag powder; in other embodiments, the solid-liquid ratio of the roasting slag and pure water can be between 3:1 and 10:1. Make your selection.
  • High pressure leaching The water leaching slag powder and the leaching additive (the leaching additive is ferrous sulfate) are dried at a weight ratio of 0.3:1:2 of ferrous sulfate, graphite slag and water leaching slag.
  • Graphite powder is mixed evenly and added to the autoclave, then pure water is added to the autoclave at a solid-to-liquid ratio of 7:1, sulfuric acid is added at an amount of 350g/L and stirred evenly, followed by high-pressure leaching at 160°C for 3 hours. The maximum pressure during the period is 1.76MPa. After leaching is completed and cooled to below 50°C, solid-liquid separation is performed to obtain metal sulfate solution and high-pressure slag.
  • the metal sulfate solution is used as the raw material for recovering nickel cobalt manganese lithium, and the high-pressure slag is dried and then processed. 4.05g of expanded graphite and air separation residue were obtained through air separation. The air separation residue was used to mix with alkali leaching residue to prepare alkali leaching residue powder. After testing, the specific surface area of the expanded graphite was 39 g/cm 2 . In other embodiments, the solid-liquid ratio of the water-leached slag powder and pure water can be selected between 6:1 and 8:1.
  • the metal sulfate solution in this embodiment can be recycled to obtain products containing nickel, cobalt, manganese, and lithium through steps such as iron removal, extraction, and synthesis to prepare ternary precursors and lithium carbonate/lithium hydroxide products; the expanded graphite can be For preparing adsorbent materials or refractory materials.
  • Example 1 The difference between this comparative example and Example 1 is that only sodium bisulfate dihydrate is added during roasting, and graphite slag is not added to assist the roasting. Finally, 2.13 g of expanded graphite was obtained, and the specific surface area of the expanded graphite was 31 g/cm 2 .
