WO2023098500A1 - 锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法 - Google Patents

锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法 Download PDF

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WO2023098500A1
WO2023098500A1 PCT/CN2022/133160 CN2022133160W WO2023098500A1 WO 2023098500 A1 WO2023098500 A1 WO 2023098500A1 CN 2022133160 W CN2022133160 W CN 2022133160W WO 2023098500 A1 WO2023098500 A1 WO 2023098500A1
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lithium
niobium
acid
tantalum
gypsum
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PCT/CN2022/133160
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English (en)
French (fr)
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殷志刚
周复
邓星星
徐川
高宜宝
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天齐创锂科技(深圳)有限公司
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Priority to AU2022402780A priority Critical patent/AU2022402780B2/en
Publication of WO2023098500A1 publication Critical patent/WO2023098500A1/zh

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    • B—PERFORMING OPERATIONS; TRANSPORTING
    • B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
    • B03B—SEPARATING SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS
    • B03B9/00—General arrangement of separating plant, e.g. flow sheets
    • B03B9/06—General arrangement of separating plant, e.g. flow sheets specially adapted for refuse
    • 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
    • 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
    • Y02W—CLIMATE CHANGE MITIGATION TECHNOLOGIES RELATED TO WASTEWATER TREATMENT OR WASTE MANAGEMENT
    • Y02W30/00—Technologies for solid waste management
    • Y02W30/20—Waste processing or separation

Definitions

  • the invention relates to a method for comprehensively recovering lithium, tantalum-niobium, silicon-aluminum powder, iron concentrate and gypsum from lithium slag, and belongs to the technical field of lithium slag treatment.
  • the comprehensive recycling of solid waste secondary resources is a major theme of environmental protection in the world today.
  • the comprehensive recycling of solid waste is conducive to alleviating the shortage of national resources and moving towards a sustainable development path. It is the only way to create a conservation-oriented society.
  • the comprehensive recycling of solid waste has achieved remarkable results.
  • people's requirements for environmental ecological protection have increased and the shortage of resources has led to the problem of lagging economic development.
  • the utilization of solid waste resources will have a huge development prospect.
  • the lithium slag extracted from spodumene is mainly used as an ingredient in low value-added fields such as cement and concrete, which makes the lithium slag extracted from spodumene unable to be quickly absorbed.
  • the stockpiling of lithium slag extracted from spodumene will undoubtedly bring problems such as environmental pollution and land occupation.
  • the demand for comprehensive utilization of lithium slag extracted from spodumene will become more urgent with the rapid development of the lithium battery industry.
  • Patent CN1297860A and patent CN1090597C disclose ceramic glazed tiles made of acidic lithium slag and its manufacturing method. It uses acidic lithium slag as the main raw material, and is equipped with wollastonite, pyrophyllite and kaolin as auxiliary materials. After grinding, pulping, Filtration, mud cake, drying, crushing, compacting, drying, biscuit firing, glaze firing and other steps, and finally develop a technology to replace a part of traditional high-quality mineral raw materials with acidic lithium slag to manufacture ceramic glazed tiles. Since these two patents only use a small amount of lithium slag, they have no technical advantages for quickly absorbing a large amount of lithium slag.
  • Patent CN103601230 discloses a method for the comprehensive utilization of lithium slag to produce chemical raw materials. This patent finally obtains calcium chloride, ammonium fluoride, white carbon black, aluminum salt and ammonium sulfate through multiple steps. This patent cannot avoid the use of a large amount of Acid solution, resulting in high acidity in the exhaust gas and difficult production operations.
  • Patent CN108273826A discloses a full-phase high-value recycling method for lithium slag. It mainly uses alkali conversion-magnetic separation to obtain pyrophyllite raw materials for glass fiber. The by-products are gypsum and magnetic separation tailings. The patent uses alkali The conversion process has the disadvantage of high cost and cannot be industrialized. Patent CN108147658A and patent WO2019/141098A1 mainly describe the use of flotation-magnetic separation process to obtain lithium pyrophyllite, gypsum, and magnetic separation tailings.
  • Patent CN214488258U discloses a comprehensive recovery and utilization system of lithium slag, which mainly adopts pre-grinding and water circulation classification rough technology to replace pulping operation, and adopts flotation and alkali-to-solid-liquid two-phase desulfurization technology, weak-strong magnetic separation Iron technology and ring water segmental circulation technology are used to comprehensively utilize the waste residue generated after lithium ore sulfuric acid process lithium extraction.
  • This patent only provides a system, which has the following disadvantages: the pre-grinding will make it more difficult to separate lithium slag and gypsum, and the final result is that the yield of silicon-aluminum fine powder in flotation operations is low and the cost remains high; in addition, Alkali conversion has the characteristics of high cost, long conversion time, and low efficiency, and it is difficult to scale up production. Secondly, the system does not recover iron and lithium from lithium slag, which undoubtedly causes waste of resources and does not realize comprehensive recycling.
  • Chinese patent CN108191226A discloses a method for producing glass fiber by using spodumene slag as a fluxing clarifier.
  • the raw material ratio is: 100-120 parts by mass of kaolin, 150-410 parts by mass of pyrophyllite, and 150-200 parts by mass of quicklime 50-70 parts by mass of dolomite, 50-70 parts by mass of colemanite, 130-310 parts by mass of lepiolite, 20-30 parts by mass of fluorite, and 10-30 parts by mass of spodumene slag.
  • Chinese patent CN1114232223A discloses a method for preparing ceramics by replacing kaolin with spodumene slag.
  • the proportion of various raw materials for the ceramic blank is: 50-75% of spodumene slag, 10-20% of quartz sand, and 1% of potassium feldspar. -10%, albite 1-10%; the proportion of various raw materials for ceramic glazes is: spodumene slag 40-60%, quartz sand 15-40%, feldspar 15-20%, porcelain stone 1- 10%.
  • Chinese patent CN113480182A discloses a kind of glass fiber with industrial waste as the main raw material and its preparation method.
  • the raw material components of glass fiber are: 0-200 parts by mass of lithium industrial tailings, 70-200 parts by mass of coal gangue, high silicon and aluminum associated 400-650 parts by mass of ore.
  • This invention proposes glass fiber and its preparation scheme with industrial waste residue as the main raw material. By rationally using industrial waste, it can replace the raw materials such as pyrophyllite, kaolin and quartz mainly used in the current glass fiber production. Reduce industrial risk of raw materials for glass fiber production.
  • CN1090597C discloses the method of utilizing acidic lithium slag to manufacture ceramic glazed tiles. This technology also has the disadvantages of high cost, difficult industrialization, and slow speed of absorbing lithium slag.
  • Patent CN1112335C provides a technology of using spodumene acid method to extract lithium waste residue to prepare gypsum reinforcing agent.
  • CN106082739A provides a technology of spodumene acid method lithium extraction waste residue mixed and dried as cement admixture technology.
  • the technical problem to be solved by the present invention is to provide a method for comprehensively recovering lithium, tantalum-niobium, silicon-aluminum micropowder, iron concentrate and gypsum from lithium slag.
  • the lithium slag of the present invention comprehensively reclaims the method for lithium, tantalum-niobium, silicon-aluminum micropowder, iron concentrate and gypsum comprising:
  • the lithium slag is re-selected to obtain concentrate 1 and tailings 1, and the concentrate 1 is separated by weak magnetic field to obtain coarse-grained tantalum-niobium rich material and coarse-grained iron concentrate;
  • step d performing weak magnetic separation on the tailings pulverized in step c to obtain fine-grained iron concentrate and tailings 3;
  • step e reselecting the concentrate 2 in step e to obtain fine-grained tantalum-niobium concentrate and high-iron lithium-rich material;
  • the lithium slag is lithium slag extracted from spitzolite;
  • the re-election in step a includes shaker gravity separation, spiral gravity separation, centrifugal gravity separation, hydrocyclone gravity separation, jig gravity gravity separation, wind gravity gravity separation, One or more combinations of heavy media reselections;
  • the collector of flotation described in b step comprises by weight:
  • the polyether or polyalcohol is at least one of polyvinyl ether, polyoxypropylene ether, polyvinyl alcohol, and polyoxyethylene ether, preferably 1 to 10 parts of polyvinyl ether, 1 to 10 parts of polyoxypropylene ether, 1-10 parts of polyvinyl alcohol;
  • the propylene oxide block copolymer is at least one of PE6100, PE6200, PE6400, PE8100;
  • the dodecylsulfonic acid or sulfuric acid includes dodecylbenzenesulfonic acid, dodecylsulfonic acid, dodecylsulfuric acid; preferably includes dodecylbenzenesulfonic acid and its salts; more preferably the 1-10 parts of dodecylbenzenesulfonic acid and its salt;
  • the mass concentration of the silica sol is preferably 5-40%;
  • the pulp concentration of the flotation is preferably 20-60%.
