Cooperative treatment method for secondary aluminum ash and overhaul slag

Abstract

The invention belongs to the technical field of aluminum industry solid hazardous waste resource utilization, and relates to a secondary aluminum ash and overhaul slag co-processing method which comprises the following steps: S1, crushing secondary aluminum ash, adding concentrated sulfuric acid, uniformly mixing, and roasting to obtain clinker; adding water into the clinker, leaching and filtering to obtain filtrate A and aluminum oxide; s2, crushing the overhaul slag, adding the filtrate A, leaching and filtering to obtain filtrate B and leaching slag; s3, adjusting the pH value of the filtrate B to 1-11, reacting, and filtering to obtain filtrate C and cryolite; and crystallizing the filtrate C to obtain sodium sulfate. The concentrated sulfuric acid mixed roasting process is adopted for the secondary aluminum ash, reaction of concentrated sulfuric acid and alpha-Al2O3 with MgAl2O4 is promoted under the high-temperature condition, aluminum in the secondary aluminum ash is converted into soluble aluminum sulfate, and therefore the leaching rate of aluminum is increased.

Description

Technical Field

[0001] The invention belongs to the technical field of solid hazardous waste resource utilization in the aluminum industry and relates to a method for coordinated treatment of secondary aluminum ash and overhaul slag. Background Art

[0002] The solid slag obtained by recycling and recasting primary aluminum ash or other aluminum waste, and recovering part of the aluminum by adding flux such as chloride and high-temperature melting is called secondary aluminum ash. In addition, secondary aluminum ash is also produced in the process of recycled aluminum and aluminum material processing, and its elemental aluminum content is only 5%-20%. The components of secondary aluminum ash are mainly elemental aluminum, aluminum oxide, aluminum nitride, aluminum carbide, fluoride, chloride, other metal elements and metal oxides. It is reactive and toxic and is listed in the "National List of Hazardous Wastes (2025 Edition)". After overhaul, aluminum electrolytic cells produce a large amount of overhaul slag, including waste cell linings, waste cathodes and cathode ash. The main chemical components of waste cell linings are aluminosilicates, and some fluorides exist in the form of NaF. The waste cathode is mainly highly graphitized carbon, and also contains Na3AlF6, NaF, CaF2, Al2O3 and other valuable metal components. The large amount of fluoride contained in secondary aluminum ash and overhaul slag poses a huge threat to environmental life and human health. Their harmless disposal and resource utilization face huge challenges and difficulties.

[0003] Prior art CN113278808B discloses a method for linking the recovery of multiple solid waste materials in the aluminum smelting process. First, the aluminum ash and cathode ash are crushed and mixed, and the elemental aluminum and high iron materials are selected by screening and magnetic separation, and then water leaching is performed, and then the water leaching residue is hot acid leached with sulfuric acid, and the carbon residue after decarbonization is leached using the obtained hot acid leaching slurry. After filtering, NaOH is added to the leaching solution for neutralization and precipitation to obtain hydroxy aluminum fluoride, and finally a mixture of aluminum fluoride and aluminum oxide is obtained after calcination. However, the XRD of the aluminum ash raw material used in the patent shows that the main form of aluminum is aluminum nitride, which has strong activity and good reactivity. However, aluminum in secondary aluminum ash usually exists mainly in the form of inert α-Al2O3 and MgAl2O4, which has a very stable structure, is insoluble in acid and alkali under normal conditions, and is difficult to leach. Therefore, it is not clear whether this method is applicable to aluminum ash of different sources and components; in addition, the process has many steps, a complex flow, and involves a variety of equipment and chemical reagents, strong technical dependence, and high cost.

