Electrochemical technology to drive spent lithium-ion batteries (LIBs) recycling: recent progress, and prospects

Resource properties of spent LIBs

As a highly recyclable secondary resource, spent LIBs contain substantial quantities of valuable metallic elements, including Co, Ni, and Li, with concentrations surpassing those in natural ores^[22,23]. The manufacturing cost of LIBs also remains highly dependent on the fluctuation of raw material prices, such as lithium. As shown in Figure 1C, the price of lithium carbonate rapidly increased to $68,100 in 2022, a 440% increase compared to the previous year. However, by 2023, the price of lithium carbonate dropped back to $46,000, a significant decrease of 32.45% compared to 2022^[24]. These substantial price fluctuations will inevitably affect the stability of supply and profitability for LIB manufacturers. Therefore, appropriate recycling technologies can not only promote the establishment of a closed-loop and stable supply chain for these materials, but also diminish the complexity of extraction from natural ores. Consequently, this approach may lead to a potential reduction in the cost of LIB materials, providing an economic advantage in the manufacturing of LIBs^[25].

Environmental problems of spent LIBs

Increasing demand for LIBs has precipitated a marked increase in spent units. Currently, only a minor portion of these batteries is subjected to recycling and resource recovery, with the majority still relegated to landfills or disposed of illegally^[26]. LIBs, which contain heavy metals such as Co, Ni, and manganese (Mn) (all recognized carcinogens), and toxic organic electrolytes, may emit harmful HF gas upon contact with water. As a result, these materials are considered hazardous waste. The improper disposal of these batteries can lead to substantial emissions of noxious substances into the environment, threatening ecological integrity and public health^[27]. Consequently, recycling these batteries not only enhances resource efficiency, but also advances the development of a green and sustainable LIBs sector^[28].

Uneven distribution of production resources

The essential raw materials used for LIB production are primarily found in a few countries. In terms of the distribution of natural resources, Chile, Australia, and Argentina hold over 70% of the world’s lithium resources, as shown in Figure 1D-F. The Democratic Republic of Congo and Australia account for 66% of global cobalt resources, while 41% of global nickel resources are concentrated in Australia and Indonesia. This uneven distribution of natural resources has led to disparities in the production output of Li, Co, and Ni worldwide^[29]. As shown in Figure 1G-I, the production of lithium in Australia and nickel in Indonesia demonstrated a yearly increase from 2016 to 2023, with the Democratic Republic of Congo consistently remaining the top global producer of cobalt^[29]. The centralization of these vital resources can significantly influence the raw material availability for LIBs, resulting in notable fluctuations in the cost and stability of the supply chain^[30]. Recycling spent LIBs presents a valuable opportunity, as it can overcome the geographical constraints on resource availability, thus stabilizing the supply chain and reducing manufacturing costs^[31-32].

Pyrometallurgical recovery technology

Pyrometallurgical technology encompasses a series of physical and chemical reactions at high temperatures, enabling the retention of metals in the positive electrode material in alloy form and facilitating their separation from impurities. Typically, a reducing agent is added during the reaction to collect different metals under a reducing atmosphere as either alloy or slag. Figure 3A and B illustrates the typical process flow. Pyrometallurgical processes consist of three main steps. The initial step involves crushing the entire battery and processing it in a rotary kiln for reduction roasting, segmented into low and medium-temperature phases. The aim of the low-temperature stage, at approximately 300 °C, involves evaporating the electrolyte and minimizing explosion risks. The medium-temperature phase is then increased to 700 °C, utilizing a negative electrode graphite, separator, and binder, which generates reducing gas upon heating, as reducing agents for preliminary reduction of Ni, Co, and Mn. The second step elevates the temperature above 1,200 °C for further reduction and smelting, allowing Ni, Co, and copper (Cu) to enter the alloy phase, which can be subsequently separated by magnetic separation^[39]. The third step applies hydrometallurgical technology for alloy leaching and separation, facilitating the recycling of valuable metals from the positive electrode material. The primary pyrometallurgical recycling processes for LIBs include high-temperature smelting and calcination, with the latter subdivided into direct, atmosphere-assisted, and salt-assisted calcination. Compared to high-temperature smelting, calcination operates at lower temperatures, potentially improving metal recovery rates by producing water-soluble salts. This method can reduce acid consumption and the emission of toxic gases in subsequent stages.

Figure 3. (A) Flowchart of the pyrometallurgical process for recycling spent LIBs. (B) A typical pyrometallurgical treatment process of spent LIBs^[39]. Copyright 2021, the American Chemical Society^[39].

High-temperature smelting technology for LIB recycling was first introduced in the Umicore battery recycling process^[40]. Instead of mechanically preprocessing individual batteries, this process utilizes specialized ultra-high temperature (UHT) technology, incorporating slagging agents, to directly smelt spent batteries at elevated temperatures. This process can produce an alloy containing valuable metals (Co, Ni, Fe, and Cu), which can be subsequently separated through hydrometallurgical technology [Figure 4A]^[41]. Plastics, solvents, and graphite will be incinerated, exiting in gaseous form, while Li, aluminum (Al), and Mn will be lost to the slag and not recovered [Figure 4B]^[42]. UHT technology facilitates the safe processing of large volumes of complex metallic waste. The mixture, containing batteries, coke, and slagging agents (limestone, sand), is introduced into the furnace, with optimal feedstock comprising 30-50 wt% spent batteries to yield economically viable cobalt and nickel products. Preheated air at 500 °C is then injected at the furnace base, which can be segmented into three temperature zones, namely, preheating, plastic pyrolysis, and smelting. The inaugural pilot plant was operational for several years in Hofors, Sweden. Drawing from this experience, a second-generation pilot plant was launched in Hoboken, Belgium, in 2011, with an annual capacity of 7,000 tons. The primary benefits of this process include the elimination of mechanical battery preprocessing and high recovery rates of Co, Ni, and Cu. However, this process necessitates extremely high temperatures, exceeding 1,500 °C, resulting in considerable lithium loss and leading to increased energy consumption and resource depletion.

Figure 4. (A) Flowsheet showing the Umicore/Rhodia’s VAL'EAS™ battery scrap recycling process^[41]. Copyright 2014 Elsevier^[41]. (B) Detailed illustration of the smelting process^[42]. Copyright 2022 Elsevier^[42]. (C) Recovery of spent LIBs through reduction roasting and facile acid leaching processes^[43]. Copyright 2020 Elsevier^[43].

