CN109811132B - Method for comprehensively recycling carbon, iron, aluminum, zinc and lead from blast furnace gas mud - Google Patents
Method for comprehensively recycling carbon, iron, aluminum, zinc and lead from blast furnace gas mud Download PDFInfo
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Abstract
The invention discloses a method for comprehensively recycling carbon, iron, zinc, aluminum and lead from blast furnace gas mud, which comprises the following steps of drying, crushing, grinding, multi-stage flotation, acid washing and suction filtration for recycling carbon; performing acid leaching, oxidation and filtration on the tailings after carbon recovery to obtain filtrate and filter residues; then adjusting the pH value of the obtained filtrate, precipitating iron hydroxide, zinc hydroxide and aluminum hydroxide step by step, and roasting to obtain iron ore concentrate, zinc ore concentrate and aluminum ore concentrate respectively; and adding a saturated ammonium acetate solution into the filtered filter residue, stirring and filtering, adding a sodium carbonate solution into the filtrate, filtering to obtain basic lead carbonate precipitate, and roasting to obtain lead concentrate. The recovery process provided by the invention is simple to operate, can comprehensively recover carbon, iron, zinc, aluminum and lead in the gas mud, realizes recycling of multiple valuable elements, reduces pollution to the environment, has important economic and environmental benefits, and is suitable for popularization and application.
Description
Technical Field
The invention belongs to the field of metallurgical solid waste recycling, relates to a comprehensive recycling method of multiple valuable elements in blast furnace gas mud, and particularly relates to a method for comprehensively recycling carbon, iron, aluminum, zinc and lead from the blast furnace gas mud.
Background
The blast furnace gas mud is a product obtained by wet dedusting, washing, precipitating and concentrating superfine particles generated by reaction in a high-temperature zone in the steel smelting process and raw material dust carried by blast furnace gas, and the main components of the blast furnace gas mud are carbon, iron, zinc and valuable or toxic metal elements such as aluminum, lead and the like with relatively high content. At present, the amount of blast furnace gas mud produced in China is over 500 million tons every year, and the blast furnace gas mud is generally used as a sintering ore ingredient except for being stockpiled as solid waste. Because the gas mud has fine granularity, the air permeability of a sinter bed can be reduced in the sintering process, and the sintering efficiency is influenced. Meanwhile, volatile elements such as zinc and lead in the gas mud are circularly enriched in the blast furnace, so that the upper part of the blast furnace is nodulated and a gas pipeline is blocked, thereby influencing the normal operation of the blast furnace.
Carbon, iron, zinc, aluminum, lead and the like contained in the blast furnace gas mud are very important industrial raw materials, if reasonable recycling is not carried out, not only is the resource waste greatly, but also heavy metal elements in the gas mud can cause serious pollution to the environment if the gas mud is not treated and directly stacked and buried or is simply used as a roadbed material. Therefore, the method comprehensively recovers the carbon, iron, zinc, aluminum, lead and other valuable metal elements in the blast furnace gas mud, can realize the recycling of metallurgical solid wastes, can also reduce environmental pollution, and realizes the green sustainable development of industry.
At present, the comprehensive recycling of the gas mud mainly comprises the following steps of recycling carbon, iron and zinc from the gas mud by a physical and chemical method: chinese patent (publication No. CN105063254B) "a method for separating iron, zinc and carbon from blast furnace gas mud" is to magnetically roast and oxidize the gas mud, then magnetically separate iron ore concentrate and zinc-rich tailings, leach and separate solid and liquid the zinc-rich material to obtain zinc-rich liquid and carbon fine powder, thus realizing the comprehensive recovery of carbon, iron and zinc in the gas mud; chinese patent (publication No. CN106119557B) "a method for comprehensively recovering zinc, iron and carbon from blast furnace gas sludge" dissolving gas sludge with sulfuric acid, filtering, and oxidizing, neutralizing, and filtering the filtrate to obtain iron slag; electrodepositing the filtrate after iron removal to obtain electrolytic zinc; mixing the filter residue obtained by dissolving the gas mud and the iron slag obtained by filtering, magnetizing and roasting to obtain iron ore concentrate; and the tailings are subjected to flotation to recover carbon. However, in the separation process of the iron ore concentrate, the iron oxide is converted into the magnetite by utilizing the reducing agent carbon in the gas mud, the carbon in the gas mud is consumed, and carbon monoxide and carbon dioxide gas are released, so that the recovery of the carbon in the gas mud is reduced, the emission of greenhouse gas is increased, and a certain damage effect on the environment is achieved; in addition, because a large amount of carbon powder with fine granularity to micro-nano scale exists in the gas mud, the gas mud is difficult to separate from magnetite during magnetic separation, so that the grade of iron ore concentrate is reduced; meanwhile, the gas mud also contains valuable metal element aluminum and toxic metal element lead with relatively high content, and if the gas mud is not recycled, the gas mud is a waste of resources and can increase the pollution to the environment.
