Method and equipment for recovering zinc and indium in zinc hydrometallurgy slag dust

Abstract

The invention discloses a method and equipment for recycling zinc and indium in zinc hydrometallurgy slag dust, and belongs to the technical field of zinc hydrometallurgy slag dust recycling, and the method for recycling zinc and indium in zinc hydrometallurgy slag dust comprises the following steps: S1, pretreatment of zinc hydrometallurgy slag dust: S101, grading screening; s102, alkali washing and impurity removal; s103, carrying out acid pickling and desorption; step S2, carrying out segmented synergistic leaching; s3, step-by-step separation and recovery; according to the method, in the pretreatment stage, through cooperative treatment of grading screening, alkali washing, acid washing and microwave roasting, the differential wave absorbing characteristic of microwaves on different mineral components is utilized, a temperature gradient is formed in the material, an inclusion is promoted to generate cracks and dissociate, the phase dissociation degree of indium is increased to 85% or above from 65%-75% of a traditional process, and the mechanical property of indium is improved; the content of part of insoluble phases is reduced by 30%, a key guarantee is provided for efficient dissolution of indium in the subsequent low-acid leaching stage, and the tedious process that indium needs to be enriched for multiple times through a rotary kiln pyrogenic process in a traditional process is avoided.

Description

Technical Field

[0001] This invention belongs to the field of wet zinc smelting slag dust recovery technology, specifically, it relates to a method and equipment for recovering zinc and indium from wet zinc smelting slag dust. Background Technology

[0002] Zinc, as an important non-ferrous metal, is widely used in alloy manufacturing, galvanizing industry and other fields. In the process of hydrometallurgical zinc smelting, about 0.3 tons of zinc leaching residue are generated for every ton of cathode zinc produced. In order to recover zinc, lead and rare metals (such as indium) from the leaching residue, the industry generally uses Welles rotary kiln or fuming furnace to achieve the volatilization of valuable metals in a strong reducing atmosphere above 1350℃, forming hydrometallurgical zinc smelting slag dust (secondary zinc oxide dust) containing zinc and indium. The sum of zinc and lead content in this slag dust is usually greater than 50%, and the indium content is 300-1500g / t. It is an important secondary resource for recovering zinc and indium. At the same time, it also contains impurities such as fluorine, chlorine and sulfur. If it is stored for a long time, it will not only waste resources, but also cause soil and water pollution risks.

[0003] While existing technologies include methods for recovering zinc and indium from wet zinc smelting slag dust, certain shortcomings and problems remain in practical operation: The phase composition of wet zinc smelting slag dust is complex; zinc mainly exists as zincite and sphalerite, while indium exists partly as stable indium oxide and partly as minerals such as zinc ferrite with a spinel structure. Conventional crushing and grinding methods are insufficient to effectively dissociate indium, limiting subsequent leaching efficiency. Furthermore, fluorine and chlorine impurities in the slag dust can corrode equipment and form complexes with metal ions during the leaching stage, potentially reducing the leaching selectivity of zinc and indium. Traditional processes in existing technologies often employ a "one-stage" method. The two-stage leaching scheme of "neutral leaching + one-stage high-acid leaching" recovers zinc first and then enriches indium. However, this process has significant shortcomings: First, the leaching rates of zinc and indium are low in the low-acid leaching stage, requiring multiple enrichment of the leaching residue through rotary kiln pyrolysis until the indium content reaches 2000-3000 g / t before recovery, resulting in high energy consumption and high costs. Second, the acid concentration in the high-acid leaching stage is too high (usually greater than 250 g / L), which not only increases acid consumption costs but also leads to the leaching of large amounts of impurities such as iron and lead, making subsequent purification difficult and requiring high reagent consumption. Furthermore, in existing processes, zinc recovery often involves direct electrolysis of the leaching solution without addressing the Fe content in the leaching solution. 3+ Pb 2+ Deep purification of impurities may result in low purity of cathode zinc; indium recovery often uses a single extraction system, and the extraction and back-extraction processes are single-stage operations. The extraction rate of indium needs to be further improved, and the crude indium obtained after reduction of the back-extraction solution has low purity. Multiple refining processes are required to achieve industrial-grade purity, and the complexity and cost of the process need to be further reduced. Summary of the Invention

[0004] The problem to be solved

[0005] In response to the problems raised in the prior art, the present invention provides a method and equipment for recovering zinc and indium from wet zinc smelting slag dust.

[0006] Technical solution

[0007] To solve the above problems, the present invention adopts the following technical solution.

[0008] This application provides a method for recovering zinc and indium from wet zinc smelting slag dust, the steps of which are as follows:

[0009] Step S1: Pretreatment of slag dust from wet zinc smelting:

[0010] S101. Grading and screening: The wet zinc smelting slag dust is graded and screened by a vibrating sieve to remove slag dust particles and impurity particles.

