CN113151669A - Method for decomposing low-grade tantalum-niobium resource and extracting tantalum-niobium by alkaline process - Google Patents
Method for decomposing low-grade tantalum-niobium resource and extracting tantalum-niobium by alkaline process Download PDFInfo
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- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B1/00—Preliminary treatment of ores or scrap
- C22B1/02—Roasting processes
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- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D1/00—Flotation
- B03D1/001—Flotation agents
- B03D1/018—Mixtures of inorganic and organic compounds
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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
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D1/00—Flotation
- B03D1/02—Froth-flotation processes
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- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
- C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
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- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B34/00—Obtaining refractory metals
- C22B34/20—Obtaining niobium, tantalum or vanadium
- C22B34/24—Obtaining niobium or tantalum
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- C—CHEMISTRY; METALLURGY
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- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
- C22B7/001—Dry processes
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- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
- C22B7/006—Wet processes
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- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
- C22B7/04—Working-up slag
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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
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D2201/00—Specified effects produced by the flotation agents
- B03D2201/02—Collectors
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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
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D2201/00—Specified effects produced by the flotation agents
- B03D2201/04—Frothers
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- Y02P10/20—Recycling
Abstract
The invention discloses a method for decomposing low-grade tantalum-niobium resources and extracting tantalum-niobium by an alkaline method, which comprises the steps of uniformly mixing ground low-grade tantalum-niobium resources with potassium carbonate and a composite oxidant, roasting, and performing water leaching on a roasted product to obtain a leaching solution containing potassium polytantalate and potassium polyniobate; performing foam flotation on the leachate to obtain a foam product enriched with tantalum and niobium; drying and calcining the foam product in sequence to obtain an oxide containing tantalum and niobium; the method has the advantages of low alkali consumption, low roasting temperature and short time consumption, the recovery rate of the tantalum-niobium resource in the low-grade tantalum-niobium resource is up to 95-98%, the method is environment-friendly, the requirement on equipment is low, and the method is suitable for industrial large-scale production.
Description
Technical Field
The invention relates to a low-grade tantalum-niobium resource processing method, in particular to a method for decomposing a low-grade tantalum-niobium resource and extracting tantalum-niobium by an alkaline method, and belongs to the field of rare and precious metal metallurgy.
Background
Tantalum (Ta) and niobium (Nb) are rare precious metals, located in the fifth and sixth periods of group VB, respectively. Because lanthanide series shrinks, the metal bond of tantalum and niobium is stronger, and the tantalum and niobium has higher melting point and larger atomic enthalpy, so the tantalum and niobium alloy is widely applied to the fields of chemical industry, metallurgy, aerospace, nuclear industry, steel industry and the like. Tantalum and niobium belong to elements which are not widely distributed, tantalum and niobium concentrate resources in China are deficient, and most of tantalum concentrate in China depends on import. The low-grade resources containing tantalum and niobium are mainly divided into two types: low grade minerals and metallurgical slag. The use of the traditional hydrofluoric acid method for the raw materials causes the large use of acid, poor impurity separation effect and low tantalum-niobium recovery rate, so that the development of an efficient, economic and green process for recovering tantalum and niobium from low-grade raw materials is urgently needed. At present, the main methods for decomposing the tantalum-niobium ore comprise an acid decomposition method, an alkali decomposition method and a chlorination method
Acid decomposition method: due to the characteristic of corrosion resistance of tantalum and niobium, under the condition of low temperature, strong acid and strong base except hydrofluoric acid are used for leaching, the leaching rate is lower than 30%, and therefore, the tantalum and niobium ores are difficult to decompose by using cheap industrial inorganic acid. At present, tantalum-niobium ores are decomposed by a hydrofluoric acid method in foreign factories, when tantalum-niobium concentrates are decomposed by hydrofluoric acid, elements such as alkaline earth, rare earth and uranium generate insoluble fluoride precipitates, tantalum, niobium and part of impurities enter a solution and are separated by a subsequent extraction process, the method is high in decomposition rate which can reach more than 98%, the process is simple, but hydrofluoric acid is high in toxicity and volatility, 10% of hydrofluoric acid waste gas is generated in the decomposition process, harm is caused to human health, and a large amount of fluoride, fluorite and fluorine-containing wastewater are generated to cause serious pollution to the ecological environment. The hydrofluoric acid-sulfuric acid method is commonly used in domestic factories, the addition of sulfuric acid not only reduces the consumption of hydrofluoric acid, but also decomposes to generate non-extracted sulfate, which is beneficial to the separation of impurities in the subsequent extraction process.