  • the grade of tantalum in the lithium slag is calculated as Ta 2 O 5 and the grade of niobium is calculated as Nb 2 O5, which is lower than 100 ppm, preferably 50-100 ppm.
  • the oxide grade of tantalum and niobium in the lithium slag of the present invention is lower than 100ppm and can be recovered, and the recovery rate can be guaranteed to be greater than 45%.
  • the recovery rate of tantalum and niobium is low because the grade is too low, or the recovery is abandoned;
  • the grade of tantalum and niobium is higher than 100ppm, the method of the present invention can certainly recover well, and the recovery rate is higher.
  • the magnetic field strength of the weak magnetic separation is 100-2000 Gauss, preferably 300-1000 Gauss; the magnetic field strength of the strong magnetic separation is 10000-20000 Gauss, preferably 12000-17000 Gauss.
  • the flotation described in step b has also added a regulator, and the regulator is:
  • the amount of the collector in step b is 50-3000 g/t, preferably 100-1000 g/t lithium slag.
  • the flotation includes roughing, sweeping and fine selection, preferably the roughing is 1 to 3 times, the sweeping is 1 to 4 times, and the fineness is 1 to 3 times; the sweeping process captures
  • the amount of collector used is 1/20-13/12 of the amount of collector used in rough selection, and no collector is added during the selection process;
  • the amount of collector used is 1/3 of the rough selection amount, and the amount of sweeping three collectors is 1/4 of the rough selection amount.
  • the particle size of the tailings 2 after the crushing in step c is below 325 mesh
  • the crushing is preferably to classify the tailings 2 to obtain particles above 325 mesh and below 325 mesh, and crush the particles above 325 mesh Afterwards, it is mixed with particles below 325 mesh; the crushing is preferably finely ground by a non-ferrous media mill.
  • 325 mesh is about 45 ⁇ m, and 325 mesh or less is about 45 ⁇ m or less. 325 mesh or more is about 45 ⁇ m or more.
  • concentration-filtration is also carried out before the drying described in step e.
  • the reselection in step a includes rough selection and refinement, preferably 1 to 3 times of rough selection and 1 to 3 times of refinement;
  • the reselection in step f includes rough selection and fine selection, preferably 1 to 3 times of rough selection and 1 to 3 times of fine selection.
  • the C8-20 fatty acids and salts thereof in the collector include octanoic acid, nonanoic acid, capric acid, undecanoic acid, dodecanoic acid, tridecanoic acid, tetradecanoic acid, pentadecanoic acid , at least one of palmitic acid, margaric acid, octadecanoic acid, nonadecanic acid, eicosic acid, oleic acid, linoleic acid, linolenic acid, arachidonic acid;
  • the aviation kerosene includes 1-10 parts of aviation kerosene wide fraction type; the aviation kerosene includes aviation kerosene preferably also includes 1-10 parts of kerosene type; 1-10 parts of heavy fraction type;
  • the monoglyceride fatty acid ester includes at least one of glyceryl oleate, glyceryl stearate, glyceryl laurate, and glyceryl palmitate; preferably includes glyceryl laurate;
  • Described quaternary ammonium salt comprises dodecyl to hexadecyl trimethyl ammonium chloride or ammonium bromide; Be preferably dodecyl, fourteen or cetyl trimethyl ammonium chloride or ammonium bromide; More preferably Dodecyltrimethylammonium chloride or ammonium bromide;
  • the alkali is at least one of sodium hydroxide, potassium hydroxide, sodium carbonate, potassium carbonate, sodium bicarbonate, potassium bicarbonate;
  • the salt is sodium salt, potassium salt, ammonium salt, calcium salt, magnesium salt at least one of .
  • the kerosene type of aviation kerosene is also called middle fraction type of aviation kerosene, with a boiling point of 150°C to 280°C, a boiling point of heavy fraction type of 190°C to 315°C, and a wide fraction type of boiling point of 60°C to 280°C.
  • the described lithium spodumene sulfuric acid extraction tailings flotation desulfurization collector is prepared by the following method:
  • the present invention thoroughly realizes the purpose of diversification and high-value utilization of deep-processed lithium slag products, and solves the major problem of refractory slag that plagues the lithium salt industry;
  • the present invention can obtain high-silicon, high-aluminum, low-iron, and low-sulfur silica-alumina micropowder, which can be used in industries such as glass fiber, ceramics, and papermaking, and is used to replace raw materials such as pyrophyllite, kaolin, talc, etc., greatly Greatly reduce the production cost of glass fiber, ceramics and paper industry;
  • the present invention obtains high-quality gypsum concentrate through flotation, and the purity of gypsum is as high as 95%. materials, etc., increasing the value of gypsum;
  • the present invention makes full use of resource characteristics, obtains iron ore concentrate through weak magnetic separation, and further improves the value of comprehensive utilization of lithium slag;
  • the present invention has obtained tantalum-niobium concentrate, calculated as tantalum-niobium oxide content of 150ppm, the annual output of lithium slag is 3 million tons, and the total amount of tantalum-niobium oxide is close to 450 tons.
  • the present invention has obtained the high-iron lithium-rich slag, the content of lithium oxide ( Li2O ) in the high-iron lithium-rich slag is 1.0-1.5%, and the yield of the high-iron lithium-rich slag is about 5-10%, calculated with 7% yield, then It produces 210,000 tons of high-iron lithium-rich slag a year, and about 2,000-3,000 tons of lithium metal.
  • the high-iron lithium-rich slag can be used as lithium ore to further recover lithium carbonate.
  • the content of SO in the tailings left by the flotation of the present invention is small, and for lithium slag with SO content greater than 10%, flotation can obtain tailings with SO content less than 0.1%.
  • Fig. 1 is a process flow diagram of a specific embodiment of the present invention.
  • Fig. 2 is a process flow diagram of recycling and extracting lithium from a high-iron lithium-rich material according to the present invention.
  • Fig. 3 is a comprehensive recovery process diagram of acid-based roasting of a high-iron lithium-rich material according to the present invention.
  • the lithium slag of the present invention comprehensively reclaims the method for lithium, tantalum-niobium, silicon-aluminum micropowder, iron concentrate and gypsum comprising:
  • the lithium slag is re-selected to obtain concentrate 1 and tailings 1, and the concentrate 1 is separated by weak magnetic field to obtain coarse-grained tantalum-niobium rich material and coarse-grained iron concentrate;
  • step d performing weak magnetic separation on the tailings pulverized in step c to obtain fine-grained iron concentrate and tailings 3;
  • step e reselecting the concentrate 2 in step e to obtain fine-grained tantalum-niobium concentrate and high-iron lithium-rich material;
  • the lithium slag is lithium slag extracted from spitzolite;
  • the re-election in step a includes shaker gravity separation, spiral gravity separation, centrifugal gravity separation, hydrocyclone gravity separation, jig gravity gravity separation, wind gravity gravity separation, One or more combinations of heavy media reselections;
  • the collector of flotation described in b step comprises by weight:
  • the polyether or polyalcohol is at least one of polyvinyl ether, polyoxypropylene ether, polyvinyl alcohol, and polyoxyethylene ether, preferably 1 to 10 parts of polyvinyl ether, 1 to 10 parts of polyoxypropylene ether, 1-10 parts of polyvinyl alcohol;
  • the propylene oxide block copolymer is at least one of PE6100, PE6200, PE6400, PE8100;
  • the dodecylsulfonic acid or sulfuric acid includes dodecylbenzenesulfonic acid, dodecylsulfonic acid, dodecylsulfuric acid; preferably includes dodecylbenzenesulfonic acid and its salts; more preferably the 1-10 parts of dodecylbenzenesulfonic acid and its salt;
  • the mass concentration of the silica sol is preferably 5-40%;
  • the pulp concentration of the flotation is preferably 20-60%.
  • the distribution ratio of each component of the collector is mixed in any proportion within the corresponding range, and the purpose of high-efficiency flotation desulfurization of lithium slag can be achieved according to the corresponding ratio, and the content of gypsum is guaranteed to be greater than 95%, and the content of impurity SiO 2 in gypsum is less than 1 %, Al 2 O 3 content ⁇ 1%, providing high-quality raw materials for the subsequent preparation of whisker gypsum.
  • the configuration of the desulfurization collector of the present invention can achieve rapid and efficient removal of sulfur from lithium slag by effectively adjusting the content of each component; through the process of the present invention, high-quality gypsum can be easily obtained.
  • the gypsum obtained by flotation in the present invention can be directly used as a raw material for producing gypsum whiskers, gypsum putty powder or filler after being filtered.
  • the filtered water produced by flotation continues to return to the flotation operation after being collected, and the present invention does not produce waste water discharge; in view of the fact that the gypsum in the product will take away part of the water, the production process of the present invention needs to add new water to ensure normal production.
