Method for extracting aluminum oxide from gasification furnace slag

Abstract

The invention discloses a method for extracting aluminum oxide from gasification furnace slag, and belongs to the technical field of ore metallurgy. The method comprises the following steps: mixing potassium carbonate, dolomite and calcium fluoride to obtain a sintering agent; heating and activating the sintering agent to obtain a pre-activated sintering agent; mixing and sintering the pre-activated sintering agent and gasification slag to obtain sintered clinker; the sintered clinker is subjected to alkaline leaching, and alkaline leaching liquid and alkaline leaching residues are obtained; carrying out acid leaching on the alkaline leaching residues, adjusting the pH value, and removing impurities to obtain an acid leaching solution; mixing the acid leaching solution and the alkali leaching solution, adjusting the pH value, and precipitating to obtain crude aluminum hydroxide; dissolving the crude aluminum hydroxide with alkali to obtain a tetrahydroxy sodium aluminate solution; introducing carbon dioxide gas into the tetrahydroxy sodium aluminate solution, and precipitating to obtain aluminum hydroxide; and finally calcining the aluminum hydroxide at high temperature to obtain aluminum oxide. According to the method, the total recovery rate of aluminum is increased, impurities such as silicon, iron and magnesium can be effectively separated, and resource utilization of the gasification furnace slag is achieved.

Description

TECHNICAL FIELD

[0001] The application belongs to the technical field of ore metallurgy, and particularly relates to a method for extracting aluminum oxide from gasification furnace slag. BACKGROUND

[0002] Gasification furnace slag is a common byproduct of coal chemical industry, and the annual discharge amount in China is nearly 100 million tons. Long-term accumulation of gasification furnace slag not only occupies land resources, but also easily causes soil and water pollution, becoming an urgent environmental and resource management problem. As a core raw material in the fields of aluminum industry, ceramics, chemical industry, etc., the annual demand for aluminum oxide in China exceeds 80 million tons, but high-quality bauxite resources are increasingly scarce, and the contradiction between supply and demand of resources is prominent. The content of aluminum oxide in gasification furnace slag can reach 20% to 40%, which has extremely high utilization value. Extracting aluminum oxide from the slag can not only realize the resource utilization of solid waste and relieve environmental pressure, but also supplement the supply of aluminum oxide raw materials, reduce the dependence on high-quality bauxite in the industry, and has significant environmental and economic values. It is of great significance to guarantee the sustainable development of the aluminum industry and improve the comprehensive utilization level of gasification furnace slag. In recent years, the efficient utilization of gasification furnace slag has become the focus of attention in the scientific research and industrial circles, and a series of new technologies and new processes have emerged. However, these emerging processes still face many challenges in practical application. Traditional alkali lime sintering method: This method uses sodium carbonate and limestone as sintering agents to convert aluminum oxide into sodium aluminate and silicon into calcium silicate to realize separation at high temperature. Although this method is mature, the reaction activity of sodium carbonate and aluminum minerals is low, and the reaction can only be promoted at a high temperature of 1350 to 1400 DEG C. High temperature not only increases energy consumption, but also easily leads to excessive sintering of the furnace charge, thereby reducing the subsequent leaching efficiency. At the same time, the calcium silicate generated by the reaction of limestone and silicon easily wraps the unreacted aluminum minerals, resulting in low aluminum leaching rate and incomplete separation of silicon and aluminum. Metal sintering method: Alkali metal salts such as potassium carbonate and sodium carbonate are used as activators to generate oxides by high-temperature decomposition, which react with aluminum oxide to generate soluble aluminate. Although the metal sintering method can reduce the sintering temperature, the cost of alkali metal salts is high, and the alkali metal salts are easily volatilized at high temperature, resulting in high consumption of alkali metal salts and low recovery rate, which is difficult to be applied on a large scale. Acid or alkali direct leaching: This method uses inorganic strong acid or strong base to convert aluminum oxide in the raw material into soluble aluminum salt or sodium aluminate solution. Since 20% to 30% of aluminum in the gasification furnace slag exists in the form of mullite with stable crystal structure, the crystal lattice is difficult to be destroyed under normal pressure conditions to form effective dissolution, resulting in that the aluminum leaching rate is generally lower than 60%. In addition, separate alkali leaching cannot recover the aluminum in the magnesium aluminum spinel phase in the gasification furnace slag, causing waste of aluminum resources. Acid leaching will cause the synchronous dissolution of silicon, iron and other impurities, which increases the cost and easily leads to aluminum loss. Based on the above technical shortcomings, it is of great value to design an extraction method that can reduce energy consumption, reduce impurities and improve yield.

