CN113976309B - Method for comprehensively recovering lithium, tantalum-niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum from lithium slag - Google Patents
Method for comprehensively recovering lithium, tantalum-niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum from lithium slag Download PDFInfo
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- 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
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Abstract
The invention relates to a method for comprehensively recovering lithium, tantalum, niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum from lithium slag, and belongs to the technical field of lithium slag treatment. The method for comprehensively recovering lithium, tantalum, niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum from lithium slag comprises the following steps: reselecting the lithium slag to obtain concentrate 1 and tailings 1, and carrying out weak magnetic separation on the concentrate 1 to obtain coarse-grain tantalum-niobium rich materials and coarse-grain iron concentrate; carrying out flotation on the tailings 1 to obtain gypsum and tailings 2; crushing the tailings 2; carrying out weak magnetic separation on the crushed tailings to obtain fine iron ore concentrate and tailings 3; strongly magnetic separating the tailings 3 to obtain concentrate 2 and tailings 4, and drying the tailings 4 to obtain silicon-aluminum micro powder; and (4) reselecting the concentrate 2 to obtain fine-grained tantalum-niobium concentrate and a high-iron lithium-rich material. The invention solves the great problem that slag which troubles the lithium salt industry is difficult to treat; can obtain high-silicon high-aluminum low-iron low-sulfur silicon-aluminum micro powder, gypsum concentrate with the purity of more than 95 percent, iron concentrate, tantalum-niobium concentrate and high-iron lithium-rich slag.
Description
Technical Field
The invention relates to a method for comprehensively recovering lithium, tantalum-niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum from lithium slag, and belongs to the technical field of lithium slag treatment.
Background
The comprehensive recycling of the secondary resources of the solid wastes is a major theme of environmental protection in the world, and the comprehensive recycling of the solid wastes is beneficial to relieving the shortage of national resources, moves towards a sustainable development road and is a necessary way for creating a conservation-oriented society. In recent years, with support of policy for comprehensive recycling of secondary resources, comprehensive recycling of solid wastes has achieved significant results. Along with the development of economic society, the problem that the economic development is lagged due to the improvement of the environmental and ecological protection requirements of people and the shortage of resources is obvious, and the utilization of solid waste resources has huge development prospect.
At present, the demand for lithium resources is increasing due to the rapid development of the lithium battery industry. The extraction of lithium salt mainly depends on ore and salt lake, wherein the extraction of lithium from the salt lake is difficult to compare with the extraction of lithium from ore due to high cost, high impurity content and the like. Spodumene is mainly adopted for extracting lithium from ores, and a lot of spodumene resources cannot be put into production on a large scale due to various reasons. And extracting lithium salt by using spodumene, wherein 7-8 tons of slag are produced when 1 ton of lithium salt is produced, and more than 200 ten thousand tons of lithium slag are produced in one year according to the existing lithium salt yield. At present, spodumene lithium extraction slag is mainly used as an ingredient in low-added-value fields such as cement and concrete, so that the spodumene lithium extraction slag cannot be rapidly consumed. The stacking of the spodumene lithium extraction slag undoubtedly brings problems of environmental pollution, land occupation and the like, and the demand for comprehensive utilization of the spodumene lithium extraction slag is more urgent along with the rapid development of the lithium battery industry.
Patent CN1297860A and patent CN1090597C disclose ceramic glazed tile made from acidic lithium slag and a manufacturing method thereof, wherein the acidic lithium slag is used as a main raw material, wollastonite, pyrophyllite and kaolin are used as auxiliary materials, and the steps of grinding, pulping, filter pressing, mud cake, drying, crushing, green pressing, drying, biscuit firing, glaze firing and the like are carried out, so that the technology for manufacturing the ceramic glazed tile by replacing a part of traditional high-quality mineral raw materials with acidic lithium slag is finally developed. These two patents have no technical advantage in rapidly dissolving a large amount of lithium slag because only a small amount of lithium slag is used. Patent CN103601230 discloses a method for producing chemical raw materials by comprehensively utilizing lithium slag, calcium chloride, ammonium fluoride, white carbon black, aluminum salt and ammonium sulfate are finally obtained through a plurality of steps, and the patent cannot avoid using a large amount of acid solution, so that the acidity in the discharged gas is high, and the difficulty of production operation is high.
Patent CN108273826A discloses a full-phase high-valued recycling method of lithium slag, which mainly adopts an alkali conversion-magnetic separation mode to obtain pyrophyllite raw materials for glass fibers, and byproducts are gypsum and magnetic separation tailings. Patent CN108147658A and patent WO2019/141098a1 mainly describe that lithium pyrophyllite, gypsum and magnetic separation tailings are obtained by a flotation-magnetic separation process, and although the lithium slag is subjected to high-valued preparation of pyrophyllite, resources such as calcium sulfate, tantalum, niobium and lithium in the pyrophyllite are not subjected to high-valued preparation. Patent CN214488258U discloses a comprehensive recycling system for lithium slag, which mainly adopts a preposed ore grinding and water circulation grading rough technology to replace pulping operation, and adopts a flotation and alkali-to-solid-liquid two-phase desulfurization technology, a weak-strong magnetic separation iron removal technology and a water circulation sectional circulation technology to comprehensively utilize waste slag generated after lithium is extracted by a lithium ore sulfuric acid process. This patent provides only one system and has the following disadvantages: the separation difficulty of the lithium slag and the gypsum is increased due to the preposed ore grinding, and the final result is that the yield of the silicon-aluminum micro powder in the flotation operation is low and the cost is high; in addition, the alkali conversion has the characteristics of high cost, long conversion time, low efficiency and the like, and is difficult to realize large-scale production; secondly, the system does not recover iron and lithium in the lithium slag, so that resource waste is undoubtedly caused, and comprehensive recycling is not realized.
Chinese patent CN108191226A discloses a method for producing glass fiber by using spodumene slag as fluxing clarifier, which comprises the following raw materials in proportion: 100-120 parts of kaolin, 150-410 parts of pyrophyllite, 150-200 parts of quick lime, 50-70 parts of dolomite, 50-70 parts of colemanite, 130-310 parts of white sepiolite, 20-30 parts of fluorite and 10-30 parts of spodumene slag. Chinese patent CN1114232223A discloses a method for preparing ceramic by replacing kaolin with spodumene slag, wherein the ceramic blank comprises the following raw materials in percentage by weight: 50-75% of spodumene slag, 10-20% of quartz sand, 1-10% of potassium feldspar and 1-10% of albite; the ceramic glaze comprises the following raw materials in percentage by weight: 40-60% of spodumene slag, 15-40% of quartz sand, 15-20% of feldspar and 1-10% of porcelain stone. Chinese patent CN113480182A discloses a glass fiber using industrial waste as main raw material and its preparation method, the raw material components of the glass fiber are: the invention provides glass fiber taking industrial waste residue as a main raw material and a preparation scheme thereof, and the glass fiber replaces raw materials such as pyrophyllite, kaolin, quartz and the like mainly used in the current glass fiber production by reasonably using industrial wastes, thereby reducing the industrial risk of the raw materials for producing the glass fiber. CN1090597C discloses a method for manufacturing ceramic glazed tiles by using acid lithium slag, which has the defects of high cost, difficult industrialization, slow lithium slag absorption speed and the like.
Patent CN1112335C provides a technology for preparing a gypsum reinforcing agent by utilizing spodumene acid method lithium extraction waste residues, and patent CN106082739A provides a technology for taking a product obtained by mixing and drying the spodumene acid method lithium extraction waste residues as a cement admixture, and the patent technologies stay in a low-valued stage and do not better realize high-valued utilization of the spodumene lithium extraction waste residues.