  • Example 1 The difference between this comparative example and Example 1 is that during water immersion, pure water is not used, but 350g/L sulfuric acid is used, and the stirring time is changed to 2 hours. Finally, 1.98 g of expanded graphite was obtained, and the specific surface area of the expanded graphite was 25 g/cm 2 .
  • Example 1 The difference between this comparative example and Example 1 is that the high-pressure leaching step is changed to normal pressure leaching, 350g/L sulfuric acid is added at a solid-liquid ratio of 5:1 for leaching, and the leaching time is 2 hours. In the end, expanded graphite was not obtained.
  • Example 1 and Comparative Example 2 By comparing the data of Example 1 and Comparative Example 1, it can be seen that if graphite slag is not added during roasting, the leaching rate of nickel cobalt manganese lithium decreases, the amount of slag increases, there are fewer expanded graphite products and the specific surface area decreases; by comparison It can be seen from the data of Example 1 and Comparative Example 2 that if pure water is not used during water immersion, but 350g/L sulfuric acid is used, the leaching rate of nickel cobalt manganese lithium will slightly decrease, the slag amount will double, and the expanded graphite product will be less and less than The surface area is significantly reduced; by comparing the data of Example 1 and Comparative Example 3, it can be seen that if the high-pressure leaching step is changed to normal pressure leaching, the leaching rate of nickel cobalt manganese lithium is significantly reduced, the slag amount is significantly doubled, and there is no expanded graphite product, and please refer to Figure 3.
  • the present disclosure provides an efficient method for leaching nickel, cobalt, manganese and lithium.
  • the leaching rate is close to 100%, and the amount of slag produced is small and the specific surface area of expanded graphite is relatively large. High, excellent performance.

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Abstract

一种三元电池粉浸出方法，包括以下步骤：配制反应浆液、除氯、除氟、碱浸、焙烧、水浸和高压浸出；其中除氯主要是通过加入氧化钙粉末和偏铝酸钠溶液实现；除氟主要是通过进行混凝沉淀实现；碱浸是指向除氟氯后液中加入氢氧化钠进行浸出；焙烧是指将碱浸渣粉末与焙烧添加剂混合均匀后进行焙烧；水浸是通过往焙烧渣中加入纯水进行浸出；高压浸出是指将水浸渣粉末与浸出添加剂混合均匀后加入高压釜内，再加入纯水和硫酸并搅拌均匀，进行高压浸出。

Description

一种三元电池粉浸出方法 技术领域