[0011] S102, Alkaline washing to remove impurities: The graded slag and dust particles are put into an alkaline washing container and sodium hydroxide solution is added. After the reaction, the mixture is filtered to obtain the first filter residue, which is then washed with ultrapure water.

[0012] S103, Acid washing and desorption: Put the washed first filter residue into an acid washing container, add dilute sulfuric acid, acid wash and filter to obtain the second filter residue, then wash the second filter residue with ultrapure water and dry the second filter residue at 80-100℃ for 2-3 hours.

[0013] S104 Microwave roasting: The dried second filter residue is placed into a microwave roasting device, and after roasting and natural cooling, pretreated residue is obtained.

[0014] Step S2: Segmented Co-leaching;

[0015] Step S3: Stepwise separation and recycling.

[0016] Furthermore, the specific steps of segmented collaborative leaching in step S2 include:

[0017] S201, First stage of low acid ultrasonic leaching: The pretreated residue is put into the ultrasonic reaction vessel, sulfuric acid solution is added, the ultrasonic generator is turned on, and after leaching for 50-70 minutes, it is filtered to obtain a first stage of leachate and a first stage of leaching residue.

[0018] S202, two-stage high-acid enhanced leaching: the first-stage leaching residue is put into the enhanced leaching container, sulfuric acid solution is added, and oxygen is introduced at the same time. After leaching for 70-90 minutes, it is filtered to obtain the second-stage leachate and the second-stage leaching residue.

[0019] Furthermore, the specific steps of the stepwise separation and recycling in step S3 include:

[0020] S301, Zinc Recovery: Adjust the pH of the first-stage leachate to 3.5-4.5, add zinc powder, stir at 40-50℃ for 30-40 minutes to remove impurities, filter to obtain purified solution, and send the purified solution into an electrolytic cell to electrolytically deposit cathode zinc, and then melt and cast to obtain zinc ingots.

[0021] S302, Indium Recovery: Mix the second-stage leaching solution with the filter residue after impurity removal in step S301, adjust the pH of the mixture to 1.5-2.5, add P204 extractant and diluent, stir and extract at 25-35℃ for 20-30 min, let stand to separate into layers to obtain a loaded organic phase, back-extract the loaded organic phase with hydrochloric acid to obtain a back-extract, add a sodium sulfite solution with a mass concentration of 10-15 g / L to the back-extract, reduce at 50-60℃ for 30-40 min, filter to obtain crude indium, and purify the crude indium by distillation under vacuum to obtain metallic indium.

[0022] Preferably, the microwave roasting process in step 104 adopts a segmented heating method: first, the temperature is raised from room temperature to 90-110℃ at a rate of 5-8℃ / min, held for 50-70min, and then lowered to room temperature at a rate of 3-5℃ / min.

[0023] Furthermore, in step S201, the ultrasonic generator adopts a pulsed working mode with a pulse period of 3-5s and a working duty cycle of 60-80%.

[0024] Furthermore, in step S301, the zinc powder has a particle size of 100-200 mesh and is vacuum annealed at 900-1000℃ for 1-2 hours before being added.

[0025] Furthermore, in step S302, the P204 extractant and the diluent sulfonated kerosene are mixed at a volume ratio of 1:3-5. The extraction process adopts a three-stage countercurrent extraction, and the back-extraction process adopts a two-stage countercurrent back-extraction.

[0026] An apparatus for recovering zinc and indium from wet zinc smelting slag dust includes a pretreatment unit, a leaching unit, a separation and recovery unit, and a control system module;

[0027] The pretreatment unit includes a vibrating sample sieve module, an alkaline washing tank module, a first filter module, an acid washing tank module, a second filter module, a drying oven module, and a microwave roasting furnace module connected in sequence. The alkaline washing tank module and the acid washing tank module are equipped with heating jackets, stirrers, and online pH monitors. The microwave roasting furnace module is equipped with a power adjustment module and a temperature control system module.

[0028] The leaching unit includes an ultrasonic reactor module, a third filter module, an enhanced leaching tank module, and a fourth filter module connected in sequence. The ultrasonic reactor module is equipped with an ultrasonic generator module, a heating module, a stirring module, and a temperature sensor module. The enhanced leaching tank module is equipped with a heating module, a stirring module, an oxygen inlet pipe, and a flow controller module.