In conclusion, if a technology for recycling lithium slag comprehensively can be developed, gypsum, tantalum-niobium, iron, silicon-aluminum micro powder and lithium in the lithium slag are all recycled, and the lithium slag is smoothly consumed, so that the development of the lithium industry can be carried out, and the healthy and rapid development of the lithium industry and the lithium slag treatment industry can be greatly promoted. All valuable components in the lithium slag are effectively utilized without generating solid wastes, so that the worries of the development of the lithium industry are solved at once.
Disclosure of Invention
The invention aims to solve the technical problem of providing a method for comprehensively recovering lithium, tantalum-niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum from lithium slag.
In order to solve the first technical problem, the method for comprehensively recovering lithium, tantalum-niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum from lithium slag comprises the following steps:
a. reselecting the lithium slag to obtain concentrate 1 and tailings 1, and carrying out weak magnetic separation on the concentrate 1 to obtain coarse-grain tantalum-niobium rich materials and coarse-grain iron concentrate;
b. carrying out flotation on the tailings 1 to obtain gypsum and tailings 2;
c. crushing the tailings 2;
d. c, carrying out weak magnetic separation on the tailings crushed in the step c to obtain fine iron ore concentrate and tailings 3;
e. strongly magnetic separating the tailings 3 to obtain concentrate 2 and tailings 4, and drying the tailings 4 to obtain silicon-aluminum micro powder;
f. reselecting the concentrate 2 in the step e to obtain fine-grained tantalum-niobium concentrate and a high-iron lithium-rich material;
wherein the lithium slag is the residue of extracting lithium from the glimmer; step a, the reselection is one or more of table reselection, spiral reselection, centrifugal reselection, hydrocyclone reselection, jigger reselection, wind power reselection and dense medium reselection;
b, the collector for flotation in the step comprises the following components in percentage by weight:
50-100 parts of at least one of C8-20 fatty acid and salt thereof; 1-30 parts of aviation kerosene; 1-30 parts of at least one of dodecyl sulfonic acid or sulfuric acid and salts thereof; 1-30 parts of at least one of polyether or polyalcohol; 1-10 parts of a propylene oxide block copolymer; 1-10 parts of sorbitol monooleate; 1-10 parts of monoglyceride; 1-30 parts of quaternary ammonium salt; 1-10 parts of hexadecyl halogenated pyridine; 5-50 parts of alkali; 10-50 parts of silica sol; 10-100 parts of water;
the polyether or the polyalcohol is at least one of polyvinyl ether, polyoxypropylene ether, polyvinyl alcohol and polyoxyethylene ether, and preferably 1-10 parts of polyvinyl ether, 1-10 parts of polyoxypropylene ether and 1-10 parts of polyvinyl alcohol;
the propylene oxide block copolymer is at least one of PE6100, PE6200, PE6400 and PE 8100;
the dodecyl sulfonic acid or sulfuric acid comprises dodecyl benzene sulfonic acid, dodecyl sulfonic acid and dodecyl sulfuric acid; preferably dodecyl benzene sulfonic acid and its salts; more preferably 1-10 parts of dodecyl benzene sulfonic acid and salts thereof;
the mass concentration of the silica sol is preferably 5-40%;
the concentration of the ore pulp subjected to flotation is preferably 20-60%.
In one embodiment, the grade of tantalum in the lithium slag is Ta2O5The grade of niobium is measured as Nb2O5Calculated by less than 100ppm, preferably 50 to 100 ppm.
The grade of the oxides of tantalum and niobium in the lithium slag is lower than 100ppm, the oxides can be recovered, and the recovery rate can be ensured to be higher than 45%; the existing method has low tantalum-niobium recovery rate or abandons the recovery due to low grade; when the grade of tantalum and niobium is higher than 100ppm, the method can also recover the tantalum and niobium well and has higher recovery rate.
In one embodiment, the magnetic field intensity of the weak magnetic separation is 100-2000 gauss, preferably 300-1000 gauss; the magnetic field intensity of the strong magnetic separation is 10000-20000 gausses, and 12000-17000 gausses are preferable.
In a specific embodiment, the flotation in the step b is further added with a modifier, wherein the modifier is:
the modifier is at least one of aluminum sol, sodium pyrophosphate, polyepoxysuccinic acid or salt thereof, polyaspartic acid or salt thereof, carboxylic acid-sulfonate copolymer TH-2000, carboxylic acid-sulfonic acid-nonionic terpolymer TH-3100, phosphonocarboxylic acid copolymer POCA, polyacrylic acid or salt thereof, maleic acid-acrylic acid copolymer sodium salt, tannin, chitosan and sodium carboxymethylcellulose, preferably aluminum sol, sodium pyrophosphate, polyacrylic acid or salt thereof, carboxylic acid-sulfonate copolymer TH-2000 and tannin, and the dosage of the modifier is preferably 0-6000 g/t of lithium slag, and more preferably 500-3000 g/t of lithium slag.
In a specific embodiment, the amount of the collector used in the step b is 50-3000 g/t, and preferably 100-1000 g/t of lithium slag.
In a specific embodiment, the flotation comprises roughing, scavenging and concentrating, preferably the roughing is performed for 1-3 times, the scavenging is performed for 1-4 times, and the concentrating is performed for 1-3 times; the using amount of the collecting agent in the scavenging process is 1/20-13/12 of that of the roughing collecting agent, and the collecting agent is not added in the concentration process; preferably, the dosage of the scavenger first collecting agent is 1/2 of the roughing dosage, the dosage of the scavenger second collecting agent is 1/3 of the roughing dosage, and the dosage of the scavenger third collecting agent is 1/4 of the roughing dosage.
In a specific embodiment, the particle size of the tailings 2 obtained after the grinding in the step c is 325 meshes or less, the grinding is preferably to classify the tailings 2 to obtain particles of 325 meshes or more and 325 meshes or less, and the particles of 325 meshes or more are mixed with the particles of 325 meshes or less after grinding; the pulverization is preferably carried out by adopting a non-ferrous medium mill.
About 45 μm in 325 mesh, and about 45 μm or less in 325 mesh. The size of the 325 mesh sieve is about 45 μm or more.
In one embodiment, the drying in step e is preceded by concentration-filtration.
In a specific embodiment, the reselection in the step a comprises rough concentration and fine concentration, preferably 1 to 3 times of rough concentration and 1 to 3 times of fine concentration;
and f, the reselection comprises rough concentration and fine concentration, preferably 1-3 times of rough concentration and 1-3 times of fine concentration.
In one embodiment, the C8-20 fatty acids and salts thereof in the collector include at least one of caprylic acid, pelargonic acid, capric acid, undecanoic acid, dodecanoic acid, tridecanoic acid, tetradecanoic acid, pentadecanoic acid, hexadecanoic acid, heptadecanoic acid, octadecanoic acid, nonadecanoic acid, eicosanoic acid, oleic acid, linoleic acid, linolenic acid, arachidonic acid;
the aviation kerosene comprises 1-10 parts of aviation kerosene wide-fraction type; the aviation kerosene preferably comprises 1-10 parts of kerosene; 1-10 parts of heavy fraction type;
the monoglyceride comprises at least one of glyceryl oleate, glyceryl stearate, glyceryl laurate and glyceryl palmitate; preferably glycerol laurate;
the quaternary ammonium salt comprises dodecyl-hexadecyl trimethyl ammonium chloride or ammonium bromide; preferably dodecyl, tetradecyl or hexadecyltrimethyl ammonium chloride or bromide; more preferably dodecyl trimethyl ammonium chloride or ammonium bromide;
the alkali is at least one of sodium hydroxide, potassium hydroxide, sodium carbonate, potassium carbonate, sodium bicarbonate and potassium bicarbonate; the salt is at least one of sodium salt, potassium salt, ammonium salt, calcium salt and magnesium salt.
