Method for recovering gold from electronic components having a composite plating layer

Abstract

The application discloses a gold recovery method for electronic components with composite plating layers. The method comprises the following steps: heat treating the electronic components; selectively leaching the gold-containing solder by using an HCl-H2O2 system; detecting the residual amount of non-gold components in the leaching residue, and feeding back the control of re-leaching according to the detection result until the residual amount of non-gold components reaches a preset threshold; using I2-KI aqueous solution to perform secondary selective gold leaching on the leaching residue reaching the threshold, and recovering the gold; at the same time, using an acidic electrolyte containing fluorine compounds to perform pulse anodic electrochemical treatment on the substrate with a Ti / Au composite metallization layer, and then separating the Au film from the substrate surface through ultrasonic treatment and recovering the Au film. The application sets selective leaching and interface peeling paths for gold in different occurrence states, reduces the interference of non-gold components on the gold leaching process, avoids the large dissolution of Au and SiC substrates, and realizes the efficient and selective recovery of gold in electronic components.

Description

Technical Field

[0001] This invention relates to the field of metal recycling technology, and in particular to a method for gold recycling from electronic components with composite coatings. Background Technology

[0002] Existing industrial gold recovery methods, including cyanidation and aqua regia leaching, typically rely on the premise that the composition of the material to be treated is relatively homogeneous, or that gold mainly exists in the interface system formed by gold and a single base metal. These methods achieve the dissolution and recovery of the precious metal through bulk leaching. However, in multilayer circuit stack structures, gold is dispersed in multiple regions, and the gold in different regions forms different adjacent metal systems with solder, Ti, SiC, and other chip metal layers. Therefore, if a traditional bulk leaching process is used, gold and its adjacent metals in each region will participate in the leaching simultaneously. This results in non-gold components in the solder, as well as various metal components such as Ti, Ni, Ge, Si, and In, entering the leaching system concurrently. This not only increases the difficulty of subsequent gold separation and purification but also, different non-gold components may interfere with gold leaching and subsequent recovery by consuming oxidants or complexing agents in the leaching system, altering the redox environment and acid-base balance, and forming insoluble precipitates, hydrolysis products, or colloids. This leads to increased leaching agent consumption, reduced gold leaching efficiency, and increased complexity in subsequent separation and purification processes. Therefore, for the aforementioned composite stacked structures with multiple metals, interfaces, and solder types, it is difficult to achieve effective separation of different metal layers and efficient gold recovery using an integral leaching process designed for homogeneous gold-containing waste. Existing selective gold stripping techniques all rely on the electrochemical reactivity of the Ni layer in acidic systems. By selectively dissolving the Ni adhesion layer, the upper gold film loses its adhesion and is peeled off, thus achieving non-dissolving gold recovery. However, the adhesion layer in the above structure is Ti, not Ni. In common acidic systems such as nitric acid and aqua regia, Ti rapidly forms a dense TiO2 passivation film on its surface, causing the aforementioned nickel etching system to completely fail. On the other hand, while existing solder reflow separation processes can separate gold from solder, they are suitable for mixtures of bulk gold and solder. For thin Ti / Au metallization layers tightly adhered to rigid SiC substrates, reflow methods cannot effectively peel the thin gold layer from the substrate surface, lacking applicability.

[0003] Therefore, there is an urgent need to establish a layered processing and selective gold recovery method for this type of multi-layered stacked structure. Summary of the Invention

[0004] To address the shortcomings of existing technologies, this invention proposes a method for gold recovery from electronic components with composite coatings.

[0005] This invention provides a method for gold recovery from electronic components with a composite plating layer, wherein the surface of the electronic component includes a gold solder layer and a Ti / Au composite metallization layer, and the method includes: S1: The electronic component is subjected to heat treatment to melt the gold solder layer, and the chip is separated from the substrate to collect the gold solder and the substrate 1 with the Ti / Au metallization layer; S2: The gold-containing solder is placed in an HCl-H2O2 system, and the first leaching residue is obtained by solid-liquid separation; S3: Place the first leaching residue in an I2-KI aqueous solution and react in the dark for 1.5~3 h, then collect the gold-containing filtrate; S4: The substrate 1 is placed in an electrolyte containing an acidic medium and a fluorine-containing compound. The substrate 1 is used as the working electrode. Alternating anodic pulses and de-energizing pulses are applied to selectively dissolve Ti and form an interfacial cleavage along the Ti / Au interface. S5: The substrate 1 after pulse electrochemical treatment is subjected to ultrasonic treatment to detach the Au film from the surface of the substrate 1, and the Au film is collected.

[0006] In some embodiments, the gold-containing solder in step S1 is one or more of AuSn, AuGe, AuSi, AuIn, and SnAgCu; the gold in the solder forms an alloy or intermetallic compound with Sn, Ge, Si, or In. By limiting the solder type to the aforementioned gold-containing solder, the first leaching solution can be selected according to the chemical properties of the corresponding non-gold components, allowing the non-gold components to preferentially dissolve, complex, undergo redox transformation, or hydrolyze and transfer to the liquid phase, while Au is retained as much as possible in a solid or gold-rich form, reducing the interference of non-gold components on subsequent gold leaching from the source.

[0007] In some embodiments, the gold-containing solder is AuSn; Step S2: The first leaching residue contains a non-metallic component, wherein the non-metallic component is Sn, and the residual amount of Sn is ≤0.52wt%.

