Titanium metallurgical chlorination tailing multi-metal comprehensive utilization method

Abstract

The application discloses a kind of titanium metallurgy chlorination tailings multi-metal comprehensive utilization method, belong to solid waste resource utilization and hydrometallurgy technical field.The method with the chlorination tailings generated in the production process of titanium metallurgy as raw material, first, using dilute hydrochloric acid to the target elements such as scandium, vanadium, iron, manganese and so on are selectively leached;Liquor is sequentially through neutralization and iron is deposited, extraction separation scandium, extraction separation vanadium, neutralization and manganese is deposited and so on, realize the step recovery of iron, scandium, vanadium, manganese;Leaching residue is further treated by concentrated acid leaching, and the obtained leaching liquor is obtained by neutralization and precipitation niobium tantalum rough concentrate;The application utilizes the difference of each metal component in leaching activity, extraction affinity and hydrolysis precipitation behavior, constructs the process route of step leaching-step separation, can realize the efficient separation and enrichment of niobium, tantalum, scandium, vanadium, iron, manganese and other valuable metals;The comprehensive utilization degree of multiple valuable metals in tailings is improved, and the resource utilization of solid waste is realized.

Description

Technical Field

[0001] This invention relates to the fields of solid waste resource utilization and hydrometallurgical technology, specifically to a method for comprehensive utilization of multiple metals in titanium metallurgical chlorination tailings. Background Technology

[0002] Titanium dioxide is an important inorganic chemical raw material, widely used in coatings, plastics, papermaking, inks and other fields. The existing titanium dioxide production processes mainly include the sulfuric acid process and the chloride process. Among them, the chloride process has the characteristics of relatively short process flow, stable product quality and high degree of continuity. In recent years, its industrial application scale has continued to expand.

[0003] Furthermore, the chlorination process typically uses titanium-containing raw materials such as high-titanium slag or natural rutile as the base material, and prepares titanium dioxide through processes such as chlorination, refining and oxidation. During the chlorination and subsequent treatment processes, solid waste containing components such as iron, manganese, scandium, vanadium, niobium, tantalum, silicon, and titanium will be generated. Different production raw materials, chlorination conditions and pretreatment methods will result in certain differences in the chemical composition and phase composition of the obtained solid waste.

[0004] Currently, the chlorinated slag produced during the production of titanium dioxide using the chlorination process can be preliminarily treated through methods such as water leaching, acid leaching, solid-liquid separation, and mineral processing. Among these methods, some easily soluble metal components can be added to the leaching solution for further recovery, and the remaining solid materials can be separated by flotation to remove residual carbon or titanium-containing components. However, even after the above treatment, a certain amount of chlorinated tailings may still be generated, which still contain valuable metals such as iron, manganese, scandium, vanadium, niobium, and tantalum.

[0005] It should be noted that this type of tailings is characterized by the coexistence of multiple metals and significant differences in the content of different metals; among them, the content of components such as iron and manganese is relatively high, while the content of valuable metals such as scandium, vanadium, niobium and tantalum is relatively low; if only one of the metals is recovered separately, the remaining valuable components are easily lost with the intermediate materials or final residue, thus making it difficult to fully realize the resource value of the tailings.

[0006] Furthermore, existing hydrometallurgical processes typically employ acid leaching to introduce the target metal into the liquid phase, followed by separation based on the extraction, precipitation, or ion exchange properties of different metals. For titanium metallurgical solid waste containing multiple metals, increasing the acid concentration to ensure sufficient leaching of each metal can lead to the simultaneous entry of different metals into the leachate, complicating the subsequent separation system and increasing the amount of extractants, neutralizers, and other reagents required. Conversely, if the acid concentration is too low, the leaching of target components such as iron, manganese, scandium, and vanadium will be insufficient, which is also detrimental to subsequent comprehensive recovery.

[0007] Furthermore, the effective separation of low-content metals such as scandium and vanadium in the liquid phase usually requires the use of different types of extraction systems. Repeatedly adjusting the solution pH with acid or alkali between extraction steps not only increases reagent consumption but also further complicates the process flow and operational control. Therefore, how to rationally arrange the separation sequence of each metal based on the differences in leaching under specific acidity conditions and the subsequent extraction and precipitation behaviors is a problem that needs to be solved to achieve the comprehensive utilization of this type of chlorination tailings. Based on this, it is necessary to provide a multi-metal comprehensive utilization method suitable for titanium metallurgical chlorination tailings. Summary of the Invention

[0008] Based on the above analysis, the purpose of this invention is to provide a method for the comprehensive utilization of multiple metals in titanium metallurgical chlorination tailings. This method involves pre-separating different metal components in the tailings through a single low-acidity leaching process, followed by staged separation based on the differences in neutralization precipitation, solvent extraction, and concentrated acid leaching conditions. This achieves the comprehensive recovery of valuable metals such as iron, scandium, vanadium, manganese, niobium, and tantalum. Furthermore, another technical problem addressed by this invention is reducing the subsequent separation load caused by all metals entering the same leaching solution. By using the primary leaching solution mainly for the sequential separation of iron, scandium, vanadium, and manganese, and using the primary leaching residue for further recovery of niobium and tantalum, two material routes with different treatment objectives can be formed in the initial leaching stage.