[0004] Prior art CN113426807B discloses a method for combined treatment of hazardous waste slag from aluminum electrolysis and comprehensive resource utilization. This technology is based on "aluminum salt leaching + neutralization precipitation". The aluminum ash is first pressurized for leaching and then leached with sulfuric acid, and then the obtained aluminum sulfate salt solution is used to leach the carbon slag. After filtering, the filtrate is neutralized with the alkaline leaching solution obtained after pressurized alkaline leaching of overhaul slag to precipitate hydroxyaluminum fluoride, which is then calcined to obtain aluminum fluoride. However, this method will introduce impurity silicon in the raw material into the solution during the alkaline leaching process, thereby affecting the product quality; at the same time, this method involves pressurized leaching, which has high requirements for equipment, increases equipment investment and operating costs, and the recovery rate of valuable components in this process is not high.

[0005] In addition, regarding the recovery form of aluminum element, the target products of the above two prior arts are both aluminum fluoride, and only aluminum and fluorine are recovered, while the remaining valuable substances, such as sodium, are discarded, resulting in a waste of some valuable substances. Summary of the invention

[0006] The purpose of the present invention is to provide a method for the coordinated treatment of secondary aluminum ash and overhaul slag, which can effectively treat the inert aluminum oxide in the aluminum ash through roasting with concentrated sulfuric acid, significantly improve the recovery rate of aluminum element, and recover more valuable substances.

[0007] In order to achieve the above object, the present invention adopts the following technical solution:

[0008] A method for co-processing secondary aluminum ash and overhaul slag comprises the following steps:

[0009] S1. Grind the secondary aluminum ash, add concentrated sulfuric acid, mix well, and then roast to obtain clinker; add water to the clinker, leaching, filtering, and obtaining filtrate A and aluminum oxide;

[0010] S2, crushing the overhaul slag, adding filtrate A, leaching, filtering, and obtaining filtrate B and leached slag;

[0011] S3. Adjust the pH value of filtrate B to 1-11 for reaction, filter and obtain filtrate C and cryolite; crystallize filtrate C to obtain sodium sulfate.

[0012] The invention mixes crushed and sieved secondary aluminum ash with concentrated sulfuric acid, and roasts to obtain clinker; crushes the roasted clinker, and then leaches it with pure water, and filters to obtain a primary leaching solution, wherein the primary leaching residue is alpha-Al2O3; uses the primary leaching solution to leach overhaul residue, and filters to obtain a secondary leaching solution, wherein the secondary leaching residue is carbon, or nepheline and feldspar; adds alkali solution to the secondary leaching solution, adjusts the pH value of the solution, and obtains a cryolite product after filtering and drying; and crystallizes the filtered filtrate to obtain sodium sulfate.

[0013] The overhaul slag includes the waste cathode and the waste tank lining refractory material part. When the waste cathode is selected, the leached slag in step S2 is carbon; when the refractory material is selected, the leached slag in step S2 is nepheline or feldspar.

[0014] The secondary aluminum ash includes secondary aluminum ash containing a large amount of α-alumina generated in the primary aluminum smelting process, aluminum alloy production and processing process, and waste aluminum recovery and regeneration process. The overhaul slag includes waste cell lining, waste cathode carbon blocks, and cathode ash generated in the overhaul process of aluminum electrolytic cells.

[0015] In one of the preferred embodiments, the secondary aluminum ash includes, by weight: 5-20 parts of aluminum, 5-30 parts of aluminum nitride, 15-60 parts of aluminum oxide, 7-15 parts of spinel (MgAl2O4), 1-5 parts of silicon dioxide, and 1-5 parts of fluoride.

[0016] In one preferred embodiment, the α-crystalline aluminum oxide accounts for 40-90% of the aluminum oxide.

[0017] In one preferred embodiment, the overhaul slag comprises, by weight: 45-85 parts of carbon, 15-26 parts of sodium fluoride, 15-55 parts of electrolyte (cryolith), and 0.1-5 parts of calcium fluoride.

[0018] In one preferred embodiment, the particle sizes of the secondary aluminum ash and the overhaul slag after crushing are both below 50 mesh, further below 150 mesh, and further below 300 mesh.

[0019] In one preferred embodiment, in said S1, the amount of concentrated sulfuric acid added is 1-5 times the mass of the secondary aluminum ash, and the concentration of the concentrated sulfuric acid is 98%.