Zhang et al. explored the recovery of valuable metals from spent LIB cathodes [Figure 4C] using a reduction roasting approach that capitalized on repurposed graphite from anode materials^[43]. This two-step method consisted of a reduction roast at 600 °C for 3 h, followed by leaching with sulfuric acid (H₂SO₄), thus eliminating the need for additional reducing agents. The findings demonstrated that this process achieved leaching rates exceeding 99% for Ni, Co, and Li, and over 97% for Mn. This innovative method not only repurposed waste graphite, enhancing energy efficiency, but also precluded the substantial generation of hydrogen gas during metal recovery. Yan et al. also suggested an improved process of carbothermal reduction for the preferential extraction of lithium, which was combined with water leaching from spent Li(Ni_xMn_yCo_1-x-y)O₂ (NMC) batteries^[44]. When the roasting temperature was below 800 °C, Co and Ni were reduced to their elemental forms, while Mn remained in oxide form, and Li was converted to Li₂CO₃. When the roasting temperature exceeded 900 °C, Li was mainly converted to Li₂O. Lithium was extracted as lithium hydroxide and lithium carbonate with an efficiency of 93% following water leaching for 30 min using a 5 mL/g liquid-to-solid ratio. The researchers also investigated the effect of impurities on the recovery rate of Li, and determined that Cu had no significant effect. However, the lithium recovery rate decreased in the presence of Al due to the formation of lithium aluminate.

The pyrometallurgical process can facilitate the rapid conversion of valuable metals into alloys at elevated temperatures, predominantly using iron (Fe) ore-based techniques for cobalt and nickel extraction. This method has a straightforward operation and efficient elimination of organic materials, binders, plastics, and other components from spent LIBs, making it appropriate for industrial-scale production. Nevertheless, this process has considerable disadvantages, including excessive energy consumption, significant material loss yielding high input costs, inadequate purity of the recovered valuable metals, and the production of hazardous substances such as dioxins and furans. These issues result in substantial secondary pollution and pose challenges in complying with contemporary environmental protection standards.

Hydrometallurgical recovery technology

Hydrometallurgy serves as a well-established and widely used technology for recycling spent LIBs. This method mainly utilizes inorganic acids, organic acids, ammonia-based reagents, and similar chemicals to leach valuable metals from pre-treated cathode active materials into a solution. Subsequently, valuable metals can be separated from this solution through processes such as precipitation and extraction. The hydrometallurgical process flow is shown in Figure 5.

Figure 5. Hydrometallurgical process flowchart of spent LIBs.

Hydrometallurgy involves two primary processes, namely, leaching and purification. The decision between common leaching methods, such as acid and alkali leaching, depends on the physical and chemical properties of the target metals. Acid leaching utilizes inorganic agents such as hydrochloric acid (HCl), H₂SO₄, nitric acid (HNO₃), and phosphoric acid (H₃PO₄), and organic agents such as citric acid, oxalic acid, and formic acid for extraction purposes, as shown in Table 1.

Leaching methods can be characterized by their high recovery efficiency, low energy consumption, and rapid reaction rates, making them highly suitable for recycling spent LIBs. However, the non-selective nature of acid leaching increases the complexity and expense of metal separation. Inorganic acid leaching typically requires high acid concentrations, resulting in significant corrosiveness and necessitating specialized equipment. Additionally, this process often mandates the extensive addition of water or alkalis for neutralization, complicating post-treatment procedures. The leaching process will emit hazardous gases, presenting risks to human health and the environment. When utilizing organic acids as leaching agents, the process remains limited to low solid-to-liquid ratios and suffers from slow leaching rates, making it impractical for large-scale industrial use. Furthermore, the prolonged degradation cycles of organic acids contribute to potential environmental pollution.

Ammonia leaching has recently attracted significant interest due to its selective extraction capabilities targeting metals such as Li, Ni, Co, Al, Fe, and Mn. Nevertheless, concerns regarding additional wastewater generation remain a challenge. The intricate binding mechanisms between the metals of value and ammonia-based leachates remain improperly understood and require more thorough investigation. Subsequent to hydrometallurgical leaching, the leachate contains valuable metal ions such as Li⁺, Ni²⁺, Co²⁺, and Mn²⁺, and impurity ions such as Fe²⁺, Al³⁺, and Cu²⁺. Conventional separation techniques such as solvent extraction, chemical precipitation, and electrodeposition have been typically employed to isolate and concentrate these valuable metals from complex mixtures.

Electrochemical pretreatment

Methods for separating current collectors from active materials include mechanical processing^[71], thermal treatment^[72], leaching^[73], and solvent treatment^[74]. Mechanical processing consists of crushing and sieving to isolate the cathode material powder, which unfortunately produces a powder material contaminated with impurities such as copper or aluminum from the current collector. Thermal treatment involves the decomposition of the binder [polyvinylidene fluoride (PVDF)] at elevated temperatures to segregate the cathode active powder, though at a significant energy cost. Leaching employs substantial quantities of acid to dissolve copper or aluminum foil, consequently producing harmful gases. Solvent treatment utilizes organic solvents to dissolve PVDF, facilitating the separation of cathode powder. However, this method consumes considerable volumes of solvents, posing a risk to human health. To enhance separation efficiency between the current collector and cathode material, researchers have explored the electrochemical technique to separate the active material and current collector.

Cao et al. developed a method for separating copper foil from graphite in the anode by using the disassembled anode of LIBs as the cathode, graphite as the anode, and employing Na₂SO₄ (1.5 g/L) solution as the electrolyte, as shown in Figure 6A^[75]. At 30 V, the complete separation of the current collector (copper foil) and graphite was achieved within 25 min. The generation of hydrogen gas on the surface of the copper foil during electrolysis created gaps, facilitating the detachment of graphite. Moreover, this method reduced the bonding force, increasing the separation efficiency. The resulting copper foil was recyclable without additional processing. Residual Li⁺ from the spent graphite entered the solution during electrolysis, and dissolved copper could be reclaimed through precipitation to a lesser extent. The purity of the regenerated graphite was approximately 95%, with energy consumption of around 4.70 kWh/kg.