[0029] The separation and recovery unit includes a purification tank module, a fifth filter module, an electrolytic cell module, a melting furnace module, a mixing tank module, an extraction tank module, a back-extraction tank module, a reduction tank module, a sixth filter module, and a vacuum distillation module. The purification tank module is equipped with a pH adjustment module and a stirring module. The electrolytic cell is equipped with an electrode plate, a temperature controller module, and a current and voltage monitoring module. The extraction tank module and the back-extraction tank module are both equipped with a stirring module and a stratification monitoring module. The vacuum distillation module is equipped with a vacuum degree controller module and a temperature sensor module.

[0030] The control system module is electrically connected to the electrical equipment in the pretreatment unit, leaching unit, and separation and recycling unit, respectively, and is used to control the operating parameters of each device and record real-time data.

[0031] Furthermore, the inner wall of the ultrasonic reactor module is coated with polytetrafluoroethylene, and the transducers of the ultrasonic generator module are evenly distributed on the side wall of the reactor, with a quantity of 4-6, and a shock-absorbing buffer layer is provided between the transducers and the side wall of the reactor.

[0032] Furthermore, a gas distributor is provided at the bottom of the enhanced leaching tank module. The gas distributor is made of porous titanium alloy with a pore size of 0.5-1mm, and the surface of the distributor is provided with an anti-clogging coating.

[0033] Beneficial effects

[0034] Compared with the prior art, the beneficial effects of the present invention are as follows:

[0035] (1) In the pretreatment stage of this invention, the "grading and screening-alkali washing-acid washing-microwave roasting" process is used to form a temperature gradient inside the material by utilizing the different absorption characteristics of microwaves on different mineral components. This causes the inclusions to crack and dissociate, increasing the phase dissociation degree of indium from 65-75% in the traditional process to more than 85%. The content of some insoluble phases is reduced by 30%, which provides a key guarantee for the efficient dissolution of indium in the subsequent low acid leaching stage. This avoids the cumbersome process of enriching indium multiple times through rotary kiln fire method in the traditional process. Attached Figure Description

[0036] Figure 1 This is a process flow diagram of the present invention. Detailed Implementation

[0037] To make the objectives, technical solutions, and advantages of the embodiments of this application clearer, the technical solutions of the embodiments of this application will be clearly and completely described below with reference to the accompanying drawings. Obviously, the described embodiments are some embodiments of this application, but not all embodiments. Generally, the components of the embodiments of this application described and shown in the accompanying drawings can be arranged and designed in various different configurations.

[0038] Therefore, the following detailed description of the embodiments of this application provided in the accompanying drawings is not intended to limit the scope of the claimed application, but merely to illustrate selected embodiments of the application. All other embodiments obtained by those skilled in the art based on the embodiments of this application without inventive effort are within the scope of protection of this application.

[0039] Example 1:

[0040] Reference Figure 1 A method for recovering zinc and indium from wet zinc smelting slag dust, comprising the following steps:

[0041] Step S1: Pretreatment of slag dust from wet zinc smelting:

[0042] S101. Grading and screening: The wet zinc smelting slag dust is graded and screened through a vibrating sieve to screen out slag dust particles with a particle size ≤11μm and remove impurity particles with a particle size >11μm.

[0043] S102, Alkaline washing to remove impurities: The graded slag and dust particles are put into an alkaline washing container, and a sodium hydroxide solution with a mass concentration of 0.3-0.5wt% is added. The liquid-solid ratio is controlled at 8-10:1 (mL / g), the temperature is 55-65℃, and the stirring speed is 70-90rpm. After alkaline washing for 40-60min, the mixture is filtered to obtain the first filter residue. The first filter residue is washed with ultrapure water until the pH of the washing solution is 7-8.

[0044] S103, Acid washing and desorption: Put the washed first filter residue into an acid washing container, add dilute sulfuric acid with a mass fraction of 18-22%, control the liquid-solid ratio at 9-11:1 (mL / g), the temperature at 55-65℃, and the stirring speed at 70-90 rpm, acid wash for 40-60 min, filter to obtain the second filter residue, wash the second filter residue with ultrapure water until the pH of the washing solution is 6-7, and dry the second filter residue at 80-100℃ for 2-3 h;

[0045] S104. Microwave Calcination: The dried second filter residue is placed into a microwave calcination device. The microwave power is set to 450-550W, the calcination temperature to 90-110℃, and the holding time to 50-70min. After natural cooling, pretreated residue is obtained.

[0046] Step S2, Segmented Co-leaching:

[0047] S201, First-stage low-acid ultrasonic leaching: The pretreated residue is put into an ultrasonic reaction vessel, and a sulfuric acid solution with a mass concentration of 130-150 g / L is added. The liquid-solid ratio is controlled at 4.5-5.5:1 (mL / g), the temperature is 70-80℃, the stirring speed is 380-420 rpm, the ultrasonic generator is turned on, the ultrasonic frequency is set to 20-40 kHz, and after leaching for 50-70 min, it is filtered to obtain a first-stage leachate and a first-stage leachate residue.