A chlorination method: the chloridizing metallurgy of tantalum and niobium refers to that chloridizing agents such as chlorine gas and the like decompose tantalum-niobium ore at high temperature to generate tantalum-niobium chlorides, and the main parts in concentrate are separated by utilizing different vapor pressures of the chlorides. For example, the chlorides of tantalum, niobium and titanium with lower boiling point can be taken out of the reaction furnace along with the gas and separated from the chlorides of rare earth, alkaline earth and alkali metals with high boiling point, and the like, and the chlorides are condensed in a condensing device. For example, the boiling chlorination method improved by Chinese patent (CN109182782A) has high decomposition rate and is not limited by mineral components, but has the disadvantages of high corrosion to equipment, complex equipment and complex process operation.
Alkali decomposition method: the alkali decomposition method includes a conventional alkali fusion method and a high-pressure alkali solution decomposition method, and is based on the principle that tantalum niobium ore is decomposed into tantalate (niobate), and then the tantalate (niobate) is reacted with acid to generate tantalum niobium oxide. The temperature of the traditional alkali melting method is above 800 ℃, and the alkali consumption is 6-8 times of the theoretical value, which is the main defect. Chinese patent (CN1605639A) develops a method for cleanly converting tantalum-niobium ores at 300 ℃, but the ratio of alkali ores is still as high as 3-7: 1. the domestic and foreign metallurgy researchers put forward a high-pressure alkali solution decomposition method, such as Chinese patent (CN103572046A), which greatly reduces the alkali consumption and the reaction temperature, but the reaction process is operated under pressure and the reaction conditions are harsh.
Disclosure of Invention
Aiming at the defects of the processing method of the low-grade tantalum-niobium resource in the prior art, the invention aims to provide the method for decomposing the low-grade tantalum-niobium resource by the alkaline method and extracting the tantalum-niobium resource.
In order to realize the technical purpose, the invention provides a method for decomposing low-grade tantalum-niobium resources and extracting tantalum-niobium by an alkaline method, which comprises the following steps:
1) uniformly mixing the ground low-grade tantalum-niobium resource with potassium carbonate and a composite oxidant, and roasting to obtain a roasted product; the composite oxidant comprises saltpeter, potassium hypochlorite and potassium permanganate;
2) treating the roasted product by water leaching to obtain leachate containing potassium polytantalate and potassium polyniobate;
3) performing foam flotation on the leachate to obtain a foam product enriched with tantalum and niobium; the froth flotation adopts a primary amine collecting agent and a ketone collecting agent as combined collecting agents;
4) and (3) drying and calcining the foam product in sequence to obtain the oxide containing tantalum and niobium.
The key of the technical scheme of the invention is as follows: on one hand, special oxidants are adopted in the process of decomposing the low-grade tantalum-niobium resources by alkaline oxidizing roasting, and the addition of the oxidants can promote the low-grade tantalum-niobium resources to be quickly and efficiently decomposed at a relatively low roasting temperature and to be converted into water-soluble potassium polytantalate (niobate) with high selectivity; on the other hand, a special combined collector is adopted in the froth flotation process, the combined collector has good flotation selectivity on multi-tantalum (niobium) acid radicals, the deep enrichment of tantalum and niobium can be realized, and the recovery efficiency of tantalum and niobium is improved.