Figure 6. (A) Graphite anode prior to electrolysis, accompanied by the recovered copper foil and anode active materials post-electrolysis^[75]. Copyright 2021 Elsevier^[73]. (B) Schematic of electrolytic separation. (C) Influence of supplied voltage on lithium dissolution and separation duration. (D) Impact of inter-electrode spacing on lithium dissolution and separation duration. (E) Concentration effect of sodium sulfate on lithium dissolution and separation duration. (F) Consequences of electrolysis duration on lithium recovery and separation time. (B-F) Figures reprinted with permission from Ref.^[74]. Copyright 2021 Elsevier^[75].

The previously mentioned electrochemical methods can also facilitate the separation of aluminum foil from active materials, though with changes in the electrode reactions. Chu et al. employed a neutral Na₂SO₄ solution as the electrolyte, positioning the dismantled cathode of NCM batteries as the electrolytic anode and utilizing either copper or aluminum as the electrolytic cathode, as shown in Figure 6B^[76]. Oxygen gas was also produced between the aluminum foil and active materials during electrolysis. Similarly, echoing the charging dynamics of LIBs, Li⁺ migrated from the cathode material into the solution, achieving the separation of lithium and transition metals. At an optimal electrolysis voltage of 30 V, the lithium dissolution efficiency reached 77%, with the separation of Ni, Co, and Mn in their solid forms [Figure 6C-F]. The anodic reactions are expressed as:

Electrochemical leaching

Hydrometallurgical techniques for recycling LIBs present widespread applicability and yield high-purity products, making these techniques a favored method. These techniques encompass dissolving valuable metals from depleted batteries into acidic or alkaline solutions, followed by purification processes such as extraction and precipitation to eliminate any impurities, culminating in the production of high-purity products. The leaching step has shown to be especially pivotal, significantly influencing the recovery efficiency of valuable metals. However, to improve leaching efficiency, substantial amounts of leaching solution are typically used, leading to the production of large volumes of wastewater. By integrating electrochemical methods for leaching positive electrode materials, these challenges can be mitigated, with electrochemical reactions relying on an external electric current.

Lei et al. devised a pioneering process for recycling LIBs by integrating reduction roasting with electrochemical leaching, as shown in Figure 7A-D^[77]. The process involved reduction roasting of the spent LiNi_1/3Co_1/3Mn_1/3O₂ material at 1,600 °C for 120 min, which released O₂ and disrupted its layered structure, resulting in a mixture of Li₂CO₃, NiO, Co₃O₄, MnO₂, and Mn₂O₃. This disruption of the layered structure enhanced valuable metal leaching. The roasted material was subsequently encased in polypropylene bags and attached to a graphite cathode, with a titanium (Ti) anode. The setup utilized G1204 [an anion exchange membrane (AEM)] as the separator and 1.5 M H₂SO₄ as the electrolyte. Subsequently, the roasted product underwent further reduction through electrochemical methods at room temperature, with a current of 0.8 A applied for 150 min. The leaching rates of Li, Co, Ni, and Mn were recorded as 100, 90.53, 90.59, and 66.40%, respectively, as shown in Figure 7E. Compared to identical leaching conditions involving direct acid leaching and the electrochemical leaching of thermally treated roasted spent cathode materials, this method proved to substantially minimize the utilization of chemical reagents such as reducing agents and acid, while also achieving higher metal recovery rates and yielding greater economic benefits.

Figure 7. (A) Flowsheet illustrating the recovery of spent LIBs through thermal reduction and electrochemical leaching processes. (B) Mechanism of the reaction for the direct electrochemical leaching of cathode materials. (C) Mechanism of the reaction for thermal reduction process followed by acid leaching. (D) Mechanism of the reaction involving thermal reduction process followed by electrochemical leaching (green represents reactants, with red indicating substances introduced during the leaching processes). (E) Comparative evaluation of metal leaching effectiveness among the three methods: (1) direct electrochemical leaching; (2) thermal reduction process combined with acid leaching; (3) thermal reduction process coupled with electrochemical leaching. (A-E) Figures reprinted with permission from Ref.^[77]. Copyright 2021 American Chemical Society^[77]. (F) Schematic representation of the electrochemically assisted leaching of metals from MOFC, where the leachate, carbon/graphite materials, and electrowon copper can be separately recovered. (G) Leaching efficiency of metals and specific energy consumption at various tested pulp densities. (F-G) Figures reprinted with permission from Ref.^[78]. Copyright 2020 Elsevier^[78].

Electrochemical leaching can also be used to process mixtures of different spent LIB materials. Diaz et al. developed a method using electrochemical techniques to extract valuable metals from mixed crushed LIB metal oxide filter cakes (MOFC), composed of LiCoO₂, LiMn_xCo_yO₂, LiNi_xMn_yCo_zO₂, LiCu_xMnO₂, LiNi_xCo_yAl_zO₂, current collectors (aluminum and copper foil), and graphite^[78]. The electrolytic cell comprised an anodic chamber and a cathodic chamber separated by a bipolar membrane (BPM), as depicted in Figure 7F. The anode comprised nickel foam, immersed in a 1 M KOH solution, while the cathode was constructed from stainless steel mesh, with the cathode chamber electrolyte ranging from 0.5-2 M H₂SO₄, enhanced by 10 mM FeSO₄ and the addition of MOFC. During the electrolysis process, Fe²⁺ served as the reducing agent, facilitating the reduction of Co³⁺ ions, and improving their leaching rate. With a slurry concentration of 65 g/L and a controlled low acid concentration, voltage, current, and electrolyte pH, the leaching rates of Li, Ni, Co and Mn exceeded 96% within 8 h, as shown in Figure 7G. The process generated electrolytic copper and graphite carbon as by-products. Compared to conventional acid leaching, this methodology noticeably decreased acid consumption and produced a lower pH leaching solution, significantly reducing wastewater treatment costs.

Electrochemical separation

After leaching the positive electrode materials, the efficient separation of diverse valuable metal elements from the intricate leach solution serves as a vital process step in the recycling of spent LIBs. Due to the high ion selectivity of electrochemical techniques, multiple metal elements in the solution can be effectively separated, with the resulting metal products post-separation exhibiting greater purity, making them suitable for direct use in the regenerative synthesis of cathode materials.