[0048] S202, Two-stage high-acid enhanced leaching: The first-stage leaching residue is put into an enhanced leaching container, and a sulfuric acid solution with a mass concentration of 180-220 g / L is added. The liquid-solid ratio is controlled at 5-6:1 (mL / g), the temperature is 85-95℃, the stirring speed is 400-450 rpm, and oxygen is introduced at the same time, with the oxygen flow rate controlled at 0.2-0.4 L / min. After leaching for 70-90 min, the solution is filtered to obtain the second-stage leachate and the second-stage leaching residue.

[0049] Step S3, Stepwise Separation and Recycling:

[0050] S301, Zinc Recovery: Adjust the pH of the first-stage leachate to 3.5-4.5, add zinc powder with a mass concentration of 5-8 g / L, stir at 40-50℃ for 30-40 min to remove impurities, filter to obtain purified solution, and send the purified solution into an electrolytic cell, controlling the electrolysis temperature at 35-45℃ and the current density at 400-500 A / m. 2 The cell voltage is 3.2-3.5V, and cathode zinc is obtained by electrolytic deposition, which is then melted and cast to obtain zinc ingots;

[0051] S302, Indium Recovery: Mix the second-stage leaching solution with the filter residue after impurity removal in S3.1, adjust the pH of the mixture to 1.5-2.5, add 20-30 vol% P204 extractant, stir and extract at 25-35℃ for 20-30 min, allow to stand and separate into layers to obtain a loaded organic phase, back-extract the loaded organic phase with 3-5 mol / L hydrochloric acid to obtain a back-extract, add a sodium sulfite solution with a mass concentration of 10-15 g / L to the back-extract, reduce at 50-60℃ for 30-40 min, filter to obtain crude indium, and purify the crude indium by distillation under vacuum to obtain metallic indium with a purity ≥99.99%.

[0052] The pretreatment stage of this invention achieves multiple optimizations through a synergistic process of "grading and sieving - alkaline washing - acid washing - microwave roasting":

[0053] Grading and screening removes slag and dust particles with a diameter ≤11μm and impurities such as SiO2 with a diameter >11μm, thereby increasing the contact area between the leaching agent and the valuable metal by 15-20%. Alkali washing can reduce the chlorine content in the slag and dust from 5-6% to below 0.35%, avoiding the corrosion of equipment and interference with the leaching reaction caused by impurities. Acid washing can remove the oxide layer and residual alkali on the surface of the slag and dust, ensuring the uniformity of subsequent microwave roasting.

[0054] Microwave calcination uses a power of 450-550W and a segmented heating mode of "heating up at 5-8℃ / min - holding at 5℃ / min - cooling down at 3-5℃ / min". This can increase the phase dissociation degree of indium from 50-60% in traditional processes to over 85%, and reduce the content of encapsulated phases such as Zn3In2O6 by 30%. This provides a key guarantee for the efficient dissolution of indium in the subsequent leaching stage. Compared with traditional pyrometallurgical enrichment, the pretreatment energy consumption is reduced by 60-70%.

[0055] Meanwhile, this invention employs a segmented leaching scheme of "one stage of low-acid ultrasonic leaching + two stages of high-acid oxygen-enhanced leaching":

[0056] A low-acid leaching process combined with pulsed ultrasound utilizes the cavitation effect of ultrasound to break down the adsorption layer on the mineral surface, promoting the selective dissolution of zinc and indium. The zinc concentration in the first-stage leaching solution reaches 85 g / L and the indium concentration reaches 420 mg / L. The single-pass leaching rates of zinc and indium reach 83-85% and 78-82%, respectively. Compared with conventional stirred leaching, the leaching efficiency is increased by 25-30%, and the leaching time is shortened to 50-70 min.

[0057] The second-stage high-acid leaching process introduces oxygen at a rate of 0.2-0.4 L / min, which, combined with a porous titanium alloy gas distributor at the bottom of the tank, ensures uniform oxygen dispersion and enhances the dissolution of residual indium in the first-stage leaching residue. The indium concentration in the second-stage leaching solution reaches 1850 mg / L, ultimately increasing the total indium leaching rate to over 88%. Simultaneously, the staged leaching process avoids the problem of excessive impurity dissolution caused by single high-acid leaching, reducing the content of impurities such as iron and lead in the leaching solution by 40-45%, effectively reducing the difficulty of subsequent purification.

[0058] In step 104, the microwave roasting process adopts a segmented heating method: first, the temperature is raised from room temperature to 90-110℃ at a rate of 5-8℃ / min, held for 50-70min, and then lowered to room temperature at a rate of 3-5℃ / min.