[0008] In this invention, due to the complex microstructure between Au and Sn in the solder, and the potential formation of gold-containing particles of varying sizes and morphologies after melting, cooling, and subsequent leaching, a small amount of Sn may remain in the Au-enriched solid phase after the first leaching. If the residual amount exceeds a preset threshold, directly performing a second-stage leaching process would lead to non-productive consumption of the leaching agent and increase the risk of gold encapsulation by colloids or precipitates. Therefore, the residual amount detection results are fed back to the first-stage leaching step, and leaching and detection are repeated until the threshold is reached before initiating the second-stage leaching process, thus achieving process control of the second-stage leaching initiation conditions.

[0009] In some embodiments, in step S1, the heat treatment is performed under an inert atmosphere, the heat treatment temperature is 290-310°C, and the holding time is 5-20 min. Nitrogen or other inert atmospheres can reduce the oxidation of solder and metallization layer during the heating process, reduce the obstruction of solder oxide film to melt separation, and avoid the formation of oxide coating layers that are difficult to remove. At the same time, controlling the heat treatment temperature allows the solder to melt without significantly damaging the SiC substrate and Ti / Au metallization layer, which is beneficial for achieving spatial separation between the solder layer and the metallization layer.

[0010] In some embodiments, in step S2, the HCl-H2O2 system comprises 1–3 mol / L HCl and 0.05–1.5 mol / L H2O2; HCl provides an acidic medium and chloride ions, causing the dissolved non-metallic components to form soluble chlorides or chloride complexes; H2O2 provides oxidation, promoting the conversion of low-valence non-metallic components to higher-valence forms that are more easily absorbed into the liquid phase. Limiting the acid concentration, oxidant concentration, and liquid-solid ratio within the above ranges can ensure the removal rate of non-metallic components while reducing the possibility of Au being oxidized and dissolved, as well as the local temperature rise caused by excessively rapid reactions.

[0011] In some embodiments, in step S2, the liquid-to-solid ratio of the gold-containing solder to the HCl-H2O2 system is 5–15 mL / g, the leaching temperature is 30–50°C, and the leaching time is 20–60 min.

[0012] In some embodiments, in step S3, the I2-KI aqueous solution contains 2–10 g / L of I2 and 20–60 g / L of KI; I2 forms I3 in the presence of KI. - Furthermore, it reacts with Au to form a soluble gold iodide complex, allowing Au to enter the second leaching solution from the first leaching residue. KI is used to improve the solubility of iodine and stabilize the iodine complex; light-protected operation reduces the photodecomposition of I2, maintaining the effective oxidation capacity of the gold leaching system. By pre-removing non-gold components, the consumption of I2 by reducing species such as Sn can be reduced, and the adsorption of gold complex ions by hydrolyzed colloids can be reduced.

[0013] In some embodiments, in step S3, the liquid-to-solid ratio of the first leaching residue to the I2-KI aqueous solution is 10-20 mL / g, and the leaching temperature is 15-35°C.

[0014] In some embodiments, in step S4, the acidic medium includes sulfuric acid, and the fluorinated compound includes sodium fluoride; and / or The electrolyte comprises 0.1–1 mol / L H₂SO₄ and 0.02–0.3 mol / L NaF.

[0015] In some embodiments, in step S4, the current density of the anode pulse is 5–50 mA / cm². 2 The energizing time is 1–10 s, the de-energizing time is 5–30 s, and the number of pulses is 50–500. During the anodic energizing phase, Ti dissolves; during the de-energizing phase, the concentration polarization near the Ti surface is alleviated, and the continuous growth and accumulation of the passivation film is reduced. Compared with continuous anodic electrolysis, the pulsed mode can extend the selective dissolution window of Ti and reduce Au pitting caused by the increase in local potential after the passivation film thickens. The current density determines the anodic dissolution rate of Ti, the energizing time determines the amount of dissolution per pulse, the de-energizing time is used for diffusion layer recovery and passivation film state adjustment, and the number of pulses is used to match the Ti layer thickness and the degree of interface removal. By synergistically controlling these parameters, Ti can be removed starting from the edge of the Au film, exposed areas, or defect areas, and extending along the Ti / Au interface, avoiding local dissolution of Au due to excessively high current density. By controlling the anodic pulse current density, energizing time, de-energizing time, and number of pulses, the possibility of anodic dissolution of Au can be reduced while ensuring continuous Ti removal.

[0016] In some embodiments, in step S5, the ultrasonic treatment frequency is 20–80 kHz, the power is 50–500 W, and the treatment time is 0.5–10 min. After Ti is gradually removed from the Ti / Au interface, the effective support area and adhesion between the Au film and SiC decrease. The microjets and local shear forces generated by ultrasonic cavitation can act on the edges of the Au film and the undercut region of the interface, causing the unsupported Au film to detach as a whole or in sheets without needing to dissolve the Au again into the leachate.

[0017] In some embodiments, the first gold recovery product is combined with the Au membrane and then smelted to obtain a gold product; The smelting is carried out under an inert atmosphere at a temperature of 1000–1200°C.

[0018] In this invention, the gold obtained through the solder path typically contains a small amount of residual solder components, while the gold obtained through the Ti / Au path is mainly a sheet-like Au film. After being recovered separately, the gold is then collected and refined. The appropriate washing, melting, or refining conditions can be selected according to the impurity composition of different gold products to reduce the separation difficulties caused by the coexistence of different impurities.

