CN110396610B - Method for treating titanium minerals and metal silicate minerals through ammonium salt pressure pyrolysis - Google Patents
Method for treating titanium minerals and metal silicate minerals through ammonium salt pressure pyrolysis Download PDFInfo
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- C01F7/02—Aluminium oxide; Aluminium hydroxide; Aluminates
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Abstract
The invention discloses a method for treating titanium minerals and metal silicates by ammonium salt pressure pyrolysis. The method comprises the steps of pretreating titanium minerals and metal silicates, uniformly mixing the pretreated titanium minerals and metal silicates with ammonium salts according to a certain proportion, carrying out low-temperature pyrolysis in a high-pressure kettle, carrying out closed roasting when the pH value of gas at a sampling port is acidic, efficiently salinizing valuable metals in the minerals by utilizing the ammonium salts under a closed condition, releasing valuable elements embedded in lattices, and carrying out acid leaching on roasted slag in dilute acid to obtain a high-concentration acid leaching solution of titanium, aluminum and magnesium; hydrolyzing titanium, aluminum and magnesium in the pickle liquor step by step, and recovering the titanium, aluminum and magnesium in a stepped manner to obtain a high-purity product; meanwhile, the ammonia water is recycled, and the ammonium salt can be recycled through evaporation and crystallization. The method realizes the step recovery of titanium, magnesium, aluminum and ammonium salt in titanium minerals and metal silicates, and realizes the purpose of recycling waste resources.
Description
Technical Field
The invention belongs to the field of industrial solid waste resource utilization, and relates to a method for extracting titanium, magnesium and aluminum by treating titanium minerals and metal silicate minerals through ammonium salt pressurization pyrolysis.
Background
Titanium is abundant on the earth, the earth crust abundance is 0.61%, the content of the titanium is higher than that of common copper, nickel, tin, lead and zinc, more than 140 kinds of known minerals are known, but the known minerals are symbiotic minerals, and the ilmenite, the titanomagnetite and the hematite are mainly present. The vanadium titano-magnetite in Panxi area of China is a world extra large deposit with multi-metal symbiosis, the storage capacity reaches 96.6 hundred million (million) t, and Titanium (TiO)2)8.73The titaniferous ore is difficult to separate by adopting a purely physical ore dressing method because the structure is compact, the gangue content is high, the crystal grain size is about 10 mu m basically, and the titaniferous ore phase is difficult to separate. At present, about 53 percent of titanium in vanadium titano-magnetite enters a blast furnace smelting link in an iron ore concentrate mode to finally form titanium-containing blast furnace slag. For years, most of the cement is directly stockpiled or used as cement production ingredients, so that the waste of resources is serious, a large amount of land is occupied, and the environment is polluted. The metal silicate minerals are oxygen-containing acid salt minerals formed by combining metal cations and silicate radicals, are main minerals forming the earth crust and the upper mantle, and many silicate minerals have high magnesium and aluminum contents, such as kaolinite, montmorillonite, zeolite and the like, which are important non-metallic mineral raw materials and materials.
At present, the direct sulfuric acid leaching method, the sub/molten salt method and the high-temperature carbonitriding-low-temperature chlorination method are mainly used in the process research of titanium minerals. The sulfuric acid direct leaching method requires high sulfuric acid concentration, has large corrosion to equipment, and simultaneously causes secondary pollution due to a large amount of acid leaching solution and acid leaching residues; the sub/molten salt method is a method for decomposing titanium-containing blast furnace slag in a system with water content less than 50% in a sub molten salt zone or no water in a molten salt zone system, but generally the viscosity of molten reactants is high, and the preparation of the hydrolyzed raw material titanyl sulfate solution is greatly influenced by molten salt reaction and washing. The high-temperature selective reduction carbonitriding-low-temperature chlorination method is used for high-temperature treatment at the temperature of more than 1500 ℃, the energy consumption is huge, the used raw materials such as nitrogen, carbon and chlorine are required to be in contact with the whole furnace slag, the utilization rate is low, and the cost is high. The flotation method or the alkaline leaching method is mostly adopted for metal silicate minerals, but metal ions inevitably exist in a mineral flotation system, and the metal ions in the solution have activation or inhibition effects on the silicate minerals and sometimes influence the dispersion behavior of mineral particles due to the change of the surface electrical property of the minerals, so that the flotation is extremely unstable; the alkali dissolution method has higher cost and more impurity elements in the product.
