A method for cascade recycling of key components of core laminated structure of waste photovoltaic modules

Abstract

The application discloses a kind of waste photovoltaic module core laminated structure key component's step recovery method, it is the granular powder material obtained by breaking, pyrolysis to waste photovoltaic module core laminated structure, sequentially carry out the selective leaching of composite organic acid solution aluminum and recover aluminum processing, the selective leaching of thiourea complexing oxidation solution silver-lead and recover silver processing, and the precipitation solidification processing of lead and the purification recovery processing of silicon, realize the step-by-step recovery of key components aluminum, silver, lead, silicon element.The recovery method of the application, according to the step separation order of "aluminum→silver→lead→silicon", realizes the selective separation and efficient recovery of corresponding metal and silicon resources to the composite waste material containing silicon matrix (containing antireflection layer SiN x , silver, aluminum, EVA and PbO.

Description

Technical Field

[0001] This invention relates to the field of resource recycling technology for waste photovoltaic modules, specifically a cascade recycling method for key components of the core laminated structure of waste photovoltaic modules. Background Technology

[0002] With the continued expansion of the global photovoltaic industry, the cumulative installed capacity of photovoltaics worldwide exceeded 1TW in 2023. The design life of photovoltaic modules is usually 25 to 30 years. The modules put into operation in the early stage are gradually entering the retirement cycle. A large-scale retirement wave will be ushered in in the next decade. Against this background, developing efficient and low-pollution photovoltaic module recycling technology is not only the core path to realize the recycling and utilization of key resources such as silicon, silver and aluminum, but also an urgent need to alleviate the environmental pollution risk of retired modules and support the sustainable development of the photovoltaic industry.

[0003] The laminated structure, as the core unit of a photovoltaic module, is composed of a glass substrate, aluminum frame, EVA film, silicon-based solar cells, backsheet, junction box, aluminum electrodes, solder ribbon, silver, and lead oxide composite. After the outer components (i.e., glass, aluminum frame, junction box, and backsheet) are removed, the laminated structure is often mechanically crushed into powder form, which is then sent to subsequent processing stages. Currently, the recycling of crushed battery cell powder faces several technical challenges: (i) During the crushing process, EVA glue, metal electrodes, and silicon battery cells are intertwined, causing the components to coexist closely, making it difficult to achieve high-purity separation through simple mechanical sorting; (ii) The powder contains various metals such as silver, aluminum, lead, and zinc in different chemical forms (elemental or oxide), resulting in complex phases. Conventional acid / alkali leaching solutions can easily lead to the co-dissolution of multiple metals, contaminating the leaching solution and increasing the burden of subsequent separation and purification, thus reducing the purity and recovery rate of valuable metals (especially silver); (iii) If lead in the form of PbO in the powder is not directionally solidified during leaching or post-treatment, it can easily enter the waste liquid or tailings, causing environmental risks and compliance issues; (iv) Residual EVA on the surface of the powder can affect the contact of the leaching agent and hinder metal dissolution. Therefore, there is an urgent need for a selective and environmentally friendly recycling method that can be designed specifically for the characteristics of crushed battery cells to achieve efficient separation and resource utilization of key components such as silver, aluminum, silicon, and lead. Summary of the Invention

[0004] This invention addresses the shortcomings of existing recycling processes for waste photovoltaic modules, such as multi-metal co-solution, difficult separation, and high pollution. It provides a tiered recycling method for key components of the core laminated structure of waste photovoltaic modules, following a tiered separation sequence of "aluminum → silver → lead → silicon," achieving the recycling of components including the silicon substrate (containing the anti-reflection layer SiN). x Selective separation and efficient recovery of corresponding metal or silicon resources from composite wastes containing SiO2, silver, aluminum, EVA, and PbO.

[0005] To achieve the above objectives, the present invention provides a tiered recycling method for key components of the core laminated structure of waste photovoltaic modules. The method involves crushing and pyrolyzing the core laminated structure of waste photovoltaic modules to obtain granular powder, which is then subjected to selective leaching of aluminum and recovery of aluminum with a composite organic acid solution, selective leaching of silver-lead and recovery of silver with a thiourea composite complex oxidation solution, precipitation and solidification of lead, and purification and recovery of silicon. This method achieves the tiered recycling of key components aluminum, silver, lead, and silicon. The composite organic acid solution is selected from one of the following: oxalic acid-phosphoric acid composite solution, oxalic acid-tartaric acid composite solution, oxalic acid-maleic acid composite solution, and oxalic acid-succinic acid composite solution. The pH of the composite organic acid solution is 1.0~2.0. The thiourea complex oxidation solution is selected from one of the following: thiourea-ferric iron complex solution, thiourea-persulfate complex solution, thiourea-sulfate complex solution, thiourea-hydrogen peroxide complex solution, and thiourea-potassium permanganate complex solution. The pH of the thiourea complex oxidation solution is 2.0~5.0.

[0006] This invention is based on a recovery route of "layer-by-layer deconstruction—selective complexation—directional dissolution—graded solidification". First, aluminum is selectively leached out through the complexation effect of a composite organic acid solution for aluminum recovery. Then, silver and lead are selectively leached out through the complexation and oxidation effects of a thiourea composite complex oxidation solution for silver recovery. After that, lead is directionally solidified through precipitation. Finally, silicon is purified and recovered, achieving efficient step-by-step separation and recovery of key components aluminum, silver, lead and silicon.

[0007] In the selective leaching of aluminum, taking an oxalic acid-phosphoric acid composite solution as an example, oxalic acid provides coordination and complexation capabilities, while phosphoric acid provides buffering and dissolution-promoting effects to achieve the directional and selective dissolution of aluminum. The specific reaction mechanism is as follows:

[0008] The aforementioned complexation reaction allows aluminum to enter the solution as a stable aluminum oxalate complex, while silver and silicon remain undissolved. Under these acidic conditions, some PbO will react and dissolve Pb. 2+ The presence of phosphoric acid causes the dissolved Pb 2+ It tends to form a loosely structured, easily separable, insoluble lead phosphate precipitate, avoiding passivation problems caused by lead oxalate formation that would affect aluminum leaching. In the subsequent silicon purification and recovery process, based on the density difference between lead and silicon, the insoluble lead phosphate is effectively separated without affecting the high-purity silicon powder product obtained from silicon recovery. At the same time, lead solidification is also achieved, thus effectively solving the problem of subsequent separation complexity caused by multi-metal co-dissolution in traditional nitric acid solutions.

[0009] In the selective leaching process of silver-lead, taking a thiourea-ferric composite solution as an example, thiourea, as a ligand for silver ions, can form a stable thiourea-silver complex under acidic conditions, thereby achieving the directional dissolution of metallic silver. This composite solution can dissolve metallic silver (Ag) without damaging the silicon matrix structure. 0 It is oxidized and dissolved, forming a soluble complex.

[0010] The complexation-oxidation dissolution reaction of silver is as follows:

[0011] Simultaneously, this composite solution can also dissolve and complex lead oxide (PbO) into the solution, but it will not dissolve the sparingly soluble lead phosphate precipitate. The reaction is as follows:

[0012] The subsequent coordination effect is as follows:

[0013] Where x depends on the concentration of thiourea in the solution and the coordination equilibrium of the solution.

[0014] Thus, both silver and lead enter the leaching solution as soluble complex ions, forming a complex solution containing silver and lead, which is then separated based on the differences in their properties; while the silicon substrate (including the antireflective layer) remains in a solid phase and does not dissolve, and is subsequently purified and recovered by separating it from the sparingly soluble lead salt.

[0015] Therefore, the recycling method of the present invention has the ability to physically and chemically deconstruct complex mixed solid phases, precisely control the chemical selectivity between metals, gently protect the silicon matrix, and directional solidification and disposal of harmful metals such as lead. It not only achieves efficient and highly selective separation and full-scale recycling of key components, but also is green and environmentally friendly throughout the process and is easy to scale up industrially.

[0016] As a limitation of the above technical solution, after the waste photovoltaic modules are dismantled to obtain the core laminated structure, they are crushed into powder, and the EVA adhesive is carbonized by pyrolysis in an inert atmosphere to obtain granular powder.

[0017] As a limitation of the above technical solution, the oxalic acid mass concentration in the composite organic acid solution is 3~5 wt%, and the mass concentration of the other component acid is 1~3 wt%, that is, the mass concentrations of phosphoric acid, tartaric acid, maleic acid and succinic acid are all 1~3 wt%.

[0018] As a limitation on the above technical solution, the ferric iron is selected from one of FeCl3, Fe2(SO4)3, and Fe(NO3)3; and / or, the persulfate is selected from sodium persulfate or ammonium persulfate; and / or, the sulfate is selected from copper sulfate or ferric sulfate. Preferably, the concentration of thiourea in the thiourea complex oxidizing solution is 0.05~0.20 mol / L, and the concentration of the other complex component is 0.01~0.10 mol / L, that is, the concentrations of ferric iron, persulfate, sulfate, hydrogen peroxide, and potassium permanganate are all 0.01~0.10 mol / L.