Disclosure of Invention
The invention mainly aims to solve the problems of low carbon recovery rate, low iron concentrate grade, difficulty in recovering aluminum and lead, easiness in causing environmental pollution and the like of the blast furnace gas sludge in the prior art, and provides a method for comprehensively recovering and utilizing carbon, iron, zinc, aluminum and lead from the blast furnace gas sludge, so that the high-efficiency recovery and utilization and low-pollution discharge of the solid waste treatment process of the blast furnace gas sludge are synchronously realized, and the economic and environmental benefits are considered.
In order to achieve the purpose, the invention adopts the technical scheme that:
a method for comprehensively recycling carbon, iron, zinc, aluminum and lead from blast furnace gas mud comprises the following steps of firstly recycling carbon by a flotation-acid cleaning-suction filtration method, and then sequentially recycling iron, zinc, aluminum and lead by a fractional precipitation method:
1) drying and grinding blast furnace gas mud, adding water, stirring into ore pulp, adjusting the pH value of the obtained ore pulp, adding a collecting agent and a foaming agent, performing flotation to collect carbon pulp, performing suction filtration and drying to obtain carbon concentrate I, pickling, and performing suction filtration again to obtain carbon concentrate II;
2) carrying out suction filtration, drying and acid leaching on the tailings obtained by flotation in the step 1), oxidizing with hydrogen peroxide, and filtering to respectively obtain filtrate I and filter residue I;
3) adding a sodium hydroxide solution into the filtrate I obtained in the step 2) to adjust the pH value for iron precipitation, controlling the temperature to promote the generation of an iron hydroxide precipitate, and filtering to obtain a filtrate II and a filter residue II; washing the obtained filter residue II to obtain iron slag, and roasting to obtain iron ore concentrate;
4) continuously adding sodium hydroxide solution into the filtrate II obtained in the step 3) to adjust the pH value for zinc precipitation, and filtering to obtain filtrate III and filter residue III; washing the obtained filter residue III to obtain zinc slag, and roasting to obtain zinc concentrate;
5) adding dilute sulfuric acid into the filtrate III obtained in the step 4) to adjust the pH value for precipitating aluminum, filtering and washing the precipitate for multiple times to obtain zinc slag, and roasting to obtain aluminum concentrate;
6) carrying out suction filtration and drying on the filter residue I obtained in the step 2), adding a saturated ammonium acetate solution, stirring, filtering, adding a sodium carbonate solution into the filtrate to adjust the pH value, and precipitating lead; filtering and washing the precipitate to obtain lead slag, and roasting to obtain lead concentrate.
In the scheme, the granularity of the gas ash in the step 1) is less than 0.1 μm.
In the scheme, the collecting agent adopted in the step 1) is diesel oil, and the foaming agent is 2# oil.
In the scheme, the concentration of the ore pulp in the step 1) is 4-6%, and the pH value is 10-13.
In the scheme, in the flotation process in the step 1), the amount of the collecting agent relative to the gas mud is 6-8 wt%; the amount of the foaming agent is 3-4 wt% relative to the amount of the gas mud.
