Method for preparing low-thallium chloride product by deep thallium removal from steel solid waste

Abstract

This invention relates to the field of steel solid waste resource utilization and heavy metal pollution control technology, specifically to a method for preparing low-thallium chloride products by deep dethallium removal from steel solid waste. The method uses sintering machine head ash, blast furnace bag ash, coke powder, and water as raw materials, and processes them through steps such as raw material treatment, coarse dethallium removal, heavy metal removal, deep dethallium removal, calcium and magnesium removal, extreme dethallium removal, and potassium and sodium product preparation. This yields potassium chloride and sodium chloride products with extremely low thallium content, and significantly reduces the thallium content in the circulating slag. By specifically setting the location, step details, and parameters for each step of coarse, deep, and extreme dethallium removal, the purity of the potassium and sodium salt products and the stability of the production process are ensured, demonstrating promising industrial application prospects.

Description

Technical Field

[0001] This invention relates to the field of resource utilization of steel solid waste and heavy metal pollution control, specifically to a method for preparing low-thallium chloride products by deep dethallium removal from steel solid waste. Background Technology

[0002] Steelmaking solid waste, such as sintering ash, blast furnace bag ash, and zinc-containing dust and sludge, typically contains valuable elements like Fe, Zn, K, and Na, possessing high resource utilization value. Current technologies for this type of steelmaking solid waste usually employ a combination of pyrometallurgical enrichment and wet separation to recover valuable components like zinc, potassium, and sodium. However, thallium (Tl), which is associated with steelmaking solid waste, can migrate and accumulate during roasting, dust removal, rinsing, purification, and evaporation crystallization processes, entering the process liquid phase, solid phase, and final products, thus posing environmental and product quality risks.

[0003] Thallium is a highly toxic and bioaccumulative heavy metal. Trace amounts of thallium in steel production raw materials are prone to volatilization and speciation under high-temperature conditions, and can enter different media during subsequent dust removal, desulfurization, and wet treatment processes. Especially in the wet resource recovery of steel solid waste, thallium may migrate with rinsing liquid, heavy metal filtrate, calcium and magnesium removal filtrate, and evaporation mother liquor, and remain in subsequent potassium and sodium salt products. Therefore, achieving effective control and deep removal of thallium in the main resource recovery process is one of the key issues in the high-value utilization of steel solid waste.

[0004] Existing technologies for thallium treatment focus on the effective transfer and enrichment of thallium. For example, Chinese patent application CN117861430A discloses a "process for enriching thallium in sintering machine head flue gas into machine head ash." This technology reuses thallium-containing wastewater from wet desulfurization and wet electrostatic precipitator systems for sintering feed, allowing thallium to circulate within the sintering-dust removal system and ultimately enrich in the machine head ash. This method can achieve the reuse of thallium-containing wastewater and reduce wastewater discharge, but its technical objective is to transfer thallium from the liquid phase and enrich it in the solid phase; it does not achieve substantial removal of thallium within the system, and the enriched high-thallium machine head ash still requires further treatment.

[0005] Chinese invention patent publication CN103693819B discloses a method for the deep treatment of thallium-containing wastewater. This method involves the deep treatment of thallium in the wastewater through starch hydrolysis, sodium hydroxide, biological thallium removal agents, and alkali addition and stirring. However, because this method treats thallium-containing wastewater, it is not suitable for the overall aquatic environment of steel solid waste recycling. In other words, directly applying the above-mentioned existing technology to the wet resource recovery process of steel solid waste still has certain shortcomings. The wet resource recovery process of steel solid waste involves many types of liquid phases, and the liquid phases at different process nodes vary greatly in terms of salinity, alkalinity, pH, and the composition of coexisting ions. For example, the rinsing liquid, heavy metal removal filtrate, and calcium and magnesium removal filtrate in the potassium and sodium recovery section all have different system characteristics, and the existence state of thallium in each liquid phase and the difficulty of removal also differ. Existing enrichment or end-of-pipe treatment technologies for desulfurization wastewater typically target a single object and lack systematic control schemes for the multi-node liquid phase in the main process of steel solid waste resource utilization. They also lack segmented thallium removal designs that are coordinated with the main potassium and sodium recovery processes.

[0006] Furthermore, existing technologies primarily focus on wastewater discharge control or the centralized transfer of thallium-containing pollutants, with less attention paid to the residual control of thallium in resource-based products such as potassium and sodium salts. If wastewater is treated only at the end of the process, it is difficult to prevent thallium from entering the product phase during the upstream impurity removal, salt separation, and crystallization processes, thereby affecting the quality of the final product.

[0007] Therefore, existing technologies lack a multi-node thallium removal method suitable for the main process of wet resource recovery of steel solid waste, especially for the potassium and sodium recovery stage, so as to control and remove thallium step by step at different liquid phase nodes and effectively reduce the thallium content in the final sodium chloride and potassium chloride products. Summary of the Invention

[0008] To address the aforementioned shortcomings of existing technologies, the present invention aims to provide a three-step thallium removal method for the potassium and sodium recovery stage in the treatment of steel solid waste. This method integrates a thallium removal unit into the main process of potassium and sodium recovery from steel solid waste. Different methods of coarse thallium removal, deep thallium removal, and extreme thallium removal are sequentially implemented at three key nodes: the rinsing filtrate, the heavy metal removal filtrate, and the calcium and magnesium removal filtrate. This forms a graded reduction and step-by-step interception thallium removal technology route, reducing the migration and accumulation of thallium in the process liquid phase, and ultimately obtaining low-thallium sodium chloride and low-thallium potassium chloride products.