Patent document "a method for producing titanium alloy directly from titanium-containing mineral" (CN1888101A), discloses a method for preparing titanium alloy directly from titanium-containing mineral, which comprises using titanium-containing mineral as raw material, adding reducing agent to perform electric arc furnace smelting for iron removal to obtain high titanium slag, adding electrolytic aluminum to the high titanium slag in a plasma high temperature furnace to reduce titanium oxide to obtain titanium aluminum alloy mother liquor, then adding 1-10% of calcium fluoride and deoxidizer, and smelting in a vacuum electric furnace to obtain titanium aluminum alloy with oxygen content of 0.05-0.15%. Although the invention realizes metal extraction and alloying in the metal thermal reduction and shortens the process period, the whole process is carried out in an electric furnace, the cost is higher, and when the method of electroslag remelting is adopted to remove impurity elements, the electrode consumption is serious under strong current.
Patent document "method for preparing titanium dioxide by using high titanium type blast furnace slag" (CN108975393A), discloses a method for obtaining carbide slag by high temperature carbonization and crushing treatment; the carbonized residue and chlorine gas are subjected to low-temperature chlorination reaction to obtain the product containing TiCl4Chlorination of the gaseous reaction mixture, crude TiCl after condensation4(ii) a The crude TiCl is treated with fatty acid4Vanadium removal treatment and rectification are carried out to obtain rectified TiCl4Rectifying TiCl4And O2Oxidation reaction is carried out to obtain the product containing TiO2And Cl2The mixed product of the oxidation reaction is cooled and filtered to obtain the titanium dioxide product. Although the method fully recycles the titanium in the high-titanium blast furnace slag, the high-temperature carbonization temperature is up to 1580-1700 ℃, the requirement on equipment is high, the process is complex, and meanwhile, the vanadium removal treatment by fatty acid flotation is required, so that the industrial application possibility is low.
Patent document "a method for recovering vanadium and silicon by asynchronously converting silicate vanadium ore" (CN105331816A) discloses a method for recovering silicon and vanadium by two-stage alkaline leaching of silicate with low concentration and high concentration, which is suitable for various silicate minerals such as stone coal and clay. The method mainly comprises the steps of finely grinding silicate minerals to-200 meshes which account for more than 80%, then leaching the silicate minerals by low-concentration alkali liquor at high temperature and high pressure, filtering, collecting filtrate, and preparing water glass with the modulus of more than or equal to 2; and performing high-temperature alkaline leaching-low-temperature alkaline leaching on filter residues to obtain high-grade concentrate powder, and adding 25-40% sulfuric acid to recover vanadium. Although the invention adopts high-low concentration two-stage alkali leaching of silicate minerals, the comprehensive utilization rate of resources is improved, the final vanadium product has more impurity elements and lower grade, can only be used for preliminary enrichment, and simultaneously has large acid-base consumption and higher cost.
In conclusion, the existing method for treating the titanium-containing minerals and the metal silicate minerals generally has the defects of low product purity, incapability of realizing clean production, high treatment cost and the like due to complex structure and serious associated phenomena of valuable metals, and a method for economically and feasibly utilizing the titanium-containing blast furnace slag has not been found so far.
Disclosure of Invention
The technical problem to be solved by the invention is as follows: how to overcome the problems of low leaching rate of titanium, magnesium and aluminum in extracted titanium minerals and metal silicate minerals and low product purity, and provides a method for treating the titanium minerals and the metal silicate minerals by ammonium salt pressure pyrolysis.
The invention provides a method for treating titanium minerals and metal silicate minerals by ammonium salt pressure pyrolysis, which is used for efficiently recovering valuable metals of titanium, magnesium and aluminum, effectively solving the problem of low product purity, obtaining high-valued products, realizing ammonium salt recycling and having very important theoretical and practical significance for efficient utilization of the titanium-containing minerals and the metal silicate.
The invention provides a method for treating titanium minerals and metal silicate minerals by ammonium salt pressure pyrolysis, which comprises the following steps:
step one, drying and scattering titanium minerals and metal silicate minerals to obtain powder materials;
step two, uniformly mixing the powder material obtained in the step one with ammonium salt, roasting for 0.5-2 h at 250-450 ℃ in a roasting device, opening a sampling port of the roasting device at an initial stage, collecting generated ammonia gas, introducing water to form ammonia water, and sealing and pressurizing for pyrolysis when the pH value of the gas at the sampling port is acidic;
step three, adding the roasting slag obtained by pyrolysis in the step two into a dilute acid solution for acid leaching and then filtering;
and step four, preserving the heat of the filtrate obtained in the step three for 0.5-1 h in a boiling state, adding water for dilution, preserving the heat for 0.5-2 h in the boiling state, then recrystallizing to obtain off-white hydrolysate, filtering and collecting filtrate and filter residue, washing the filter residue with dilute acid to remove impurities, and calcining to obtain the titanium dioxide.