[0009] To solve the above-mentioned technical problems, the present invention provides the following technical solution:

[0010] A method for comprehensive utilization of multiple metals in titanium metallurgical chlorination tailings includes the following steps:

[0011] S1. Add a 2-10 wt.% hydrochloric acid solution to the chlorinated tailings produced during the chloride process of titanium dioxide production, and carry out a first leaching under heating and stirring conditions; after the first leaching is completed, filter to obtain the first leaching solution and the first leaching residue.

[0012] Furthermore, in step S1, by controlling the concentration of hydrochloric acid, iron, manganese, scandium and vanadium in the tailings are preferentially introduced into the primary leaching solution, while niobium and tantalum are mainly retained in the primary leaching residue. This arrangement can form a primary leaching solution treatment route and a primary leaching residue treatment route in the primary leaching stage, providing a material basis for the subsequent separate recovery of different metal components.

[0013] Further, the concentration of hydrochloric acid in step S1 is preferably 6-9 wt.%; the liquid-to-solid ratio of the single leaching is preferably 2.5-4.0 L / kg; the leaching temperature is preferably 70-90℃; and the leaching time is preferably 60-90 min.

[0014] S2. Add neutralizing agent A to the primary leachate obtained in step S1 and adjust the pH to 3.0-4.0; after filtration, scandium vanadium solution and crude iron hydroxide precipitate are obtained.

[0015] Furthermore, in step S2, the pH is preferably adjusted to 3.4 to 3.7; within this pH range, the iron in the primary leaching solution is precipitated and separated from the liquid phase by neutralization treatment, while scandium and vanadium continue to be retained in the resulting scandium-vanadium solution for subsequent extraction and separation.

[0016] Further, the neutralizing agent A is selected from at least one of Ca(OH)2, CaCO3, MgO and NaOH; preferably, step S2 uses carbide slag with Ca(OH)2 as the main component as neutralizing agent A to achieve neutralization of the primary leachate and precipitation of iron components.

[0017] S3. The scandium-vanadium solution obtained in step S2 is extracted using organic phase B; after extraction, phase separation is performed to obtain scandium-containing organic phase and scandium leaching residue.

[0018] Further, the organic phase B consists of 5-15 vol.% di(2-ethylhexyl) phosphate, 2-5 vol.% tributyl phosphate, and the balance sulfonated kerosene; preferably, the organic phase B consists of 8-12 vol.% di(2-ethylhexyl) phosphate, 3-4 vol.% tributyl phosphate, and the balance sulfonated kerosene.

[0019] Furthermore, in step S3, the volume ratio A / O of the scandium-vanadium solution to the organic phase B is preferably 6:1 to 8:1; through the above organic phase composition and ratio conditions, scandium preferentially enters the organic phase, and vanadium is mainly retained in the scandium extraction residue.

[0020] S4. Extract the scandium raffinate obtained in step S3 using organic phase C; after extraction, separate the phases to obtain a vanadium-containing organic phase and a vanadium raffinate.

[0021] Further, the organic phase C consists of 5-15 vol.% secondary carbon amine extractant, 2-5 vol.% secondary octanol, and the balance sulfonated kerosene; preferably, the secondary carbon amine extractant is secondary carbon amine N1923, and the organic phase C consists of 8-12 vol.% secondary carbon amine N1923, 3-4 vol.% secondary octanol, and the balance sulfonated kerosene.

[0022] Further, in step S4, the volume ratio A / O of the scandium leaching residue to the organic phase C is preferably 4:1 to 6:1; under the above conditions, vanadium enters the vanadium-containing organic phase and a vanadium leaching residue is obtained for subsequent manganese precipitation treatment.

[0023] It should be noted that the scandium extract residue obtained from phase separation in step S3 is kept at the pH state after phase separation before entering step S4; in step S4, organic phase C is used to extract the scandium extract residue under this pH condition; this setting allows the scandium extraction step and the vanadium extraction step to be connected continuously, and reduces the acid-base adjustment operation between the two extraction steps.

[0024] S5. Add neutralizing agent A to the vanadium extraction residue obtained in step S4 and adjust the pH to 9.5-10.5; after filtration, crude manganese hydroxide precipitate is obtained.

[0025] Furthermore, in step S5, the pH is preferably adjusted to 9.8-10.5; the neutralizing agent A is preferably carbide slag with Ca(OH)2 as the main component, so that the manganese in the vanadium extraction residue forms a precipitate and is separated from the liquid phase.

[0026] With this setup, the leaching solution undergoes a series of processes: iron neutralization and precipitation, scandium extraction and separation, vanadium extraction and separation, and manganese neutralization and precipitation. This allows the different metal components in the leaching solution to be separated in the order of iron, scandium, vanadium, and manganese, avoiding the concentration of multiple target metals in the same final product.

[0027] S6. Add concentrated acid to the primary leaching residue obtained in step S1, and carry out secondary leaching under heating and stirring conditions; after the secondary leaching is completed, filter to obtain secondary leachate.

[0028] Furthermore, in step S6, the liquid-to-solid ratio of the secondary leaching is 1.2–3.0 L / kg, and the leaching time is 30–120 min;

[0029] Further, the concentrated acid is a hydrochloric acid solution or a sulfuric acid solution with a concentration >20 wt.%; preferably, the concentrated acid is sulfuric acid, and a secondary leaching is performed by first treating with concentrated sulfuric acid and then adding water for further leaching.