本公开涉及资源回收领域，特别是涉及一种三元电池粉浸出方法。
背景技术
随着三元锂离子电池消费市场的扩大，三元锂电池制造企业日益兴盛起来，然而电池制作中往往会产生大量失效的电池正极浆料。这些正极浆料含有大量的镍钴锰锂金属资源与石墨渣，若直接弃置不仅会造成环境污染，而且会造成资源浪费，因此研究如何将正极浆料中的有价金属进行高效回收利用的技术迫在眉睫，一批废旧电池回收企业也逐渐兴起。
现有的工业废旧电池回收利用技术包括硫酸常压浸出法和焙烧-水浸联用法。硫酸常压浸出法是通过多级低酸浸出与高酸浸出联用的方法还原浸出三元电池粉中的镍钴锰锂金属为金属硫酸盐溶液，其耗费时间漫长、硫酸消耗量巨大，并且由于其金属浸出率不高，浸出后会产生大量夹杂有镍钴锰锂金属的石墨渣堆置，造成环境污染和资源浪费；焙烧-水浸联用法是通过湿法火法混合冶金来提取镍钴锰锂金属，但是此方法会产生大量夹杂镍钴锰锂金属的焙烧渣堆置，同样存在污染环境、浸出率不高的问题。
发明内容
基于此，本公开的目的在于，提供一种三元电池粉浸出方法，解决了现有三元电池粉浸出方法因浸出难、耗时长、浸出率低而导致的浸出渣大、造成环境污染和资源浪费的问题，具有浸出快、浸出率高的优点。
三元电池粉携带有部分溶剂中的Cl，高温及高压时Cl的渗透能力很大，它可以通过金属表面的氧化膜直接同金属反应生成相应的氯化物并蒸发，蒸发的氯化物会与氧气反应生成氯气和相应的氧化物。在这个过程中氧化层起着类似于催化作用的作用：即氯化物在靠近金属表面的地方被氧化生成氯气，新生成的氯气又重新返回金属表面，基于此，Cl对金属的腐蚀能以较大的速率进行。本公开通过生石灰和偏铝酸钠去除三元电池粉浆液中的氯离子，其原理为钙铝沉淀法，即利用铝盐与钙盐与游离态氯离子结合，形成吸附氯离子的弗氏盐沉淀，最后通过滤沉淀物的方式简单高效地达到除氯的效果，为后续的高温焙烧过程和高压浸出过程减免了可燃性杂质气体的干扰、降低爆炸风险，还使马弗炉和反应釜免受氯的腐蚀作用，且其最终得到的偏铝酸钠还可以进行回用，绿色环保。其具体反应原理为：
CaO+H2O＝Ca(OH)2；
AlO2-+2H2O＝Al(OH)3+OH-；
4Ca(OH)2+2Al(OH)3+2Cl-＝2OH-+Ca4Al2Cl2(OH)12↓。
虽然钛表面会自动形成一层稳定性好、结合力强的氧化膜，但是在还原性酸溶液中，氟化物很容易与氢离子结合形成氟化氢，优先吸附在钛材氧化膜上，排挤氧原子导致钛合金表面的钝化膜形成可溶性氟化物而发生腐蚀，遭到破坏。本公开通过添加生石灰和十八水硫酸铝混凝剂实现两步除F，是为了保护高压浸出的钛材反应釜，其中生石灰主要起到初步除F的作用，生石灰初步除氟原理为：2F-+Ca(OH)2＝CaF2↓+2OH-；而十八水硫酸铝作为混凝剂可在混凝时产生轻细絮体矾花提产生混凝作用高沉降速度，提高溶液中氟离子的脱除率，同时硫酸铝中铝离子与酸中氟络合生成稳定性很强的氟铝阴配离子【AlFn3-n】，主要形式为AlF2+～AlF3；由于生石灰和十八水硫酸铝价格低廉易购买、该工艺步骤简单且易工业化，在生石灰与十八水硫酸的共同作用下，游离氟可以降低到生产所允许的浓度，壳达到防止钛材反应釜腐蚀的目的；额外地，通过采用十八水硫酸铝引入铝离子便于制备偏铝酸钠以便回用除氯。
在高压浸出的过程中，铝的存在会导致氢气的产生，提高高压浸出过程的爆炸风险。本公开通过采用碱浸的方式除铝，利用铝离子的酸性将其浸出为水溶性的偏铝酸根离子，与镍钴锰酸锂等不溶物通过固液分离可以选择性除杂，其原理为：Al3++4OH-＝AlO2-+2H2O。
本公开以一定比例的石墨渣与焙烧添加剂进行高温协同焙烧得到金属硫酸盐溶液和焙烧渣。焙烧具体反应机理如下：
12LiNi0.8Coo.1Mn0.1O2+36NaHSO4·2H2O+3C＝6Li2SO4+18Na2SO4+9.6NiSO4+1.2CoSO4+1.2MnSO4+3CO2+90H2O。其中二水硫酸氢钠在高温焙烧中反应分解为焦硫酸钠参与反应，从而破坏镍钴锰酸锂的层状结构；石墨渣通过引入碳元素加速层状结构的分解与转化，有效改善硫化焙烧反应的热力学条件，强化浸出过程。在该本公开的焙烧体系中，焙烧温度超过350℃时，镍钴锰酸锂、硫酸钴/镍/锰与碳可发生固固反应，且镍钴锰酸锂的转化随着温度的升高而加强，可实现在较温和的条件下完成层状结构的完全破坏。本公开的硫碳协同焙烧体系在焙烧升温过程中，使三元电池粉发生物理化学反应逐步生成NiSO4、CoSO4、MnSO4和Li2SO4，并且由于硫酸镍钴锰锂盐与石墨渣之间的水溶性差异，在室温下即可通过纯水浸出硫酸盐溶液。
本公开通过焙烧渣与一定量的硫酸亚铁、硫酸、石墨渣配料进行高压浸出从而深度回收镍钴锰锂金属。高压浸出是通过控制高温条件实现加压效果，加速浸出过程，达到深度浸出镍钴锰锂的效果；其原理为：
C+2H2SO4＝CO2+2SO2+2H2O，
2LiNi0.8Co0.1Mn0.1O2+4SO2+H2O＝Li2SO4+1.6NiSO4+0.2CoSO4+0.2MnSO4+H2SO4；