[0020] If the acid-to-solid ratio is too low, the reaction will be incomplete and the aluminum conversion rate will be low, thereby reducing the subsequent aluminum leaching rate. However, if the acid-to-solid ratio is too high, concentrated sulfuric acid will be wasted and the amount of alkali required to be added will also increase.

[0021] In one preferred embodiment, in said S1, the calcination time is 1-8 hours, further 2-5 hours; and the calcination temperature is 150-500°C.

[0022] If the roasting time is too short, the reaction is incomplete; if it is too long, energy consumption will increase and more hydrogen fluoride will be volatilized, resulting in a waste of fluorine; if the temperature is too low, it is difficult to induce the reaction. Properly increasing the temperature is conducive to accelerating mass transfer and thus accelerating the reaction. However, if the temperature is too high, concentrated sulfuric acid will decompose into sulfur dioxide and water, reducing the amount of concentrated sulfuric acid actually participating in the reaction, resulting in a waste of concentrated sulfuric acid.

[0023] In one of the preferred embodiments, in said S1, the roasting process further comprises: absorbing the gas generated in the roasting process by pure water.

[0024] After the soluble gas is dissolved in pure water, the solution can participate in the subsequent leaching step, and the insoluble gas (such as hydrogen and methane) can be used as energy fuel after being collected.

[0025] In one preferred embodiment, in S1, the leaching temperature is 25-90°C, the leaching time is 30-300 min, the stirring speed is 100-600 rpm, and the leaching liquid-solid ratio is 2-20.

[0026] If the leaching temperature is too low, the time is too short, or the rotation speed is too low, it will affect the efficiency and element extraction rate, while if it is too high, it will increase energy consumption; if the liquid-to-solid ratio is too low, it will lead to incomplete leaching of valuable elements, while if it is too high, it will dilute the concentration of various ions in the solution, which is not conducive to the subsequent production of cryolite.

[0027] In one preferred embodiment, in S1, the leaching temperature is 50-90°C.

[0028] In one preferred embodiment, in S1, the leaching time is 60-180 min.

[0029] In one preferred embodiment, in S1, the stirring speed is 200-500 rpm.

[0030] In one preferred embodiment, in S1, the leaching liquid-to-solid ratio is 5-10.

[0031] In one preferred embodiment, in S2, the leaching temperature is 25-90°C, the leaching time is 30-300min, the stirring speed is 100-600rpm, and the leaching liquid-solid ratio is 2-20.

[0032] In one preferred embodiment, in S2, the leaching temperature is 50-90°C.

[0033] In one preferred embodiment, in S2, the leaching time is 60-180 min.

[0034] In one preferred embodiment, in S2, the stirring speed is 200-500 rpm.

[0035] In one preferred embodiment, in S2, the leaching liquid-to-solid ratio is 5-10.

[0036] In a preferred embodiment, in S3, the pH value of the filtrate B is adjusted to 3-9.

[0037] In the process of the present invention, the pH of the obtained filtrate B is less than 1. Too low a pH affects the yield, and too high a pH causes the produced cryolite to dissolve.

[0038] In one preferred embodiment, in S3, an alkali solution is added to adjust the pH value of the filtrate B.

[0039] In one preferred embodiment, the alkali solution is one of a solution of sodium hydroxide, sodium carbonate, sodium bicarbonate, ammonia water, ammonium carbonate, and ammonium bicarbonate, and the concentration of the alkali solution is 1-4 mol / L.

[0040] In one preferred embodiment, in S3, the reaction temperature is 25-90°C, further 40-75°C; the reaction time is 30-300 min, further 60-180 min.

[0041] Low temperature affects efficiency, while too high temperature will produce hydroxyaluminum fluoride, affecting the purity of cryolite; short time will result in incomplete reaction, while too long time will increase energy consumption.

[0042] In a preferred embodiment, in S3, the crystallization method is evaporation crystallization or freezing crystallization.

[0043] In one preferred embodiment, in said S3, the solution after crystallization filtration is returned to step S1 for leaching the roasted clinker.