  • the grade of tantalum in the lithium slag is calculated as Ta 2 O 5 and the grade of niobium is calculated as Nb 2 O5, which is lower than 100 ppm, preferably 50-100 ppm.
  • the oxide grade of tantalum and niobium in the lithium slag of the present invention is lower than 100ppm and can be recovered, and its recovery rate can be guaranteed to be greater than 45%.
  • the grade is too low, the recovery rate of tantalum and niobium is low, or the recovery is abandoned;
  • the method of the present invention can also recover well when the niobium grade is higher than 100 ppm, and the recovery rate is higher.
  • the magnetic field strength of the weak magnetic separation is 100-2000 Gauss, preferably 300-1000 Gauss; the magnetic field strength of the strong magnetic separation is 10000-20000 Gauss, preferably 12000-17000 Gauss.
  • the flotation described in step b has also added a regulator, and the regulator is:
  • the amount of the collector in step b is 50-3000 g/t, preferably 100-1000 g/t lithium slag.
  • the flotation includes roughing, sweeping and fine selection, preferably the roughing is 1 to 3 times, the sweeping is 1 to 4 times, and the fineness is 1 to 3 times; the sweeping process captures
  • the amount of collector is 1/20-12/13 of the amount of collector used in rough selection, and no collector is added during the selection process;
  • the amount of collector used is 1/3 of the rough selection amount, and the amount of sweeping three collectors is 1/4 of the rough selection amount.
  • the mesh size of the tailings 2 after the crushing in step c is below 325 mesh, and the crushing is preferably to classify the tailings 2 to obtain particles with a mesh number of 325 mesh or more and a mesh number of 325 mesh or less.
  • the particles with a mesh size above 325 mesh are pulverized and mixed with particles with a mesh size below 325 mesh; the pulverization is preferably finely ground by a non-ferrous media mill.
  • Classification equipment can use spiral classifier, cyclone, linear screen and so on.
  • concentration-filtration is also carried out before the drying described in step e.
  • the reselection in step a includes rough selection and refinement, preferably 1 to 3 times of rough selection and 1 to 3 times of refinement;
  • the reselection in step f includes rough selection and fine selection, preferably 1 to 3 times of rough selection and 1 to 3 times of fine selection.
  • the C8-20 fatty acids and salts thereof in the collector include octanoic acid, nonanoic acid, capric acid, undecanoic acid, dodecanoic acid, tridecanoic acid, tetradecanoic acid, pentadecanoic acid , at least one of palmitic acid, margaric acid, octadecanoic acid, nonadecanic acid, eicosic acid, oleic acid, linoleic acid, linolenic acid, arachidonic acid;
  • the aviation kerosene includes 1-10 parts of aviation kerosene wide fraction type; the aviation kerosene includes aviation kerosene preferably also includes 1-10 parts of kerosene type; 1-10 parts of heavy fraction type;
  • the monoglyceride fatty acid ester includes at least one of glyceryl oleate, glyceryl stearate, glyceryl laurate, and glyceryl palmitate; preferably includes glyceryl laurate;
  • Described quaternary ammonium salt comprises dodecyl to hexadecyl trimethyl ammonium chloride or ammonium bromide; Be preferably dodecyl, fourteen or cetyl trimethyl ammonium chloride or ammonium bromide; More preferably Dodecyltrimethylammonium chloride or ammonium bromide;
  • the alkali is at least one of sodium hydroxide, potassium hydroxide, sodium carbonate, potassium carbonate, sodium bicarbonate, potassium bicarbonate;
  • the salt is sodium salt, potassium salt, ammonium salt, calcium salt, magnesium salt at least one of .
  • the kerosene type of aviation kerosene is also called middle fraction type of aviation kerosene, with a boiling point of 150°C to 280°C, a boiling point of heavy fraction type of 190°C to 315°C, and a wide fraction type of boiling point of 60°C to 280°C.
  • the described lithium spodumene sulfuric acid extraction tailings flotation desulfurization collector is prepared by the following method:
  • 1-5 parts of aviation kerosene wide fraction type Preferably, 1-5 parts of aviation kerosene wide fraction type; 1-5 parts of kerosene type; 1-5 parts of heavy fraction type;
  • dodecylbenzenesulfonic acid and its salt Preferably, 1-5 parts of dodecylbenzenesulfonic acid and its salt; 1-5 parts of dodecylsulfuric acid and its salt; 1-5 parts of dodecylsulfonic acid and its salt;
  • dodecyltrimethylammonium chloride or dodecyltrimethylammonium bromide Preferably, 1-5 parts of dodecyltrimethylammonium chloride or dodecyltrimethylammonium bromide; 1-5 parts of tetradecyltrimethylammonium chloride or tetradecyltrimethylammonium bromide 10 parts, cetyltrimethylammonium chloride or cetyltrimethylammonium bromide 1-5 parts.
  • 1-2 parts of aviation kerosene wide fraction type 1-2 parts of kerosene type; 1-2 parts of heavy fraction type;
  • dodecylbenzenesulfonic acid and its salt 1 part of dodecylbenzenesulfonic acid and its salt; 1 part of dodecylsulfuric acid and its salt; 1 part of dodecylsulfonic acid and its salt;
  • polyvinyl ether Preferably 1-2 parts of polyvinyl ether; 1-2 parts of polyoxypropylene ether; 1-2 parts of polyvinyl alcohol;
  • dodecyltrimethylammonium chloride or dodecyltrimethylammonium bromide 1 part of tetradecyltrimethylammonium chloride or tetradecyltrimethylammonium bromide, 1 part 1 part of hexaalkyltrimethylammonium chloride or cetyltrimethylammonium bromide.
  • the method further includes subjecting the high-iron lithium-rich material to alkaline or acid roasting-leaching-solid-liquid separation-purifying and concentrating the filtrate after solid-liquid separation to obtain lithium salt products .
  • the high-iron lithium-rich material can adopt the lithium extraction process similar to the existing spodumene lithium extraction process.
  • the difference between the high-iron lithium-rich material and spodumene is that it does not need to be transformed and roasted, and can be directly roasted by alkali method or acid method.
  • Alkali method or acid method roasting process the acid used in acid method roasting is concentrated sulfuric acid, in principle, choose 80% or more concentrated sulfuric acid, preferably 98% concentrated sulfuric acid.
  • the solid after solid-liquid separation is silica-alumina micropowder, which can be recycled.
  • lithium slag was extracted from spodumene from a company in Sichuan.
  • the main minerals are quartz, calcite, gypsum, gibbsite, andalusite, corundum, glass phase, ⁇ -spodumene, ⁇ -spodumene, zeolite, and Feldspar, tantalite (trace), niobite (trace).
  • the grade of raw material Ta2O5 is 90ppm
  • the grade of Nb2O5 is 50ppm
  • the grade of SO3 is 6.2%.
  • the raw materials are directly re-selected in the spiral chute, and the re-separation concentrate in the spiral chute enters the shaking table for selection, and the shaking table concentrate is directly separated by weak magnetic field.
  • the magnetic field strength is 1000 gauss.
  • the grade of 2 O 5 is 18.56%
  • the grade of Nb 2 O 5 is 9.56%
  • the recovery rate of tantalum and niobium is 46.12% and 32.68% respectively
  • the magnetic separation concentrate is coarse-grained iron concentrate
  • TFe is 52.13%
  • the recovery rate is 12.89%.
  • Preparation of collector First, mix 20 parts of sodium hydroxide and 50 parts of silica sol with a mass fraction of 40%, heat to 80° C. and stir for 5 hours to obtain paste A;
  • fatty acid/fatty acid salt of C 8-20 (the fatty acid/fatty acid salt of C 8-20 in this embodiment is caprylic acid and lauric acid mixed according to 1:1), 1 part of aviation kerosene wide fraction type, 1 part of aviation kerosene middle fraction type, 1 part of aviation kerosene heavy fraction type, 1 part of sodium dodecylbenzenesulfonate, 1 part of sodium dodecyl sulfate, 1 part of polyvinyl ether, 1 part of polyoxypropylene ether, poly 1 part of vinyl alcohol, 1 part of ethylene oxide-propylene oxide block copolymer EO-PO-EO (PE6100 used in this experiment), 1 part of sorbitol monooleate, 1 part of monoglyceride oleate (this experiment)
  • the monoglycerol oleate in this experiment is glyceryl oleate, glyceryl stearate, glyceryl laurate, glyceryl
  • the paste A and the paste B are evenly mixed to obtain the lithium slag desulfurization collector C.
  • Gravity separation tailings directly enter the flotation, adjust the slurry concentration to 35%, and add 2000g of adjusting agent and 300g of collector in turn according to the ton of ore to carry out the first roughing; the second roughing adjusting agent dosage is 500g, The amount of the collector is 100g, and the coarser 1 concentrate and the rougher 2 concentrate are mixed to obtain the rougher concentrate, and the adjustment agent for the flotation rougher is aluminum sol.