[0009] To achieve the above objectives, the present invention adopts the following technical solution: A method for preparing low-thallium chloride products by deep dethallium removal from steel solid waste includes the following steps: I. Raw material processing: Sintering machine head ash, blast furnace bag ash, coke powder and water are mixed evenly and granulated. After enrichment by pyrometallurgy, zinc-rich dust and roasting residue are obtained. The zinc-rich dust is washed to obtain rinsing filter residue and rinsing filtrate with thallium content of 40~43mg / L.

[0010] II. Crude thallium removal: The rinsing filtrate obtained in step I is mixed with the biosorbent at a liquid-solid ratio of (7~9) ml: 1 g. The adsorption reaction is carried out at 18~23℃ and pH 7.5~8.5 for 50~75 min. Then, solid-liquid separation is performed to obtain crude thallium removal filtrate and thallium-containing precipitate residue.

[0011] III. Gravity removal treatment: Add a precipitant to the crude thallium-removed filtrate obtained in step II and adjust the pH to alkaline to perform gravity removal treatment, thereby obtaining gravity-removed filtrate and gravity-removed filter cake.

[0012] IV. Deep thallium removal: Add sodium sulfide to the filtrate obtained in step III at a solid-liquid ratio of (0.2~0.3) g: 1 L, stir at 22~28℃ for 8~12 min to carry out sulfidation reaction, then let stand and age for 25~35 min, and then perform solid-liquid separation to obtain deep thallium removal filtrate and thallium-containing precipitate residue.

[0013] V. Calcium and magnesium removal treatment: Add calcium and magnesium removal agents to the deep thallium removal filtrate obtained in step IV to form precipitates of calcium, magnesium and silicon impurities in the filtrate. After solid-liquid separation, calcium and magnesium removal filtrate and calcium and magnesium removal filter residue are obtained.

[0014] VI. Limiting thallium removal: Sodium sulfide and sodium carbonate are added sequentially to the calcium- and magnesium-free filtrate obtained in step V at solid-liquid ratios of (0.2~0.29) g: 1 L and (2.9~3.5) g: 1 L. After stirring at 22~28℃ for 8~12 min, the mixture is allowed to stand for 55~65 min, and then solid-liquid separation is performed to obtain the limiting thallium-free filtrate and the thallium-containing precipitate.

[0015] VII. Preparation of potassium and sodium products: The thallium-free filtrate obtained in step VI is evaporated and crystallized to obtain sodium chloride and potassium chloride products, wherein the thallium content in the sodium chloride product is 0.005-0.02 mg / kg and the thallium content in the potassium chloride product is 0.005-0.03 mg / kg.

[0016] Preferably, the thallium content in the filtrate obtained in step III is 10~12 mg / L, and the thallium content in the filter cake is 1200~1300 mg / kg.

[0017] Preferably, the thallium content in the calcium-magnesium-removed filtrate obtained in step V is 7.02~8.38 mg / L, and the thallium content in the calcium-magnesium-removed filter residue is 140~160 mg / kg.

[0018] Preferably, in step I, the sintering machine head ash, blast furnace bag ash, coke powder and water are mixed evenly in a weight ratio of (5~10):(1~5):(3~6):(8~12).

[0019] Preferably, the pyrometallurgical enrichment is performed by placing the granulated raw material mixture into a roasting furnace and roasting it at a temperature of 1150~1250℃ for 0.8~1.2 hours to obtain roasting residue and flue gas. The flue gas is then separated from the dust by a dust removal device to obtain zinc-rich dust and dust-removed gas. The gas is then treated by a flue gas treatment device before being discharged.

[0020] Preferably, the rinsing process involves mixing zinc-rich dust with deionized water or 1 wt.% NaOH solution at a ratio of (130~160) g: (550~650) ml, stirring at 22~28℃ and 180~220 rpm for 35~45 min, and then performing vacuum filtration to obtain rinsing residue and rinsing filtrate.

[0021] Preferably, the weight removal process in step III specifically involves adding a precipitant to the crude thallium-removed filtrate obtained in step II at a solid-liquid ratio of 0.3-3 g / L and adjusting the pH to alkaline. After stirring continuously for 20-30 minutes, the mixture is allowed to stand for 30-60 minutes to obtain a weight-removed filtrate and a weight-removed filter cake. The precipitant is one or a mixture of sodium hydroxide and sodium carbonate, with a pH of 9-11. Solid-liquid separation is then achieved through pressure filtration.

[0022] Preferably, the calcium and magnesium removal treatment in step V specifically involves: adding a calcium removal agent to the deep thallium removal filtrate obtained in step IV at a solid-liquid ratio of 0.3~1.8 g / L, stirring at 20~30℃ for 10~15 min, then adding a magnesium removal agent at a solid-liquid ratio of 1~3 g / L, and continuing to stir at the same temperature for 10~15 min under conditions of pH 10~12, and then separating the solid and liquid by filtration to obtain a calcium and magnesium removal filtrate and a calcium and magnesium removal filter residue; wherein the calcium removal agent is one or a mixture of two of sodium carbonate or sodium oxalate, and the magnesium removal agent is sodium hydroxide or potassium hydroxide.

