Stepwise Recovery of Valuable Metals from Spent Lithium-Ion Batteries Through In Situ Thermal Reduction and Selective Leaching

Abstract

The sustainable recycling of valuable metals from spent lithium-ion batteries (LIBs) is imperative for closing the resource loop. This study presents an integrated strategy for the stepwise recovery of metals from spent cathode sheets by in situ thermal reduction and selective leaching. The in situ thermal reduction converted the cathode material into a mixture of Li₂CO₃, LiAlO₂, Ni, Co, NiO, and CoO while simultaneously liberating the cathode materials from the Al current collector through binder removal. A combined process of water leaching, wet sieving, and filtration successfully achieved the separation and enrichment of Li-rich aqueous solution (near 60% Li), Al-rich coarse fraction (over 87% Al), and fine powder enriched with transition metals (over 90% of Ni, Co, and Mn). The pyrolysis gases released from binder decomposition were the key driver for forming Li₂CO₃, whereas the concurrent generation of LiF and LiAlO₂ limited direct water leaching efficiency. An alkaline leaching step was therefore introduced to co-extract Al and the associated Li from LiAlO₂, followed by an acid leaching step that recovered over 96% of the transition metals from the treated residue without external reductants. Complete mass balance analysis shows that the integrated process achieved overall recoveries of 91.86% for Li, 91.93% for Ni, 92.23% for Co, and 92.61% for Mn from all the combined leachate streams. Consequently, this work provides a reagent-saving, stepwise hydrometallurgical process for the comprehensive recycling of valuable metals from spent LIBs.

1. Introduction

The widespread adoption of lithium-ion batteries (LIBs) in electric vehicles, consumer electronics, and energy storage systems has precipitated a global wave of battery retirement after 3–10 years of service [1,2]. This presents both a critical opportunity for resource recovery and a pressing environmental challenge. It is estimated that the mass of spent electric vehicle battery packs may reach up to 8 million tons by 2040 [3]. These end-of-life units represent a valuable secondary source of critical metals such as Li, Ni, Co, and Mn, yet also contain hazardous components including heavy metals and organic compounds [4,5,6]. Thus, developing efficient recycling strategies is crucial for advancing a circular economy.

Valuable metals in spent LIBs are predominantly concentrated in the cathode sheet, making it the primary target for recovery. However, efficient recycling faces two major challenges: the strong adhesion of cathode particles to the Al current collector by organic binders like PVDF, which complicates their separation [7,8], and the presence of key transition metals (e.g., Ni³⁺, Co³⁺, and Mn⁴⁺) in high oxidation states within the cathode lattice, which hinders their direct dissolution in subsequent hydrometallurgical processing [9,10]. Therefore, an effective initial treatment must simultaneously achieve material liberation from the current collector and reduction of these high-valence metals.

In situ thermal reduction under an oxygen-free atmosphere offers a promising solution to this dual requirement [11,12]. This process utilizes a thermal field to decompose the organic binder, thereby liberating the cathode material [13], while using intrinsic reductants (e.g., PVDF [14], graphite [15], carbon black [16], Al foils [17], and separator [18]) to collapse the cathode structure and reduce the high-valence metals. During pyrolysis, organic components generate reducing gases that promote gas–solid reactions, while solid carbonaceous residues facilitate solid–solid reduction [19]. Our prior works also confirmed the feasibility of this approach, demonstrating that direct pyrolysis of intact cathode sheets can effectively induce in situ thermal reduction [20,21]. This eliminates the need for external chemical reductants in subsequent leaching stages. Moreover, the associated carbothermal reduction converts lithium into water-soluble Li₂CO₃, enabling its priority recovery via simple water leaching. However, during thermal treatment, Al foils can react to form insoluble phases such as LiAlO₂, which compromises lithium extraction efficiency [22,23]. Therefore, developing a tailored recovery strategy to efficiently separate and reclaim all valuable metals from the thermally reduced products of high-Al-content cathodes remains a critical task. Furthermore, while the overall feasibility has been established, the individual contributions and synergistic mechanisms of the key internal components, namely Al foil, carbon, and PVDF binder, in driving the reduction process are not yet fully elucidated.