[0019] The conditions for obtaining pyrolysis granular powder and the composition of composite organic acid solution and thiourea composite complex oxidation solution were improved. The composite organic acid solution was used to improve the high selective complexation of aluminum, and the thiourea composite complex oxidation solution was used to achieve the directional dissolution of silver and control the synergistic complexation of lead, thereby optimizing the cascade recovery effect.

[0020] As a limitation of the above technical solution, the granular powder is first impregnated in a composite organic acid solution to carry out a selective leaching reaction of aluminum, resulting in a solid mixture containing silver, silicon, lead oxide and sparingly soluble lead salt, and a filtrate containing aluminum. NaOH solution is added to the filtrate to convert aluminum ions into sodium aluminate, and then CO2 gas is continuously introduced to precipitate Al(OH)3. Preferably, the solid-liquid ratio of the granular powder to the composite organic acid solution is 1g:10~20ml, and / or the leaching temperature of the selective leaching reaction of aluminum is 40~70 ℃, and the stirring leaching time is 30~120min.

[0021] The selectively leached aluminum ions are precipitated as aluminum hydroxide. The precipitation reaction is as follows:

[0022] The purity of the obtained aluminum hydroxide precipitate can reach ≥94.8%. This treatment can realize continuous feeding, reaction and separation of materials, forming an efficient resource recovery process.

[0023] As a limitation of the above technical solution, the solid obtained after solid-liquid separation following selective leaching of aluminum is impregnated in a thiourea complex complex oxide solution for selective leaching of silver-lead reaction. After the reaction is completed, solid-liquid separation is performed, and the filtrate is treated by electrodeposition to obtain metallic silver. Preferably, the amount of thiourea complex complex oxidation solution added is 1g:2~10ml in a solid-liquid ratio, and / or the leaching temperature of the selective leaching silver-lead reaction process is 40~70℃, and the stirring leaching time is 20~120min.

[0024] The filtrate obtained from the selective leaching of silver-lead reaction is a solution containing silver-lead complex ions. This is because there is a significant difference in the electrode reduction potentials of silver and lead (Ag... + / Ag = +0.80 V, Pb 2+Since Pb = -0.13 V, silver can be preferentially deposited by constant current electrodeposition, while lead remains in the solution, thus achieving the separation of silver and lead. The purity of the deposited silver product can reach ≥99.8 wt%.

[0025]

[0026] As a limitation of the above technical solution, a precipitant is added to the electrodeposition residue to complete the solidification of lead; preferably, a precipitant is added to make the pH of the reaction solution 6.5~7.5, and lead is precipitated at a temperature of 20~60 ℃.

[0027] As a limitation of the above technical solution, the precipitant is selected from one of the following: carbonate precipitant, sulfate precipitant, phosphate precipitant, and silicate precipitant.

[0028] Lead solidification employs precipitation methods, such as carbonate precipitation (forming PbCO3), sulfate precipitation (forming PbSO4), phosphate precipitation (forming Pb3(PO4)2), and silicate solidification (forming PbSiO3), to directionally precipitate lead into lead salts that are highly stable and not easily dissolved again, thus achieving the harmless solidification treatment of lead.

[0029]

[0030] This method avoids the problem of secondary H2S pollutants that may be generated by existing sulfide precipitation methods, making it more environmentally friendly. In addition, compared with nitric acid co-leaching solution, it does not cause the problem of multiple metals such as silver and lead being co-dissolved and difficult to separate. It realizes a segmented synergistic recovery mode of "silver first purification → lead later solidification", which has high separation efficiency, strong selectivity, simple post-processing and low environmental risk.

[0031] As a limitation of the above technical solution, the solid obtained after solid-liquid separation following selective leaching of silver-lead reaction is subjected to ball milling, gravity separation, and hot water washing to complete the purification and recovery of silicon; preferably, the ball milling conditions are a ball-to-material ratio of 5:1 to 15:1 and a rotation speed of 200 to 400 rpm; and / or, the hot water temperature is 60 to 95 ℃, and the washing is performed 2 to 4 times.

[0032] By using ball milling, particles are mechanically broken down to increase the specific surface area, release any trace impurities that may be trapped inside, and improve the surface properties of the particles. Then, by taking advantage of the density difference between silicon and lead, insoluble lead salts are separated by gravity separation. Finally, the high temperature and fluidity of hot water are used to remove soluble ionic impurities and some residual organic matter adsorbed on the surface of silicon particles, thus achieving deep purification and recovery of high-purity silicon products.

[0033] As a limitation of the above technical solution, the cascade recycling method for key components of the core laminated structure of waste photovoltaic modules includes the following operational steps: a. Pyrolysis and sorting of core laminated structure powder: After removing the junction box, glass, welding strip, back sheet, welding strip and aluminum frame from the waste photovoltaic modules, the core laminate is crushed into 0.5~2 mm powder; the powder is taken and pyrolyzed at 500±50℃ for 5~30 min under an inert atmosphere to carbonize the EVA adhesive. After cooling, it is sorted and sieved to obtain granular powder; b. Selective leaching and recovery of aluminum: The granular powder is impregnated in a composite organic acid solution and leached with stirring at 50±5℃ for 80±20 min. After the leaching reaction is completed, solid-liquid separation is performed, and the solid is a mixture containing silver, silicon, lead oxide, and insoluble lead salts. The filtrate is an aluminum-containing solution. 1±0.5 mol / L NaOH solution is added to the filtrate to convert aluminum ions into sodium aluminate. Subsequently, CO2 gas is continuously introduced to precipitate Al(OH)3. Preferably, the purity of Al(OH)3 is not less than 94.8%. c. Selective leaching of silver-lead and recovery of silver: The solid obtained from solid-liquid separation in step b is impregnated in a thiourea complex oxidizing solution and leached with stirring at 50±5 ℃ for 60±10 min. After the leaching reaction is completed, solid-liquid separation is performed, and the filtrate is subjected to electrodeposition treatment at a constant current of -10±2 mA / cm for 2±0.5 h to obtain metallic silver; preferably, the purity of silver is not less than 99.8%, and the recovery rate is not less than 92.4%; d. Directional solidification of lead: Add a precipitant to the electrodeposition residue until the pH of the reaction solution is 7±0.5, and then react at 40±5 ℃ for 30±5 min to precipitate lead; preferably, the lead residue is not higher than 0.1 ppm; e. Silicon recovery and purification: The solid obtained from solid-liquid separation in step c is ball-milled for 30±5 min, then water is added to make a slurry. The slurry is gravity-separated using a hydrocyclone. Silicon powder is enriched in the overflow liquid, and sparingly soluble lead salts are enriched in the bottom liquid. The silicon powder overflow liquid is collected and washed 2-4 times with hot water at 80±5 ℃, stirring for 20±5 min each time. After drying, silicon powder is obtained. Preferably, the purity of silicon powder is not less than 98.2%, and the recovery rate is not less than 95.9%.

[0034] Further improvements and optimizations to the overall recycling method of this invention will enable the operation and conditions of each processing step to fully exert synergistic effects, achieving efficient and green separation and purification of key components aluminum, silver, lead and silicon, as well as resource utilization.

[0035] In summary, the cascade recycling method for key components of the core laminated structure of waste photovoltaic modules of the present invention has the following advantages.

[0036] 1. It has strong metal selectivity and can achieve efficient separation through "stepwise deconstruction and directional dissolution".

[0037] This invention employs a composite organic acid solution for selective leaching of aluminum, followed by a thiourea composite complex oxide solution for directional dissolution of silver and lead. This avoids the subsequent separation difficulties caused by the simultaneous dissolution of multiple elements such as Ag, Al, and Pb in traditional nitric acid co-leaching processes, achieving a precise, step-by-step separation: "aluminum removed first → silver preferentially dissolved → lead solidified later → silicon retained." This innovative sequence ensures that each step of recovery involves a single phase, reducing separation complexity and improving overall recovery efficiency.

[0038] 2. The recycling process is environmentally friendly and can form a highly efficient, streamlined process for resource recycling.