According to the technical scheme, the low-grade tantalum-niobium resource, the potassium carbonate and the composite oxidant are roasted at high temperature, the tantalum-niobium can be converted into soluble potassium salt, the rare earth metals, the ferro-manganese and the like are mainly converted into insoluble oxides, and the main reaction mechanism is as follows (wherein O is2Provided by a composite oxidant):
(Mn,Fe)(TaO3)2+4K2CO3+O2＝4K2TaO3+2(Mn,Fe)2O3+CO2↑；
(Mn,Fe)(NbO3)2+4K2CO3+O2＝4K2NbO3+2(Mn,Fe)2O3+CO2↑；
2REFCO3+Na2CO3＝RE2O3+2NaF+3CO2↑；
2REPO4+3Na2CO3＝RE2O3+2Na3PO4+3CO2↑；
conversion of aluminum, silicon, tungsten, tin to soluble potassium salts:
SiO2+Na2CO3＝Na2SiO3+CO2↑；
SnO2+Na2CO3＝Na2SnO3+CO2↑；
FeWO4+Na2CO3+1/2O2＝Na2WO4+Fe2O3+CO2↑；
MnWO4+Na2CO3+1/2O2＝Na2WO4+Mn2O3+CO2↑；
zirconium dioxide is an amphoteric oxide which forms a zirconate by co-melting with a base, but the zirconate readily hydrolyzes to ZrO when exposed to water2·xH2O and precipitating:
ZrO2+2NaOH＝Na2ZrO3+H2O；
Na2ZrO3+H2O＝ZrO2+NaOH；
the roasted product obtained by high-temperature roasting is leached by water, and the water leaching slag contains elements such as iron, manganese, titanium, rare earth metal, zirconium and the like and a small amount of insoluble potassium metatantalate (niobium) and still has recovery value due to the fact that the water leaching slag contains a small amount of tantalum and niobium, and can be directly used as smelting slag to return to the roasting process, or return to the roasting after leaching and recovering part of iron, manganese and titanium by dilute acid; the water leaching solution contains elements such as tungsten, tin, silicon, aluminum and the like and potassium tantalate (niobate), in the subsequent flotation separation process, the tantalum-niobium enriched component is separated by adopting a special flotation collector through high-selectivity flotation, and the flotation foam product is dried and calcined to obtain the high-purity tantalum-niobium oxide.
As a preferable scheme, the low-grade tantalum-niobium resource is ground to a particle size of less than 0.079mm, and the mass percentage content of the particle size is more than 90%. After the low-grade tantalum-niobium resources are ground to a proper granularity, the interaction between potassium carbonate and an oxidant in a high-temperature solid-phase reaction and the low-grade tantalum-niobium resources is facilitated, and the high-temperature solid-phase reaction efficiency is improved.
As a preferable scheme, the addition amount of the composite oxidant is 1-6 wt% of the total mass of the potassium carbonate and the low-grade tantalum-niobium resource.
As a preferable scheme, the mass ratio of the low-grade tantalum-niobium resource to the potassium carbonate is 1: 1-3. The mass ratio of the low-grade tantalum-niobium resource to the potassium carbonate is preferably 1: 1.2-1.5, and the dosage of the potassium carbonate can be greatly reduced under the condition that an oxidant is used cooperatively to promote the rapid and efficient decomposition of the low-grade tantalum-niobium resource.
As a preferable scheme, the composite oxidant consists of the following components in parts by weight: 40-60 parts of saltpeter; 8-20 parts of potassium hypochlorite; 20-40 parts of potassium permanganate. The further preferable composite oxidant consists of the following components in parts by weight: 50-60 parts of saltpeter, 10-15 parts of potassium hypochlorite and 20-30 parts of potassium permanganate. The preferable composite oxidant can promote the low-grade tantalum-niobium resource to be highly selectively converted into water-soluble potassium poly-tantalum (niobium) at a lower potassium carbonate ratio and a relatively lower temperature, and simultaneously oxidize rare earth metals, iron, manganese, titanium and the like into metal oxides which cannot be soaked in water, so that the separation of tantalum-niobium from the rare earth metals, iron, manganese, titanium and the like in the low-grade tantalum-niobium resource can be realized. The combination of the saltpeter, the potassium hypochlorite and the potassium permanganate not only can strengthen the mineral decomposition of low-grade tantalum-niobium resources, but also can reduce the dosage of the oxidant and the danger of the oxidant by using a mixed oxidant compared with using a single oxidant.
As a preferred embodiment, the conditions of the roasting treatment are as follows: the temperature is 600-800 ℃, and the time is 30-100 min. The more preferable roasting treatment conditions are as follows: the temperature is 650-750 ℃ and the time is 40-60 min. If the roasting temperature is too low and the roasting time is too short, the tantalum and niobium in the low-grade tantalum and niobium resource are difficult to be efficiently converted into the potassium tantalate (niobate) which is easy to dissolve in water; if the roasting temperature is too high or the roasting time is too long, energy consumption is too high, and the cost is increased.
As a preferable scheme, the water immersion treatment conditions are as follows: the solid-liquid ratio is 1g: 3-6 mL, the temperature is 50-95 ℃, and the time is 0.5-1.5 h. The solid-to-liquid ratio is preferably 1g:5 mL. The preferable temperature is 80-95 ℃ and the time is 1 h.