其中，SO2作还原剂，增加了Ni/Co/Mn/Li的浸出率，使得镍钴锰酸锂不断浸出，反应 界面不断向内收缩，参与的固体物质形成疏松多孔的灰层，石墨渣转化为疏松多孔的蠕状膨胀石墨，因此，在高压浸出的过程中，随着石墨增加，金属浸出率增加；而硫酸亚铁主要是作为防腐剂加入，可以在不引入新的杂质离子的同时保护钛材反应釜及搅拌装置免受高压下的腐蚀作用；在硫酸亚铁的作用下钛材反应釜表面会在原有的钛材氧化膜上形成一层新的铁氧化物保护膜，通过该保护膜实现防腐效果，该保护膜主要为无定形或微晶的水合氧化铁FeOOH，可能是通过以下反应进行：4Fe2++8OH-+O2＝4FeOOH+H2O；需要注意的是，成膜前钛材反应釜应清洗干净，表面较脏时可用酸洗。
本公开的高压渣进行风选分离得到可外售的蠕状膨胀石墨与风选渣，风选渣可返回焙烧再次回用提取镍钴锰锂，并且碱浸步骤得到的偏铝酸钠溶液可回用于除氯及初步除氟步骤中，具有节能高效环保的特点。
一种三元电池粉浸出方法，包括以下步骤：
将所述三元电池粉配制成反应浆液；
向所述反应浆液中加入氧化钙粉末，搅拌反应，反应完全后，再加入偏铝酸钠溶液和氧化钙粉末，搅拌反应，反应完全后，进行固液分离得到除氯后液；
调节所述除氯后液的pH至预设pH值后，进行混凝沉淀，混凝沉淀后静置两小时，进行固液分离得到除氟氯后液和污泥；
向所述除氟氯后液中加入氢氧化钠，碱浸，进行固液分离得到偏铝酸钠溶液与碱浸渣，将所述碱浸渣烘干粉碎得到碱浸渣粉末；
将所述碱浸渣与焙烧添加剂混合均匀后，焙烧，得到焙烧渣；
向所述焙烧渣中加入纯水，水浸，进行固液分离得到金属硫酸盐溶液和水浸渣，将所述水浸渣烘干粉碎得到水浸渣粉末；
将所述水浸渣与浸出添加剂混合均匀后，加入纯水和硫酸，搅拌，高压浸出，进行固液分离得到金属硫酸盐溶液和高压渣。
本公开所述的一种三元电池粉浸出方法，解决了现有三元电池粉浸出方法因浸出难、耗时长、浸出率低而导致的浸出渣大、造成环境污染和资源浪费的问题，具有浸出快、浸出率高的优点。
在一实施例，配制所述反应浆液的方法为：按预设制浆条件向所述三元电池粉中加入纯水，搅拌均匀；所述预设制浆条件包括3:1～10:1的液固比，200～400rpm的搅拌速度和5～10min的搅拌时间。液固比小于3:1会导致制浆不充分，反应升温过快，浆液浓度高等导致反应不充分；液固比大于10:1会导致浆液浓度低，体积量大，耗水占体积等导致不经济；转速高于400rpm容易导致物料的飞溅和转轴的磨损，转速低于200rpm容易导致溶液混合不均匀一 级反应不充分。
在一实施例，制备所述除氯后液时，进行第一次所述搅拌反应前加入的所述氧化钙粉末的用量依据钙氟摩尔比为1:1～2:1确定，进行第一次所述搅拌反应后加入的所述偏铝酸钠溶液的用量依据铝氯质量比为1:10确定、所述氧化钙粉末的用量依据钙铝摩尔比为2:1确定。钙氟摩尔比小于1：1则反应不完全，钙氟摩尔大于2:1会导致物料浪费、提高生产成本；铝氯质量比小于1：10则反应不完全，铝氯质量比大于1：10会导致物料浪费、提高生产成本；钙铝摩尔比小于2：1则反应不完全，钙铝摩尔比大于2:1会导致物料浪费、提高生产成本。
在一实施例，制备所述除氯后液时，进行第一次所述搅拌反应的方法为在200～400rpm的搅拌速度下搅拌30min，进行第二次所述搅拌反应的方法为在200～400rpm的搅拌速度下搅拌60min。
在一实施例，调节所述pH时，所述预设pH值为12。
在一实施例，进行所述混凝沉淀的方法为向所述除氯后液中加入混凝剂并调节pH至6～8，搅拌反应。
在一实施例，所述混凝剂的加入量为1000mg/L，所述混凝剂包括氯化铝、硫酸铝、硫酸铁、氯化铁中的一种或几种，调节所述pH的试剂为氧化钙和硫酸。
在一实施例，进行所述搅拌反应的方法为先在200～400rpm的搅拌速度下搅拌5min，再在40～70rpm的搅拌速度下搅拌60min，所述静置的时间为2h。先进行快速的搅拌，使絮凝颗粒初步形成，进行慢速搅拌，使絮凝颗粒不断结大形成絮凝体。在慢速搅拌的过程中，若转速大于70rpm会导致结大的絮凝体被打碎，降低絮凝效果，若转速小于40rpm会导致固体颗粒不能充分接触，影响絮凝效果，若搅拌时间长于60min会导致原本结大的絮凝体重新分散，影响絮凝效果。
在一实施例，加入所述氢氧化钠时，所述氢氧化钠的用量依据钠铝摩尔比为5:1～7:1确定。钠铝摩尔比小于5：1则反应不完全，钠铝摩尔比大于7:1会导致物料浪费、提高生产成本。
在一实施例，进行所述碱浸的方法为在200～400rpm的搅拌速度下搅拌1～2h。搅拌时间小于1h则反应不完全，反应时间大于2h则影响生产效率，增加生产能耗。
在一实施例，所述焙烧添加剂为石墨渣粉末和二水硫酸氢钠，所述石墨渣、二水硫酸氢钠与碱浸渣的重量比为0.2:1:5。
在一实施例，进行所述焙烧的方法为在600℃下焙烧2～3h。焙烧时间小于1h则反应不完全，焙烧时间大于2h则影响生产效率，增加生产能耗。
在一实施例，所述焙烧渣与所述纯水的固液比为3:1～10:1，进行所述水浸的方法为在200～400rpm的搅拌速度下搅拌45min。预设固液比小于3:1会导致浸出不充分，导致资源的浪费；预设固液比大于10:1会导致金属硫酸盐溶液浓度低，体积量大，耗水占体积等导致不经济。
在一实施例，所述浸出添加剂为硫酸亚铁和石墨粉末，所述硫酸亚铁、石墨渣与水浸渣的重量比为0.3:1:2