Preferably, the flotation process comprises one-time roughing and three-time concentrating; the acid cleaning of the carbon concentrate I adopts dilute sulfuric acid, the mass concentration of the dilute sulfuric acid is 20-25%, and the mass ratio of the dilute sulfuric acid to the carbon concentrate I is (5-10): 1.
In the scheme, dilute sulfuric acid with the mass concentration of 20-25% is adopted in the acid leaching step in the step 2); the mass ratio of the acid to the gas mud is (12-13): 1; the volume ratio of the hydrogen peroxide to the dilute sulfuric acid is 1 (1.2-1.3), and the concentration of the hydrogen peroxide is 25-35%; (ii) a The temperature adopted by acid leaching is controlled to be 60-70 ℃.
Preferably, the acid leaching and hydrogen peroxide oxidation steps comprise the steps of performing suction filtration and drying on the tailings obtained by flotation, adding dilute sulfuric acid and hydrogen peroxide, and reacting for 0.5-1 h under stirring at the temperature of 60-70 ℃.
In the scheme, the pH value in the step 3) is 2.3-4.0, the temperature is controlled to be 65-70 ℃, the roasting temperature is 850-880 ℃, and the roasting time is not less than 2 hours.
In the scheme, the pH value in the step 4) is 8.0-10.5, the roasting temperature is 125-150 ℃, and the roasting time is not less than 2 hours.
In the scheme, the pH value in the step 5) is 5.2-7.8, the roasting temperature is 450-470 ℃, and the roasting time is not less than 2 hours.
In the scheme, the mass ratio of the saturated ammonium acetate solution to the filter residue I in the step 6) is (3-4) to 1; the concentration of the sodium carbonate solution is 1.0-1.5 mol/L.
In the scheme, the pH value in the step 6) is 7.2-8.7, the roasting temperature is 420-450 ℃, and the roasting time is not less than 2 hours.
The principle of the invention is as follows:
the method adopts a flotation process to efficiently recover the carbon concentrate, and further removes metal oxide impurities mixed in the carbon concentrate by combining an acid washing-suction filtration process, so that the grade of the carbon concentrate is improved, the dosage of medicaments such as hydrogen peroxide and the like in the acid leaching oxidation process can be effectively reduced, the process cost is reduced, and the recovery rate and the grade of subsequent iron can be ensured; then, dissolving valuable metals into metal cations by acid leaching of tailings obtained by flotation at the water bath temperature of 60-70 ℃, converting ferrous ions in gas mud into ferric ions by hydrogen peroxide, oxidizing zinc sulfide into soluble zinc sulfate, oxidizing lead sulfide into insoluble lead sulfate, and filtering to separate metal cations except lead from insoluble impurities; dissolving lead sulfate in insoluble impurities into lead acetate by using saturated ammonium acetate; according to the difference of pH ranges of precipitates generated by hydroxides of iron, zinc and aluminum and basic lead carbonate, respectively extracting valuable or harmful metal elements such as iron, zinc, aluminum and lead by fractional precipitation; because the pH values of the aluminum and zinc precipitates are relatively close and the pH range of the aluminum precipitates is relatively narrow, the aluminum and zinc are separated by adopting a reverse adjustment method, and the problems of low fault tolerance rate, low product recovery rate and purity (zinc is easily doped in aluminum and the like) and the like of the traditional pH adjustment method are effectively avoided; finally, the high-efficiency and comprehensive recycling and harmless treatment of carbon, iron, zinc, aluminum and lead in the gas mud are realized.
Compared with the prior art, the invention has the beneficial effects that:
1) compared with the existing blast furnace gas mud treatment method, the method firstly recovers carbon, avoids the problems that carbon is consumed by hydrogen peroxide in the acid leaching oxidation process and the like, can effectively improve the recovery rate and grade of carbon and iron in the blast furnace gas mud, reduces the consumption of hydrogen peroxide and saves the treatment cost.
2) The method further realizes the recovery of zinc, aluminum and lead, particularly better avoids the problems of low fault tolerance rate and the like of the traditional pH adjusting method during the recovery of aluminum and zinc, has larger operable space, higher efficiency and more convenience, effectively reduces the production cost of enterprises, increases the economic efficiency of the enterprises, and better realizes the resource recycling of valuable and harmful elements in metallurgical solid wastes.