[0044] In one preferred embodiment, when the secondary aluminum ash or overhaul slag raw material used contains lithium, sodium carbonate is added to the solution after crystallization and filtration in S3 to precipitate lithium to obtain lithium carbonate.

[0045] The present invention is further explained below:

[0046] Specifically, the present invention comprises the following steps:

[0047] S1. Grind and crush the secondary aluminum ash and the overhaul slag respectively to obtain fine powder of a certain particle size, wherein the secondary aluminum ash is powder A and the overhaul slag is powder B;

[0048] S2, mixing concentrated sulfuric acid with the powder A obtained in step S1 and then roasting; in the roasting step, the aluminum element is converted into soluble aluminum sulfate:

[0049] 2Al+3H2SO4→Al2(SO4)3+3H2↑;

[0050] 2AlN+4H2SO4→Al2(SO4)3+(NH4)2SO4;

[0051] Al2O3+3H2SO4→Al2(SO4)3+3H2O↑;

[0052] 2NaAl 11 O 17 +34H2SO4→11Al2(SO4)3+Na2SO4+34H2O↑;

[0053] 2Na3AlF6+6H2SO4→Al2(SO4)3+3Na2SO4+12HF↑;

[0054] 2K3AlF6+6H2SO4→Al2(SO4)3+3K2SO4+12HF↑;

[0055] S3, crushing the clinker obtained after roasting in step S2, leaching with pure water, filtering, and obtaining filtrate A and filter cake A;

[0056] First leaching, dissolving aluminum sulfate:

[0057] Al2(SO4)3→2Al 3+ +3SO4 2- ;

[0058] S4, using the filtrate A obtained in step S3 as a leaching agent to leach the powder B obtained in step S1, filtering to obtain a filtrate B and a filter cake B;

[0059] The second leaching is to dissolve the valuable elements in the overhaul slag; Al and F are both dissolved; the secondary leaching slag is carbon, or nepheline and feldspar;

[0060] Al 3+ +2NaF→2Na + +AlF2 + ;

[0061] 2Al 3+ +Na3AlF6→3Na + +3AlF2 + ;

[0062] S5, using alkaline solution to adjust the pH of the filtrate B obtained in step S4, filtering, and obtaining filtrate C and filter cake C;

[0063] Adjusting the pH to precipitate cryolite containing Al, Na and F;

[0064] Al 3+ +3Na + +6F - →Na3AlF6;

[0065] S6. Crystallizing the filtrate C obtained in step S5, and obtaining filtrate D and filter cake D after solid-liquid separation to obtain sodium sulfate.

[0066] Compared with the prior art, the present invention has the following advantages:

[0067] (1) The present invention adopts a method for the coordinated disposal and resource utilization of aluminum ash and overhaul slag to recycle the valuable elements in the secondary aluminum ash, especially aluminum in the form of inert α-Al2O3 and MgAl2O4, and coordinately treats the overhaul slag to ultimately obtain α-Al2O3, cryolite, sodium sulfate, and purified carbon, nepheline, feldspar and other value-added products.

[0068] (2) Furthermore, the present invention adopts a concentrated sulfuric acid mixed roasting process for the secondary aluminum ash, which promotes the reaction of concentrated sulfuric acid, α-Al2O3 and MgAl2O4 under high temperature conditions, converting the aluminum therein into soluble aluminum sulfate, thereby improving the leaching rate of aluminum. Compared with direct acid leaching or alkaline leaching, the leaching rate of aluminum element in aluminum ash is greatly improved, effectively increasing by 15%-35%.

[0069] (3) Furthermore, the present invention utilizes the primary leaching solution to cooperatively leach the overhaul slag, thereby achieving the purification of the materials in the overhaul slag and supplementing fluorine, sodium and aluminum for the subsequent cryolite precipitation step. BRIEF DESCRIPTION OF THE DRAWINGS

[0070] Figure 1 It is a schematic diagram of the process flow of the present invention.

[0071] Figure 2 This is the XRD diagram of the α-Al2O3 product obtained in Example 1 of the present invention.

[0072] Figure 3 This is the XRD diagram of the graphite product obtained in Example 1 of the present invention.