[0059] In step S201, the ultrasonic generator adopts a pulse working mode with a pulse period of 3-5s and a working duty cycle of 60-80%.

[0060] In step S301, the zinc powder has a particle size of 100-200 mesh and is vacuum annealed at 900-1000℃ for 1-2 hours before being added.

[0061] In step S302, P204 extractant and sulfonated kerosene diluent are mixed at a volume ratio of 1:3-5. The extraction process adopts three-stage countercurrent extraction, and the back-extraction process adopts two-stage countercurrent back-extraction.

[0062] An apparatus for recovering zinc and indium from wet zinc smelting slag dust includes a pretreatment unit, a leaching unit, a separation and recovery unit, and a control system module;

[0063] The pretreatment unit includes a vibrating sieve, an alkaline washing tank, a first filter, an acid washing tank, a second filter, a drying oven, and a microwave roasting furnace connected in sequence. Both the alkaline washing tank and the acid washing tank are equipped with heating jackets, stirrers, and online pH monitors. The microwave roasting furnace is equipped with a power adjustment module and a temperature control system module. The leaching unit includes an ultrasonic reactor, a third filter, an enhanced leaching tank, and a fourth filter connected in sequence. The ultrasonic reactor is equipped with an ultrasonic generator, a heating device, a stirring device, and a temperature sensor. The enhanced leaching tank is equipped with a heating device, a stirring device, an oxygen inlet pipe, and a flow controller. The separation and recovery unit includes a removal tank, a fifth filter, an electrolytic cell, a melting furnace, a mixing tank, an extraction tank, a back-extraction tank, a reduction tank, a sixth filter, and a vacuum distillation unit. The removal tank is equipped with a pH adjustment device and a stirring device. The electrolytic cell is equipped with electrode plates, a temperature controller, and a current and voltage monitor. The extraction tank and the back-extraction tank are both equipped with stirring devices and a stratification monitor. The vacuum distillation unit is equipped with a vacuum controller and a temperature sensor. The control system module is electrically connected to the electrical equipment in the pretreatment unit, the leaching unit, and the separation and recovery unit, respectively, to control the operating parameters of each device and record real-time data.

[0064] The inner wall of the ultrasonic reactor is coated with polytetrafluoroethylene. The transducers of the ultrasonic generator are evenly distributed on the side wall of the reactor, with a quantity of 4-6. A shock-absorbing buffer layer is provided between the transducers and the side wall of the reactor.

[0065] The bottom of the enhanced leaching tank is equipped with a gas distributor. The gas distributor is made of porous titanium alloy with a pore size of 0.5-1mm, and the surface of the distributor is coated with an anti-clogging coating.

[0066] In practical implementation, this embodiment addresses the slag and dust generated by a hydrometallurgical zinc smelting enterprise in Liupanshui, Guizhou Province, and utilizes the method and equipment of this invention to recover zinc and indium. The process parameters are as follows:

[0067] Pretreatment of wet zinc smelting slag dust: Turn on the vibrating sieve and feed the wet zinc smelting slag dust into the sieve at a uniform speed. Control the sieving rate to 20kg / h and screen out slag dust particles with a particle size ≤11μm. Impurity particles with a particle size >11μm are collected and sent to the waste slag treatment system.

[0068] The graded slag particles were put into an alkaline washing tank, and 800L of sodium hydroxide solution with a mass concentration of 0.4wt% was added. The heating jacket was turned on to raise the temperature to 60℃, and the stirring speed was adjusted to 80rpm. After alkaline washing for 50min, the residue was filtered through the first filter to obtain the first filter residue. The first filter residue was washed three times countercurrently with ultrapure water, with a water volume of 50L each time, until the pH of the washing solution was 7.5, thus completing the removal of fluoride and chlorine.

[0069] The washed first filter residue was put into an acid washing tank, and 900L of 20% dilute sulfuric acid was added. The mixture was heated to 60°C and stirred at 80 rpm for 50 minutes. After acid washing, the residue was filtered through a second filter to obtain the second filter residue. The second filter residue was washed with ultrapure water until the pH of the washing solution reached 6.5. The second filter residue was then sent to a drying oven and dried at 90°C for 2.5 hours to obtain 88.2 kg of dried filter residue (moisture content ≤0.5%).

[0070] The dried second filter residue was placed in a microwave roasting furnace, the microwave power was set to 500W, and a segmented heating mode was adopted: the temperature was raised from room temperature to 100℃ at a rate of 6℃ / min, held for 60min, and then lowered to room temperature at a rate of 4℃ / min. After natural cooling, 87.5kg of pretreated residue was obtained.