  • the rougher concentrate is selected once, and the selected middle ore returns to the first rougher; the rougher tailings are swept three times, the amount of the first sweeping agent is 500g, the amount of collector is 50g, and the second time
  • the amount of sweeping adjustment agent is 250g, the amount of collector is 30g, the amount of third sweeping adjustment agent is 250g, and the amount of collector is 20g. Sweeping ore returns to the first roughing operation, and finally a closed circuit circulation is formed.
  • Desulfurization gypsum concentrate and desulfurization tailings are obtained, and the purity of CaSO 4 .2H 2 O in the gypsum concentrate is greater than 95%.
  • the flotation tailings are classified into -325 mesh samples and +325 mesh samples; the +325 mesh samples are directly fed into the ceramic mill, and the grinding to -325 mesh accounts for 100%; the tailings are classified into -325 mesh samples and after grinding-
  • the 325-mesh samples were mixed and directly entered into weak magnetic separation with a magnetic field strength of 2000 Gauss to obtain fine-grained iron concentrate with a TFe of 42.23% and a yield of 8.2%.
  • the ore pulp after weak magnetic separation directly enters the strong magnetic separator.
  • the magnetic field strength of strong magnetic separation roughing is 1.2T, and that of sweeping is 1.7T.
  • the magnetic separation tailings are directly concentrated, filtered, and dried to obtain silica-alumina micropowder 1 with a yield of 75%, a grade of SO 3 of 0.15%, and a grade of Fe 2 O 3 of 0.32%.
  • the magnetic products rich in tantalum, niobium and lithium directly enter the shaking table for gravity separation, the concentrate of the first gravity separation is directly selected, the ore of the first gravity separation returns to gravity separation, and the tailings of the first gravity separation are directly used as tailings; Gravity separation concentrate is used for the second selection, and the second gravity separation concentrate is the final fine-grained tantalum and niobium concentrate, and the second gravity separation and tailings are directly returned to the first gravity separation. After gravity separation, fine-grained tantalum-niobium concentrate and high-iron lithium-rich material were obtained respectively.
  • the fine-grained tantalum-niobium concentrate had a Ta 2 O 5 grade of 10.52% and a Nb 2 O 5 grade of 4.78%, and the recovery rates of tantalum and niobium were 14.73% and 18.79% respectively.
  • the yield of silicon-aluminum micropowder in the whole process of extracting lithium from high-iron lithium-rich material by sulfuric acid method is 92%, and then the lithium carbonate product is obtained after concentration-removal of sodium, lithium precipitation and other processes, and the recovery rate of Li 2 O in the whole process is 20.5% (the operating recovery rate is 82%).
  • Example 2 Others are similar to Example 1, except that the only difference is 5 parts of aviation kerosene wide fraction type; 3 parts of aviation kerosene medium fraction type, and 3 parts of aviation kerosene heavy fraction type.
  • Example 4 Others are similar to Example 1, except that the regulator of Example 4 is chitosan.
  • the regulator of embodiment 4 is sodium carboxymethyl cellulose.
  • Example 1 Others are similar to Example 1, the only difference in Embodiment 6 is that the magnetic field strength of the weak magnetic separation is 1500 Gauss; the only difference in Embodiment 7 is that the magnetic field strength of the strong magnetic separation is 11000 Gauss;
  • Embodiment 8 the collector consumption for the first time is 500g/t
  • the collector consumption for the second time is 250g/t
  • the collector consumption for the third time is 250g/t. Consumption is 100g/t
  • the only difference in embodiment 9 is that the first time the amount of sweeping modifier is 800g/t, the second time the amount of sweeping modifier is 500g/t, and the third time the amount of sweeping modifier is 200g/t. t.

Abstract