As a preferred scheme, the combined collector consists of the following components in percentage by mass: 30-70% of primary amine collecting agent; and 30-70% of ketone collecting agent. The preferred combined collector can be used for high-selectivity flotation separation of poly-tantalum (niobate) in a complex system containing tungsten, tin, silicon, aluminum and the like. The primary amine collecting agent and the ketone collecting agent are compounded for use, the formed micelle interface arrangement is more compact than that of a single surfactant, the surface hydrophobicity of the poly-tantalum (niobium) acid radical can be obviously improved, the floatability is improved, the critical micelle concentration and the surface tension of a compound system can be greatly reduced by combining the two collecting agents, the medicament cost can be reduced, and the pressure on the environment can be reduced.
As a preferable scheme, the primary amine collector is at least one of dioctyl amine, dodecyl amine, tetradecylamine and hexadecyl amine.
As a preferred scheme, the ketone collector is at least one of methyl isobutyl ketone, cyclohexanone and diisopropyl ketone.
The amine collecting agent and the ketone collecting agent are common collecting agents in the prior art, and the key point of the invention is that the amine collecting agent and the ketone collecting agent are combined to be used for flotation separation of multi-tantalum (niobium) acid radicals, so that a better flotation separation effect is achieved, and high-efficiency enrichment of the multi-tantalum (niobium) acid radicals is realized.
In a preferable scheme, the addition amount of the combined collector in the leaching solution is 500 mg/L-10 g/L. The addition amount of the combined collector in the leachate is preferably 3 g/L-6 g/L.
As a preferred scheme, the froth flotation adopts at least one of pine oil, soda and sodium hexametaphosphate as a foaming agent.
Preferably, the addition amount of the foaming agent in the leachate is 100mg/L to 200 mg/L.
As a preferable scheme, the foam product is dried at the temperature of 100-120 ℃, and the main function is to volatilize the solvent; and further calcining at 400-500 ℃, wherein the calcination mainly causes the decomposition of organic matters and the oxidation of tantalum-niobium metal into corresponding oxides.
The low-grade tantalum-niobium resource is refractory low-grade ore containing tantalum and niobium or smelting slag containing tantalum and niobium generated in the process of smelting other metals, wherein the mass percentage content of tantalum and niobium is generally within the range of 2-15%.
The method for decomposing low-grade tantalum-niobium resources and extracting tantalum-niobium by an alkaline method comprises the following steps:
(1) grinding the low-grade tantalum-niobium resource to a granularity smaller than 0.079mm (more than 90%), and uniformly mixing the low-grade tantalum-niobium resource with potassium carbonate and a composite oxidant (comprising 40-60: 8-20: 20-40 mass ratio of saltpeter, potassium hypochlorite and potassium permanganate), wherein the mixing ratio meets the following requirements: adding the composite oxidant in an amount of 1-6 wt% of the total mass of the potassium carbonate and the low-grade tantalum-niobium resource, wherein the mass of the potassium carbonate is 1-3 times of the mass of the low-grade tantalum-niobium resource, placing the mixed material into a muffle furnace, and roasting at 600-800 ℃ for 30-100 min to obtain a roasted product;
(2) leaching the roasted product in the step (1) by using water under the leaching conditions: 3-6 mL of solid-liquid ratio, 50-95 ℃ of temperature and 0.5-1.5 h of time, leaching water-soluble potassium tantalate (niobate) and adding a small amount of potassium metatantalate (niobate) in the water leaching slag obtained by filtering into the raw materials for roasting again;
(3) adding a flotation reagent (comprising 30-70% of primary amine collecting agent and 30-70% of ketone collecting agent in an addition amount of 500 mg/L-10 g/L) which is selective to poly-tantalum (niobium) acid radicals and a bubble dispersing agent (terpineol, soda, sodium hexametaphosphate and the like in an addition amount of 100 mg/L-200 mg/L) into the water leaching solution obtained in the step (2), introducing air to obtain a tantalum (niobium) enriched foam product, drying at the temperature of 100-120 ℃, and calcining at the temperature of 400-500 ℃ to obtain the high-purity tantalum-niobium oxide.
