Rutile dissociation method based on coordination of acid leaching and alkali washing

Abstract

The invention belongs to the technical field of comprehensive utilization of mineral resources, and provides a rutile dissociation method based on acid leaching-alkali washing synergy, which comprises the following steps: (1) reacting rutile raw ore, acid liquor and a leaching agent to obtain acid leaching filter residues; and (2) carrying out reaction on the acid leaching filter residues and alkali liquor to obtain rutile concentrate, and completing dissociation. In order to solve the problem that the fine disseminated rutile ore cannot be effectively disseminated by conventional ore grinding, so that enrichment is difficult in a traditional physical ore dressing method, a process route of'moderate ore grinding and chemical deep dissociation 'is creatively adopted to replace a traditional mode depending on physical ore dressing. The moderate ore grinding avoids the risks of pulp viscosity increase and titanium loss caused by excessive ore grinding, and the chemical deep dissociation realizes the chemical selective separation of micro-fine rutile and gangue minerals, thereby effectively solving the problem that the micro-fine disseminated rutile is difficult to enrich through a conventional method.

Description

Technical Field

[0001] This invention relates to the field of comprehensive utilization technology of mineral resources, and in particular to a rutile dissociation method based on acid leaching-alkali washing synergy. Background Technology

[0002] Rutile (TiO2) is the main mineral in which titanium is found in nature and a key raw material for extracting titanium and titanium dioxide. With the increasing depletion of high-grade rutile resources, the development and utilization of low-grade and difficult-to-process rutile ore (titanium grade 3-6%) has become a key research focus in the industry. These ores typically have the following characteristics: ① Low titanium grade (TiO2 content is only 3%-6%), closely associated with gangue components such as iron, silicon, and aluminum; ② Fine mineral embedding (rutile particles are as fine as 1-10μm, or even nanoscale, and exist in a "jagged" or inclusion form with gangue minerals such as sericite, chlorite, and sphene); ③ Limitations of conventional grinding (even with ultrafine grinding to -325 mesh or even -500 mesh, the degree of liberation between rutile and gangue minerals is still less than 60%, and the proportion of liberated rutile particles is low); ④ Ineffectiveness of traditional beneficiation methods (physical methods such as gravity separation, magnetic separation, and flotation rely on differences in mineral density / magnetism / surface properties, have poor selectivity for finely embedded rutile, and cannot achieve effective enrichment).

[0003] For such difficult-to-process rutile ores, existing chemical treatment processes mostly employ a single "acid leaching" or "alkali leaching" process. While acid leaching (such as hydrochloric acid or sulfuric acid) can dissolve reactive impurities like iron and aluminum, it is almost ineffective against silicate inclusions (such as sphene and mica) and ilmenite intergrowths embedded in the fine rutile, and incomplete acid leaching leads to severe impurity residue. Alkali leaching primarily targets silicate decomposition, but the high specific surface area of ​​the fine particles exacerbates titanium dissolution and loss (titanium loss rate >5%), and it cannot selectively remove impurities such as iron and aluminum. Furthermore, the fine embedding characteristics make slurry filtration and washing difficult, and the separation efficiency of acid and alkali leachates from solid particles is low, further limiting the application effectiveness of chemical processes. Summary of the Invention

[0004] The purpose of this invention is to overcome the technical bottlenecks of conventional grinding and beneficiation, and to provide a rutile liberation method based on acid leaching-alkali washing synergy.

[0005] To achieve the above-mentioned objectives, the present invention provides the following technical solution: This invention provides a rutile dissociation method based on acid leaching-alkali washing synergy, comprising the following steps: (1) The raw rutile ore, acid solution and leaching agent are reacted to obtain acid leaching filter residue; (2) The acid leaching residue and alkaline solution are reacted to obtain rutile concentrate, thus completing the dissociation.

[0006] Preferably, the titanium grade of the rutile ore in step (1) is 3-6%; The mass percentage of components with a particle size ≤200 mesh in raw rutile ore is 60-80%.