[0039] This invention utilizes a composite organic acid solution to selectively leach aluminum, followed by the addition of sodium hydroxide solution to convert aluminum ions into sodium aluminate. Carbon dioxide is then introduced for carbonation decomposition, resulting in the precipitation of aluminum hydroxide, thus achieving targeted aluminum resource recovery. The leached solids (mainly containing silicon, silver, lead oxide, and insoluble lead salts) are subsequently subjected to cascaded recovery and purification. The thiourea composite complex oxidation solution used for the selective leaching of silver-lead can be regenerated into Fe through aeration or electrochemical methods. 3+ The solution is stable and recyclable; lead is solidified through precipitation into a highly stable, insoluble lead salt that is not easily dissolved again. Compared with the sulfide precipitation method, this avoids the risk of H2S emission pollution, significantly improving environmental friendliness and safety. This enables continuous feeding, reaction, and separation of materials to be recycled, forming a highly efficient resource recycling process.

[0040] 3. It has a high recovery rate of multiple valuable elements, high product purity, and combines resource utilization benefits.

[0041] The cascade recycling process of this invention can achieve: ① Aluminum leaching rate ≥ 93.4%, and the purity of the obtained Al(OH)3 ≥ 94.8%; ② Silver electrodeposition recovery rate ≥ 92.4%, product purity ≥ 99.8 wt%; ③ The residual rate of lead after directional curing is ≤0.1 ppm, and the insoluble lead salt precipitate is stable and leak-free; ④ The silicon purity is ≥98.2% and the recovery rate is ≥95.9%, which can be reused as metallurgical grade silicon resources.

[0042] The overall recycling method can transform low-value waste components into high-value-added products, resulting in significant economic benefits.

[0043] 4. Avoid using highly hazardous chemicals such as HF; the process operation is highly safe.

[0044] This invention does not use fluorine-containing etchants in the silicon purification process. Instead, it adopts a physical-weak chemical synergistic purification method of mechanical ball milling + gravity separation + hot water washing, which effectively removes surface adsorbed impurities. It eliminates the need for high-risk reagents such as HF and NH4F, significantly reducing operational risks and equipment corrosion resistance requirements. The process is safer and the equipment cost is lower.

[0045] 5. The process can be scaled up industrially, has a high degree of continuity, and low energy consumption and cost.

[0046] The present invention relates to mild reaction conditions, such as leaching at 40~70 °C and operation under normal pressure conditions. The reagents used are common and renewable, and there is no difficult-to-treat waste gas or waste liquid. The leaching-separation-recovery process can be scaled up industrially using a continuous reaction vessel, electrodeposition tank and conventional solid-liquid separation equipment. The process is simple and suitable for large-scale processing of retired photovoltaic modules.

[0047] 6. The recycling methods have a wide range of applications.

[0048] The recycling method of this invention is applicable to composite waste containing polymetallic oxides and silicon substrates, such as decommissioned crystalline silicon photovoltaic cells and their crushed powder (typically containing Al, Ag, PbO, and Si), solar electrode or conductive paste waste powder (mainly containing Ag and PbO), and semiconductor processing silicon wafer scraps or cutting waste (mainly containing Si and SiO2 / SiN). x The waste materials include films and conductive glass fragments with metal oxide coatings (mainly containing metal oxides and silicon substrates). The target components contained in these waste materials can be selectively separated and recycled according to their actual composition through the step-by-step processing of this invention. Attached Figure Description

[0049] Figure 1 This is a schematic diagram of the recycling process of key components in the core laminated structure of photovoltaic modules according to the present invention.

[0050] Figure 2 The graph shows the changes in silicon and silver content on the surface of the granular powder before (a) and after (b) the leaching reaction in step c of Example 1.

[0051] Figure 3 The image shows the SEM image and elemental distribution map of the silicon powder recovered in Example 1.

[0052] Figure 4 The images shown are SEM and EDS images of the silver recovered in Example 1.

[0053] Figure 5 The bar chart shows the comparison of element recovery rate and purity between the various embodiments and comparative examples.

[0054] Figure 6 The residual Pb concentrations for each embodiment and comparative example are given. Detailed Implementation

[0055] The technical solution of the present invention will be clearly and completely described below with reference to the embodiments. Obviously, the described embodiments are only some embodiments of the present invention, and not all embodiments. Based on the embodiments of the present invention, all other embodiments obtained by those of ordinary skill in the art without creative effort are within the scope of protection of the present invention.

[0056] Unless otherwise specified, the experimental methods described below are conventional methods; the raw materials or test materials used are typical products purchased from the market, unless otherwise specified. In the quantitative experiments of the following examples and comparative examples, three replicate experiments were conducted, and the results were averaged.

[0057] To accurately evaluate the performance of the process described in this invention and ensure the comparability and repeatability of the data, all key indicators involved in the examples and comparative examples were determined using the following national standards, industry standards, or recognized chemical analysis standard methods.

[0058] Aluminum leaching rate: EDTA titration method was used. An appropriate amount of leachate was accurately measured and digested with nitric acid-perchloric acid to completely destroy the organic complexing agent, thus converting all aluminum to Al. 3+ In an acetate-ammonium acetate buffer system with pH ≈ 4.5, using PAN as an indicator, aluminum was complexed with excess EDTA standard solution, and then the excess EDTA was back-titrated with copper sulfate standard solution. The aluminum content was calculated based on the amount of EDTA consumed.

[0059] Aluminum leaching rate (%) = (mass of aluminum in the leachate / theoretical total mass of aluminum in the initial material) × 100%.

[0060] Silicon purity and recovery rate: A gravimetric method combined with inductively coupled plasma atomic emission spectrometry (ICP-OES) was used. A silicon powder sample was accurately weighed, dissolved in a mixture of hydrofluoric acid and nitric acid to volatilize the silicon as SiF4. The remaining residue was ignited at 950°C to constant weight, and the insoluble content was calculated. Alternatively, the sample was alkali-fused and then acidified, and the content of key impurity elements was determined using ICP-OES. Silicon purity was calculated by subtraction.

[0061] Silicon purity (%) = 100% - Σ impurity content.

[0062] Silicon recovery rate (%) = (final recovered silicon powder mass / theoretical total mass of silicon in the initial material) × 100%.

[0063] Silver purity: The silver product obtained by electrodeposition is dissolved in nitric acid using the fire assay method (standard method) and then the silver content and impurity elements such as gold, copper, and lead are determined by ICP-OES to calculate the purity.

[0064] Silver recovery rate: The silver concentration in the leaching solution and electrodeposition residue was determined by atomic absorption spectrometry (AAS) or ICP-OES. Silver recovery rate (%) = (mass of metallic silver obtained by electrodeposition / theoretical total mass of silver in the initial material) × 100%.

[0065] Lead residual concentration: Graphite furnace atomic absorption spectrometry (GFAAS) or inductively coupled plasma mass spectrometry (ICP-MS) was used. The clarified residual liquid after lead curing was taken, stabilized with an appropriate amount of acid, and precisely diluted before direct injection for determination. This method is suitable for the highly sensitive and accurate detection of trace lead (< 0.1 ppm).

[0066] Aluminum hydroxide purity: determined by EDTA titration (same as "Aluminum Leaching Rate"). After dissolving the Al(OH)3 sample obtained from precipitation, the aluminum content was determined according to the above method and converted into aluminum hydroxide purity.

[0067] The key components of the core laminated structure of waste photovoltaic modules are recycled in stages. A schematic diagram of the recycling process is shown below. Figure 1 As shown.

[0068] The waste photovoltaic modules used in the examples and comparative examples were from the same batch. Their composition was determined, and the results were as follows: glass substrate 70 wt%, aluminum frame 18 wt%, EVA film 5.1 wt%, silicon-based solar cell 3.65 wt%, backsheet 1.5 wt%, junction box 1 wt%, aluminum electrode 0.53 wt%, solder ribbon 0.167 wt%, silver 0.050 wt%, and lead oxide 0.003 wt%.

[0069] The mass fraction of macroscopic components such as glass substrate, aluminum frame, back plate, junction box, and solder strip is obtained by manually disassembling and weighing them, and then dividing the weight by the total mass.

[0070] The mass fraction of EVA film was determined by Soxhlet extraction combined with thermogravimetric analysis (TGA). EVA was extracted with an organic solvent (such as xylene), the solvent was evaporated and the film was weighed, and then its decomposition weight loss was measured by TGA under an inert atmosphere.

[0071] Silicon content can be determined by gravimetric method, which involves alkali melting, acidification, and dehydration followed by calcination at 950℃ to constant weight, and the silicon content is calculated based on the mass of silicon dioxide; or by alkali melting-inductively coupled plasma atomic emission spectrometry (ICP-OES).

[0072] The content of silver, aluminum, and lead oxide can be determined by inductively coupled plasma atomic emission spectrometry (ICP-OES) or graphite furnace atomic absorption spectrometry (GFAAS).

[0073] Example 1 A cascade recycling process for key components of the core laminated structure of waste photovoltaic modules is carried out according to the following steps.