Electrochemical deposition

Electrochemical deposition involves the process in which metal ions of value within the leaching solution obtain electrons under an applied electric field, subsequently depositing on the cathode to form a coating. The order in which electrochemical deposition of varied metal ions occurs will be dictated by the properties of the metal ions, electrolyte composition, pH, temperature, and current density. Consequently, by modifying these conditions, specific metal ions can be selectively reduced, facilitating the separation of valuable metals. Compared to alternative separation techniques, electrochemical deposition offers benefits such as straightforward raw materials, reduced costs, a concise process, high product purity, and excellent performance.

Prabaharan et al. employed electrochemical techniques to extract Co and Mn from spent LIBs [Figure 8A-C]^[79]. At a concentration of 2 M H₂SO₄, with a current density of 400 A/cm², and after an electrolysis period of 3 h, the leaching rate of Co and Mn reached close to 100%. Additionally, Cu from the batteries was concurrently deposited at the cathode during the leaching, removing the need for further Cu separation and recovery processes. Electrochemical deposition facilitated the recovery of Mn and Co from the leaching solution. In this process, Co was deposited at the cathode through electrochemical reduction, while Mn was oxidized to form MnO₂ at the anode. The respective reactions during electrochemical deposition are expressed as follows.

Figure 8. (A) The method of electrodeposition for the separation of cobalt from spent LIBs. Characteristics of cobalt metal deposition at (B) a pH of 2.5 and (C) a pH of 4.5. (A-C) Figures reprinted with permission from Ref.^[79]. Copyright 2017 Elsevier^[79] (D) Schematic illustration of the electrodeposition apparatus^[83]. Copyright 2019 Royal Society of Chemistry^[83].

Cathode:

(3) $$ \begin{equation} \begin{aligned} \mathrm{Co}^{2+}+2 \mathrm{e}^{-} \rightarrow \mathrm{Co} \end{aligned} \end{equation} $$

(4) $$ \begin{equation} \begin{aligned} 2 \mathrm{H}^{+}+2 \mathrm{e}^{-} \rightarrow \mathrm{H}_{2} \end{aligned} \end{equation} $$

Anode:

(5) $$ \begin{equation} \begin{aligned} 2 \mathrm{H}_{2} \mathrm{O} \rightarrow \mathrm{O}_{2}+4 \mathrm{H}^{+}+2 \mathrm{e}^{-} \end{aligned} \end{equation} $$

(6) $$ \begin{equation} \begin{aligned} \mathrm{Mn}^{2+}+2 \mathrm{H}_{2} \mathrm{O} \rightarrow \mathrm{MnO}_{2}+4 \mathrm{H}^{+}+2 \mathrm{e}^{-} \end{aligned} \end{equation} $$

The pH of the electrolyte was found to significantly influence the quality of deposited cobalt metal. As the pH increased from 1 to 4.5, the current efficiency of Co deposition improved from 33.2% to 97.1%, although the purity of the product declined to 85.3%. When the pH was maintained between 2 and 2.5, the deposited cobalt metal demonstrated enhanced brightness, and the impurity content remained below 0.8%. Optimal conditions for complete deposition of Co and Mn included an electrolyte pH range of 2 to 2.5, a temperature of 90 °C, and a current density of 200 A/m², while the recovery rates for Co, Cu, and Mn exceeded 96%, 97%, and 99%, respectively, leading to product purities of 99.2% for cobalt metal, 99.5% for copper metal, and 96% for MnO₂.

Freitas et al. employed HCl and H₂O₂ to leach spent LiCoO₂ powder materials, and subsequently used the leaching solution as the electrolyte for cobalt recovery using the electrochemical method^[80]. As the pH increased, the reductive potency of hydrogen ions diminished, enhancing the current efficiency of cobalt deposition. The maximum current efficiency, amounting to 96.90%, was achieved at a pH of 5.40, while cobalt deposition nucleation at the cathode was continuous. By contrast, at a pH of 2.70, the nucleation shift to an instantaneous mode affected the morphology of the cobalt deposits^[81-82].

In addition to metals such as Ni, Co, Mn, and Li, the leachate from spent LIBs cathode materials contains impurities including Fe, Cu, and Al. Typically, Al can be precipitated by adjusting the pH of the leachate. However, due to the similar precipitation pH values of Cu²⁺, Co²⁺, and Ni²⁺, pH adjustment alone remains insufficient for Cu removal. Freitas et al. explored the impact of Cu presence in the leachate on Co electrodeposition^[80]. The leaching of spent cathode materials containing Co and Cu was conducted using H₂SO₄ and H₂O₂, with resultant leachate serving as the electrolyte for electrodeposition, utilizing Pt as the anode. Co and Cu were electrodeposited onto platinum, glassy carbon, or aluminum cathodes. At a pH of 5.4, the Co/Cu multilayers that formed on the cathodes showed increased porosity, primarily due to the continuous nucleation mechanism. At a pH of 2.7, the Co/Cu multilayers exhibited reduced porosity, reflecting a shift from continuous to instantaneous nucleation. Therefore, the separation of Cu and Co through electrodeposition demonstrated difficulties. Because Cu²⁺ has a higher standard electrode potential for reduction compared to hydrogen evolution potential, copper ions(II) can be selectively separated, unlike Ni²⁺, Co²⁺, and Mn²⁺. Peng et al. successfully employed an electrochemical method to selectively remove copper ions (II) from the leachate, thus achieving the preparation of high-purity cobalt [Figure 8D]^[83].

If a notable difference in the reduction potentials of valuable metal ions in the leach solution is present, electrodeposition can separate and recover different valuable metals, achieving high metal purity. However, if the reduction potentials of valuable metal ions are similar (i.e., Co and Ni), electrodeposition will occur simultaneously for metals with close reduction potentials, leading to decreased metal purity. Therefore, in the recycling process of spent LIB cathode materials, Ni and Co will be initially separated from the leaching solution via extraction, followed by electrodeposition to independently prepare elemental Ni and Co on the cathode.

Electrodialysis

Electrodialysis involves the movement of various metal ions at distinct rates across a semi-permeable membrane, driven by an external electric field and influenced by their charge and size disparities. This method facilitates the selective migration of ions, enabling the efficient separation of valuable metals. In the recycling of spent LIBs cathode materials, electrodialysis has been extensively employed for the separation and recovery of critical metals such as Ni, Co, and Li, offering vital technical support for the recycling processes of LIBs.