[0071] Segmented co-leaching: The pretreated residue was put into an ultrasonic reactor, and 437.5L of sulfuric acid solution with a mass concentration of 140g / L (liquid-solid ratio 5:1mL / g) was added. The mixture was heated to 75℃, the stirring speed was adjusted to 400rpm, and the ultrasonic generator was turned on (pulse period 4s, duty cycle 70%, ultrasonic frequency 30kHz). After leaching for 60min, the residue was filtered through a third filter to obtain 420L of first-stage leachate and 21.3kg of first-stage leachate residue.

[0072] The first stage leaching residue was added to an enhanced leaching tank, and 127.8L of sulfuric acid solution with a mass concentration of 200g / L (liquid-solid ratio 6:1mL / g) was added. The mixture was heated to 90℃, stirred at 420rpm, and oxygen was simultaneously introduced through a porous titanium alloy gas distributor at the bottom of the tank, with the oxygen flow rate controlled at 0.3L / min. After leaching for 80min, the residue was filtered through a fourth filter to obtain 120L of second stage leaching solution (In concentration 1850mg / L, Zn concentration 25g / L) and 8.5kg of second stage leaching residue.

[0073] Stepwise separation and recovery: A first-stage leachate (420L) is sent to a purification tank, where the pH is adjusted to 4.0 with a 10wt% sodium hydroxide solution. Then, 3.36kg (8g / L) of 150-mesh zinc powder (previously vacuum-annealed at 950℃ for 1.5h) is added, and the mixture is stirred at 45℃ for 35min to remove impurities (removing Fe). 3+ Pb 2+(Impurities, etc.) are filtered through a fifth filter to obtain 415L of purified liquid and 2.8kg of filtered residue. The purified liquid is then fed into an electrolytic cell, where the electrolysis temperature is controlled at 40℃ and the current density at 450A / m³. 2 With a cell voltage of 3.3V, electrolytic deposition was carried out for 24 hours to obtain 28.5 kg of cathode zinc. The cathode zinc was then sent to a melting furnace and melted at 500℃ to obtain 28.2 kg of zinc ingots.

[0074] The second-stage leachate (120 L) and the impurity-removed filter residue (2.8 kg) were added to a mixing tank and stirred for 30 min. The pH of the mixture was then adjusted to 2.0 with 5 wt% sulfuric acid. The mixture was then transferred to an extraction tank, and 120 L of 25 vol% P204 extractant was added. The mixture was stirred and extracted at 30 °C for 25 min, and then allowed to stand for 30 min to separate the layers, yielding 118 L of the supported organic phase. The supported organic phase was then transferred to a back-extraction tank, and 4 mol / L hydrochloric acid was added. 60L of indium was stirred and back-extracted for 20 min, then allowed to stand for 20 min to separate into layers, yielding 58L of back-extracted solution. The back-extracted solution was then transferred to a reduction tank, and 5.8L of sodium sulfite solution with a mass concentration of 12g / L was added. The solution was reduced at 55℃ for 35 min, and then filtered through a sixth filter to obtain 261g of crude indium. The crude indium was then transferred to a vacuum distillation apparatus and purified by distillation at a vacuum of 1×10-3 Pa and a temperature of 1200℃ to obtain 252g of metallic indium with a purity ≥99.99%.

[0075] In this embodiment, the zinc recovery rate reached 96.5%, the indium recovery rate reached 88.2%, and the energy consumption per ton of slag and dust was reduced by 25% and the acid consumption was reduced by 30% compared with the traditional process.

[0076] Example 2:

[0077] A method for recovering zinc and indium from wet zinc smelting slag dust is basically the same as in Example 1, except that:

[0078] This embodiment addresses the slag and dust generated by a hydrometallurgical zinc smelting plant in Yunnan Province. The process parameters are as follows:

[0079] Pretreatment stage: Alkaline washing with sodium hydroxide concentration of 0.3wt%, liquid-to-solid ratio of 9:1, temperature of 55℃, time of 60min; Acid washing with dilute sulfuric acid concentration of 18%, liquid-to-solid ratio of 10:1, temperature of 55℃, time of 60min; Microwave roasting with power of 450W, segmented heating rate of 5℃ / min, roasting temperature of 90℃, holding time of 70min.

