CN111893308A - Method for comprehensively utilizing red mud without tailings - Google Patents

Method for comprehensively utilizing red mud without tailings Download PDF

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CN111893308A
CN111893308A CN202010904906.1A CN202010904906A CN111893308A CN 111893308 A CN111893308 A CN 111893308A CN 202010904906 A CN202010904906 A CN 202010904906A CN 111893308 A CN111893308 A CN 111893308A
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red mud
leaching
slag
pellets
tailings
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朱德庆
潘建
郭正启
杨聪聪
李思唯
李启厚
李紫云
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Central South University
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Central South University
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    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
    • C22B7/001—Dry processes
    • C—CHEMISTRY; METALLURGY
    • C21—METALLURGY OF IRON
    • C21B—MANUFACTURE OF IRON OR STEEL
    • C21B11/00—Making pig-iron other than in blast furnaces
    • C21B11/10—Making pig-iron other than in blast furnaces in electric furnaces
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B1/00—Preliminary treatment of ores or scrap
    • C22B1/02—Roasting processes
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B1/00—Preliminary treatment of ores or scrap
    • C22B1/14—Agglomerating; Briquetting; Binding; Granulating
    • C22B1/16—Sintering; Agglomerating
    • C22B1/20—Sintering; Agglomerating in sintering machines with movable grates
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B1/00—Preliminary treatment of ores or scrap
    • C22B1/14—Agglomerating; Briquetting; Binding; Granulating
    • C22B1/24—Binding; Briquetting ; Granulating
    • C22B1/2406—Binding; Briquetting ; Granulating pelletizing
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B1/00—Preliminary treatment of ores or scrap
    • C22B1/14—Agglomerating; Briquetting; Binding; Granulating
    • C22B1/24—Binding; Briquetting ; Granulating
    • C22B1/242—Binding; Briquetting ; Granulating with binders
    • C22B1/243—Binding; Briquetting ; Granulating with binders inorganic
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
    • C22B3/06—Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic acid solutions, e.g. with acids generated in situ; in inorganic salt solutions other than ammonium salt solutions
    • C22B3/10—Hydrochloric acid, other halogenated acids or salts thereof
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
    • C22B3/12—Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic alkaline solutions
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
    • C22B7/006—Wet processes
    • C22B7/007—Wet processes by acid leaching
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
    • C22B7/006—Wet processes
    • C22B7/008—Wet processes by an alkaline or ammoniacal leaching
    • Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
    • Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
    • Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
    • Y02P10/00—Technologies related to metal processing
    • Y02P10/20—Recycling

Abstract

The invention provides a method for comprehensively utilizing red mud without tailings, which comprises the following steps: adding an additive and a binder into the pretreated red mud, mixing, pelletizing and drying to obtain red mud dry balls; preheating and roasting the dry pellets to obtain preheated pellets; adding a reducing agent into the preheated pellets for pre-reduction treatment to obtain pre-reduced pellets; adding a reducing agent into the obtained pre-reduced pellets, mixing, and then carrying out melt separation treatment in an electric furnace to obtain pig iron and melt separation slag; mixing the obtained molten slag with sodium carbonate and limestone, and then carrying out slag modification treatment to obtain modified slag; performing ball milling and alkaline leaching on the modified slag to obtain alkaline leaching slag and aluminate solution; and (3) carrying out dilute acid leaching desiliconization and aluminum leaching on the alkaline leaching residue to obtain acid leaching solution and perovskite concentrate. The method can realize the high-efficiency separation and recovery of main valuable elements such as iron, aluminum, silicon, titanium and the like in the red mud, has low energy consumption, does not pollute the environment, and is easy to realize industrial production.

Description

Method for comprehensively utilizing red mud without tailings
Technical Field
The invention relates to the field of comprehensive utilization of red mud, in particular to a method for comprehensively utilizing red mud without tailings.
Background
The red mud is residue generated when bauxite is leached by strong alkali in the production process of alumina, and 1-1.6 tons of red mud can be generated when one ton of alumina is produced. At present, the amount of red mud discharged in the world is 1.2 hundred million tons per year. China is a large alumina producing country, and only 2012, China produces nearly 5000 million tons of red mud. In addition, the red mud has high contents of alkali metals, heavy metals and radioactive elements, and if the method of stacking, landfill and the like is adopted, certain harm is caused to the environment, and a large amount of metal resources are wasted. Red mud generally contains Fe2O3，Al2O3，Na2O and TiO2When valuable metals are recycled independently, one element cannot effectively solve the problems of process economy and huge red mud stacking amount, so that the comprehensive utilization of the red mud can be effectively realized only by adopting a technology of jointly recycling various metals.
The recovery of valuable metals in the red mud is mainly aimed at the comprehensive utilization of iron and aluminum, and the recovery of scandium is also considered in part of processes. At present, the following processes are mainly studied.