[0023] Preferably, the preparation of potassium and sodium products in step VII is as follows: The thallium-free filtrate obtained in step VI is continuously evaporated and concentrated at 123~126℃. Sodium chloride first reaches saturation and crystallizes out. When the potassium chloride concentration reaches a weight ratio of (60~65):100 with water, it is considered that potassium chloride is close to the saturation solubility at this temperature. The mixture is sent to a centrifuge to separate the lower layer of NaCl crystals and the upper layer of saturated KCl solution. The NaCl crystals are dried at 100~120℃ for 30~60 min to obtain sodium chloride product. The upper layer of saturated KCl solution is introduced into a cooling tank and cooled from 123~126℃ to 10~25℃ at a rate of 1~2℃ / min. After KCl crystallization continues for 4~6 hours, centrifugation is performed to obtain KCl crystals and mother liquor. The mother liquor is returned for re-evaporation and concentration. The KCl crystals are dried at 80~100℃ for 30~45 min to obtain potassium chloride product.

[0024] Preferably, the roasting residue obtained in step I is ground to below 0.074 mm, and then the refined iron powder is separated by magnetic separation. The tailings are then used as raw materials for building materials.

[0025] Preferably, the rinsed filter residue obtained in step I is dried and ground before being used as a raw material in the nano zinc oxide preparation process.

[0026] Preferably, the thallium-containing precipitate obtained in steps II, IV and VI is subjected to sealing and solidification treatment.

[0027] Preferably, the filter cake obtained in step III and the filter residue obtained in step V for removing calcium and magnesium are both returned to step I as recycled raw materials. They are then used together with the granulated raw material mixture as raw materials in the pyrometallurgical enrichment step for mixed roasting.

[0028] Preferably, the stirring speed in steps IV and VI is 130~165 r / min.

[0029] Preferably, the solid-liquid separation described in steps II, IV and VI is performed by plate and frame filter press.

[0030] Preferably, in step II, before mixing the rinsing filtrate with the biosorbent, the following steps are performed: ozone gas is bubbled into the rinsing filtrate for 8-10 minutes to ensure that the ozone concentration is 10-20 mg / L (effective oxidation Tl). + For Tl 3+ (and generate ·OH free radicals to enhance the reaction).

[0031] Preferably, in step VI, the sodium sulfide is microencapsulated sodium sulfide. The microencapsulated sodium sulfide is prepared by first dissolving sodium sulfide in water to a concentration of 0.5-1.0 mol / L, then mixing this sodium sulfide solution with a chitosan-polyacrylic acid (CS-PAA) emulsion at a volume ratio of 1:(4-5), and emulsifying at a stirring speed of 8000-10000 rpm for 10-15 minutes. Then, 0.4-0.6 ml of glutaraldehyde is added per 100 ml of emulsion, and the mixture is slowly emulsified at a stirring speed of 800-1000 rpm. Glutaraldehyde was added dropwise, and the mixture was stirred continuously for 1.8–2.2 hours after addition. The resulting microcapsules were then collected by centrifugation and washed 2–5 times with anhydrous ethanol before vacuum drying to obtain microencapsulated sodium sulfide. (In a low-pH acidic environment, the amino groups of chitosan are protonated, causing the microspheres to expand and the pores to increase, thereby accelerating the release of sodium sulfide. In a high-pH alkaline environment, the microcapsule structure is dense, and the release rate is slowed down, thus playing a "slow-release" role. Therefore, the release is slower under alkaline conditions, achieving intelligent control of "adjusting the release rate according to the environment").

[0032] Preferably, the biosorbent is a known biosorbent such as Sargassum fusiforme, brown algae, freshwater Chlorella, hydrolyzed feather powder, or lignocellulose.

[0033] As a preferred option, the thallium-containing precipitate obtained in step II is subjected to low-temperature pyrolysis to prepare an adsorbent material, which is then reused at the front end to work together with the bioadsorbent as an adsorbent for crude thallium removal.

[0034] As a preferred option, the thallium-containing precipitate obtained in step IV is subjected to electrochemical leaching followed by membrane separation to recover high-purity thallium metal, thereby realizing the resource-based treatment of the hazardous waste precipitate.

[0035] Preferably, the thallium-containing precipitate obtained in step VI is placed in a flowing electrode electrochemical reactor and a DC voltage of 1.8~2.2V is applied to oxidize and regenerate Tl2S into soluble Tl. + They migrate directionally to the cathode chamber and accumulate there; while sulfur ions are reduced to S. 2- Then, the cathode enrichment solution was evaporated and crystallized to obtain high-purity thallium chloride (TlCl) product.

[0036] The technical effects of this invention are as follows: 1. This invention sets the thallium removal process at three key liquid phase nodes: rinsing filtrate, heavy filtrate removal filtrate, and calcium and magnesium filtrate removal filtrate. This breaks away from the existing technology that relies solely on a single impurity removal unit or end-of-line treatment unit to remove thallium. This allows thallium to be gradually reduced before entering subsequent processes, thereby effectively suppressing the migration and accumulation of thallium throughout the entire process. This achieves graded thallium removal and full-process control in the potassium and sodium recovery stage.

[0037] 2. The rinsing filtrate contains a high concentration of thallium and complex impurities. This invention employs a biosorbent for coarse thallium removal, which facilitates rapid reduction of thallium load at the upstream stage. After heavy removal, sodium sulfide precipitation is used in the liquid phase to further reduce the thallium concentration. After calcium and magnesium removal, sodium sulfide and sodium carbonate co-precipitate are used in the liquid phase to achieve ultimate removal of residual thallium. The thallium removal method of this invention is matched with the process nodes, making the treatment more targeted. Furthermore, the three treatment methods specifically designed in this invention focus on different stages of different steps, achieving a progressive effect and enhancing adaptability.