[0007] Preferably, the acid solution in step (1) is a hydrochloric acid or sulfuric acid solution; The hydrochloric acid has a mass fraction of 15-30%; the sulfuric acid solution has a mass fraction of 40-60%.

[0008] Preferably, the mass-to-volume ratio of the rutile ore and acid solution in step (1) is 1g:3~6mL.

[0009] Preferably, the leaching agent in step (1) is sodium fluoride and / or fluorosilicic acid.

[0010] Preferably, when the leaching agent is sodium fluoride, the mass of sodium fluoride is 0.5-6% of the mass of the rutile ore; When the leaching agent is fluorosilicic acid, the mass of fluorosilicic acid is 0.2-4% of the mass of the raw rutile ore.

[0011] Preferably, the reaction temperature in step (1) is 60~120℃, the rotation speed is 300~600rpm, and the time is 3~6h.

[0012] Preferably, the mass fraction of the alkaline solution in step (2) is 10-20%.

[0013] Preferably, the mass-to-volume ratio of the acid leaching residue and the alkaline solution in step (2) is 1g:3~6mL.

[0014] Preferably, the reaction temperature in step (2) is 60~100℃, the rotation speed is 300~600rpm, and the time is 3~6h.

[0015] This invention provides a rutile dissociation method based on acid leaching-alkali washing synergy, comprising the following steps: (1) reacting rutile ore, acid solution, and leaching agent to obtain acid leaching residue; (2) reacting the acid leaching residue with alkali solution to obtain rutile concentrate, thus completing the dissociation. This invention achieves selective dissociation and precise separation of the major titanium component and associated impurities in rutile by stepwise control of the acid / alkali reaction conditions. Its innovative points are summarized as follows: 1. Targeted solutions to the problem of micro-embedding Innovative Process Route: Addressing the problem that conventional grinding cannot effectively liberate finely disseminated rutile ore, making it difficult to enrich rutile ore using traditional physical beneficiation methods, a creative "moderate grinding + deep chemical liberation" process route is adopted to replace traditional methods relying on physical beneficiation. Moderate grinding avoids the risks of increased slurry viscosity and titanium loss caused by over-grinding, while deep chemical liberation achieves the chemically selective separation of finely disseminated rutile from gangue minerals, effectively solving the problem of enriching finely disseminated rutile ore using conventional methods.

[0016] Adapting to the characteristics of fine particles: By controlling the grinding fineness and liquid-solid ratio, the characteristics of fine particles are adapted, reducing the viscosity of the slurry and the difficulty of filtration. In contrast to the traditional process of pursuing high fineness grinding but with poor results, this is more in line with the actual situation of such difficult-to-process rutile ore.

[0017] 2. Deep impurity removal and efficient titanium retention

[0018] First, a strong acid (such as hydrochloric acid / sulfuric acid) of specific concentration and temperature is used to directionally dissolve active metal impurities such as iron and aluminum (forming soluble chlorides / sulfates), while simultaneously destroying the crystal structure of impurities such as silicates and carbonates. The subsequent alkaline washing step selectively decomposes residual silicates (generating soluble sodium silicate) and incompletely reacted calcium and magnesium compounds (forming hydroxide precipitates or soluble salts) through an alkaline solution (such as sodium hydroxide), thereby achieving the complete removal of encapsulated impurities and isomorphic impurities.

[0019] Innovative application of auxiliary leaching agents: Sodium fluoride or fluorosilicic acid is introduced as an auxiliary leaching agent during the acid leaching process, utilizing fluoride ions (F... - It specifically binds to the Si-O bonds in the silicate lattice, preferentially disrupting the structure of gangue minerals such as sericite, chlorite, and sphene, releasing encapsulated rutile particles and dissolving embedded iron and aluminum impurities. This targeted auxiliary leaching method significantly improves the removal of reactive impurities such as iron and aluminum, solving the problem of insufficient dissolution of encapsulated impurities in finely embedded rutile by traditional acid leaching.