[0074] a. Pyrolysis and sorting of core laminated structure powder: After removing the junction box, glass, welding strip, backsheet, welding strip, and aluminum frame from the waste photovoltaic modules, the core laminate is crushed into 0.5~2 mm powder. 800 g of powder was taken, and the proportions of each element in the powder were determined and calculated to be 38.552% silicon, 0.559% silver, 0.031% lead oxide, and 5.607% aluminum, with EVA accounting for 55.251%. The powder was pyrolyzed at 500 ℃ for 10 min under an inert argon atmosphere to carbonize the EVA adhesive. After cooling, it was sorted by airflow and sieved through a 200 μm vibrating screen to obtain approximately 358 g of particulate powder. The particulate powder was determined to contain 1.25% silver, 86.15% silicon, 12.53% aluminum, and 0.07% lead oxide.

[0075] b. Selective leaching and recovery of aluminum: 100 g of granular powder was impregnated in 1.5 L of oxalic acid-phosphoric acid composite solution (the composite solution contained 3 wt% oxalic acid and 2 wt% phosphoric acid, pH 1.0, and all solutions were in deionized water). The leaching was carried out by stirring at 50 ℃ for 90 min. After the leaching reaction was completed, the solid and liquid were separated. The solid was a mixture containing silver, silicon, lead oxide and insoluble lead phosphate. The filtrate was an aluminum-containing solution. The leaching rate of aluminum in the filtrate was determined to be 95.8%. 1 mol / L NaOH solution was added to the filtrate, and then CO2 gas was continuously introduced. 36.1 g of Al(OH)3 precipitate was obtained, and the purity of Al(OH)3 was determined to be 96.1%.

[0076] c. Selective leaching of silver-lead and recovery of silver: The solid obtained from solid-liquid separation in step b was completely impregnated in 300 ml of a thiourea-ferric trivalent solution (thiourea 0.1 mol / L, Fe2(SO4)3 0.05 mol / L, pH 4.0 in the composite solution). The solution was stirred and leached at 50 ℃ for 60 min. After the leaching reaction was completed, the solid and liquid were separated, and the solid obtained were silica slag and insoluble lead phosphate. The filtrate was a solution containing silver-lead complex ions. The filtrate was subjected to electrodeposition treatment at a constant current of -10 mA / cm for 2 h to obtain 1.188 g of silver powder with a purity of 99.92% and a recovery rate of 95.0%.

[0077] d. Directional curing of lead: The electrodeposition residue from step c, as determined by ICP-OES, contained lead ions (Pb). 2+The initial concentration was approximately 187.1 ppm. Subsequently, Na₂CO₃ solution (NaHCO₃ solution or (NH₄)₂CO₃ solution could also be used) was added to the electrodeposition residue to adjust the pH to 7.0, and the reaction was carried out at 40 °C for 30 min to obtain PbCO₃ precipitate. After the reaction, solid-liquid separation was performed, and the supernatant was again analyzed by ICP-OES to determine the Pb concentration in the electrodeposition residue. 2+ Residue <0.08 ppm.

[0078] e. Silicon recovery and purification: The solid obtained from the solid-liquid separation in step c was ball-milled (ball-to-material ratio 10:1, rotation speed 300 rpm) for 30 min, and then water was added to mix and form a slurry. The slurry was then gravity-separated using a hydrocyclone. Utilizing the density difference between silicon powder and lead salt, silicon powder was enriched in the overflow liquid, while insoluble lead phosphate was enriched in the bottom liquid, thus achieving efficient separation. The silicon powder overflow liquid was collected and washed three times with 80 ℃ hot water for 20 min each time. After drying, 84.5 g of silicon powder was obtained, with a purity of 98.6% and a recovery rate of 96.7%.

[0079] The changes in silicon and silver content on the granular powder before and after the leaching reaction in step c are as follows: Figure 2 As shown, Figure 2 Image (a) is a SEM image of the surface of the pre-processed powder particles, showing the local morphology and elemental distribution of the silicon solar cell. Figure 2 (b) is a SEM image of the surface of the processed powder particles, showing the surface morphology and element distribution of the recovered high-purity silicon powder (lead was not selected when taking the picture due to its low content, so it is not shown in the figure; it mainly reflects the removal of silver). This indicates that the recovered silicon has a very high purity.

[0080] Figure 3 The SEM image of the surface morphology of the silicon powder recovered in step e and the corresponding elemental distribution map are shown in the figure. As can be seen from the figure, the obtained silicon powder particles have complete morphology and clean surface, and no obvious adhering substances or heterogeneous phase precipitation are observed. The corresponding elemental distribution map shows that the silicon element signal is uniformly distributed and there are no other impurities, indicating that the silicon phase has high purity and the residue of impurity elements in the silicon matrix is ​​effectively controlled.

[0081] Figure 4 The images show the SEM and EDS spectra of the silver recovered in step c. The EDS spectrum shows that silver is the main constituent element, while other elements show weak signal responses, indicating extremely high phase purity of silver. (Comparison) Figure 2 The morphology and elemental distribution before and after the leaching reaction show that silver and silicon elements are mixed together before step c recovery. Figure 2In the middle (a), the recovered silver element is concentrated in the silver recovery product area, while the silicon element is concentrated in the silicon powder area, indicating that the process has successfully achieved efficient separation of silver and silicon and the recovered product has high purity.

[0082] Example 2 A tiered recycling process for key components of the core laminated structure of waste photovoltaic modules is carried out according to the following steps: a. Pyrolysis and sorting of core laminated structure powder: After removing the junction box, glass, welding strip, backsheet, welding strip and aluminum frame from the waste photovoltaic modules, the core laminate is crushed into 0.5~2 mm powder; 800 g of powder is taken, and the proportion of each element in the powder is determined by measurement and calculation as follows: silicon 38.467%, silver 0.558%, lead oxide 0.022%, aluminum 5.578%, and EVA accounts for 55.375% of the powder; under an inert argon atmosphere, the powder is pyrolyzed at 550 ℃ for 20 min to carbonize the EVA adhesive. After cooling, it is sorted by airflow and sieved by a 200 μm vibrating screen to obtain about 357 g of granular powder. The granular powder is determined to contain 1.25% silver, 86.20% silicon, 12.50% aluminum and 0.05% lead oxide.

[0083] b. Selective leaching and recovery of aluminum: 100 g of granular powder was impregnated in 1.7 L of an oxalic acid-tartaric acid composite solution (containing 3 wt% oxalic acid, 2.5 wt% tartaric acid, pH 1.0). After stirring at 45 ℃ for 100 min, solid-liquid separation was performed. The solid obtained was a mixture containing silver, silicon, lead oxide, and insoluble lead tartrate. The filtrate was an aluminum-containing solution. The aluminum leaching rate of the filtrate was determined to be 93.4%. Adding 1 mol / L NaOH solution and then continuously bubbling CO2 gas resulted in the precipitation of 35.6 g of Al(OH)3, with a purity of 94.8%. c. Selective Leaching of Silver-Lead and Silver Recovery: The solid obtained from the solid-liquid separation in step b was completely impregnated in 300 ml of a thiourea-ferric composite solution (thiourea 0.1 mol / L, Fe2(SO4)3 0.08 mol / L, pH 3.0). The leaching reaction was completed by stirring at 45 ℃ for 70 min. After solid-liquid separation, the solids were silica slag and insoluble lead tartrate. The filtrate was a silver-lead complex ion solution. The filtrate was subjected to electrodeposition at a constant current of -8 mA / cm for 2.5 h, yielding 1.174 g of silver powder with a purity of 99.85% and a recovery rate of 93.8%. d. Directional curing of lead: The electrodeposition residue from step c, as determined by ICP-OES, contained lead ions (Pb). 2+The initial concentration was approximately 135.1 ppm. Subsequently, Na₂SO₄ solution was added to the electrodeposition residue to bring the pH to 6.5, and the reaction was carried out at 35 °C for 35 min to obtain PbSO₄ precipitate. After the reaction, solid-liquid separation was performed, and the supernatant was again analyzed by ICP-OES to determine the Pb content. 2+ Residue <0.1 ppm.

[0084] e. Silicon recovery and purification: The solid obtained from the solid-liquid separation in step c was ball-milled (ball-to-material ratio 15:1, speed 400 rpm) for 35 min, and then water was added to mix and form a slurry. The slurry was then gravity-separated using a hydrocyclone. Silicon powder was enriched in the overflow liquid, and insoluble lead tartrate was enriched in the bottom liquid, achieving efficient separation. The silicon powder overflow liquid was collected and washed three times with hot water at 75 ℃ for 25 min each time. After drying, 84.2 g of silicon powder was obtained, with a purity of 98.2% and a recovery rate of 95.9%.