[0073] Figure 4 This is the XRD diagram of the cryolite product obtained in Example 1 of the present invention.

[0074] Figure 5 This is the XRD diagram of the sodium sulfate product obtained in Example 1 of the present invention. DETAILED DESCRIPTION

[0075] The present invention is not limited to the following specific embodiments. A person skilled in the art can implement the present invention in various other specific embodiments according to the contents disclosed in the present invention, or any simple changes or modifications made to the design structure and ideas of the present invention fall within the protection scope of the present invention. It should be noted that the embodiments and features in the embodiments of the present invention can be combined with each other without conflict.

[0076] In the embodiment, the aluminum ash components include: aluminum: 5%-20%, aluminum nitride: 5%-30%, aluminum oxide: 15%-60%, mainly α-crystalline, spinel (MgAl2O4): about 7%-15%, silicon dioxide: 1%-5%, fluoride: 1%-5%. The spent cathode components include: carbon: 45%-85%, sodium fluoride: about 15%-26%, electrolyte (cryolite): about 15%-55%, and a small amount of calcium fluoride, about 0.1-5%.

[0077] Example 1

[0078] according to Figure 1 The process flow diagram shown in the figure, the specific implementation steps of this embodiment are as follows:

[0079] S1. Using secondary aluminum ash (Al 28.83%, F 21.78%, O 21.48%, Na 16.35%, N2.83%, Ca 2.45%, K 1.73%, Fe 0.26%, of which α-Al2O3 content is 44.56%) and waste cathode carbon blocks (C43.00%, O 20.04%, F 16.00%, Na 9.32%, Al 4.93%, Si 2.90%, Li 1.03%, Ca0.98%, Fe 0.76%, K 0.58%) of an electrolytic aluminum plant as raw materials, the secondary aluminum ash and the waste cathode carbon blocks are crushed to less than 150 mesh respectively and set aside.

[0080] S2. Take 180g of concentrated sulfuric acid in a corundum crucible, slowly add 60g of aluminum ash (the mass ratio of acid to material is 3g:1g) into the crucible, stir and mix, and then put the corundum crucible into a pit furnace, roast at 150°C for 4 hours, take out, and obtain clinker.

[0081] S3. After the clinker is crushed, the clinker powder is slowly added into a beaker containing pure water (pure water is preheated to 90°C) at a liquid-solid ratio of 12ml:1g under stirring, and a water bath magnetic stirrer is used for reaction. The stirring speed is adjusted to 300rpm, and the reaction is carried out at 90°C for 180min. The primary leaching liquid is obtained by filtering, and the filter cake is shown in FIG. Figure 2 , which is α-Al2O3.

[0082] S4. Use the primary leachate as the leachate, use a water bath magnetic stirrer to react, slowly add the waste cathode carbon block fine powder to the primary leachate at a liquid-to-solid ratio of 6 ml:1 g, adjust the stirring speed to 300 rpm, react at 90 ° C for 120 min, filter, and obtain the secondary leachate. The filter cake is Figure 3 , which is graphite.

[0083] S5. Use a water bath magnetic stirrer to react, adjust the stirring speed to 200 rpm, the temperature to 50 ° C, transfer the secondary leachate to a beaker, insert a pH electrode into the solution to detect pH changes in real time, use a peristaltic pump to slowly pump 1 mol / L sodium hydroxide lye into the secondary leachate, adjust the solution pH to 9.0, react at 60 ° C for 60 min, age for 60 min, filter, and obtain the filtrate. The filter cake is Figure 4 , which is Na3AlF6.

[0084] S6, transfer the filtrate to a rotary evaporator to concentrate and crystallize until crystals are continuously precipitated, filter, and return the filtrate to step S3 to participate in the leaching reaction with pure water, and the filter cake is Figure 5 , is Na2SO4.

[0085] In this Example 1, the leaching rate of aluminum in aluminum ash is 95.34%, the leaching rate of fluorine in the spent cathode is 97.45%, and the cryolite recovery rate is 86.31%.