The kerosene type of the aviation kerosene is also called aviation kerosene middle-fraction type, the boiling point of which is 150-280 ℃, the boiling point of which of the heavy fraction type is 190-315 ℃, and the boiling point of which of the wide fraction type is 60-280 ℃.
In a specific embodiment, the collector for flotation and desulfurization of the tailings from the sulfuric acid lithium extraction of spodumene is prepared by the following method:
a. mixing the alkali and the silica sol according to a mass ratio, stirring for 0.5-24 h at 50-80 ℃, and reacting to obtain a reagent A;
mixing all the components except the alkali and the silica sol together according to the mass ratio, stirring for 1-2 h at the temperature of 80-100 ℃, and reacting to obtain a reagent B;
b. and uniformly mixing the reagent A and the reagent B to obtain the pasty collector for flotation and desulfurization of the spodumene tailing extracted by the sulfuric acid process.
In a specific embodiment, the method further comprises the steps of subjecting the high-iron lithium-rich material to an alkaline or acid roasting-leaching-solid-liquid separation-purifying and concentrating filtrate after the solid-liquid separation to obtain a lithium salt product.
The high-iron lithium-rich material can adopt a lithium extraction process similar to the lithium extraction of the existing spodumene, and the high-iron lithium-rich material is directly roasted by an alkaline method or an acid method without carrying out transformation roasting unlike the spodumene. The acid adopted in the acid method roasting is concentrated sulfuric acid, and the concentrated sulfuric acid with the concentration of more than 80 percent is selected in principle, and the concentrated sulfuric acid with the concentration of 98 percent is preferred.
In a specific embodiment, the method for purifying the filtrate comprises the steps of adding an oxidizing agent into the filtrate, adding alkali into the oxidized filtrate to adjust the pH value to be more than 1, and performing solid-liquid separation to obtain the purified filtrate; preferably, the pH is 1 to 5, and more preferably 1 to 4.7.
The alkali can be basic carbonate, hydroxide, lime, ammonia water, etc., and relates to sodium carbonate, potassium carbonate, sodium bicarbonate, potassium bicarbonate, calcium carbonate, calcium hydroxide, sodium hydroxide, potassium hydroxide, lime, ammonia water, etc.
In one embodiment, the oxidizing agent is hydrogen peroxide, hypochlorite (e.g., sodium salt, calcium salt), chlorate (e.g., sodium salt, potassium salt), perchlorate (e.g., sodium salt, potassium salt), persulfate (e.g., sodium salt, potassium salt, ammonium salt), oxygen, ozone, or the like, preferably hydrogen peroxide, hypochlorite (e.g., sodium salt, calcium salt), chlorate (e.g., sodium salt, potassium salt), perchlorate (e.g., sodium salt, potassium salt), persulfate (e.g., sodium salt, potassium salt, ammonium salt); more preferably hydrogen peroxide, calcium hypochlorite and sodium chlorate; the time for oxidation is 0.5-24 h, preferably 0.5-10 h, and more preferably 1-5 h.
Has the advantages that:
1. the method thoroughly realizes the purposes of diversified and high-value utilization of the lithium slag deep-processed products, and solves the great problem that slag in the lithium salt industry is difficult to treat;
2. the silicon-aluminum micro powder with high silicon, high aluminum, low iron and low sulfur can be obtained, and can be used in the industries of glass fiber, ceramics, paper making and the like to replace raw materials such as pyrophyllite, kaolin, talc and the like, so that the production cost of the industries of glass fiber, ceramics and paper making is greatly reduced;
3. according to the invention, high-quality gypsum concentrate is obtained through flotation, the purity of the gypsum is up to more than 95%, the gypsum belongs to high-purity gypsum, the gypsum can be used as putty powder, and can be used for developing whisker gypsum materials, coatings, mould materials and the like, so that the value of the gypsum is improved;
4. the invention fully utilizes the resource characteristics, obtains iron ore concentrate through weak magnetic separation, and further improves the value of comprehensive utilization of the lithium slag;
5. the invention obtains the tantalum-niobium concentrate, and the annual output of the lithium slag is 300 ten thousand tons calculated by the content of tantalum-niobium oxide, so that the total amount of the tantalum-niobium oxide is close to 450 tons, and the market value is about 4.5 hundred million RMB.
6. The invention obtains the high-iron lithium-rich slag, and lithium oxide (Li) in the high-iron lithium-rich slag2O) content is 1.0-1.5%, the yield of the high-iron lithium-rich slag is about 5-10%, and calculated by 7% yield, 21 ten thousand tons of high-iron lithium-rich slag and about 2000 tons of lithium metal are generated in one year, the potential market value is about 20-30 hundred million RMB, and the high-iron lithium-rich slag can be used as lithium ore for further recycling lithium carbonate.
7. The invention flotates SO in the residual tailings3Low content of SO3The SO can be obtained by flotation of the lithium slag with the content of more than 10 percent3Tailings with content less than 0.1%.
Drawings
FIG. 1 is a process flow diagram of one embodiment of the present invention.
FIG. 2 is a process flow chart of the lithium extraction by recovering the high-iron lithium-rich material.
FIG. 3 is a comprehensive recovery process diagram of acid roasting of a high-iron lithium-rich material of the invention.
Detailed Description
In order to solve the technical problems, the method for comprehensively recovering lithium, tantalum-niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum from lithium slag comprises the following steps:
a. reselecting the lithium slag to obtain concentrate 1 and tailings 1, and carrying out weak magnetic separation on the concentrate 1 to obtain coarse-grain tantalum-niobium rich materials and coarse-grain iron concentrate;
b. carrying out flotation on the tailings 1 to obtain gypsum and tailings 2;
c. crushing the tailings 2;
d. c, carrying out weak magnetic separation on the tailings crushed in the step c to obtain fine iron ore concentrate and tailings 3;
e. strongly magnetic separating the tailings 3 to obtain concentrate 2 and tailings 4, and drying the tailings 4 to obtain silicon-aluminum micro powder;
f. reselecting the concentrate 2 in the step e to obtain fine-grained tantalum-niobium concentrate and a high-iron lithium-rich material;
wherein the lithium slag is the residue of extracting lithium from the glimmer; step a, the reselection is one or more of table reselection, spiral reselection, centrifugal reselection, hydrocyclone reselection, jigger reselection, wind power reselection and dense medium reselection;
b, the collector for flotation in the step comprises the following components in percentage by weight:
50-100 parts of at least one of C8-20 fatty acid and salt thereof; 1-30 parts of aviation kerosene; 1-30 parts of at least one of dodecyl sulfonic acid or sulfuric acid and salts thereof; 1-30 parts of at least one of polyether or polyalcohol; 1-10 parts of a propylene oxide block copolymer; 1-10 parts of sorbitol monooleate; 1-10 parts of monoglyceride; 1-30 parts of quaternary ammonium salt; 1-10 parts of hexadecyl halogenated pyridine; 5-50 parts of alkali; 10-50 parts of silica sol; 10-100 parts of water;
the polyether or the polyalcohol is at least one of polyvinyl ether, polyoxypropylene ether, polyvinyl alcohol and polyoxyethylene ether, and preferably 1-10 parts of polyvinyl ether, 1-10 parts of polyoxypropylene ether and 1-10 parts of polyvinyl alcohol;
the propylene oxide block copolymer is at least one of PE6100, PE6200, PE6400 and PE 8100;
the dodecyl sulfonic acid or sulfuric acid comprises dodecyl benzene sulfonic acid, dodecyl sulfonic acid and dodecyl sulfuric acid; preferably dodecyl benzene sulfonic acid and its salts; more preferably 1-10 parts of dodecyl benzene sulfonic acid and salts thereof;
the mass concentration of the silica sol is preferably 5-40%;
the concentration of the ore pulp subjected to flotation is preferably 20-60%.