The method according to the present invention, wherein as an option, the method may further comprise:
step five, dropwise adding the filtrate obtained in the step four into the ammonia water formed in the step two until the pH value is 5-6, allowing hydrolysate to appear, filtering and collecting the filtrate and filter residues, washing the filter residues with water to remove impurities, and calcining to obtain an aluminum oxide product;
step six, dropwise adding the filtrate obtained in the step five into the ammonia water formed in the step two until the pH value is more than or equal to 12, generating hydrolysate, filtering and collecting the filtrate and filter residue, washing the filter residue with water to remove impurities, and calcining to obtain a magnesium oxide product;
and step seven, adding the ammonia water formed in the step two into the filtrate obtained in the step six, and evaporating and crystallizing to obtain ammonium salt crystals.
According to the method, as an option, in the fifth step, the filter residue is washed and subjected to impurity removal for 4-8 times, and then calcined at 1000-1300 ℃ for 1-5 hours;
and sixthly, washing the filter residue to remove impurities for 3-6 times, and calcining for 2-4 hours at 400-800 ℃.
According to the method, as an option, in the first step, titanium minerals and metal silicate minerals are ball-milled until the particle size of-0.074 mm is larger than or equal to 75%, and are dried and scattered at 50-110 ℃ to obtain a gray powder material with the water content of less than or equal to 2 wt%.
According to the method, as an option, in the second step, the judgment standard for controlling the opening and closing of the sampling port valve is that a pH meter detects that the pH of the gas at the sampling port is less than or equal to 7, so that valuable elements in minerals are fully salted to generate soluble salts. The roasting device in the second step can be any high-pressure roasting device in the field, including but not limited to an autoclave and the like.
According to the method, as an option, in the second step, the powder material and ammonium salt are uniformly mixed according to the proportion of 1: 0.5-1: 5, wherein the ammonium salt is ammonium sulfate and/or ammonium chloride.
The method according to any of the invention, wherein as an option, step three is firingAdding 2-30% of H into the slag according to the liquid-solid ratio of 10: 1-2: 12SO4Or HCl solution, and acid leaching for 0.5-4 h at 40-90 ℃ and a stirring speed of 100-400 rpm.
The method according to any one of the invention, wherein, as an option, the filtrate in the fourth step is diluted by adding water with 0.4-2 times volume fraction.
According to any one of the methods provided by the invention, as an option, in the fourth step, the filter residue is washed by dilute sulfuric acid or dilute hydrochloric acid with the concentration of 1% -5% for 6-10 times to remove impurities, and then calcined at 850-1000 ℃ to obtain the titanium dioxide.
The titanium minerals and the metal silicate minerals are mainly perovskite, black titanium, diopside, calcium aluminum garnet and magnesia-alumina spinel, and insoluble valuable elements are converted into water-soluble sulfate/chloride through pressure pyrolysis of ammonium salts under a closed condition, so that the aim of efficiently leaching the valuable elements is fulfilled.
In order to recycle the ammonium salt crystal product more conveniently, preferably, sulfuric acid is used for acid leaching and washing when ammonium sulfate is added in ammonium salt roasting; when ammonium chloride is added during roasting of ammonium salt, hydrochloric acid is preferably used for acid leaching and washing.
In the present invention, ammonium sulfate and/or ammonium chloride are thermally decomposed to produce SO2、SO3Or HCl gas, can realize the full salinization of valuable metals of minerals by adjusting a sampling port valve, and simultaneously decompose the generated NH3And the ammonium salt is returned to participate in salt reaction and is evaporated and crystallized, so that the recycling of the ammonium salt is realized, the cost is saved, and the method is more environment-friendly.
The technical scheme of the invention has the following beneficial effects:
(1) SO is added by controlling the opening and closing of a valve of the sampling port2、SO3Or the HCl gas is completely sealed in the reaction kettle, so that the HCl gas is fully salted with the titanium, magnesium and aluminum valuable metals to generate soluble sulfate/chloride, the leaching rate of the titanium, magnesium and aluminum is greatly improved, and the resource utilization rate is effectively improved;
(2) the invention carries out titanium-magnesium-aluminum gradient recovery on titanium minerals and metal silicate minerals, and the leaching rate of titanium is more than or equal to 90 percent after roasting ammonium salt and leaching with dilute acidThe leaching rate of aluminum is more than or equal to 80 percent, the leaching rate of magnesium is more than or equal to 95 percent, and the leaching rate of aluminum is higher than or equal to NH3The pH value is adjusted by recycling, and high-value products of titanium, aluminum and magnesium are recycled in a gradient way, so that the economic benefit is obvious;
(3) the invention only needs low concentration of H2SO4The titanium, the magnesium and the aluminum in the roasted product can be leached by HCl, so that the acid consumption is effectively reduced;
(4) according to the invention, ammonia gas is collected by controlling the valve of the sampling port, and is added into the final filtrate to be evaporated and crystallized to generate ammonium salt crystals, so that the recycling of the ammonium salt crystals is realized, the cost is reduced, and the method is green and environment-friendly.
Drawings
FIG. 1 is a schematic process flow diagram of a method for treating titanium minerals and metal silicate minerals by ammonium salt pressure pyrolysis.