The direct magnetic separation iron recovery process comprises the following steps: li Chao Xiang reports that Pingguo aluminum adopts an SLON type high gradient magnetic separator to directly magnetically separate red mud, and adopts a primary coarse-primary fine full magnetic separation method, and the process mainly has the problems that the recovery rate of iron is low, and aluminum and titanium are not effectively recovered.
Direct reduction sintering-magnetic separation: mixing and grinding red mud, limestone, sodium carbonate and coal, then carrying out direct reduction sintering at 800-1000 ℃, crushing and leaching sintered blocks to obtain 89% aluminum leaching, carrying out high-gradient magnetic separation on the slag, smelting the obtained magnetic substances at high temperature, and leaching non-magnetic substances to extract titanium by adopting sulfuric acid. The process can obtain 70% Ti leaching, and has the problems of low iron recovery rate and high acid consumption.
Reduction smelting-slag leaching process: the red mud and carbon are mixed and smelted, so that iron in the red mud is separated out, and the scandium-iron separation effect is achieved. The slag is subjected to alkaline leaching or acid leaching to recover scandium, and the process has the problems of low aluminum recovery rate and high alkali consumption and acid consumption.
The process of using the melting reduction-furnace slag as the raw material/refractory material of the compound fertilizer comprises the following steps: for example, according to the Chinese invention patent CN110093471A, red mud and water are stirred, size-mixed and then subjected to filter pressing to obtain pre-dealkalized red mud, the pre-dealkalized red mud is mixed with a reducing agent and a calcium-containing additive to obtain pellets, then the pellets are subjected to high-temperature melting-magnetic separation to obtain reduced iron and furnace slag, and the furnace slag is used as a raw material/refractory material of a compound fertilizer.
Sulfating roasting-leaching process: for example, the process disclosed in the invention patent CN111137907A comprises the steps of mixing Bayer red mud with concentrated sulfuric acid, completely salinizing at low temperature to obtain sulfated red mud, roasting the sulfated red mud at 900 ℃ of 600-.
The direct red mud leaching process comprises the following steps: for example, the Chinese invention patent CN107267757A directly leaches the red mud by hydrochloric acid, and obtains Al by evaporation roasting2O3And Fe2O3Is leached outThe slag is leached with sulphuric acid to convert the titanium to titanyl sulfate. The process has the problems of low recovery rate of iron and aluminum and high acid consumption.
The existing red mud recovery technology is difficult to realize the efficient separation and recovery of various valuable metals, and has the problems of high energy consumption and low red mud utilization rate, so that a method with high efficiency and high red mud recovery utilization rate is urgently needed to relieve the red mud treatment pressure.
Disclosure of Invention
The invention provides a method for comprehensively utilizing red mud without tailings, and aims to provide a process which has low energy consumption, does not pollute the environment, is easy to realize industrial production, and realizes the high-efficiency separation and recovery of various valuable metals in the red mud.
In order to achieve the aim, the invention provides a method for comprehensively utilizing red mud without tailings, which comprises the following steps:
s1, pelletizing:
adding an additive and a binder into the pretreated red mud, mixing, pelletizing and drying to obtain red mud dry balls;
s2, preheating and roasting:
preheating and consolidating the dry pellets obtained in the step S1 to obtain preheated pellets;
s3, pre-reduction treatment:
adding reducing coal into the preheated pellets obtained in the S2 for pre-reduction treatment to obtain pre-reduced pellets; wherein the pre-reduction temperature is 800-1300 ℃, and the pre-reduction time is 15-180 min;
s4, melt separation treatment:
adding a reducing agent into the pre-reduced pellets obtained in the step S3, mixing, and performing melt separation treatment to obtain pig iron and melt separation slag; wherein the melt separation temperature is 1450-1600 ℃, and the melt separation time is 20-40 min;
s5, modification treatment:
mixing the molten slag obtained in the step S4 with sodium carbonate and limestone, and then carrying out modification treatment to obtain modified slag; wherein the modification temperature is 900-1300 ℃, and the modification time is 15-60 min;
s6, ball milling and alkaline leaching:
carrying out ball milling and alkaline leaching on the modified slag obtained in the step S5 to obtain alkaline leaching slag and aluminate solution;
s7, acid leaching:
and (4) carrying out acid leaching on the alkaline leaching residue obtained in the step (S6) to obtain acid leaching solution and perovskite concentrate.
Preferably, the red mud comprises the following components in percentage by mass: 30-65% Fe2O35 to 30% of Al2O30 to 5% of Na20 to 10% of TiO25 to 20% of SiO2And the balance being CaO and other impurities.
Preferably, in the step S1, the pretreatment is to mix the red mud with water and then perform high-pressure roll milling; wherein the mass of the water accounts for 6.5-7.5% of the mass of the red mud; the pressure of the high-pressure roller mill is 250-292 kN/m.
Preferably, in S1, the additive is one of limestone, quicklime and slaked lime, and the binder is bentonite.