The components of the collecting agent are proportioned in any proportion in a corresponding range, and the aim of efficient flotation and desulfurization of lithium slag can be fulfilled according to the corresponding proportion, so that the content of gypsum is more than 95%, and the impurity SiO in the gypsum is ensured2Content < 1%, Al2O3The content is less than 1 percent, and provides high-quality raw materials for the subsequent preparation of the whisker gypsum.
The desulfurization collecting agent is prepared by effectively adjusting the content of each component, so that sulfur in the lithium slag can be quickly and efficiently removed; by the process, the high-quality gypsum can be easily obtained.
It is worth mentioning that the gypsum obtained by the flotation can be directly used as raw materials for producing gypsum whiskers, gypsum putty powder or fillers after being filtered. The filtered water produced by flotation is collected and then continuously returned to the flotation operation, and the invention does not produce wastewater discharge; as the gypsum in the product needs to take away part of water, the production process of the invention needs to be supplemented with new water finally to ensure the normal operation of the production.
In one embodiment, the grade of tantalum in the lithium slag is Ta2O5The grade of niobium is measured as Nb2O5Calculated by less than 100ppm, preferably 50 to 100 ppm.
The grade of the oxides of tantalum and niobium in the lithium slag is lower than 100ppm, and the oxides can be recovered, and the recovery rate can be ensured to be more than 45%; the existing method has low tantalum-niobium recovery rate or abandons the recovery due to low grade; the method of the invention can recover tantalum and niobium with grade higher than 100ppm and has higher recovery rate.
In one embodiment, the magnetic field intensity of the weak magnetic separation is 100-2000 gauss, preferably 300-1000 gauss; the magnetic field intensity of the strong magnetic separation is 10000-20000 gausses, and 12000-17000 gausses are preferable.
In a specific embodiment, the flotation in the step b is further added with a modifier, wherein the modifier is:
the modifier is at least one of aluminum sol, sodium pyrophosphate, polyepoxysuccinic acid or salt thereof, polyaspartic acid or salt thereof, carboxylic acid-sulfonate copolymer TH-2000, carboxylic acid-sulfonic acid-nonionic terpolymer TH-3100, phosphonocarboxylic acid copolymer POCA, polyacrylic acid or salt thereof, maleic acid-acrylic acid copolymer sodium salt, tannin, chitosan and sodium carboxymethylcellulose, preferably aluminum sol, sodium pyrophosphate, polyacrylic acid or salt thereof, carboxylic acid-sulfonate copolymer TH-2000 and tannin, and the dosage of the modifier is preferably 0-6000 g/t of lithium slag, and more preferably 500-3000 g/t of lithium slag.
In a specific embodiment, the amount of the collector used in the step b is 50-3000 g/t, and preferably 100-1000 g/t of lithium slag.
In a specific embodiment, the flotation comprises roughing, scavenging and concentrating, preferably the roughing is performed for 1-3 times, the scavenging is performed for 1-4 times, and the concentrating is performed for 1-3 times; the using amount of the collecting agent in the scavenging process is 1/20-12/13 of that of the roughing collecting agent, and the collecting agent is not added in the concentration process; preferably, the dosage of the scavenger first collecting agent is 1/2 of the roughing dosage, the dosage of the scavenger second collecting agent is 1/3 of the roughing dosage, and the dosage of the scavenger third collecting agent is 1/4 of the roughing dosage.
In one embodiment, the mesh number of the crushed tailings 2 in the step c is 325 meshes or less, and the crushing is preferably to classify the tailings 2 to obtain particles with the mesh number of 325 meshes or more and 325 meshes or less, crush the particles with the mesh number of 325 meshes or more and mix the crushed particles with the mesh number of 325 meshes or less; the pulverization is preferably carried out by adopting a non-ferrous medium mill.
The grading equipment can adopt a spiral classifier, a swirler, a linear sieve and the like.
In one embodiment, the drying in step e is preceded by concentration-filtration.
In a specific embodiment, the reselection in the step a comprises rough concentration and fine concentration, preferably 1 to 3 times of rough concentration and 1 to 3 times of fine concentration;
and f, the reselection comprises rough concentration and fine concentration, preferably 1-3 times of rough concentration and 1-3 times of fine concentration.
In one embodiment, the C8-20 fatty acids and salts thereof in the collector include at least one of caprylic acid, pelargonic acid, capric acid, undecanoic acid, dodecanoic acid, tridecanoic acid, tetradecanoic acid, pentadecanoic acid, hexadecanoic acid, heptadecanoic acid, octadecanoic acid, nonadecanoic acid, eicosanoic acid, oleic acid, linoleic acid, linolenic acid, arachidonic acid;
the aviation kerosene comprises 1-10 parts of aviation kerosene wide-fraction type; the aviation kerosene preferably comprises 1-10 parts of kerosene; 1-10 parts of heavy fraction type;
the monoglyceride comprises at least one of glyceryl oleate, glyceryl stearate, glyceryl laurate and glyceryl palmitate; preferably glycerol laurate;
the quaternary ammonium salt comprises dodecyl-hexadecyl trimethyl ammonium chloride or ammonium bromide; preferably dodecyl, tetradecyl or hexadecyltrimethyl ammonium chloride or bromide; more preferably dodecyl trimethyl ammonium chloride or ammonium bromide;
the alkali is at least one of sodium hydroxide, potassium hydroxide, sodium carbonate, potassium carbonate, sodium bicarbonate and potassium bicarbonate; the salt is at least one of sodium salt, potassium salt, ammonium salt, calcium salt and magnesium salt.
The kerosene type of the aviation kerosene is also called aviation kerosene middle-fraction type, the boiling point of which is 150-280 ℃, the boiling point of which of the heavy fraction type is 190-315 ℃, and the boiling point of which of the wide fraction type is 60-280 ℃.
In a specific embodiment, the collector for flotation and desulfurization of the tailings from the sulfuric acid lithium extraction of spodumene is prepared by the following method:
a. mixing the alkali and the silica sol according to a mass ratio, stirring at 50-80 ℃ for 0.5-24 h, and reacting to obtain a reagent A;
mixing all the components except the alkali and the silica sol together according to the mass ratio, stirring for 1-2 h at the temperature of 80-100 ℃, and reacting to obtain a reagent B;
b. and uniformly mixing the reagent A and the reagent B to obtain the pasty flotation desulfurization collecting agent for the spodumene sulfuric acid method lithium extraction tailings.
In one embodiment, 50 to 100 parts of at least one of C8-20 fatty acid and salt thereof; 1-15 parts of aviation kerosene; 1-15 parts of at least one of dodecyl sulfonic acid or sulfuric acid and salts thereof; 1-15 parts of at least one of polyether or polyalcohol; 1-5 parts of a propylene oxide block copolymer; 1-5 parts of sorbitol monooleate; 1-5 parts of monoglyceride; 1-20 parts of quaternary ammonium salt; 1-5 parts of hexadecyl halogenated pyridine; 10-50 parts of alkali; 10-50 parts of silica sol; 10-50 parts of water;
preferably 1-5 parts of aviation kerosene wide-cut type; 1-5 parts of kerosene type; 1-5 parts of heavy fraction type;
preferably 1-5 parts of dodecyl benzene sulfonic acid and salts thereof; 1-5 parts of dodecyl sulfuric acid and salts thereof; 1-5 parts of dodecyl sulfonic acid and salts thereof;
preferably 1-5 parts of polyvinyl ether; 1-5 parts of polyoxypropylene ether; 1-5 parts of polyvinyl alcohol;
preferably 1-5 parts of dodecyl trimethyl ammonium chloride or dodecyl trimethyl ammonium bromide; tetradecyltrimethylammonium chloride or bromide 1-10 parts, and hexadecyltrimethylammonium chloride or bromide 1-5 parts.