[0019] In summary, compared with the prior art, the present invention achieves the following technical effects: (1) The present invention treats the solder layer and the Ti / Au-SiC metallization layer separately by pretreatment and gold-containing region diversion, avoiding Sn, Ti, Au and other chip metals from entering the same leaching system at the same time, reducing the complexity of subsequent separation and purification; by removing non-gold components in the solder through primary selective leaching, and combined with residual detection and feedback control, the secondary leaching gold is started within the preset impurity tolerance range, reducing non-productive consumption of leaching agent and gold loss caused by colloidal adsorption.

[0020] (2) This invention removes the Ti adhesion layer through pulsed electrochemical processes, causing the Ti / Au interface to be cleaved and promoting Au film desorption. This reduces Au anodic dissolution while enabling direct collection of the Au film. The Au film, which has lost interfacial support, is recovered non-dissolvingly through ultrasonic, vibration, or liquid flow impact, reducing the number of Au transfers in the liquid phase, reducing gold loss, and facilitating the acquisition of high-purity flake gold products.

[0021] (3) In the gold recovery process, the present invention does not require the complete SiC substrate to be broken up as a whole, and reduces substrate damage by controlling the heat treatment and electrochemical treatment conditions, thereby preserving the possibility of SiC substrate recycling.

[0022] (4) The two recycling paths of the present invention are independent of each other and ultimately achieve unified refining. This allows for the adoption of appropriate recycling processes for different gold sources, and also enables the comprehensive recovery of Au in solder and Au in Ti / Au metallization layers. Compared with aqua regia leaching, cyanide leaching, or high-temperature roasting routes, the present invention reduces the separation burden caused by multi-metal co-leaching and can reduce the environmental and equipment load caused by cyanide-containing waste liquid, nitrogen oxide-containing tail gas, and high-temperature treatment. Attached Figure Description

[0023] To more clearly illustrate the technical solutions of the embodiments of the present invention, the accompanying drawings used in the embodiments will be briefly introduced below. It should be understood that the following drawings only show some embodiments of the present invention and should not be regarded as a limitation on the scope. For those skilled in the art, other related drawings can be obtained based on these drawings without creative effort.

[0024] Figure 1 This is a process system flow diagram of the present invention.

[0025] Figure 2 This is a schematic diagram of a discarded wafer structure for a semiconductor laser epitaxial wafer. Detailed Implementation

[0026] To enable those skilled in the art to better understand the present invention, the technical solutions in the embodiments of the present invention are clearly and completely described. 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 skilled in the art without creative effort should fall within the scope of protection of the present invention.

[0027] The overall process of this invention revolves around the shunting and recycling of different gold-containing regions in a multilayer circuit stack. First, the gold-containing solder layer is separated from the Ti / Au-SiC metallization layer through heat treatment, cutting, or peeling. Specifically, the solder path selectively removes non-gold components such as Sn, Ge, Si, or In, and then determines whether to proceed to secondary iodine-potassium iodide immersion based on the residual amount. The Ti / Au path selectively removes the Ti adhesion layer through pulsed electrochemical methods, and then directly collects the Au film after desorption using ultrasound or liquid flow impact. Finally, the gold products obtained from both paths are refined and collected separately, thereby reducing interference from multi-metal co-immersion while achieving the recycling of Au and SiC substrates. See the process system flowchart below. Figure 1 .

[0028] Unless otherwise specified, the experimental methods used in the following examples are conventional methods. Unless otherwise specified, all materials and reagents used are commercially available.

[0029] Example 1 This embodiment provides a method for recovering gold from a multilayer circuit stack structure. The material to be recovered is a 4-inch SiC-based semiconductor laser epitaxial wafer, and its structure from top to bottom is as follows: Figure 2 As shown, the components are: a laser chip (GaAs-based, with Ti / Pt / Au ohmic contacts on the surface, chip size 1 mm × 1 mm × 100 μm), an AuSn20 eutectic solder layer (Au 80 wt%, Sn 20 wt%, thickness 20 μm, melting point 280℃), a front Ti / Au metallization layer (Ti 100 nm, Au 0.2 μm), a SiC substrate (4H-SiC, heavily N-type doped, resistivity 0.02 Ω·cm, diameter 100 mm, thickness 350 μm), and a back Ti / Au metallization layer (same as the front, symmetrical).

[0030] The specific steps are as follows: S1: Place the wafer to be recycled into a tube furnace under nitrogen protection (flow rate 200 mL / min), heat to 300℃ at a rate of 5℃ / min, and hold for 10 min to completely melt the AuSn20 solder layer. Remove the wafer and, while still hot, use a vacuum pick to remove the laser chips one by one. The gold contact layer on the chips is carried away with the chips and collected separately. Use a quartz scraper to scrape off the molten solder remaining on the Ti / Au surface, collecting it as the first gold-containing material (solder layer material), weighing approximately 0.35 g. The Ti / Au-SiC metallized substrate after the solder has been scraped off is used as the second gold-containing material. Check the integrity of the SiC substrate: visually inspect for cracks and chipping, and the diameter should still be 100 mm when measured with calipers.

[0031] S2: Place the first gold-containing material in a 100 mL polytetrafluoroethylene beaker.

[0032] Preparation of the first leachate: Dilute 17.2 mL of 36 wt% HCl to 100 mL, add 10 mL of 30 wt% H2O2, and stir until well mixed. Add 3.5 mL of the first leachate (liquid-to-solid ratio 10:1, mL:g) to a beaker, place in a constant temperature water bath at 40℃, and stir magnetically at 200 rpm for 30 min.