[0020] Excellent titanium retention: Acid leaching pretreatment reduces impurity content and inclusion structure stability, providing a better reaction contact interface for subsequent alkaline washing; alkaline washing specifically removes insoluble impurities remaining after acid leaching. The two processes form a closed-loop process of "dissolving active impurities first, then breaking stable inclusions," improving dissociation efficiency by more than 30% compared to single processes. By precisely controlling the process conditions of acid leaching and alkaline washing, the endpoint pH of acid leaching is controlled at 1.5~3.5 to avoid titanium lattice corrosion, and alkaline washing precisely removes active silicon and enriches high-valence elements. At the same time, the titanium loss rate is ≤1%, which is a significant improvement over traditional processes (titanium loss rate ≥5%), ensuring efficient retention of titanium components.

[0021] 3. Enrichment and recycling of valuable elements

[0022] Achieving efficient enrichment: The process achieves efficient separation and enrichment of elements such as Ti, Nb, Sc, and rare earth elements from impurities, with enrichment factors reaching 5-10 times. For associated valuable elements in low-grade rutile ore, a deep chemical dissociation process is used to enrich them in the slag phase, providing a solid foundation for subsequent separation and recovery of valuable elements. This enables comprehensive multi-element recovery from low-grade rutile ore, enhancing the comprehensive utilization value of the resource.

[0023] The comprehensive recycling approach not only focuses on the purification of titanium, the main component of rutile, but also emphasizes the recycling and utilization of associated valuable elements. This approach changes the traditional process's neglect of valuable elements and embodies a more comprehensive resource utilization concept.

[0024] 4. Process adaptability and green environmental protection

[0025] Process flexibility: By controlling the grinding fineness and liquid-solid ratio, the process adapts to the characteristics of fine particles, demonstrating its adaptability and flexibility to raw materials with different properties. It can adjust process parameters according to actual conditions to ensure processing results.

[0026] Environmentally friendly: Acid and alkali waste liquids are recycled, and fluorides in the waste liquids can be recovered, reducing waste emissions and environmental pollution, which meets the requirements of green and environmentally friendly development. It also reduces the consumption of reagents and the generation of waste, and the overall treatment cost is 15% to 20% lower than that of traditional chemical purification processes.

[0027] 5. Significant economic value

[0028] This process transforms low-grade, difficult-to-process rutile ore with a titanium content of 3-6% into titanium concentrate with a TiO2 content of ≥16% and rich in valuable elements such as Nb, Sc, and rare earth elements. Simultaneously, the liberation degree of rutile is increased to over 75%, providing a foundation for further improving the titanium grade and extracting and comprehensively utilizing associated valuable elements. Resource utilization rate is increased by over 20%, significantly enhancing economic benefits and providing an economically feasible technical solution for the development and utilization of low-grade rutile ore. Attached Figure Description

[0029] Figure 1 This is a schematic flowchart of the rutile dissociation method based on acid leaching-alkali washing synergy of the present invention. Detailed Implementation

[0030] This invention provides a rutile dissociation method based on acid leaching-alkali washing synergy, comprising the following steps: (1) The raw rutile ore, acid solution and leaching agent are reacted to obtain acid leaching filter residue; (2) The acid leaching residue and alkaline solution are reacted to obtain rutile concentrate, thus completing the dissociation.

[0031] In this invention, the titanium grade of the rutile ore in step (1) is preferably 3-6%, more preferably 4-5%, and even more preferably 4.5%.

[0032] In this invention, the mass percentage of the component with a particle size ≤200 mesh in the rutile ore is preferably 60-80%, more preferably 65-75%, and even more preferably 68-72%.

[0033] In this invention, a specific mesh size is selected to avoid excessive grinding that could increase the viscosity of the slurry and increase the risk of titanium loss. The traditional process of -325 mesh or higher fineness is not pursued in order to balance the degree of liberation and the efficiency of slurry processing.

[0034] In this invention, the acid solution in step (1) is a hydrochloric acid or sulfuric acid solution.