In the leaching solution of spent LIB cathode materials, lithium ions coexist with other transition metal ions. To efficiently separate lithium from transition metal ions, Iizuka et al. introduced a method utilizing BPM electrodialysis in conjunction with metal ion complexation, which specifically enhanced the separation of Li⁺ and Co²⁺ in the leaching solution, with ethylenediaminetetraacetic acid (EDTA) serving as the complexing agent^[84,85]. At pH levels above 4, the formation of anionic groups CoY^2- resulted from the complexation of EDTA with Co²⁺, leaving Li⁺ as the cations in the solution. Thus, exploiting the charge disparity, Co²⁺ and Li⁺ separation could be realized through electrodialysis. Figure 9A presents the configuration of the electrodialysis system, which included three electrolysis cells separated by BPMs, AEMs, and cation exchange membranes (CEMs). The cobalt recovery cell was linked to the anode (stainless steel), while the lithium recovery cell was connected to the cathode (Ti/Pt). The leaching solution and EDTA were introduced into the feed cell, where under the influence of an electric field, CoY^2- migrated to the cobalt recovery cell through the AEM, and Li⁺ moved to the lithium recovery cell through the CEM, thus separating Li⁺ from Co²⁺. Figure 9B-G illustrates the variations in Li⁺ and Co²⁺ concentrations across the three cells over time, demonstrating how the Li and Co concentrations progressively increased in their respective recovery cells. The selectivity of Li and Co achieved a notable value of 99%.

Figure 9. (A) Method principle for Co and Li separation based on electrodialysis. Variations over time in Li quantities within the cells; impact of initial pH levels: (B) feed cell, (C) Li recovery cell, and (D) Co recovery cell. Variations over time in Co quantities within the cells; impact of initial pH levels: (E) feed cell, (F) Li recovery cell, and (G) Co recovery cell. (A-G) Figures reprinted with permission from Ref.^[85]. Copyright 2023 Elsevier^[85].

In the electrodialysis process, the ion exchange membrane could easily capture and stabilize cobalt and lithium ions, leading to lower recovery rates for these metals. To address this issue, a semi-batch experiment was conducted. Initially, the feed solution was added to the feed cell and operated for 30 min, after which it was replenished with a fresh solution. This replacement cycle was repeated every 30 min for a total duration of 150 min. After each 30 min of operation, the metal concentration in the feed cell significantly decreased, approaching zero over time [Figure 10A]. Despite the increase in cycles, the metal ion removal rate from the feed solution remained consistent. Conversely, the concentration of lithium ions in the lithium recovery cell and cobalt in the cobalt recovery cell linearly increased with operation time, as shown in Figure 10B and C. These findings suggested that this method not only effectively separated lithium and cobalt ions from the feed solution but also concentrated the metals through semi-batch operation.

Figure 10. Changes with time in amounts of Li and Co in cells. (A) feed cell, (B) Li recovery cell, and (C) Co recovery cell^[84]. Copyright 2013 Elsevier^[84].

Chan et al. pioneered a method employing electrodialysis technology for the separation of Li, Co, Mn, and Ni from the LiNi_0.33Mn_0.33Co_0.33O₂ (NMC111) cathode material of spent LIBs^[86]. This process encompassed the EDTA-mediated complexation of Ni and Co, followed by three-stage electrodialysis for the separation of Li, Co, Mn, and Ni, the subsequent decomplexation of Ni and Co with EDTA, and a purification step to recover the final products. As shown in Figure 11A, the electrodialysis apparatus comprised Pt/Ti anodes, 316 stainless steel cathodes, and five chambers spatially arranged between two CEMs and two AEMs.

Figure 11. (A) Schematic diagram showing the electrodialysis experimental setup for the separation of Li, Ni, Mn, and Co from NMC111. Principles of the three-stage electrodialysis process for separating Li, Ni, Mn, and Co from NMC111: (B) separation of Ni in Stage 1; (C) Separation of Co in Stage 2; (D) separation of Li from Mn in Stage 3. (A-D) Figures reprinted with permission from Ref.^[86]. Copyright 2012 Elsevier^[86].

The separation of Li, Co, Mn, and Ni from NMC111 via electrodialysis consisted of a three-stage process: Ni in Stage 1; Co in Stage 2; and Li and Mn in Stage 3. During Stage 1, as shown in Figure 11B, the leachate comprised Li⁺, Mn²⁺, Co²⁺, Ni²⁺, and SO₄^2-. The introduction of EDTA selectively complexed with Ni to form anionic Ni-EDTA species (NiHY^- and NiY^2-). Driven by an electric field, NiHY^-, NiY^2-, and SO₄^2- ions traversed the AEM into the Ni-EDTA complex chamber, where they reacted with H⁺ migrating from the electrode rinsing chamber to form NiH₂Y and H₂SO₄. Concurrently, Li⁺, Mn²⁺, Co²⁺, and Na⁺ ions migrated through the CEM to the metal chamber, reacting with SO₄^2- from the electrode rinsing chamber to form Li₂SO₄, MnSO₄, CoSO₄, and Na₂SO₄. These salts served as the input for Stage 2. As a result, Ni was effectively separated from Li, Mn, and Co in the first stage.

In Stage 2 [Figure 11C], the solution originating from the metal chamber of Stage 1 contained Li⁺, Mn²⁺, Co²⁺, Na⁺, and SO₄^2- ions. The addition of EDTA preferentially formed Co-EDTA complex anions (CoHY^- and CoY^2-). Similar to Stage 1, CoHY^-, CoY^2-, and SO₄^2- ions were transported through the AEM to the Co-EDTA complex chamber, where they combined with H⁺ ions migrating from the electrode rinsing chamber to form CoH₂Y and H₂SO₄. Concurrently, Li⁺, Mn²⁺, and Na⁺ were transported through CEM to the metal chamber, and combined with SO₄^2- ions migrating from the electrode rinsing chamber to form Li₂SO₄, MnSO₄, and Na₂SO₄. These products served as the feed solution in Stage 3, achieving the separation of Co from Li and Mn in Stage 2.