Detailed Description
In order to make the technical problems, technical solutions and advantages of the present invention more apparent, the following description should be taken in conjunction with the accompanying drawings and the embodiments.
As a preferred method for implementing the invention, the method for carrying out pressure pyrolysis treatment on titanium minerals and metal silicate minerals by using ammonium salt comprises the following steps:
step one, ball-milling titanium minerals and metal silicate minerals until the diameter of minus 0.074mm is more than or equal to 75%, drying and scattering at 50-110 ℃ to obtain a gray powder material with the water content less than or equal to 2 wt%;
step two, uniformly mixing the powder material obtained in the step one with ammonium salt (ammonium sulfate and/or ammonium chloride) according to the proportion of 1: 0.5-1: 5, roasting in an autoclave at the temperature of 250-450 ℃ for 0.5-2 h, opening a sampling port at the initial stage, collecting generated ammonia gas, introducing distilled water to form ammonia water, and sealing and pressurizing for pyrolysis when the pH value of the gas at the sampling port is acidic;
step three, adding 2-30% by mass of H into the roasting slag obtained in the step two according to the liquid-solid ratio of 10: 1-2: 12SO4Or the HCl solution is subjected to acid leaching for 0.5 to 4 hours at the temperature of 40 to 90 ℃ and the stirring speed of 100 to 400rpm, and then the solution is filtered, wherein the leaching rate of titanium in the filtrate is more than or equal to 96 percent, the leaching rate of aluminum is more than or equal to 92 percent, and the leaching rate of magnesium is more than or equal to 97 percent;
step four, preserving the heat of the filtrate obtained in the step three for 0.5-1 hour in a boiling state, adding 0.4-2 times of distilled water for dilution, preserving the heat for 0.5-2 hours in the boiling state, recrystallizing to obtain off-white hydrolysate, filtering and collecting filtrate and filter residue, washing the filter residue with 1-5% dilute sulfuric acid or dilute hydrochloric acid to remove impurities for 6-10 times, and calcining at 850-1000 ℃ to obtain titanium dioxide with the purity of more than 98%;
further, the invention can also continue the following steps five-seven:
step five, dropwise adding the filtrate obtained in the step four into the ammonia water formed in the step two until the pH value is 5-6, allowing hydrolysate to appear, filtering and collecting the filtrate and filter residue, washing the filter residue with distilled water to remove impurities for 4-8 times, and calcining at 1000-1300 ℃ for 1-5 hours to obtain an alumina product with the purity of more than 95%;
step six, dropwise adding the filtrate obtained in the step five into the ammonia water formed in the step two until the pH is more than or equal to 12, generating hydrolysate, filtering and collecting the filtrate and filter residue, washing the filter residue with distilled water for 3-6 times to remove impurities, and calcining at 400-800 ℃ for 2-4 h to obtain a magnesium oxide product with the purity of more than 97%;
and step seven, adding the ammonia water formed in the step two into the filtrate obtained in the step six, and evaporating and crystallizing to obtain ammonium salt crystals, thereby realizing recycling.
And in the second step, the judgment standard for controlling the opening and closing of the valve at the sampling port is that a pH meter detects that the pH of the gas at the sampling port is less than or equal to 7 so as to realize the full salinization of valuable elements in the minerals and generate soluble salts.
Example 1
The Ninggang steel group steel slag has the following chemical component analysis:
TABLE 1 analysis results of steel slag components
The sample is steel slag from Ning Steel group oxygen-blown converter, the analysis result of chemical composition is shown in Table 1, and the main minerals are calcium aluminum garnet and cancrinite.
The method for recovering the titanium, magnesium and aluminum valuable metals in a gradient manner according to the method shown in the figure 1 comprises the following specific steps:
(1) roasting with ammonium sulfate
Ball-milling the steel slag until the particle size is 81.09% at-0.074 mm, drying and scattering the steel slag at 90 ℃, uniformly mixing the powder material and ammonium sulfate according to the proportion of 1:3, pyrolyzing the mixture in an autoclave at the temperature of 300 ℃ for 1.5h, opening a sampling port at the initial stage, collecting generated ammonia gas, introducing distilled water to form ammonia water, and sealing and pressurizing the mixture for pyrolysis when the pH value of the gas at the sampling port is less than or equal to 7.
(2) Acid leaching
And (2) mixing the roasting slag obtained in the step (1) according to a liquid-solid ratio of 5: 1 to 10% by mass of H2SO4And (3) carrying out acid leaching in the solution in a water bath with the temperature of 70 ℃ and the stirring speed of 350rpm for 3h, filtering the leaching solution after the acid leaching is finished, and measuring by ICP (inductively coupled plasma) to calculate the leaching rates of the titanium, the magnesium and the aluminum in the leaching solution to be 97.69%, 94.29% and 98.45% respectively.