[0035] In this invention, the mass fraction of the hydrochloric acid is preferably 15-30%, more preferably 20-25%, and even more preferably 22-23%; the mass fraction of the sulfuric acid solution is preferably 40-60%, more preferably 45-55%, and even more preferably 46-54%.

[0036] In this invention, the mass-to-volume ratio of rutile ore and acid in step (1) is preferably 1g:3~6mL, more preferably 1g:3.5~5.5mL, and even more preferably 1g:4~5mL.

[0037] In this invention, the leaching agent in step (1) is sodium fluoride and / or fluorosilicic acid.

[0038] In this invention, when the leaching agent is sodium fluoride, the mass of sodium fluoride is preferably 0.5-6% of the mass of the raw rutile ore, more preferably 1-5%, and even more preferably 2-4%.

[0039] In this invention, when the leaching agent is fluorosilicic acid, the mass of fluorosilicic acid is preferably 0.2-4% of the mass of the raw rutile ore, more preferably 1-3%, and even more preferably 1.5-2.5%.

[0040] In this invention, the temperature of the reaction in step (1) is preferably 60~120℃, more preferably 70~110℃, and even more preferably 100~105℃; the rotation speed is preferably 300~600rpm, more preferably 350~550rpm, and even more preferably 400~500rpm; and the time is preferably 3~6h, more preferably 4~5h, and even more preferably 4.4~4.6h.

[0041] In this invention, the mechanism of acid leaching in step (1) is as follows: fluoride ions (F - It specifically binds to the Si-O bonds in the silicate lattice (forming [SiF6)). 3-Complexes or disruption of Si-O-Al bridging bonds) preferentially disrupt the structure of gangue minerals such as sericite, chlorite, and sphene, releasing encapsulated rutile particles and dissolving embedded iron and aluminum impurities (forming FeCl3 / Fe2(SO4)3 and AlCl3 / Al2(SO4)3); simultaneously, F - It forms soluble fluoride-iron complexes with iron (such as [FeF6)). 3- The acid solution synergistically promotes the deep dissolution of iron and aluminum impurities. The acid solution dissolves the generated active metal ions and assists in the decomposition of other easily soluble impurities. By controlling the final pH to 1.5–3.5 (to avoid titanium lattice corrosion), selective removal of impurities embedded in the fine rutile is achieved.

[0042] In this invention, after the reaction in step (1) is completed, solid-liquid separation is performed to obtain acid leaching residue (mainly composed of fine-grained rutile, active silicon, and enriched high-valence elements such as Ti, Nb, Sc, and rare earth elements) and acid leaching filtrate (containing metal salts such as Fe and Al and a small amount of fluoride). The obtained acid leaching residue is washed to remove residual acid and soluble metal salts from the surface until the pH of the washing liquid is ≥5, and then the reaction in step (2) is carried out. The washing water is returned to the acid leaching process for recycling.

[0043] In this invention, the mass fraction of the alkaline solution in step (2) is preferably 10-20%, more preferably 12-18%, and even more preferably 14-16%.

[0044] In this invention, the mass-to-volume ratio of the acid leaching residue and the alkaline solution in step (2) is preferably 1g:3~6mL, more preferably 1g:3.5~5.5mL, and even more preferably 1g:4~5mL; the alkali used in the alkaline solution is sodium hydroxide and / or potassium hydroxide.

[0045] In this invention, the reaction temperature in step (2) is preferably 60~100℃, more preferably 70~90℃, and even more preferably 80~85℃; the rotation speed is preferably 300~600rpm, more preferably 400~500rpm, and even more preferably 440~460rpm; and the time is preferably 3~6h, more preferably 3.5~5.5h, and even more preferably 4~5h.