[0085] Example 3 A tiered recycling process for key components of the core laminated structure of waste photovoltaic modules is carried out according to the following steps: a. Pyrolysis and sorting of core laminated structure powder: After removing the junction box, glass, welding strip, backsheet, welding strip and aluminum frame from the waste photovoltaic modules, the core laminate is crushed into 0.5~2 mm powder. 800 g of powder is taken, and the proportion of each element in the powder is determined by measurement and calculation as follows: silicon 38.350%, silver 0.547%, lead oxide 0.031%, aluminum 5.571%, and EVA accounts for 55.501% of the powder. Under an inert argon atmosphere, it is pyrolyzed at 450 ℃ for 30 min to carbonize the EVA adhesive. After cooling, it is sorted by airflow and sieved by a 200 μm vibrating screen to obtain about 356 g of particulate powder. The particulate powder is determined to contain 1.23% silver, 86.18% silicon, 12.52% aluminum and 0.07% lead oxide.

[0086] b. Selective leaching and recovery of aluminum: 100 g of granular powder was impregnated in 2.0 L of oxalic acid-maleic acid composite solution (containing 5 wt% oxalic acid, 1.5 wt% maleic acid, and pH 1.5). After stirring at 55 ℃ for 70 min, solid-liquid separation was performed. The solid was a mixture containing silver, silicon, lead oxide, and insoluble lead maleate. The filtrate was an aluminum-containing solution. The aluminum leaching rate of the filtrate was determined to be 95.1%. 1 mol / L NaOH solution was added, followed by continuous bubbling of CO2 gas, which precipitated 35.9 g of Al(OH)3. The purity of Al(OH)3 was determined to be 95.9%.

[0087] c. Selective Leaching of Silver-Lead and Silver Recovery: The solid obtained from the solid-liquid separation in step b was completely impregnated in 300 ml of a thiourea-sodium persulfate composite solution (thiourea 0.2 mol / L, Na₂S₂O₈ 0.03 mol / L, pH 3.0). After stirring at 55 ℃ for 50 min, solid-liquid separation was performed, yielding a silica slag as the solid and a silver-lead complex ion solution as the filtrate. The filtrate was then subjected to electrodeposition at a constant current of -12 mA / cm for 1.5 h, yielding 1.140 g of silver powder with a purity of 99.80% and a recovery rate of 92.5%. d. Directional curing of lead: The electrodeposition residue from step c, as determined by ICP-OES, contained lead ions (Pb). 2+ The initial concentration was approximately 189.2 ppm. Subsequently, NH4H2PO4 solution was added to the electrodeposition residue to bring the pH to 7.5, and the reaction was carried out at 45 °C for 25 min to obtain Pb3(PO4)2 precipitate. After the reaction was complete, solid-liquid separation was performed, and the supernatant was again analyzed by ICP-OES to determine the Pb content. 2+ Residue <0.1 ppm.

[0088] e. Silicon recovery and purification: The solid silicon slag obtained from the solid-liquid separation in step c was ball-milled (ball-to-material ratio 5:1, rotation speed 300 rpm) for 25 min, and then water was added to mix and form a slurry. The slurry was then gravity-separated using a hydrocyclone. Silicon powder was enriched in the overflow liquid, and insoluble lead maleate was enriched in the bottom liquid, achieving efficient separation. The silicon powder overflow was collected and washed three times with hot water at 85 ℃ for 15 min each time. After drying, 84.3 g of silicon powder was obtained, with a purity of 98.4% and a recovery rate of 96.3%.

[0089] In each embodiment, the mother liquor after Al(OH)3 precipitation in step b can be treated uniformly by adding dilute acid (such as sulfuric acid or hydrochloric acid) to adjust the pH to 6-9, and adding a small amount of flocculant (such as PAC, polyaluminum chloride) to promote the sedimentation of residual trace suspended solids. The resulting supernatant, after testing, meets the requirements of relevant regulations such as the "Integrated Wastewater Discharge Standard" (GB 8978) and can be safely discharged. The mother liquor after separating the P lead salt precipitate in step d can be evaporated and concentrated to recover the salt byproducts (such as sodium carbonate, sodium sulfate, and ammonium dihydrogen phosphate), and the remaining condensate is discharged in compliance with standards.

[0090] Comparative Example Different processing methods were used to recycle the core laminated structure of photovoltaic modules, which were compared with the recycling method of the present invention.

[0091] Comparative Example 1 The core laminated structure of photovoltaic modules is recycled using the traditional nitric acid co-leaching method, and the specific operation is as follows.

[0092] (1) The core laminate structure of the waste photovoltaic module was crushed, pyrolyzed and sorted according to the same operation as step a in Example 1. From 800 g of powder with a particle size of 0.5~2 mm (the proportion of each element was measured and calculated to be 37.884% silicon, 0.532% silver, 0.035% lead oxide, 5.548% aluminum, and 56.001% EVA in the powder), about 352 g of particulate powder (containing 1.21% silver, 86.10% silicon, 12.61% aluminum and 0.08% lead oxide) was obtained.

[0093] (2) 100 g of granular powder was placed in 1.5 L of 3 mol / L HNO3 solution with pH 1.0. After stirring at 50 ℃ for 90 min, the solid and liquid were separated. The solid was silicon slag, and the filtrate was a solution containing aluminum, silver and lead. The aluminum leaching rate in the filtrate was determined to be 96.2%. 1 mol / L NaOH solution was added to the filtrate, and then CO2 was continuously passed through. 444.3 g of Al(OH)3 was precipitated, and the purity was determined to be 79.2%.

[0094] (3) Electrodeposition was performed on the filtrate after Al(OH)3 precipitation. Electrodeposition was carried out at a constant current of -10 mA / cm for 2 hours to obtain 1.040 g of silver powder. The purity was determined to be 96.84% and the recovery rate was 83.2%.

[0095] (4) The electrodeposition residue was subjected to directional solidification of lead according to the same procedure as step d in Example 1. ICP-OES analysis revealed that the lead ion concentration (Pb) in the residue was... 2+ The initial concentration was approximately 130.5 ppm. After precipitating PbCO3 from the electrodeposition residue, solid-liquid separation was performed after the reaction was complete. The supernatant was then analyzed by ICP-OES to determine the Pb concentration. 2+ The residue was as high as 15 ppm.

[0096] (5) The solid silicon slag obtained after solid-liquid separation in step (2) was ball-milled (ball-to-material ratio 10:1, rotation speed 300 rpm) for 30 min, then washed with hot water at 80 ℃ for 20 min each time, and dried to obtain 82.9 g of silicon powder. The purity was determined to be 88.6% and the recovery rate was 85.3%.

[0097] Nitric acid is a strong inorganic acid with high oxidizing properties. It cannot selectively dissolve specific metals, but will cause multiple metals such as aluminum, silver, lead, and zinc to dissolve simultaneously, thus corroding the antireflective coating (SiN) on the surface of the silicon wafer. x(SiO2) and even part of the silicon matrix, therefore, after aluminum and silver are precipitated sequentially from nitric acid solution, multiple ions such as lead and zinc still exist, which will significantly interfere with the subsequent Na2CO3 precipitation step. Therefore, Pb 2+ After precipitation with PbCO3, some Pb was found under a high nitrate concentration background. 2+ It will exist in a complexed or colloidal state, causing Pb 2+ The residue was as high as 15 ppm. This result indicates that non-selective leaching solutions cannot effectively separate aluminum and lead, causing a large number of impurity elements and reaction products to adhere to the surface of silicon particles. Simple hot water washing cannot effectively remove them, and the loss of the silicon matrix itself leads to a decrease in recovery rate.

[0098] Comparative Example 2 The core laminated structure of photovoltaic modules is recycled using a single organic acid, and the specific operation is as follows.

[0099] (1) The core laminate structure of the waste photovoltaic module was crushed, pyrolyzed and sorted according to the same operation as step a in Example 1 to obtain granular powder. The granular powder contained 1.22% silver, 86.13% silicon, 12.59% aluminum and 0.06% lead oxide.

[0100] (2) 100 g of granular powder was placed in 1.5 L of oxalic acid solution with a pH of 1.5 and stirred at 50 °C for 90 min. After solid-liquid separation, the solid was a mixture containing silicon, silver, unreacted aluminum, residual PbO, and newly formed lead oxalate precipitate. The filtrate was an aluminum-containing solution. The leaching rate of aluminum in the filtrate was determined to be 70.1%. 1 mol / L NaOH solution was added to the filtrate, and CO2 was continuously introduced. 28.2 g of Al(OH)3 precipitated, and the purity of Al(OH)3 was determined to be 90.5%.

[0101] (3) The solid obtained from the solid-liquid separation in step (2) was completely immersed in 300 ml of thiourea-ferric solution (thiourea 0.1 mol / L, Fe2(SO4)3 0.05 mol / L, pH 4.0). After leaching at 50 ℃ for 60 min, the solid and liquid were separated. The solid obtained was a mixture of silicon, incompletely reacted aluminum, residual PbO and lead oxalate precipitate. The filtrate contained silver-lead and Al produced by aluminum dissolution. 3+ The complex ion solution was used. The filtrate was subjected to electrodeposition at a constant current of -10 mA / cm for 2 h to obtain 1.075 g of silver powder with a purity of 97.30% and a recovery rate of 85.7%.