[0003] Patent CN104609451A discloses a preparation process of dry sintering aluminum trioxide, comprising the following steps: (1) grinding step: adding limestone, sodium carbonate and nitrogen-containing compound into bauxite, mixing according to a certain proportion, and grinding into raw meal powder; grinding the mixed raw meal with a certain fineness according to the proportion; (2) the raw meal obtained in step 1 is transported to the preheater located at the tail of the rotary kiln, wherein the preheater comprises two cyclones and a decomposition furnace, the two cyclones are communicated, and the decomposition furnace is communicated with one of the cyclones; (3) the auxiliary feeding port is arranged at the discharge port of one of the cyclones, and the reducing agent coke powder is added through the auxiliary feeding port; the raw meal preheated by the preheater and the coke powder are calcined in the rotary kiln; (4) the calcined product of step 3 is cooled to obtain aluminum trioxide clinker. The cooled aluminum trioxide clinker is dissolved by alkali solution, and finally the self-pulverization rate of the aluminum trioxide clinker is more than 92%, and the clinker dissolution rate is more than 90%. Although the technology realizes the extraction of aluminum trioxide by sintering and acid leaching, it is essentially a traditional lime sintering method designed for bauxite raw materials. The technology has the following limitations: first, the method is only suitable for bauxite, and has poor adaptability to gasification furnace slag with high silicon-aluminum ratio and containing a large amount of mullite and other stable crystal phases, and the aluminum conversion rate is low; second, the purification step is simple, only iron is removed by acid leaching, and there is no deep purification step for other impurities such as magnesium and calcium, so the removal effect is not good, and it is difficult to obtain high-purity products. Therefore, the existing technology cannot achieve the goal of efficiently and highly purifying aluminum from gasification furnace slag. SUMMARY

[0004] According to the deficiencies of the prior art, the present application optimizes the proportioning of sintering agent and pre-activates it to obtain a pre-activated sintering agent, then mixes the pre-activated sintering agent with gasification furnace slag and sintering, the obtained sintered clinker is subjected to a step-by-step leaching and purification process, i.e. alkali leaching, acid leaching, pH adjustment precipitation and sodium hydroxide dissolution purification to obtain aluminum hydroxide, and finally the aluminum hydroxide is calcined at high temperature to obtain aluminum trioxide. By optimizing the proportioning of sintering agent and enriching the purification steps, the technical defects of low aluminum leaching rate and incomplete impurity separation in the background technology are effectively overcome.

[0005] Specifically, the technical scheme of the present application includes the following contents: A method for extracting aluminum trioxide from gasification furnace slag, comprising the following steps: Mixing potassium carbonate, dolomite and calcium fluoride in a mass ratio of 6:4~6:0.5, ball milling and drying to obtain a sintering agent; Grinding the sintering agent after heating to obtain a pre-activated sintering agent; Mixing the pre-activated sintering agent with the gasification furnace slag in a mass ratio of 1.5:1 to obtain a sintered clinker; The sintered clinker is mixed with deionized water at a mass ratio of 8:1, then leached, and filtered to obtain an alkali leaching solution and an alkali leaching residue; The alkali leaching residue is mixed with hydrochloric acid at a mass ratio of 15:1, then leached and filtered, the pH of the filtrate is adjusted to 11.2, and then filtered again, and the filtrate is collected to obtain an acid leaching solution; The alkali leaching solution is mixed with the acid leaching solution, the pH is adjusted, and then filtered to obtain crude aluminum hydroxide; The crude aluminum hydroxide is mixed with an aqueous sodium hydroxide solution at a mass ratio of 10:1, then leached, and filtered to obtain a sodium tetrahydroxyaluminate solution; The sodium tetrahydroxyaluminate solution is heated and stirred, and carbon dioxide gas is introduced until the pH is 10.5, then filtered, and the filter cake is dried at 105°C for 12 hours to obtain aluminum hydroxide; The aluminum hydroxide is calcined at high temperature to obtain aluminum oxide.

[0006] Further, the process of mixing the potassium carbonate, dolomite and calcium fluoride, and then ball milling and drying includes first ball milling at a speed of 400 r / min for 2 hours, with the mass ratio of zirconium oxide grinding balls to material being 10:1, and then vacuum drying at 60°C for 2 hours.

[0007] Further, the process of grinding the sintering agent after heating includes first heating at a temperature increasing rate of 5°C / min to 900°C, maintaining the temperature for 1.5 hours, and then taking out and grinding to pass through a 100-mesh sieve after natural cooling.

[0008] Further, the process of sintering the pre-activated sintering agent mixed with the gasification slag includes first heating at a temperature increasing rate of 5°C / min to 1000-1100°C, sintering for 2 hours, and then taking out and grinding to pass through a 100-mesh sieve after natural cooling.

[0009] Further, the process of leaching and filtering the sintered clinker mixed with deionized water includes first stirring leaching at 90°C at a rotation speed of 200 r / min for 60 minutes, and then vacuum suction filtration and washing the filter residue with 50 mL of deionized water preheated to 80°C in three times, retaining the filter residue to obtain an alkali leaching residue, and collecting the filtrate to obtain an alkali leaching solution.