[0038] 3. By moving the thallium control node forward, this invention reduces the enrichment and transfer of thallium during subsequent calcium and magnesium removal and evaporation crystallization processes, allowing the thallium content in the final potassium chloride product to be controlled at 0.005–0.03 mg / kg and the thallium content in the sodium chloride product to be controlled at 0.005–0.02 mg / kg. This significantly improves product quality and increases product added value, while effectively reducing the burden on subsequent processes and improving product quality.

[0039] 4. The bio-adsorption, sulfide precipitation, and co-precipitation processes employed in this invention can all be implemented using existing wet recovery processes with stirring, dosing, and filtration equipment. Only adaptive additions or optimizations are needed on the existing process basis. This has the characteristics of good process compatibility, small equipment modification, and good industrial application prospects. Such processes are easy to integrate into existing production lines and have lower implementation costs. Moreover, the thallium-containing precipitates generated at the three nodes of this invention are all separated in solid form. By collecting and sealing them separately, the continued circulation and migration of thallium in the liquid phase is avoided, which is conducive to subsequent centralized disposal and environmental risk control. It is also conducive to the centralized management of thallium-containing solid phases. Furthermore, in the preferred embodiment, these solid-form thallium-enriched slags are treated differently, realizing the resource utilization of thallium.

[0040] 5. The present invention, through a preferred technical solution, involves ozone oxidation in the crude thallium removal process and microencapsulation of sodium sulfide before its addition in the extreme thallium removal process. This is because ozone, under alkaline conditions, can effectively remove the difficult-to-treat thallium (TH)... + Oxidation into Tl, which is more easily precipitated 3+ It can be more easily adsorbed and removed by biosorbents, Tl 3+The higher charge density results in stronger electrostatic attraction with functional groups such as carboxyl (-COOH), hydroxyl (-OH), and amino (-NH2) groups on the surface of the biosorbent, making it easier for ion exchange or complexation reactions to occur, thereby significantly improving the adsorption capacity of the biosorbent. Furthermore, in the preferred embodiment, sodium sulfide is pre-encapsulated in chitosan-polyacrylic acid (CS-PAA) microcapsules (microcapsule particle size controlled within 20-60 μm). Under the specific alkaline environment of this invention, a slow-release effect can be achieved (under acidic conditions, the microcapsules will release relatively quickly). The release rate of the microcapsules is naturally suppressed under alkaline conditions, avoiding the escape of H2S or non-selective precipitation caused by a large-scale release of sodium sulfide at once. Combined with stirring, a stable and controllable sulfur ion supply is ensured throughout the 8-12 minute stirring process, thereby further enhancing the stability of the reaction and strengthening the dethallium removal effect of the ultimate dethallium removal step. Attached Figure Description

[0041] Figure 1 This is a schematic flowchart of a deep thallium removal method according to one embodiment of the present invention.

[0042] Figure 2 This is a sample image and XRD analysis diagram of potassium chloride product according to one embodiment of the present invention.

[0043] Figure 3 This is a sample image and XRD analysis diagram of sodium chloride product according to one embodiment of the present invention. Detailed Implementation

[0044] The present invention will be further described in detail below with reference to embodiments. It should be understood that the following embodiments are for illustrative purposes only and are not intended to limit the scope of protection of the present invention. Various equivalent substitutions or modifications made by those skilled in the art without departing from the inventive concept should fall within the scope of protection of the present invention.

[0045] Unless otherwise specified, the raw materials, equipment, and testing methods used in the implementation of this invention can all be conventional materials, equipment, and testing methods in the field. Unless otherwise specified, the percentages, ratios, and concentrations mentioned in the embodiments refer to mass percentages, mass-volume ratios, or mass concentrations.

[0046] Example 1

[0047] This embodiment illustrates the specific application of the three-step thallium desulfurization method of the present invention in the potassium and sodium recovery process of steel solid waste. It is a method for deep desulfurization of steel solid waste to prepare low-thallium chloride products. The method includes the following steps: (1) Pyrometallurgical enrichment and rinsing: 8 parts by weight of sintering machine head ash, 3 parts by weight of blast furnace bag ash, 4 parts by weight of coke powder and 10 parts by weight of water were mixed evenly and granulated. The granulated raw material mixture was placed in a roasting furnace and roasted at 1200℃ for 1 hour to obtain roasting residue and flue gas. The flue gas was separated from the dust by a dust removal device to obtain zinc-rich dust and dust-removed gas. The gas was treated by a flue gas treatment device and then discharged. The zinc-rich dust was stirred at 200 rpm for 40 min and then vacuum filtered to obtain rinsed filter residue and rinsed filtrate. The thallium content in the rinsed filtrate was 41.97 mg / L.

[0048] (2) Crude thallium removal: The rinsing filtrate obtained in step (1) is introduced into the crude thallium removal reaction unit. A biosorbent (e.g., hydrolyzed feather powder is used in this embodiment) is added to the rinsing filtrate, wherein the ratio of biosorbent to rinsing filtrate is 1g:8ml. Adsorption is carried out at a temperature of 20℃ and a pH of 8 for 60min. After adsorption, a plate and frame filter press is used for solid-liquid separation to obtain crude thallium removal filtrate and thallium-containing precipitate. The obtained thallium-containing precipitate is sealed and solidified, and the obtained crude thallium removal filtrate enters the gravimetric removal process.