Preferably, in the step S1, the drying temperature is 200-300 ℃, and the drying time is more than 30 min.
Preferably, the preheating roasting temperature is 950-1200 ℃, the time is 5-30 min, and the roasting wind speed is 2.0-2.6 m/s.
Preferably, the mass ratio of the reduced coal to the preheated pellets in the pre-reduction treatment is 0.5-2: 1.
Preferably, the mass ratio of the reducing agent to the pre-reduced pellets in the melting separation treatment is 0.05-0.2: 1, and the reducing agent is a carbon-containing fuel.
Preferably, in the step S6, the ball milling is specifically to configure the modified slag into ore pulp with a concentration of 30-80%, and then grind the obtained ore pulp into ball milling slag, wherein particles with a particle size of less than 0.074mm in the ball milling slag account for 80-95%.
Preferably, in the step S6, the alkaline leaching is specifically to add the ball milling residue into an alkaline solution according to a liquid-solid mass ratio of 1-10: 1, wherein the concentration of the alkaline solution is 20-160 g/L; wherein the alkaline leaching temperature is 60-95 ℃, and the alkaline leaching time is 20-120 min.
Preferably, in the step S7, the acid leaching is specifically to add the alkali leaching residue into an acid liquor according to a liquid-solid mass ratio of 1-10: 1, wherein the concentration of the acid liquor is 10-392 g/L; wherein the acid leaching temperature is 25-95 ℃, and the acid leaching time is 20-120 min.
The method comprises the steps of treating red mud by adopting a pre-reduction-melting separation method, carrying out pre-reduction at 800-1300 ℃ by taking coal or coke as a reducing agent, reducing iron oxide in the red mud into metallic iron, and then mixing pre-reduced pellets with limestone or quicklime and the reducing agent (adjusting the slag type of the pre-reduced pellets by adding CaO to promote most of SiO2Form 2 CaO. SiO with CaO2) And carrying out melt separation at 1450-1600 ℃ to obtain pig iron, so that the aluminum oxide and the titanium oxide in a chemical combination state enter slag, and iron and aluminum titanium are separated. Mixing the molten slag with sodium carbonate and limestone at 1100-1300 ℃ for modification, and carrying out alkaline leaching on the modified slag to obtain a sodium aluminate solution and alkaline leaching slag, thereby realizing the separation of aluminum and titanium. The alkaline leaching residue is subjected to acid leaching to obtain perovskite, the acid leaching solution contains a certain amount of sodium silicate and can be used for producing water glass and the like, titanium is recycled through acid leaching, impurities such as silicon are utilized, and the effects of harmless treatment of alumina red mud and comprehensive recycling of valuable elements are achieved.
The scheme of the invention has the following beneficial effects:
the method provided by the invention realizes harmless treatment of red mud and comprehensive recycling of valuable elements, and the pig iron obtained by melting can be used for electric furnace steelmaking, wherein the recovery rate of iron is more than 94%, and the grade of the pig iron is more than 92%; the alkaline leaching residue obtained by alkaline leaching contains a certain amount of elements such as titanium, silicon, calcium, magnesium and the like, and aluminum is recovered by a sodium aluminate solution, wherein the leaching rate of the aluminum reaches more than 80 percent, so that the separation of the aluminum and the titanium is realized; the perovskite concentrate obtained by acid leaching can be used for preparing light energy materials, the recovery rate of titanium is more than 85%, and the obtained acid leaching solution can be used for preparing water glass.
The method provided by the invention realizes effective recovery of valuable metals such as iron, aluminum, titanium and the like, has low energy consumption, does not pollute the environment and is easy to realize industrial production.
Drawings
FIG. 1 is a process flow diagram of the present invention.
Detailed Description
In order to make the technical problems, technical solutions and advantages of the present invention more apparent, the following detailed description is given with reference to the accompanying drawings and specific embodiments.
Example 1
The mass percentages of the components in the embodiment are as follows: fe2O350% of Al2O315% of Na2O is 1.5%, TiO212 percent of the total weight of the alloy, and the balance of SiO2CaO, and the like; the red mud is used as raw material. Rolling once under 292kN/m pressure by a high-pressure roller mill, uniformly mixing the red mud after the roller mill with limestone and bentonite, pelletizing on a disc machine to obtain green pellets with the diameter of 8-16 mm and the water content of 16%, and drying the green pellets at the temperature of 200-300 ℃ for more than 30min to obtain dry pellets;
preheating the dry balls on a chain grate machine for consolidation, preheating at 950 deg.C for 15min to obtain balls with compressive strength of 500 (N.pieces)-1) The pellets are preheated.
