40results about How to "Improve leaching rate" patented technology

This invention relates to a method for the combined leaching of high-silicon zinc roasted ore and high-sulfur zinc oxide powder, belonging to the field of hydrometallurgical technology. The high-sulfur zinc oxide powder is leached neutrally to obtain zinc oxide leaching residue and zinc powder leaching solution. The zinc powder leaching solution is then mixed with high-silicon zinc roasted ore slurry for controlled silica leaching of the high-silicon zinc roasted ore to obtain zinc ore leaching residue and zinc ore leaching solution. The zinc powder leaching residue and zinc ore leaching residue are mixed and then subjected to low-acid leaching to obtain low-leaching residue and low-leaching solution. The low-leaching residue is then subjected to hot acid leaching to obtain hot-leaching residue and hot-leaching solution. The hot-acid residue is then subjected to high-acid leaching to obtain lead-silver slag product and high-leaching solution. The hot-leaching solution is neutralized and reduced by high-sulfur zinc oxide powder to obtain neutralized solution and neutralized residue. The neutralized solution is then subjected to mineralization and iron precipitation to obtain hematite product and iron-removed solution. This invention can effectively control the leaching of silicon in the high-silicon zinc roasted ore leaching process, and simultaneously realize the oxidative leaching of sulfides in zinc oxide powder and the reductive leaching of zinc ferrite in zinc roasted ore, thus simplifying the leaching process of zinc roasted ore and zinc oxide powder.

This invention discloses a method for preparing battery-grade nickel sulfate and iron phosphate using nickel pig iron and ferric phosphate slag as raw materials. The preparation process involves first mixing the ferric phosphate slag with cationic alkyl polyacrylamide to form a uniform mixture. This mixture is then added to dilute acid and stirred to disperse it. Nickel pig iron powder is added, and the mixture is filtered to obtain a leaching solution. The leaching solution is then subjected to nickel-iron separation using resin to obtain an iron-containing solution and a nickel-containing resin. The nickel-containing resin is desorbed to obtain a desorbed solution. A nickel-based pH adjuster is added to this desorbed solution, and the solution is filtered to obtain a nickel sulfate solution. After evaporation and concentration to precipitate crystals, the solution is dried and crushed to obtain battery-grade nickel sulfate. Hydrogen peroxide is added to the iron-containing solution to oxidize ferrous iron to ferric iron. A phosphorus source is added, the pH is adjusted, and the solution is reacted, filtered, and dried to obtain iron phosphate dihydrate. This is then calcined to obtain battery-grade iron phosphate. This invention uses nickel pig iron and ferrophosphate slag as raw materials, and utilizes the ferrophosphate slag to promote the full leaching of iron from nickel pig iron, and nickel pig iron to promote the leaching of phosphate and iron from ferrophosphate. With the addition of cationic alkyl polyacrylamide, it achieves high leaching rates of iron and nickel in a short time, and simultaneously achieves the separation of non-ferrous and nickel metal ions during the acid leaching process, so as to finally obtain high-purity, qualified battery-grade nickel sulfate and ferrophosphate.

This invention provides a method for complexing and leaching beryllium from lepidolite leaching residue. The lepidolite leaching residue is mixed with a sulfuric acid solution containing DTPA, and the pH is adjusted to 0.1-1.5, the DTPA concentration to 0.03-0.1 mol / L, and the solid-liquid ratio to 1g:8-15mL. The resulting slurry is ball-milled, and the exposed beryllium is in situ complexed by DTPA through mechanochemical processes to disrupt the mineral phase lock-in. After solid-liquid separation, a beryllium-containing filtrate and purified residue are obtained. This invention achieves a beryllium leaching rate of over 60%, effectively solving the problems of deep beryllium lock-in and deactivation of chelating agents under strong acid conditions. The process operates at room temperature and low energy consumption, showing promising prospects for industrial application.