Compared with the prior art, the technical scheme of the invention has the beneficial effects that:
the technical scheme of the invention adopts an alkaline oxidizing roasting-water leaching-flotation three-step dressing-smelting combined method for low-grade tantalum-niobium resources, the method can promote the low-grade tantalum-niobium resources to be quickly and efficiently decomposed under low alkali and relatively low temperature by adopting the special oxidant, and the tantalum-niobium is converted into potassium tantalate (niobate) which is easy to dissolve in water with high selectivity, and the rare earth metal, iron, manganese and the like are converted into metal oxides, thereby realizing the separation of tantalum and niobium from rare earth metals and transition metals such as iron, manganese and the like by a simple water immersion method, on the basis, the tantalum and niobium is further enriched by a flotation separation method, and by adopting a special combined collector with high selectivity to multi-tantalum (niobium) acid radicals in the water leaching solution, the flotation separation of tantalum and niobium and elements such as tungsten, tin, silicon, aluminum and the like can be realized, the tantalum and niobium is enriched deeply, the tantalum and niobium recovery efficiency is improved, and the tantalum and niobium recovery rate reaches 95% -98%.
The technical scheme of the invention has the advantages of low alkali consumption, low roasting temperature, short time consumption in the recovery process of the low-grade tantalum-niobium resource, high recovery rate of the tantalum-niobium resource in the low-grade tantalum-niobium resource, environmental friendliness, low requirement on equipment and suitability for industrial large-scale production; compared with the traditional acid method, the method avoids using HF and high-concentration sulfuric acid, is environment-friendly, has low requirement on equipment, and is suitable for industrial continuous production; compared with the traditional alkali melting method and the high-pressure alkali solution decomposition method, the method has the advantages of weak alkalinity, less alkali consumption, short roasting time and low equipment requirement, and the tantalum-niobium oxide obtained by decomposing the potassium alkali has higher purity than sodium alkali, short roasting time and high recovery rate aiming at low-grade refractory tantalum-niobium ore.
Detailed Description
The invention will be further explained and illustrated with reference to specific examples. These examples are only for better understanding of the present invention and do not limit the scope of the present invention.
In the following examples, low-grade tantalum-niobium resources are ground to a particle size of less than 0.079mm, wherein the mass percentage of the particle size is 95%.
Example 1
And uniformly mixing 50g of ground low-grade tantalum-niobium ore (the components are shown in the table 1), 60g of potassium carbonate, 3g of niter, 1g of potassium hypochlorite and 2g of potassium permanganate, putting the mixture into a muffle furnace for roasting, controlling the roasting temperature to be 750 ℃, and roasting for 40min to obtain a roasted product. Leaching the roasted product with water at 90 ℃, wherein the solid-to-liquid ratio is 1: and 5, leaching water-soluble potassium tantalate (niobate) solution. Adding a flotation reagent into the water immersion liquid, wherein the flotation reagent comprises laurylamine and methyl isobutyl ketone according to the mass ratio of 3:7 and the total concentration of 4g/L, adding 200mg/L of pinitol oil serving as a bubble dispersant, introducing air, obtaining a tantalum (niobium) enriched foam product on the upper layer of the solution, drying at 100 ℃, and calcining at 500 ℃ to obtain the high-purity tantalum niobium oxide.
The tantalum recovery rate is 96.8 percent and the niobium recovery rate is 95.8 percent through calculation.
TABLE 1 compositions, in wt%, of certain low grade tantalum niobium ores
Example 2
And uniformly mixing 50g of ground low-grade tantalum-niobium ore (the components are shown in the table 1), 65g of potassium carbonate, 3g of niter, 1g of potassium hypochlorite and 2g of potassium permanganate, putting the mixture into a muffle furnace for roasting, controlling the roasting temperature to be 800 ℃ and the roasting time to be 40min, and obtaining a roasted product. Leaching the roasted product with water at 95 ℃, wherein the solid-liquid ratio is 1: and 5, leaching water-soluble potassium tantalate (niobate) solution. Adding a flotation reagent into the water immersion liquid, wherein the flotation reagent comprises laurylamine and methyl isobutyl ketone according to the mass ratio of 4:6 and the total concentration of 2g/L, adding 200mg/L of pinitol oil serving as a bubble dispersant, introducing air, obtaining a tantalum (niobium) enriched foam product on the upper layer of the solution, drying at 100 ℃, and calcining at 450 ℃ to obtain the high-purity tantalum niobium oxide.
The tantalum recovery rate is 96.4 percent and the niobium recovery rate is 97.2 percent through calculation.