[0086] Example 2

[0087] On the basis of Example 1, only the following changes were made: in step S2, the mass ratio of the acid material was 2.5 g of concentrated sulfuric acid: 1 g of aluminum ash, and the roasting time was 2 hours. The rest was the same as Example 1. In this Example 2, the leaching rate of aluminum in aluminum ash was 93.32%, the leaching rate of fluorine in the waste cathode was 96.53%, and the cryolite recovery rate was 87.63%.

[0088] Example 3

[0089] On the basis of Example 1, only the following changes were made: in step S3, the liquid-solid ratio was 10 ml: 1 g, the temperature was 50°C, and the reaction time was 90 min. The rest was the same as Example 1. In this Example 2, the leaching rate of aluminum in aluminum ash was 90.85%, the leaching rate of fluorine in the waste cathode was 94.26%, and the cryolite recovery rate was 84.07%.

[0090] Example 4

[0091] On the basis of Example 1, only the following changes were made: in step S4, the liquid-solid ratio was 10 ml: 1 g, the temperature was 80°C, and the reaction time was 120 min. The rest was the same as Example 1. In this Example 2, the leaching rate of aluminum in aluminum ash was 92.34%, the leaching rate of fluorine in the waste cathode was 95.33%, and the cryolite recovery rate was 84.58%.

[0092] Example 5

[0093] In step S5, 2 mol / L sodium hydroxide is used to adjust the pH to 6.0, react at 40° C. for 120 min, and age for 60 min. Others are the same as in Example 1. In this Example 2, the leaching rate of aluminum in aluminum ash is 94.36%, the leaching rate of fluorine in the waste cathode is 95.84%, and the cryolite recovery rate is 88.05%.

[0094] Example 6

[0095] S1. Using secondary aluminum ash (Al 51.47%, O 29.46%, Mg 6.16%, Cl2.98%, N 4.35%, F 1.03%, K 2.27%, Si 0.72%, Fe 0.42%, Ca 0.22%, of which α-Al2O3 content is 61.62%) of an aluminum alloy production enterprise and waste cathode carbon blocks (C 43.00%, O 20.04%, F 16.00%, Na 9.32%, Al 4.93%, Si2.90%, Li 1.03%, Ca 0.98%, Fe 0.76%, K 0.58%) as raw materials, the secondary aluminum ash and the waste cathode carbon blocks are crushed to less than 150 mesh respectively and set aside.

[0096] S2. Take 100g of concentrated sulfuric acid in a corundum crucible, slowly add 20g of aluminum ash (the mass ratio of acid to material is 5g:1g) into the crucible, stir and mix, and then put the corundum crucible into a pit furnace, roast at 250°C for 6 hours, take it out, and obtain clinker.

[0097] S3. After the clinker is crushed, slowly add the clinker powder into a beaker filled with pure water (pure water is preheated to 90°C) at a liquid-to-solid ratio of 15ml:1g under stirring, use a water bath magnetic stirrer to react, adjust the stirring speed to 300rpm, react at 90°C for 240min, and filter.

[0098] S4. Use the primary leachate filtered in S3 as the leachate, use a water bath magnetic stirrer to react, slowly add the waste cathode fine powder to the primary leachate at a liquid-solid ratio of 5 ml:1 g, adjust the stirring speed to 300 rpm, react at 90°C for 120 min, and filter.

[0099] S5. Use a water bath magnetic stirrer to react, adjust the stirring speed to 200 rpm, and the temperature to 50°C. Transfer the secondary leachate obtained by filtration in S4 to a beaker, insert a pH electrode into the solution to detect pH changes in real time, use a peristaltic pump to slowly pump 4 mol / L sodium hydroxide lye into the leachate, adjust the solution pH to 7.0, react at 60°C for 90 min, age for 60 min, and filter.

[0100] S6. Transfer the filtrate to a rotary evaporator to concentrate and crystallize until crystals continue to precipitate, and filter.