Adding the preheated pellets and the reducing coal into a reducing furnace according to the coal mine mass ratio of 2:1, and reducing for 90min at 1050 ℃ to obtain pre-reduced pellets with the metallization rate of 80%;
adding a reducing agent coke according to 10 percent of the mass of the pre-reduced pellets, uniformly mixing the reducing agent and the pre-reduced pellets, putting the mixture into an electric furnace for smelting, smelting for 30min at the temperature of 1550 ℃, obtaining the pig iron with the grade of 95 percent, the iron recovery rate of 94 percent and Al in the molten slag2O3The content of Na is 48%2O content of 5.5%, TiO2And SiO2The content is 15 percent and 12 percent, iron basically enters pig iron in the melting process, and aluminum, silicon and titanium enter a slag phase to realize iron-aluminum separation;
and (4) melting slag modification. Considering that part of alumina and silica in the molten slag form aluminosilicate, it is converted into sodium aluminate and calcium silicate by adjusting the alkali ratio and calcium ratio. Adding sodium carbonate and limestone according to the alkali ratio of 1 and the calcium ratio of 2, uniformly mixing the three, oxidizing and roasting at 1300 ℃ for 15min, and taking out and naturally cooling; crushing the obtained modified slagAfter the granularity is less than 1mm, the ore pulp enters a ball mill for fine grinding, the concentration of the ore pulp is controlled to be 50 percent, and 90 percent of particles with the granularity of less than 0.074mm are ground; leaching ball-milling residue for 120min at 95 deg.C under the conditions of liquid-solid ratio of 10:1(mL: g) and sodium hydroxide concentration of 80g/L to obtain leaching residue containing Al2O3TiO content of 13%219% of SiO2The content is 30.4 percent, the leaching rate of the aluminum is 80 percent, the leaching rate of the titanium and the silicon is about 5 percent, the separation of the aluminum, the titanium and the silicon is realized, and the obtained sodium aluminate solution is used for producing the alumina;
leaching the alkaline leaching residue after aluminum extraction for 30min under the conditions that the liquid-solid ratio is 10:1(mL: g), the hydrochloric acid concentration is 87.6g/L and the leaching temperature is 30 ℃ to further remove aluminum and silicon, wherein the obtained leaching residue is perovskite concentrate containing TiO246% of SiO2The recovery rate of titanium is 0.45%, and the recovery rate of titanium is 95%, so that the separation of silicon and titanium is effectively realized. The obtained perovskite can be used for preparing photoelectric materials, and the leaching solution can be crystallized to obtain water glass.
Example 2
The mass percentages of the components in the embodiment are as follows: fe2O350% of Al2O315% of Na2O is 1.5%, TiO212 percent of the total weight of the alloy, and the balance of SiO2And impurities such as CaO, and the like, and the red mud is used as a raw material. Grinding the red mud subjected to the roller grinding once under the pressure of 292kN/m by using a high-pressure roller, uniformly mixing the red mud subjected to the roller grinding with limestone and bentonite, pelletizing on a disc machine to obtain green pellets, wherein the diameter of the obtained green pellets is 8-16 mm, the moisture of the green pellets is 16%, and drying the green pellets at the temperature of 200-300 ℃ for more than 30min to obtain dry pellets;
preheating the dry balls on a chain grate machine for consolidation, preheating at 1050 deg.C for 12min to obtain balls with compressive strength of 525 (N.N)-1) The pellets are preheated.
Adding the preheated pellets and the reducing coal into a reducing furnace according to the coal mine mass ratio of 2:1, and reducing for 60min at 1050 ℃ to obtain pre-reduced pellets with the metallization rate of 75%;
adding reducing agent coke according to 10 percent of the mass of the pre-reduced pellets, and reducingThe agent and the pre-reduced pellets are evenly mixed and put into an electric furnace for smelting, the smelting is carried out for 30min at the temperature of 1600 ℃, the obtained pig iron grade is 96 percent, the iron recovery rate is 98 percent, and Al in the molten slag is2O3The content of Na is 48%2O content of 5.5%, TiO2And SiO2The content is 15 percent and 12 percent, iron basically enters pig iron in the melting process, and aluminum, silicon and titanium enter a slag phase to realize iron-aluminum separation;
and (4) melting slag modification. Considering that part of alumina and silica in the molten slag form aluminosilicate, it is converted into sodium aluminate and calcium silicate by adjusting the alkali ratio and calcium ratio. Adding sodium carbonate and limestone according to the alkali ratio of 1 and the calcium ratio of 2, uniformly mixing the three, oxidizing and roasting at 1200 ℃ for 20min, and taking out and naturally cooling; crushing the obtained modified slag until the granularity is less than 1mm, and then, performing fine grinding in a ball mill, wherein the concentration of ore pulp is controlled to be 50%, and 90% of particles with the granularity of less than 0.074mm are ground; leaching ball-milling residue for 120min at 90 deg.C under the conditions of liquid-solid ratio of 10:1(mL: g) and sodium hydroxide concentration of 160g/L to obtain leaching residue containing Al2O3TiO content of 13%219% of SiO2The content is 18 percent, the leaching rate of aluminum is 85 percent, the leaching rate of titanium and silicon is about 5 percent, the separation of aluminum, titanium and silicon is realized, and the obtained sodium aluminate solution is used for producing alumina;
leaching the alkaline leaching residue after aluminum extraction for 30min under the conditions that the liquid-solid ratio is 10:1(mL: g), the hydrochloric acid concentration is 43.8g/L and the leaching temperature is 30 ℃ to further remove aluminum and silicon, wherein the obtained leaching residue is perovskite concentrate containing TiO248% of SiO2The recovery rate of titanium was 96% at 0.13%, and silicon and titanium were effectively separated. The obtained perovskite can be used for preparing photoelectric materials, and the leaching solution can be crystallized to obtain water glass.