3) Can effectively avoid the pollution of heavy metal elements to the environment in the process of stockpiling or low-value utilization of the blast furnace gas mud, and has important economic and environmental benefits.
Drawings
Fig. 1 is a process flow diagram of a comprehensive recycling method of multiple valuable elements in blast furnace gas sludge as described in example 1.
Detailed Description
In order to better understand the present invention, the following examples are further provided to illustrate the present invention, but the present invention is not limited to the following examples.
Example 1
A comprehensive recycling method of multiple valuable elements in blast furnace gas mud is disclosed, the process flow chart of which is shown in figure 1, and the method specifically comprises the following steps:
1) putting blast furnace gas mud into a drying oven, drying at 120 ℃ for 36h, and grinding for 2h by using a planetary ball mill to obtain gas ash with the particle size of less than 0.1 mu m;
2) adding water into the gas ash obtained in the step 1) to make the concentration of the ore pulp be 6wt%, uniformly stirring, and adding collecting agent diesel oil and foaming agent 2# oil (the using amount of the collecting agent relative to the gas mud is 7 wt%; the amount of the foaming agent is 4wt percent relative to the gas mud), then quicklime is added to adjust the pH value of the obtained slurry to 11, and the carbon slurry and the tailings are collected by once roughing and three times of fine selection; carrying out suction filtration on the collected carbon slurry, drying at 120 ℃ for 24 hours, adding dilute sulfuric acid with the mass concentration of 20 wt% and the mass being 5 times that of the carbon concentrate for pickling, and carrying out suction filtration to obtain carbon concentrate;
3) performing suction filtration on the tailings obtained by flotation in the step 1), drying at 120 ℃ for 24 hours, then placing the tailings into a beaker, adding 20 wt% of dilute sulfuric acid with the mass concentration being 12 times of the mass of the gas mud, then adding 30 wt% of hydrogen peroxide with the volume ratio of 1:1.2 to the sulfuric acid, placing the beaker into a constant-temperature water bath kettle, controlling the water bath temperature at 60 ℃, rotating the electric stirrer at 150r/min, simultaneously performing acid leaching and oxidation for 1 hour, and filtering to obtain filtrate I and filter residue I respectively;
4) dropwise adding a sodium hydroxide solution into the filtrate I obtained in the step 3) to adjust the pH value to 4.0, and controlling the temperature to be 70 ℃ (otherwise, ferric hydroxide colloid is generated, is difficult to separate and remove, and has a large influence on the extraction result), so as to promote the generation of ferric hydroxide precipitate; taking down the beaker, standing for 10 minutes, filtering and washing the precipitate for multiple times to obtain filter residue II (iron slag) and filtrate II; putting the iron slag into a muffle furnace, and roasting for 2 hours at 850 ℃ to obtain iron ore concentrate;
5) dropwise adding sodium hydroxide into the filtrate II obtained in the step 4) to adjust the pH value to 10.0, and promoting the generation of zinc hydroxide precipitate; taking down the beaker, standing for 10 minutes, filtering and washing the precipitate for multiple times to obtain filtrate III and filter residue III (zinc residue); putting the zinc slag into a muffle furnace, and roasting for 2 hours at 150 ℃ to obtain zinc concentrate;
6) adding dilute sulfuric acid into the filtrate III obtained in the step 5) to adjust the pH value to 7.0, and promoting the generation of aluminum hydroxide precipitate; taking down the beaker, standing for 10 minutes, filtering, washing the precipitate for multiple times by using a decantation method, and collecting the obtained filter residue (aluminum slag) and filtrate IV; putting the zinc slag into a muffle furnace, and roasting for 2 hours at 450 ℃ to obtain aluminum concentrate;
7) carrying out suction filtration and drying on the filter residue I obtained in the step 3), adding a saturated ammonium acetate solution, wherein the mass ratio of the ammonium acetate solution to the filter residue I is 4:1, stirring and filtering, adding a sodium carbonate solution with the concentration of 1mol/L into the filtrate, adjusting the pH value to 8.2, and promoting the generation of basic lead carbonate precipitate; standing for 10 minutes, filtering, and washing the precipitate for multiple times by using a decantation method to obtain lead slag and filtrate V; putting the lead slag into a muffle furnace, and roasting for 2 hours at the temperature of 420 ℃ to obtain lead concentrate;
8) and (3) combining the filtrate IV obtained in the step 6) and the filtrate V obtained in the step 7), dropwise adding dilute sulfuric acid with the mass concentration of 20-25%, adjusting the pH value to 7.0 to be neutral, or directly returning to the step 3) to be used as a leaching agent.