The invention relates to a utilization method of low-grade easily dissolvable stone type niobium concentrate, and belongs to the field of mineral processing and hydrometallurgy. The method solves at least one of the problems of low recovery rate of Fe and Nb elements, complex process flow, high energy consumption and cost, harsh reaction conditions, high environmental pollution risk and the like in the utilization of the low-grade easily dissolvable stone type niobium concentrate in the prior art. The invention relates to a utilization method of low-grade easily dissolvable stone type niobium concentrate, which comprises the following steps: uniformly mixing niobium concentrate with a reducing agent, pelletizing, and reducing and roasting to obtain roasted ore; the roasted ore is smashed and then subjected to magnetic separation, and iron ore concentrate and tailings are obtained; acidizing and roasting to obtain re-roasted ore; circulating grass leaching to obtain grass leaching liquid and grass leaching residues; the grass extract is subjected to extraction and back extraction through extraction resin, and niobium-containing back extraction liquid is obtained. According to the method, the leaching rate of niobium and the recovery rate of the iron element are increased, the technological process is simplified, the energy consumption and the production cost are reduced, and green and environment-friendly production is achieved.

The application discloses a method for obtaining high-grade tin concentrate from silver separation residue, and specifically comprises the following steps: uniformly mixing silver separation residue, an additive and carbon powder according to a proportion, preparing spherical material, and obtaining roasting residue through rotary kiln roasting; grinding the obtained roasting residue into powder, using a reducing agent solution to perform tin leaching on the powdered roasting residue, obtaining leaching liquid and leaching residue, and sending the leaching residue to other valuable element recovery; adding an oxidizing agent to the obtained leaching liquid, selectively precipitating tin in the solution, and obtaining tin precipitation residue and tin precipitation liquid. The high-grade tin concentrate can be obtained after the tin precipitation residue is dried and washed. The method directly obtains the tin concentrate from the silver separation residue, and has the advantages of high grade, few impurities, simple overall process operation, small reagent consumption in the tin precipitation process, low cost, short reaction time and no need of heating.

The invention provides a method for recovering lithium from waste lithium iron phosphate black powder, which comprises the following steps: S1, heating, stirring and mixing ethylene glycol and ferric chloride hexahydrate to obtain a eutectic solvent; s2, adding waste lithium iron phosphate black powder into the eutectic solvent, and reacting to obtain reaction slurry; and S3, carrying out solid-liquid separation on the reaction slurry to obtain a lithium-containing leachate. According to the method, the lithium in the waste lithium iron phosphate black powder can be selectively leached, and the leaching rate of the lithium can be increased.

The invention relates to a method for recovering nickel and cobalt from cobalt nickel hydroxide slag. The method comprises the following steps: mixing the cobalt nickel hydroxide slag with an ammonium solution, and carrying out ammonia leaching reaction to obtain an ammonia complexing solution; the ammonia complexing solution is sequentially subjected to oxidation decomplexing and a first sulfuration reaction, and a manganese and copper removal ammonia complexing solution is obtained; the manganese and copper removal ammonia complexing solution is subjected to a reduction reaction to obtain nickel-cobalt powder and a zinc ammonia complexing solution; the zinc-ammonia complexing solution is subjected to a second vulcanization reaction to obtain an ammonium solution, and the ammonium solution is recycled; according to the method, nickel and cobalt in cobalt nickel hydroxide are efficiently recycled, the purity of nickel and cobalt powder preferably reaches 99.5% or above, the recovery rate of nickel preferably reaches 94.2% or above, the recovery rate of cobalt preferably reaches 91.8% or above, and the method is simple in technological process, low in cost, environmentally friendly and good in industrial application prospect.

The invention relates to the technical field of lithium dihydrogen phosphate preparation, and discloses a method for preparing lithium dihydrogen phosphate by recycling lithium batteries, which comprises the following steps: S1, crushing waste batteries; s2, screening and separating components with different particle sizes to obtain battery powder; s3, leaching is conducted, and a lithium-containing leaching solution is obtained; s4, adding an extracting agent to obtain high-purity lithium-containing filtrate; s5, obtaining phosphate for depositing lithium; s6, obtaining a lithium phosphate precipitate; and S7, obtaining a lithium dihydrogen phosphate solution. A small part of the lithium-containing leachate is treated with alkali liquor, metal impurities are removed, phosphate used for precipitating lithium is generated through phosphorus and lithium of the leachate, the prepared phosphate is mixed with the purified lithium-containing filtrate, lithium phosphate is directly precipitated, dependence on external raw materials and the production cost are reduced, and the method is suitable for industrial production. And lithium in the whole process is synchronously recovered through shunting treatment, so that lithium loss caused by shunting is avoided, and the overall recovery rate of lithium is improved.