[0102] (4) The residual solution after electrodeposition in step (3) was analyzed by ICP-OES, and the lead ion (Pb) content was found to be [missing information]. 2+The initial concentration was approximately 166.3 ppm. Subsequently, Na₂CO₃ solution was added to the electrodeposition residue until the pH reached 7.0, and the reaction was carried out at 40 °C for 30 min to obtain PbCO₃ precipitate. After the reaction, solid-liquid separation was performed, and the supernatant was again analyzed by ICP-OES to determine the Pb concentration. 2+ The residue was 5 ppm.

[0103] (5) The solid obtained from the solid-liquid separation in step (3) is ball-milled (ball-to-material ratio 10:1, rotation speed 300 rpm) for 30 min, and then water is added to mix and make a slurry. The slurry is gravity-separated using a hydrocyclone, and the silicon powder is enriched in the overflow liquid. The silicon powder overflow liquid is collected and washed three times with hot water at 80 ℃ for 20 min each time. After drying, 84.7 g of silicon powder is obtained, the purity is determined to be 92.5%, and the recovery rate is 91.0%.

[0104] Comparative Example 2, using a single oxalic acid solution, showed significantly lower results than the embodiments of this invention. On one hand, oxalic acid has limited leaching capacity for aluminum; on the other hand, some PbO reacts to form lead oxalate precipitate, which creates a dense and firmly adhered passivation layer on the aluminum particle surface, hindering further aluminum dissolution and resulting in a lower aluminum leaching rate (70.1%). This demonstrates the inherent defect of the single oxalic acid system in generating a harmful passivation layer. In contrast, the composite organic acid solution of this invention utilizes another component, such as phosphate, to guide lead to form a loose and easily separable lead phosphate precipitate, thereby avoiding the passivation problem and achieving efficient aluminum leaching and high-purity silicon recovery. This highlights the design advantage of this invention, which uses composite organic acids to control the precipitate morphology for efficient separation.

[0105] Comparative Example 3 Sodium cyanide solution is the most important solution for extracting gold and silver in hydrometallurgical processes, and it is widely used due to its high leaching efficiency and mature technology. This comparative example uses a sodium cyanide-hydrogen peroxide system to treat pyrolysis-sorted battery cell particles; the specific operation is as follows.

[0106] (1) Pyrolysis and sorting of core laminated structure powder: After removing the junction box, glass, welding strip, back sheet, welding strip and aluminum frame of the waste photovoltaic module, the core laminate is crushed into 0.5~2 mm powder; 800 g of powder (after measurement and calculation, the proportion of each element in the powder is 37.829% silicon, 0.527% silver, 0.031% lead oxide, 5.489% aluminum, and EVA accounts for 56.124% of the powder) is pyrolyzed at 500 ℃ for 30 min in an inert argon atmosphere to carbonize the EVA adhesive. After cooling, it is sorted by airflow and sieved by a 200 μm vibrating screen to obtain about 351 g of particulate powder (after measurement, it contains 1.20% silver, 86.22% silicon, 12.51% aluminum and 0.07% lead oxide).

[0107] (2) Comprehensive leaching and preliminary separation of sodium cyanide: 100g of the granular powder obtained in step (1) was immersed in 1.5L sodium cyanide solution (NaCN concentration 0.3 wt%, pH 11.0 maintained with NaOH), and 0.5 wt% hydrogen peroxide was added as a leaching aid. The mixture was stirred and leached for 8h at a temperature of 25℃. After the reaction was completed, the solid and liquid were separated. The solid was mainly silicon and PbO, and the filtrate was a cyanide complex solution containing aluminum and silver.

[0108] In this alkaline sodium cyanide system, aluminum can react with cyanide ions (CN ions). - ) and hydroxide ions (OH) - A coordination reaction occurs, forming soluble aluminum cyanide complexes (such as NaAl(CN)4) and aluminates (such as NaAlO2). After leaching with sodium cyanide solution, the elemental silver in the powder is converted into soluble sodium silver cyanide complexes.

[0109] (3) Selective precipitation and recovery of aluminum: CO2 gas was slowly introduced into the filtrate obtained in step (2), and 38.3 g of Al(OH)3 precipitate was generated. The purity of Al(OH)3 was determined to be 88.5%. After filtration and separation, aluminum hydroxide solid was obtained. After washing and drying, it can be used as aluminum recovery product. The aluminum recovery rate was calculated to be 93.7%.

[0110] CO2, as a mild acidifier, reacts with alkaline substances in the system and lowers the pH of the solution. When the pH drops to 9.0, aluminates and aluminum cyanide complexes will hydrolyze, preferentially forming aluminum hydroxide (Al(OH)3) precipitate. The remaining filtrate after aluminum precipitation is rich in silver cyanide complexes and can be sent to subsequent processes for silver recovery.

[0111] (4) Zinc powder replacement to recover silver: Add excess zinc powder to the filtrate obtained after recovering aluminum in step (3) and carry out a replacement reaction. After reacting for 2 hours, filter to obtain silver-zinc precipitate. Wash the precipitate with dilute sulfuric acid to dissolve excess zinc powder and obtain 1.004 g of silver powder. The silver recovery rate was determined to be 79.2% and the purity was 94.62%.

[0112] Obviously, this recovery rate and purity are significantly lower than those of the example. The main reason is that zinc powder replacement introduces zinc impurities, and other metal ions coexisting in the cyanide solution may also be replaced.

[0113] (5) Residual liquid treatment: The residual liquid after cyanide leaching and replacement is highly toxic cyanide-containing wastewater and must be specially treated. The residual liquid is treated by alkaline chlorination. First, add chlorine or hypochlorite to pH>11 to oxidize cyanide to cyanate. Then adjust the pH to 8 and further oxidize it to carbon dioxide and nitrogen.

[0114] (6) Separation of lead and purification of silicon: The solid obtained in step (2) (mainly containing silicon and lead oxide) was immersed in 300 ml of 2 mol / L nitric acid solution and stirred at 50 °C for 60 min. The solid and liquid were separated, and the filter residue was silicon powder and the filtrate was lead nitrate solution. The lead ion concentration (Pb) was determined by ICP-OES. 2+ The initial concentration was approximately 180.9 ppm. Subsequently, Na₂CO₃ solution was slowly added to the filtrate until the pH reached 7.5, and the mixture was stirred for 30 min to allow lead to precipitate as PbCO₃. After the reaction was complete, solid-liquid separation was performed, and the supernatant was again analyzed by ICP-OES to determine the Pb concentration in the filtrate. 2+ The residue was <0.1 ppm. The obtained silicon powder was washed with deionized water and dried to obtain 78.3 g of silicon powder. The purity of the silicon powder was determined to be 90.5%, and the recovery rate was 82.2%. The results of this comparative example, which used a traditional sodium cyanide-hydrogen peroxide system to treat photovoltaic module powder, showed significantly lower silver recovery efficiency (79.2%, purity 94.62%) and silicon product purity (90.5%) compared to the previous example. Furthermore, the process generated highly toxic cyanide-containing wastewater, requiring complex alkaline chlorination for subsequent treatment. Additionally, complex steps were needed to separate lead oxide and silicon, increasing recovery costs. These results highlight the inherent disadvantages of the sodium cyanide system, including insufficient selective separation capability, significant environmental hazards, and high subsequent treatment costs. This underscores the significant technical advantages of the cascade separation method of this invention in achieving efficient, green, and highly selective recovery of key components.

[0115] Comparative Example 4 The core laminated structure of photovoltaic modules is recycled by first dissolving and recovering silver, and then dissolving and recovering lead. The specific operation is as follows.

[0116] (1) The core laminate structure of the waste photovoltaic module was crushed, pyrolyzed and sorted according to the same operation as step a in Example 1. 800 g of powder with a particle size of 0.5~2 mm (the proportion of each element in the powder was measured and calculated to be 38.234% silicon, 0.546% silver, 0.027% lead oxide, 5.569% aluminum, and 55.624% EVA) to obtain about 355 g of particulate powder (the proportion of each element in the powder was measured to be 1.23% silver, 86.16% silicon, 12.55% aluminum and 0.06% lead oxide).

[0117] (2) 100 g of granular powder was soaked in 300 ml of thiourea-ferric solution (thiourea 0.1 mol / L, Fe2(SO4)3 0.05 mol / L, pH 4.0 in thiourea-ferric solution), and after leaching at 50 ℃ for 60 min, solid-liquid separation was performed. The solid was a mixture containing silicon, residual silver, aluminum and lead compounds, and the filtrate was a complex ion solution containing silver and aluminum. The silver leaching rate was determined to be 78.6%.