Example 3
And uniformly mixing 50g of ground low-grade tantalum-niobium ore (the components are shown in the table 1), 70g of potassium carbonate, 3g of niter, 1g of potassium hypochlorite and 2g of potassium permanganate, putting the mixture into a muffle furnace, and roasting for 50min at the roasting temperature of 700 ℃ to obtain a roasted product. Leaching the roasted product with water at 85 ℃, wherein the solid-to-liquid ratio is 1: and 5, leaching water-soluble potassium tantalate (niobate) solution. Adding a flotation reagent into the water extract, wherein the flotation reagent comprises dodecylamine and methyl isobutyl ketone according to equal mass ratio and the total concentration is 2g/L, adding 500mg/L of pinitol oil serving as a bubble dispersant, introducing air, obtaining a tantalum (niobium) enriched foam product on the upper layer of the solution, drying at 100 ℃, and calcining at 400 ℃ to obtain the high-purity tantalum niobium oxide.
The tantalum recovery rate is 97.0 percent and the niobium recovery rate is 96.5 percent through calculation.
Example 4
And uniformly mixing 50g of ground low-grade tantalum-niobium ore (the components are shown in the table 1), 75g of potassium carbonate, 3g of niter, 1g of potassium hypochlorite and 2g of potassium permanganate, putting the mixture into a muffle furnace, and roasting for 60min at the roasting temperature of 650 ℃ to obtain a roasted product. Leaching the roasted product with water at 85 ℃, wherein the solid-to-liquid ratio is 1: and 5, leaching water-soluble potassium tantalate (niobate) solution. Adding a flotation reagent into the water extract, wherein the flotation reagent comprises laurylamine and methyl isobutyl ketone according to the mass ratio of 6:4 and the total concentration of 1g/L, adding 500mg/L of pinitol oil serving as a bubble dispersant, introducing air, obtaining a tantalum (niobium) enriched foam product on the upper layer of the solution, drying at 100 ℃, and calcining at 450 ℃ to obtain the high-purity tantalum niobium oxide.
The tantalum recovery rate is 95.7 percent and the niobium recovery rate is 96.5 percent through calculation.
Example 5
And uniformly mixing 50g of ground low-grade tantalum-niobium ore (the components are shown in the table 1), 80g of potassium carbonate, 3g of niter, 1g of potassium hypochlorite and 2g of potassium permanganate, putting the mixture into a muffle furnace for roasting, controlling the roasting temperature to be 650 ℃, and roasting for 60min to obtain a roasted product. Leaching the roasted product with water at 95 ℃, wherein the solid-liquid ratio is 1: and 5, leaching water-soluble potassium tantalate (niobate) solution. Adding a flotation reagent into the water extract, wherein the flotation reagent comprises laurylamine and methyl isobutyl ketone according to the mass ratio of 7:3 and the total concentration of 6g/L, adding 500mg/L of pinitol oil serving as a bubble dispersant, introducing air, obtaining a tantalum (niobium) enriched foam product on the upper layer of the solution, drying at 100 ℃, and calcining at 500 ℃ to obtain the high-purity tantalum niobium oxide.
The tantalum recovery rate is 98.6 percent and the niobium recovery rate is 99.0 percent through calculation.
Comparative example 1
And uniformly mixing 50g of ground low-grade tantalum-niobium ore (the components are shown in the table 1), 25g of potassium carbonate, 3g of niter, 1g of potassium hypochlorite and 2g of potassium permanganate, putting the mixture into a muffle furnace for roasting, controlling the roasting temperature to be 750 ℃, and roasting for 40min to obtain a roasted product. Leaching the roasted product with water at 90 ℃, wherein the solid-to-liquid ratio is 1: and 5, leaching water-soluble potassium tantalate (niobate) solution. Adding a flotation reagent into the water extract, wherein the flotation reagent comprises laurylamine and methyl isobutyl ketone according to the mass ratio of 7:3 and the concentration of 4g/L, adding 500mg/L of pinitol oil serving as a bubble dispersant, introducing air, obtaining a tantalum (niobium) enriched foam product on the upper layer of the solution, drying at 100 ℃, and calcining at 500 ℃ to obtain the high-purity tantalum niobium oxide.
The calculated results show that the recovery rate of tantalum is 67.9 percent and the recovery rate of niobium is 65.6 percent. The potassium carbonate ratio is too low, so that the tantalum and niobium in the low-grade tantalum-niobium ore are difficult to be converted into potassium tantalate (niobate).