一种锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，属于锂渣处理技术领域。该锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法包括：将锂渣重选得到精矿1和尾矿1，精矿1弱磁分离得到粗粒钽铌富料和粗粒铁精矿；将尾矿1进行浮选，获得石膏和尾矿2；将尾矿2粉碎；将粉碎后的尾矿进行弱磁分离，得到细粒铁精矿和尾矿3；将尾矿3强磁分离得到精矿2和尾矿4，尾矿4干燥即得硅铝微粉；将精矿2重选得到细粒钽铌精矿和高铁富锂料。解决了困扰锂盐行业的渣难处理的重大难题；可以获得高硅高铝低铁低硫的硅铝微粉、纯度高达95％以上的石膏精矿、铁精矿、钽铌精矿、高铁富锂渣。

Description

锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法
相关申请的交叉引用
本申请要求在2021年12月01日提交的中国专利申请CN2021114556395的权益和优先权，并且在此全文引用该申请以作参考以及其它全部用途。
技术领域
本发明涉及一种锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，属于锂渣处理技术领域。
背景技术
固体废弃物二次资源的综合回收利用是当今世界环保的一大主题，固体废弃物综合回收利用有利于缓解国家资源短缺，走向可持续发展道路，是创造节约型社会的必由之路。近年来，伴随政策对二次资源综合回收利用的扶持，固体废弃物的综合回收取得了显著成效。伴随经济社会的发展，人们对环境生态保护要求的提高及资源的短缺导致经济发展滞后的问题凸显，固废资源化利用将有巨大的发展前景。
目前，由于锂电行业的快速发展，导致对锂资源需求增长。锂盐的提取主要依赖于矿石和盐湖，其中盐湖提锂由于成本高、杂质含量偏高等导致其难以与矿石提锂媲美。矿石提锂主要采用锂辉石，且很多锂辉石资源由于种种原因导致无法大规模的投产。采用锂辉石提锂盐，每生产1吨锂盐会产出7～8吨渣，以现有的锂盐产量计算，一年产生超过200万吨锂渣。目前，锂辉石提锂渣主要作为水泥、混凝土等低附加值领域的配料，导致锂辉石提锂渣无法被快速消纳。锂辉石提锂渣的堆存无疑会带来环境污染、土地占用等问题，对锂辉石提锂渣进行综合利用的需求将伴随锂电行业的快速发展变得更加紧迫。
专利CN1297860A与专利CN1090597C公开了用酸性锂渣制造的陶瓷釉面砖及其制造方法，其采用酸性锂渣为主要原料，配以硅灰石、叶蜡石和高岭土作辅料，经粉磨、制浆、压滤、泥饼、干燥、粉碎、压坯、干燥、素烧、釉烧等步骤，最终开发除一种酸性锂渣取代一部分传统的优质矿物原料来制造陶瓷釉面砖的技术。这两篇专利由于只使用了少量的锂渣，对快速消纳大量锂渣没有技术优势。专利CN103601230公开了一种锂渣综合利用生产化工原料的方法，该专利通过多个步骤最终获得了氯化钙、 氟化铵、白炭黑、铝盐和硫酸铵，该专利无法避免使用大量的酸溶液，导致排放气体中酸度较高，生产作业难度大。
专利CN108273826A公布了一种锂渣的全相高值化回收利用方法，其主要采用碱转化-磁选的方式获得玻纤用叶蜡石原料，副产品为石膏、磁选尾渣，该专利采用碱转化工艺，存在成本高的缺点，无法产业化。专利CN108147658A和专利WO2019/141098A1主要阐述采用浮选-磁选的工艺获得锂质叶蜡石、石膏、磁选尾渣，这两篇专利虽然将锂渣进行了高值化制备叶蜡石，但尚未将其中的硫酸钙、钽铌、锂等资源高值化。专利CN214488258U公开了一种锂渣的综合回收利用系统，主要采用前置磨矿和水循环分级粗略技术代替造浆作业、以及采用浮选和碱转固液两相脱硫技术、弱-强磁选除铁技术和环水分段循环技术，对锂矿石硫酸法工艺提锂后产生的废渣进行综合利用。该专利只是提供了一个系统，存在以下缺点：其中磨矿前置会导致锂渣与石膏分离难度加大，最终的结果是浮选作业硅铝微粉产率偏低，成本居高不下；另外，碱转化存在成本高、转化时间长、效率低等特点，生产上难以规模化；其次，该系统没有回收锂渣中得铁和锂，无疑造成资源浪费，没有实现综合回收利用。
中国专利CN108191226A公开了一种用锂辉石矿渣作助熔澄清剂生产玻璃纤维的方法，其原料配比为：高岭土100-120质量份、叶蜡石150-410质量份、生石灰150-200质量份、白云石50-70质量份、硬硼钙石50-70质量份、白泡石130-310质量份、萤石20-30质量份、锂辉石矿渣10-30质量份。中国专利CN1114232223A公开了一种由锂辉石矿渣代替高岭土制备陶瓷的方法，陶瓷胚料的各种原料占比为：锂辉石矿渣50-75％、石英砂10-20％、钾长石1-10％、钠长石1-10％；陶瓷釉料的各种原料占比为：锂辉石矿渣40-60％、石英砂15-40％、长石15-20％、瓷石1-10％。中国专利CN113480182A公开了一种以工业废料为主原料的玻璃纤维及其制备方法，玻璃纤维原料组分为：锂工业尾矿0-200质量份、煤矸石70-200质量份、高硅铝伴生矿400-650质量份，该发明提出了以工业废渣为主原料的玻璃纤维及其制备方案，通过合理使用工业废弃物，替代当前玻璃纤维生产主要使用的叶蜡石、高岭土和石英等原料，降低玻璃纤维生产原料的工业风险。CN1090597C公布了利用酸性锂渣制造陶瓷釉面砖的方法，这种技术同样存在成本高、难产业化、消纳锂渣速度慢等缺点。
专利CN1112335C提供了一种利用锂辉石酸法提锂废渣制备石膏增强剂的技术，CN106082739A提供了一种锂辉石酸法提锂废渣混合进行烘干所得的产物作为水泥掺合料的技术，这些专利技术还停留在低值化阶段，没有更好得实现锂辉石提锂渣高值 化利用。
综上所述，如果能够开发一种锂渣资源化综合回收利用得技术，将锂渣中的石膏、钽铌、铁、硅铝微粉及锂全部资源化，顺利消纳锂渣，从而能够承接锂产业发展，将会极大的促进锂工业及锂渣处理行业的健康快速发展。锂渣中的各有价组分均得到有效利用且无固废产生，将一举解决锂工业发展得后顾之忧。
发明内容
本发明要解决的技术问题是提供一种锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法。
为解决上述第一个技术问题，本发明的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法包括：
a.将锂渣重选得到精矿1和尾矿1，所述精矿1弱磁分离得到粗粒钽铌富料和粗粒铁精矿；
b.将所述尾矿1进行浮选，获得石膏和尾矿2；
c.将所述尾矿2粉碎；
d.将c步骤粉碎后的尾矿进行弱磁分离，得到细粒铁精矿和尾矿3；
e.将尾矿3强磁分离得到精矿2和尾矿4，尾矿4干燥即得硅铝微粉；
f.将步骤e中精矿2重选得到细粒钽铌精矿和高铁富锂料；
其中，所述锂渣为辉锂石提锂渣；a步骤所述重选为摇床重选、螺旋重选、离心重选、水力旋流器重选、跳汰机重选、风力重选、重介质重选中的一种或多种组合；
b步骤所述浮选的捕收剂按重量计包括：
C8-20的脂肪酸及其盐中的至少一种50～100份；航空煤油1～30份；十二烷基的磺酸或硫酸及其盐中的至少一种1～30份；聚醚或聚醇中的至少一种1～30份；环氧丙烷嵌段共聚物1～10份；山梨醇单油酸酯1～10份；单甘油脂肪酸酯1～10份；季铵盐1～30份；十六烷基卤化吡啶1～10份；碱5～50份；硅溶胶10～50份；水10～100份；
所述聚醚或聚醇为聚乙烯醚、聚氧丙烯醚、聚乙烯醇、聚氧乙烯醚中的至少一种，优选为聚乙烯醚1～10份、聚氧丙烯醚1～10份、聚乙烯醇1～10份；
所述环氧丙烷嵌段共聚物为PE6100、PE6200、PE6400、PE8100中的至少一种；
所述十二烷基的磺酸或硫酸包括十二烷基苯磺酸、十二烷基磺酸、十二烷基硫酸；优选包括十二烷基苯磺酸及其盐；更优选所述十二烷基苯磺酸及其盐1～10份；
所述硅溶胶质量浓度优选为5～40％；
所述浮选的矿浆浓度优选为20～60％。
在一种具体实施方式中，所述锂渣中钽的品位以Ta 2O 5计、铌的品位以Nb 2O5计，低于100ppm，优选50～100ppm。
本发明锂渣中钽铌的氧化物品位低于100ppm还能回收，且其回收率能够保证大于45％；而现有的方法，由于品位太低，导致钽铌回收率低，或放弃回收；钽铌品位高于100ppm时，本发明的方法当然也能很好的回收，且回收率更高。
在一种具体实施方式中，所述弱磁分离的磁场强度为100～2000高斯，优选300～1000高斯；所述强磁分离的磁场强度为10000～20000高斯，优选12000～17000高斯。
在一种具体实施方式中，b步骤所述浮选还添加了调整剂，所述调整剂为：
铝溶胶、焦磷酸钠、聚环氧琥珀酸或其盐、聚天冬氨酸或其盐、羧酸-磺酸盐共聚物TH-2000、羧酸-磺酸-非离子三元共聚物TH-3100、膦酰基羧酸共聚物POCA、聚丙烯酸或其盐、马来酸-丙烯酸共聚物钠盐、丹宁、壳聚糖、羧甲基纤维素钠中的至少一种，优选铝溶胶、焦磷酸钠、聚丙烯酸或其盐、羧酸-磺酸盐共聚物TH-2000、丹宁，所述调整剂的用量优选为0～6000g/t锂渣，更优选为500～3000g/t锂渣。
在一种具体实施方式中，b步骤所述捕收剂的用量为50～3000g/t，优选为100～1000g/t锂渣。
在一种具体实施方式中，所述浮选包括粗选、扫选和精选，优选所述粗选1～3次，扫选1～4次，精选1～3次；扫选过程捕收剂用量为粗选捕收剂用量的1/20～13/12，精选过程不添加捕收剂；优选所述扫选一捕收剂用量为粗选用量的1/2，扫选二捕收剂用量为粗选用量的1/3，扫选三捕收剂用量为粗选用量的1/4。
在一种具体实施方式中，c步骤所述粉碎后尾矿2粒度325目以下，所述粉碎优选为将尾矿2分级得到325目以上、325目以下的微粒，将325目以上的微粒粉碎后与325目以下的微粒混合；所述粉碎优选采用非铁介质磨机细磨。