[0033] Under acidic conditions, Sn in AuSn solder undergoes oxidative dissolution to form Sn. 2+ Subsequently, H2O2 further promoted Sn 2+ To Sn 4+ The transformation promotes the migration of Sn from the solder phase to the liquid phase. After the reaction is complete, the solution changes from colorless and transparent to light yellow, and grayish-brown sponge-like solid residue can be seen at the bottom of the beaker.

[0034] S3: The material obtained in S2 was vacuum filtered, and the filtrate (first leachate, containing SnCl2 / SnCl4) and filter residue (first leachate residue) were collected separately. The first leachate residue was washed three times with deionized water (20 mL each time), then washed once with anhydrous ethanol (10 mL), and dried in a vacuum drying oven at 60℃ for 2 h, and weighed. Approximately 50 mg of the dried first leachate residue was taken, digested with aqua regia, and diluted to 50 mL. The Sn residue content was determined by ICP-OES. The results showed that the Au content in the first leachate residue was 92.3 wt%, and the Sn residue content Cx = 0.31 wt%.

[0035] S4: Weigh the first leaching residue from S3 and place it in a 100 mL brown light-proof bottle. Prepare the second leaching solution: an aqueous solution of 5 g / L I2 + 40 g / L KI (weigh 0.5 g of I2 and 4.0 g of KI and dissolve them in 100 mL of deionized water, stirring until completely dissolved). Add 48 mL of the second leaching solution (liquid-solid ratio 15:1, mL:g) to the light-proof bottle, and stir magnetically at 200 rpm at room temperature (25℃) for 2 h in the dark.

[0036] Reaction principle: 2Au + I2 + 2KI → 2KAuI2, gold reacts with iodine-potassium iodide to form a soluble complex KAuI2 which enters the liquid phase.

[0037] S5: Vacuum filter the leachate obtained in S4 to remove insoluble residues and collect the Au-containing filtrate. Add sodium sulfite solution (10 wt% Na2SO3 aqueous solution) to the filtrate to reduce gold: Adjust the pH to 8-9 with 0.1 mol / L NaOH solution, add Na2SO3 at 1.5 times the theoretical amount of Au (approximately 0.48 g of Na2SO3 per 1 g Au), heat in a water bath to 60℃, and stir for 30 min.

[0038] Reaction principle: 2KAuI2 + Na2SO3 + H2O → 2Au↓ + Na2SO4 + 2KI + 2HI, gold precipitates as brown sponge gold powder. After natural cooling to room temperature, the gold powder is collected by vacuum filtration, washed three times with deionized water and once with anhydrous ethanol, dried under vacuum at 60℃ for 4 h, and weighed to obtain 0.258 g of the first gold recovery product. The gold recovery rate of the solder pathway is 92.1% (0.258 / 0.280 g).

[0039] S6: Pulsed electrochemical removal of Ti / Au-SiC metallization layer The working electrode is a Ti / Au-SiC substrate (front side up) obtained in S1, fixed in a PTFE electrode holder, exposing the front Ti / Au layer. Copper wires are attached to the edge of the Au layer using conductive silver paste for electrical connection. The counter electrode is a platinum sheet electrode (4 cm² area). 2 The electrodes are placed parallel to each other opposite the working electrode, with a spacing of 2 cm; the reference electrode is a saturated calomel electrode (SCE), which is connected to the electrolytic cell through a salt bridge.

[0040] Electrolyte preparation: 0.5 mol / L H2SO4 + 0.1 mol / L NaF (dilute 27.2 mL of concentrated H2SO4 to 1 L, add 4.2 g of NaF, and stir to dissolve). Pour the electrolyte into a 500 mL double-jacketed glass electrolytic cell and connect an external constant temperature water bath to maintain 25℃.

[0041] Pulsed electrochemical parameters: constant current pulse mode, anodic pulse current density J = 20 mA / cm² 2 (The effective area of ​​Ti / Au in a single exposure is 15 cm²) 2 (corresponding current 300 mA), power-on time ton = 5 s, power-off time toff = 10 s, pulse period T = ton + toff = 15 s, number of pulses N = 200 times (total processing time 50 min).

[0042] During the anodic pulse phase, Ti undergoes anodic dissolution (Ti → Ti). 4+ + 4e - ), Ti 4+ With F - Complexation to form [TiF6] 2- Entering the electrolyte, F - Simultaneously, the TiO2 passivation film is destroyed; during the power-off phase, concentration polarization is eliminated, and the growth rate of the new passivation film on the Ti surface decreases, creating conditions for the next pulse dissolution. Au in 0.5 mol / L H2SO4 (Cl-free) - In the system, at 20 mA / cm 2 No significant anodic dissolution occurs at the current density. Ti preferentially dissolves in the exposed area and in local defect regions between the Au film and the Ti layer, gradually forming voids or interfacial undercut regions extending along the Ti / Au interface as electrochemical treatment progresses.

[0043] S7: After 200 pulsed electrochemical treatments (S6), the edges of the Au film showed obvious peeling. The substrate was removed from the electrolytic cell, rinsed with deionized water, immersed in a beaker containing 200 mL of deionized water, and placed in an ultrasonic cleaner for 2 min at 40 kHz and 200 W. The ultrasonic vibration caused the Au film, which had formed an interface undercut, to desorb, resulting in a sheet-like Au film. Inspection of the front side of the SiC substrate: the surface was the natural color of SiC (dark gray), with no obvious Au residue and no obvious corrosion.