在一实施例，所述水浸渣粉末与所述纯水的固液比为6:1～8:1，所述硫酸的加入量为250～400g/L，所述硫酸的浓度为98％。
在一实施例，进行所述高压浸出的方法为在160～180℃下浸出2～3h，浸出的最大压力不大于2MPa。
在一实施例，还包括以下步骤：将所述碱浸得到的偏铝酸钠溶液用于制备所述除氯后液。
在一实施例，还包括以下步骤：对所述金属硫酸盐溶液进行除铁、萃取和合成，得到含镍、钴、锰、锂的产品。
在一实施例，将所述高压渣烘干并进行风选得到膨胀石墨和风选渣，将所述风选渣作为所述碱浸渣用于制备所述碱浸渣粉末。
本公开的有益效果在于：
1.本公开利用生石灰初步除氟，再用硫酸铝作为混凝剂实现深度除氟，除氟效果明显，可排除氟化氢在焙烧过程中的干扰，避免氟离子对高压反应釜的腐蚀作用。氟和钙生成的氟化钙pKsp值较大，用钙盐除氟效果佳，生石灰不仅来源广泛而且价格低廉，处理工艺简单，并具备变废为宝、节约资源等优点，加入生石灰的同时起到了中和酸性废水的作用。而硫酸铝作为混凝剂可发生水解后生成具有吸附作用的矾花，从而氟化物被凝聚沉淀，引进的铝离子可用于制备除氯试剂中的偏铝酸钠。
2、以生石灰和铝酸钠作为处理剂处理氯离子，可排除后端焙烧和高压浸出时氯离子转化为氯化氢等可燃气带来的干扰，保证实验安全可靠降低爆炸风险。两者协同除氯兼顾氯离子处理效果优良、环境友好且经济成本低的优点，操作简单便于工业化，该方法具有广阔的应用前景。
3、通过碱浸实现选择性除铝，同时制备的偏铝酸钠溶液可作为除氯试剂回用。镍钴锰酸锂层状结构尚未被破坏，以粉末的形式赋存在碱浸渣中，有利于下一步焙烧处理。
4、通过二水硫酸氢钠作为焙烧添加剂，通过固定比例添加石墨渣配料辅助焙烧可完全破坏电池粉的层状结构，石墨渣与二水硫酸氢钠焙烧起协同作用，能够有效改善原焙烧体系 的热力学条件，降低被烧反应温度，强化有价金属的浸出，使Li、Ni、Co、Mn元素的赋存状态发生转变，由高价转变为2价，全部转变为可溶于水的硫酸复式盐，通过水浸达到镍钴锰锂的高浸出率(均大于99％)，实现元素的高效回收。
5、高压浸出过程添加定量的石墨渣可充当还原剂的作用，可有效推动高压浸出反应，同时以其为原料通过高温高压反应后风选得到了蠕状膨胀石墨，该膨胀石墨耐高温且疏松多孔、比表面积大，可用于制备相关耐火材料或吸附材料，从而实现石墨渣的资源化。
6、以硫酸亚铁作为高压浸出防腐剂，不仅可以保护反应釜免受腐蚀，而且避免引入新的杂质元素，利于后端除杂。
7、通过高压浸出实现了焙烧渣中镍钴锰锂的深度回收，对比常规酸浸不仅节省时间而且提高效能，实现变废为宝。本公开通过控制升温实现高压(不高于2Mpa)，安全可控，使氧气在较高的分压下参与浸出反应，节能环保。
为了更好地理解和实施，下面结合附图详细说明本公开。
附图说明
图1为实施例1所述的一种三元电池粉浸出方法流程示意图；
图2为实施例1的膨胀石墨的SEM图；
图3为原料、常压低酸浸出渣、高压浸出渣的XRD图。
具体实施方式
实施例1
本实施例提供一种三元电池粉浸出方法，请参阅图1，包括以下步骤：
配制反应浆液：按预设制浆条件，向200g三元电池粉中加入1000ml纯水，以300rpm的搅拌速度搅拌5min，得到反应浆液；检测反应浆液中氯离子和氟离子的浓度，经检测，反应浆液中氟离子浓度为458mg/L，氯离子浓度为14.97mg/L；在其他实施例中，所述预设制浆条件可以为3:1～10:1的液固比，200～400rpm的搅拌速度和5～10min的搅拌时间。
除氯及初步除氟：按1.5:1的钙氟摩尔比向反应浆液中加入氧化钙粉末，以300rpm的搅拌速率搅拌30min，接着再按1:10的铝氯质量比、按2:1的钙铝摩尔比向反应浆液中加入氧化钙粉末和偏铝酸钠溶液，接着以300rpm的搅拌速率搅拌60min，进行固液分离得到除氯后液，经检测，除氯后液中氯离子的浓度为0.96mg/L；在其他实施例中，进行第一次所述搅拌反应前加入的所述氧化钙粉末的用量可以在1:1～2:1的钙氟摩尔比范围内进行选择；在其他实施例中，进行第一次所述搅拌反应后加入的所述偏铝酸钠溶液的用量可以在1:10 的铝氯质量比范围内进行选择、所述氧化钙粉末的用量可以在2:1的钙铝摩尔比范围内进行选择。
除氯后液除氟：用氧化钙粉末调节除氯后液的pH至12，按1000mg/L的用量向除氯后液中加入1g混凝剂(本实施例的混凝剂优选为十八水硫酸铝)，并通过氧化钙和硫酸调节pH至6～8，以300rpm的转速快速搅拌5分钟，再以60r/min的转速慢速搅拌60分钟，即进行混凝沉淀，混凝沉淀后静置两小时，进行固液分离得到除氟氯后液和污泥，污泥进行填埋处理；经检测，除氟氯后液中氟离子的浓度为0.62mg/L，铝离子浓度为23g/L；在其他实施例中，混凝剂包括氯化铝、硫酸铝、硫酸铁、氯化铁中的一种或几种。
碱浸：按6:1的钠铝摩尔比，向氟氯后液中加入氢氧化钠，按以300rpm的转速搅拌60min，进行固液分离得到偏铝酸钠溶液与碱浸渣，其中偏铝酸钠溶液可以应用于前述的除氯及初步除氟的过程中，碱浸渣通过烘箱烘干后进行球磨，得到碱浸渣粉末；在其他实施例中，钠铝摩尔比可以在5:1～7:1的范围内进行选择。