[0049] (3) De-gravity treatment: The crude thallium-removed filtrate obtained in step (2) is introduced into the de-gravity unit. Sodium hydroxide is added to the crude thallium-removed filtrate at a solid-liquid ratio of 1.5 g / L and the pH value is adjusted to 10. After stirring continuously for 25 min, it is allowed to stand for 50 min to allow the heavy metal ions in the liquid phase to form a precipitate, and a de-gravity filtrate and a de-gravity filter cake are obtained. Then, after filtration by a plate and frame filter press, a de-gravity filtrate and a de-gravity filter cake are obtained. The de-gravity filtrate and the de-gravity filter cake are tested and the thallium content in the de-gravity filtrate is 11.68 mg / L and the thallium content in the de-gravity filter cake is 1247 mg / kg.

[0050] (4) Deep thallium removal: The filtrate obtained in step (3) is introduced into the deep thallium removal unit. Sodium sulfide is added to the filtrate at a dosage of 0.24 g / L. The mixture is reacted at 25°C with a stirring rate of 150 r / min for 10 min, and then allowed to stand for 30 min. Solid-liquid separation is then performed using a plate and frame filter press to obtain the deep thallium removal filtrate and the thallium-containing precipitate. The obtained thallium-containing precipitate is sealed and solidified, and the deep thallium removal filtrate enters the calcium and magnesium removal process.

[0051] (5) Calcium and magnesium removal treatment: The deep thallium removal filtrate obtained in step (4) is introduced into the calcium and magnesium removal unit. Sodium oxalate, a calcium removal agent, is added to the deep thallium removal filtrate at a solid-liquid ratio of 1.2 g / L. The mixture is stirred at 25°C for 12 min. Then, sodium hydroxide, a magnesium removal agent, is added at a solid-liquid ratio of 1.1 g / L. The mixture is stirred at pH 11 for 13 min to allow calcium and magnesium impurities (including silicon) in the liquid phase to precipitate. After solid-liquid separation by filtration, calcium and magnesium removal filtrate and calcium and magnesium removal filter residue are obtained. The thallium content of the calcium and magnesium removal filtrate and the calcium and magnesium removal filter residue are tested respectively. The thallium content in the calcium and magnesium removal filter residue is 150 mg / kg, and the thallium content in the calcium and magnesium removal filtrate is 7.65 mg / L.

[0052] (6) Limiting thallium removal: The calcium and magnesium removed filtrate obtained in step (5) is introduced into the limiting thallium removal unit. Sodium sulfide is added at 0.24 g / L at 25°C, followed by sodium carbonate at 3.2 g / L. The mixture is stirred at 150 r / min for 10 min, then allowed to stand for 60 min. Solid-liquid separation is then performed using a plate and frame filter press to obtain the limiting thallium removal filtrate and the thallium-containing precipitate. The obtained thallium-containing precipitate is sealed and solidified, while the limiting thallium removal filtrate is introduced into the evaporation and crystallization process.

[0053] (7) Evaporation crystallization for salt production: The thallium-depleted filtrate obtained in step (6) is sent to the evaporation crystallization system for evaporation concentration and salt separation crystallization. The thallium-depleted filtrate is continuously evaporated and concentrated at 125°C. Sodium chloride first reaches saturation and crystallizes out. When the potassium chloride concentration is detected to reach a weight ratio of about 63:100 with water, it is considered that potassium chloride is close to the saturation solubility at this temperature. The mixed liquid is sent to a centrifuge to separate the lower layer of NaCl crystals and the upper layer of KCl saturated solution. The NaCl crystals are dried at 100~120°C for 50 min to obtain sodium chloride product. The upper layer of KCl saturated solution is introduced into a cooling tank and cooled from 125°C to 20°C at a rate of 1.5°C / min. After KCl crystallization continues for 5 hours, centrifugation is performed to obtain KCl crystals and mother liquor. The mother liquor is returned to be evaporated and concentrated again. The KCl crystals are dried at 80~100°C for 40 min to obtain potassium chloride product. The thallium content of sodium chloride and potassium chloride products was tested, and the results showed that the thallium content in the potassium chloride product was 0.03 mg / kg and the thallium content in the sodium chloride product was 0.02 mg / kg.

[0054] The actual potassium chloride product and its XRD pattern obtained in this embodiment are shown below. Figure 2 As shown, through Figure 2It can be seen that the potassium chloride particles are fine and uniform in color, without obvious lumps or impurities, which meets the typical appearance characteristics of industrial or agricultural potassium chloride. Furthermore, the XRD pattern shows that KCl is the main component of the product, corresponding to multiple high-intensity diffraction peaks. Small amounts of NaCl and sodium potassium sulfate (K3Na(SO4)2) are present, but there are no diffraction peaks for heavy metals. Moreover, all diffraction peaks in the pattern are sharp and of high intensity, indicating that all components in the sample are crystalline with good crystallinity, meeting the composition requirements for industrial or agricultural potassium chloride. The obtained sodium chloride product and its XRD pattern are shown below. Figure 3 As shown, through Figure 3 It can be seen that the sodium chloride crystals are uniform in size, pure in color, and free of discolored particles. Figure 3 The XRD pattern showed only characteristic diffraction peaks of NaCl (sodium chloride), and multiple diffraction peaks were of high intensity and clearly positioned. There were no diffraction peaks of heavy metals, indicating that the sample contained almost only sodium chloride, which was of high purity. Furthermore, the diffraction peaks were sharp and of obvious intensity, indicating that the sodium chloride in the sample was crystalline and had good crystallinity.

[0055] This indicates that this method can be used to prepare composite industrial, agricultural, or edible potassium chloride and sodium chloride products from steel waste.