[0030] Furthermore, when using sulfuric acid for secondary leaching, first add a sulfuric acid solution with a concentration of 90-95 wt.% to the primary leaching residue at a liquid-to-solid ratio of 1.25-1.50 L / kg and react for 20-40 min; then add water to adjust the liquid-to-solid ratio to 2.0-3.0 L / kg, and then stir and leach at 60-90℃ for 20-40 min, and filter to obtain the secondary leachate.

[0031] This setup allows for the initial leaching residue to be treated under high acidity conditions, followed by the addition of water to adjust the material system and continued heating for leaching. The niobium and tantalum retained in the solid phase during the initial leaching stage then enter the secondary leaching solution, providing conditions for subsequent niobium and tantalum precipitation and enrichment.

[0032] S7. Add neutralizing agent A to the secondary leachate obtained in step S6 and adjust the pH to 1.5-3.0; after filtration, niobium-tantalum crude concentrate is obtained.

[0033] Furthermore, in step S7, the pH of the secondary leaching solution is preferably adjusted to 1.8–2.5; within this pH range, niobium and tantalum in the secondary leaching solution precipitate, thereby obtaining niobium-tantalum crude concentrate.

[0034] Furthermore, the chlorinated tailings of the present invention are tailings generated during the production of titanium dioxide using the chlorination process; preferably, the dry basis composition of the chlorinated tailings satisfies: Fe > 15 wt.%, Mn > 4 wt.%, 0.2 wt.% < (Nb + Ta) < 0.4 wt.%, 0.015 wt.% < Sc < 0.04 wt.%, 0.15 wt.% < V < 0.40 wt.%, SiO2 > 15 wt.%.

[0035] It should be noted that in this invention, wt.% represents the mass percentage of the corresponding component; vol.% represents the volume percentage of the corresponding component in the organic phase; the liquid-to-solid ratio represents the ratio of liquid volume to dry solid mass, in L / kg; and the A / O ratio represents the ratio of aqueous phase volume to organic phase volume.

[0036] Furthermore, in this invention, "primary leachate" and "primary leachate residue" refer to the liquid phase and solid phase obtained by filtration after primary leaching in step S1, respectively; "scandium extraction residue" and "vanadium extraction residue" refer to the aqueous phase obtained after phase separation in the corresponding extraction steps, respectively; and "secondary leachate" refers to the leachate obtained after processing the primary leachate residue in step S6.

[0037] It should be noted that the steps in this invention are connected according to the material flow direction; wherein, the primary leaching solution enters the cascade separation route of iron, scandium, vanadium and manganese, and the primary leaching residue enters the niobium and tantalum recovery route; both routes start from the material obtained in step S1, but their subsequent processing objects and target metals are different.

[0038] Compared with the prior art, the present invention can achieve the following beneficial effects:

[0039] 1. This invention controls the concentration of hydrochloric acid in the primary leaching stage, allowing iron, manganese, scandium, and vanadium to preferentially enter the primary leaching solution, while keeping niobium and tantalum mainly in the primary leaching residue. This creates solid-liquid separation of different metal components during the primary leaching stage, avoiding the simultaneous entry of all valuable metals into the same leaching solution and thus reducing the load on subsequent separation.

[0040] 2. The present invention performs neutralization and iron precipitation, scandium extraction and separation, vanadium extraction and separation, and neutralization and manganese precipitation sequentially on the primary leachate; each treatment step is set sequentially according to the separation conditions of different metals, and crude iron hydroxide precipitate, scandium-containing organic phase, vanadium-containing organic phase, and crude manganese hydroxide precipitate can be obtained from the primary leachate respectively.

[0041] 3. In this invention, a scandium extraction system composed of di(2-ethylhexyl) phosphate and tributyl phosphate is continuously set up with a vanadium extraction system composed of secondary carbon amine and secondary octanol; the scandium extract residue can enter the vanadium extraction step without the need for acid or alkali adjustment between the two extraction steps, thereby reducing intermediate pH adjustment operations.

[0042] 4. The present invention further uses concentrated acid to perform secondary leaching on the primary leaching residue; preferably, concentrated sulfuric acid is used for pretreatment followed by water leaching, so that the niobium and tantalum retained in the solid phase in the primary leaching stage enter the secondary leaching solution, and niobium-tantalum crude concentrate is obtained through subsequent neutralization and precipitation.

[0043] 5. This invention forms an overall process route of "one-time dilute acid leaching and diversion - one-time leaching solution staged separation - one-time leaching residue enhanced leaching"; this route can process iron, manganese, scandium, vanadium, niobium and tantalum in the same titanium metallurgical chlorination tailings in stages, which is conducive to improving the comprehensive utilization of multiple valuable metals in the tailings and realizing the resource utilization of solid waste. Attached Figure Description

[0044] Figure 1 This is a process flow diagram of a method for comprehensive utilization of multiple metals in titanium metallurgical chlorination tailings according to the present invention. Detailed Implementation

[0045] The present application will now be described in further detail with reference to the accompanying drawings. It should be noted that the following specific embodiments are only used to further illustrate the present application and should not be construed as limiting the scope of protection of the present application. Those skilled in the art can make some non-essential improvements and adjustments to the present application based on the above application content.