Herein, this study aims to systematically investigate the specific roles of inherent reducing agents during oxygen-free pyrolysis of the spent cathode sheet and to develop a stepwise, integrated recovery process based on the phase distribution of thermal reduction products. The proposed strategy utilizes in situ reduction to transform cathode materials into separable phases, followed by sequential separation steps, including gentle water leaching for lithium, physical sieving, and targeted alkali–acid leaching, to achieve efficient and comprehensive recovery of all valuable metals. This work seeks to provide important insights into the in situ thermal reduction mechanism in battery waste and to advance a more sustainable and cost-effective recycling paradigm.

2. Materials and Methods

2.1. Materials and Reagents

The spent cathode sheets used in this study were manually dismantled from fully discharged LIBs. Analytical-grade HCl and HNO₃, obtained from Xilong Reagent Corporation (Guangzhou, China), were used to prepare aqua regia for digesting cathode materials, and the leachate was measured by inductively coupled plasma–mass spectrometry (ICP-MS) to determine their original metal content. The main metal composition of the spent cathode sheet is presented in

Table 1. Main metal composition of the spent cathode sheet (wt.%).

Element	Li	Mn	Co	Ni	Al
Content	5.19	15.06	16.29	17.44	5.58

. The results confirm that the cathode active material is LiNi_1/3Co_1/3Mn_1/3O₂ (denoted as NCM), while the Al content originates from the current collector foil. Analytical-grade NaOH solid and H₂SO₄ were used to prepare 1 M aqueous solutions with deionized water for subsequent leaching experiments.

2.2. Experimental Process

An integrated recovery process (Figure 1), comprising in situ thermal reduction, water leaching, alkaline leaching, and acid leaching, was applied to recover valuable metals from the spent cathode sheets.

2.2.1. In Situ Thermal Reduction

Thermal treatment was conducted in a tube furnace (MXG1200-80, Shanghai Micro-X Furnace, Shanghai, China) under a continuous high-purity nitrogen (N₂) flow of 200 mL/min to maintain an oxygen-free atmosphere. Approximately 10 g of each cathode sheet piece was placed in an alumina crucible and positioned in the center zone of the tube. The temperature was raised from room temperature to the target temperature (500, 600, or 700 °C) at a heating rate of 10 °C/min, held for a specified duration (30 min or 2 h), and then cooled to room temperature under a N₂ protection. To specifically investigate the role of carbonaceous materials, the cathode powder containing conductive carbon black and PVDF pyrolytic carbon was gently dislodged from the Al foil after pyrolysis at 500 °C for 30 min and used for subsequent experiments. To study the combined effect of carbon and Al foil, all materials obtained after the 500 °C–30 min pyrolysis were directly subjected to further roasting at 600 °C for 2 h. The roasted products were collected for subsequent steps. To manage the off-gases generated during PVDF pyrolysis, the furnace exhaust was connected to a gas treatment system consisting of condensation collectors for pyrolytic oils, followed by gas-washing bottles containing deionized water and NaOH solution to remove acidic gases, including HF and other fluorinated species [24,25]. The furnace was operated under a continuous N₂ flow throughout the experiment within a fume hood.

2.2.2. Water Leaching and Separation

The water leaching experiments were conducted at room temperature. A pre-weighed amount of roasted product was mixed with deionized water at a liquid-to-solid ratio of 10:1 mL/g. This step serves to selectively dissolve water-soluble lithium compounds (primarily Li₂CO₃) generated during the prior thermal reduction, while the hydraulic agitation simultaneously promotes the detachment of cathode particles from the Al foil. The mixture was mechanically stirred at 400 rpm for 1.5 h to suspend the materials and hydraulically separate cathode particles from the Al foil. For the three-stage leaching experiment, the solid residue from each stage was filtered and subjected to a fresh leaching cycle under identical conditions. The leachates from all stages were separately collected, filtered, and evaporated at 90 °C to yield solid products for analysis. Following the final leaching, the bulk residue was wet-sieved through a 0.5 mm sieve using deionized water. The oversize solid was dried directly to obtain a coarse fraction (+0.5 mm), while the undersize solid was filtered and dried to recover a fine fraction (−0.5 mm). After filtration, the solid residue was rinsed with deionized water using a consistent protocol: approximately half the volume of the original leachate was used to wash the reaction vessel and transfer the solids to the filter, and an additional half-volume was used to rinse the filter cake. All washing solutions were combined with the corresponding leachate prior to volume make-up for ICP-MS analysis.