[0080] Leaching stages: First stage leaching: sulfuric acid concentration 130 g / L, liquid-to-solid ratio 4.5:1, temperature 70℃, ultrasonic frequency 20 kHz, pulse period 3 s, duty cycle 60%, time 70 min; Second stage leaching: sulfuric acid concentration 180 g / L, liquid-to-solid ratio 5:1, temperature 85℃, oxygen flow rate 0.2 L / min, time 90 min;

[0081] Separation and recovery stage: Zinc recovery pH 3.5, zinc powder concentration 5 g / L, temperature 40℃, electrolysis current density 400 A / m 2 The tank voltage is 3.2V; the indium recovery mixture has a pH of 1.5, a P2O4 concentration of 20 vol% (1:3 volume ratio with sulfonated kerosene), and undergoes three-stage countercurrent extraction, two-stage countercurrent back-extraction, and back-extraction hydrochloric acid concentration of 3 mol / L, sodium sulfite concentration of 10 g / L, and a reduction temperature of 50℃.

[0082] This embodiment ultimately yielded 26.8 kg of zinc ingots and 215 g of metallic indium, with a 23% reduction in slag dust energy consumption and a 28% reduction in acid consumption.

[0083] Example 3:

[0084] A method for recovering zinc and indium from wet zinc smelting slag dust is basically the same as in Example 1, except that:

[0085] This embodiment addresses the slag and dust generated by a hydrometallurgical zinc smelting plant in Guangxi. The process parameters are as follows:

[0086] Pretreatment stage: Alkaline washing with sodium hydroxide concentration of 0.5wt%, liquid-to-solid ratio of 10:1, temperature of 65℃, time of 40min; Acid washing with dilute sulfuric acid concentration of 22%, liquid-to-solid ratio of 11:1, temperature of 65℃, time of 40min; Microwave roasting with power of 550W, segmented heating rate of 8℃ / min, roasting temperature of 110℃, and holding time of 50min.

[0087] Leaching stages: First stage leaching: sulfuric acid concentration 150 g / L, liquid-to-solid ratio 5.5:1, temperature 80℃, ultrasonic frequency 40 kHz, pulse period 5 s, duty cycle 80%, time 50 min; Second stage leaching: sulfuric acid concentration 220 g / L, liquid-to-solid ratio 6:1, temperature 95℃, oxygen flow rate 0.4 L / min, time 70 min;

[0088] Separation and recovery stage: Zinc recovery pH 4.5, zinc powder concentration 7 g / L, temperature 50℃, electrolysis current density 500 A / m 2 The tank voltage is 3.5V; the indium recovery mixture has a pH of 2.5, a P2O4 concentration of 30 vol% (1:5 volume ratio with sulfonated kerosene), and undergoes three-stage countercurrent extraction, two-stage countercurrent back-extraction, and back-extraction hydrochloric acid concentration of 5 mol / L, sodium sulfite concentration of 15 g / L, and a reduction temperature of 60℃.

[0089] This embodiment ultimately yielded 30.1 kg of zinc ingots and 285 g of metallic indium, with a 28% reduction in slag dust energy consumption and a 32% reduction in acid consumption.

[0090] The embodiments described above are merely preferred embodiments of the present invention, and while the descriptions are specific and detailed, they should not be construed as limiting the scope of the present invention. It should be noted that those skilled in the art can make various modifications, improvements, and substitutions without departing from the concept of the present invention, and these all fall within the protection scope of the present invention.

Claims

1. A method for recovering zinc and indium from a zinc hydrometallurgy residue dust, characterized in that, The steps are as follows: Step S1, pretreatment of zinc residue dust by wet method: S101, grading screening: the zinc residue dust by wet method is graded and screened by a vibrating sample screen, the residue dust particles are screened, and the impurity particles are removed; S102, alkali washing and impurity removal: the graded residue dust particles are put into an alkali washing container, and sodium hydroxide solution is added, after reaction, filtration is performed to obtain first filter residue, and the first filter residue is washed with ultrapure water; S103, acid washing and desorption: the washed first filter residue is put into an acid washing container, dilute sulfuric acid is added, after acid washing, filtration is performed to obtain second filter residue, the second filter residue is washed with ultrapure water, and the second filter residue is dried at 80-100℃ for 2-3h; S104, microwave roasting: the dried second filter residue is put into a microwave roasting device, after roasting, natural cooling is performed to obtain pretreated residue dust; Step S2, staged and coordinated leaching; Step S3, stepwise separation and recovery.

2. The method according to claim 1, characterized in that: The specific steps of the staged and coordinated leaching in step S2 include: S201, one-stage low-acid ultrasonic leaching: the pretreated residue dust is put into an ultrasonic reaction container, sulfuric acid solution is added, an ultrasonic generator is started, after leaching for 50-70min, filtration is performed to obtain one-stage leaching liquid and one-stage leaching residue; S202, two-stage high-acid intensified leaching: the one-stage leaching residue is put into an intensified leaching container, sulfuric acid solution is added, and oxygen is introduced, after leaching for 70-90min, filtration is performed to obtain two-stage leaching liquid and two-stage leaching residue.