This invention provides a method and equipment for recovering lithium, fluorine, and aluminum from lithium-rich waste aluminum electrolyte, relating to the field of metal material recycling. The method for recovering lithium, fluorine, and aluminum from lithium-rich waste aluminum electrolyte includes the following steps: S1, the waste aluminum electrolyte is crushed, ground, sieved, and then dried for later use. The method provided by this invention utilizes ultrasonic-assisted leaching of lithium-containing waste aluminum electrolyte with organic acid (malonic acid). Compared with traditional inorganic acid leaching, organic acid has advantages such as good biodegradability, weak acidity, low corrosiveness, recyclability, and environmental friendliness. During neutralization, cryolite byproducts are obtained, which can be returned to the aluminum electrolysis cell for reuse, realizing the recycling of fluorine and aluminum. Finally, high-purity lithium phosphate is obtained through precipitation, effectively utilizing the lithium resources in the waste aluminum electrolyte.

The application discloses a method for extracting lithium, cesium and rubidium from pegmatite lithium concentrate, comprising the following steps: step 1, uniformly mixing mixed roasting additives with pegmatite lithium concentrate to obtain a mixture; step 2, performing alkali roasting on the mixture to obtain a calcine containing lithium, rubidium and cesium; step 3, finely grinding the calcine containing lithium, rubidium and cesium, and then adding a leaching agent to leach the calcine to obtain a solution containing lithium, rubidium and cesium; and step 4, extracting lithium, rubidium and cesium from the solution containing lithium, rubidium and cesium. By adding mixed roasting additives to the pegmatite lithium concentrate raw material, low-temperature alkali roasting can be realized. In addition, the method provided by the application can be used for large-scale industrial production, is energy-saving and environment-friendly, low in cost, and has a wide popularization prospect.

The application provides a method for strengthening leaching of ionic rare earth, which is based on regulation of ZETA potential and seepage rate of clay mineral by cation and anion coordination adsorption to improve leaching efficiency of ionic rare earth, and comprises the following steps: providing ionic rare earth ore, citrate solution and non-ammonium inorganic salt solution, wherein the liquid-solid ratio of the citrate solution to the ionic rare earth ore is 0.1-0.5:1, and the liquid-solid ratio of the non-ammonium inorganic salt solution to the ionic rare earth ore is 0.9-0.5:1; using the citrate solution to perform first leaching treatment on the ionic rare earth ore to obtain first leaching solution; using the non-ammonium inorganic salt solution to perform second leaching treatment on the ionic rare earth ore to obtain second leaching solution; performing first post-processing treatment on the first leaching solution to obtain rare earth; and performing second post-processing treatment on the second leaching solution to obtain rare earth. The method for leaching ionic rare earth has the advantage of high leaching rate.

The invention belongs to the technical field of hydrometallurgy and comprehensive utilization of resources, and relates to a method for extracting lithium from lepidolite at normal pressure and high efficiency based on synergistic activation and targeted impurity removal. The method comprises the following steps of mechanochemical activation-pre-deconstruction, hydrochloric acid normal-pressure reinforced leaching, fluorine and aluminum synergistic precipitation and recycling, deep removal of iron, calcium and magnesium targeted precipitation and lithium recovery. According to the method disclosed by the invention, the lithium extraction process is optimized, and the lepidolite concentrate is subjected to crystal pre-deconstruction by combining mechanical and chemical activation, so that extremely high reaction activity is created for subsequent acid leaching reaction under normal pressure; then, fluorine, aluminum and iron are removed by accurately controlling reaction conditions, and finally, a composite precipitator is added to deeply remove calcium and magnesium. The lithium extraction and impurity removal method is simple in technological process, high in lithium recovery rate, low in impurity content and wide in applicability.

The invention provides a clay type lithium ore external field enhanced resource comprehensive utilization method, and relates to the field of hydrometallurgy. The method comprises the following steps: uniformly mixing clay type lithium ore with concentrated sulfuric acid, and carrying out low-temperature curing to obtain clinker; mixing the clinker with water, and then carrying out leaching reaction to obtain a lithium-rich aluminum leaching solution; carrying out evaporative crystallization on the lithium-rich aluminum leachate to obtain a mixed crystal of lithium sulfate and aluminum sulfate; performing high-temperature pyrolysis reaction on the mixed crystal to obtain an oxide; carrying out high-energy grinding and leaching on the oxide, and filtering to obtain a lithium-containing solution and crude aluminum oxide; and carrying out post-treatment on the lithium-containing solution to obtain the battery-grade lithium carbonate. According to the method, the technological process of the clay type lithium ore is optimized, synchronous leaching of valuable elements lithium and aluminum and efficient separation of the lithium element and the aluminum element are achieved, the problem that the cost is high when lithium is independently extracted is solved, cooperative utilization of clay type lithium ore resources is achieved, and the cost is reduced.