2.2.3. Alkaline–Acid Leaching

The fine fraction was leached in a 1 M NaOH solution at a liquid-to-solid ratio of 10:1 mL/g. The leaching was conducted at 80 °C for 3 h with magnetic stirring at 400 rpm. After leaching, the slurry was filtered, and the solid residue was thoroughly washed with deionized water until the pH of the filtrate was near neutral. All washings were collected and combined with the alkaline leachate for ICP-MS analysis. The washed residue was then dried to obtain the alkaline leaching residue. The obtained Al/Li-rich solution was subjected to pH adjustment to approximately 6–7 using dilute H₂SO₄, causing Al³⁺ to precipitate as Al(OH)₃, while Li⁺ remained in solution [22]. The Al(OH)₃ precipitate was separated by filtration, yielding a purified lithium solution for subsequent recovery.

The alkali leaching residue was subsequently leached in a 1 M H₂SO₄ solution at a liquid-to-solid ratio of 10:1 mL/g. The process was conducted at 80 °C for 1 h with magnetic stirring at 400 rpm. Upon completion, the slurry was filtered to obtain the acid leachate and the final solid residue. The acid leachate and all washings were combined for subsequent analysis.

2.3. Analytical Methods

The phase composition of solid products was characterized by X-ray diffraction (XRD, BRUKER D8 ADVANCE, Karlsruhe, Germany). The morphology and elemental distribution were examined using scanning electron microscopy (SEM, TESCAN GAIA3 XMH, Brno, Czech Republic) with energy-dispersive X-ray spectroscopy (EDS, OXFORD Ultim Max 65, London, UK). The concentrations of Li, Ni, Co, Mn, and Al in all solutions were determined by ICP-MS (AGILENT 7900, Santa Clara, CA, USA). The relevant metal leaching efficiencies were calculated using the following equation:

$$E={\displaystyle \frac{C_f{V}_f}{mw}}100\%$$

where C_f (g/L) is the metal concentration in the final leachate after dilution to a fixed volume, V_f (L) is the fixed volume to which the leachate was diluted, m (g) is the mass of the solid sample before leaching, and w (%) is the initial mass fraction of the target metal in the solid sample. Fresh deionized water was used as the leachant, so the initial metal concentration was zero.

Key leaching experiments (one-step water leaching) were performed in duplicate, with results presented as mean ± SD. Other leaching experiments were single runs, as indicated in the respective figure captions. All reported efficiencies are supported by complementary XRD and/or SEM analysis of the corresponding solid residues, providing multi-instrument validation of the findings.

3. Results and Discussion

3.1. Thermogravimetric and Phase Transition Analysis of Spent Cathode Sheets

The thermal behavior of the spent cathode sheet under a N₂ atmosphere was analyzed by TG-DSC (Figure 2). The initial weight loss between 25 °C and 150 °C corresponds to the evaporation of residual water and electrolyte. A slight mass gain occurred at 150 °C and 350 °C, followed by a loss from 350 °C to 500 °C due to PVDF binder decomposition [19]. This step is critical, as it not only removes the binder but also initiates the release of reductive gases. Between 500 °C and 690 °C, no significant mass change was observed. However, the DSC curve exhibited two distinct thermal events: an exothermic peak near 600 °C, indicating the collapse of the cathode structure, and an endothermic peak at 653 °C, corresponding to the melting of the Al foil. Subsequently, a rapid mass loss from 690 °C to 900 °C resulted from the continuous release of lattice oxygen from the cathode material during thermal reduction.