In Stage 3 [Figure 11D], the solution from the metal chamber of Stage 2 contained Li⁺, Mn²⁺, Na⁺, and SO₄^2-. Since Mn²⁺ ions did not completely form Mn-EDTA complex anions, changing the membrane configuration in the electrodialysis cell with monovalent CEMs enhanced selectivity for monovalent over divalent cations. Consequently, SO₄^2- ions passed through AEM to the H₂SO₄ compartment, joining H⁺ ions migrating from the electrode rinse compartment to produce H₂SO₄. Simultaneously, Li⁺ and Na⁺ ions were transported through the monovalent CEM to the Li₂SO₄ compartment, where they merged with SO₄^2- ions migrating from the electrode rinsing compartment, forming Li₂SO₄ and Na₂SO₄. Mn²⁺ ions were hindered by monovalent CEM and retained in the feed compartment. Thus, Li was separated from Mn in Stage 3.

During the electrodialysis process, a 99.3% separation of Ni was achieved in the first stage when the molar ratio of EDTA to Ni was 1.1. In the second stage, with a molar ratio of EDTA to Co of 1.2, 87.3% of Co was separated. In the third stage, monovalent CEM facilitated the sequential separation of approximately 99% of Li and Mn. Post-experiment, Ni and Co were desorbed from EDTA at a pH below 0.5, allowing for the recovery of EDTA solids. Selective precipitation was then utilized to recover all four metals with purities exceeding 99%.

Due to the intricate composition of spent LIB cathode materials, which resulted in a diverse array of metal cations in the leachate, simple electrodialysis techniques were found to be inadequate for separation. The addition of chelating agents to form complexes with specific metal ions, converting them into anions for separation from other cations, offers a viable approach. Nevertheless, this method remains constrained by protracted separation times and the increased consumption of reagents.

Electrochemical lithium stripping

Electrochemical lithium stripping, grounded in the charge and discharge mechanisms of LIBs, employs electrical current instead of chemical reagents to drive reactions, thus facilitating selective lithium-ion removal from positive electrode materials. This approach can significantly diminish reagent consumption and circumvent lithium loss, thereby enhancing lithium recovery rates.

To recover the cathode material LiMn₂O₄, a slurry electrolysis method was proposed for the recuperation of Li and Mn [Figure 12A]^[87]. In the electrolytic cell, a filter cloth (< 50 μm) was employed to segregate the anode and cathode chamber. The electrolyte comprised a mixed solution of H₂SO₄ and MnSO₄, with graphite (cathode) and ruthenium-plated titanium electrode (anode), which was operated at 90 °C. Spent LiMn₂O₄ was introduced into the cathode chamber and dispersed. LiMn₂O₄ powder dissolved in the cathode chamber during electrolysis, releasing Mn²⁺ and Li⁺ ions [Figure 12B]. The influence of H₂SO₄ and MnSO₄ concentrations, along with slurry and current densities, on the rates at which lithium and manganese leached is shown in Figure 12C-J. Optimal conditions included a slurry concentration of 75 g/L, a current density of 40 mA/cm², a reaction time of 20 h, and a 1 M solution of H₂SO₄ and MnSO₄. Under these conditions, the leaching rates of lithium and manganese reached 99% and 92%, respectively, with cathode and anode current efficiencies of 77% and 62%. After electrolysis, Mn²⁺ was oxidized to Mn⁴⁺ and precipitated as MnO₂ in the anode chamber, while Li⁺ in the cathode reacted with Na₂CO₃ to produce Li₂CO₃ [Figure 12B]. The purities of the recovered MnO₂ and Li₂CO₃ were 92.33% and 99.59%, respectively, allowing for their direct reuse as raw materials to synthesize LiMn₂O₄ cathode materials. A preliminary economic analysis of recovering 1 ton of LiMn₂O₄ using this method was conducted on the recovery of lithium, indicating total revenue from the recovered products of $2,353.59, with reagent and energy costs of $442.43 and $392.00, respectively, demonstrating the economic viability of this process.

Figure 12. (A) Schematic representation of cathode material LiMn₂O₄ electrolysis. (B) Mechanism diagram of the cathode material LiMn₂O₄ electrolysis. Impact of electric current density on (C) the leaching efficiency, and (D) electrode current efficiency and cell voltage. Influence of H₂SO₄ concentration on (E) the leaching efficiency of Mn and (F) current efficiency. Impact of density on (G) leaching efficiency and (H) current efficiency. Influence of MnSO₄ concentration on the (I) leaching efficiency and (J) current efficiency. (A-J) Figures reprinted with permission from Ref.^[87]. Copyright 2019 American Chemical Society^[87].

Li et al.^[88] introduced a method for recycling LiFePO₄ cathode materials that integrated the charging dynamics of LiFePO₄ batteries with a slurry electrolysis strategy, as illustrated in Figure 13A. This method employed an electrolytic cell divided into anode and cathode chambers by an AEM, with sodium chloride serving as the electrolyte, and Ru-coated Ti electrodes facilitating electrolysis. Within the anode chamber, used LiFePO₄ underwent oxidation under a positive potential to yield FePO₄, which liberated lithium ions into the electrolyte. Due to the insolubility of FePO₄ in water, it could be easily isolated through filtration. Concurrently, in the cathode chamber, a hydrogen evolution reaction produced sodium hydroxide as a secondary product. The impact of variables such as the leaching time, electric current density, sodium chloride concentration, and LiFePO₄ concentration on lithium extraction is detailed in Figure 13B-E. This method successfully extracted over 98% of lithium into the electrolyte and recovered more than 96% of iron in the form of FePO₄, consuming 2,613.33 k/Wh of energy. Electron-driven reactions during slurry electrolysis negated the necessity for chemical reagents, such as oxidizing agents or acids. Filtration of the anode chamber yielded FePO₄, and the addition of sodium carbonate (Na₂CO₃) to the filtrate, thus, facilitated the recovery of lithium ions as LiCO₃.

Figure 13. (A) Diagram illustrating the electrolysis process. Impact of (B) leaching duration, (C) electric current density, (D) sodium chloride concentration, and (E) LiFePO₄ concentration on the leaching efficiency of lithium^[87]. (A-E) Figures reprinted with permission from Ref.^[88]. Copyright 2019 Elsevier^[88].