[0056] Example 2

[0057] This embodiment illustrates an example of adding a preliminary oxidation step in step (2) of Embodiment 1. Other settings in this embodiment are the same as in Embodiment 1, except that in step (2), before mixing the rinsing filtrate with the biosorbent, ozone gas is introduced into the rinsing filtrate for 9 minutes, with an ozone concentration greater than 10 mg / L. The thallium content in the de-accumulated filtrate obtained in step (3) was measured to be 9.32 mg / L, and the thallium content in the de-accumulated filter cake was 1031 mg / kg. This demonstrates that by treating with ozone gas before biosorption, the thallium removal rate can be improved. This is because ozone can pre-oxidize the thallium. + Effective oxidation to Tl 3+ Trivalent thallium ions are more likely to be adsorbed by biosorbents and enter the thallium-containing precipitate.

[0058] Example 3

[0059] This embodiment is used to illustrate the addition of sodium sulfide in the form of microcapsules in step (6) of embodiment 1. The other settings in this embodiment are the same as in embodiment 1. The difference is that before step (6), sodium sulfide capsules are prepared first, and then sodium sulfide is added in the form of sodium sulfide microcapsules (the actual amount of sodium sulfide powder added is still 0.24 g / L). The steps for preparing sodium sulfide microcapsules are as follows: first, sodium sulfide and water are prepared into an 8 mol / L solution, and then the sodium sulfide solution is mixed with chitosan-polyacrylic acid (CS-PAA) emulsion at a volume ratio of 1:4.5 and emulsified at a stirring speed of 9000 rpm for 12 min. Then, glutaraldehyde is slowly added dropwise at a stirring speed of 900 rpm at a volume of 0.5 ml per 100 ml. After the addition is completed, the mixture is stirred for 2 hours. Then, the obtained microcapsules are collected by centrifugation sedimentation, and then washed three times with anhydrous ethanol and vacuum dried to obtain microencapsulated sodium sulfide. Thallium content was determined by testing the sodium chloride and potassium chloride products. The thallium content in the potassium chloride product was 0.02 mg / kg, and the thallium content in the sodium chloride product was 0.01 mg / kg. It can be seen that the thallium content is lower than that in Example 1. This is because, under a high-pH alkaline environment, the microcapsule structure is dense and not easily ruptured, thus slowing the release rate and achieving a sustained-release effect. Therefore, the release is slower under alkaline conditions, avoiding the instantaneous release of sodium sulfide and reducing the risk of potential H2S escape, improving the selectivity of Tl2S precipitation, and reducing co-precipitated impurities. This enhances the thallium removal effect.

[0060] Comparative Example 1 This comparative example serves as a comparative test to illustrate the impact of omitting the coarse thallium removal step on subsequent processes and product quality. Except for omitting step S1 (coarse thallium removal), the remaining operational steps in this comparative example are the same as in Example 1. Testing of the filtrate and filter cake revealed that, without the coarse thallium removal step, the thallium content in the filtrate of this comparative example was 21.98 mg / L, significantly higher than the 11.68 mg / L in Example 1 after coarse thallium removal; simultaneously, the thallium content in the filter cake was 3998.1 mg / kg, also significantly higher than the 1247 mg / kg in Example 1.

[0061] The comparison between this comparative example and Example 1 further illustrates that if coarse thallium removal is not carried out first at the rinsing filtrate stage, the thallium load in the subsequent heavy removal process will increase significantly, which will not only exacerbate the migration of thallium in the process system, but also be detrimental to the stable implementation of subsequent deep and extreme thallium removal.

[0062] Comparative Example 2 This comparative example is used to illustrate the treatment effect comparison test without setting a deep thallium removal step. Except for omitting step (4) deep thallium removal, the other operation steps in this comparative example are the same as in Example 1, that is, the filtrate after heavy removal directly enters the calcium and magnesium removal process. The calcium and magnesium slag and filtrate obtained from this comparative example were tested, and the thallium content in the calcium and magnesium slag was 269 mg / kg and the thallium content in the calcium and magnesium filtrate was 20.61 mg / L; while after setting a deep thallium removal step in Example 1, the thallium content in the calcium and magnesium slag decreased to 150 mg / kg and the thallium content in the calcium and magnesium filtrate decreased to 7.65 mg / L.

[0063] The comparison between this comparative example and Example 1 demonstrates that incorporating a sodium sulfide deep dethallium removal process in the liquid phase after weight removal can significantly reduce the migration of thallium to the calcium and magnesium removal process, thereby improving the purity of the subsequent liquid phase.

[0064] Comparative Example 3 This comparative example is used to demonstrate the impact of omitting the thallium removal step on the final product. Except for omitting step (6) of thallium removal, the remaining operating steps in this comparative example are the same as in Example 1; that is, the calcium- and magnesium-free filtrate obtained in this comparative example directly enters the evaporation and crystallization process. The thallium content of the obtained sodium chloride and potassium chloride products was tested, and it was found that the thallium content in the obtained sodium chloride product was 241.8 mg / kg, and the thallium content in the potassium chloride product was 13 mg / kg. In contrast, after setting step S3 in Example 1, the thallium content in the sodium chloride product could be controlled at 0.02 mg / kg, and the thallium content in the potassium chloride product could be controlled at 0.03 mg / kg. This comparative result shows that if thallium removal is not performed on the liquid phase after calcium and magnesium removal before evaporation and crystallization, residual thallium will enter the final product during the salt separation process, leading to a significant decrease in product quality.