[0046] It should be noted that the metal content in the following examples is based on the dry basis of the corresponding sample; the leaching rate, extraction rate and precipitation rate are calculated according to the mass change of the corresponding metal in the material before and after treatment; the liquid-to-solid ratio in each example is expressed as the ratio of liquid volume to dry basis solid mass.

[0047] In this embodiment, the titanium metallurgical chlorination tailings are taken from the tailings formed during the production of titanium dioxide by the chlorination process; the specific dry basis composition of the chlorination tailings used in Examples 1 to 3 is shown in each example; the hydrochloric acid, sulfuric acid, NaOH, CaCO3, Ca(OH)2 and MgO used are all commercially available reagents that can meet the requirements of hydrometallurgical tests, and are prepared according to the concentration required for each step.

[0048] Furthermore, the organic reagents used for extraction include di(2-ethylhexyl) phosphate, tributyl phosphate, secondary carbon amine N1923, sec-octanol, and sulfonated kerosene; wherein, di(2-ethylhexyl) phosphate and tributyl phosphate are used to prepare organic phase B, secondary carbon amine N1923 and sec-octanol are used to prepare organic phase C, and sulfonated kerosene is used as an organic phase diluent.

[0049] Furthermore, carbide slag is used as a neutralizing agent in some embodiments; the carbide slag has Ca(OH)2 as its main alkaline component; it is added gradually according to the pH change of the solution to be treated in order to achieve the pH range set in the corresponding steps.

[0050] It should be noted that the specific manufacturers and brands of the experimental materials do not constitute a limitation of the technical solution of this invention; similar commercially available materials may be used, provided that the corresponding effective components and functions are met.

[0051] In this embodiment, the contents of Fe, Mn, Sc, V, Nb, and Ta in the raw materials, leachate, leachate residue, raffinate, and precipitate were determined using corresponding elemental quantitative analysis methods; major components such as SiO2 were determined using corresponding component analysis methods; and all samples in the same set of examples and comparative examples used the same sample pretreatment method and detection method.

[0052] Furthermore, the pH of the solutions in each step was measured using a calibrated pH meter; the concentrations of hydrochloric acid and sulfuric acid were calculated according to the prepared concentrations; and the A / O ratio during extraction was the ratio of the volume of the aqueous phase to the volume of the organic phase.

[0053] Furthermore, the leaching rate of the target element is calculated as "the mass of the target element in the leachate / the mass of the target element in the solid raw material before leaching × 100%"; when the leaching residue mass balance is used for verification, the ratio of the difference in the mass of the target element in the raw material and the leaching residue to the mass of the target element in the raw material is used for verification.

[0054] Furthermore, the extraction rate of the target element is calculated as "the difference between the mass of the target element in the aqueous phase before extraction and the mass of the target element in the raffinate / the mass of the target element in the aqueous phase before extraction × 100%"; the precipitation rate of the target element is calculated as "the difference between the mass of the target element in the liquid phase before precipitation and the mass of the target element in the filtrate after filtration / the mass of the target element in the liquid phase before precipitation × 100%".

[0055] Furthermore, the Fe and Mn contents in crude ferric hydroxide and crude manganese hydroxide were converted according to the stoichiometric relationship between Fe(OH)3 and Mn(OH)2 after testing; the Nb and Ta contents in niobium-tantalum crude concentrate were converted according to the stoichiometric relationship between Nb2O5 and Ta2O5, and the sum of the two converted values ​​was recorded as the content in terms of (Nb,Ta)2O5.

[0056] It should be noted that the above detection and calculation methods are used to ensure that the data between different embodiments and comparative examples have a unified basis for comparison; for the same set of experiments, each sample is measured using the same pretreatment and detection aperture.

[0057] Example 1

[0058] In this embodiment, chlorinated tailings generated during the production of titanium dioxide using a certain chlorination process are used as the raw material for treatment. The dry basis composition of the chlorinated tailings is Fe 18.2 wt.%, Mn 5.4 wt.%, Nb+Ta 0.32 wt.%, Sc 0.028 wt.%, V 0.25 wt.%, and SiO2 21.6 wt.%, of which Nb content is 0.28 wt.% and Ta content is 0.04 wt.%. The remaining components mainly include TiO2, Al2O3, and CaO.

[0059] Further, a dilute acid leaching was performed; hydrochloric acid with a concentration of 8 wt.% was added to the chlorinated tailings, the liquid-to-solid ratio was controlled at 3.5 L / kg, and the mixture was stirred and leached at 80°C for 80 min; after leaching, the mixture was filtered to obtain a primary leaching solution and a primary leaching residue; the leaching rates of Fe, Mn, Sc, and V were 91.2%, 92.5%, 89.3%, and 90.8%, respectively.

[0060] Further, the primary leachate was neutralized to precipitate iron; carbide slag was added to the primary leachate and the pH was adjusted to 3.5; after filtration, scandium-vanadium solution and crude iron hydroxide precipitate were obtained; the Fe precipitation rate was 97.5%, and the Fe(OH)3 content in the dry basis of the obtained crude iron hydroxide precipitate was 86.4 wt.%.