(3) High value product recovery
Preserving the heat of the acid leaching filtrate obtained in the step (2) for 0.5h in a boiling state, adding 1 time of dilution water, preserving the heat for 1h, recrystallizing to obtain off-white hydrolysate, filtering and collecting the filtrate and filter residue, washing the filter residue with 2% dilute sulfuric acid for 5 times to remove impurities, and calcining at 920 ℃ to obtain titanium dioxide with the purity of 98.59%; dropwise adding the ammonia water collected in the step (1) into the filtrate to adjust the pH value to 5.5, generating hydrolysate, filtering and collecting the filtrate and filter residue, washing the filter residue with distilled water to remove impurities for 5 times, and calcining at 1100 ℃ for 3h to obtain an alumina product with the purity of 96.34%; dropwise adding the ammonia water collected in the step (1) into the filtrate to adjust the pH value to 13, generating hydrolysate, filtering and collecting filtrate and filter residue, washing the filter residue with distilled water to remove impurities for 5 times, and calcining at 650 ℃ for 3 hours to obtain a magnesium oxide product with the purity of 98.57%; and adding the residual ammonia water into the filtrate, and evaporating and crystallizing to obtain ammonium sulfate crystals, thereby realizing the cyclic utilization of the ammonium sulfate crystals.
Example 2
The titanium-containing blast furnace slag of the Panzhihua steel group has the following chemical component analysis:
TABLE 2 analysis results of components of titanium-containing blast furnace slag
The sample is taken from titanium-containing blast furnace slag obtained after iron making by Panzhihua steel group, the analysis result of chemical components is shown in Table 2, and the main minerals are perovskite and diopside.
The method for recovering the titanium, magnesium and aluminum valuable metals in a gradient manner according to the method shown in the figure 1 comprises the following specific steps:
(1) roasting with ammonium sulfate
Ball-milling the titanium-containing blast furnace slag until the diameter of the slag is 77.28 percent when the slag is-0.074 mm, drying and scattering the slag at 95 ℃, uniformly mixing the powder material and ammonium sulfate according to the proportion of 1:2, pyrolyzing the slag in a high-pressure kettle for 2 hours at the temperature of 350 ℃, opening a sampling port at the initial stage, collecting generated ammonia gas, introducing distilled water to form ammonia water, and sealing and pressurizing for pyrolysis when the pH value of the gas at the sampling port is less than or equal to 7.
(2) Acid leaching
And (2) mixing the roasting slag obtained in the step (1) according to a liquid-solid ratio of 8: 1 to 8% by mass of H2SO4And (3) carrying out acid leaching in the solution for 2.5h in a water bath with the temperature of 80 ℃ and the stirring speed of 390rpm, filtering the leaching solution after the acid leaching is finished, and measuring and calculating leaching rates of titanium, magnesium and aluminum valuable metals in the leaching solution to be 96.85%, 93.21% and 97.18% respectively through ICP.
(3) High value product recovery
Preserving the heat of the acid leaching filtrate obtained in the step (2) for 0.7h in a boiling state, adding 1.5 times of dilution water, preserving the heat for 1.2h, recrystallizing to obtain off-white hydrolysate, filtering and collecting the filtrate and filter residue, washing the filter residue with 3.5% dilute sulfuric acid to remove impurities for 5 times, and calcining at 950 ℃ to obtain titanium dioxide with the purity of 98.81%; dropwise adding the ammonia water collected in the step (1) into the filtrate to adjust the pH value to 5.3, generating hydrolysate, filtering and collecting the filtrate and filter residue, washing the filter residue with distilled water to remove impurities for 6 times, and calcining at 1250 ℃ for 2 hours to obtain an alumina product with the purity of 97.26%; dropwise adding the ammonia water collected in the step (1) into the filtrate to adjust the pH value to 12.5, generating hydrolysate, filtering and collecting filtrate and filter residue, washing the filter residue with distilled water to remove impurities for 5 times, and calcining at 700 ℃ for 2 hours to obtain a magnesium oxide product with the purity of 97.59%; and adding the residual ammonia water into the filtrate, and evaporating and crystallizing to obtain ammonium sulfate crystals, thereby realizing the cyclic utilization of the ammonium sulfate crystals.
Example 3
The chemical component analysis of the high titanium slag of the river steel company Limited is as follows:
TABLE 3 analysis results of the composition of high titanium slag
The sample was obtained from high titanium slag smelted by electric furnace of Hei Steel products Ltd, and the analysis results of the chemical components are shown in Table 3, and the main minerals were black titanium ore, silicate glass body and free TiO2。
The method for recovering the titanium, magnesium and aluminum valuable metals in a gradient manner according to the method shown in the figure 1 comprises the following specific steps:
(1) roasting with ammonium sulfate
Ball-milling high-titanium slag until the particle diameter is 71.63% at-0.074 mm, drying and scattering at 100 ℃, uniformly mixing the powder material and ammonium sulfate according to the proportion of 1:4, pyrolyzing in an autoclave at 400 ℃ for 2.5h, opening a sampling port at the initial stage, collecting generated ammonia gas, introducing distilled water to form ammonia water, and sealing and pressurizing for pyrolysis when the pH value of the gas at the sampling port is less than or equal to 7.