[0046] In this invention, during the reaction process of step (2), the alkali solution selectively decomposes the residual active silicon (such as SiO2 → soluble sodium silicate), while promoting the formation of stable composite oxides or titanate phases of high-valence elements such as titanium, niobium, scandium, and rare earth elements, which are enriched in the solid slag. By precisely controlling the alkali concentration and reaction conditions, the dissolution and removal of active silicon in fine-grained rutile and the stable retention of the slag phase of high-valence elements are achieved, avoiding titanium dissolution loss (titanium loss rate ≤1%). The final pH is controlled at 10~13. After the reaction is completed, solid and liquid are separated to obtain alkali-washed filter residue (high-purity fine-grained rutile) and alkali-washed filtrate (containing sodium silicate and a small amount of impurities).

[0047] In this invention, after the reaction in step (2) is completed, the alkaline washing filter residue is repeatedly washed countercurrently with deionized water (until the washing liquid pH ≤ 8 and the conductivity ≤ 50 μS / cm) to remove residual alkaline solution and soluble salts on the surface; the washing water is returned to the alkaline washing process for recycling. After washing, it is dried to constant weight. The drying temperature is preferably 100~120℃, more preferably 105~115℃, and even more preferably 108~112℃; after drying, rutile concentrate is obtained (the titanium grade is increased to ≥ 16%, the monomer liberation degree of rutile is increased from the original 6% to more than 75%, and the associated Nb, Sc, rare earth elements are effectively enriched, such as Nb2O5 enriched by 10~15 times and Sc2O3 enriched by 5~10 times).

[0048] In this invention, the obtained rutile concentrate can be further processed to obtain high-grade rutile products through selective floc flotation or electrostatic separation, or to selectively dissolve high-valence elements such as Ti, Nb, Sc, and rare earth elements through acid leaching or roasting-acid leaching processes to achieve the separation and recovery of valuable components; or it can be directly used as a titanium-rich, niobium-, scandium-, and rare earth raw material for the preparation of high-end materials; the Fe and Al salts and fluorides in the acid leaching filtrate can be recovered as iron and aluminum compounds (such as fluoroaluminates) through crystallization, precipitation, and other methods; the sodium silicate in the alkaline washing filtrate can be recovered as silica through acid neutralization or used to prepare water glass, thereby achieving waste liquid recycling and resource maximization.

[0049] A schematic diagram of the rutile dissociation method based on acid leaching-alkali washing synergy of the present invention is shown below. Figure 1 As shown.

[0050] The technical solutions provided by the present invention will be described in detail below with reference to the embodiments, but they should not be construed as limiting the scope of protection of the present invention.

[0051] Example 1

[0052] Raw material: A low-grade rutile ore (titanium grade 4.26%, rutile particle size 0.5-8μm, with fine inclusions of sericite and chlorite, -200 mesh accounting for 55%, but the degree of liberation of individual particles is only 6%).

[0053] step: Grind the ore to a minimum of -200 mesh, making up 55% of the total (avoid over-grinding). Acid leaching: Mineral powder and 50% sulfuric acid solution were mixed at a ratio of 1g:3mL, and sodium fluoride (the mass of sodium fluoride was 6% of the raw rutile ore) was added. The mixture was reacted at 110℃ and 500rpm for 4h, with the pH controlled at 2.5. After the reaction, the leaching rate of Al2O3 was 92%, the leaching rate of Fe2O3 was 95%, and the release rate of encapsulated rutile was significantly improved. Washing: The acid leaching filter residue is washed countercurrently with deionized water until the pH of the washing solution is ≥5 to remove residual acid and soluble metal salts from the surface. The washing water is returned to the acid leaching process for recycling.

[0054] Alkali washing: Mix the acid leaching residue with 15% sodium hydroxide solution at a ratio of 1g:5mL, and react at 90℃ and 500rpm for 4h; The final titanium concentrate product has a titanium grade of 16.2%. The alkali-washed filter residue is enriched by 11.6 times for Nb2O5, 8.9 times for Sc2O3, and 10.8 times for rare earth elements. The degree of rutile liberation is 78%, and the titanium loss rate is 0.6%, which solves the problem of the inability to effectively enrich finely embedded rutile.