In the embodiment, the adopted blast furnace gas mud is provided by a certain steel plant in Hubei; ICP component analysis shows that the elements contained in the blast furnace gas mud comprise carbon, iron, zinc, aluminum, silicon, sulfur, calcium, potassium, sodium, lead and the like; XRD phase analysis shows that zinc and lead in the blast furnace gas mud mainly exist in the forms of zinc sulfide and lead sulfide; the major recoverable value components are shown in table 1.
TABLE 1 information of the main recoverable valuable components in the blast furnace gas sludge of certain iron and steel works in Hubei
| Composition (I) | C | TFe | ZnO | Al2O3 | PbO |
| Content (wt%) | 51.025 | 15.902 | 7.99 | 4.171 | 1.544 |
Tests show that the recovery rate of carbon in the blast furnace gas mud is 89.5 percent by adopting the recovery process in the embodiment; the iron recovery rate is 76.8%; the zinc recovery rate is 89.4%; the recovery rate of aluminum is 76.1 percent; the lead recovery rate is 72.1%; from the further analysis of the composition, the recovered carbon was 89.2% pure; the purity of iron is 88.7%; the purity of zinc is 92.4%; the purity of the aluminum is 84.4%; the purity of the lead is 77.4 percent, the high-efficiency recovery of carbon, iron, zinc, aluminum and lead can be realized, and the purity is better.
Example 2
A comprehensive recycling method of multiple valuable elements in blast furnace gas mud specifically comprises the following steps:
1) putting blast furnace gas mud into a drying oven, drying at 120 ℃ for 36h, and grinding for 2h by using a planetary ball mill to obtain gas ash with the particle size of less than 0.1 mu m;
2) adding water into the gas ash obtained in the step 1) to enable the concentration of ore pulp to be 6%, uniformly stirring, and adding collecting agent diesel oil and foaming agent No. 2 oil into the obtained pulp, wherein the using amount of the collecting agent relative to the gas mud is 7% by mass; the amount of the foaming agent relative to the gas mud is 4 percent by mass. Adding quicklime to adjust the pH value of the obtained slurry to 12, performing primary roughing and three times of fine selection, and collecting carbon slurry and tailings; carrying out suction filtration on the collected carbon slurry, drying at 120 ℃ for 24 hours, adding dilute sulfuric acid with the mass concentration of 20 wt% and the mass being 5 times that of the carbon concentrate for pickling, and carrying out suction filtration to obtain carbon concentrate;
3) carrying out suction filtration on the tailings obtained by flotation in the step 1), drying at 120 ℃ for 24 hours, then placing the tailings into a beaker, adding 25 wt% dilute sulfuric acid with the concentration being 12 times of the mass of the gas ash, then adding 30 wt% hydrogen peroxide with the volume ratio of 1:1.3 to the sulfuric acid, placing the beaker into a constant-temperature water bath kettle, controlling the water bath temperature to be 65 ℃, rotating the electric stirrer at 150r/min, simultaneously carrying out acid leaching and oxidation for 1 hour, and filtering to obtain filtrate I and filter residue I respectively;
4) dropwise adding sodium hydroxide into the filtrate I obtained in the step 2) to adjust the pH value to 3.6, controlling the temperature to be 70 ℃, and promoting the generation of ferric hydroxide precipitate; taking down the beaker, standing for 10 minutes, filtering and washing the precipitate for multiple times to obtain filter residue II (iron slag) and filtrate II; putting the iron slag into a muffle furnace, and roasting for 1 hour at 880 ℃ to obtain iron ore concentrate;