(2) Acid leaching
And (2) mixing the roasting slag obtained in the step (1) according to a liquid-solid ratio of 10:1 to 12% by mass of H2SO4And (3) carrying out acid leaching on the solution in a water bath with the temperature of 85 ℃ and the stirring speed of 355rpm for 3.5h, filtering the leaching solution after the acid leaching is finished, and measuring and calculating the leaching rates of the titanium, magnesium and aluminum valuable metals in the leaching solution to be 97.38%, 94.62% and 98.57% respectively through ICP (inductively coupled plasma).
(3) High value product recovery
Preserving the heat of the acid leaching filtrate obtained in the step (2) for 0.8h in a boiling state, adding 0.6 times of dilution water, preserving the heat for 0.6h, recrystallizing to obtain off-white hydrolysate, filtering and collecting the filtrate and filter residue, washing the filter residue with 4% dilute sulfuric acid to remove impurities for 7 times, and calcining at 960 ℃ to obtain titanium dioxide with the purity of 99.34%; dropwise adding the ammonia water collected in the step (1) into the filtrate to adjust the pH value to be 5.1, generating hydrolysate, filtering and collecting the filtrate and filter residue, washing the filter residue with distilled water to remove impurities for 7 times, and calcining at 1150 ℃ for 1.5h to obtain an alumina product with the purity of 96.44%; dropwise adding the ammonia water collected in the step (1) into the filtrate to adjust the pH value to 13.2, generating hydrolysate, filtering and collecting filtrate and filter residue, washing the filter residue with distilled water to remove impurities for 6 times, and calcining at 750 ℃ for 1.5h to obtain a magnesium oxide product with the purity of 98.16%; and adding the residual ammonia water into the filtrate, and evaporating and crystallizing to obtain ammonium sulfate crystals, thereby realizing the cyclic utilization of the ammonium sulfate crystals.
Example 4
Kaolin from Fujian company has the following chemical composition analysis:
TABLE 4 Kaolin composition analysis results
The sample is obtained from Kaolin of Zhangzhou company of Fujian, the chemical composition analysis results are shown in Table 4, and the main minerals are halloysite and illite
The method for recovering the titanium, magnesium and aluminum valuable metals in a gradient manner according to the method shown in the figure 1 comprises the following specific steps:
(1) roasting with ammonium sulfate
Ball-milling kaolin until the particle size of-0.074 mm is 82.54%, drying and scattering at 105 ℃, uniformly mixing the powder material and ammonium sulfate according to the proportion of 1:3.5, pyrolyzing in an autoclave at the temperature of 420 ℃ for 2.3h, opening a sampling port at the initial stage, collecting generated ammonia gas, introducing distilled water to form ammonia water, and sealing and pressurizing for pyrolysis when the pH value of gas at the sampling port is less than or equal to 7.
(2) Acid leaching
And (2) mixing the roasting slag obtained in the step (1) according to a liquid-solid ratio of 12: 1 to 7% by mass of H2SO4And (3) carrying out acid leaching on the solution in a water bath with the temperature of 93 ℃ and the stirring speed of 410rpm for 4h, filtering the leaching solution after the acid leaching is finished, and measuring and calculating leaching rates of titanium, magnesium and aluminum valuable metals in the leaching solution to be 98.38%, 95.62% and 99.17% respectively through ICP.
(3) High value product recovery
Preserving the heat of the acid leaching filtrate obtained in the step (2) for 0.4h in a boiling state, adding 1.3 times of dilution water, preserving the heat for 1.2h, recrystallizing to obtain off-white hydrolysate, filtering and collecting the filtrate and filter residue, washing the filter residue with 3.5% dilute sulfuric acid to remove impurities for 6 times, and calcining at 980 ℃ to obtain 98.52% titanium dioxide powder; dropwise adding the ammonia water collected in the step (1) into the filtrate to adjust the pH value to 5.4, generating hydrolysate, filtering and collecting the filtrate and filter residue, washing the filter residue with distilled water to remove impurities for 8 times, and calcining at 1230 ℃ for 1.2h to obtain an alumina product with the purity of 97.29%; dropwise adding the ammonia water collected in the step (1) into the filtrate to adjust the pH value to 13.7, generating hydrolysate, filtering and collecting filtrate and filter residue, washing the filter residue with distilled water to remove impurities for 7 times, and calcining at 770 ℃ for 1.8h to obtain a magnesium oxide product with the purity of 97.96%; and adding the residual ammonia water into the filtrate, and evaporating and crystallizing to obtain ammonium sulfate crystals, thereby realizing the cyclic utilization of the ammonium sulfate crystals.