[0065] Comparative Example 4 This comparative example is used to demonstrate the comparative effect of using only sodium sulfide without sodium carbonate in the ultimate thallium removal stage. In this comparative example, step (6) is changed to only adding 0.24 g / L sodium sulfide to the calcium-magnesium filtrate, reacting for 10 min and aging for 60 min, while the other steps and parameters are the same as in Example 1. The thallium content of the final sodium chloride and potassium chloride products was tested, and it was found that the thallium content in the obtained sodium chloride product was 102.6 mg / kg, and the thallium content in the potassium chloride product was 8 mg / kg. By comparing the data with Comparative Example 3 and Example 1, it was found that although sodium sulfide was added as the main material for thallium removal, and sodium carbonate was only added to adjust the alkalinity of the system, the thallium content in the final product was only slightly lower than that in Comparative Example 3, but still much higher than that in Example 1. Such technical effect was not expected before the experiment. Analysis shows that sodium carbonate not only regulates the alkalinity of the system, but also works synergistically with sodium sulfide to promote the removal of residual thallium. Without the addition of sodium carbonate, residual calcium and magnesium ions will co-precipitate or encapsulate with thallium sulfide, thus hindering the removal of thallium by sodium sulfide. In other words, the sequential addition of sodium sulfide and sodium carbonate does not simply mean each plays its own independent role, but rather they work closely together. The combined use of sodium sulfide and sodium carbonate is more conducive to achieving the ultimate thallium removal before evaporation and crystallization.

[0066] In summary, the three-step thallium removal method for the potassium and sodium recovery stage in the steel solid waste treatment process provided by this invention achieves progressive reduction and segmented interception of thallium in the main potassium and sodium recovery process by sequentially implementing coarse thallium removal, deep thallium removal, and extreme thallium removal treatments at three key nodes: rinsing filtrate, heavy filtrate removal, and calcium and magnesium filtrate removal. This method can significantly reduce the migration and accumulation of thallium in subsequent processes and effectively control the thallium content in the final sodium chloride and potassium chloride products, demonstrating good process practicality and application value.

[0067] The above description is merely a preferred embodiment of the present invention and is not intended to limit the scope of protection of the present invention. Any modifications, equivalent substitutions, and improvements made within the spirit and principles of the present invention should be included within the scope of protection of the present invention.

Claims

1. A method for preparing low-thallium chloride products from steel solid waste through deep dethallium removal, characterized in that, Includes the following steps: I. Raw material processing: Sintering machine head ash, blast furnace bag ash, coke powder and water are mixed evenly and granulated. After enrichment by pyrometallurgy, zinc-rich dust and roasting residue are obtained. The zinc-rich dust is washed to obtain rinsing filter residue and rinsing filtrate with thallium content of 40~43mg / L. II. Crude thallium removal: The rinsing filtrate obtained in step I is mixed with the biosorbent at a liquid-solid ratio of (7~9) ml: 1 g. The adsorption reaction is carried out at 18~23℃ and pH 7.5~8.5 for 50~75 min. Then, solid-liquid separation is performed to obtain crude thallium removal filtrate and thallium-containing precipitate residue. III. Gravity removal treatment: Add a precipitant to the crude thallium-removed filtrate obtained in step II and adjust the pH to alkaline to perform gravity removal treatment, to obtain gravity-removed filtrate and gravity-removed filter cake; IV. Deep thallium removal: Add sodium sulfide to the filtrate obtained in step III at a solid-liquid ratio of (0.2~0.3) g: 1 L, stir at 22~28℃ for 8~12 min to carry out sulfidation reaction, then let stand and age for 25~35 min, and then perform solid-liquid separation to obtain deep thallium removal filtrate and thallium-containing precipitate residue; V. Calcium and magnesium removal treatment: Add calcium removal agent and magnesium removal agent to the deep thallium removal filtrate obtained in step IV to form calcium, magnesium and silicon impurities in the heavy thallium removal filtrate into precipitates. After solid-liquid separation, calcium and magnesium removal filtrate and calcium and magnesium removal filter residue are obtained. VI. Limiting thallium removal: Sodium sulfide and sodium carbonate are added sequentially to the calcium- and magnesium-free filtrate obtained in step V at solid-liquid ratios of (0.2~0.29) g: 1 L and (2.9~3.5) g: 1 L. After stirring at 22~28℃ for 8~12 min, the mixture is allowed to stand for 55~65 min, and then solid-liquid separation is performed to obtain the limiting thallium removal filtrate and the thallium-containing precipitate. VII. Preparation of potassium and sodium products: The thallium-free filtrate obtained in step VI is evaporated and crystallized to obtain sodium chloride and potassium chloride products, wherein the thallium content in the sodium chloride product is 0.005-0.02 mg / kg and the thallium content in the potassium chloride product is 0.005-0.03 mg / kg.

2. The method for preparing low-thallium chloride products from steel solid waste as raw material according to claim 1, characterized in that, The thallium content in the filtrate obtained in step III is 10~12 mg / L, and the thallium content in the filter cake is 1200~1300 mg / kg; The thallium content in the calcium-magnesium-removed filtrate obtained in step V is 7.02~8.38 mg / L, and the thallium content in the calcium-magnesium-removed filter residue is 140~160 mg / kg.