[0055] Example 2

[0056] Raw material: Rutile ore with a titanium grade of 3.96% (rutile and hematite are closely associated and embedded in gangues such as sericite, chlorite and quartz. After simple magnetic separation, the titanium grade of the ore is increased to 5.2%, with -325 mesh accounting for 95%, but the degree of liberation of the monomers is only 55%).

[0057] step: Grind appropriately to -200 mesh, accounting for 65%; Acid leaching: Mineral powder was mixed with 60% sulfuric acid solution at a ratio of 1g:5mL, and 2% fluorosilicic acid was added. The mixture was reacted at 100℃ and 400rpm for 6 hours. After the reaction, the leaching rate of Al2O3 was 93%, and the leaching rate of Fe2O3 was 96%. Subsequent alkaline washing and other steps were the same as in Example 1, and the final titanium concentrate product was obtained with a titanium grade of 18.6%. The Nb2O5 enrichment in the alkaline washing filter residue was 12.6 times, the Sc2O3 enrichment was 8.6 times, and the rare earth enrichment was 10.6 times. The degree of rutile dissociation was 75%, and the titanium loss rate was 0.8%, realizing the chemical dissociation and deep purification of finely embedded rutile.

[0058] The above description is only a preferred embodiment of the present invention. It should be noted that for those skilled in the art, several improvements and modifications can be made without departing from the principle of the present invention, and these improvements and modifications should also be considered within the scope of protection of the present invention.

Claims

1. A method for the dissociation of rutile based on the synergy of acid leaching-alkaline washing, characterized by, The method comprises the following steps: (1) reacting rutile raw ore, acid liquor and leaching agent to obtain acid leaching residue; (2) reacting the acid leaching residue and alkali liquor to obtain rutile concentrate, and completing dissociation.

2. The method for rutile dissociation based on the synergy of acid leaching and alkali washing according to claim 1, characterized in that, The titanium grade of the rutile raw ore in step (1) is 3-6%. The mass percentage of components with particle size of less than 200 μm in the rutile raw ore is 60-80%.

3. The method for rutile dissociation based on the synergy of acid leaching and alkali washing according to claim 2, characterized in that, The acid liquor in step (1) is hydrochloric acid or sulfuric acid solution. The mass fraction of the hydrochloric acid is 15-30%, and the mass fraction of the sulfuric acid solution is 40-60%.

4. The method for rutile dissociation based on the synergy of acid leaching and alkali washing according to claim 3, characterized in that, The mass-volume ratio of the rutile raw ore to the acid liquor in step (1) is 1 g:3-6 mL.

5. The method for rutile dissociation based on the synergy of acid leaching and alkali washing according to claim 4, characterized in that, The leaching agent in step (1) is sodium fluoride and / or fluorosilicic acid.

6. The method for rutile dissociation based on the synergy of acid leaching and alkali washing according to claim 5, characterized in that, When the leaching agent is sodium fluoride, the mass of sodium fluoride is 0.5-6% of the mass of the rutile raw ore. When the leaching agent is fluorosilicic acid, the mass of fluorosilicic acid is 0.2-4% of the mass of the rutile raw ore.

7. The method for the leaching-alkaline washing synergic based rutile dissociation according to claim 6, characterized in that, The temperature of the reaction in step (1) is 60-120°C, the rotation speed is 300-600 rpm, and the time is 3-6 h.

8. The method for rutile dissociation based on the synergy of acid leaching and alkali washing according to claim 7, characterized in that, The mass fraction of the alkali liquor in step (2) is 10-20%.

9. The method for rutile dissociation based on the synergy of acid leaching and alkali washing according to claim 8, characterized in that, The mass-volume ratio of the acid leaching residue to the alkali liquor in step (2) is 1 g:3-6 mL.

10. The method for the leaching-alkaline washing synergic rutile dissociation according to claim 9, characterized in that, The temperature of the reaction in step (2) is 60-100°C, the rotation speed is 300-600 rpm, and the time is 3-6 h.

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