Example 3
The mass percentages of the components in the embodiment are as follows: fe2O350% of Al2O315% of Na2O is 1.5%, TiO212 percent of the total weight of the alloy, and the balance of SiO2Red mud containing impurities such as CaO and the like is used as a raw material. Is ground by a high-pressure rollerRolling twice under the pressure of 250kN/m, uniformly mixing the red mud subjected to roller grinding with limestone and bentonite, pelletizing on a disc machine to obtain green pellets with the diameter of 8-16 mm and the moisture of 17%, and drying the green pellets at the temperature of 200-300 ℃ for more than 30min to obtain dry pellets;
preheating the dry balls on a chain grate machine for consolidation, preheating at 1250 deg.C for 10min to obtain balls with compression strength of 605 (N.n)-1) The pellets are preheated.
Adding the preheated pellets and the reducing coal into a reducing furnace according to the coal mine mass ratio of 1.5:1, and reducing for 90min at 1300 ℃ to obtain pre-reduced pellets with the metallization rate of 88%;
adding a reducing agent coke according to 10 percent of the mass of the pre-reduced pellets, uniformly mixing the reducing agent and the pre-reduced pellets, putting the mixture into an electric furnace for smelting, smelting for 40min at 1500 ℃, obtaining pig iron with the grade of 94 percent, the iron recovery rate of 95 percent and Al in the molten slag2O3Content of 46% Na2O content of 5.5%, TiO2And SiO2The content is 17 percent and 15 percent, iron basically enters pig iron in the melting process, and aluminum, silicon and titanium enter a slag phase to realize iron-aluminum separation;
and (4) melting slag modification. Considering that part of alumina and silica in the molten slag form aluminosilicate, it is converted into sodium aluminate and calcium silicate by adjusting the alkali ratio and calcium ratio. Adding sodium carbonate and limestone according to the alkali ratio of 1 and the calcium ratio of 2, uniformly mixing the three, oxidizing and roasting at 900 ℃ for 60min, and taking out and naturally cooling; crushing the obtained modified slag until the granularity is less than 1mm, and then, performing fine grinding in a ball mill, wherein the concentration of ore pulp is controlled to be 50%, and the particles with the granularity of less than 0.074mm account for 85%; leaching ball-milling residue for 120min at 80 deg.C under conditions of liquid-solid ratio of 10:1(mL: g) and sodium hydroxide concentration of 160g/L to obtain leaching residue containing Al2O3Content of 15% TiO2Content of 22% SiO2The content is 20%, the leaching rate of aluminum is 82%, the leaching rate of titanium and silicon is about 5%, the separation of aluminum, titanium and silicon is realized, and the obtained sodium aluminate solution is used for producing alumina;
for the alkaline leaching residue after the aluminum extraction,leaching for 30min under the conditions that the liquid-solid ratio is 10:1(mL: g), the concentration of hydrochloric acid is 10.9g/L and the leaching temperature is 30 ℃ to further remove aluminum and silicon, wherein the obtained leaching residue is perovskite concentrate containing TiO244% of SiO2The recovery rate of titanium was 94% at 0.13%, and silicon and titanium were effectively separated. The obtained perovskite can be used for preparing photoelectric materials, and the leaching solution can be crystallized to obtain water glass.
Comparative example 1
The mass percentages of the components in the embodiment are as follows: fe2O350% of Al2O315% of Na2O is 1.5%, TiO212 percent of the total weight of the alloy, and the balance of SiO2CaO, and the like; the red mud is used as raw material. Rolling once under 292kN/m pressure by a high-pressure roller mill, uniformly mixing the red mud after the roller mill with limestone and bentonite, pelletizing on a disc machine to obtain green pellets with the diameter of 8-16 mm and the water content of 16%, and drying the green pellets at the temperature of 200-300 ℃ for more than 30min to obtain dry pellets;
preheating the dry balls on a chain grate machine for consolidation, preheating at 950 deg.C for 15min to obtain balls with compressive strength of 500 (N.pieces)-1) The pellets are preheated.