[0044] S8: Vacuum filter the deionized water suspension containing the Au membrane obtained in S7. Wash the Au membrane on the filter membrane three times with deionized water and once with anhydrous ethanol. Dry the filter membrane together in a vacuum drying oven at 60°C for 2 hours. Remove the Au membrane sheet from the filter membrane with tweezers.

[0045] The SiC substrate with the front Au film removed was flipped over so that the back side was facing up. Pulsed electrochemical removal of Ti was performed using the same parameters (freshly prepared electrolyte). Ultrasonic desorption was then performed using the same parameters. The substrate was collected and weighed. The gold recovery rate via the Ti / Au pathway was 0.059 g, with a theoretical gold content of 0.2 μm Au layer thickness × 78.5 cm² area. 2Au density 19.32 g / cm³ 3 ×2 surfaces = 0.061g, with a gold recovery rate of approximately 97.2% via the Ti / Au pathway.

[0046] S9: Combine the first gold recovery product (sponge gold from the solder path, 0.258 g) and the second gold recovery product (Au film from the Ti / Au path, 0.059 g), with a total mass of 0.317 g. Place in a quartz crucible, add 0.05 g of borax as a flux, and place in a high-temperature furnace that can be purged with argon gas. Under argon protection (flow rate 500 mL / min), heat to 1100℃ at a rate of 10℃ / min and hold for 20 min to completely melt the gold. Allow to cool naturally to room temperature, remove the gold ingot, and weigh 0.315 g. Take a small sample from the surface of the gold ingot and analyze its purity using ICP-OES.

[0047] The combined gold recovery rate of the two recycling paths was 93.0%. After smelting, 0.315 g of gold product was obtained, with a comprehensive gold recovery rate of 92.4% calculated based on the final gold product quality. The gold product purity was 99.92 wt%. No macroscopic cracks, edge chipping, or obvious corrosion were observed on the SiC substrate.

[0048] Example 2 The same batch of gold-containing material (solder layer material, AuSn20 eutectic solder, 0.35 g per group, containing 0.280 g Au) was divided into 7 groups (A1~A7). Different Sn residues were obtained by changing the primary selective leaching conditions. The leaching conditions for each group are shown in Table 1. Table 1

[0049] After the first-stage selective leaching of each group, solid-liquid separation was performed. The resulting filter residue was washed three times with deionized water, dried to constant weight, and the residual Sn content (C) was determined by ICP-OES. x .

[0050] Subsequently, the first leaching residues from each group were uniformly subjected to secondary gold selective leaching in S5: the second leaching solution was a 5 g / L I2 + 40 g / L KI aqueous solution with a liquid-to-solid ratio of 15:1, at room temperature (25℃), magnetically stirred at 200 rpm, and reacted in the dark for 2 h. The S5 conditions were identical for all groups, with the only difference being the residual Sn content in the first leaching residue.

[0051] After S5 leaching, the following indicators were tested: (1) Au leaching rate: The Au concentration in the leachate was determined by ICP-OES, and the Au leaching rate was calculated as: Au mass in leachate / Au mass in first leaching residue × 100%.

[0052] (2) I2 consumption: The remaining I2 after leaching was titrated with 0.1 mol / L Na2S2O3 standard solution (I2 + 2 Na2S2O3 → 2NaI + Na2S4O6), the actual I2 consumption was calculated, and compared with the theoretical consumption (calculated according to the stoichiometry of Au leaching amount: 2Au + I2 → 2AuI, that is, 0.5 mol I2 is consumed for every 1 mol Au) to obtain the I2 consumption / theoretical amount ratio.

[0053] (3) Gold recovery rate: The S5 leachate was reduced with sodium sulfite according to S6, the gold powder was collected by filtration, dried and weighed, and the gold recovery rate was calculated as: gold powder mass / Au mass in the first leaching residue × 100%.

[0054] The results for each group are shown in Table 2: Table 2

[0055] As shown in Table 2, with the increase of Sn residue C in the first leaching residue... x As C increases, the Au leaching rate gradually decreases in the second-stage gold leaching process, while the non-productive consumption of I2 gradually increases. x When the concentration is ≤ 0.52 wt%, the Au leaching rate is ≥ 91.4%, the actual I2 consumption is only 1.03~1.07 times the theoretical amount, and the excess consumption is ≤ 7%, mainly due to Sn. 2+ Caused by a redox reaction with I₂. When 0.52... <C x When the concentration of Sn is ≤ 2 wt%, the Au leaching rate decreases from 87.6% to 80.3%, and when the concentration of Sn is > 2 wt%, the Au leaching rate drops sharply to below 69%. This indicates that a high residual Sn concentration significantly interferes with the subsequent gold leaching process.

[0056] Further analysis of I2 consumption reveals that when Cx > 0.52 wt%, the proportion of excess I2 consumption increases rapidly with increasing Sn residue: C x =1.85 wt% in excess of 34%, C x = 8.76 wt% in excess of 118%, C x = 19.8 wt% is in excess by 376%. This indicates that as the residual Sn content increases, the non-productive consumption of I2 by Sn increases significantly, thereby reducing the utilization rate of effective I2 in the leaching gold system.

[0057] In summary, a Sn residue of 0.52 wt% is the key critical range affecting the subsequent iodine-potassium iodide leaching gold effect. When C xWhen the concentration of Sn is ≤0.52 wt%, the Au leaching rate can still be maintained above 91.4%, and the excess I2 consumption does not exceed 7%, indicating that the consumption of Sn by the gold leaching oxidant and the colloidal adsorption are both at a low level; while when C x When the concentration is >0.52 wt%, both the excessive consumption of I2 and the decrease in Au leaching rate are significantly aggravated.