325目约45μm，325目以下，是约45μm以下。325目以上是约45μm以上。
在一种具体实施方式中，e步骤所述的干燥前还先进行了浓缩-过滤。
在一种具体实施方式中，a步骤所述的重选包括粗选和精选，优选粗选1～3次，精选1～3次；
f步骤所述的重选包括粗选和精选，优选粗选1～3次，精选1～3次。
在一种具体实施方式中，所述捕收剂中C8-20的脂肪酸及其盐包括辛酸、壬酸、 癸酸、十一酸、十二酸、十三酸、十四酸、十五酸、十六酸、十七酸、十八酸、十九酸、二十酸、油酸、亚油酸、亚麻酸、花生四烯酸中的至少一种；
所述航空煤油包括航空煤油宽馏分型1～10份；所述航空煤油包括航空煤油优选还包括煤油型1～10份；重馏分型1～10份；
所述单甘油脂肪酸酯包括甘油油酸酯、甘油硬脂酸酯、甘油月桂酸酯、甘油棕榈酸酯中的至少一种；优选包括甘油月桂酸酯；
所述季铵盐包括十二～十六烷基三甲基氯化铵或溴化铵；优选为十二、十四或十六烷基三甲基氯化铵或溴化铵；更优选为十二烷基三甲基氯化铵或溴化铵；
所述碱为氢氧化钠、氢氧化钾、碳酸钠、碳酸钾、碳酸氢钠、碳酸氢钾中的至少一种；所述盐为钠盐、钾盐、铵盐、钙盐、镁盐中的至少一种。
航空煤油的煤油型也叫航空煤油中馏分型，沸点150℃～280℃，重馏分型沸点190℃～315℃，宽馏分型沸点60℃～280℃。
在一种具体实施方式中，所述的锂辉石硫酸法提锂尾渣浮选脱硫捕收剂采用如下方法制备得到：
a.将所述碱和硅溶胶按照质量比混合，50～80℃搅拌0.5～24h，反应得到试剂A；
将除碱和硅溶胶以外的其余全部组分按照质量比混合在一起，在80～100℃条件下搅拌1～2h，反应得到试剂B；
b.将试剂A和试剂B混合均匀得到膏状的锂辉石硫酸法提锂尾渣浮选脱硫捕收剂。
有益效果：
1.本发明彻底实现了锂渣深度加工产品多元化和高值化利用的目的，解决了困扰锂盐行业的渣难处理的重大难题；
2.本发明可以获得高硅高铝低铁低硫的硅铝微粉，该硅铝微粉能够用于玻纤、陶瓷、造纸等行业，用于替代叶蜡石、高岭土、滑石等原料，极大的降低了玻纤、陶瓷和造纸行业的生产成本；
3.本发明通过浮选获得了高品质的石膏精矿，石膏纯度高达95％以上，属于高纯石膏，石膏既能够当腻子粉使用，又能够用来开发晶须石膏材料、涂料、模具用料等，提高了石膏的价值；
4.本发明充分利用资源特性，通过弱磁分离获得了铁精矿，进一步提高了锂渣综合利用的价值；
5.本发明获得了钽铌精矿，以钽铌氧化物含量150ppm计算，锂渣年产量300万吨，则钽铌氧化物总量接近450吨。
6.本发明获得了高铁富锂渣，高铁富锂渣中氧化锂(Li 2O)含量1.0～1.5％，高铁富锂渣产率在5-10％左右，以7％产率计算，则一年产生高铁富锂渣21万吨，约2000-3000吨锂金属，高铁富锂渣可作为锂矿进一步回收碳酸锂。
7.本发明浮选剩余的尾矿中SO 3含量少，对于SO 3含量大于10％的锂渣，浮选能够获得SO 3含量小于0.1％的尾矿。
附图说明
图1为本发明的一种具体实施方式的工艺流程图。
图2为本发明的一种高铁富锂料回收提锂的工艺流程图。
图3为本发明的一种高铁富锂料酸法焙烧综合回收工艺图。
具体实施方式
为解决本发明的第一个技术问题，本发明的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法包括：
a.将锂渣重选得到精矿1和尾矿1，所述精矿1弱磁分离得到粗粒钽铌富料和粗粒铁精矿；
b.将所述尾矿1进行浮选，获得石膏和尾矿2；
c.将所述尾矿2粉碎；
d.将c步骤粉碎后的尾矿进行弱磁分离，得到细粒铁精矿和尾矿3；
e.将尾矿3强磁分离得到精矿2和尾矿4，尾矿4干燥即得硅铝微粉；
f.将步骤e中精矿2重选得到细粒钽铌精矿和高铁富锂料；
其中，所述锂渣为辉锂石提锂渣；a步骤所述重选为摇床重选、螺旋重选、离心重选、水力旋流器重选、跳汰机重选、风力重选、重介质重选中的一种或多种组合；
b步骤所述浮选的捕收剂按重量计包括：
C8-20的脂肪酸及其盐中的至少一种50～100份；航空煤油1～30份；十二烷基的磺酸或硫酸及其盐中的至少一种1～30份；聚醚或聚醇中的至少一种1～30份；环氧丙烷嵌段共聚物1～10份；山梨醇单油酸酯1～10份；单甘油脂肪酸酯1～10份；季铵盐1～30份；十六烷基卤化吡啶1～10份；碱5～50份；硅溶胶10～50份；水10～100份；
所述聚醚或聚醇为聚乙烯醚、聚氧丙烯醚、聚乙烯醇、聚氧乙烯醚中的至少一种， 优选为聚乙烯醚1～10份、聚氧丙烯醚1～10份、聚乙烯醇1～10份；
所述环氧丙烷嵌段共聚物为PE6100、PE6200、PE6400、PE8100中的至少一种；
所述十二烷基的磺酸或硫酸包括十二烷基苯磺酸、十二烷基磺酸、十二烷基硫酸；优选包括十二烷基苯磺酸及其盐；更优选所述十二烷基苯磺酸及其盐1～10份；
所述硅溶胶质量浓度优选为5～40％；
所述浮选的矿浆浓度优选为20～60％。
捕收剂的各组分配比在相应的范围内按照任意比例配比，按照相应配比就能够实现锂渣高效浮选脱硫的目的，保证石膏含量大于95％，石膏中杂质SiO 2含量＜1％、Al 2O 3含量＜1％，为后续制备晶须石膏提供高品质原料。
需要指出的是，本发明的脱硫捕收剂的配置，通过有效调配各组分的含量，能够实现锂渣中硫快速、高效脱出；通过本发明的工艺，能够轻松获得高品质石膏。
值得一提的是，本发明浮选获得的石膏经过过滤后，可以直接作为生产石膏晶须的原料、石膏腻子粉或填料。浮选产生的过滤水，通过收集后继续返回浮选作业，本发明不产生废水排放；鉴于产品中石膏要带走部分水，最终本发明生产工艺需要补充新水，以保证生产正常进行。
在一种具体实施方式中，所述锂渣中钽的品位以Ta 2O 5计、铌的品位以Nb 2O5计，低于100ppm，优选50～100ppm。
本发明锂渣中钽铌的氧化物品位低于100ppm还能回收，其回收率能够保证大于45％；而现有的方法，由于品位太低，导致钽铌回收率低，或放弃回收；钽铌品位高于100ppm本发明的方法当然也能很好的回收，且回收率更高。
在一种具体实施方式中，所述弱磁分离的磁场强度为100～2000高斯，优选300～1000高斯；所述强磁分离的磁场强度为10000～20000高斯，优选12000～17000高斯。
在一种具体实施方式中，b步骤所述浮选还添加了调整剂，所述调整剂为：
铝溶胶、焦磷酸钠、聚环氧琥珀酸或其盐、聚天冬氨酸或其盐、羧酸-磺酸盐共聚物TH-2000、羧酸-磺酸-非离子三元共聚物TH-3100、膦酰基羧酸共聚物POCA、聚丙烯酸或其盐、马来酸-丙烯酸共聚物钠盐、丹宁、壳聚糖、羧甲基纤维素钠中的至少一种，优选铝溶胶、焦磷酸钠、聚丙烯酸或其盐、羧酸-磺酸盐共聚物TH-2000、丹宁，所述调整剂的用量优选为0～6000g/t锂渣，更优选为500～3000g/t锂渣。
在一种具体实施方式中，b步骤所述捕收剂的用量为50～3000g/t，优选为100～1000g/t锂渣。
在一种具体实施方式中，所述浮选包括粗选、扫选和精选，优选所述粗选1～3次，扫选1～4次，精选1～3次；扫选过程捕收剂用量为粗选捕收剂用量的1/20～12/13，精选过程不添加捕收剂；优选所述扫选一捕收剂用量为粗选用量的1/2，扫选二捕收剂用量为粗选用量的1/3，扫选三捕收剂用量为粗选用量的1/4。
在一种具体实施方式中，c步骤所述粉碎后尾矿2的目数325目以下，所述粉碎优选为将尾矿2分级得到目数325目以上、目数325目以下的微粒，将目数325目以上的微粒粉碎后与目数325目以下的微粒混合；所述粉碎优选采用非铁介质磨机细磨。
分级设备可采用螺旋分级机、旋流器、直线筛等。
在一种具体实施方式中，e步骤所述的干燥前还先进行了浓缩-过滤。
在一种具体实施方式中，a步骤所述的重选包括粗选和精选，优选粗选1～3次，精选1～3次；
f步骤所述的重选包括粗选和精选，优选粗选1～3次，精选1～3次。
在一种具体实施方式中，所述捕收剂中C8-20的脂肪酸及其盐包括辛酸、壬酸、癸酸、十一酸、十二酸、十三酸、十四酸、十五酸、十六酸、十七酸、十八酸、十九酸、二十酸、油酸、亚油酸、亚麻酸、花生四烯酸中的至少一种；
所述航空煤油包括航空煤油宽馏分型1～10份；所述航空煤油包括航空煤油优选还包括煤油型1～10份；重馏分型1～10份；
所述单甘油脂肪酸酯包括甘油油酸酯、甘油硬脂酸酯、甘油月桂酸酯、甘油棕榈酸酯中的至少一种；优选包括甘油月桂酸酯；
所述季铵盐包括十二～十六烷基三甲基氯化铵或溴化铵；优选为十二、十四或十六烷基三甲基氯化铵或溴化铵；更优选为十二烷基三甲基氯化铵或溴化铵；
所述碱为氢氧化钠、氢氧化钾、碳酸钠、碳酸钾、碳酸氢钠、碳酸氢钾中的至少一种；所述盐为钠盐、钾盐、铵盐、钙盐、镁盐中的至少一种。
航空煤油的煤油型也叫航空煤油中馏分型，沸点150℃～280℃，重馏分型沸点190℃～315℃，宽馏分型沸点60℃～280℃。
在一种具体实施方式中，所述的锂辉石硫酸法提锂尾渣浮选脱硫捕收剂采用如下方法制备得到：
a.将所述碱和硅溶胶按照质量比混合，50～80℃搅拌0.5～24h，反应得到试剂A；
将除碱和硅溶胶以外的其余全部组分按照质量比混合在一起，在80～100℃条件下搅拌1～2h，反应得到试剂B；
b.将试剂A和试剂B混合均匀得到膏状的锂辉石硫酸法提锂尾渣浮选脱硫捕收剂。
在一种具体实施方式中，C8-20的脂肪酸及其盐中的至少一种50～100份；航空煤油1～15份；十二烷基的磺酸或硫酸及其盐中的至少一种1～15份；聚醚或聚醇中的至少一种1～15份；环氧丙烷嵌段共聚物1～5份；山梨醇单油酸酯1～5份；单甘油脂肪酸酯1～5份；季铵盐1～20份；十六烷基卤化吡啶1～5份；碱10～50份；硅溶胶10～50份；水10～50份；
优选航空煤油宽馏分型1～5份；煤油型1～5份；重馏分型1～5份；