[0118] The crushed particles contain a large amount of metallic aluminum. When immersed in a thiourea-ferric solution, aluminum, being a reactive metal, will preferentially react with Fe. 3+ A redox reaction occurs (2Al + 3Fe). 3+ → 2Al 3+ + 3Fe 2+ This results in a large consumption of the oxidant Fe in the solution. 3 + This results in insufficient concentration of the oxidant used to oxidize elemental silver; simultaneously, the Al produced during dissolution... 3+ It may also compete with thiourea for coordination. Therefore, when materials containing a large amount of metallic aluminum directly enter the thiourea-ferric leaching system, the preferential oxidation and competitive coordination of aluminum will severely inhibit the effective leaching of silver. This not only slows down the silver leaching rate and significantly reduces the leaching percentage, but also causes the oxidant (Fe) to... 3+ The ineffective consumption of Al and complexing agents (thiourea) increases reagent costs. More importantly, a large amount of Al... 3+ Entering the solution will deteriorate the subsequent electrodeposition environment and affect the purity of the silver product.

[0119] (3) The filtrate obtained from the solid-liquid separation in step (2) was subjected to electrodeposition treatment. Electrodeposition was carried out at a constant current of -10 mA / cm for 2 h to obtain 0.926 g of silver powder. The purity was determined to be 97.80% and the recovery rate was 73.6%.

[0120] In addition to silver, the filtrate also contains a large amount of Al produced by the dissolution of aluminum. 3+ Although aluminum has a very negative standard electrode potential (approximately -1.66 V) and will not be electrochemically reduced at the potential for silver deposition, the local pH at the cathode surface increases due to the hydrogen evolution reaction. When this pH reaches the precipitation solubility product of Al(OH)3, Al... 3+ It will generate amorphous aluminum hydroxide through chemical precipitation, which will be trapped or adsorbed in the silver deposition layer, thus contaminating the silver products.

[0121] (4) The solid obtained from the solid-liquid separation in step (2) was completely immersed in an oxalic acid-phosphoric acid composite solution (containing 3 wt% oxalic acid, 2 wt% phosphoric acid, and pH 1.0) at a solid-liquid ratio of 1 g: 15 mL. Aluminum was leached at a temperature of 50 °C. After stirring for 90 min, the solid and liquid were separated. The solid obtained was a mixture containing silicon, residual silver, some aluminum, and lead compounds. The filtrate was a solution containing aluminum and lead. The aluminum leaching rate of the filtrate was determined to be 90.5%. 1 mol / L NaOH solution was added to the filtrate, and CO2 was continuously introduced. 35.7 g of Al(OH)3 was precipitated. The purity of Al(OH)3 was 92.0%.

[0122] Since the silver leaching reaction in step (2) is carried out under acidic conditions, some aluminum will be dissolved irregularly and the solid surface properties will change, resulting in the aluminum leaching reaction in step (4) being incomplete and the rate being uneven. The purity of the obtained Al(OH)3 product is reduced due to the inclusion of residues.

[0123] (5) Combine the electrodeposition residue after silver recovery in step (3) and the filtrate residue after aluminum recovery in step (4), and determine the lead ion (Pb) content by ICP-OES. 2+ The initial concentration was approximately 205.4 ppm. Subsequently, Na₂CO₃ solution was added to pH 7.0 to solidify the lead. After the reaction, solid-liquid separation was performed, and the supernatant was again analyzed by ICP-OES. However, due to the chaotic metal dissolution pathway in the preceding steps, the lead dispersion resulted in a lower concentration of Pb in the final residual solution. 2+ The concentration fluctuates greatly, approximately 3 ppm, resulting in unstable curing effects.

[0124] (6) Silicon recovery and purification: The final silicon slag after leaching in steps (2) and (4) was ball-milled (ball-to-material ratio 10:1, rotation speed 300 rpm) for 30 min, and then water was added to make a slurry. The slurry was separated by gravity using a hydrocyclone. Silicon powder was enriched in the overflow liquid, while insoluble lead salts were enriched in the bottom liquid, thus achieving separation. The silicon powder overflow was collected and washed three times with 80 ℃ hot water for 20 min each time. After drying, 83.9 g of silicon powder was obtained. The purity of silicon powder was determined to be 94.0%, and the recovery rate was 91.5%.

[0125] Due to the two acidic leaching processes, conducted in the order of "strong acid first, then weak acid," the silicon substrate was subjected to a more adverse effect compared to the previous example. Firstly, the strong acid thiourea-Fe... 3+In the solution, some silicon substrates are slightly corroded, and the surface becomes rough and porous. Subsequently, in the oxalic acid-phosphoric acid solution in step (4), although the acidity is weak, the acid will penetrate into the damaged silicon substrate surface, further dissolving impurities and expanding the cracks. This "first severe damage, then penetration" effect results in more impurity ions being encapsulated or adsorbed inside and on the surface of the silicon particles. In addition, a large amount of impurities such as aluminum and lead, and their reaction products (such as hydroxides) that were not completely removed in the silver dissolution step, remain. Therefore, the silicon substrate surface is subjected to repeated chemical impacts, and the subsequent impurity removal step results in heavy surface contamination. Even after the same ball milling and hot water washing, these impurities that have penetrated deep into the silicon substrate surface or formed strong bonds are difficult to completely remove, resulting in a significant reduction in silicon powder purity. At the same time, during the washing process, some silicon particles that have become smaller or more porous due to chemical erosion are more easily lost, causing a decrease in recovery rate. The results of this comparative example demonstrate the importance of the "aluminum first, silver later" sequence of the present invention for protecting the integrity of the silicon substrate. The tiered recycling sequence of the present invention is based on the coupling results of metal activity differences, coordination stability, redox potential and solid phase film structure stability. If the step sequence is changed, problems such as aluminum and silver co-dissolving, difficulty in post-processing lead, and silicon loss due to damage to the silicon surface film will occur.

[0126] The recycling results of each embodiment and the comparative example were compared, and the results are shown in Table 1 below. Figure 5 , Figure 6 .

[0127] Table 1. Comparison of Recycling Results between Examples and Comparative Examples

[0128] As shown in Table 1, the recovery results of the embodiments demonstrate comprehensive and significant advantages in all key performance indicators, with overall performance far superior to the comparative schemes. Specifically, in the aluminum leaching stage, the leaching rates of Examples 1-3 were consistently above 93.4%, reflecting the high efficiency and high selectivity of the composite organic acid solution. In contrast, the leaching rate of the single oxalic acid solution in the comparatives was low due to insufficient complexing ability, and although the nitric acid solution had a high leaching rate, its indiscriminate solubility due to its strong acidity placed a heavy burden on subsequent separation. In the most valuable silver recovery, the advantages of the embodiments are extremely prominent. Their silver recovery rate (92.5%~95.0%) and product purity (99.80%~99.92%) significantly outperformed all the comparative examples. This directly verifies the core advantage of the "selective leaching of aluminum first" stepwise separation strategy, creating a clean reaction environment for the subsequent silver extraction step, thereby ensuring the efficient and high-purity recovery of silver. Conversely, Comparative Example 4, which involved dissolving silver before aluminum, provided strong counter-evidence, resulting in the lowest silver recovery rate (73.6%). This indicates that changing the order of the recovery process leads to severe interference from aluminum impurities in the silver leaching and electrodeposition processes. Furthermore, the examples also demonstrated superior quality and environmental friendliness in the final product, achieving the highest purity (≥98.2%) and recovery rate (≥95.9%) of the obtained silicon powder, with lead residue concentration effectively controlled below 0.1 ppm, achieving stable solidification of harmful elements. Moreover, the use of highly toxic sodium cyanide was avoided throughout the process, and no NO was generated. x Or toxic gases such as H2S.

[0129] Figure 5 The bar chart shows the comparison of elemental recovery rate and purity between the various embodiments and comparative examples. Figure 6 Pb for each embodiment and comparative example 2+ The residual concentration graphs show that Example 1 is far ahead in terms of recycling efficiency and product quality, as well as the significant advantages of each example in lead detoxification, demonstrating economic efficiency and environmental friendliness suitable for industrial applications.

[0130] In summary, the recovery method of this invention achieves mild and directional separation of difficult-to-separate components through mechanism matching, exhibits strong metal selectivity, and enables highly efficient separation through "stepwise deconstruction-directional dissolution." It also boasts high recovery rates of multiple valuable elements, high product purity, and resource utilization benefits. Furthermore, it is environmentally friendly, avoids the use of hazardous chemicals such as HF, and ensures high process safety. The process flow can be industrially scaled up, has a high degree of continuity, low energy consumption and cost, and a wide range of applications. It demonstrates strong process selectivity, good series connection, and excellent overall recovery performance.