3. The method for preparing low-thallium chloride products from steel solid waste as raw material by deep dethallium removal according to claim 1 or 2, characterized in that, In step I, sintering machine head ash, blast furnace bag ash, coke powder and water are mixed evenly in a weight ratio of (5~10):(1~5):(3~6):(8~12); The pyrometallurgical enrichment process involves placing the granulated raw material mixture into a roasting furnace and roasting it at 1150~1250℃ for 0.8~1.2 hours to obtain roasting residue and flue gas. The flue gas is then separated from the dust by a dust removal device to obtain zinc-rich dust and dust-removed gas. The gas is then treated by a flue gas treatment device before being discharged. The rinsing process specifically involves mixing zinc-rich dust with deionized water or 1wt.% NaOH solution at a ratio of (130~160)g:(550~650)ml, stirring at 22~28℃ and 180~220rpm for 35~45min, and then performing vacuum filtration to obtain rinsing residue and rinsing filtrate.

4. The method for preparing low-thallium chloride products from steel solid waste as raw material according to claim 1 or 2, characterized in that, The specific process of weight removal in step III is as follows: a precipitant is added to the crude thallium-removed filtrate obtained in step II at a solid-liquid ratio of 0.3~3 g / L and the pH is adjusted to be alkaline. After stirring continuously for 20~30 min, the mixture is allowed to stand for 30~60 min to obtain a weight-removed filtrate and a weight-removed filter cake. The precipitant is one or a mixture of sodium hydroxide and sodium carbonate, with a pH of 9~11. Solid-liquid separation is then performed by pressure filtration.

5. The method for preparing low-thallium chloride products from steel solid waste as raw material by deep dethallium removal according to claim 1 or 2, characterized in that, The calcium and magnesium removal process in step V specifically involves adding a calcium removal agent to the deep thallium removal filtrate obtained in step IV at a solid-liquid ratio of 0.3~1.8 g / L, stirring at 20~30℃ for 10~15 min, then adding a magnesium removal agent at a solid-liquid ratio of 1~3 g / L, and continuing to stir at the same temperature for 10~15 min under conditions of pH 10~12. After solid-liquid separation by filtration, calcium and magnesium removed filtrate and calcium and magnesium removed filter residue are obtained. The calcium removal agent is one or a mixture of sodium carbonate or sodium oxalate, and the magnesium removal agent is sodium hydroxide or potassium hydroxide.

6. The method for preparing low-thallium chloride products from steel solid waste as raw material by deep dethallium removal according to claim 1 or 2, characterized in that, The preparation of potassium and sodium products in step VII is as follows: The thallium-free filtrate obtained in step VI is continuously evaporated and concentrated at 123~126℃. Sodium chloride first reaches saturation and crystallizes out. When the potassium chloride concentration reaches a weight ratio of (60~65):100 with water, it is considered that potassium chloride is close to the saturation solubility at this temperature. The mixture is sent to a centrifuge to separate the lower layer of NaCl crystals and the upper layer of saturated KCl solution. The NaCl crystals are dried at 100~120℃ for 30~60 min to obtain sodium chloride product. The upper layer of saturated KCl solution is introduced into a cooling tank and cooled from 123~126℃ to 10~25℃ at a rate of 1~2℃ / min. After KCl crystallization continues for 4~6 hours, centrifugation is performed to obtain KCl crystals and mother liquor. The mother liquor is returned for re-evaporation and concentration. The KCl crystals are dried at 80~100℃ for 30~45 min to obtain potassium chloride product.

7. The method for preparing low-thallium chloride products from steel solid waste as raw material by deep dethallium removal according to claim 1 or 2, characterized in that, The roasting residue obtained in step I is ground to below 0.074 mm and then separated into refined iron powder by magnetic separation. The tailings are then used as raw materials for building materials. The rinsed filter residue obtained in step I is dried and ground before being used as a raw material in the preparation process of nano zinc oxide. The thallium-containing precipitate obtained in steps II, IV, and VI is then sealed and solidified. The filter cake obtained in step III and the filter residue obtained in step V for removing calcium and magnesium are both returned to step I as recycled raw materials. Together with the granulated raw material mixture, they are used as raw materials for mixed roasting in the pyrometallurgical enrichment step.

8. The method for preparing low-thallium chloride products from steel solid waste as raw material by deep dethallium removal according to claim 1 or 2, characterized in that, The stirring in steps IV and VI is performed at a speed of 130-165 r / min. The solid-liquid separation described in steps II, IV and VI is all performed by plate and frame filter press.

9. The method for preparing low-thallium chloride products from steel solid waste as raw material by deep dethallium removal according to claim 1 or 2, characterized in that, In step II, before mixing the rinsing filtrate with the biosorbent, the following steps are performed: ozone gas is introduced into the rinsing filtrate for 8-10 minutes to ensure that the dissolved concentration of ozone is 10-20 mg / L.

10. The method for preparing low-thallium chloride products from steel solid waste as raw material by deep dethallium removal according to claim 1 or 2, characterized in that, In step VI, the sodium sulfide is microencapsulated sodium sulfide. The microencapsulated sodium sulfide is prepared by first preparing a 0.5~1.0 mol / L solution of sodium sulfide and water, then mixing the sodium sulfide solution with chitosan-polyacrylic acid emulsion at a volume ratio of 1:(4~5), and emulsifying at a stirring rate of 8000~10000 rpm for 10~15 min. Then, glutaraldehyde is slowly added dropwise at a stirring rate of 800~1000 rpm at a rate of 0.4~0.6 ml per 100 ml. After the addition is completed, stirring is continued for 1.8~2.2 hours. The obtained microcapsules are then collected by centrifugation sedimentation, washed 2~5 times with anhydrous ethanol, and then vacuum dried to obtain microencapsulated sodium sulfide.

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