[0061] Further, the obtained scandium-vanadium solution was subjected to scandium extraction; organic phase B consisted of 10 vol.% di(2-ethylhexyl) phosphate, 3.5 vol.% tributyl phosphate and the remainder sulfonated kerosene; the A / O ratio was controlled at 7:1 for extraction; after phase separation, a scandium-containing organic phase and scandium extraction residue were obtained, and the single-stage extraction rate of Sc was 93.4%.

[0062] Furthermore, the obtained scandium extraction residue was directly used for vanadium extraction without adjusting the pH with acid or alkali; the organic phase C consisted of 10 vol.% secondary carbon primary amine N1923, 3.5 vol.% secondary octanol and the remainder sulfonated kerosene; the A / O ratio was controlled at 5:1 for extraction; after phase separation, a vanadium-containing organic phase and vanadium extraction residue were obtained, with a single-stage extraction rate of 94.2%.

[0063] Further, the vanadium extraction residue was neutralized and manganese was precipitated; carbide slag was added to the vanadium extraction residue and the pH was adjusted to 10.0; after filtration, crude manganese hydroxide precipitate was obtained; the Mn precipitation rate was 91.5%, and the content of Mn(OH)2 in the dry basis of the obtained crude manganese hydroxide precipitate was 77.8 wt.%.

[0064] Furthermore, the primary leaching residue was subjected to a secondary concentrated acid leaching. Sulfuric acid with a concentration of 93 wt.% was added to the primary leaching residue at a liquid-to-solid ratio of 1.3 L / kg, and the reaction was allowed to proceed for 30 min. Subsequently, water was added to adjust the liquid-to-solid ratio to 2.5 L / kg, and the mixture was stirred and leached at 75°C for 30 min. After leaching, the residue was filtered to obtain the secondary leachate. The leaching rates of Nb and Ta were 88.2% and 86.9%, respectively.

[0065] Further, the secondary leachate was neutralized to precipitate niobium and tantalum; a 30 wt.% NaOH solution was added to the secondary leachate to adjust the pH to 2.0; after filtration, a niobium-tantalum crude concentrate was obtained; the precipitation rates of Nb and Ta were 92.5% and 91.8%, respectively, and the content of (Nb,Ta)₂O₅ in the dry basis of the obtained niobium-tantalum crude concentrate was 2.6 wt.%.

[0066] With this setup, this embodiment uses a single dilute acid leaching process to form two processing routes: a primary leaching solution and a primary leaching residue. The primary leaching solution sequentially separates Fe, Sc, V, and Mn, while the primary leaching residue undergoes a second concentrated acid leaching process and neutralization precipitation to obtain niobium-tantalum crude concentrate.

[0067] Example 2

[0068] In this embodiment, chlorinated tailings are used as the raw material for treatment; the dry basis composition of the chlorinated tailings is Fe 16.9 wt.%, Mn 4.8 wt.%, Nb+Ta 0.26 wt.%, Sc 0.019 wt.%, V 0.17 wt.%, SiO2 19.3 wt.%; wherein, the Nb content is 0.24 wt.% and the Ta content is 0.02 wt.%; the remaining components mainly include TiO2, Al2O3 and CaO.

[0069] Further, 6 wt.% hydrochloric acid was added to the chlorinated tailings; the liquid-to-solid ratio was controlled at 2.5 L / kg, and the mixture was stirred and leached at 70℃ for 90 min; after filtration, a primary leaching solution and a primary leaching residue were obtained; the leaching rates of Fe, Mn, Sc and V were 85.6%, 86.3%, 85.2% and 86.0%, respectively.

[0070] Further, CaCO3 was added to the primary leachate and the pH was adjusted to 3.4; after filtration, scandium-vanadium solution and crude iron hydroxide precipitate were obtained; the Fe precipitation rate was 97.0%, and the Fe(OH)3 content in the dry basis of the crude iron hydroxide precipitate was 85.5 wt.%.

[0071] Furthermore, the scandium-vanadium solution was extracted using organic phase B, which consisted of 8 vol.% di(2-ethylhexyl) phosphate, 3 vol.% tributyl phosphate, and the remainder sulfonated kerosene; the A / O ratio was controlled at 6:1; after phase separation, a scandium-containing organic phase and scandium extraction residue were obtained, and the single-stage extraction rate of Sc was 92.0%.

[0072] Furthermore, without adjusting the pH of the scandium extraction residue with acid or alkali, vanadium extraction was directly performed using organic phase C, which consisted of 8 vol.% secondary carbon amine N1923, 3 vol.% secondary octanol, and the remainder sulfonated kerosene; the A / O ratio was controlled at 4:1; after phase separation, a vanadium-containing organic phase and vanadium extraction residue were obtained, with a single-stage extraction rate of 92.3%.

[0073] Further, carbide slag was added to the vanadium extraction residue and the pH was adjusted to 9.8; after filtration, crude manganese hydroxide precipitate was obtained; the Mn precipitation rate was 90.2%, and the content of Mn(OH)2 in the crude manganese hydroxide precipitate on a dry basis was 76.1 wt.%.

[0074] Further, sulfuric acid with a concentration of 90 wt.% was added to the primary leaching residue at a liquid-to-solid ratio of 1.25 L / kg, and the reaction was carried out for 20 min; then water was added to adjust the liquid-to-solid ratio to 2.0 L / kg, and the mixture was stirred and leached at 60 °C for 40 min; after filtration, a secondary leaching solution was obtained; the leaching rates of Nb and Ta were 85.4% and 85.1%, respectively.