[0010] Further, the process of leaching and filtering the alkali leaching residue mixed with hydrochloric acid, and then filtering again after adjusting the pH includes first stirring leaching at 80°C at a rotation speed of 200 r / min for 90 minutes, and then vacuum suction filtration and washing the filter residue with 20 mL of 0.5 mol / L hydrochloric acid aqueous solution preheated to 80°C in two times; the filtrate and the washing liquid are collected and heated to 80°C, 5 mol / L sodium hydroxide aqueous solution is added under the condition of stirring at 200 r / min, the pH of the solution is adjusted to 11.2, and then the temperature and rotation speed are kept unchanged for 30 minutes. After stirring, vacuum suction filtration is immediately performed, and the precipitate is washed with 30 mL of deionized water preheated to 60°C in two times, and the washing liquid is combined with the filtrate.

[0011] Further, the process of mixing the alkali leaching solution with the acid leaching solution, adjusting the pH and standing for filtration comprises: first heating to 80℃, continuously stirring at a rotating speed of 150 r / min, and monitoring the pH of the solution in real time; pumping the acid leaching solution into the alkali leaching solution at a set segmented flow rate using a constant flow pump: when the pH is greater than 10.0, the flow rate is controlled at 25 mL / min; when the pH drops to 10.0, the liquid addition is paused, and the stirring is continued for 10 minutes; when 10.0 > pH > 7.0, the flow rate is controlled at 15 mL / min; when the pH reaches 7.0, the liquid addition is paused again, and the stirring is continued for 10 minutes; when 7.0 > pH > 6.0, the flow rate is controlled at 10 mL / min. If the pH of the mixed solution is still higher than 6.0 after the acid leaching solution is completely pumped in, 1 mol / L hydrochloric acid solution is added dropwise at a flow rate of 10 mL / min until the pH of the solution stabilizes at 6.0. After the pH stabilizes, the stirring speed is reduced to 100 r / min, and the stirring is continued at 80℃ for 60 minutes. Then, the stirring is stopped, the solution is allowed to stand at 80℃ for 30 minutes, vacuum filtration is performed, and the filter cake is washed with 50 mL of deionized water preheated to 60℃ for 3 times.

[0012] Further, the process of mixing the crude aluminum hydroxide with the sodium hydroxide aqueous solution and then leaching and filtering comprises: first stirring at a rotating speed of 100 r / min at 80℃ for 1 hour, and then performing vacuum filtration to collect the filtrate.

[0013] Further, the process of filtering after heating and stirring the sodium tetrahydroxyaluminate solution and passing carbon dioxide gas into the solution comprises: first heating to 70℃ and stirring at a rotating speed of 100 r / min, while passing carbon dioxide gas into the solution at a ventilation rate of 0.5 L / min using a mass flow controller, until the pH of the solution drops to 10.5, then stopping the ventilation, allowing the solution to stand at 70℃ for 30 minutes, performing vacuum filtration, and washing the precipitate with 2000 mL of deionized water preheated to 70℃ until the pH of the washing liquid is 7.0.

[0014] Further, the conditions for high-temperature calcination of aluminum hydroxide include a calcination temperature of 600℃ and a calcination time of 2 hours.

[0015] Compared with the prior art, the present application has the following advantages: (1) The present application uses potassium carbonate, dolomite and calcium fluoride as sintering agents for the difficult-to-dissolve aluminum-containing phases such as mullite and magnesium-aluminum spinel in the gasification furnace slag. Potassium carbonate, as the core activator, reacts with aluminum minerals at high temperatures to form soluble aluminate; the calcium and magnesium elements provided by dolomite can combine with silicon to form stable silicates, thereby achieving silicon-aluminum separation; calcium fluoride acts as a fluxing agent, effectively reducing the sintering temperature and further promoting silicon-aluminum separation. The synergistic effect of the three optimizes the sintering effect, realizes efficient conversion of difficult-to-dissolve aluminum, and has wide adaptability to raw materials.

[0016] (2) The present application preheats and activates the sintering agent to decompose dolomite into high-activity calcium oxide and magnesium oxide, thereby reducing the sintering temperature and improving the reactivity of the sintering agent, so that the sintering agent fully reacts with the slag and the energy consumption of the reaction is reduced.

[0017] (3) The sintering temperature of 1000-1100 DEG C is used to ensure that the reaction is completely performed while avoiding over-sintering of the materials and volatilization of potassium carbonate, and the process cost is low.

[0018] (4) The synergistic extraction process of the multi-step purification process of "alkali leaching-acid leaching-merging and precipitation" extracts aluminum in different chemical phases in the gasification slag, thereby improving the total recovery rate of aluminum and solving the problem of low recovery rate of aluminum by direct leaching by acid or alkali method. At the same time, through multi-step pH precise control, the main impurities such as silicon, iron and magnesium are effectively separated.