In one embodiment, 50 to 100 parts of at least one of C8-20 fatty acid and salt thereof; 1-6 parts of aviation kerosene; 1-3 parts of at least one of dodecyl sulfonic acid or sulfuric acid and salts thereof; 1-6 parts of at least one of polyether or polyalcohol; 1-2 parts of a propylene oxide block copolymer; 1 part of sorbitol monooleate; 1 part of monoglyceride; 1-3 parts of quaternary ammonium salt; 1 part of hexadecyl halogenated pyridine; 15-20 parts of alkali; 10 parts of silica sol; 40-50 parts of water;
preferably 1-2 parts of aviation kerosene wide-cut type; 1-2 parts of kerosene; 1-2 parts of heavy fraction type;
preferably 1 part of dodecyl benzene sulfonic acid and salts thereof; 1 part of dodecyl sulfuric acid and salts thereof; 1 part of dodecyl sulfonic acid and salts thereof;
preferably 1-2 parts of polyvinyl ether; 1-2 parts of polyoxypropylene ether; 1-2 parts of polyvinyl alcohol;
preferably 1 part of dodecyl trimethyl ammonium chloride or dodecyl trimethyl ammonium bromide; tetradecyltrimethylammonium chloride or bromide 1 part, hexadecyltrimethylammonium chloride or bromide 1 part.
In a specific embodiment, the method further comprises the steps of subjecting the high-iron lithium-rich material to an alkaline or acid roasting-leaching-solid-liquid separation-purifying and concentrating filtrate after the solid-liquid separation to obtain a lithium salt product.
As shown in figure 2, the high-iron lithium-rich material can adopt a lithium extraction process similar to the lithium extraction of the existing spodumene, and the high-iron lithium-rich material is directly roasted by an alkaline method or an acid method without transformation roasting unlike the spodumene. The acid adopted in the acid method roasting is concentrated sulfuric acid, and the concentrated sulfuric acid with the concentration of more than 80 percent is selected in principle, and the concentrated sulfuric acid with the concentration of 98 percent is preferred. The solid obtained after solid-liquid separation is the silicon-aluminum micro powder as shown in figure 3, and can be recycled.
In a specific embodiment, the method for purifying the filtrate comprises the steps of adding an oxidant into the filtrate, adding alkali into the oxidized filtrate to adjust the pH value to be more than 1, and carrying out solid-liquid separation to obtain the purified filtrate; preferably, the pH is 1 to 5, and more preferably 1 to 4.7.
As shown in FIG. 3, the solid obtained by adding alkali to adjust the pH value and then performing solid-liquid separation can be recycled and used as a cement retarder.
The alkali can be basic carbonate, hydroxide, lime, ammonia water, etc., and relates to sodium carbonate, potassium carbonate, sodium bicarbonate, potassium bicarbonate, calcium carbonate, calcium hydroxide, sodium hydroxide, potassium hydroxide, lime, ammonia water, etc.
In one embodiment, the oxidizing agent is hydrogen peroxide, hypochlorite (e.g., sodium salt, calcium salt), chlorate (e.g., sodium salt, potassium salt), perchlorate (e.g., sodium salt, potassium salt), persulfate (e.g., sodium salt, potassium salt, ammonium salt), oxygen, ozone, or the like, preferably hydrogen peroxide, hypochlorite (e.g., sodium salt, calcium salt), chlorate (e.g., sodium salt, potassium salt), perchlorate (e.g., sodium salt, potassium salt), persulfate (e.g., sodium salt, potassium salt, ammonium salt); more preferably hydrogen peroxide, calcium hypochlorite and sodium chlorate; the time for oxidation is 0.5-24 h, preferably 0.5-10 h, and more preferably 1-5 h.
The following examples are provided to further illustrate the embodiments of the present invention and are not intended to limit the scope of the present invention.
Example 1
As shown in figure 1, spodumene extracted lithium slag from Sichuan company is mainly prepared from quartz, calcite, gypsum, gibbsite, andalusite, corundum, glass phase, alpha-spodumene, beta-spodumene, zeolite, orthoclase, tantalite (trace), and niobite (trace). Raw material Ta2O5Grade of 90ppm, Nb2O5Grade 50ppm, SO3The grade was 6.2%.
(I) gravity separation-flux weakening separation
The raw material is directly reselected by adopting a spiral chute, the concentrate reselected by the spiral chute enters a table concentrator for concentration, the table concentrator concentrate is directly subjected to weak magnetic separation, the magnetic field intensity is 1000 gauss, the magnetic separation tailings are coarse-grained tantalum-niobium concentrate, Ta2O5Grade 18.56% and Nb2O5The grade is 9.56 percent, and the recovery rates of tantalum and niobium are 46.12 percent and 32.68 percent respectively; the magnetic separation concentrate is coarse fraction iron concentrate, the TFe is 52.13 percent, and the recovery rate is 12.89 percent.
(II) desulfurization by flotation
Preparing a collecting agent: firstly, mixing 20 parts of sodium hydroxide and 50 parts of silica sol with the mass fraction of 40%, heating to 80 ℃, and stirring for 5 hours to obtain a paste A;
secondly, mixing C8-20100 parts of fatty acid/fatty acid salt of (2) (C of this example)8-20The fatty acid/fatty acid salt of (a) is caprylic acid and lauric acid according to a 1: 1 part of mixed) aviation kerosene wide fraction type 1 part, aviation kerosene middle fraction type 1 part, aviation kerosene heavy fraction type 1 part, sodium dodecyl benzene sulfonate 1 part, sodium dodecyl sulfate 1 part, polyvinyl ether 1 part, polyoxypropylene ether 1 part, polyvinyl alcohol 1 part, ethylene oxide-propylene oxide block copolymer EO-PO-EO (PE 6100)1 part, sorbitol monooleate 1 part, monoglycerol oleate (the monoglycerol oleate in the experiment is glyceryl oleate, glyceryl stearate, methyl glyceryl stearate, ethyl glyceryl oleate, ethyl glyceryl stearate, ethyl glyceryl oleate, ethyl glyceryl stearate, ethyl glyceryl oleate, ethyl glyceryl stearate and ethyl glyceryl oleate, ethyl glyceryl stearate, ethyl glyceryl oleate, ethyl glyceryl stearate, ethyl glyceryl oleate, ethyl glyceryl stearate and ethyl glyceryl stearate, ethyl glyceryl oleate, ethyl glyceryl stearate and ethyl glyceryl oleate, ethyl glyceryl stearate and ethyl glyceryl stearate, ethyl glyceryl oleate, ethyl glyceryl stearate, ethyl,The components of the lauric glyceride and the palmitic glyceride comprise the following components in percentage by weight: 1 part of mixture) 1 part of mixture, 1 part of dodecyl ammonium chloride, 1 part of hexadecyl trimethyl ammonium chloride, 1 part of hexadecyl pyridine chloride and 50 parts of water are completely and uniformly mixed, heated to 80 ℃ and stirred for 2 hours to obtain a paste B.
And finally, uniformly mixing the paste A and the paste B to obtain the lithium slag desulfurization collecting agent C.