Based on this method, a process route was proposed for the recycling of LiFePO₄ cathode materials through slurry electrolysis, as illustrated in Figure 14. The spent LiFePO₄ underwent oxidation and delithiation in the anode chamber due to electron activity. Concurrently, in the cathode chamber, the hydrogen evolution reaction occurred, producing NaOH. A minor amount of Cl₂ released during the slurry electrolysis process could be absorbed by the NaOH solution in the cathode chamber. More than 99% of Li could be leached into the electrolyte, resulting in an anolyte containing 7.55 g/L of Li. The high efficiency and selectivity of this process could enable the recovery of the most valuable elements without the need for prior separation and purification. The purity of the recovered FePO₄ reached 99.15%, allowing its use as a raw material for synthesizing LiFePO₄. The generated NaOH in the cathode electrolyte aided in removing Fe from the anode electrolyte. After filtration, Li₂CO₃ precipitated through evaporation and the addition of Na₂CO₃, reaching a purity of 99.69%. The resultant mother liquor, with a Li content of 2.53 g/L and Na content of 35.72 g/L, could be recycled as a cathode electrolyte. An economic analysis for recycling 1 ton of spent LiFePO₄ material was performed, and the costs for reagents and energy were identified as $71.45 and $395.15, respectively, while the revenue from the sale of Li₂CO₃, FePO₄, and NaOH products was approximately $2,846.01, demonstrating the economic viability of this recycling technology.

Figure 14. Flowsheet of the slurry electrolysis for recycling spent LiFePO₄^[88]. Copyright 2019 Elsevier^[88].

Employing NaCl as the electrolyte can decrease processing costs, while using lithium salts as electrolytes in both the anode and cathode can remove impurity elements and enhance conductivity. Li et al. utilized Li₂SO₄ and LiOH as the anode and cathode electrolytes, respectively, employing platinized titanium electrodes for electrolysis to achieve higher current efficiency^[89]. The study added LiFePO₄ to the anode electrolyte, which underwent continuous leaching for Li⁺ and FePO₄. Throughout the electrolysis process, Li⁺ moved from the anode compartment to the cathode compartment through a CEM, driven by an electric field. Subsequent to electrolysis, LiOH·H₂O was produced through evaporation and crystallization. Under the optimal conditions of a current density of 60 mA/cm², an Li₂SO₄ concentration of 50 g/L, an LiOH concentration of 20 g/L, and an LiFePO₄ concentration of 100 g/L, the leaching rate of lithium surpassed 96%. The process achieved a current efficiency of 85.81%, with an energy consumption of 2,775.07 k/Wh.

Electrochemical regeneration

One of the primary failure mechanisms of LIB cathodes is the ongoing loss and consumption of Li⁺ during cycling. Consequently, cathode material repair can be achieved by directly replenishing Li. Electrochemical regeneration processes often utilize conventional three-electrode or two-electrode systems, with the degraded cathode material acting as the working electrode for Li⁺ insertion and structural restoration. The application of a cathodic potential to the depleted cathode serves as the principal driving force, effectively initiating the regeneration process and significantly reducing the activation energy required for Li⁺ migration to lithium vacancies. This can lead to lithiation and structural repair of the material. After replenishing Li, the cathode material generally requires annealing in a specific atmosphere to completely restore its crystal structure, thus tightly organizing Li atoms within the structure.

Zhang et al. presented a technique for rejuvenating LiCoO₂ cathode materials by electrochemically driving Li⁺ insertion into the depleted Li_xCoO₂^[90]. As shown in Figure 15, the electrochemical apparatus included the cathode (spent Li_xCoO₂), anode (platinum foil), and reference electrode [Ag/AgCl saturated with KCl, 0.199 V vs. Normal Hydrogen Electrode (NHE)], with the electrolyte consisting of Li₂SO₄ solution at a specific concentration. The regeneration mechanism, as shown in Figure 15A, involved Li⁺ migration toward vacancies within the Li_xCoO₂ structure, prompted by an electric current. Once a threshold voltage was achieved, Li⁺ insertion concluded. Subsequent annealing at 700 °C expelled crystalline water, revitalizing the LiCoO₂ lattice structure. The completion time for full reduction decreased with increasing Li⁺ concentration (0.1 to 1 M) and current density (0.12 to 0.42 mA/cm²). The surface morphology and crystalline configuration of the regenerated material mirrored those of commercial LiCoO₂, achieving charge capacities of up to 136 mAh/g. Notably, the presence of impurities in the spent material significantly impaired the electrochemical properties of the regenerated LiCoO₂. Thus, purifying the regenerated cathode material of extraneous elements could pose a significant challenge for subsequent investigations.

Figure 15. (A) Method for electrochemically inserting Li⁺ into Li_xCoO₂ electrodes from the spent LIBs, thus generating new LiCoO₂ materials. (B) The proposed mechanism for directly regenerating LiCoO₂ materials through the electrochemical insertion of Li⁺ ion into the Li_xCoO₂ electrode. Cycling performance of the Li/LiCoO₂ electrode from (C) the regenerated LiCoO₂ electrode and (D) commercial LiCoO₂ electrode at C/5 current rate. (A-D) Figures reprinted with permission from Ref.^[89]. Copyright 2020 American Chemical Society^[90].

Yang et al. recycled spent Li_xCoO₂ from dismantled LIBs, employing it as the cathode^[91]. LiCoO₂ was used as the anode with the electrolyte of Li₂SO₄, facilitating the integration of Li⁺ into Li_xCoO₂ through electrolysis, as shown in Figure 16A. After the working electrode voltage reached the cutoff point, Li_xCoO₂ fully transitioned to LiCoO₂. The reduced material underwent sintering at 700 °C for 6 h, yielding the regenerated product. The process then involved utilizing the LiCoO₂ electrode to replenish the depleted Li⁺ in the electrolyte. As shown in Figure 16B, compared to conventional pyrometallurgy and hydrometallurgy, this method drastically reduced energy consumption to 6 mJ/kg and presented significant prospects for advancement.

Figure 16. (A) The electrolytic cell used in the re-lithiation process under higher current density (0.4 mA/cm²). (B) Comparative analysis of energy consumption between pyrometallurgy, hydrometallurgy, and direct recycling approaches for LiCoO₂. Figures reprinted with permission from Ref.^[91]. Copyright 2020 Wiley^[91].

In addition to directly replenishing the failed cathode material through electrochemical methods, another technique involves synthesizing the cathode material directly through electrochemical processes post-leaching, minimizing subsequent elemental separation steps and facilitating the direct reuse of the cathode material.