[0019] (5) Compared with the prior art, the present application realizes the efficient extraction of aluminum in the gasification slag at a temperature of 1000-1100 DEG C through the synergistic effect of pre-activated sintering agent and step-by-step acid and alkali leaching, the total recovery rate of aluminum is not less than 86%, and the purity of aluminum oxide product is not less than 97%, which provides a feasible technical path for the resource utilization of gasification slag. DETAILED DESCRIPTION

[0020] The technical solutions of the present application will be described clearly and completely by the embodiments of the present application. Obviously, the described embodiments are only a part of the embodiments of the present application, not all the embodiments. Based on the embodiments in the present application, all other embodiments obtained by those skilled in the art without creative labor belong to the scope of protection of the present application.

[0021] Unless otherwise specified, the raw materials and reagents used in the present application are commercially available or can be prepared by known methods.

[0022] Preparation Example 1: The preparation of the sintering agent includes the following process: Potassium carbonate 60g, dolomite 40g and calcium fluoride 5g were weighed according to the mass ratio of 6:4:0.5, put into a planetary ball mill, and ball milled at a speed of 400r / min for 2 hours, and the mass ratio of zirconia grinding ball to material was 10:1. After ball milling, the mixture was transferred to a ceramic evaporation dish and placed in a vacuum drying box at 60 DEG C and-0.1MPa vacuum degree for 2 hours to obtain the sintering agent.

[0023] Preparation Example 2: The preparation of the sintering agent includes the following process: Potassium carbonate 60 g, dolomite 50 g, calcium fluoride 5 g were weighed according to the mass ratio of 6:5:0.5, and were placed in a planetary ball mill, and were ball milled at a speed of 400 r / min for 2 hours, and the mass ratio of zirconia grinding balls to materials was 10:1. After ball milling, the mixture was transferred to a ceramic evaporating dish, and was placed in a vacuum drying oven, and was dried at 60°C under a vacuum degree of-0.1 MPa for 2 hours to obtain a sintering agent.

[0024] Preparation Example 3: Preparation of the sintering agent, including the following flow: Potassium carbonate 60 g, dolomite 60 g, calcium fluoride 5 g were weighed according to the mass ratio of 6:6:0.5, and were placed in a planetary ball mill, and were ball milled at a speed of 400 r / min for 2 hours, and the mass ratio of zirconia grinding balls to materials was 10:1. After ball milling, the mixture was transferred to a ceramic evaporating dish, and was placed in a vacuum drying oven, and was dried at 60°C under a vacuum degree of-0.1 MPa for 2 hours to obtain a sintering agent.

[0025] Preparation Example 4: Preparation of the sintering agent, including the following flow: Potassium carbonate 60 g, dolomite 60 g were weighed according to the mass ratio of 1:1, and no calcium fluoride was added, and were placed in a planetary ball mill, and were ball milled at a speed of 400 r / min for 2 hours, and the mass ratio of zirconia grinding balls to materials was 10:1. After ball milling, the mixture was transferred to a ceramic evaporating dish, and was placed in a vacuum drying oven, and was dried at 60°C under a vacuum degree of-0.1 MPa for 2 hours to obtain a sintering agent.