Further phase analysis (Figure 3a–d) shows that after treatment at 500 °C for 30 min, the characteristic peaks of the cathode material broadened without new phases forming, indicating minor structural distortion but no intensive reduction, consistent with the dominant role of binder removal at this stage. At 600 °C, the original cathode peaks disappeared, and new peaks corresponding to NiO, CoO, metallic Ni and Co, Li₂CO₃, and LiAlO₂ appeared, confirming cathode decomposition and the involvement of Al foil in the reaction. The coexistence of oxides and metals signifies a progressing reduction process. At 700 °C, the metallic Ni and Co peaks became dominant, demonstrating that higher temperatures promote deeper thermal reduction. Notably, all these transformations were driven solely by the in situ components within the cathode sheet, eliminating the need for external carbon addition and simplifying the process. Furthermore, SEM analysis revealed that the pristine cathode material (Figure 3e) possessed a dense surface, whereas after heat treatment at 500 °C (Figure 3f), cracks emerged on the surface of the spent cathode sheet. This observation further confirms that the removal of the binder during thermal reduction treatment promotes the disintegration of the electrode material, thereby facilitating the subsequent recovery of different components.

3.2. Roles of Different In Situ Components in Thermal Reduction

After binder removal at 500 °C, the adhesion between cathode particles and the Al foil was significantly weakened, allowing easy mechanical separation [24]. The distinct roles of each in situ component during reduction at 600 °C for 2 h were systematically studied, and the phase component of each roasted product is shown in Figure 4a–d. The baseline sample (Figure 4a, 500 °C) showed only broadened cathode peaks. With only solid carbon (pyrolytic carbon and carbon black, NCM+C, Figure 4b), the cathode peaks weakened alongside the emergence of NiO and CoO, indicating carbon-induced structural collapse and partial reduction. Adding Al foil (NCM+C+Al, Figure 4c) further reduced the cathode peak intensity, but the products remained NiO and CoO. In both cases, no Li₂CO₃ or LiAlO₂ was detected. In contrast, when PVDF was present (NCM+C+Al+PVDF, Figure 4d), distinct Li₂CO₃ and LiAlO₂ peaks appeared, and metallic Ni and Co were also detected. The formation of LiAlO₂ can be attributed to the following mechanism: during thermal reduction, lattice oxygen is continuously released from the cathode material under the action of reducing substances, leading to the oxidation of Al foil to form Al₂O₃ [26]. Subsequently, Li₂O derived from the cathode material reacts with Al₂O₃ to generate LiAlO₂ [17]. Additionally, Li continuously migrates from the interior to the surface of cathode particles to form Li₂CO₃ [27], which can also react with Al₂O₃. These results demonstrate that gaseous products from PVDF decomposition are crucial for lithium extraction (forming Li₂CO₃) and the reaction with Al to form LiAlO₂, while also enabling deeper reduction of transition metals to their metallic state.

To quantify the lithium recovery potential enabled by this phase transformation, a one-step water leaching test was performed (Figure 4e). This direct water leaching strategy capitalizes on the in situ-generated Li₂CO₃, selectively extracting lithium while leaving transition metals in the solid residue. The sample treated at 500 °C showed a low Li leaching efficiency of 13.31% due to its intact layered structure. Efficiencies slightly increased to 15.86% (NCM+C) and 21.95% (NCM+C+Al), indicating that solid carbon and Al foil facilitate structural collapse but are ineffective at converting lithium into water-soluble carbonate. Remarkably, when PVDF was present, the efficiency surged to 48.95%. This trend unequivocally proves that pyrolysis gases are the key to forming water-soluble Li₂CO₃, which is essential for the subsequent selective lithium recovery. However, the formation of Li₂CO₃ is accompanied by the generation of water-insoluble LiAlO₂, which cannot be directly recovered by water, presenting a key challenge for achieving higher lithium recovery rates through water leaching alone.