[0058] To further determine the effect of Sn residue after primary leaching on secondary iodine-potassium iodide leaching, three additional critical range experiments (A8-A10) were conducted in addition to the original A1-A7 groups. For each group, 0.35 g of AuSn20 solder material was used, and primary selective leaching was performed at a liquid-to-solid ratio of 10:1, a stirring speed of 300 rpm, and a leaching temperature of 40±1℃. Group A8 was leached with 2.0 mol / L HCl and 0.30 mol / L H2O2 for 28 min; Group A9 was leached with the same leaching solution for 26 min; and Group A10 was leached with the same leaching solution for 24 min. Immediately after primary leaching, solid-liquid separation was performed. The filter residue was washed three times with deionized water and dried to constant weight. The Sn residue was determined using ICP-OES. Each experiment was performed in triplicate, and the values ​​in the table are the average values ​​of the three experiments.

[0059] Table 3

[0060] The results from groups A8 to A10 show that when the Sn residue is 0.38–0.48 wt%, the Au leaching rate is not less than 91%, the I2 excess consumption is not more than 6%, and the gold recovery rate is not less than 90%. When the Sn residue increases to 0.55 wt%, the Au leaching rate decreases to 90.8%, the gold recovery rate decreases to 89.7%, and the leachate begins to show slight turbidity. Combining the data from groups A1 to A7, it was determined that entering the secondary gold leaching step when the Sn residue does not exceed 0.52 wt% can balance the Au leaching rate, I2 utilization rate, and gold recovery rate; when the Sn residue is greater than 0.52 wt%, it returns to S2 for a second primary leaching. Therefore, this application sets 0.52 wt% as the preset threshold for the Sn residue after the first primary leaching. When the detection shows: C x When C ≤ 0.52 wt%, the first leaching residue is deemed to meet the conditions for entering the secondary selective gold leaching step; when C x When the Sn content is >0.52 wt%, the process returns to the first-stage selective leaching step to further reduce the Sn residue.

[0061] Comparative Example 1 The only difference between this comparative example and Example 1 is the S6 pulse electrochemical parameters; the other steps are the same as in Example 1.

[0062] Take the obtained Ti / Au-SiC metallized substrate, rinse it with deionized water and dry it at 60°C. Fix the substrate face up in a PTFE fixture, exposing approximately 15 cm of the face. 2 The Ti / Au metallized region was covered, and the remaining areas were masked with PTFE tape. The Ti / Au metallized layer on the substrate was used as the working electrode, and copper wires were led out from the edge of the Au layer using conductive silver paste. A platinum sheet with an area of ​​4 cm² was used as the counter electrode. 2 The electrodes were arranged parallel to each other opposite the working electrode, with a spacing of 2 cm. A saturated calomel electrode served as a reference electrode and was connected to the electrolytic cell via a salt bridge. The electrolyte was a mixed aqueous solution of 0.5 mol / L H₂SO₄ and 0.1 mol / L NaF, with a volume of 300 mL, and the temperature was maintained at 25 ± 2 °C. The electrolyte was chloride-free to reduce the possibility of anodic dissolution of Au. Before electrolysis, the exposed area of ​​the working electrode, the substrate mass, and the surface and edge conditions of the Au layer were recorded.

[0063] The working electrode is continuously energized at a constant current with a current density of 20 mA / cm². 2 The corresponding operating current is approximately 300 mA, and the continuous energizing time is 1000 s, with no power-off phase. To correspond with the parameter representation in Example 1, this comparative example is denoted as a continuous energizing mode with ton=15 s and tof=0 s; the actual energizing process is a continuous constant energizing for 1000 s, rather than repeating 15 s pulses. During continuous electrolysis, the working electrode potential and electrolytic cell current are recorded every 100 s, and the electrolyte color and the state of bubbles on the substrate surface are observed.

[0064] After continuous electrolysis, the substrate was removed and all electrolyte was collected. A quantitative electrolyte sample was taken, acidified, and diluted to a fixed volume, and then the Ti and Au contents were determined by ICP-OES. The Ti removal rate was calculated using the following formula: Ti removal rate = (mass of Ti in the electrolyte / theoretical mass of the Ti layer in the exposed area before treatment) × 100%. Wherein, the theoretical mass of the Ti layer in the exposed area before treatment is based on a Ti layer thickness of 100 nm and an exposed area of ​​15 cm². 2 Ti density 4.506 g / cm³ 3 Calculation: Au dissolution loss is calculated based on the Au content in the electrolyte.

[0065] After electrolysis, the substrate surface was rinsed three times with 50 mL of deionized water each time. The substrate was then immersed in 200 mL of deionized water and sonicated at 40 kHz and 200 W for 2 min. The sonicated solution was filtered through a 0.45 μm filter membrane. The Au membrane on the filter membrane was washed three times with deionized water and once with anhydrous ethanol, then dried under vacuum at 60 ℃ for 2 h and weighed. The substrate and Au membrane were observed using a stereomicroscope, and the Au membrane desorption area, central adhesion area, and pitting corrosion on the Au membrane surface were recorded. The back side of the substrate was flipped over and treated under the same continuous electrolysis, sonication, and filtration conditions.