[0075] Further, a 30 wt.% NaOH solution was added to the secondary leachate and the pH was adjusted to 1.8; after filtration, a niobium-tantalum crude concentrate was obtained; the precipitation rates of Nb and Ta were 90.5% and 90.3%, respectively, and the content of (Nb,Ta)₂O₅ in the dry basis of the niobium-tantalum crude concentrate was 2.2 wt.%.

[0076] With this configuration, this embodiment can still recover the corresponding valuable metals according to the primary leaching solution route and the primary leaching residue route, even under the condition of lower values ​​within the preferred parameter range.

[0077] Example 3

[0078] In this embodiment, chlorinated tailings are used as the raw material for treatment; the dry basis composition of the chlorinated tailings is Fe 21.3 wt.%, Mn 5.8 wt.%, Nb+Ta 0.35 wt.%, Sc 0.032 wt.%, V 0.33 wt.%, SiO2 24.6 wt.%; wherein, the Nb content is 0.30 wt.% and the Ta content is 0.05 wt.%; the remaining components mainly include TiO2, Al2O3, CaO and residual carbon.

[0079] Further, hydrochloric acid with a concentration of 9 wt.% was added to the chlorinated tailings; the liquid-to-solid ratio was controlled at 4.0 L / kg, and the mixture was stirred and leached at 90°C for 60 min; after filtration, a primary leaching solution and a primary leaching residue were obtained; the leaching rates of Fe, Mn, Sc and V were 93.0%, 94.1%, 92.0% and 93.5%, respectively.

[0080] Further, Ca(OH)2 was added to the primary leachate and the pH was adjusted to 3.7; after filtration, scandium-vanadium solution and crude iron hydroxide precipitate were obtained; the Fe precipitation rate was 98.2%, and the Fe(OH)3 content in the crude iron hydroxide precipitate on a dry basis was 87.5 wt.%.

[0081] Furthermore, the scandium-vanadium solution was extracted using organic phase B, which consisted of 12 vol.% di(2-ethylhexyl) phosphate, 4 vol.% tributyl phosphate, and the remainder sulfonated kerosene; the A / O ratio was controlled at 8:1; after phase separation, a scandium-containing organic phase and scandium extraction residue were obtained, and the single-stage extraction rate of Sc was 94.0%.

[0082] Furthermore, without adjusting the pH of the scandium extraction residue with acid or alkali, vanadium extraction was directly performed using organic phase C, which consisted of 12 vol.% secondary carbon amine N1923, 4 vol.% secondary octanol, and the remainder sulfonated kerosene; the A / O ratio was controlled at 6:1; after phase separation, a vanadium-containing organic phase and vanadium extraction residue were obtained, with a single-stage extraction rate of 95.0%.

[0083] Further, carbide slag was added to the vanadium extraction residue and the pH was adjusted to 10.5; after filtration, crude manganese hydroxide precipitate was obtained; the Mn precipitation rate was 93.0%, and the content of Mn(OH)2 in the crude manganese hydroxide precipitate on a dry basis was 78.5 wt.%.

[0084] Further, sulfuric acid with a concentration of 95 wt.% was added to the primary leaching residue at a liquid-to-solid ratio of 1.5 L / kg and reacted for 40 min; then water was added to adjust the liquid-to-solid ratio to 3.0 L / kg, and leaching was carried out by stirring at 90°C for 20 min; after filtration, a secondary leaching solution was obtained; the leaching rates of Nb and Ta were 90.2% and 89.5%, respectively.

[0085] Further, a 30 wt.% NaOH solution was added to the secondary leachate and the pH was adjusted to 2.5; after filtration, a niobium-tantalum crude concentrate was obtained; the precipitation rates of Nb and Ta were 93.0% and 92.5%, respectively, and the content of (Nb,Ta)₂O₅ in the dry basis of the niobium-tantalum crude concentrate was 2.4 wt.%.

[0086] With this configuration, this embodiment can achieve the stepwise separation of Fe, Sc, V and Mn in the primary leaching solution under the condition of higher values ​​within the preferred parameter range, and further recover Nb and Ta from the primary leaching residue.

[0087] Comparative Example 1

[0088] This comparative example uses the same raw materials and other process conditions as Example 1; the difference is that the concentration of hydrochloric acid used in the first leaching was adjusted from 8 wt.% to 1.0 wt.%.

[0089] In this comparative example, the leaching rates of Fe, Mn, Sc, and V after one leaching were 18.4%, 15.2%, 31.5%, and 38.7%, respectively. Due to the low acidity of the first leaching, Fe, Mn, Sc, and V could not fully enter the first leaching solution, and the relative enrichment of Nb and Ta in the first leaching residue was also affected. The content of (Nb,Ta)₂O₅ in the dry basis of the subsequently obtained niobium-tantalum crude concentrate was approximately 0.9 wt.

[0090] Therefore, when the concentration of hydrochloric acid in the primary leaching is lower than the range defined in this invention, it is difficult to achieve effective separation of Fe, Mn, Sc, and V from the primary leaching residue; thus, it is not conducive to the implementation of the two treatment routes of the primary leaching solution and the primary leaching residue.