[0101] The filter cakes obtained by filtering in steps S3, S4, S5 and S6 are washed and dried to obtain α-Al2O3, C powder, cryolite and sodium sulfate. In this embodiment 3, the leaching rate of aluminum in aluminum ash is 83.92%, the leaching rate of fluorine in waste cathode is 92.49%, and the cryolite recovery rate is 87.69%.

[0102] Example 7

[0103] On the basis of Example 6, only the following changes are made: in step S2, the mass ratio of acid material is 4g concentrated sulfuric acid: 1g aluminum ash, and the roasting time is 4 hours; in step S3, the liquid-solid ratio is 10ml: 1g, the temperature is 70°C, and the reaction time is 180min; in step S4, the liquid-solid ratio is 10ml: 1g, the temperature is 70°C, and the reaction time is 120min; in step S5, 3mol / L sodium hydroxide is used to adjust the pH to 5.0, react at 40°C for 180min, and age for 60min. In this Example 4, the leaching rate of aluminum in aluminum ash is 81.04%, the leaching rate of fluorine in waste cathode is 90.76%, and the cryolite recovery rate is 85.13%.

[0104] Comparative Example 1

[0105] The secondary aluminum ash (below 150 mesh) used in Example 1 was used as raw material, and the aluminum ash was directly leached using sulfuric acid solution. 100 ml of sulfuric acid solution with a concentration of 30% was prepared, and 10 g of aluminum ash was slowly added under stirring. A water bath magnetic stirrer was used for reaction, and the stirring speed was adjusted to 300 rpm. The reaction was carried out at 100° C. for 5 hours, and filtered. Then, the subsequent steps S4-S6 were carried out with the leaching solution, which was the same as Example 1. In this comparative example 1, the leaching rate of aluminum in the aluminum ash was only 78.96%.

[0106] Comparative Example 2

[0107] The secondary aluminum ash (below 150 mesh) used in Example 6 was used as raw material, and the aluminum ash was directly leached using sulfuric acid solution. 100 ml of sulfuric acid solution with a concentration of 30% was prepared, and 10 g of aluminum ash was slowly added under stirring. A water bath magnetic stirrer was used for reaction, and the stirring speed was adjusted to 300 rpm. The reaction was carried out at 100° C. for 5 hours, and filtered. Then, the subsequent steps S4-S6 were carried out with the leaching solution, which was the same as Example 6. In this comparative example 2, the leaching rate of aluminum in the aluminum ash was only 50.52%.

[0108] Comparative Example 3

[0109] The secondary aluminum ash (below 150 mesh) and waste cathode carbon block fine powder used in Example 1 were used as raw materials, and the liquid-to-solid ratio was 3 cm3 / g, soaked in industrial water at 25°C, mechanically stirred for 4 hours, and then filtered to obtain water-leaching residue and water-leaching liquid; the obtained water-leaching residue was mixed with a 2.5 mol / L sulfuric acid solution with a liquid-to-solid ratio of 10 cm 3 / g, the temperature is 100°C, leaching for 3h, hot acid leaching pulp is obtained, and leachate is filtered. Then the leachate is used to perform subsequent steps S5-S6, the same as Example 1. In this comparative example 3, the leaching rate of aluminum in aluminum ash is only 45.99%, and the filter cake obtained after filtering the pulp contains multiple components such as α-alumina, carbon, and cryolite, which is difficult to use directly and needs further separation.

[0110] Comparative Example 4

[0111] Take the secondary aluminum ash (below 150 mesh) and waste cathode carbon block fine powder used in Example 1 as raw materials, add concentrated sulfuric acid at an acid-to-material ratio of 3g:1g, stir and mix, and then put the corundum crucible into a pit furnace, roast at 200°C for 4 hours, take out, and obtain clinker. Then leaching is carried out, and the leaching process is the same as step S3. The leachate is reacted at 90°C for 120 minutes, filtered, and a secondary leachate is obtained. Then proceed to subsequent steps S5-S6, the same as Example 1. In this comparative example 4, the leaching rate of aluminum in the aluminum ash is 71.63%.