焙烧：以0.2:5:1的石墨渣、二水硫酸氢钠与碱浸渣的重量比将碱浸渣粉末和焙烧添加剂(焙烧添加剂为烘干后的石墨粉末和二水硫酸氢钠)加入至坩埚中混合均匀，混合均匀后将坩埚置于马弗炉内，在600℃中焙烧2h，得到焙烧渣；
水浸：按5：1的预设固液比往焙烧渣中加入纯水，以300rpm的转速搅拌45min，进行固液分离后得到金属硫酸盐溶液和水浸渣，其中金属硫酸盐溶液作为回收镍钴锰锂的原料，水浸渣通过烘箱烘干后进行球磨，得到水浸渣粉末；在其他实施例中，焙烧渣和纯水的固液比可以在3:1～10:1之间进行选择。
高压浸出：以0.3:1:2的硫酸亚铁、石墨渣与水浸渣的重量比为0.3:1:2将所述水浸渣粉末和浸出添加剂(浸出添加剂为硫酸亚铁和烘干后的石墨粉末)混合均匀后加入高压釜内，再向高压釜内按8:1的固液比加入纯水、按350g/L的加入量加入硫酸并搅拌均匀，在160℃下高压浸出2h，期间最高压力为1.76MPa，浸出完成并冷却至50℃以下后，进行固液分离得到金属硫酸盐溶液和高压渣，其中金属硫酸盐溶液作为回收镍钴锰锂的原料，高压渣烘干后进行风选得到3.97g膨胀石墨和风选渣，风选渣用于与碱浸渣混合制备碱浸渣粉末，经检测，所述膨胀石墨的比表面积面积为41g/cm2，膨胀石墨的SEM图如图2所示。在其他实施例中，水浸渣粉末与纯水的固液比可以在6:1～8:1之间进行选择。
本实施例的金属硫酸盐溶液可以通过除铁、萃取和合成等步骤回收得到含镍、钴、锰、锂的产品，制备三元前驱体与碳酸锂/氢氧化锂产品；所述膨胀石墨可用于制备吸附材料或耐火材料。
实施例2
本实施例提供一种三元电池粉浸出方法，包括以下步骤：
配制反应浆液：按预设制浆条件，向200g三元电池粉中加入1000ml纯水，以300rpm的搅拌速度搅拌5min，得到反应浆液；检测反应浆液中氯离子和氟离子的浓度，经检测，反应浆液中氟离子浓度为376mg/L，氯离子浓度为23.51mg/L；在其他实施例中，所述预设制浆条件可以为3:1～10:1的液固比，200～400rpm的搅拌速度和5～10min的搅拌时间。
除氯及初步除氟：按1.5:1的钙氟摩尔比向反应浆液中加入氧化钙粉末，以300rpm的搅拌速率搅拌30min，接着再按1:10的铝氯质量比、2:1的钙铝摩尔比向反应浆液中加入氧化钙粉末和偏铝酸钠溶液，接着以300rpm的搅拌速率搅拌60min，进行固液分离得到除氯后液，经检测，除氯后液中氯离子的浓度为0.87mg/L；在其他实施例中，进行第一次所述搅拌反应前加入的所述氧化钙粉末的用量可以在1:1～2:1的钙氟摩尔比范围内进行选择；在其他实施例中，进行第一次所述搅拌反应后加入的所述偏铝酸钠溶液的用量可以在1:10的铝氯质量比范围内进行选择、所述氧化钙粉末的用量可以在2:1的钙铝摩尔比范围内进行选择。
除氯后液除氟：用氧化钙粉末调节除氯后液的pH至12，按1000mg/L的用量向除氯后液中加入1g混凝剂(本实施例的混凝剂优选为十八水硫酸铝)，并通过氧化钙和硫酸调节pH至6～8，以300rpm的转速快速搅拌5分钟，再以60r/min的转速慢速搅拌60分钟，即进行混凝沉淀，混凝沉淀后静置两小时，进行固液分离得到除氟氯后液和污泥，污泥进行填埋处理；经检测，除氟氯后液中氟离子的浓度为0.91mg/L，铝离子浓度为23g/L；在其他实施例中，混凝剂包括氯化铝、硫酸铝、硫酸铁、氯化铁中的一种或几种。
碱浸：按6:1的钠铝摩尔比，向氟氯后液中加入氢氧化钠，以300rpm的转速搅拌60min，进行固液分离得到偏铝酸钠溶液与碱浸渣，其中偏铝酸钠溶液应用于前述的除氯及初步除氟的过程中，碱浸渣通过烘箱烘干后进行球磨，得到碱浸渣粉末；在其他实施例中，钠铝摩尔比可以在5:1～7:1的范围内进行选择。
焙烧：以0.2:5:1的石墨渣、二水硫酸氢钠与碱浸渣的重量比将碱浸渣粉末和焙烧添加剂(焙烧添加剂为烘干后的石墨粉末和二水硫酸氢钠)加入至坩埚中混合均匀，混合均匀后将坩埚置于马弗炉内，在600℃中焙烧3h，得到焙烧渣；
水浸：按6：1的固液比往焙烧渣中加入纯水，以300rpm的转速搅拌45min，进行固液分离后得到金属硫酸盐溶液和水浸渣，其中金属硫酸盐溶液作为回收镍钴锰锂的原料，水浸渣通过烘箱烘干后进行球磨，得到水浸渣粉末；在其他实施例中，焙烧渣和纯水的固液比可以在3:1～10:1之间进行选择。
高压浸出：以0.3:1:2的硫酸亚铁、石墨渣与水浸渣的重量比为0.3:1:2将所述水浸渣粉末和浸出添加剂(浸出添加剂为硫酸亚铁和烘干后的石墨粉末)混合均匀后加入高压釜内，再向高压釜内按6:1的固液比加入纯水、按350g/L的加入量加入硫酸并搅拌均匀，在160℃下高压浸出3h，期间最高压力为1.76MPa，浸出完成并冷却至50℃以下后，进行固液分离得到金属硫酸盐溶液和高压渣，其中金属硫酸盐溶液作为回收镍钴锰锂的原料，高压渣烘干后进行风选得到4.05g膨胀石墨和风选渣，风选渣用于与碱浸渣混合制备碱浸渣粉末，经检测，所述膨胀石墨的比表面积面积为45g/cm2；在其他实施例中，水浸渣粉末与纯水的固液比可以在6:1～8:1之间进行选择。
本实施例的金属硫酸盐溶液可以通过除铁、萃取和合成等步骤回收得到含镍、钴、锰、锂的产品，制备三元前驱体与碳酸锂/氢氧化锂产品；所述膨胀石墨可用于制备吸附材料或耐火材料。
实施例3
本实施例提供一种三元电池粉浸出方法，包括以下步骤：