优选十二烷基苯磺酸及其盐1～5份；十二烷基硫酸及其盐1～5份；十二烷基磺酸及其盐1～5份；
优选聚乙烯醚1～5份；聚氧丙烯醚1～5份；聚乙烯醇1～5份；
优选十二烷基三甲基氯化铵或十二烷基三甲基溴化铵1～5份；十四烷基三甲基氯化铵或十四烷基三甲基溴化铵1～10份、十六烷基三甲基氯化铵或十六烷基三甲基溴化铵1～5份。
在一种具体实施方式中，C8-20的脂肪酸及其盐中的至少一种50～100份；航空煤油1～6份；十二烷基的磺酸或硫酸及其盐中的至少一种1～3份；聚醚或聚醇中的至少一种1～6份；环氧丙烷嵌段共聚物1～2份；山梨醇单油酸酯1份；单甘油脂肪酸酯1份；季铵盐1～3份；十六烷基卤化吡啶1份；碱15～20份；硅溶胶10份；水40～50份；
优选航空煤油宽馏分型1～2份；煤油型1～2份；重馏分型1～2份；
优选十二烷基苯磺酸及其盐1份；十二烷基硫酸及其盐1份；十二烷基磺酸及其盐1份；
优选聚乙烯醚1～2份；聚氧丙烯醚1～2份；聚乙烯醇1～2份；
优选十二烷基三甲基氯化铵或十二烷基三甲基溴化铵1份；十四烷基三甲基氯化铵或十四烷基三甲基溴化铵1份、十六烷基三甲基氯化铵或十六烷基三甲基溴化铵1份。
在一种具体实施方式中，所述方法还包括将所述高铁富锂料进行碱法或酸法焙烧-浸出-固液分离-将固液分离后的滤液净化、浓缩工艺后获得锂盐产品。
如图2所示高铁富锂料可以采用现有锂辉石提锂相似的提锂工艺，高铁富锂料与锂辉石不同的是不用进行转型焙烧，直接进行碱法或酸法焙烧。碱法或酸法焙烧工艺， 酸法焙烧采用的酸为浓硫酸，原则上选择80％以上的浓硫酸，优选98％的浓硫酸。如图3所示固液分离后的固体即为硅铝微粉，可以回收利用。
下面结合实施例对本发明的具体实施方式做进一步的描述，并不因此将本发明限制在所述的实施例范围之中。
实施例1
如图1所示采用四川某公司锂辉石提锂渣，主要矿物石英、方解石、石膏、三水铝石、红柱石、刚玉、玻璃相、α锂辉石、β锂辉石、沸石、正长石，钽铁矿(微量)、铌铁矿(微量)。原料Ta 2O 5品位为90ppm，Nb 2O 5品位50ppm，SO 3品位为6.2％。
(一)重选-弱磁分离
原料直接采用螺旋溜槽进行重选，螺旋溜槽重选精矿进入摇床精选，摇床精矿直接进行弱磁分离，磁场强度为1000高斯，磁选尾矿为粗粒钽铌精矿，Ta 2O 5品位18.56％、Nb 2O 5品位9.56％，钽铌回收率分别为46.12％和32.68％；磁选精矿为粗粒级铁精矿，TFe为52.13％，回收率为12.89％。
(二)浮选脱硫
制备捕收剂：首先，将20份氢氧化钠和50份质量分数为40％的硅溶胶混合，加热到80℃搅拌5h，得到膏状物A；
其次，将C 8-20的脂肪酸/脂肪酸盐100份(本实施例的C 8-20的脂肪酸/脂肪酸盐是辛酸和月桂酸按照1：1混合)、航空煤油宽馏分型1份、航空煤油中馏分型1份、航空煤油重馏分型1份、十二烷基苯磺酸钠1份、十二烷基硫酸钠1份、聚乙烯醚1份、聚氧丙烯醚1份、聚乙烯醇1份、环氧乙烷－环氧丙烷嵌段共聚物EO-PO-EO(本次实验采用的是PE6100)1份、山梨醇单油酸酯1份、单甘油油酸酯(本次实验的单甘油油酸酯是甘油油酸酯、甘油硬脂酸酯、甘油月桂酸酯、甘油棕榈酸酯，各组分按照1：1混合)1份、十二烷基氯化铵1份、十六烷基三甲基氯化铵1份、十六烷基氯化吡啶1份、水50份完全混合均匀，加热到80℃并搅拌2h，得到膏状物B。
最后，将膏状物A和膏状物B混合均匀，即得到锂渣脱硫捕收剂C。
重选尾矿直接进入浮选，调整矿浆浓度为35％，按照吨给矿计，依次加入2000g调整剂、300g捕收剂进行第一次粗选；第二次粗选调整剂用量为500g，捕收剂用量为100g，粗选1精矿和粗选2精矿混合得到粗选精矿，浮选粗选中调整剂为铝溶胶。粗选精矿精选一次，精选中矿返回第一次粗选；粗选尾矿进行三次扫选作业，第一次扫选调整剂用量为500g，捕收剂用量为50g，第二次扫选调整剂用量为250g，捕收剂用 量为30g，第三次扫选调整剂用量为250g，捕收剂用量为20g，扫选中矿返回第一次粗选作业，最终形成闭路循化，得到脱硫石膏精矿和脱硫尾矿，石膏精矿中CaSO 4.2H 2O纯度大于95％。
(三)分级-磨矿-弱磁分离
浮选尾矿分级为-325目的样品和+325目的样品；其中+325目的样品直接进入陶瓷磨机，磨矿至-325目占100％；将尾矿分级-325目样品与磨矿后-325目的样品混合，直接进入弱磁分离，磁场强度为2000高斯，得到细粒级铁精矿，TFe为42.23％，收率8.2％。
(四)强磁分离-重选
弱磁分离后的矿浆直接进入强磁选机，强磁分离粗选磁场强度为1.2T，扫选为1.7T，粗选和扫选混合，强磁选后分别得到磁选尾矿和富钽铌锂磁性产品。磁选尾矿直接进行浓缩、过滤、干燥得到硅铝微粉1，产率75％，SO 3品位0.15％，Fe 2O 3品位0.32％。富钽铌锂磁性产品直接进入摇床重选，第一次重选精矿直接进行精选，第一次重选中矿返回重选，第一次重选尾矿直接作为尾矿；第一次重选精矿进行第二次精选，第二次重选精矿为最终细粒钽铌精矿，第二次重选中矿和尾矿直接返回第一次重选。重选后分别得到细粒钽铌精矿和高铁富锂料，细粒钽铌精矿Ta 2O 5品位10.52％、Nb 2O 5品位4.78％，钽铌回收率分别为14.73％和18.79％，高铁富锂料Li 2O品位1.58％，锂回收率为25％。
(五)高铁富锂料硫酸法提锂
如图2和3所示，取1000克高铁富锂料与50克98％的浓硫酸混合，置于马弗炉中，300℃条件下恒温焙烧2h；将焙烧料冷却后与水按照固液质量比为1：1混合浸出，搅拌浸出2h，浸出温度为40℃；固液分离后得到滤液和硅铝微粉2，滤液中加入5ml双氧水，氧化反应0.5h，添加碳酸钙调节pH值到3，过滤得到净化锂液和钙铁渣(水泥缓凝剂)。高铁富锂料硫酸法提锂全流程硅铝微粉产率92％，再经浓缩-脱钠，沉锂等工艺后得到碳酸锂产品，Li 2O全流程回收率为20.5％(作业回收率为82％)。
实施例2
其它与实施例1相似，唯一不同的是航空煤油宽馏分型5份；航空煤油中馏分型3份、航空煤油重馏分型3份。
实施例3
其它与实施例1相似，唯一不同的是不添加调整剂。
实施例4-5
其它与实施例1相似，唯一不同的是实施例4的调整剂为壳聚糖。实施例4的调整剂为羧甲基纤维素钠。
实施例6-7
其它与实施例1相似，实施例6唯一不同的是弱磁分离的磁场强度为1500高斯；实施例7唯一不同的是强磁分离的磁场强度为11000高斯；
实施例8-9
其它与实施例1相似，实施例8唯一不同的是第一次扫选捕收剂用量为500g/t，第二次扫选捕收剂用量为250g/t，第三次扫选捕收剂用量为100g/t；实施例9唯一不同的是第一次扫选调整剂用量为800g/t，第二次扫选调整剂用量为500g/t，第三次扫选调整剂用量为200g/t。
表1实施例1-9粗粒钽铌精矿品位及收率
实施例 Ta 2O 5品位％ Nb 2O 5品位％ Ta 2O 5收率％ Nb 2O 5收率％
1 18.56 9.56 46.12 32.68
2 19.88 10.51 47.22 33.69
3 18.06 9.16 45.35 31.68
4 20.51 9.92 46.17 33.68
5 19.51 10.51 46.12 33.98
6 19.51 9.26 45.42 31.68
7 18.51 9.99 45.45 32.66
8 17.56 10.51 44.19 32.64
9 18.59 9.88 46.17 30.69
表2实施例1-9细粒钽铌精矿品位及收率
实施例 Ta 2O 5品位％ Nb 2O 5品位％ Ta 2O 5收率％ Nb 2O 5收率％
1 10.52 4.78 14.73 18.79
2 11.5 4.99 13.41 17.99
3 11.12 4.98 14.71 18.78
4 11.12 5.18 13.93 16.79
5 10.12 4.18 14.03 16.09
6 9.98 4.98 14.79 18.77
7 10.11 4.95 14.78 17.79
8 12.52 4.08 15.93 17.47
9 10.59 4.98 14.73 18.99
表3实施例1-9铁精矿品位及收率
Figure PCTCN2022133160-appb-000001
表4实施例1-9硅铝微粉、碳酸锂品位及收率
Figure PCTCN2022133160-appb-000002
Figure PCTCN2022133160-appb-000003
表5实施例1-9石膏品位及收率
Figure PCTCN2022133160-appb-000004

Claims (31)

  1. 锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，所述方法包括：
    a.将锂渣重选得到精矿1和尾矿1，所述精矿1弱磁分离得到粗粒钽铌富料和粗粒铁精矿；
    b.将所述尾矿1进行浮选，获得石膏和尾矿2；
    c.将所述尾矿2粉碎；
    d.将c步骤粉碎后的尾矿进行弱磁分离，得到细粒铁精矿和尾矿3；
    e.将尾矿3强磁分离得到精矿2和尾矿4，尾矿4干燥即得硅铝微粉；
    f.将步骤e中精矿2重选得到细粒钽铌精矿和高铁富锂料；
    其中，所述锂渣为锂辉石提锂渣；a步骤所述重选为摇床重选、螺旋重选、离心重选、水力旋流器重选、跳汰机重选、风力重选、重介质重选中的一种或多种组合；
    b步骤所述浮选的捕收剂按重量计包括：
    C8-20的脂肪酸及其盐中的至少一种50～100份；航空煤油1～30份；十二烷基的磺酸或硫酸及其盐中的至少一种1～30份；聚醚或聚醇中的至少一种1～30份；环氧丙烷嵌段共聚物1～10份；山梨醇单油酸酯1～10份；单甘油脂肪酸酯1～10份；季铵盐1～30份；十六烷基卤化吡啶1～10份；碱5～50份；硅溶胶10～50份；水10～100份；