[0112] Since the main component of the waste cathode is carbon, the waste cathode carbon block and the secondary aluminum ash are mixed and roasted together with concentrated sulfuric acid. If there is not enough concentrated sulfuric acid, they cannot be mixed evenly, and an excess of concentrated sulfuric acid is required to mix evenly, resulting in a waste of concentrated sulfuric acid. At the same time, concentrated sulfuric acid has strong oxidizing properties and carbon has reducing properties. During the roasting process, redox reactions between concentrated sulfuric acid and carbon may occur, resulting in a large amount of carbon and sulfuric acid loss and waste. In addition, during the mixing process, a large amount of carbon in the waste cathode will wrap up the valuable components in the secondary aluminum ash, and carbon is a hydrophobic substance, which ultimately leads to incomplete reaction between concentrated sulfuric acid and aluminum ash.

[0113] It should be noted that the above embodiments are merely examples for clearly illustrating the present invention, and are not intended to limit the embodiments of the present invention. For those skilled in the art, other different forms of changes or modifications can be made based on the above description. It is impossible to list all the embodiments here. Any obvious changes or modifications derived from the technical solution of the present invention are still within the scope of protection of the present invention.

Claims

1. A method for co-processing secondary aluminum ash and overhaul slag, characterized in that: The following steps are involved: S1. Grind the secondary aluminum ash, add concentrated sulfuric acid, mix well, and then roast to obtain clinker; add water to the clinker, leaching, filtering, and obtaining filtrate A and aluminum oxide; S2, crushing the overhaul slag, adding filtrate A, leaching, filtering, and obtaining filtrate B and leached slag; S3, adjusting the pH value of the filtrate B to 1-11 for reaction, and then filtering to obtain filtrate C and cryolite; The filtrate C was crystallized to obtain sodium sulfate.

2. The collaborative processing method according to claim 1, characterized in that: In parts by weight, the secondary aluminum ash comprises: 5-20 parts of aluminum, 5-30 parts of aluminum nitride, 15-60 parts of aluminum oxide, 7-15 parts of spinel, 1-5 parts of silicon dioxide, and 1-5 parts of fluoride; among the aluminum oxide, α-crystalline aluminum oxide accounts for 40-90%; the overhaul slag comprises: 45-85 parts of carbon, 15-26 parts of sodium fluoride, 15-55 parts of electrolyte, and 0.1-5 parts of calcium fluoride.

3. The collaborative processing method according to claim 1, characterized in that: In the S1, the amount of concentrated sulfuric acid added is 1-5 times the mass of the secondary aluminum ash; the roasting time is 1-8 hours, and the roasting temperature is 150-500°C.

4. The collaborative processing method according to claim 1, characterized in that: In S1, the roasting process also includes: absorbing the gas generated in the roasting process by pure water, and the obtained solution is used in the subsequent leaching step of S1.

5. The collaborative processing method according to claim 1, characterized in that: In S1, the leaching temperature is 25-90°C, the leaching time is 30-300min, the stirring speed is 100-600rpm, and the leaching liquid-solid ratio is 2-20.

6. The collaborative processing method according to claim 1, characterized in that: In S2, the leaching temperature is 25-90°C, the leaching time is 30-300min, the stirring speed is 100-600rpm, and the leaching liquid-solid ratio is 2-20.

7. The collaborative processing method according to claim 1, characterized in that: In S3, alkali solution is added to adjust the pH value of the filtrate B to 3-9; the alkali solution is one of the solutions of sodium hydroxide, sodium carbonate, sodium bicarbonate, ammonia water, ammonium carbonate, and ammonium bicarbonate, and the concentration of the alkali solution is 1-4 mol / L.

8. The collaborative processing method according to claim 1, characterized in that: In S3, the reaction temperature is 25-90°C and the reaction time is 30-300 min.

9. The collaborative processing method according to any one of claims 1 to 8, characterized in that: In S3, the crystallization method is evaporation crystallization or freezing crystallization.

10. The collaborative processing method according to claim 9, characterized in that: In step S3, the solution after crystallization and filtration is returned to step S1 for leaching the roasted clinker.

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