[0091] Comparative Example 2

[0092] This comparative example uses the same raw materials and other process conditions as Example 1; the difference is that the concentration of hydrochloric acid used in the first leaching was adjusted from 8 wt.% to 15 wt.%.

[0093] In this comparative example, the leaching rates of Fe, Mn, Sc, and V after one leaching were 93.5%, 94.0%, 92.8%, and 93.7%, respectively; meanwhile, the leaching rates of Nb and Ta reached 11.5% and 9.2%, respectively. Some Nb and Ta entered the primary leaching solution and dispersed during subsequent neutralization, iron precipitation, and extraction treatments, resulting in a decrease in the retention of Nb and Ta in the primary leaching residue. The final niobium-tantalum crude concentrate had a dry basis content of 1.4 wt.% based on (Nb,Ta)₂O₅.

[0094] Therefore, when the concentration of hydrochloric acid in the primary leaching is higher than the range defined in this invention, although Fe, Mn, Sc and V can enter the primary leaching solution, Nb and Ta will also increase their proportion entering the primary leaching solution; thus weakening the selective diversion effect of step S1 on different metal components.

[0095] Comparative Example 3

[0096] This comparative example uses the same raw materials as Example 1 and follows steps S1 to S6 as in Example 1; the difference is that in step S7, the pH of the secondary leachate is adjusted from 2.0 to 4.0.

[0097] In this comparative example, the precipitation rates of Nb and Ta were 95.6% and 94.8%, respectively; however, as the pH increased, more impurities such as Fe and Al in the secondary leachate underwent co-precipitation; the content of (Nb,Ta)₂O₅ in the dry basis of the obtained precipitate was 0.86 wt.%, and the filtration performance of the obtained precipitate decreased.

[0098] Therefore, the pH of the secondary leaching solution in step S7 not only affects the precipitation degree of Nb and Ta, but also affects the co-precipitation of impurity components; controlling the pH in the range of 1.5 to 3.0 is beneficial to obtaining precipitates with a high degree of niobium-tantalum enrichment.

[0099] To facilitate comparison of the separation and enrichment of valuable metals under different process conditions, the main detection results of Examples 1-3 and Comparative Examples 1-3 are summarized in Table 1; where “—” indicates that the data was not detected or recorded separately in the corresponding example or comparative example.

[0100] Table 1 Summary of main detection results for examples and comparative examples

[0101]

[0102] Based on Table 1 above, as well as the various embodiments and comparative examples, the following conclusions can be drawn:

[0103] As shown in Table 1, when Fe, Mn, Sc and V were leached once using 6-9 wt.% hydrochloric acid in Examples 1-3, they all had high leaching rates. After further treatment of the leaching solution by neutralization and precipitation of iron, extraction of scandium, extraction of vanadium and neutralization and precipitation of manganese, the corresponding enriched products of iron, scandium, vanadium and manganese could be obtained respectively. At the same time, after secondary leaching and neutralization precipitation of the leaching residue with concentrated sulfuric acid, niobium-tantalum crude concentrate could be obtained.

[0104] Furthermore, compared with Example 1, Comparative Example 1 showed a significant decrease in the primary leaching rate of Fe, Mn, Sc and V after reducing the concentration of hydrochloric acid in the primary leaching to 1 wt.%. This indicates that when the primary leaching acidity is too low, it is not conducive to the entry of the above metals into the primary leaching solution and affects the separation and processing of the subsequent two material routes.

[0105] It should be noted that in Comparative Example 2, after increasing the hydrochloric acid concentration in the primary leaching to 15 wt.%, although Fe, Mn, Sc, and V maintained high leaching rates, the primary leaching rates of Nb and Ta reached 11.5% and 9.2%, respectively, and the content of (Nb,Ta)₂O₅ in the final niobium-tantalum crude concentrate decreased to 1.4 wt.%. This indicates that excessively high acidity in the primary leaching increases the proportion of Nb and Ta entering the primary leaching solution, which is not conducive to the retention and subsequent enrichment of Nb and Ta in the primary leaching residue.

[0106] Furthermore, in Comparative Example 3, after increasing the precipitation pH of the secondary leachate to 4.0, although the precipitation rates of Nb and Ta increased, the content of (Nb,Ta)₂O₅ in the resulting precipitate decreased to 0.86 wt.%, and a large amount of Fe, Al and other impurities co-precipitated. This indicates that the precipitation pH of the secondary leachate affects both the precipitation degree of niobium and tantalum and the enrichment degree of the resulting precipitate.

[0107] With this setup, Examples 1-3 and Comparative Examples 1-3 verified the technical solution of the present invention from two aspects: the acidity of the primary leaching and the pH of the niobium-tantalum precipitation. The control of the hydrochloric acid concentration in the primary leaching is conducive to the formation of solid-liquid separation between Fe, Mn, Sc, V and Nb, Ta, and subsequently forms a primary leaching liquid cascade separation route and a primary leaching residue niobium-tantalum recovery route.

[0108] The above-described embodiments are merely illustrative of several implementations of the present invention, and while the descriptions are relatively specific and detailed, they should not be construed as limiting the scope of the present invention. It should be noted that those skilled in the art can make various modifications and improvements without departing from the concept of the present invention, and these modifications and improvements all fall within the scope of protection of the present invention.