3.3. Lithium Leaching Behavior and Product Separation

The water leaching behavior was further investigated via a three-stage leaching process on the reduced cathode sheet (NCM+C+Al+PVDF condition). As shown in Figure 5a, the Li leaching efficiency decreased sharply with each stage, reaching only 6.93% and 3.48% in the second and third stages, respectively. Phase analysis of the evaporated products (Figure 5b) revealed that Li₂CO₃ was the dominant phase in the first stage, accompanied by minor LiF. LiF became the main product in the second stage, while the third-stage residue primarily contained LiAl₂(OH)₇•2H₂O and LiF. This shift is attributed to the rapid dissolution of the readily soluble Li₂CO₃ during the initial leaching, leaving behind less soluble LiF and LiAl₂(OH)₇•2H₂O in later stages. The persistence and increased predominance of LiF and LiAl₂(OH)₇•2H₂O across the stages confirm the role in inhibiting the kinetics and overall efficiency of lithium recovery via water leaching. Morphology analysis (Figure 6) corroborated this evolution: the first-stage product exhibited rod-like Li₂CO₃ crystals, the second-stage product consisted of flocculent LiAl₂(OH)₇•2H₂O mixed with needle-like LiF, and the third-stage product was predominantly composed of elongated LiF crystals. Collectively, approximately 60% of the total lithium, which is in the forms of Li₂CO₃, LiF, and LiAl₂(OH)₇•2H₂O, can be preferentially recovered through this mild water leaching process.

After water leaching, the residue was wet-sieved (0.5 mm) to separate the metals (Figure 7a). The leachate was enriched with 59.36% of the total Li, along with only 1.18% Al and less than 1% of each transition metal (Ni, Co, Mn), demonstrating successful selective Li extraction. The coarse fraction (+0.5 mm) was rich in Al (87.18%), containing less than 8% of other metals. SEM-EDS analysis of the coarse fraction revealed large, clean surfaces of oxidized Al foil (Figure 7c), confirming the extensive detachment of cathode material achieved. Sparse Ni, Co, and Mn signals were detected on these foils, attributed to residual cathode particles mechanically embedded within the foil [28]. In contrast, the fine fraction (−0.5 mm) contained most of the transition metals (94.35% Ni, 94.18% Co, 91.82% Mn), alongside 32.59% Li and 11.64% Al. This effective physical separation, achieved after gentle hydraulic agitation during water leaching, produces three fractions with unique metallic enrichments ideal for downstream recovery. XRD analysis (Figure 7b) confirmed that the fine powder consisted of Ni, Co, NiO, CoO, and LiAlO₂, with no Li₂CO₃ detected, verifying its prior complete leaching. The presence of LiAlO₂ in this fraction aligns with the earlier hypothesis that it is a water-insoluble product formed during thermal reduction.

3.4. Metal Recovery from the −0.5 mm Fine Fraction

The fine fraction, enriched with over 90% of the transition metals, was processed for metal recovery via a two-step leaching method. Due to the presence of water-insoluble LiAlO₂, alkaline leaching was first applied (Figure 8a), which effectively dissolved 87.76% Al and 58.87% Li from the fine-fraction water leaching residue, confirming LiAlO₂ decomposition and successfully recovering the lithium trapped in this phase. This step opens a route for separate Al and Li recovery from the alkaline solution by pH adjustment [22]. The alkaline leaching residue (Figure 8b) retained the spherical morphology of the cathode particles. Crucially, given the prior in situ thermal reduction that converted the cathode into acid-soluble forms, subsequent acid leaching was performed without any external reductant. This step achieved high extraction rates: 99.31% Li, 94.14% Al, 100% Mn, 97.61% Co, and 96.95% Ni. The final acid residue (Figure 8c) contained only flocculent material, confirming efficient metal dissolution.

Table 2. Overall mass balance and metal recovery distribution across the integrated process (%).

Stream	Li	Ni	Co	Mn	Al
Input (spent cathode sheet)	100	100	100	100	100
Outputs:
(1) Li-rich solution	59.36	0.43	0.29	0.79	1.18
(2) Al-rich coarse fraction	8.05	5.22	5.53	7.39	87.18
(3) Alkaline leachate	19.19	0.78	0.64	1.42	10.22
(4) Acid leachate	13.31	90.71	91.3	90.4	1.34
(5) Final residue	0.09	2.85	2.24	0	0.08
Total recovery (sum of leachates)	91.86	91.93	92.23	92.61	12.74

presents the complete mass balance and metal distribution across each stage of the integrated recycling process. After three-stage water leaching, 59.36% Li was selectively recovered in the Li-rich solution, while the majority of Al (87.18%) reported to the coarse fraction (+0.5 mm). The fine fraction (−0.5 mm), enriched with over 90% of Ni, Co, and Mn, was subsequently processed by alkaline leaching, which extracted 19.19% Li and 10.22% Al. The subsequent acid leaching without any external reductant recovered 90.71% Ni, 91.30% Co, 90.40% Mn, and 13.31% Li from the alkali-treated residue. The final residue contained minimal amounts of all metals (<3%), confirming the high efficiency of the overall process. It should be noted that the total recovery values shown represent the sum of metals recovered in the three leachate streams (water leachate, alkaline leachate, and acid leachate). The aluminum-rich coarse fraction, while not included in this summation, contains 87.18% of the initial aluminum and can be further processed for Al recovery.