配制反应浆液：按预设制浆条件，向200g三元电池粉中加入1000ml纯水，以300rpm的搅拌速度搅拌5min，得到反应浆液；检测反应浆液中氯离子和氟离子的浓度，经检测，反应浆液中氟离子浓度为610mg/L，氯离子浓度为14.97mg/L；在其他实施例中，所述预设制浆条件可以为3:1～10:1的液固比，200～400rpm的搅拌速度和5～10min的搅拌时间。
除氯及初步除氟：按1.5:1的钙氟摩尔比向反应浆液中加入氧化钙粉末，按第一预设搅拌方法，以300rpm的搅拌速率搅拌30min，接着再按1:10的铝氯质量比、按2:1的钙铝摩尔比向反应浆液中加入氧化钙粉末和偏铝酸钠溶液，接着以300rpm的搅拌速率搅拌60min，进行固液分离得到除氯后液，经检测，除氯后液中氯离子的浓度为0.92mg/L；在其他实施例中，进行第一次所述搅拌反应前加入的所述氧化钙粉末的用量可以在1:1～2:1的钙氟摩尔比范围内进行选择；在其他实施例中，进行第一次所述搅拌反应后加入的所述偏铝酸钠溶液的用量可以在1:10的铝氯质量比范围内进行选择、所述氧化钙粉末的用量可以在2:1的钙铝摩尔比范围内进行选择。
除氯后液除氟：用氧化钙粉末调节除氯后液的pH至12，按1000mg/L的用量向除氯后液中加入1g混凝剂(本实施例的混凝剂优选为十八水硫酸铝)，并通过氧化钙和硫酸调节pH至6～8，以300rpm的转速快速搅拌5分钟，再以60r/min的转速慢速搅拌60分钟，即进行混凝沉淀，混凝沉淀后静置两小时，进行固液分离得到除氟氯后液和污泥，污泥进行填埋处理；经检测，除氟氯后液中氟离子的浓度为0.77mg/L，铝离子浓度为23g/L；在其他实施例 中，混凝剂包括氯化铝、硫酸铝、硫酸铁、氯化铁中的一种或几种。
碱浸：按6:1的钠铝摩尔比，向氟氯后液中加入氢氧化钠，以300rpm的转速搅拌60min，进行固液分离得到偏铝酸钠溶液与碱浸渣，其中偏铝酸钠溶液可以应用于前述的除氯及初步除氟的过程中，碱浸渣通过烘箱烘干后进行球磨，得到碱浸渣粉末；在其他实施例中，钠铝摩尔比可以在5:1～7:1的范围内进行选择。
焙烧：以0.2:5:1的石墨渣、二水硫酸氢钠与碱浸渣的重量比将碱浸渣粉末和焙烧添加剂(焙烧添加剂为烘干后的石墨粉末和二水硫酸氢钠)加入至坩埚中混合均匀，混合均匀后将坩埚置于马弗炉内，按预设焙烧方法，在600℃中焙烧2h，得到焙烧渣；
水浸：按5：1的预设固液比往焙烧渣中加入纯水，以300rpm的转速搅拌45min，进行固液分离后得到金属硫酸盐溶液和水浸渣，其中金属硫酸盐溶液作为回收镍钴锰锂的原料，水浸渣通过烘箱烘干后进行球磨，得到水浸渣粉末；在其他实施例中，焙烧渣和纯水的固液比可以在3:1～10:1之间进行选择。
高压浸出：以0.3:1:2的硫酸亚铁、石墨渣与水浸渣的重量比为0.3:1:2将所述水浸渣粉末和浸出添加剂(浸出添加剂为硫酸亚铁和烘干后的石墨粉末)混合均匀后加入高压釜内，再向高压釜内按7:1的固液比加入纯水、按350g/L的加入量加入硫酸并搅拌均匀，在160℃下高压浸出3h，期间最高压力为1.76MPa，浸出完成并冷却至50℃以下后，进行固液分离得到金属硫酸盐溶液和高压渣，其中金属硫酸盐溶液作为回收镍钴锰锂的原料，高压渣烘干后进行风选得到4.05g膨胀石墨和风选渣，风选渣用于与碱浸渣混合制备碱浸渣粉末，经检测，所述膨胀石墨的比表面积面积为39g/cm2。在其他实施例中，水浸渣粉末与纯水的固液比可以在6:1～8:1之间进行选择。
本实施例的金属硫酸盐溶液可以通过除铁、萃取和合成等步骤回收得到含镍、钴、锰、锂的产品，制备三元前驱体与碳酸锂/氢氧化锂产品；所述膨胀石墨可用于制备吸附材料或耐火材料。
对比例1
本对比例与实施例1的区别在于：焙烧时只添加二水硫酸氢钠，不添加石墨渣辅助焙烧。最终得到膨胀石墨2.13g，膨胀石墨的比表面积面积为31g/cm2。
对比例2
本对比例与实施例1的区别在于：水浸时，不采用纯水，而采用350g/L的硫酸，搅拌时间改为2h。最终得到膨胀石墨1.98g，膨胀石墨的比表面积面积为25g/cm2。
对比例3
本对比例与实施例1的区别在于：高压浸出步骤改为常压浸出，以5:1的固液比加入350g/L的硫酸进行浸出，浸出时间为2h。最终未得到膨胀石墨。
采用相同的回收方法对实施例1-3以及对比例1-3所得的金属硫酸盐溶液中的镍、钴、锰和锂进行回收，回收结果如下表1所示。
表1实施例1-3以及对比例1-3的镍、钴、锰、锂的回收结果
通过对比实施例1和对比例1的数据可知，若焙烧时不加石墨渣辅助，镍钴锰锂浸出率有所下降，渣量有所增加，膨胀石墨产品更少且比表面积降低；通过对比实施例1和对比例2的数据可知，若水浸时，不采用纯水，而采用350g/L的硫酸，镍钴锰锂浸出率稍有下降，渣量翻倍，膨胀石墨产品更少且比表面积明显降低；通过对比实施例1和对比例3的数据可知，若高压浸出步骤改为常压浸出，镍钴锰锂浸出率明显下降，渣量显著翻倍且无膨胀石墨产品，并且可参阅图3，相对于常压浸出(对应图中低酸浸出渣曲线)，高压浸出后(对应图中高压浸出渣曲线)镍钴锰酸锂的三元结构基本被完全破坏，高压渣中以碳为主，回收效果高；由上表实例数据可知，本公开提供了一种高效浸出镍钴锰锂的方法，其浸出率均接近于100％，且产生的渣量小且膨胀石墨比表面积较高，性能优良。