Adding the preheated pellets and the reducing coal into a reducing furnace according to the coal mine mass ratio of 2:1, and reducing for 90min at 1050 ℃ to obtain pre-reduced pellets with the metallization rate of 80%;
adding a reducing agent coke according to 10 percent of the mass of the pre-reduced pellets, uniformly mixing the reducing agent and the pre-reduced pellets, putting the mixture into an electric furnace for smelting, smelting for 30min at the temperature of 1550 ℃, obtaining the pig iron with the grade of 95 percent, the iron recovery rate of 94 percent and Al in the molten slag2O3The content of Na is 48%2O content of 5.5%, TiO2And SiO2The content is 15 percent and 12 percent, iron basically enters pig iron in the melting process, and aluminum, silicon and titanium enter a slag phase to realize iron-aluminum separation;
the molten slag is directly leached without modification. Crushing the obtained melt separation slag until the granularity is less than 1mm, then, performing fine grinding in a ball mill, controlling the concentration of ore pulp to be 50%, and grinding particles with the granularity of less than 0.074mm90 percent of the total weight; leaching ball-milling residue for 120min at 95 deg.C under the conditions of liquid-solid ratio of 10:1(mL: g) and sodium hydroxide concentration of 80g/L to obtain leaching residue containing Al2O338% of TiO219% of SiO2The content was 15%, the leaching rate of aluminum was 50%, and the leaching rates of titanium and silicon were about 5%, and separation of aluminum, titanium and silicon was not achieved.
Leaching the alkaline leaching residue after aluminum extraction for 30min under the conditions that the liquid-solid ratio is 10:1(mL: g), the hydrochloric acid concentration is 87.6g/L and the leaching temperature is 30 ℃, wherein the obtained leaching residue is titanium dioxide and aluminosilicate and contains TiO226% of SiO210.45% of Al2O3At 15%, the recovery rate of titanium was 75%, and separation of silicon, aluminum and titanium was not efficiently achieved.
Comparative example 2
The mass percentages of the components in the embodiment are as follows: fe2O350% of Al2O315% of Na2O is 1.5%, TiO212 percent of the total weight of the alloy, and the balance of SiO2CaO, and the like; the red mud is used as raw material. Rolling once under 292kN/m pressure by a high-pressure roller mill, uniformly mixing the red mud after the roller mill with limestone and bentonite, pelletizing on a disc machine to obtain green pellets with the diameter of 8-16 mm and the water content of 16%, and drying the green pellets at the temperature of 200-300 ℃ for more than 30min to obtain dry pellets;
preheating the dry balls on a chain grate machine for consolidation, preheating at 950 deg.C for 15min to obtain balls with compressive strength of 500 (N.pieces)-1) The pellets are preheated.
Adding the preheated pellets and the reducing coal into a reducing furnace according to the coal mine mass ratio of 2:1, and reducing for 90min at 1050 ℃ to obtain pre-reduced pellets with the metallization rate of 80%;
adding a reducing agent coke according to 10 percent of the mass of the pre-reduced pellets, uniformly mixing the reducing agent and the pre-reduced pellets, putting the mixture into an electric furnace for smelting, smelting for 30min at the temperature of 1550 ℃, obtaining the pig iron with the grade of 95 percent, the iron recovery rate of 94 percent and Al in the molten slag2O3The content of Na is 48%2O content of 5.5％，TiO2And SiO2The content is 15 percent and 12 percent, iron basically enters pig iron in the melting process, and aluminum, silicon and titanium enter a slag phase to realize iron-aluminum separation;
the molten slag is directly leached out through modification by a single additive. Considering that a part of alumina and silica in the molten slag form aluminosilicate, it is converted into sodium aluminate and cristobalite by adjusting the alkali ratio. Adding sodium carbonate according to the alkali ratio of 1, uniformly mixing the two, carrying out oxidizing roasting at 1300 ℃ for 15min, taking out the mixture, naturally cooling the mixture, crushing the obtained modified slag until the granularity is less than 1mm, and then carrying out fine grinding in a ball mill, wherein the concentration of the ore pulp is controlled to be 50%, and the particles with the ore grinding granularity of less than 0.074mm account for 90%; leaching ball-milling residue for 120min at 95 deg.C under the conditions of liquid-solid ratio of 10:1(mL: g) and sodium hydroxide concentration of 80g/L to obtain leaching residue containing Al2O3TiO content of 25%2Content of 24% SiO2The content was 20%, the leaching rate of aluminum was 65% and the leaching rates of titanium and silicon were about 7%, and separation of aluminum, titanium and silicon was not achieved.
Leaching the alkaline leaching residue after aluminum extraction for 30min under the conditions that the liquid-solid ratio is 10:1(mL: g), the hydrochloric acid concentration is 87.6g/L and the leaching temperature is 30 ℃, wherein the obtained leaching residue is titanium dioxide and aluminosilicate and contains TiO232% of SiO216.5% of Al2O3At 10%, the recovery rate of titanium was 80%, and separation of silicon, aluminum and titanium was not effectively achieved.