5) dropwise adding sodium hydroxide into the filtrate II obtained in the step 4) to adjust the pH value to 10.2, and promoting the generation of zinc hydroxide precipitate; taking down the beaker, standing for 10 minutes, filtering and washing the precipitate for multiple times to obtain filtrate III and filter residue III (zinc residue); putting the zinc slag into a muffle furnace, and roasting for 2 hours at 150 ℃ to obtain zinc concentrate;
6) adding dilute sulfuric acid into the filtrate III obtained in the step 5) to adjust the pH value to 6.7, and promoting the generation of aluminum hydroxide precipitate; taking down the beaker, standing for 10 minutes, filtering, washing the precipitate for multiple times by using a decantation method, and collecting the obtained filter residue (aluminum slag) and filtrate IV; putting the zinc slag into a muffle furnace, and roasting for 2 hours at 470 ℃ to obtain aluminum concentrate;
7) carrying out suction filtration and drying on the filter residue I obtained in the step 3), and pouring a saturated ammonium acetate solution into the filter residue I, wherein the mass ratio of the ammonium acetate solution to the filter residue I is 4:1, stirring and filtering, adding a sodium carbonate solution with the concentration of 1mol/L into the filtrate, adjusting the pH value to 8.0, and promoting the generation of basic lead carbonate precipitate; standing for 10 minutes, filtering, and washing the precipitate for multiple times by using a decantation method to obtain lead slag and filtrate V; and (3) putting the lead slag into a muffle furnace, and roasting for 1 hour at the temperature of 450 ℃ to obtain lead concentrate.
8) And (3) combining the filtrate IV obtained in the step 6) and the filtrate V obtained in the step 7), dropwise adding dilute sulfuric acid with the mass concentration of 20-25%, adjusting the pH value to 7.0 to be neutral, or directly returning to the step 3) to be used as a leaching agent.
In the embodiment, the adopted blast furnace gas mud is provided by a certain steel plant in Hubei; ICP component analysis shows that the elements contained in the blast furnace gas mud comprise carbon, iron, zinc, aluminum, silicon, sulfur, calcium, potassium, sodium, lead and the like; XRD phase analysis shows that zinc and lead in the blast furnace gas mud mainly exist in the forms of zinc sulfide and lead sulfide; the major recoverable value components are shown in table 2.
TABLE 2 information of the main recoverable valuable components in the blast furnace gas sludge of certain iron and steel works in Hubei
| Composition (I) | C | TFe | ZnO | Al2O3 | PbO |
| Content (wt%) | 43.663 | 16.741 | 5.844 | 5.674 | 1.417 |
Tests show that the recovery rate of carbon in the blast furnace gas mud is 86.3 percent by adopting the recovery process in the embodiment; the iron recovery rate is 80.2%; the zinc recovery rate is 88.1 percent; the aluminum recovery rate is 79.2%; the lead recovery rate is 70.4%; from the further compositional analysis results, the recovered carbon grade was 88.7%; the purity of iron is 86.5%; the purity of the zinc is 90.6 percent; the purity of the aluminum is 81.6 percent; the purity of the lead is 80.1 percent, the high-efficiency recovery of carbon, iron, zinc, aluminum and lead can be realized, and the grade is good.