Directly carrying out flotation on gravity tailings, adjusting the concentration of ore pulp to be 35%, and sequentially adding 2000g of regulator and 300g of collecting agent according to ton ore feeding for primary rough separation; the using amount of the conditioning agent for the second roughing is 500g, the using amount of the collecting agent is 100g, the concentrate for the roughing 1 and the concentrate for the roughing 2 are mixed to obtain a roughed concentrate, and the conditioning agent in the flotation roughing is alumina sol. The roughing concentrate is selected once, and the selected middlings are returned to the first roughing; carrying out three times of scavenging operation on the rough concentration tailings, wherein the using amount of a scavenging regulator for the first time is 500g, the using amount of a collecting agent is 50g, the using amount of a scavenging regulator for the second time is 250g, the using amount of the collecting agent is 30g, the using amount of a scavenging regulator for the third time is 250g, and the using amount of the collecting agent is 20g, returning scavenging middlings to the first rough concentration operation, finally forming closed circulation to obtain desulfurized gypsum concentrate and desulfurized tailings, and carrying out CaSO (calcium ammonium sulfate) in the gypsum concentrate4.2H2The purity of O is more than 95 percent.
(III) grading-ore grinding-weak magnetic separation
Classifying the flotation tailings into a sample of-325 meshes and a sample of +325 meshes; wherein, a sample with a size of +325 meshes directly enters a ceramic mill, and is ground until the size of-325 meshes accounts for 100%; and mixing the tailings graded-325-mesh sample with the grinded-325-mesh sample, and directly performing weak magnetic separation with the magnetic field intensity of 2000 Gauss to obtain fine-grained iron ore concentrate, wherein the TFe is 42.23%, and the yield is 8.2%.
(IV) Strong magnetic separation-reselection
Directly feeding the ore pulp subjected to weak magnetic separation into a strong magnetic separator, performing strong magnetic separation and roughing on the ore pulp with the magnetic field intensity of 1.2T and scavenging on the ore pulp with the magnetic field intensity of 1.7T, mixing roughing and scavenging, and performing strong magnetic separation to obtain magnetic separation tailings and tantalum-niobium-lithium-rich magnetic products. Directly concentrating, filtering and drying the magnetic separation tailings to obtain silicon-aluminum micro powder 1 with the yield of 75 percent and SO3Grade 0.15%, Fe2O3The grade is 0.32%. Directly carrying out reselection on the lithium-rich tantalum-niobium magnetic product in a shaking table, directly carrying out concentration on concentrate in first reselection, returning middlings in the first reselection for reselection, and directly using tailings in the first reselection as tailings; and performing secondary concentration on the concentrate obtained by the primary gravity separation, wherein the concentrate obtained by the secondary gravity separation is the final fine-grained tantalum-niobium concentrate, and the middling and tailings obtained by the secondary gravity separation are directly returned to the primary gravity separation. Respectively obtaining fine-grained tantalum-niobium concentrate and high-iron lithium-rich material after gravity separation, and fine-grained tantalum-niobium concentrate Ta2O5Grade 10.52% and Nb2O5The grade is 4.78 percent, the recovery rates of tantalum and niobium are 14.73 percent and 18.79 percent respectively, and the high-iron lithium-rich material Li2The O grade is 1.58 percent, and the lithium recovery rate is 25 percent.
(V) extracting lithium from high-iron lithium-rich material by sulfuric acid process
As shown in fig. 2 and 3, 1000g of high-iron lithium-rich material is mixed with 50g of 98% concentrated sulfuric acid, and the mixture is placed in a muffle furnace and roasted for 2 hours at a constant temperature of 300 ℃; cooling the roasted material, and then mixing the cooled roasted material with water according to the solid-liquid mass ratio of 1: 1, mixing and leaching, stirring and leaching for 2 hours, wherein the leaching temperature is 40 ℃; and (3) performing solid-liquid separation to obtain filtrate and silicon-aluminum micro powder 2, adding 5ml of hydrogen peroxide into the filtrate, performing oxidation reaction for 0.5h, adding calcium carbonate to adjust the pH value to 3, and filtering to obtain purified lithium liquid and calcium-iron slag (cement retarder). The yield of silicon-aluminum micro powder in the whole process of extracting lithium by a sulfuric acid method of the high-iron lithium-rich material is 92 percent, and then lithium carbonate product, namely Li, is obtained after the processes of concentration, sodium removal, lithium precipitation and the like2The recovery rate of the O whole flow is 20.5 percent (the operation recovery rate is 82 percent).
Example 2
The others are similar to example 1, except that 5 parts of aviation kerosene wide cut type; 3 parts of aviation kerosene middle fraction type and 3 parts of aviation kerosene heavy fraction type.
Example 3
The other is similar to example 1, the only difference being that no modifier is added.
Examples 4 to 5
The other example is similar to example 1, except that the conditioning agent of example 4 is chitosan. The conditioning agent of example 4 was sodium carboxymethylcellulose.
Examples 6 to 7
Otherwise similar to example 1, example 6 only differs in that the field strength for the field weakening separation is 1500 gauss; example 7 the only difference is that the field strength for the strong magnetic separation is 11000 gauss;
examples 8 to 9
Otherwise similar to example 1, the only difference in example 8 is that the first scavenging collector dose was 500g/t, the second scavenging collector dose was 250g/t and the third scavenging collector dose was 100 g/t; the only difference in example 9 is that the first scavenging modifier is 800g/t, the second scavenging modifier is 500g/t and the third scavenging modifier is 200 g/t.
TABLE 1 EXAMPLES 1-9 coarse tantalate-niobium concentrate grades and yields
| Examples | Ta2O5Grade% | Nb2O5Grade% | Ta2O5Yield% | Nb2O5Yield% |
| 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 |
Table 2 examples 1-9 fine-grained tantalum niobium concentrate grades and yields
| Examples | Ta2O5Grade% | Nb2O5Grade% | Ta2O5Yield% | Nb2O5Yield% |
| 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 |
Table 3 examples 1-9 iron concentrate grades and yields
TABLE 4 EXAMPLES 1-9 SILICON-ALUMINUM FINE POWDER, LITHIUM CARBONATE GRADE AND YIELD
TABLE 5 examples 1-9 Gypsum grade and yield
Claims (41)
1. The method for comprehensively recovering lithium, tantalum, niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum from lithium slag is characterized by comprising the following steps:
a. reselecting the lithium slag to obtain concentrate 1 and tailings 1, and performing weak magnetic separation on the concentrate 1 to obtain coarse-grain tantalum-niobium rich materials and coarse-grain iron concentrate;
b. carrying out flotation on the tailings 1 to obtain gypsum and tailings 2;
c. crushing the tailings 2;
d. c, carrying out weak magnetic separation on the tailings crushed in the step c to obtain fine iron ore concentrate and tailings 3;
e. strongly magnetic separating the tailings 3 to obtain concentrate 2 and tailings 4, and drying the tailings 4 to obtain silicon-aluminum micro powder;
f. reselecting the concentrate 2 in the step e to obtain fine-grained tantalum-niobium concentrate and a high-iron lithium-rich material;
wherein the lithium slag is spodumene lithium extraction slag; step a, the reselection is one or more of table reselection, spiral reselection, centrifugal reselection, hydrocyclone reselection, jigger reselection, wind power reselection and dense medium reselection;
b, the collector for flotation in the step comprises the following components in percentage by weight:
50-100 parts of at least one of C8-20 fatty acid and salt thereof; 1-30 parts of aviation kerosene; 1-30 parts of at least one of dodecyl sulfonic acid or sulfuric acid and salts thereof; 1-30 parts of at least one of polyether or polyalcohol; 1-10 parts of a propylene oxide block copolymer; 1-10 parts of sorbitol monooleate; 1-10 parts of monoglyceride; 1-30 parts of quaternary ammonium salt; 1-10 parts of hexadecyl halogenated pyridine; 5-50 parts of alkali; 10-50 parts of silica sol; 10-100 parts of water;
the polyether or the polyalcohol is at least one of polyvinyl ether, polyoxypropylene ether, polyvinyl alcohol and polyoxyethylene ether;
the propylene oxide block copolymer is at least one of PE6100, PE6200, PE6400 and PE 8100;
the dodecyl sulfonic acid or sulfuric acid comprises dodecyl benzene sulfonic acid, dodecyl sulfonic acid and dodecyl sulfuric acid.