[0026] Example 1: A method for extracting aluminum oxide from gasification furnace slag, including the following steps: The sintering agent of Preparation Example 1 was placed in a corundum crucible, and was placed in a box-type resistance furnace, and was heated to 900°C at a heating rate of 5°C / min, and the temperature was maintained for 1.5 hours, and after natural cooling, it was ground through a 100-mesh screen to obtain a pre-activated sintering agent; The pre-activated sintering agent was mixed with the gasification furnace slag ground through a 100-mesh screen according to a mass ratio of 1.5:1, and was placed in a corundum crucible, and was placed in a box-type resistance furnace, and was heated to 1000°C at a heating rate of 5°C / min, and the temperature was maintained for 2 hours, and after natural cooling, it was taken out and ground through a 100-mesh screen to obtain a sintered clinker; The sintered clinker was mixed with deionized water according to a mass ratio of 8:1, and was placed in a beaker, and the beaker was placed on a heat collecting type constant temperature heating magnetic stirrer, and was stirred at 90°C at a speed of 200 r / min for 60 minutes, and then a vacuum pump was connected to a Buchner funnel using a 0.45 μm PTFE membrane as a filter medium, and the filter residue was washed with 50 mL of deionized water preheated to 80°C for 3 times, and the filter residue was retained to obtain an alkali leaching residue, and the filtrate was collected to obtain an alkali leaching solution; The alkali leaching residue was mixed with 5 mol / L hydrochloric acid aqueous solution at a mass ratio of 15:1, and then placed in a three-necked flask. The three-necked flask was placed in a constant temperature water bath with a mechanical stirrer, and stirred at 200 r / min for 90 minutes at 80°C. Then, a Buchner funnel was connected with a circulating water vacuum pump, and a 0.45 μm PTFE membrane was used as the filter medium for vacuum filtration. The residue was washed with 20 mL of 0.5 mol / L hydrochloric acid aqueous solution preheated to 80°C for 2 times, and then the filtrate and washing liquid were collected and transferred to a beaker. The beaker was placed on a heat-type constant temperature heating magnetic stirrer, heated to 80°C, and stirred at 200 r / min. The pH was monitored by a pH meter, and 5 mol / L sodium hydroxide aqueous solution was added to adjust the pH of the solution to 11.2. The temperature and stirring speed were kept unchanged for 30 minutes, and then the vacuum filtration was performed using a Buchner funnel connected with a circulating water vacuum pump, and a 0.45 μm PTFE membrane was used as the filter medium. The precipitate was washed with 30 mL of deionized water preheated to 60°C for 2 times, and the washing liquid was combined with the filtrate. The combined filtrate was collected to obtain the acid leaching liquid; The alkali leaching liquid was transferred to a three-necked flask, which was placed on a heat-type constant temperature heating magnetic stirrer and heated to 80°C with continuous stirring at 150 r / min. The acid leaching liquid was pumped into the alkali leaching liquid at a set segmented flow rate using a constant flow pump under the monitoring of a pH meter: the flow rate was controlled at 25 mL / min when the pH was greater than 10.0; when the pH dropped to 10.0, the liquid addition was paused, and stirring was continued for 10 minutes; the flow rate was controlled at 15 mL / min when the pH was between 10.0 and 7.0; when the pH reached 7.0, the liquid addition was paused again, and stirring was continued for 10 minutes; the flow rate was controlled at 10 mL / min when the pH was between 7.0 and 6.0. If the pH of the mixed solution was still higher than 6.0 after the acid leaching liquid was completely pumped in, 1 mol / L hydrochloric acid aqueous solution was added dropwise at a flow rate of 10 mL / min until the pH of the solution stabilized at 6.0. After the pH stabilized, the stirring speed was reduced to 100 r / min, and the temperature and stirring speed were maintained for 60 minutes. Then, the stirring was stopped, and the mixture was left to stand at 80°C for 30 minutes. Then, vacuum filtration was performed, and the filter cake was washed with 50 mL of deionized water preheated to 60°C for 3 times. The filter cake was collected to obtain the crude aluminum hydroxide; The crude aluminum hydroxide was mixed with 4 mol / L sodium hydroxide aqueous solution at a mass ratio of 10:1, and then placed in a three-necked flask. The three-necked flask was placed on a heat-type constant temperature heating magnetic stirrer, and stirred at 100 r / min for 1 hour at 80°C. Then, a Buchner funnel was connected with a circulating water vacuum pump, and a 0.45 μm PTFE membrane was used as the filter medium for vacuum filtration. The filtrate was collected to obtain a sodium tetrahydroxyaluminate solution; The sodium tetrahydroxyaluminate solution was transferred to a three-necked flask, which was placed on a heat collecting constant temperature heating magnetic stirrer, heated to 70℃, kept stirring at 100 r / min, and pure carbon dioxide gas with a purity of ≥99.9% was introduced into the solution at a rate of 0.5 L / min using a mass flow controller under the monitoring of a pH meter until the pH value of the solution decreased to 10.5, then the solution was kept at 70℃ for 30 minutes, and vacuum filtration was performed using a 0.45 μm PTFE membrane as the filter medium, the precipitate was washed with 2000 mL of deionized water preheated to 70℃ until the pH of the washing liquid was 7.0, and the obtained filter cake was transferred to a petri dish and dried in an electric heating air oven at 105℃ for 12 hours to obtain aluminum hydroxide. The aluminum hydroxide was loaded into a corundum crucible and placed in a box-type resistance furnace, which was heated to 600℃ at a rate of 5℃ / min, and the temperature was maintained for 2 hours, and then the aluminum hydroxide was naturally cooled to obtain aluminum oxide.

[0027] Example 2: The sintering agent of Preparation Example 1 in Example 1 was replaced by the sintering agent of Preparation Example 2, and the other conditions were kept the same as in Example 1.

[0028] Example 3: The sintering agent of Preparation Example 1 in Example 1 was replaced by the sintering agent of Preparation Example 3, and the other conditions were kept the same as in Example 1.

[0029] Example 4: The sintering temperature of the pre-activated sintering agent and the gasified slag in Example 1 was replaced by 1100℃ instead of 1000℃, and the other conditions were kept the same as in Example 1.

[0030] Example 5: The sintering temperature of the pre-activated sintering agent and the gasified slag in Example 2 was replaced by 1100℃ instead of 1000℃, and the other conditions were kept the same as in Example 2.

[0031] Example 6: The sintering temperature of the pre-activated sintering agent and the gasified slag in Example 3 was replaced by 1100℃ instead of 1000℃, and the other conditions were kept the same as in Example 3.

[0032] Comparative Example 1: The sintering agent of Preparation Example 1 in Example 1 was replaced by the sintering agent of Preparation Example 4, and the other conditions were kept the same as in Example 1.