Comparative example 3
The mass percentages of the components in the embodiment are as follows: fe2O350% of Al2O315% of Na2O is 1.5%, TiO212 percent of the total weight of the alloy, and the balance of SiO2CaO, and the like; the red mud is used as raw material. Rolling once under 292kN/m pressure by a high-pressure roller mill, uniformly mixing the red mud after the roller mill with limestone and bentonite, pelletizing on a disc machine to obtain green pellets with the diameter of 8-16 mm and the water content of 16%, and drying the green pellets at the temperature of 200-300 ℃ for more than 30min to obtain dry pellets;
the dry balls are preheated and solidified on a chain grate machine at a preheating temperaturePreheating at 950 deg.C for 15min to obtain compressive strength of 500(N pieces)-1) The pellets are preheated.
Adding the preheated pellets and the reducing coal into a reducing furnace according to the coal mine mass ratio of 2:1, and reducing for 90min at 1050 ℃ to obtain pre-reduced pellets with the metallization rate of 80%;
adding a reducing agent coke according to 10 percent of the mass of the pre-reduced pellets, uniformly mixing the reducing agent and the pre-reduced pellets, putting the mixture into an electric furnace for smelting, smelting for 30min at the temperature of 1550 ℃, obtaining the pig iron with the grade of 95 percent, the iron recovery rate of 94 percent and Al in the molten slag2O3The content of Na is 48%2O content of 5.5%, TiO2And SiO2The content is 15 percent and 12 percent, iron basically enters pig iron in the melting process, and aluminum, silicon and titanium enter a slag phase to realize iron-aluminum separation;
the molten slag is directly leached out through modification by a single additive. Considering that part of alumina and silica in the molten slag form aluminosilicate, it is converted into alumina and calcium silicate by adjusting the calcium ratio. Adding limestone according to the calcium ratio of 2, uniformly mixing the limestone and the limestone, oxidizing and roasting the mixture at 1300 ℃ for 15min, taking out the mixture, naturally cooling the mixture, crushing the obtained modified slag until the granularity is less than 1mm, and then performing fine grinding in a ball mill, wherein the concentration of the ore pulp is controlled to be 50%, and the granularity of the ground particles is less than 0.074mm and accounts for 90%; leaching ball-milling residue for 120min at 95 deg.C under the conditions of liquid-solid ratio of 10:1(mL: g) and sodium hydroxide concentration of 80g/L to obtain leaching residue containing Al2O3Content of 30% TiO2Content of 22% SiO2The content was 24%, the leaching rate of aluminum was 55%, and the leaching rates of titanium and silicon were about 8%, and separation of aluminum, titanium and silicon was not achieved.
Leaching the alkaline leaching residue after aluminum extraction for 30min under the conditions that the liquid-solid ratio is 10:1(mL: g), the hydrochloric acid concentration is 87.6g/L and the leaching temperature is 30 ℃, wherein the obtained leaching residue is titanium dioxide and calcium aluminosilicate and contains TiO230% of SiO215% of Al2O3At 12%, the recovery rate of titanium was 70%, and separation of silicon, aluminum and titanium was not efficiently achieved.
While the foregoing is directed to the preferred embodiment of the present invention, it will be understood by those skilled in the art that various changes and modifications may be made without departing from the spirit and scope of the invention as defined in the appended claims.

Claims (10)

1. A method for comprehensively utilizing red mud without tailings is characterized by comprising the following steps:
s1, pelletizing:
adding an additive and a binder into the pretreated red mud, mixing, pelletizing and drying to obtain red mud dry balls;
s2, preheating and roasting:
preheating and consolidating the dry pellets obtained in the step S1 to obtain preheated pellets;
s3, pre-reduction treatment:
adding reducing coal into the preheated pellets obtained in the S2 for pre-reduction treatment to obtain pre-reduced pellets; wherein the pre-reduction temperature is 800-1300 ℃, and the pre-reduction time is 15-180 min;
s4, melt separation treatment:
adding a reducing agent into the pre-reduced pellets obtained in the step S3, mixing, and performing melt separation treatment to obtain pig iron and melt separation slag; wherein the melt separation temperature is 1450-1600 ℃, and the melt separation time is 20-40 min;
s5, modification treatment:
mixing the molten slag obtained in the step S4 with sodium carbonate and limestone, and then carrying out modification treatment to obtain modified slag; wherein the modification temperature is 900-1300 ℃, and the modification time is 15-60 min;
s6, ball milling and alkaline leaching:
carrying out ball milling and alkaline leaching on the modified slag obtained in the step S5 to obtain alkaline leaching slag and aluminate solution;
s7, acid leaching:
and (4) carrying out acid leaching on the alkaline leaching residue obtained in the step (S6) to obtain acid leaching solution and perovskite concentrate.
2. The method for comprehensively utilizing red mud without tailings according to claim 1, wherein in the step S1, the pretreatment is that the red mud is mixed with water and then subjected to high-pressure roll milling; wherein the mass of the water accounts for 6.5-7.5% of the mass of the red mud; the pressure of the high-pressure roller mill is 250-292 kN/m.