以上所述实施例仅表达了本公开的几种实施方式，其描述较为具体和详细，但并不能因此而理解为对发明专利范围的限制。应当指出的是，对于本领域的普通技术人员来说，在不脱离本公开构思的前提下，还可以做出若干变形和改进，则本公开也意图包含这些改动和变形。

Claims (19)

  1. 一种三元电池粉浸出方法，其特征在于，包括以下步骤：
    将所述三元电池粉配制成反应浆液；
    向所述反应浆液中加入氧化钙粉末，搅拌反应，反应完全后，再加入偏铝酸钠溶液和氧化钙粉末，搅拌反应，反应完全后，进行固液分离得到除氯后液；
    调节所述除氯后液的pH至预设pH值后，进行混凝沉淀，混凝沉淀后静置两小时，进行固液分离得到除氟氯后液和污泥；
    向所述除氟氯后液中加入氢氧化钠，碱浸，进行固液分离得到偏铝酸钠溶液与碱浸渣，将所述碱浸渣烘干粉碎得到碱浸渣粉末；
    将所述碱浸渣与焙烧添加剂混合均匀后，焙烧，得到焙烧渣；
    向所述焙烧渣中加入纯水，水浸，进行固液分离得到金属硫酸盐溶液和水浸渣，将所述水浸渣烘干粉碎得到水浸渣粉末；
    将所述水浸渣与浸出添加剂混合均匀后，加入纯水和硫酸，搅拌，高压浸出，进行固液分离得到金属硫酸盐溶液和高压渣。
  2. 根据权利要求1所述的一种三元电池粉浸出方法，其特征在于，配制所述反应浆液的方法为：按预设制浆条件向所述三元电池粉中加入纯水，搅拌均匀；所述预设制浆条件包括3:1～10:1的液固比，200～400rpm的搅拌速度和5～10min的搅拌时间。
  3. 根据权利要求1所述的一种三元电池粉浸出方法，其特征在于，制备所述除氯后液时，进行第一次所述搅拌反应前加入的所述氧化钙粉末的用量依据钙氟摩尔比为1:1～2:1确定，进行第一次所述搅拌反应后加入的所述偏铝酸钠溶液的用量依据铝氯质量比为1:10确定、所述氧化钙粉末的用量依据钙铝摩尔比为2:1确定。
  4. 根据权利要求1所述的一种三元电池粉浸出方法，其特征在于，制备所述除氯后液时，进行第一次所述搅拌反应的方法为在200～400rpm的搅拌速度下搅拌30min，进行第二次所述搅拌反应的方法为在200～400rpm的搅拌速度下搅拌60min。
  5. 根据权利要求1所述的一种三元电池粉浸出方法，其特征在于，调节所述pH时，所述预设pH值为12。
  6. 根据权利要求1所述的一种三元电池粉浸出方法，其特征在于，进行所述混凝沉淀的方法为向所述除氯后液中加入混凝剂并调节pH至6～8，搅拌反应。
  7. 根据权利要求6所述的一种三元电池粉浸出方法，其特征在于，所述混凝剂的加入量为1000mg/L，所述混凝剂包括氯化铝、硫酸铝、硫酸铁、氯化铁中的一种或几种，调节所述pH的试剂为氧化钙和硫酸。
  8. 根据权利要求6所述的一种三元电池粉浸出方法，其特征在于，进行所述搅拌反应的方法为先在200～400rpm的搅拌速度下搅拌5min，再在40～70rpm的搅拌速度下搅拌60min，所述静置的时间为2h。
  9. 根据权利要求1所述的一种三元电池粉浸出方法，其特征在于，加入所述氢氧化钠时，所述氢氧化钠的用量依据钠铝摩尔比为5:1～7:1确定。
  10. 根据权利要求1所述的一种三元电池粉浸出方法，其特征在于，进行所述碱浸的方法为在200～400rpm的搅拌速度下搅拌1～2h。
  11. 根据权利要求1所述的一种三元电池粉浸出方法，其特征在于，所述焙烧添加剂为石墨渣粉末和二水硫酸氢钠，所述石墨渣、二水硫酸氢钠与碱浸渣的重量比为0.2:1:5。
  12. 根据权利要求1所述的一种三元电池粉浸出方法，其特征在于，进行所述焙烧的方法为在600℃下焙烧2～3h。
  13. 根据权利要求1所述的一种三元电池粉浸出方法，其特征在于，所述焙烧渣与所述纯水的固液比为3:1～10:1，进行所述水浸的方法为在200～400rpm的搅拌速度下搅拌45min。
  14. 根据权利要求1所述的一种三元电池粉浸出方法，其特征在于，所述浸出添加剂为硫酸亚铁和石墨粉末，所述硫酸亚铁、石墨渣与水浸渣的重量比为0.3:1:2。
  15. 根据权利要求1所述的一种三元电池粉浸出方法，其特征在于，所述水浸渣粉末与所述纯水的固液比为6:1～8:1，所述硫酸的加入量为250～400g/L，所述硫酸的浓度为98％。
  16. 根据权利要求1所述的一种三元电池粉浸出方法，其特征在于，进行所述高压浸出的方法为在160～180℃下浸出2～3h，浸出的最大压力不大于2MPa。
  17. 根据权利要求1所述的一种三元电池粉浸出方法，其特征在于，还包括以下步骤：将所述碱浸得到的偏铝酸钠溶液用于制备所述除氯后液。
  18. 根据权利要求1所述的一种三元电池粉浸出方法，其特征在于，还包括以下步骤：对所述金属硫酸盐溶液进行除铁、萃取和合成，得到含镍、钴、锰、锂的产品。
  19. 根据权利要求1所述的一种三元电池粉浸出方法，其特征在于，将所述高压渣烘干并进行风选得到膨胀石墨和风选渣，将所述风选渣作为所述碱浸渣用于制备所述碱浸渣粉末。
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