Example 5
The titanium-containing blast furnace slag of the Panzhihua steel group has the following chemical component analysis:
TABLE 5 analysis results of components of titanium-containing blast furnace slag
The sample is taken from titanium-containing blast furnace slag obtained after iron making by Panzhihua steel group, the analysis result of chemical components is shown in Table 2, and the main minerals are perovskite, magnesia-alumina spinel and titanium-rich diopside.
The method for recovering the titanium, magnesium and aluminum valuable metals in a gradient manner according to the method shown in the figure 1 comprises the following specific steps:
(1) roasting of ammonium chloride
Ball-milling titanium-containing blast furnace slag until the particle size of the slag is 83.67% with the particle size of-0.074 mm, drying and scattering the slag at 110 ℃, uniformly mixing the powder material and ammonium chloride according to the proportion of 1:0.8, pyrolyzing the slag in an autoclave at the temperature of 310 ℃ for 2.5h, opening a sampling port at the initial stage, collecting generated ammonia gas, introducing distilled water to form ammonia water, and sealing and pressurizing the slag for pyrolysis when the pH value of the gas at the sampling port is less than or equal to 7.
(2) Acid leaching
And (2) mixing the roasting slag obtained in the step (1) according to a liquid-solid ratio of 6: 1, adding the mixture into an HCl solution with the mass fraction of 3.5%, carrying out acid leaching in a water bath with the temperature of 73 ℃ and the stirring speed of 375rpm for 3.6h, filtering the leaching solution after the acid leaching is finished, and measuring and calculating the leaching rates of the titanium, magnesium and aluminum valuable metals in the leaching solution to be 99.27%, 94.92% and 98.25% respectively through ICP.
(3) High value product recovery
Preserving the heat of the acid leaching filtrate obtained in the step (2) for 0.9h in a boiling state, adding 1.3 times of dilution water, preserving the heat for 1.1h, recrystallizing to obtain off-white hydrolysate, filtering and collecting the filtrate and filter residue, washing the filter residue with 2.4% diluted hydrochloric acid to remove impurities for 6 times, and calcining at 960 ℃ to obtain titanium dioxide with the purity of 99.12%; dropwise adding the ammonia water collected in the step (1) into the filtrate to adjust the pH value to be 5.6, generating hydrolysate, filtering and collecting the filtrate and filter residue, washing the filter residue with distilled water to remove impurities for 7 times, and calcining at 1280 ℃ for 2 hours to obtain an alumina product with the purity of 98.86%; dropwise adding the ammonia water collected in the step (1) into the filtrate to adjust the pH value to 12.6, generating hydrolysate, filtering and collecting filtrate and filter residue, washing the filter residue with distilled water to remove impurities for 6 times, and calcining at 710 ℃ for 2 hours to obtain a magnesium oxide product with the purity of 98.47%; and adding the residual ammonia water into the filtrate, and evaporating and crystallizing to obtain ammonium chloride crystals, thereby realizing the recycling of the ammonium chloride crystals.
Finally, it should be noted that the above embodiments are only used for illustrating the technical solutions of the present invention and are not limited. Although the present invention has been described in detail with reference to the embodiments, it will be understood by those skilled in the art that various changes may be made and equivalents may be substituted without departing from the spirit and scope of the invention as defined in the appended claims.
Claims (8)
1. A method for treating titanium minerals and metal silicate minerals by ammonium salt pressure pyrolysis comprises the following steps:
step one, drying and scattering titanium minerals and metal silicate minerals to obtain powder materials;
step two, uniformly mixing the powder material obtained in the step one with ammonium salt, roasting for 0.5-2 h at 250-450 ℃ in a roasting device, opening a sampling port of the roasting device at an initial stage, collecting generated ammonia gas, introducing water to form ammonia water, and sealing and pressurizing for pyrolysis when the pH value of the gas at the sampling port is acidic;
step three, adding 2-30% by mass of H into the roasting slag obtained by pyrolysis in the step two2SO4Or filtering after acid leaching of HCl solution;
and step four, preserving the heat of the filtrate obtained in the step three for 0.5-1 h in a boiling state, adding water with the volume fraction of 0.4-2 times that of the filtrate for dilution, preserving the heat for 0.5-2 h in the boiling state, then performing recrystallization to obtain offwhite hydrolysate, filtering and collecting filtrate and filter residues, washing the filter residues with dilute acid to remove impurities, and calcining to obtain the titanium dioxide.