Claims

1. A method for comprehensive utilization of multiple metals in titanium metallurgical chlorination tailings, characterized in that, Includes the following steps: S1. Add a 2-10 wt.% hydrochloric acid solution to the chlorinated tailings produced during the chloride process of titanium dioxide production, and carry out a first leaching under heating and stirring conditions. Filter to obtain a first leaching solution and a first leaching residue. S2. Add neutralizing agent A to the primary leachate, adjust the pH to 3.0-4.0, filter, and obtain scandium-vanadium solution and crude ferric hydroxide precipitate; S3. The scandium-vanadium solution is extracted using organic phase B, and phase separation is performed to obtain a scandium-containing organic phase and scandium extraction residue; the organic phase B consists of 5-15 vol.% di(2-ethylhexyl) phosphate, 2-5 vol.% tributyl phosphate and the balance sulfonated kerosene; S4. The scandium extraction residue is extracted using organic phase C, and phase separation is performed to obtain a vanadium-containing organic phase and a vanadium extraction residue; the organic phase C consists of 5-15 vol.% secondary carbon primary amine extractant, 2-5 vol.% secondary octanol and the balance sulfonated kerosene; S5. Add neutralizing agent A to the vanadium extraction residue, adjust the pH to 9.5-10.5, filter, and obtain crude manganese hydroxide precipitate; S6. Add concentrated acid to the primary leaching residue, and perform secondary leaching under heating and stirring conditions. Filter to obtain secondary leachate; the concentrated acid is a hydrochloric acid solution or sulfuric acid solution with a concentration >20 wt.%. S7. Add neutralizing agent A to the secondary leachate, adjust the pH to 1.5-3.0, filter, and obtain niobium-tantalum crude concentrate; The neutralizing agent A is selected from at least one of Ca(OH)2, CaCO3, MgO, NaOH, and carbide slag.

2. The method for comprehensive utilization of multiple metals in titanium metallurgical chlorination tailings according to claim 1, characterized in that, The dry basis composition of the chlorinated tailings satisfies: Fe > 15 wt.%, Mn > 4 wt.%, 0.2 wt.% < (Nb + Ta) < 0.4 wt.%, 0.015 wt.% < Sc < 0.04 wt.%, 0.15 wt.% < V < 0.40 wt.%, SiO2 > 15 wt.%.

3. The method for comprehensive utilization of multiple metals in titanium metallurgical chlorination tailings according to claim 1, characterized in that, In step S1, the concentration of hydrochloric acid is 6-9 wt.%; the liquid-to-solid ratio of a single leaching is 2.5-4.0 L / kg; the leaching temperature is 70-90℃; and the leaching time is 60-90 min.

4. The method for comprehensive utilization of multiple metals in titanium metallurgical chlorination tailings according to claim 3, characterized in that, In step S2, the pH is adjusted to 3.4-3.7; the neutralizing agent A is carbide slag with Ca(OH)2 as the main alkaline component.

5. The method for comprehensive utilization of multiple metals in titanium metallurgical chlorination tailings according to claim 1, characterized in that, In step S3, the organic phase B consists of 8-12 vol.% di(2-ethylhexyl) phosphate, 3-4 vol.% tributyl phosphate, and the balance sulfonated kerosene; the volume ratio A / O of the scandium vanadium solution to the organic phase B is 6:1 to 8:

1.

6. The method for comprehensive utilization of multiple metals in titanium metallurgical chlorination tailings according to claim 1, characterized in that, In step S4, the secondary carbon primary amine extractant is secondary carbon primary amine N1923, and the organic phase C consists of 8-12 vol.% secondary carbon primary amine N1923, 3-4 vol.% secondary octanol and the balance sulfonated kerosene; the volume ratio A / O of the scandium extract residue to the organic phase C is 4:1 to 6:

1.

7. The method for comprehensive utilization of multiple metals in titanium metallurgical chlorination tailings according to claim 1, characterized in that, In step S5, the pH is adjusted to 9.8-10.5; the neutralizing agent A is carbide slag with Ca(OH)2 as the main alkaline component.

8. The method for comprehensive utilization of multiple metals in titanium metallurgical chlorination tailings according to claim 1, characterized in that, In step S6, the liquid-to-solid ratio for the secondary leaching is 1.2–3.0 L / kg, and the leaching time is 30–120 min.

9. The method for comprehensive utilization of multiple metals in titanium metallurgical chlorination tailings according to claim 8, characterized in that, In step S6, the secondary leaching is carried out using sulfuric acid; a sulfuric acid solution with a concentration of 90-95 wt.% is added to the primary leaching residue at a liquid-to-solid ratio of 1.25-1.50 L / kg, and the reaction is carried out for 20-40 min; then water is added to adjust the liquid-to-solid ratio to 2.0-3.0 L / kg, and the mixture is stirred and leached at 60-90℃ for 20-40 min, then filtered to obtain the secondary leaching solution.

10. The method for comprehensive utilization of multiple metals in titanium metallurgical chlorination tailings according to claim 9, characterized in that, In step S7, the pH of the secondary leachate is adjusted to 1.8-2.5, and then filtered to obtain niobium-tantalum crude concentrate.

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