2. The method for comprehensively recovering lithium, tantalum, niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum by using the lithium slag according to claim 1, wherein the polyether or the polyalcohol is 1-10 parts of polyvinyl ether, 1-10 parts of polyoxypropylene ether and 1-10 parts of polyvinyl alcohol.
3. The method for comprehensively recovering lithium, tantalum and niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum by using the lithium slag according to claim 1 or 2, wherein the dodecyl sulfonic acid or sulfuric acid comprises dodecyl benzene sulfonic acid and salts thereof.
4. The method for comprehensively recovering lithium, tantalum, niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum by using the lithium slag according to claim 3, wherein the dodecyl benzene sulfonic acid and the salt thereof account for 1-10 parts.
5. The method for comprehensively recovering lithium, tantalum and niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum from lithium slag according to claim 1 or 2, wherein the mass concentration of the silica sol is 5-40%.
6. The method for comprehensively recovering lithium, tantalum and niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum from lithium slag according to claim 1 or 2, wherein the concentration of the ore pulp subjected to flotation is 20-60%.
7. The method for comprehensively recovering lithium, tantalum and niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum from lithium slag according to claim 1 or 2, wherein the grade of tantalum in the lithium slag is Ta2O5The grade of niobium is measured as Nb2Less than 100ppm calculated as O5.
8. The method for comprehensively recovering lithium, tantalum-niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum from lithium slag according to claim 7, wherein the grade of tantalum in the lithium slag is Ta2O5The grade of niobium is measured as Nb250 to 100ppm of O5.
9. The method for comprehensively recovering lithium, tantalum, niobium, silicon-aluminum micropowder, iron ore concentrate and gypsum from lithium slag according to claim 1 or 2, wherein the magnetic field intensity of weak magnetic separation is 100-2000 gauss; the magnetic field intensity of the strong magnetic separation is 10000-20000 gausses.
10. The method for comprehensively recovering lithium, tantalum, niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum by using the lithium slag according to claim 9, wherein the magnetic field intensity of weak magnetic separation is 300-1000 gauss.
11. The method for comprehensively recovering lithium, tantalum, niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum by using the lithium slag according to claim 9, wherein the magnetic field intensity of the strong magnetic separation is 12000-17000 gauss.
12. The method for comprehensively recovering lithium, tantalum and niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum from lithium slag according to claim 1 or 2, characterized in that a regulator is further added in the flotation in the step b, and the regulator is as follows:
at least one of aluminum sol, sodium pyrophosphate, polyepoxysuccinic acid or salt thereof, polyaspartic acid or salt thereof, carboxylic acid-sulfonate copolymer TH-2000, carboxylic acid-sulfonic acid-nonionic terpolymer TH-3100, phosphonocarboxylic acid copolymer POCA, polyacrylic acid or salt thereof, maleic acid-acrylic acid copolymer sodium salt, tannin, chitosan and sodium carboxymethylcellulose.
13. The method for comprehensively recovering lithium, tantalum, niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum by using the lithium slag according to claim 12, wherein the regulator is: aluminum sol, sodium pyrophosphate, polyacrylic acid or salt thereof, carboxylic acid-sulfonate copolymer TH-2000 and tannin.
14. The method for comprehensively recovering lithium, tantalum, niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum by using the lithium slag according to claim 12, wherein the using amount of the regulator is 0-6000 g/t of the lithium slag.
15. The method for comprehensively recovering lithium, tantalum, niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum by using the lithium slag according to claim 14, wherein the using amount of the regulator is 500-3000 g/t of the lithium slag.
16. The method for comprehensively recovering lithium, tantalum, niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum from lithium slag according to claim 1 or 2, wherein the using amount of the collecting agent in the step b is 50-3000 g/t.
17. The method for comprehensively recovering lithium, tantalum, niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum by using the lithium slag according to claim 16, wherein the using amount of the collecting agent in the step b is 100-1000 g/t of the lithium slag.
18. The method for comprehensively recovering lithium, tantalum and niobium, silicon-aluminum micropowder, iron ore concentrate and gypsum by using the lithium slag according to claim 16, wherein the flotation comprises roughing, scavenging and concentrating; the using amount of the collecting agent in the scavenging process is 1/20-13/12 of that of the rough separation collecting agent, and the collecting agent is not added in the fine separation process.
19. The method for comprehensively recovering lithium, tantalum, niobium, silicon-aluminum micropowder, iron ore concentrate and gypsum by using the lithium slag according to claim 18, is characterized by comprising the steps of roughing 1-3 times, scavenging 1-4 times and concentrating 1-3 times.
20. The method for comprehensively recovering lithium, tantalum and niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum by using the lithium slag according to claim 19, wherein the dosage of the scavenging one-collecting agent is 1/2 of the roughing dosage, the dosage of the scavenging two-collecting agent is 1/3 of the roughing dosage, and the dosage of the scavenging three-collecting agent is 1/4 of the roughing dosage.
21. The method for comprehensively recovering lithium, tantalum and niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum by using the lithium slag according to claim 1 or 2, wherein the granularity of the crushed tailings in the step c is more than 325 meshes.
22. The method for comprehensively recovering lithium, tantalum and niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum from lithium slag according to claim 21, wherein the crushing in the step c is to classify the tailings 2 into particles of 325 meshes or more and 325 meshes or less, and the particles of 325 meshes or more are crushed and then mixed with the particles of 325 meshes or less.
23. The method for comprehensively recovering lithium, tantalum and niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum from lithium slag according to claim 21 is characterized in that a non-ferrous medium mill is adopted for fine grinding.
24. The method for comprehensively recovering lithium, tantalum and niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum by using the lithium slag according to claim 1 or 2, characterized in that concentration and filtration are performed before drying in the step e.
25. The method for comprehensively recovering lithium, tantalum and niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum by using the lithium slag according to claim 1 or 2, wherein the reselection in the step a comprises roughing and concentrating;
and f, the reselection comprises roughing and concentrating.
26. The method for comprehensively recovering lithium, tantalum, niobium, silicon-aluminum micropowder, iron ore concentrate and gypsum by using the lithium slag according to claim 25, wherein the reselection in the step a comprises 1-3 times of roughing and 1-3 times of concentrating.
27. The method for comprehensively recovering lithium, tantalum and niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum from lithium slag according to claim 25, wherein the reselection in the step f comprises 1-3 times of rough concentration and 1-3 times of fine concentration.