[0033] Comparative Example 2: The pre-activated sintering agent and the gasified slag were replaced by the un-pre-activated sintering agent and the gasified slag in Example 1, and the other conditions were kept the same as in Example 1.

[0034] Comparative Example 3: The heating of the sintering agent in Example 1 to 900℃ is replaced by heating to 800℃, and the rest of the conditions remain the same as in Example 1.

[0035] Comparative Example 4: The pre-activated sintering agent and the gasified slag in Example 1 are replaced by 1:1, and the rest of the conditions remain the same as in Example 1.

[0036] Comparative Example 5: The mixing of the sintered clinker with deionized water in Example 1 is replaced by stirring and leaching at 80℃ in a beaker, and the rest of the conditions remain the same as in Example 1.

[0037] Comparative Example 6: The mixing of the sintered clinker with deionized water in Example 1 is replaced by stirring and leaching, and the rest of the conditions remain the same as in Example 1.

[0038] Comparative Example 7: The mixing of the sintered clinker with deionized water in Example 1 is replaced by stirring and leaching, and the rest of the conditions remain the same as in Example 1.

[0039] Comparative Example 8: The mixing of the sintered clinker with deionized water in Example 1 is replaced by stirring and leaching, and the rest of the conditions remain the same as in Example 1.

[0040] Comparative Example 9: The mixing of the sintered clinker with deionized water in Example 1 is replaced by stirring and leaching, and the rest of the conditions remain the same as in Example 1.

[0041] The main content and key impurity content of the aluminum trioxide products obtained in Examples 1-6 and Comparative Examples 1-9 were determined by Inductively Coupled Plasma-Optical Emission Spectrometer (ICP-OES). At the same time, the total recovery rate of aluminum was calculated by calculating the mass balance of aluminum elements in the entire process. The specific method is as follows: the mass of the raw material gasified slag and the final aluminum trioxide product is accurately weighed, and the purity of the aluminum trioxide is determined by ICP-OES, and the total recovery rate is calculated according to the formula "recovery rate (%) = (product mass x product aluminum trioxide purity) / (raw material mass x aluminum trioxide content in raw material) x 100%". The test results are as follows: Table 1 Product purity and impurity content test table Sample origin Al2O3purity (mass %) Main impurity content (mass fraction / %) Total aluminum recovery rate (%) SiO2Fe2O3MgOCaO Example 1 97.3 0.3 10.1 20.0 0.15 86.2 Example 2 97.2 0.3 20.3 10.1 0.15 88.5 Example 3 97.4 0.2 5.0 14.0 0.09 89.5 Example 4 97.2 0.3 5.0 30.3 0.07 88.2 Example 5 97.2 0.3 20.3 15.0 0.10 87.4 Example 6 97.3 0.3 7.3 10.2 0.17 89.3 Comparative Example 1 93.4 13.1 20.4 50.2 0.52 75.6 Comparative Example 2 94.5 8.1 10.5 95.0 0.15 72.3 Comparative Example 3 94.2 3.1 52.3 50.8 0.40 74.8 Comparative Example 4 93.8 3.4 0.5 20.2 0.45 74.2 Comparative Example 5 94.5 2.9 5.0 30.0 0.38 78.1 Comparative Example 6 94.8 1.8 0.3 8.3 0.55 68.9 Comparative Example 7 93.6 2.6 1.5 50.2 0.50 73.5 Comparative Example 8 90.1 1.7 5.0 45.0 0.42 70.0 Comparative Example 9 90.9 1.6 21.4 0.3 0.48 68.3 From the above detection data, it can be seen that: (1) The total aluminum recovery rate of Examples 1-6 is stable at 86.2%-89.5%, and the purity of aluminum oxide is higher than 97%, and the impurity content is significantly lower than that of each comparative example, indicating that through the synergistic process of pre-activated sintering agent and stepwise acid and alkali leaching purification, the impurity elements such as silicon, iron, magnesium and calcium can be effectively separated. This proves that the method of the present application has stable and excellent effect in the range of the set process parameters.

[0042] (2) By comparing Examples 1-6, it can be seen that the purity of aluminum oxide of the example with higher quality dolomite is higher than that of the example with lower quality dolomite, and the silicon oxide impurity content is also lower, indicating that higher proportion of dolomite can optimize the impurity removal effect. There is no obvious difference in purity, impurity and recovery rate between Comparative Examples 1, 2, 3 and Comparative Examples 4, 5, 6, indicating that the sintering temperature of the reaction system is 1000-1100℃, and it is feasible to select the sintering temperature of 1000℃.

[0043] (3) It can be seen from Comparative Example 1 that using sintering agent without calcium fluoride will cause the purity and recovery rate of aluminum to decrease, and the content of impurity silicon oxide to increase significantly, proving that calcium fluoride as a flux is crucial for promoting sintering reaction and realizing effective silicon-aluminum separation.