Comparative example 1
A method for comprehensively recovering multiple valuable elements from blast furnace gas mud takes the blast furnace gas mud adopted in example 1 as a sample, and specifically comprises the following steps:
1) putting blast furnace gas mud into a drying oven, drying at 120 ℃ for 36h, and grinding for 2h by using a planetary ball mill to obtain gas ash with the particle size of less than 0.1 um;
2) adding water into the gas ash obtained in the step 1), stirring to make the concentration of the ore pulp be 6%, adding diesel oil as a collecting agent, and adding 2# oil as a foaming agent (the using amount of the collecting agent relative to the gas mud is 7 wt%; the amount of the foaming agent is 4wt% relative to the gas mud), adding quicklime, adjusting the pH to 10, performing primary roughing and primary fine selection, and collecting carbon slurry and tailings; carrying out suction filtration on the collected carbon slurry, and drying at 120 ℃ for 24 hours to obtain carbon concentrate;
3) filtering the tailings left in the step 1), drying at 120 ℃ for 24 hours, then placing into a beaker, adding 20% dilute sulfuric acid with the volume 5 times that of the gas ash, adding 30% hydrogen peroxide with the volume ratio of 1:30 to the sulfuric acid, placing the beaker into a constant-temperature water bath kettle, controlling the water bath temperature at 60 ℃, the rotation speed of an electric stirrer at 150r/min, carrying out acid leaching and oxidation for 30min, filtering, and respectively storing filtrate and filter residue;
4) dropwise adding ammonia water into the filtrate prepared in the step 3), adjusting the pH value to 4.0, controlling the temperature to be 70 ℃ to generate ferric hydroxide precipitate, taking down the beaker, standing for 10 minutes, filtering and washing the precipitate for multiple times to obtain iron slag and iron-removed liquid; putting the iron slag into a muffle furnace, and roasting for 2 hours at 850 ℃ to obtain iron ore concentrate;
5) continuously dropwise adding ammonia water into the filtrate subjected to the filtration and deironing in the step 4), and adjusting the pH value to 4.8 to generate aluminum hydroxide precipitate; taking down the beaker, standing for 10 minutes, filtering and washing the precipitate for multiple times to obtain aluminum slag and a liquid after aluminum removal; placing the aluminum slag into a muffle furnace, and roasting for 2 hours at 450 ℃ to obtain aluminum concentrate;
6) adding sodium hydroxide into the filtrate obtained after the filtration and the aluminum removal in the step 5), and adjusting the pH value to 8.0 to generate zinc hydroxide precipitate; taking down the beaker, standing for 10 minutes, filtering and washing the precipitate for multiple times by using a decantation method to obtain zinc slag and a liquid after zinc removal. Putting the zinc slag into a muffle furnace, and roasting for 2 hours at 125 ℃ to obtain zinc concentrate;
7) carrying out suction filtration and drying on the filter residue obtained in the step 3), pouring a saturated ammonium acetate solution into the filter residue I, wherein the mass ratio of the ammonium acetate solution to the filter residue I is 4:1, stirring and filtering the mixture, adding a sodium carbonate solution with the concentration of 1mol/L into the filtrate, and adjusting the pH value to 8.3 to generate a basic lead carbonate precipitate; standing for 10 minutes, filtering and washing the precipitate for multiple times by using a decantation method to obtain lead slag and lead-removed liquid; putting the lead slag into a muffle furnace, and roasting for 2 hours at the temperature of 420 ℃ to obtain lead concentrate;
8) and (3) combining the filtrate obtained after zinc removal by filtration in the step 6) and the filtrate obtained after lead removal by filtration in the step 7), dropwise adding 15-20% of dilute sulfuric acid, and adjusting the pH value to 6.5-7.0 to be neutral, or directly returning to the step 3) to be used as a leaching agent.
Tests show that by adopting the recovery process disclosed by the embodiment, the recovery rate of carbon in the blast furnace gas mud is 82.5%; the iron recovery rate is 68.3%; the zinc recovery rate is 86.1 percent; the aluminum recovery rate is 81.2%; lead recovery was 70%; from the further analysis results, the recovered carbon had a grade of 81.3% and contained 10.2% of iron; the purity of iron is 81.2%; the purity of the zinc is 85.6 percent, and the zinc contains 3.3 percent of iron; the purity of the aluminum is 70.7 percent, and the aluminum contains 4.7 percent of iron and 5.1 percent of zinc; the lead purity was 73.1%. Compared with the example 1, the recovery rate and the purity of carbon, iron and aluminum are obviously reduced, and the recovery rate, the concentration and the like of zinc also have a certain tendency to be reduced.