The present application belongs to the technical field of lithium resource leaching, and relates to a method for preparing battery-grade lithium carbonate from low-grade lepidolite. The method comprises the following steps: raw material pretreatment, acid leaching reaction, membrane separation and purification, valuable metal separation and recovery, deep impurity removal and refinement, lithium precipitation and product forming. In the lithium extraction method, low-grade lepidolite, dilute hydrochloric acid and a catalyst are added into a specific plastic reaction kettle at a certain liquid-solid ratio for low-temperature and normal-pressure reaction, and then battery-grade lithium carbonate and other high-value-added products are obtained through subsequent solid-liquid separation, washing, membrane separation, byproduct separation and lithium precipitation processes. The method is energy-saving and environment-friendly, can improve the leaching rate of lithium and the comprehensive utilization rate of lepidolite, has high automation degree, low cost and is suitable for industrial production.

The invention relates to an aluminum and lithium co-extraction method of clay type lithium ore. The aluminum and lithium co-extraction method comprises the following steps: (1) pretreating the clay type lithium ore to obtain crushed ore; (2) the crushed ore is subjected to two-step countercurrent hydrothermal leaching, and aluminum-lithium-rich leaching liquid and leaching residues are obtained; and (3) carrying out adsorption lithium extraction on the aluminum-lithium-rich lixivium to obtain a lithium-rich solution and an adsorption raffinate, carrying out post-treatment on the lithium-rich solution to obtain a lithium salt product, and carrying out post-treatment on the adsorption raffinate to obtain a meta-aluminate product. The method realizes efficient separation and resource utilization of multiple components such as lithium, aluminum and the like, has high lithium salt purity and stable meta-aluminate product quality, and has the characteristics of short process, low energy consumption and suitability for industrial production.

The invention discloses a method and a device for preparing hydrophosphate based on battery powder lithium extraction slag, an operation mode of multi-stage series countercurrent leaching-segmented precise feeding-deep impurity removal and purification-closed cycle crystallization is adopted, a leaching agent adopts first-stage and second-stage segmented feeding in the leaching process, the mass transfer efficiency is greatly enhanced, and the leaching efficiency is improved. The leaching effect of the leaching agent is maximized, the leaching rate of phosphorus is remarkably improved, the problems of low phosphorus leaching rate and high product impurity content in the prior art are effectively solved, no extra wastewater is generated in the whole technological process, the leaching agent and the crystallization mother liquor are in dynamic circulation in the whole production process, and the production cost is reduced. The method has the advantages that recycling of water resources and residual effective components is achieved, environmental protection and economic benefits are achieved, the problems that in the prior art, the technological process is complex, and cost is high are effectively solved, the industrial-grade hydrophosphate product meeting the industrial standard can be directly prepared from the battery powder lithium extraction slag, and the method is suitable for industrial production. And all-component high-value recovery of the waste lithium iron phosphate battery is realized.

The utility model relates to the technical field of battery recovery, in particular to a battery graphite cathode material recovery device which comprises a separation tank body, the lower end of the separation tank body is communicated with the lower end of a dilution tank, and the upper end of the dilution tank is communicated with the lower end of a continuous overflow filtering device. The upper end of the continuous overflow filtering device is communicated with an inlet of the airflow pressurizing device, and an outlet of the airflow pressurizing device is communicated with the upper end of the separation tank body. According to the utility model, firstly, the separation tank body is used for separating crushed graphite and copper ions in copper foil, then the continuous overflow filtering device is used for separating graphite with smaller particle size in a leaching solution again, and the recycling of the leaching solution is realized through the pressurizing device, so that the separation efficiency of the graphite and the copper foil is improved, and meanwhile, the amount of waste liquid is reduced.