[0044] (4) It can be seen from Comparative Example 2 that using the sintering agent without heating activation will reduce the purity and recovery rate of aluminum, and the contents of impurities such as silicon oxide and iron oxide will significantly increase, indicating that the pre-activation step can improve the reactivity of the sintering agent, and help to more completely separate the impurities such as iron in the subsequent leaching.

[0045] (5) It can be seen from Comparative Example 3 that reducing the heating activation temperature of the sintering agent from 900°C to 800°C will reduce the purity and recovery rate of aluminum, and the contents of impurities such as silicon oxide and iron oxide will significantly increase, indicating that insufficient pre-activation temperature will reduce the reactivity of the sintering agent, and the pre-activation at 900°C can ensure the high reactivity of the sintering agent.

[0046] (6) It can be seen from Comparative Example 4 that changing the mass ratio of the sintering agent to the gasification furnace slag from 1.5:1 to 1:1 will reduce the purity and recovery rate of aluminum, and the contents of impurities, especially silicon oxide, will significantly increase. It indicates that insufficient sintering agent will lead to reduced aluminum conversion rate and incomplete separation of silicon and aluminum, and the mass ratio of 1.5:1 of the sintering agent to the gasification furnace slag is the optimal ratio for achieving sufficient reaction and efficient separation.

[0047] (7) It can be seen from Comparative Example 5 that reducing the alkali leaching temperature from 90°C to 80°C will reduce the purity and recovery rate of aluminum, and the content of impurity silicon oxide will significantly increase, indicating that reducing the alkali leaching temperature will lead to reduced leaching efficiency, and the condition of 90°C can improve the separation of silicon and aluminum and the alkali leaching efficiency.

[0048] (8) It can be seen from Comparative Example 6 that removing the acid leaching step will reduce the purity and recovery rate of aluminum, and the contents of impurities, especially silicon oxide and magnesium oxide, will significantly increase, proving that the acid leaching step is a key link in the process and is indispensable for recovering aluminum in the spinel phase of the alkali leaching residue and effectively removing magnesium impurities.

[0049] (9) It can be seen from Comparative Example 7 that changing the concentration of the hydrochloric acid aqueous solution used in the acid leaching step from 5 mol / L to 3 mol / L will reduce the purity and recovery rate of aluminum, and the contents of impurities such as silicon oxide and iron oxide will significantly increase, indicating that reducing the concentration of hydrochloric acid will lead to incomplete dissolution of the insoluble aluminum phase and poor impurity separation, and 5 mol / L hydrochloric acid can achieve efficient acid leaching.

[0050] (10) It can be seen from Comparative Example 8 that changing the stepwise flow rate to a fixed flow rate in the mixing step of the acid leaching solution and the alkali leaching solution will significantly reduce the purity and recovery rate of aluminum, and the contents of impurities such as silicon oxide, iron oxide, and magnesium oxide will significantly increase, indicating that not controlling the pH in stages will lead to co-precipitation of impurities and reduced product purity, and the stepwise flow rate control helps to improve the purity and recovery rate.

[0051] (11) It can be seen from Comparative Example 9 that changing the end point pH of carbon dioxide introduction from 10.5 to 9.5 will greatly reduce the purity and recovery of aluminum, and the contents of impurities silicon oxide and iron oxide will significantly increase, indicating that reducing the end point pH of carbon dioxide introduction will exceed the best precipitation stability interval of aluminum hydroxide and cause more impurities to coprecipitate or adsorb, and the end point pH of 10.5 can ensure complete precipitation of aluminum hydroxide and maintain good impurity removal performance.

[0052] The above-described embodiments illustrate the technical solutions and beneficial effects of the present application. It should be understood that the above-described embodiments are only specific embodiments of the present application and are not intended to limit the present application. Without departing from the spirit and scope of the present application, various changes and improvements can be made to the present application, and these changes and improvements all fall within the scope of the claimed present application.

Claims

1. A method of extracting aluminum trioxide from gasification furnace slag, characterized by, The method comprises the following steps: The sintering agent is obtained by mixing potassium carbonate, dolomite and calcium fluoride in a mass ratio of 6:4~6:0.5, ball milling and drying; The pre-activated sintering agent is obtained by grinding the sintering agent after heating; The sintered clinker is obtained by sintering the pre-activated sintering agent mixed with the gasification furnace slag in a mass ratio of 1.5:1; The sintered clinker is mixed with deionized water in a mass ratio of 8:1, and then leaching and filtering are performed to obtain an alkali leaching solution and an alkali leaching residue; The alkali leaching residue is mixed with hydrochloric acid in a mass ratio of 15:1, and then leaching and filtering are performed; the pH of the filtrate is adjusted to 11.2, and then the filtrate is collected again to obtain an acid leaching solution; The alkali leaching solution and the acid leaching solution are mixed, the pH is adjusted, and then standing and filtering are performed to obtain crude aluminum hydroxide; The crude aluminum hydroxide is mixed with a sodium hydroxide aqueous solution in a mass ratio of 10:1, and then leaching and filtering are performed to obtain a sodium tetrahydroxyaluminate solution; The sodium tetrahydroxyaluminate solution is heated and stirred, and carbon dioxide gas is introduced until the pH is 10.5; then filtering is performed, and the filter cake is dried at 105℃ for 12 hours to obtain aluminum hydroxide; The aluminum hydroxide is calcined at high temperature to obtain aluminum oxide.