The invention can be realized by all the listed raw materials, and the invention can be realized by the upper and lower limit values and interval values of all the raw materials; the examples are not to be construed as limiting the scope of the invention. The upper and lower limit values and interval values of the process parameters can realize the invention, and the embodiments are not listed.
Claims (7)
1. A method for comprehensively recycling carbon, iron, zinc, aluminum and lead from blast furnace gas mud is characterized in that firstly, a flotation-acid cleaning-suction filtration method is used for recycling carbon, and then a fractional precipitation method is used for sequentially recycling iron, zinc, aluminum and lead elements;
the flotation-acid washing-suction filtration method comprises the following steps: carrying out flotation on the blast furnace gas mud to obtain carbon concentrate I and tailings, and then carrying out acid pickling and suction filtration on the carbon concentrate I to obtain carbon concentrate II;
the flotation step comprises: drying and grinding blast furnace gas mud, adding water, stirring to prepare ore pulp, adjusting the pH value of the obtained ore pulp, adding a collecting agent and a foaming agent, performing direct flotation to obtain carbon pulp and tailings, collecting the obtained carbon pulp, performing suction filtration and drying to obtain carbon concentrate I;
the fractional precipitation method comprises the following steps:
1) carrying out suction filtration, drying and acid leaching on the tailings, oxidizing the tailings with hydrogen peroxide, and filtering to obtain filtrate I and filter residue I; dilute sulfuric acid is adopted in the acid leaching step;
2) adding a sodium hydroxide solution into the filtrate I obtained in the step 1) to adjust the pH value for iron precipitation, controlling the temperature to promote the generation of precipitate, and filtering to obtain a filtrate II and a filter residue II; washing the obtained filter residue II to obtain iron slag, and roasting to obtain iron ore concentrate;
3) continuously adding sodium hydroxide solution into the filtrate II obtained in the step 2) to adjust the pH value for zinc precipitation, and filtering to obtain filtrate III and filter residue III; washing the obtained filter residue III to obtain zinc slag, and roasting to obtain zinc concentrate;
4) adding dilute sulfuric acid into the filtrate III obtained in the step 3) to adjust the pH value for precipitating aluminum, filtering, washing and precipitating for multiple times to obtain aluminum slag, and roasting to obtain aluminum concentrate;
5) carrying out suction filtration and drying on the filter residue I obtained in the step 1), adding a saturated ammonium acetate solution, stirring, filtering, adding a sodium carbonate solution into the filtrate to adjust the pH value, and precipitating lead; filtering, washing and precipitating to obtain lead slag, and roasting to obtain lead concentrate.
2. The method of claim 1, wherein the adopted collector is diesel oil, and the foaming agent is 2# oil; the concentration of the ore pulp is 4-6 wt%, and the pH value is 10-13.
3. The method according to claim 1, wherein in the direct flotation process, the amount of the collecting agent is 6-8 wt% relative to the blast furnace gas mud; the amount of the foaming agent is 3-4 wt% relative to the blast furnace gas mud.
4. The method according to claim 1, characterized in that the forward flotation process comprises a rougher and a cleaner step; the acid washing step of the carbon concentrate I adopts dilute sulfuric acid, the mass concentration of the dilute sulfuric acid is 20-25%, and the mass ratio of the dilute sulfuric acid to the carbon concentrate I is (5-10): 1.
5. The method according to claim 1, wherein the mass concentration of the dilute sulfuric acid in the step 1) is 20-25%, and the mass ratio of the dilute sulfuric acid to the blast furnace gas mud is (12-13): 1; the volume ratio of the hydrogen peroxide to the dilute sulfuric acid is 1 (1.2-1.3); the temperature adopted for acid leaching is 60-70 ℃.
6. The method according to claim 1, wherein the temperature in step 2) is controlled to be 65-70 ℃, and the pH value is 2.3-4.0; the pH value in the step 3) is 8.0-10.5; the pH value in the step 4) is 5.2-7.8.
7. The method according to claim 1, wherein the mass ratio of the ammonium acetate solution to the filter residue I in the step 5) is (3-4): 1; the pH value is 7.2-8.7.
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