Comparative example 2
And uniformly mixing 50g of ground low-grade tantalum-niobium ore (the components are shown in the table 1), 60g of potassium carbonate, 3g of niter, 1g of potassium hypochlorite and 2g of potassium permanganate, putting the mixture into a muffle furnace for roasting, controlling the roasting temperature to be 550 ℃ and the roasting time to be 40min, and obtaining a roasted product. Leaching the roasted product with water at 90 ℃, wherein the solid-to-liquid ratio is 1: and 5, leaching water-soluble potassium tantalate (niobate) solution. Adding a flotation reagent into the water extract, wherein the flotation reagent comprises laurylamine and methyl isobutyl ketone according to the mass ratio of 3:7 and the concentration of 4g/L, adding 500mg/L of pinitol oil serving as a bubble dispersant, introducing air, obtaining a tantalum (niobium) enriched foam product on the upper layer of the solution, drying at 100 ℃, and calcining at 500 ℃ to obtain the high-purity tantalum niobium oxide.
The tantalum recovery rate was calculated to be 63.9% and the niobium recovery rate was calculated to be 61.5%. The calcination temperature is too low, so that the tantalum and niobium in the low-grade tantalum-niobium ore are difficult to be sufficiently converted into potassium tantalate (niobate).
Comparative example 3
And uniformly mixing 50g of ground low-grade tantalum-niobium ore (the components are shown in the table 1), 60g of potassium carbonate, 3g of niter, 1g of potassium hypochlorite and 2g of potassium permanganate, putting the mixture into a muffle furnace for roasting, controlling the roasting temperature to be 750 ℃, and roasting for 40min to obtain a roasted product. Leaching the roasted product with water at 95 ℃, wherein the solid-liquid ratio is 1: and 5, leaching water-soluble potassium tantalate (niobate) solution. Adding a flotation reagent into the water immersion liquid, wherein the flotation reagent comprises laurylamine and methyl isobutyl ketone according to the mass ratio of 4:6 and the total concentration of 200mg/L, adding 500mg/L of pinitol oil serving as a bubble dispersant, introducing air, obtaining a tantalum (niobium) enriched foam product on the upper layer of the solution, drying at 100 ℃, and calcining at 500 ℃ to obtain the high-purity tantalum niobium oxide.
The calculated results show that the recovery rate of tantalum is 70.3 percent and the recovery rate of niobium is 65.7 percent. The flotation requirement dosage is too low, and the potassium tantalate (niobate) is difficult to effectively separate by flotation.
Comparative example 4
And uniformly mixing 50g of ground low-grade tantalum-niobium ore (the components are shown in the table 1), 75g of potassium carbonate, 3g of niter, 1g of potassium hypochlorite and 2g of potassium permanganate, putting the mixture into a muffle furnace for roasting, controlling the roasting temperature to be 750 ℃ and the roasting time to be 60min, and obtaining a roasted product. Leaching the roasted product with water at 85 ℃, wherein the solid-to-liquid ratio is 1: and 5, leaching water-soluble potassium tantalate (niobate) solution. Adding a flotation reagent dodecylamine with the concentration of 1g/L into the water leaching solution, adding 500mg/L of pinitol oil serving as a bubble dispersant, introducing air, obtaining a tantalum (niobium) enriched foam product on the upper layer of the solution, drying at 100 ℃, and calcining at 450 ℃ to obtain the high-purity tantalum niobium oxide.
The calculated results show that the recovery rate of tantalum is 85.7 percent and the recovery rate of niobium is 86.5 percent. The flotation recovery rate is lower by adopting single dodecylamine as the flotation reagent.
Comparative example 5
And uniformly mixing 50g of ground low-grade tantalum-niobium ore (the components are shown in the table 1), 70g of potassium carbonate and 5g of niter, putting the mixture into a muffle furnace for roasting, controlling the roasting temperature to be 750 ℃ and the roasting time to be 50min, and obtaining a roasted product. Leaching the roasted product with water at 85 ℃, wherein the solid-to-liquid ratio is 1: and 5, leaching water-soluble potassium tantalate (niobate) solution. Adding a flotation reagent into the water extract, wherein the flotation reagent comprises lauryl amine, methyl isobutyl ketoamine and the like in a mass ratio, the total concentration is 2g/L, adding 500mg/L of pinitol oil serving as a bubble dispersant, introducing air, obtaining a tantalum (niobium) enriched foam product on the upper layer of the solution, drying at 100 ℃, and calcining at 400 ℃ to obtain the high-purity tantalum niobium oxide.