    所述聚醚或聚醇为聚乙烯醚、聚氧丙烯醚、聚乙烯醇、聚氧乙烯醚中的至少一种；
    所述环氧丙烷嵌段共聚物为PE6100、PE6200、PE6400、PE8100中的至少一种；
    所述十二烷基的磺酸或硫酸包括十二烷基苯磺酸、十二烷基磺酸、十二烷基硫酸；。
  2. 根据权利要求1所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，所述聚醚或聚醇为聚乙烯醚1～10份、聚氧丙烯醚1～10份、聚乙烯醇1～10份。
  3. 根据权利要求1或2所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，所述十二烷基的磺酸或硫酸包括十二烷基苯磺酸及其盐。
  4. 根据权利要求3所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，所述十二烷基苯磺酸及其盐1～10份。
  5. 根据权利要求1或2所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏 的方法，其特征在于，所述硅溶胶质量浓度为5～40％。
  6. 根据权利要求1或2所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，所述浮选的矿浆浓度为20～60％。
  7. 根据权利要求1或2所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，所述锂渣中钽的品位以Ta 2O 5计、铌的品位以Nb 2O 5计低于100ppm。
  8. 根据权利要求7所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，所述锂渣中钽的品位以Ta 2O 5计、铌的品位以Nb 2O 5计为50～100ppm。
  9. 根据权利要求1或2所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，所述弱磁分离的磁场强度为100～2000高斯；所述强磁分离的磁场强度为10000～20000高斯。
  10. 根据权利要求9所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，所述弱磁分离的磁场强度为300～1000高斯。
  11. 根据权利要求9所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，所述强磁分离的磁场强度为12000～17000高斯。
  12. 根据权利要求1或2所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，b步骤所述浮选还添加了调整剂，所述调整剂为：
    铝溶胶、焦磷酸钠、聚环氧琥珀酸或其盐、聚天冬氨酸或其盐、羧酸-磺酸盐共聚物TH-2000、羧酸-磺酸-非离子三元共聚物TH-3100、膦酰基羧酸共聚物POCA、聚丙烯酸或其盐、马来酸-丙烯酸共聚物钠盐、丹宁、壳聚糖、羧甲基纤维素钠中的至少一种。
  13. 根据权利要求12所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，所述调整剂为：铝溶胶、焦磷酸钠、聚丙烯酸或其盐、羧酸-磺酸盐共聚物TH-2000、丹宁。
  14. 根据权利要求12所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，所述调整剂的用量为0～6000g/t锂渣。
  15. 根据权利要求14所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，所述调整剂的用量为500～3000g/t锂渣。
  16. 根据权利要求1或2所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，b步骤所述捕收剂的用量为50～3000g/t。
  17. 根据权利要求16所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，b步骤所述捕收剂的用量为100～1000g/t锂渣。
  18. 根据权利要求16所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，所述浮选包括粗选、扫选和精选；扫选过程捕收剂用量为粗选捕收剂用量的1/20～13/12，精选过程不添加捕收剂。
  19. 根据权利要求18所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，所述粗选1～3次，扫选1～4次，精选1～3次。
  20. 根据权利要求19所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，所述扫选一捕收剂用量为粗选用量的1/2，扫选二捕收剂用量为粗选用量的1/3，扫选三捕收剂用量为粗选用量的1/4。21.根据权利要求1或2所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，c步骤所述粉碎后尾矿2粒度325目以上。
  21. 根据权利要求21所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，c步骤所述粉碎为将尾矿2分级得到325目以上、325目以下的微粒，将325目以上的微粒粉碎后与325目以下的微粒混合。
  22. 根据权利要求21所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，所述粉碎采用非铁介质磨机细磨。
  23. 根据权利要求1或2所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，e步骤所述的干燥前还先进行了浓缩-过滤。
  24. 根据权利要求1或2所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，a步骤所述的重选包括粗选和精选；
    f步骤所述的重选包括粗选和精选。
  25. 根据权利要求25所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，a步骤所述的重选包括粗选1～3次，精选1～3次。
  26. 根据权利要求25所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，f步骤所述的重选包括粗选1～3次，精选1～3次。
  27. 根据权利要求1或2所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，所述捕收剂中C8-20的脂肪酸及其盐包括辛酸、壬酸、癸酸、十一酸、十二酸、十三酸、十四酸、十五酸、十六酸、十七酸、十八酸、十九酸、二十酸、油酸、亚油酸、亚麻酸、花生四烯酸中的至少一种；
    所述航空煤油包括航空煤油宽馏分型1～10份；所述单甘油脂肪酸酯包括甘油油酸酯、甘油硬脂酸酯、甘油月桂酸酯、甘油棕榈酸酯中的至少一种；
    所述季铵盐包括十二～十六烷基三甲基氯化铵或溴化铵；
    所述碱为氢氧化钠、氢氧化钾、碳酸钠、碳酸钾、碳酸氢钠、碳酸氢钾中的至少一种；所述盐为钠盐、钾盐、铵盐、钙盐、镁盐中的至少一种。
  28. 根据权利要求28所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，所述航空煤油包括航空煤油还包括煤油型1～10份；重馏分型1～10份。
  29. 根据权利要求28所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，所述单甘油脂肪酸酯包括甘油月桂酸酯。
  30. 根据权利要求28所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，所述季铵盐为十二、十四或十六烷基三甲基氯化铵或溴化铵。
  31. 根据权利要求28所述的锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法，其特征在于，所述季铵盐为十二烷基三甲基氯化铵或溴化铵。
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CN113976309A (zh) * 2021-12-01 2022-01-28 天齐创锂科技(深圳)有限公司 锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的方法
CN216459396U (zh) * 2021-12-01 2022-05-10 天齐创锂科技(深圳)有限公司 锂渣综合回收锂、钽铌、硅铝微粉、铁精矿和石膏的系统

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CN116462522A (zh) * 2023-06-19 2023-07-21 湖南永杉锂业有限公司 一种制备匣钵的方法
CN116462522B (zh) * 2023-06-19 2023-08-22 湖南永杉锂业有限公司 一种制备匣钵的方法

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