3. The method according to claim 1, characterized in that: The specific steps of the stepwise separation and recovery in step S3 include: S301, zinc recovery: the one-stage leaching liquid is adjusted to pH 3.5-4.5, zinc powder is added, impurities are removed by stirring at 40-50℃ for 30-40min, filtration is performed to obtain purified liquid, the purified liquid is sent into an electrolytic cell, cathode zinc is obtained by electrolytic deposition, and zinc ingot is obtained after melting and casting; S302, indium recovery: the two-stage leaching liquid is mixed with the filter residue after impurity removal in step S301, the mixed liquid is adjusted to pH 1.5-2.5, P204 extractant and diluent are added, stirring extraction is performed at 25-35℃ for 20-30min, after standing and layering, the loaded organic phase is obtained, the loaded organic phase is back-extracted with hydrochloric acid to obtain back-extraction liquid, sodium sulfite solution with a mass concentration of 10-15g / L is added to the back-extraction liquid, reduction is performed at 50-60℃ for 30-40min, filtration is performed to obtain crude indium, and the crude indium is distilled and purified under vacuum conditions to obtain metallic indium.

4. The method according to claim 1, characterized in that: The microwave roasting process in step 104 adopts staged temperature rising: first, the temperature is raised from room temperature to 90-110℃ at a rate of 5-8℃ / min, and then the temperature is lowered to room temperature at a rate of 3-5℃ / min after holding for 50-70min.

5. The method according to claim 2, characterized in that: The ultrasonic generator in step S201 adopts a pulse working mode, the pulse period is 3-5s, and the working duty cycle is 60-80%.

6. The method according to claim 3, characterized in that: In step S301, the particle size of the zinc powder is 100-200 mesh, and the zinc powder is vacuum annealed at 900-1000℃ for 1-2h before being added.

7. The method according to claim 3, characterized in that: In step S302, the P204 extractant is mixed with sulfonated kerosene as diluent at a volume ratio of 1:3-5, the extraction process adopts three-stage countercurrent extraction, and the back-extraction process adopts two-stage countercurrent back-extraction.

8. A recovery apparatus for use in the recovery process of claim 3, characterized by The application relates to a leaching system for extracting rare earth elements from ores, which comprises a pretreatment unit, a leaching unit, a separation and recovery unit and a control system module. The pretreatment unit comprises vibration sample divider modules, alkali washing tank modules, first filter modules, acid washing tank modules, second filter modules, drying box modules and microwave roasting furnace modules which are sequentially connected; the alkali washing tank modules and the acid washing tank modules are respectively provided with heating jackets, stirrers and pH online monitors; and the microwave roasting furnace modules are respectively provided with power adjusting modules and temperature control system modules. The leaching unit comprises ultrasonic reaction kettle modules, third filter modules, intensified leaching tank modules and fourth filter modules which are sequentially connected; the ultrasonic reaction kettle modules are respectively provided with ultrasonic wave generator modules, heating modules, stirring modules and temperature sensor modules; and the intensified leaching tank modules are respectively provided with heating modules, stirring modules, oxygen inlet pipelines and flow controllers. The separation and recovery unit comprises impurity removal tank modules, fifth filter modules, electrolytic cell modules, smelting furnace modules, mixing tank modules, extraction tank modules, back-extraction tank modules, reduction tank modules, sixth filter modules and vacuum distillation modules; the impurity removal tank modules are respectively provided with pH adjusting modules and stirring modules; the electrolytic cell modules are respectively provided with electrode plates, temperature controller modules and current-voltage monitors; the extraction tank modules and the back-extraction tank modules are respectively provided with stirring modules and layering monitors; and the vacuum distillation modules are respectively provided with vacuum degree controllers and temperature sensor modules. The control system module is electrically connected with the electric equipment in the pretreatment unit, the leaching unit and the separation and recovery unit respectively, is used for controlling the operation parameters of the equipment, and records real-time data.

9. The equipment for recovering zinc and indium from zinc hydrometallurgy residue dust according to claim 8, characterized in that: The inner wall of the ultrasonic reaction kettle module is provided with a polytetrafluoroethylene coating; the transducers of the ultrasonic wave generator module are uniformly distributed on the side wall of the reaction kettle, the number of the transducers is 4-6, and a shock-absorbing buffer layer is arranged between the transducers and the side wall of the reaction kettle.

10. The equipment for recovering zinc and indium from zinc hydrometallurgy residue dust according to claim 8, characterized in that: The bottom of the intensified leaching tank module is provided with a gas distributor which is made of porous titanium alloy and has a pore diameter of 0.5-1 mm; and an anti-blocking coating is arranged on the surface of the gas distributor.

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