Wang et al. developed a method based on suspension electrolysis technology to directly leach LiCoO₂ in the anode region while re-synthesizing LiCoO₂ in the cathode region [Figure 17A]^[92]. This technique employed spent cathode material powders, derived from the crushed and mixed cathodes of various mobile phone batteries, containing elemental compositions of Co (46.50 wt%), Li (5.85 wt%), Mn (3.22 wt%), Ni (2.69 wt%), Cu (0.07 wt%), Al (11.35 wt%), and Fe (0.17 wt%). In the electrolytic apparatus, an acrylic fabric separated the cell into the anode and cathode compartments, utilizing platinum-coated titanium electrodes. The electrolyte was composed of NH₄HCO₃, (NH₄)₂SO₃, and NaF, where NH₄HCO₃ acted as a coordinating agent, (NH₄)₂SO₃ as a reducing agent, and NaF improved the conductivity of the electrolyte. The influence of NH₄HCO₃ and (NH₄)₂SO₃ concentrations, current, temperature, and holding time on the leaching rate and regeneration rate are shown in Figure 17B-G. The optimal conditions were 2.5 M NH₄HCO₃, 0.7 M/L (NH₄)₂SO₃, solid-to-liquid ratio of 4 g/L, temperature of 60 °C, current of 3.5 A, and holding time of 90 min. Under these conditions, Co and Li exhibited leaching rates of 55.4 and 71.8 wt%, respectively, with an LiCoO₂ recovery rate of 51.5 wt%. The primary reactions are detailed in:

Figure 17. (A) Flow diagram depicting the recovery process of LiCoO₂ from discarded mobile phones. Impact of the (B) NH₄HCO₃ concentration, (C) (NH₄)₂SO₃ concentration, (D) current, (E) temperature, (F) solid-to-liquid ratio, and (G) holding time on the leaching rates of lithium and cobalt, and the regeneration rate of LiCoO₂. (A-G) Figures reprinted with permission from Ref.^[84]. Copyright 2023 Elsevier^[84].

Based on the re-synthesis mechanism of LiCoO₂, Li et al. suggested regenerating LiCoO₂ through controlled electrochemical deposition reactions^[93]. Following the leaching of spent LiCoO₂ with HNO₃, a solution of 4 M LiOH was introduced, and the pH of the leaching solution was adjusted to 11, to serve as the electrolyte for the electrolysis process. The electrolytic cell employed a Pt plate cathode and Ni plate anode, as shown in Figure 18. During electrolysis, a current density of 1.0 mA/cm² was applied at 100 °C for 20 h. Co (OH)₂ suspended in the 4 M LiOH solution was converted into HCoO₂^- and migrated towards the anode due to the electric field. After reaching saturation near the anode, HCoO₂^- ions deposited on the nickel plate as Co (OH)₂. Concurrently, electron transfer from Co (OH)₂ to the electrode occurred, changing the oxidation state of Co to +3 and forming a CoOOH film. In a highly concentrated LiOH solution, a Li⁺/H⁺ exchange occurred, leading to the formation of LiCoO₂. The primary reactions during electrolysis are expressed as:

Figure 18. (A) Illustration of the re-synthesis mechanisms of LiCoO₂. (B) Schematic showing the recycling instrument using electrochemical deposition. (C) Charge-discharge characteristics of the spent and regenerated LiCoO₂ at a rate of 0.1 C. (D) Cyclic voltammogram of the LiCoO₂ material at a scan rate of 5 mV/s. (E) Cycling performance of regenerated LiCoO₂ at a rate of 0.1 C. (A-E) Figures reprinted with permission from Ref.^[93]. Copyright 2011 Elsevier^[93].

The initial capacities for charging and discharging of the regeneration LiCoO₂ were 130.8 and 127.2 mAh/g, respectively, in contrast to the capacity of commercial LCO of 140 mAh/g, as illustrated in Figure 18C. After 30 cycles, the capacity degradation remained below 4%, with a charging efficiency of 99.1%, as shown in Figure 18E. This method directly regenerated the positive electrode material from the filtrate of spent LiCoO₂. However, the electrochemical performance of the synthesized LiCoO₂ did not meet that of its commercial counterpart. The recycling process necessitated high temperatures and prolonged reaction times, leading to increased energy consumption. Moreover, the use of platinum electrodes may be impractical for large-scale applications.

Materials synthesized through electrochemical recycling can be applied in electrochemical sensors, supercapacitors, pseudocapacitors, and catalysts. Gonçalves et al. utilized AISI 430 stainless steel as the substrate and the leachate of spent LIB cathode materials as the electrolyte to prepare cobalt-containing materials through potentiostatic electrolysis^[94]. These materials were used as non-enzymatic electrochemical sensors for glucose detection. The sensor demonstrated a sensitivity of 70.2 μA/(mmol·cm²) within a linear range of 1-10 mmol/L, providing a novel approach for developing sensors from recycled materials. Barbieri et al. investigated the recycling of cobalt from spent LIB cathodes^[95]. The materials were recovered in the form of β-Co(OH)₂ and Co₃O₄ through thermal treatment, and both chemical and electrochemical precipitation methods were employed to obtain the starting materials. The final products were used to fabricate electrodes comprising Co₃O₄, carbon black, and epoxy resin. These electrodes demonstrated high reversibility, charging efficiency, and specific capacitance (13.0 F/g at 1.0 mV/s), indicating their potential use as pseudocapacitors. Aboelazm et al. utilized electrodeposition under a magnetic field to form cobalt oxides from spent LIBs^[96]. The resulting nanostructures exhibited a large electroactive surface area, which contributed to their excellent electrochemical performance and capacitive behavior. Electrochemical tests demonstrated a capacitive retention of 96% after 5,000 cycles and a charge storage capacity of 1,273 F/g at a current density of 1 A/g. Falqueto et al. utilized recycled cobalt from spent LIBs as a catalyst for azo dye degradation. Metallic cobalt was obtained through electrodeposition (-1.1 V, 15 C/cm², pH 2.7) from a solution leached from the spent LIB cathodes^[97]. When combined with 0.27 mol/L H₂O₂, the material achieved a decolorization efficiency of 85% for azo dye after 210 min under UV radiation, indicating promising catalytic properties.