2. The method of claim 1, further comprising:
step five, dropwise adding the filtrate obtained in the step four into the ammonia water formed in the step two until the pH is = 5-6, generating hydrolysate, filtering and collecting the filtrate and filter residues, washing the filter residues with water to remove impurities, and calcining to obtain an aluminum oxide product;
step six, dropwise adding the filtrate obtained in the step five into the ammonia water formed in the step two until the pH value is more than or equal to 12, generating hydrolysate, filtering and collecting the filtrate and filter residue, washing the filter residue with water to remove impurities, and calcining to obtain a magnesium oxide product;
and step seven, adding the ammonia water formed in the step two into the filtrate obtained in the step six, and evaporating and crystallizing to obtain ammonium salt crystals.
3. The method according to claim 2, characterized in that in the fifth step, the filter residue is calcined at 1000-1300 ℃ for 1-5 hours after being washed and impurity-removed for 4-8 times;
and sixthly, washing the filter residue to remove impurities for 3-6 times, and calcining for 2-4 hours at 400-800 ℃.
4. The method according to any one of claims 1 to 3, wherein in step one, the titanium minerals and the metal silicate minerals are ball-milled to-0.074 mm or more and 75% or more, dried and broken up at 50 to 110 ℃ to obtain a gray powder material with a moisture content of 2wt% or less.
5. The method according to any one of claims 1 to 3, wherein the judgment standard for controlling the opening and closing of the sampling port valve in the second step is that a pH meter detects that the pH of the gas at the sampling port is less than or equal to 7, so that valuable elements in the minerals are fully salted to generate soluble salts.
6. The method according to any one of claims 1 to 3, wherein in the second step, the powder material and ammonium salt are uniformly mixed according to a ratio of 1: 0.5-1: 5, wherein the ammonium salt is ammonium sulfate and/or ammonium chloride.
7. The method according to any one of claims 1 to 3, wherein in the third step, 2 to 30 mass percent of H is added to the roasting slag according to the liquid-solid ratio of 10:1 to 2:12SO4Or HCl solution, and acid leaching for 0.5-4 h at 40-90 ℃ and a stirring speed of 100-400 rpm.
8. The method according to any one of claims 1 to 3, wherein in the fourth step, the filter residue is washed by dilute acid with the concentration of 1-5% for 6-10 times to remove impurities, and is calcined at 850-1000 ℃ to obtain the titanium dioxide.
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Citations (7)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN101525143A (en) * | 2009-01-09 | 2009-09-09 | 东北大学 | Method for preparing magnesia, silicon dioxide and nickel oxide products from lateritic nickel ore |
| CN102424431A (en) * | 2011-09-01 | 2012-04-25 | 东北大学 | Method for preparing nickel oxide, magnesium oxide and silicon oxide products from laterite-nickel ore |
| CN104803400A (en) * | 2014-01-24 | 2015-07-29 | 中国科学院过程工程研究所 | Ammonium magnesium sulfate preparation method |
| CN104909396A (en) * | 2014-03-13 | 2015-09-16 | 中国科学院过程工程研究所 | Deep decomposition method of aluminum ammonium sulfate |
| CN105293576A (en) * | 2015-11-27 | 2016-02-03 | 四川大学 | Method for preparing high-grade artificial rutile from high-calcium and high-magnesium rock-mineral ilmenite |
| CN107267766A (en) * | 2017-08-28 | 2017-10-20 | 四川大学 | A kind of method of Selective Separation vanadium, titanium, iron from modified vanadium slag |
| CN107381585A (en) * | 2017-07-17 | 2017-11-24 | 北京矿冶研究总院 | Method for extracting aluminum silicon by pyrolyzing fly ash ammonium sulfate roasted clinker |
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Patent Citations (7)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN101525143A (en) * | 2009-01-09 | 2009-09-09 | 东北大学 | Method for preparing magnesia, silicon dioxide and nickel oxide products from lateritic nickel ore |
| CN102424431A (en) * | 2011-09-01 | 2012-04-25 | 东北大学 | Method for preparing nickel oxide, magnesium oxide and silicon oxide products from laterite-nickel ore |
| CN104803400A (en) * | 2014-01-24 | 2015-07-29 | 中国科学院过程工程研究所 | Ammonium magnesium sulfate preparation method |
| CN104909396A (en) * | 2014-03-13 | 2015-09-16 | 中国科学院过程工程研究所 | Deep decomposition method of aluminum ammonium sulfate |
| CN105293576A (en) * | 2015-11-27 | 2016-02-03 | 四川大学 | Method for preparing high-grade artificial rutile from high-calcium and high-magnesium rock-mineral ilmenite |
| CN107381585A (en) * | 2017-07-17 | 2017-11-24 | 北京矿冶研究总院 | Method for extracting aluminum silicon by pyrolyzing fly ash ammonium sulfate roasted clinker |
| CN107267766A (en) * | 2017-08-28 | 2017-10-20 | 四川大学 | A kind of method of Selective Separation vanadium, titanium, iron from modified vanadium slag |
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