[0131] Please note that the technical features of the above embodiments can be combined arbitrarily. For the sake of brevity, not all possible combinations of the technical features in the above embodiments have been described. However, as long as the combination of these technical features does not contradict each other, it should be considered within the scope of this specification. The above embodiments only illustrate several implementation methods of this application, and their descriptions are relatively specific and detailed, but they should not be construed as limiting the scope of the invention patent. It should be pointed out that for those skilled in the art, several modifications and improvements can be made without departing from the concept of this application, and these all fall within the protection scope of this application. Therefore, the protection scope of this patent application should be determined by the appended claims.

Claims

1. A method for the cascaded recycling of key components of the core laminated structure of waste photovoltaic modules, characterized in that: The granular powder obtained from the core laminated structure of waste photovoltaic modules through crushing and pyrolysis is then subjected to selective leaching of aluminum and recovery of aluminum with a composite organic acid solution, selective leaching of silver-lead and recovery of silver with a thiourea composite complex oxidation solution, precipitation and solidification of lead and purification and recovery of silicon, so as to achieve the step-by-step recovery of key components aluminum, silver, lead and silicon. The composite organic acid solution is selected from one of the following: oxalic acid-phosphoric acid composite solution, oxalic acid-tartaric acid composite solution, oxalic acid-maleic acid composite solution, and oxalic acid-succinic acid composite solution. The pH of the composite organic acid solution is 1.0~2.

0. The thiourea complex oxidation solution is selected from one of the following: thiourea-ferric iron complex solution, thiourea-persulfate complex solution, thiourea-sulfate complex solution, thiourea-hydrogen peroxide complex solution, and thiourea-potassium permanganate complex solution. The pH of the thiourea complex oxidation solution is 2.0~5.

0.

2. The cascade recycling method for key components of the core laminated structure of waste photovoltaic modules according to claim 1, characterized in that: After the waste photovoltaic modules are dismantled to obtain the core laminated structure, they are crushed into powder and then carbonized in an inert atmosphere to obtain granular powder.

3. The cascade recycling method for key components of the core laminated structure of waste photovoltaic modules according to claim 1, characterized in that: The oxalic acid concentration in the complex organic acid solution is 3-5 wt%, and the concentration of the other component acid is 1-3 wt%.

4. The cascade recycling method for key components of the core laminated structure of waste photovoltaic modules according to claim 1, characterized in that: The ferric iron is selected from one of FeCl3, Fe2(SO4)3, and Fe(NO3)3; and / or, the persulfate is selected from sodium persulfate or ammonium persulfate; and / or, the sulfate is selected from copper sulfate or ferric sulfate.

5. The cascade recycling method for key components of the core laminated structure of waste photovoltaic modules according to claim 1, characterized in that: The concentration of thiourea in the thiourea complex oxidation solution is 0.05~0.20 mol / L, and the concentration of the other complex component is 0.01~0.10 mol / L.

6. The cascade recycling method for key components of the core laminated structure of waste photovoltaic modules according to claim 1, characterized in that: The granular powder is first impregnated in a composite organic acid solution to selectively leach aluminum, resulting in a solid mixture containing silver, silicon, lead oxide and sparingly soluble lead salts. The filtrate is an aluminum-containing solution. NaOH solution is added to the filtrate to convert aluminum ions into sodium aluminate. Then, CO2 gas is continuously introduced to precipitate Al(OH)3.

7. The cascade recycling method for key components of the core laminated structure of waste photovoltaic modules according to claim 6, characterized in that: The solid-liquid ratio of the granular powder to the composite organic acid solution is 1g: 10~20ml, and / or the leaching temperature of the selective leaching aluminum reaction process is 40~70℃, and the stirring leaching time is 30~120min.

8. The cascade recycling method for key components of the core laminated structure of waste photovoltaic modules according to claim 6, characterized in that: The solid obtained after selective leaching of aluminum and solid-liquid separation is impregnated in a thiourea complex complex oxide solution for selective leaching of silver-lead. After the reaction is completed, solid-liquid separation is performed, and the filtrate is treated by electrodeposition to obtain metallic silver.

9. The cascade recycling method for key components of the core laminated structure of waste photovoltaic modules according to claim 8, characterized in that: The amount of thiourea complex complex oxidation solution added is 1g : 2~10ml in a solid-liquid ratio, and / or the leaching temperature for the selective leaching silver-lead reaction process is 40~70℃, and the stirring leaching time is 20~120min.

10. The method for cascade recycling of key components of the core laminated structure of waste photovoltaic modules according to claim 8, characterized in that: Add a precipitant to the electrodeposition residue to complete the lead curing.

11. The cascade recycling method for key components of the core laminated structure of waste photovoltaic modules according to claim 10, characterized in that: A precipitant is added to adjust the pH of the reaction solution to 6.5-7.5, and lead is precipitated at a temperature of 20-60 °C.

12. The cascade recycling method for key components of the core laminated structure of waste photovoltaic modules according to claim 10, characterized in that: The precipitant is selected from one of the following: carbonate precipitant, sulfate precipitant, phosphate precipitant, and silicate precipitant.

13. The cascade recycling method for key components of the core laminated structure of waste photovoltaic modules according to claim 8, characterized in that: The solid obtained after selective leaching of silver-lead reaction and solid-liquid separation is subjected to ball milling, gravity separation and hot water washing to complete the purification and recovery of silicon.

14. The cascade recycling method for key components of the core laminated structure of waste photovoltaic modules according to claim 13, characterized in that: The ball milling conditions are a ball-to-material ratio of 5:1 to 15:1, a rotation speed of 200 to 400 rpm, and / or a hot water temperature of 60 to 95 ℃, and washing 2 to 4 times.

15. The method for cascade recycling of key components of the core laminated structure of waste photovoltaic modules according to any one of claims 1 to 14, characterized in that, The following steps are included: a. Pyrolysis and sorting of core laminated structure powder: After removing the junction box, glass, welding strip, back sheet and aluminum frame of the waste photovoltaic module, the core laminate is crushed into 0.5~2 mm powder; the powder is taken and pyrolyzed at 500 ±50 ℃ for 5~30 min in an inert atmosphere to carbonize the EVA adhesive. After cooling, it is sorted and sieved to obtain granular powder. b. Selective leaching and recovery of aluminum: The granular powder is immersed in a composite organic acid solution and leached with stirring at a temperature of 50±5 ℃ for 80±20 min. After the leaching reaction is completed, solid-liquid separation is performed. The solid is a mixture containing silver, silicon, lead oxide and sparingly soluble lead salts, and the filtrate is an aluminum-containing solution. 1±0.5 mol / L NaOH solution is added to the filtrate to convert aluminum ions into sodium aluminate. Subsequently, CO2 gas is continuously introduced to precipitate Al(OH)3. c. Selective Leaching of Silver-Lead and Recovery of Silver: The solid obtained from solid-liquid separation in step b is impregnated in a thiourea complex oxidizing solution and leached with stirring at 50±5 ℃ for 60±10 min. After the leaching reaction is completed, solid-liquid separation is performed, and the filtrate is subjected to electrodeposition treatment at a constant current of -10±2 mA / cm². 2 Electrodeposition for 2±0.5 h yielded metallic silver; d. Directional solidification of lead: Add precipitant to the electrodeposition residue until the pH of the reaction solution is 7±0.5, and then react at 40±5 ℃ for 30±5 min to precipitate lead; e. Silicon recovery and purification: The solid obtained from solid-liquid separation in step c is ball-milled for 30±5 min, and then water is added to make a slurry. The slurry is gravity-separated using a hydrocyclone. Silicon powder is enriched in the overflow liquid, and sparingly soluble lead salts are enriched in the bottom liquid. The silicon powder overflow liquid is collected and washed with hot water at 80±5 ℃ 2~4 times, stirring for 20±5 min each time. After drying, silicon powder is obtained.

16. The method for cascade recycling of key components of the core laminated structure of waste photovoltaic modules according to claim 15, characterized in that: In step b, the purity of Al(OH)3 is not less than 94.8%.

17. The cascade recycling method for key components of the core laminated structure of waste photovoltaic modules according to claim 15, characterized in that: In step c, the purity of silver is not less than 99.8%, and the recovery rate is not less than 92.4%.

18. The method for cascade recycling of key components of the core laminated structure of waste photovoltaic modules according to claim 15, characterized in that: In step d, the lead residue is not higher than 0.1 ppm.

19. The cascade recycling method for key components of the core laminated structure of waste photovoltaic modules according to claim 15, characterized in that: In step e, the purity of the silicon powder is not less than 98.2%, and the recovery rate is not less than 95.9%.

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