Regarding the fate of fluorine, during thermal reduction, a portion is incorporated into solid products as LiF. This LiF partially dissolves during water leaching, while the remaining fraction persists through alkaline leaching due to its insolubility in alkaline media and is ultimately dissolved during acid leaching [22]. Thus, all fluorine is either captured in the gas treatment system or retained in the solid products and managed within the hydrometallurgical process.

From an energy perspective, the thermal reduction temperature of 600 °C is considerably lower than that of conventional pyrometallurgical processes (>1000 °C), suggesting potential energy benefits. In terms of reagent consumption, based on the liquid-to-solid ratio of 10:1 mL/g, processing 1 kg of spent cathode sheet consumes approximately 0.4 kg of NaOH solid for alkaline leaching and about 0.54 L of concentrated H₂SO₄ for acid leaching. More importantly, the proposed process eliminates the need for external reductants (e.g., H₂O₂) during acid leaching, as the prior in situ thermal reduction effectively converts the cathode materials into acid-soluble forms using only the intrinsic components of the spent cathode sheet. Regarding waste stream management, the alkaline leachate is processed by pH adjustment to precipitate Al(OH)₃, yielding a purified lithium solution for subsequent recovery. The acid leachate, containing over 96% of Ni, Co, and Mn, can be directed to downstream recovery processes such as solvent extraction or selective precipitation. The final solid residue contains less than 3% metals by mass. It should be noted that the reagent consumption values reported above are based on laboratory-scale conditions with a fixed liquid-to-solid ratio of 10:1 mL/g, which was selected to ensure complete leaching and demonstrate process feasibility. Further optimization of leaching parameters (e.g., reducing the liquid-to-solid ratio, reagent concentration, and leaching time) is expected to significantly lower reagent consumption in potential scale-up applications.

4. Conclusions

This study successfully developed an integrated in situ thermal reduction and hydrometallurgical process for the selective and efficient recovery of valuable metals from spent cathode sheets. The key conclusions are as follows:

(1)

The internal components of the cathode sheet effectively drive thermal reduction without external reagents. Solid carbon and Al foil participate in cathode structure collapse, reducing transition metals, and most critically, gaseous products from PVDF decomposition are indispensable for extracting lithium as water-soluble Li₂CO₃, creating the foundation for selective recovery.

(2)

The formation of Li₂CO₃ enabled its preferential separation via simple water leaching, which avoided the fine dust generation of dry crushing and achieved in situ detachment of cathode materials. Under optimal conditions, a one-step water leaching achieved 48.95% Li leaching efficiency while leaving over 99% of the transition metals in the solid residue.

(3)

A three-stage water leaching process achieved a cumulative lithium recovery of 59.36%. Subsequent wet sieving efficiently separated the products into a Li-rich solution, an Al-rich coarse fraction, and nickel, cobalt, and a transition-metal-enriched fine powder (−0.5 mm, >91% of Ni, Co, Mn), facilitating targeted downstream processing.

(4)

A two-step alkali–acid leaching process effectively recovered metals from the fine fraction. Notably, the acid leaching step required no external reductant, benefiting from the pre-reduction during in situ thermal treatment, and achieved high leaching efficiencies of 96.95% Ni, 97.61% Co, 100% Mn, and 99.31% Li from the alkali-treated residue.

(5)

Based on the complete mass balance analysis, the integrated process achieved overall recoveries of 91.86% for Li, 91.93% for Ni, 92.23% for Co, and 92.61% for Mn from the combined leachate streams. Additionally, 87.18% of Al was concentrated in the coarse fraction for subsequent recovery, with minimal residual metals (<3%) remaining in the final solid residue.