28. The method for comprehensively recovering lithium, tantalum and niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum by using the lithium slag according to claim 1 or 2, wherein the C8-20 fatty acid and the salt thereof in the collector comprise at least one of caprylic acid, pelargonic acid, capric acid, undecanoic acid, dodecanoic acid, tridecanoic acid, tetradecanoic acid, pentadecanoic acid, hexadecanoic acid, heptadecanoic acid, octadecanoic acid, nonadecanoic acid, eicosanoic acid, oleic acid, linoleic acid, linolenic acid and arachidonic acid;
the aviation kerosene comprises 1-10 parts of aviation kerosene wide-fraction type; the monoglyceride comprises at least one of glyceryl oleate, glyceryl stearate, glyceryl laurate and glyceryl palmitate;
the quaternary ammonium salt comprises dodecyl-hexadecyl trimethyl ammonium chloride or ammonium bromide;
the alkali is at least one of sodium hydroxide, potassium hydroxide, sodium carbonate, potassium carbonate, sodium bicarbonate and potassium bicarbonate; the salt is at least one of sodium salt, potassium salt, ammonium salt, calcium salt and magnesium salt.
29. The method for comprehensively recovering lithium, tantalum, niobium, silicon-aluminum micropowder, iron ore concentrate and gypsum from lithium slag according to claim 28, wherein the aviation kerosene comprises 1-10 parts of aviation kerosene; 1-10 parts of heavy fraction.
30. The method for comprehensively recovering lithium, tantalum-niobium, silicon-aluminum micropowder, iron ore concentrate and gypsum by using the lithium slag according to claim 28, wherein the monoglyceride comprises glycerol laurate.
31. The method for comprehensively recovering lithium, tantalum and niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum by using the lithium slag according to claim 28, wherein the quaternary ammonium salt is dodecyl, tetradecyl or hexadecyl trimethyl ammonium chloride or ammonium bromide.
32. The method for comprehensively recovering lithium, tantalum, niobium, silicon-aluminum micropowder, iron ore concentrate and gypsum by using the lithium slag according to claim 28, wherein the quaternary ammonium salt is dodecyl trimethyl ammonium chloride or ammonium bromide.
33. The method for comprehensively recovering lithium, tantalum and niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum from lithium slag according to claim 1 or 2, characterized by further comprising the steps of carrying out alkali method or acid method roasting-leaching-solid-liquid separation on the high-iron lithium-rich material, purifying the filtrate after the solid-liquid separation, and concentrating to obtain a lithium salt product.
34. The method for comprehensively recovering lithium, tantalum, niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum from lithium slag according to claim 33, wherein the method for purifying the filtrate comprises the steps of adding an oxidant into the filtrate, adding alkali into the oxidized filtrate to adjust the pH value to be more than 1, and carrying out solid-liquid separation to obtain the purified filtrate.
35. The method for comprehensively recovering lithium, tantalum and niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum from lithium slag according to claim 34 is characterized in that the method for purifying the filtrate comprises the steps of adding an oxidant into the filtrate, and adding alkali into the oxidized filtrate to adjust the pH value to be 1-5.
36. The method for comprehensively recovering lithium, tantalum, niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum from lithium slag according to claim 34, wherein the method for purifying the filtrate comprises the steps of adding an oxidant into the filtrate, and adding alkali into the oxidized filtrate to adjust the pH value to be 1-4.7.
37. The method for comprehensively recovering lithium, tantalum, niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum by using the lithium slag according to claim 34, wherein the oxidant is at least one of hydrogen peroxide, hypochlorite, chlorate, perchlorate, persulfate, oxygen and ozone, and the oxidation time is 0.5-24 hours.
38. The method for comprehensively recovering lithium, tantalum, niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum by using the lithium slag according to claim 37, wherein the oxidant is at least one of hydrogen peroxide, hypochlorite, chlorate, perchlorate and persulfate.
39. The method for comprehensively recovering lithium, tantalum and niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum from lithium slag according to claim 37, wherein the oxidant is at least one of hydrogen peroxide, calcium hypochlorite and sodium chlorate.
40. The method for comprehensively recovering lithium, tantalum, niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum by using the lithium slag according to claim 37, wherein the oxidation time is 0.5-10 h.
41. The method for comprehensively recovering lithium, tantalum, niobium, silicon-aluminum micro powder, iron ore concentrate and gypsum by using the lithium slag according to claim 37, wherein the oxidation time is 1-5 h.
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| CA3236807A CA3236807A1 (en) | 2021-12-01 | 2022-11-21 | Method for comprehensively recycling lithium, tantalum and niobium, silicon aluminum micropowder, iron concentrate, and gypsum from lithium slags |
| PCT/CN2022/133160 WO2023098500A1 (en) | 2021-12-01 | 2022-11-21 | Method for comprehensively recovering lithium, tantalum-niobium, silicon-aluminum micro-powder, iron ore concentrate and gypsum from lithium slag |
| AU2022402780A AU2022402780B2 (en) | 2021-12-01 | 2022-11-21 | Method for comprehensively recovering lithium, tantalum-niobium, silicon-aluminum micro-powder, iron ore concentrate and gypsum from lithium slag |
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| CN114713360B (en) * | 2022-04-14 | 2023-10-10 | 成都德菲环境工程有限公司 | Extraction process of usable substances in pyrite cinder |
| CN116462522B (en) * | 2023-06-19 | 2023-08-22 | 湖南永杉锂业有限公司 | Method for preparing sagger |
| CN117165787A (en) * | 2023-09-19 | 2023-12-05 | 天齐锂业股份有限公司 | Method for extracting lithium from spodumene and simultaneously recovering low-iron low-sulfur silicon-aluminum micropowder, high-purity gypsum, tantalum-niobium concentrate and lithium-enriched iron material |
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| CN103418488B (en) * | 2013-08-23 | 2015-02-25 | 中国地质科学院矿产综合利用研究所 | Comprehensive recovery process of lithium polymetallic ore associated with fine niobium and tantalum |
| CN103789553B (en) * | 2013-11-28 | 2015-07-22 | 中南大学 | Method for comprehensively utilizing lepidolite mineralogical reconstruction lithium-extraction slags |
| CN104209179A (en) * | 2014-09-26 | 2014-12-17 | 湖北鑫鹰环保科技有限公司 | Production method for preferably selecting lepidolite from tantalum and niobium ores |
| CN105195268A (en) * | 2015-09-18 | 2015-12-30 | 江苏融达新材料股份有限公司 | Production method for grinding lithium slag and mineral slag composite powder by means of vertical grinding machine |
| AU2017235956A1 (en) * | 2016-09-29 | 2018-04-12 | Poseidon Nickel Limited | Method of Processing Lithium-Bearing Ores |
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| CN108273826B (en) * | 2018-01-17 | 2018-12-11 | 成都绿锂环保科技有限公司 | A kind of complete mutually high-valued recoverying and utilizing method of lithium slag |
| CA3089904A1 (en) * | 2018-02-02 | 2019-08-08 | Tianqi Lithium Kwinana Pty Ltd | A process for extracting values from lithium slag |
| CN110433958A (en) * | 2019-06-18 | 2019-11-12 | 中国地质科学院矿产综合利用研究所 | Niobium-tantalum-containing lithium polymetallic ore step non-tailing recovery process |
| BR112021026615A2 (en) * | 2019-08-29 | 2022-03-15 | Basf Se | Use of a composition, direct flotation process for the beneficiation of lithium silicate and magnesium silicate, reverse flotation process for the removal of lithium silicate and magnesium silicate, and, composition |
| CN214488258U (en) * | 2021-01-15 | 2021-10-26 | 中冶长天国际工程有限责任公司 | Comprehensive recycling system for lithium slag |
| CN112958273B (en) * | 2021-03-30 | 2023-02-03 | 广东省科学院资源综合利用研究所 | Mineral separation method for pegmatite type lithium polymetallic ore |
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| CN216459396U (en) * | 2021-12-01 | 2022-05-10 | 天齐创锂科技(深圳)有限公司 | System for lithium, tantalum niobium, silicon-aluminum miropowder, iron ore concentrate and gypsum are retrieved to lithium sediment is synthesized |
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