The present application belongs to the field of mineral resources utilization, and particularly relates to a method for extracting potassium from potassium-rich slate, which comprises the following steps: subjecting the potassium-rich slate to low-temperature hydrothermal treatment at a temperature T1 to obtain leaching solution A and leaching residue A; in the low-temperature hydrothermal system, the Ca / Si weight ratio is controlled to be 2.2-3.0; the temperature T1 is 120-200 DEG C; subjecting the leaching residue A and an alkali activator to high-temperature hydrothermal treatment at a temperature T2 to obtain leaching solution B and leaching residue B; in the high-temperature hydrothermal system, the Ca / Si weight ratio is 2.2-3.0; the temperature T2 is 1.1-2 times the temperature T1. In view of the problem of difficulty in extracting potassium from potassium-rich slate, the present application proposes an innovative method: first, subjecting the raw material to low-temperature hydrothermal treatment to change its phase and microstructure, and then cooperating with the subsequent high-temperature hydrothermal process, the two synergistic effects can effectively destroy the crystal structure of the potassium-rich slate, thereby significantly improving the extraction efficiency and effect of potassium.

The invention discloses an ammonia leaching process of copper oxide ore, which comprises the following steps: crushing and grinding the copper oxide ore, adding an activating agent into the obtained ore pulp for activating treatment, sequentially adding ammonia liquor and sulfide for leaching treatment and precipitation reaction, and filtering, and copper sulfide concentrate and filtrate are obtained. Compared with a conventional ammonia leaching process, the ammonia leaching process of the copper oxide ore has the advantages that the copper oxide ore is sequentially subjected to crushing, ore grinding, activating treatment, leaching treatment and precipitation reaction, the copper leaching rate of the copper oxide ore can be remarkably increased, meanwhile, agent consumption and environmental pollution are reduced, and the ammonia leaching process is suitable for efficient recovery of complex and difficult-to-treat copper oxide ore; the method has the advantages of simplicity in operation, low cost, high leaching efficiency, good leaching effect, environmental friendliness and the like, and is high in use value and good in application prospect.

A high-silicon copper-containing calcine leaching residue resource utilization method, comprising: mixing high-silicon copper-containing calcine leaching residue with solid strong alkali, roasting at 140-160 DEG C, obtaining roasting alkali residue; the roasting alkali residue is added with calcium oxide and water to carry out concentrated alkali leaching, solid-liquid separation, obtaining alkali leaching residue and alkali leaching liquid; the alkali leaching liquid is added with a sulfidation agent to precipitate copper, obtaining copper precipitation residue and copper precipitation liquid; the copper precipitation liquid is added with calcium oxide to remove silicon, obtaining silicon precipitation residue and silicon removal liquid; copper is enriched in the copper precipitation residue; the "low-temperature roasting + concentrated alkali leaching" process of the application can effectively destroy the stable structure of high-silicon copper-containing calcine leaching residue, so that the copper wrapped by silicon is exposed, the chemical reaction of the high-silicon copper-containing calcine leaching residue leaching process is significantly strengthened, and the copper leaching rate can be greatly improved; the roasting temperature of the application is only 140-160 DEG C, which is much lower than the general alkali roasting temperature, the energy consumption is lower, the possible high-temperature injury is reduced, and the safety is higher.

The application relates to a method and device for leaching gallium and germanium from zinc leaching residue by using a strong magnetic field and composite ultrasonic waves, which comprises the following steps: S1, drying and grinding zinc powder replacement residue cakes to obtain zinc powder replacement residue powder; S2, mixing the zinc powder replacement residue powder obtained in the step S1 with low-concentration sulfuric acid, sealing, heating, inputting oxygen, stirring, ultrasonic treatment in a magnetic field environment, and filtering to obtain first-stage acid leaching residue and first-stage acid leaching solution; S3, mixing the first-stage acid leaching residue obtained in the step S2 with high-concentration sulfuric acid, sealing, heating, inputting oxygen, stirring, ultrasonic treatment in a magnetic field environment, and filtering to obtain second-stage acid leaching residue and second-stage acid leaching solution; and S4, drying the second-stage acid leaching residue obtained in the step S3 to obtain zinc powder replacement residue for leaching gallium and germanium. Compared with the prior art, the application combines oxygen pressure, ultrasonic waves, a magnetic field and stirring and other process conditions, so that more metal elements are exposed from the inside of the zinc powder replacement residue, and the leaching efficiency is improved.