2. A process for extracting alumina trihydrate from gasification slag as claimed in claim 1 wherein, The process of ball milling and drying the mixture of potassium carbonate, dolomite and calcium fluoride comprises the following steps: first, ball milling at a speed of 400r / min for 2 hours, and then drying at 60℃ for 2 hours.

3. A process for extracting alumina trihydrate from gasification slag as claimed in claim 1 wherein, The process of grinding the sintering agent after heating comprises the following steps: first, heating to 900℃, then maintaining the temperature for 1.5 hours, and then naturally cooling and grinding to pass through a 100-mesh sieve.

4. A process for extracting alumina trihydrate from gasification furnace slag as claimed in claim 1 wherein, The process of sintering the pre-activated sintering agent mixed with the gasification furnace slag comprises the following steps: first, heating to 1000~1100℃, then sintering for 2 hours, and then naturally cooling and grinding to pass through a 100-mesh sieve.

5. A process for extracting alumina trihydrate from gasification slag as claimed in claim 1 wherein, The process of leaching and filtering the sintered clinker mixed with deionized water comprises the following steps: first, stirring and leaching at 90℃ and a rotating speed of 200r / min for 60 minutes, then vacuum suction filtration, retaining the filter residue to obtain the alkali leaching residue, and collecting the filtrate to obtain the alkali leaching solution.

6. A process for extracting alumina trihydrate from gasification slag as claimed in claim 1 wherein, The process of leaching and filtering the mixture of the alkali leaching residue and hydrochloric acid, and adjusting the pH of the filtrate again comprises the following steps: first, stirring and leaching at 80℃ and a rotating speed of 200r / min for 90 minutes, then vacuum suction filtration, collecting the filtrate, heating to 80℃, adding a 5mol / L sodium hydroxide aqueous solution under the conditions of 200r / min stirring, adjusting the pH of the solution to 11.2, then maintaining the temperature and rotating speed, and stirring for 30 minutes, and then vacuum suction filtration.

7. A process for extracting alumina trihydrate from gasification furnace slag as claimed in claim 1 wherein, The process of adjusting the pH, standing and filtering after mixing the alkali leaching solution and the acid leaching solution comprises the following steps: first, heating to 80℃, continuously stirring at a rotating speed of 150r / min, then pumping the acid leaching solution into the alkali leaching solution at a set segmented flow rate: controlling the flow rate to be 25mL / min when the pH>10.0; when the pH drops to 10.0, stopping adding the liquid, and continuing to stir for 10 minutes; controlling the flow rate to be 15mL / min when 10.0>pH>7.0; when the pH reaches 7.0, stopping adding the liquid again, and continuing to stir for 10 minutes; 7.0>pH>6.0 stage control flow rate 10 mL / min; if the pH of the mixed solution is still higher than 6.0 after the acid leaching solution is completely pumped in, 1 mol / L hydrochloric acid aqueous solution is added dropwise at a flow rate of 10 mL / min until the pH of the solution is stabilized at 6.0; after the pH is stabilized, the stirring speed is reduced to 100 r / min, and stirring is continued at 80℃ for 60 minutes, then the stirring is stopped, and the solution is statically placed at 80℃ for 30 minutes, and vacuum filtration is performed.

8. A process for extracting alumina trihydrate from gasification slag as claimed in claim 1 wherein, The process of mixing the crude aluminum hydroxide with the aqueous sodium hydroxide solution and then leaching and filtering includes first stirring at 80℃ at a rotation speed of 100 r / min for 1 hour, and then performing vacuum filtration.

9. A process for extracting alumina trihydrate from gasification slag as claimed in claim 1 wherein, The process of heating and stirring the sodium tetrahydroxyaluminate solution and then filtering after carbon dioxide gas is introduced includes first heating to 70℃ and stirring at a rotation speed of 100 r / min, while introducing carbon dioxide gas into the solution at a ventilation rate of 0.5 L / min, until the pH of the solution is reduced to 10.5, then stopping the ventilation, and then statically placing the solution at 70℃ for 30 minutes, and performing vacuum filtration, and washing the precipitate with deionized water until the pH of the washing solution is 7.

0.

10. A process for extracting alumina trihydrate from gasification slag as claimed in claim 1 wherein, The conditions of high-temperature calcination of the aluminum hydroxide include a calcination temperature of 600℃ and a calcination time of 2 hours.