[0066] The gold recovery rate along the solder path was 92.1%, and the gold recovery rate along the Ti / Au path was approximately 75%. Only partial desorption of the Au film on the front side was observed, with a Ti removal rate of approximately 60%. A small amount of Au was detected in the electrolyte. The Au films collected from both sides were combined with the gold powder from the solder path and processed according to the S10 melting and ICP-OES detection method described in Example 1. The gold product purity was 99.75 wt%, and the total gold recovery rate, corrected for actual gold content, was approximately 89.1%. In continuous electrolysis mode, the TiO2 passivation film grew and thickened rapidly under continuous anodic polarization, causing the Ti dissolution current efficiency to drop to approximately 40%. Furthermore, the thickening of the passivation film increased the local potential, triggering pitting dissolution of Au.

[0067] Comparative Example 2 This comparative example is used to simulate the existing technology of mechanically crushing and then leaching a multilayer circuit stack structure. Except for mechanical crushing, aqua regia leaching, and mixed solution separation, this comparative example does not perform the gold-containing region diversion, primary selective leaching of non-gold components of solder, Sn residue feedback control, and pulsed electrochemical Ti removal treatment of the Ti / Au-SiC metallization layer as described in Example 1.

[0068] Waste semiconductor laser wafers from the same batch as those in Example 1, with the same structure and comparable quality, were used as the materials to be processed. Before processing, the total mass of the wafers to be processed was weighed using an analytical balance, and parallel samples were taken from the same batch of wafers. After complete acid digestion, the total gold content was determined using ICP-OES, which served as the benchmark value for the gold content in this comparative example. The wafers to be processed were placed in a mechanical crushing device and crushed first by low-speed shearing, and then further crushed by impact until the maximum size of the resulting fragments was less than 5 mm. During the crushing process, the laser chip, solder layer, front Ti / Au metallization layer, and back Ti / Au metallization layer were not separated or collected separately. After crushing, all fragments were mixed evenly, the mass of the crushed material was recorded, and random samples were taken to observe the size and composition of the fragments. Concentrated hydrochloric acid and concentrated nitric acid were measured separately and mixed in an acid-resistant container at a volume ratio of 3:1 to obtain aqua regia. Take the crushed mixture and add it to a 500 mL three-necked flask equipped with a reflux condenser, thermometer, mechanical stirrer, and tail gas absorption device. Add freshly prepared aqua regia at a liquid-to-solid ratio of approximately 10:1, with a total addition volume of 200 mL. Start the mechanical stirrer at 300 rpm. Heat the reaction system to 60±2℃ using a water bath and leach at this temperature for 4 hours. Record the reaction temperature, solution color, bubble generation, and material state every 30 minutes during the leaching process. Pass the tail gas from the three-necked flask sequentially into a gas washing bottle containing alkaline absorbent to absorb the nitrogen oxide-containing tail gas generated during the leaching process. After the reaction is complete, stop heating and stirring, and allow the reaction slurry to cool to room temperature.

[0069] During the overall leaching process described above, components such as Sn in the solder, Ti and Au in the Ti / Au metallization layer, and Pt and Au in the chip contact layer enter the mixed solid-liquid system. Incompletely dissolved components proceed with the insoluble residue to the subsequent solid-liquid separation step. The cooled leaching slurry is subjected to solid-liquid separation using an acid-resistant filtration device with a 0.45 μm pore size acid-resistant membrane as the filter medium. The filter residue is rinsed three times with deionized water, using 50 mL of deionized water each time. The washings are combined with the leaching solution to obtain a mixed leaching solution. The filter residue is dried at 60°C to constant weight and then weighed.

[0070] The mixed leachate was brought to a final volume and sampled. After dilution with deionized water at a predetermined ratio, the contents of Au, Sn, Ti, Pt, Ge, Ni, and Cu were determined by ICP-OES. The dried insoluble residue was observed under a stereomicroscope, and the residual gold content was determined by XRF or acid digestion-ICP-OES. Based on the detection results of the leachate and insoluble residue, the distribution of each metal between the leachate and the residue was recorded. The tested mixed leachate was placed in an acid-resistant beaker equipped with mechanical stirring. A 10 wt% urea aqueous solution was slowly added at room temperature while stirring continuously for 5 minutes until no obvious bubbles were generated, in order to reduce the influence of residual nitric acid in the mixed leachate on the subsequent reduction reaction. Subsequently, the mixed leachate was heated and concentrated to about one-third of its initial volume, and chloride ions were added with hydrochloric acid to maintain the solution in a chloride medium. A 100 g / L ferrous sulfate aqueous solution was prepared. Based on the Au content in the mixed leachate measured by ICP-OES, the theoretically required molar amount of ferrous sulfate was calculated, and ferrous sulfate solution was added at 1.5 times the theoretical amount. The reaction temperature was controlled at 60℃, the stirring speed at 300 rpm, and the reaction was carried out for 60 min. After the reaction was completed, stirring was stopped, and the mixture was allowed to stand for 30 min to allow the gold precipitate to grow fully.

[0071] A brown gold precipitate was collected by filtration through a 0.45 μm pore size membrane. The gold precipitate was washed three times with 50 mL of 0.5 mol / L hydrochloric acid, three times with deionized water, and once with anhydrous ethanol. After washing, the precipitate was dried in a vacuum drying oven at 60 °C for 4 h to obtain the primary gold precipitate.