This invention provides a method for the resource utilization of potassium-containing metallurgical solid waste. The method comprises seven steps: ash transportation and storage, pulping, leaching, filtration, carbon removal, water treatment, water storage, and irrigation. This invention achieves a potassium removal rate of over 95% and a sodium removal rate of over 89% after treating the potassium-containing metallurgical solid waste. The leaching residue is recycled in the sintering process of the metallurgical industry, reducing the alkali content of the sintered ore, improving the efficiency of blast furnace ironmaking, and realizing the low-cost recovery and effective utilization of potassium resources from potassium-containing metallurgical solid waste.

The present application belongs to the technical field of metallurgical engineering, and mainly relates to a method for wet mercury extraction from cinnabar based on composite leaching and modified resin adsorption. The method comprises the following steps: adding cinnabar concentrate into dilute sulfuric acid, filtering the impurity-removed cinnabar after stirring and reaction; mixing the impurity-removed cinnabar with a composite leaching solution, and stirring and leaching to obtain a mercury-containing complex solution; the composite leaching solution is a mixed solution of a complexing agent and ferric sulfate; after the mercury-containing complex solution is adsorbed by a modified thiourea type chelating resin, a mercury-containing desorption solution is obtained by desorption, and the mercury-containing desorption solution is subjected to electrolytic deposition to obtain metallic mercury. The method realizes efficient extraction of cinnabar reselected and flotation concentrate, recycling of waste water and harmless utilization of solid waste, and has significant advantages in leaching efficiency, environmental protection and raw material adaptability, and has broad industrial application prospects.

The application belongs to the technical field of metallurgy, and particularly discloses a method for selectively separating vanadium and chromium from vanadium-chromium slag, which specifically comprises the following steps: mixing vanadium-chromium slag with mixed salt to obtain composite mixed raw material, wherein the mixed salt is a mixture of sodium salt and calcium salt, or a mixture of sodium salt and magnesium salt; under an oxygen atmosphere, the composite mixed raw material is selectively oxidized and roasted at 850-950 DEG C for 1-4h, so that vanadium is converted into water-soluble sodium vanadate and / or sodium-calcium composite vanadate capable of releasing vanadium, to obtain composite roasted clinker; water is used as leaching agent to water soak the composite roasted clinker, and then solid-liquid separation is performed, so that vanadium and chromium are separated. The application can effectively shorten the process flow, realize efficient separation of vanadium and chromium, and effectively reduce the production cost.

The invention relates to a method for recovering valuable metals in cobalt slag through ultrasonic strengthening, and belongs to the technical field of hydrometallurgy. According to the method, the cobalt slag produced in the zinc hydrometallurgy purification process is used as a raw material, two-step leaching is conducted on the cobalt slag under the ultrasonic wave condition, the first-step sulfuric acid concentration is 1.5-2.5 mol / L, and the second-step sulfuric acid concentration is 1, 5-2.0 mol / L. Compared with a traditional cobalt slag leaching method, the using amount of a reducing agent is reduced by 40%-50%, the reaction time is shortened to be within 50%, and the method is suitable for industrial production. According to the method, the high leaching rate of valuable metal in the cobalt slag is achieved under the normal pressure and the low temperature, and the slag amount is reduced by 24%-31%. By means of the method, segmented leaching of iron, zinc, cobalt and manganese is achieved, the obtained first-segment leaching solution does not contain manganese basically, and the technical problem that manganese and zinc need to be further separated from the leaching solution is solved; and iron in the cobalt slag is basically completely leached in the first step, so that the influence of iron on cobalt leaching is effectively avoided or reduced.

This invention discloses a method for vanadium extraction by co-roasting vanadium slag with plant ash extract. This invention utilizes plant ash extract as an additive to replace NaCl, and replaces traditional sodium roasting with potassium roasting for vanadium extraction. This completely eliminates the environmental pollution caused by traditional vanadium extraction processes. Furthermore, this invention achieves a vanadium leaching rate as high as 98%, ensuring high vanadium leaching even when completely replacing NaCl as an additive. In addition, using plant ash extract as an additive offers the advantages of readily available and inexpensive raw materials compared to NaCl.