3. The method for comprehensively utilizing red mud without tailings as claimed in claim 1, wherein in the step S1, the additive is one of limestone, quicklime and slaked lime, and the binder is bentonite.
4. The method for comprehensively utilizing red mud without tailings according to claim 1, wherein the drying temperature in S1 is 200-300 ℃ and the drying time is more than 30 min.
5. The method for comprehensively utilizing red mud without tailings as claimed in claim 1, wherein the preheating roasting temperature is 950-1200 ℃, the time is 5-30 min, and the roasting wind speed is 2.0-2.6 m/s.
6. The method for comprehensively utilizing red mud without tailings as claimed in claim 1, wherein the mass ratio of the reduced coal to the preheated pellets in the pre-reduction treatment is 0.5-2: 1.
7. The method for comprehensively utilizing red mud without tailings as claimed in claim 1, wherein the mass ratio of the reducing agent to the pre-reduced pellets in the melting treatment is 0.05-0.2: 1, and the reducing agent is a carbon-containing fuel.
8. The method for comprehensively utilizing red mud without tailings as claimed in claim 1, wherein in the step S6, the ball milling is specifically to configure the modified slag into ore pulp with a concentration of 30-80%, and then grind the obtained ore pulp into ball milling slag, wherein particles with a particle size of less than 0.074mm account for 80-95%.
9. The method for comprehensively utilizing red mud without tailings as claimed in claim 1, wherein in the step S6, the alkaline leaching is specifically to add the ball-milled residue into an alkaline solution according to a liquid-solid mass ratio of 1-10: 1, wherein the concentration of the alkaline solution is 20-160 g/L; wherein the alkaline leaching temperature is 60-95 ℃, and the alkaline leaching time is 20-120 min.
10. The method for comprehensively utilizing red mud without tailings as claimed in claim 1, wherein in the step S7, the acid leaching is specifically to add the alkaline leaching residue into an acid solution according to a liquid-solid mass ratio of 1-10: 1, wherein the concentration of the acid solution is 10-392 g/L; wherein the acid leaching temperature is 25-95 ℃, and the acid leaching time is 20-120 min.
CN202010904906.1A 2020-09-01 2020-09-01 Method for comprehensively utilizing red mud without tailings Pending CN111893308A (en)

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Cited By (3)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
CN112591831A (en) * 2020-12-04 2021-04-02 兰州有色冶金设计研究院有限公司 Harmless treatment method for strongly acidic mine tailings
CN114988443A (en) * 2022-06-02 2022-09-02 中南大学 Method for recovering aluminum oxide from aluminum-rich slag
CN115261540A (en) * 2022-05-27 2022-11-01 中国恩菲工程技术有限公司 Method for recovering iron and tailings in red mud

Citations (6)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
CN103290207A (en) * 2013-06-14 2013-09-11 中南大学 Comprehensive recycling method of iron and aluminum from red mud
CN106367554A (en) * 2016-08-30 2017-02-01 宝钢发展有限公司 Method for extracting iron and valuable metal from secondary resources and producing slag wool
CN109439894A (en) * 2018-10-31 2019-03-08 安徽工业大学 A method of utilizing iron, aluminium in waste resource recycling red mud
EP1633679B1 (en) * 2003-06-16 2019-03-13 The University Of Leeds Extraction process for reactive metal oxides
CN109913641A (en) * 2019-03-18 2019-06-21 中南大学 A method of comprehensive utilization high alumina iron ore
CN110923442A (en) * 2019-12-17 2020-03-27 吕鲁平 Method for recovering titanium and iron from ilmenite

Patent Citations (6)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
EP1633679B1 (en) * 2003-06-16 2019-03-13 The University Of Leeds Extraction process for reactive metal oxides
CN103290207A (en) * 2013-06-14 2013-09-11 中南大学 Comprehensive recycling method of iron and aluminum from red mud
CN106367554A (en) * 2016-08-30 2017-02-01 宝钢发展有限公司 Method for extracting iron and valuable metal from secondary resources and producing slag wool
CN109439894A (en) * 2018-10-31 2019-03-08 安徽工业大学 A method of utilizing iron, aluminium in waste resource recycling red mud
CN109913641A (en) * 2019-03-18 2019-06-21 中南大学 A method of comprehensive utilization high alumina iron ore
CN110923442A (en) * 2019-12-17 2020-03-27 吕鲁平 Method for recovering titanium and iron from ilmenite

Cited By (3)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
CN112591831A (en) * 2020-12-04 2021-04-02 兰州有色冶金设计研究院有限公司 Harmless treatment method for strongly acidic mine tailings
CN115261540A (en) * 2022-05-27 2022-11-01 中国恩菲工程技术有限公司 Method for recovering iron and tailings in red mud
CN114988443A (en) * 2022-06-02 2022-09-02 中南大学 Method for recovering aluminum oxide from aluminum-rich slag

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Application publication date: 20201106