[0072] The dried primary gold precipitate was placed in a quartz crucible, and 0.05 g of borax was added as a flux. Under argon protection, the temperature was increased to 1100℃ at a rate of 10℃ / min and held for 20 min. After cooling, crude gold product was obtained. The crude gold product was placed in 50 mL of 0.5 mol / L hydrochloric acid and stirred at 60℃ for 30 min. It was then filtered and washed three times with deionized water. Next, 50 mL of 10% (v / v) nitric acid aqueous solution was added, and the mixture was treated at 80℃ for 30 min. The mixture was then filtered and washed three times with deionized water. The hydrochloric acid and nitric acid washings were repeated once each. Finally, the mixture was washed once with anhydrous ethanol and dried under vacuum at 60℃ for 4 h to obtain the gold product. The mass of the gold product was weighed, and the Au purity was determined using ICP-OES. The mass of Au in the gold product was calculated using the following formula: Au mass = Gold product mass × Au mass fraction.

[0073] The gold product obtained from the reduction of the mixed leaching solution, the gold-containing material collected during filtration and washing, and the residual gold in the insoluble residue were all separately assessed for quality. The total gold recovery rate was calculated using the following formula: Total gold recovery rate = Total mass of Au in all recovered gold products / Total mass of Au in the material to be processed before crushing × 100%. The total mass of Au in the material to be processed before crushing was based on the ICP-OES test results of a parallel sample from the same batch after complete acid digestion. The total mass of Au in all recovered gold products included the Au in the gold product obtained from the reduction of the mixed leaching solution, as well as the Au recovered separately from the residue or washing solution. If residual gold in the residue was not further recovered, it was considered as unrecovered gold and included in the recovery rate loss, not in the gold product quality.

[0074] The comparative example yielded a total gold recovery rate of 89.2%, with an Au purity of 97.8 wt% in the gold product. Multiple elements, including Au, Sn, Ti, Pt, Ge, Ni, and Cu, were detected simultaneously in the mixed leaching solution, requiring at least several subsequent steps such as solution conditioning, gold reduction, precipitation washing, crude gold smelting, and acid leaching purification. Nitrogen oxide-containing tail gas was generated during the leaching process, necessitating treatment by a tail gas absorption device. The total gold recovery rate and gold product purity of this comparative example were lower than those of Example 1, and the reuse of the intact SiC substrate could not be achieved. These results indicate that direct mechanical crushing and whole-body leaching allows multiple metals to simultaneously enter the same leaching system, increasing the difficulty of gold separation and purification, and leading to recovery losses.

[0075] The above description is only a preferred embodiment of the present invention and is not intended to limit the present invention. Any modifications, equivalent substitutions, improvements, etc., made within the spirit and principles of the present invention should be included within the protection scope of the present invention.

Claims

1. A method for gold recovery from electronic components with composite plating, characterized in that, The surface of the electronic component includes a gold solder layer and a Ti / Au composite metallization layer, and the method includes: S1: The electronic component is subjected to heat treatment to melt the gold solder layer, and the chip is separated from the substrate to collect the gold solder and the substrate 1 with the Ti / Au metallization layer; S2: The gold-containing solder is placed in an HCl-H2O2 system, and the first leaching residue is obtained by solid-liquid separation; S3: Place the first leaching residue in an I2-KI aqueous solution and react in the dark for 1.5~3 h, then collect the gold-containing filtrate; S4: The substrate 1 is placed in an electrolyte containing an acidic medium and a fluorine-containing compound. The substrate 1 is used as the working electrode. Alternating anodic pulses and de-energizing pulses are applied to selectively dissolve Ti and form an interfacial cleavage along the Ti / Au interface. S5: The substrate 1 after pulse electrochemical treatment is subjected to ultrasonic treatment to detach the Au film from the surface of the substrate 1, and the Au film is collected.

2. The gold recovery method according to claim 1, characterized in that, The gold-containing solder in step S1 is one or more of AuSn, AuGe, AuSi, AuIn, and SnAgCu.

3. The gold recovery method according to claim 2, characterized in that, The gold-containing solder is AuSn; Step S2: The first leaching residue contains a non-metallic component, wherein the non-metallic component is Sn, and the residual amount of Sn is ≤0.52wt%.

4. The gold recovery method according to claim 1, characterized in that, In step S1, the heat treatment is carried out under an inert atmosphere, the heat treatment temperature is 290-310℃, and the holding time is 5-20 min.

5. The gold recovery method according to claim 1, characterized in that, In step S2, the HCl-H2O2 system contains 1 to 3 mol / L HCl and 0.05 to 1.5 mol / L H2O2.

6. The gold recovery method according to claim 1, characterized in that, In step S2, the liquid-to-solid ratio of the gold-containing solder to the HCl-H2O2 system is 5–15 mL / g, the leaching temperature is 30–50℃, and the leaching time is 20–60 min.

7. The gold recovery method according to claim 1, characterized in that, In step S3, the I2-KI aqueous solution contains 2-10 g / L of I2 and 20-60 g / L of KI.

8. The gold recovery method according to claim 1, characterized in that, In step S3, the liquid-to-solid ratio of the first leaching residue to the I2-KI aqueous solution is 10-20 mL / g, and the leaching temperature is 15-35℃.

9. The gold recovery method according to claim 1, characterized in that, In step S4, the acidic medium includes sulfuric acid, and the fluorine-containing compound includes sodium fluoride; and / or The electrolyte comprises 0.1–1 mol / L H₂SO₄ and 0.02–0.3 mol / L NaF.

10. The gold recovery method according to claim 1, characterized in that, In step S4, the current density of the anode pulse is 5–50 mA / cm². 2 The power-on time is 1–10 s, the power-off time is 5–30 s, and the number of pulses is 50–500.

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