The calculated results show that the recovery rate of tantalum is 81.5 percent and the recovery rate of niobium is 85.6 percent. It is difficult to completely convert the tantalum and niobium in the low-grade tantalum-niobium ore into potassium tantalate (niobate) by using single saltpeter as an oxidant.
In conclusion, through analysis of comparative examples, the low-grade tantalum-niobium resource is enriched and separated by a roasting-water leaching-flotation three-step dressing-smelting combined method, so that the decomposition efficiency can be obviously improved, and the method is low in alkali consumption, short in roasting time, environment-friendly, low in equipment requirement and suitable for industrial continuous production.
Claims (10)
1. A method for decomposing low-grade tantalum-niobium resources and extracting tantalum-niobium by an alkaline method is characterized by comprising the following steps: the method comprises the following steps:
1) uniformly mixing the ground low-grade tantalum-niobium resource with potassium carbonate and a composite oxidant, and roasting to obtain a roasted product; the composite oxidant comprises saltpeter, potassium hypochlorite and potassium permanganate;
2) treating the roasted product by water leaching to obtain leachate containing potassium polytantalate and potassium polyniobate;
3) performing foam flotation on the leachate to obtain a foam product enriched with tantalum and niobium; the froth flotation adopts a primary amine collecting agent and a ketone collecting agent as combined collecting agents;
4) and (3) drying and calcining the foam product in sequence to obtain the oxide containing tantalum and niobium.
2. The method for decomposing low-grade tantalum-niobium resources and extracting tantalum-niobium by the alkaline method according to claim 1, is characterized in that: the low-grade tantalum-niobium resource is ground to a particle size of less than 0.079mm, and the mass percentage content of the particle size is more than 90%.
3. The method for decomposing low-grade tantalum-niobium resources and extracting tantalum-niobium by the alkaline method according to claim 1, is characterized in that:
the addition amount of the composite oxidant is 1-6 wt% of the total mass of the potassium carbonate and the low-grade tantalum-niobium resource;
the mass ratio of the low-grade tantalum-niobium resource to the potassium carbonate is 1: 1-3.
4. The method for decomposing low-grade tantalum-niobium resources and extracting tantalum-niobium by the alkaline method according to claim 1 or 3, which is characterized by comprising the following steps of: the composite oxidant comprises the following components in parts by weight: 40-60 parts of saltpeter; 8-20 parts of potassium hypochlorite; 20-40 parts of potassium permanganate.
5. The method for decomposing low-grade tantalum-niobium resources and extracting tantalum-niobium by the alkaline method according to claim 1, is characterized in that: the roasting treatment conditions are as follows: the temperature is 600-800 ℃, and the time is 30-100 min.
6. The method for decomposing low-grade tantalum-niobium resources and extracting tantalum-niobium by the alkaline method according to claim 1, is characterized in that: the water immersion treatment conditions are as follows: the solid-liquid ratio is 1g: 3-6 mL, the temperature is 50-95 ℃, and the time is 0.5-1.5 h.
7. The method for decomposing low-grade tantalum-niobium resources and extracting tantalum-niobium by the alkaline method according to claim 1, is characterized in that: the combined collector comprises the following components in percentage by mass: 30-70% of primary amine collecting agent; and 30-70% of ketone collecting agent.
8. The method for decomposing low-grade tantalum-niobium resources and extracting tantalum-niobium by the alkaline method according to claim 7, is characterized in that:
the primary amine collecting agent is at least one of dioctyl amine, dodecylamine, tetradecylamine and hexadecylamine;
the ketone collector is at least one of methyl isobutyl ketone, cyclohexanone and diisopropyl ketone.
9. The method for decomposing low-grade tantalum-niobium resources and extracting tantalum-niobium by the alkaline method according to claim 1 or 7, is characterized in that: the addition amount of the combined collector in the leachate is 500 mg/L-10 g/L.
10. The method for decomposing low-grade tantalum-niobium resources and extracting tantalum-niobium by the alkaline method according to claim 1, is characterized in that: the froth flotation adopts at least one of pinitol oil, soda and sodium hexametaphosphate as a foaming agent;
the addition amount of the foaming agent in the leaching solution is 100 mg/L-200 mg/L.
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