The application relates to a multi-material valuable component efficient leaching device and method based on a pressurized dissolved gas cavitation effect, which comprises a cylindrical pressure-resistant container fixed on a support in a horizontal mode, the end port of the pressure-resistant container is provided with a sealing door; a joint pipe A is arranged at the upper part of one end of the pressure-resistant container, a quick opening and closing valve is connected to the joint pipe A, a joint pipe B is arranged at the upper part of the other end of the pressure-resistant container, an air inlet ball valve is connected to the joint pipe B, and a pressure gauge is arranged in the middle part; a joint pipe C is arranged at the lower part of one end of the pressure-resistant container, a liquid discharge ball valve is connected to the joint pipe C, and a liquid inlet pipe is arranged at the upper part of the end; an aeration assembly connected with the air inlet ball valve is arranged in the pressure-resistant container; a pressurized dissolved gas-instantaneous pressure release cavitation mechanism is adopted to form microscale damage force, break through the microscale mass transfer resistance of a solid matrix, efficiently release the wrapped target components, greatly improve the leaching rate and shorten the extraction time; and process parameters can be flexibly adjusted to adapt to various materials from fragile plant tissues to hard mineral ores.

The application discloses a method for preparing aragonite type calcium carbonate by fixing carbon dioxide with steel slag, which comprises the following steps: (1) crushing and grinding the steel slag into steel slag powder, and then activating the steel slag powder in ultrasonic waves to obtain ultrasonic-activated steel slag powder; (2) mixing the ultrasonic-activated steel slag powder with water, adding pure acetic acid solvent to warm and acidify, and preparing acidified steel slag; (3) CO2 gas quenching the acidified steel slag, adding water to the acetic acid leaching solution and the steel slag powder in a mass ratio of 3-10:1, adjusting the pH of the leaching solution to 7-9, stopping the CO2 input after 10-90 min of stirring, and filtering to obtain calcium bicarbonate filtrate; and (4) crystallizing and filtering the calcium bicarbonate filtrate to obtain aragonite type calcium carbonate. The method can fix carbon dioxide in tail gas on site by using the waste steel slag stacked in a steel plant, achieves the effect of "waste treatment by waste", the leaching reagent can be recycled, and the method is economic and environment-friendly, and solves the technical problems of utilization of steel slag waste and carbon dioxide resources.

The application discloses a method and device for catalytic leaching of argillaceous uranium ore based on ozone, and relates to the technical field of uranium leaching and reaction equipment. The application utilizes biomass raw materials and iron-based material powder to prepare a magnetic composite material with catalysis-adsorption dual functions through in-situ reduction by high-temperature heat treatment. The slurry of argillaceous uranium ore and the magnetic composite material are mixed in a reactor for reaction. Ozone is introduced through an aeration disc to avoid bubble aggregation, significantly increase the gas-liquid contact area, and improve the ozone utilization rate. A spiral guide plate is arranged in the reactor to effectively prolong the gas-liquid contact time and improve the ozone mass transfer efficiency. A magnetic recovery device is used to recover the magnetic material to ensure the catalyst activity, reduce the catalyst loss rate, and realize efficient catalysis. Through optimization of the reactor structure, the catalyst preparation process and the process flow, the ozone utilization rate and the leaching efficiency of the argillaceous uranium ore are significantly improved, and the catalyst recycling and the recycling of residual gas are realized.

The invention relates to the technical field of hydrometallurgy, in particular to a method for strengthening leaching of polymetallic ore associated mica type copper oxide ore. The method for strengthening leaching of the polymetallic ore associated mica type copper oxide ore comprises the following steps that the polymetallic ore associated mica type copper oxide ore is sequentially crushed and ground, and ore powder is obtained; leaching the ore powder by adopting an acid leaching agent; wherein the acid leaching agent comprises sulfuric acid, a fluorine ion providing agent and polyvinylpyrrolidone; and after leaching is finished, solid-liquid separation is conducted, and copper-containing leaching liquid and leaching residues are obtained. According to the method, the problem that the wrapped copper in the polymetallic ore associated mica type copper oxide ore is difficult to leach out is effectively solved through the synergistic effect of the fluorine ions and the polyvinylpyrrolidone in the acid leaching environment, efficient recovery of the copper is achieved, the copper leaching rate is high, the leaching efficiency is high, the process is simple, the cost is low, the energy consumption is low, and environment friendliness is achieved.