Reductant-Free Cobalt Recovery from Similar Copper–Cobalt Oxide Ores via Synergistic Reductive-Acid Leaching

Abstract

Enhancing cobalt recovery from complex low-grade copper–cobalt oxide ores represents a pressing industrial challenge in the Democratic Republic of the Congo (DRC). This study comparatively examined the sulfuric acid leaching behaviors of two copper–cobalt oxide ores with similar mineralogical characteristics and developed a synergistic blending strategy to enhance cobalt recovery. At an endpoint pH of 1.5, both ores achieved high copper extraction (>90%). However, the reductive Re-ore attained >75% Co leaching efficiency, contrasting sharply with the oxidative Ox-ore’s limited 22% recovery. This disparity was attributed to refractory high-valent cobalt phases in Ox-ore requiring a reductive environment. Blending Re-ore with Ox-ore generated a pronounced synergistic effect, progressively lowering the slurry potential with increasing Re-ore mass ratio. At 50% Re-ore incorporation, the leaching efficiency of cobalt from Ox-ore surged from 22.5% to 76% owing to the decrease in slurry potential to 557 mV, substantially exceeding proportional predictions. Residue characterizations confirmed that cobalt phases in Ox-ore were solubilized in the reductive environment facilitated by the residual sulfide phase in Re-ore. Enhanced dissolution of Fe²⁺ and Mn²⁺ further correlated with potential-dependent cobalt recovery. This ore-blending strategy provides a cost-efficient alternative to chemical reductants by leveraging intrinsic ore properties to optimize cobalt extraction from challenging oxidized ores.

1. Introduction

Cobalt is an essential metal used in renewable energy, communications, aerospace, and defense technologies. Many countries have classified cobalt as a critical mineral for national economic and security interests [1,2]. However, global cobalt resources are highly unevenly distributed. The Democratic Republic of the Congo (DRC) accounts for approximately 50% of the global reserves [3,4]. In 2024, the DRC produced around 220,000 tons of cobalt, establishing itself as the world’s largest producer. The mineralogy of copper and cobalt resources in the DRC is complex, with diverse occurrence states [5,6]. Moreover, these resources are often associated with impurities such as iron, manganese, calcium, and magnesium, which lead to low cobalt recovery in traditional leaching processes [4,7]. Cobalt in copper–cobalt ores often exists as refractory minerals such as stainierite or adsorbed forms, resulting in significantly low leaching efficiency as compared to copper [8]. Therefore, improving cobalt recovery from complex low-grade copper–cobalt ores has become an urgent industry problem in the DRC.

Copper and cobalt resources in the DRC mainly include oxide deposits, sulfide deposits, and mixed oxygen–sulfur deposits [9]. With increasing mining of oxidized ores and blended ores, the extraction processes are shifting toward hydrometallurgy or combined beneficiation and metallurgical technology [10,11]. At present, hydrometallurgy is the main method for leaching copper–cobalt oxide ores. Agitated leaching is suitable for high-grade copper–cobalt ores and leachable minerals, while heap leaching is generally applicable to low-grade copper–cobalt ores [12,13,14]. Usually, sulfuric acid is used as the leaching agent, and the leaching efficiency of copper and cobalt can be increased by optimizing the leaching conditions such as the temperature, liquid–solid ratio, and acid concentration [15,16,17,18]. For instance, with an optimized dosage of sulfuric acid and leaching time, the leaching efficiencies of copper and cobalt reached 95.88% and 91.88% from a low-grade copper–cobalt oxide ore in the DRC [19]. In addition, a two-stage leaching process was used to enhance the selective recovery of copper and cobalt. After copper leaching, iron powders were used as a reductant to selectively leach cobalt. Since copper–cobalt ores often contain gangue minerals such as carbonates, their high acid consumption will increase production costs [20,21]. Zhang et al. studied copper and cobalt recovery from copper–cobalt ore in the DRC by ammonium chloride roasting and found that 90% copper and 95% cobalt can be extracted via roasting at 300 °C for 3 h under the optimal NH₄Cl-to-ore mass ratio of 1:2 [22]. However, high ammonia costs and stringent environmental regulations limit the practical application of this method.

Although copper oxide minerals such as malachite and azurite are readily soluble in acidic solutions, tri-valent cobalt phases (e.g., CoO(OH)x) need reductive leaching using agents such as SO₂, Fe²⁺, and sulfite salts [23,24]. For instance, Jie et al. used SO₂ as a reductant in a sulfuric acid system, achieving leaching efficiencies of 94.6% for copper and 78.9% for cobalt [25]. Li et al. carried out reductive-acid leaching of a low-grade copper–cobalt oxide ore at 80 °C for 4 h, where SO₂ was introduced to control the endpoint potential of the leachate to 340–350 mV, resulting in copper and cobalt recovery of 74.23% and 59.31%, respectively [26]. Uahengo et al. studied reductive leaching in the presence of ferrous sulfate, ammonium ferrous sulfate, or ferric sulfate additives and found that ferrous sulfate promoted the leaching of cobalt and manganese, while ferric sulfate enabled selective leaching of copper [27]. Liu et al. reported that the addition of 1.5 g/L ferrous sulfate increased the cobalt leaching efficiency from 43.32% to 78.25% [28]. Yu et al. used sodium sulfite as a reductant to leach a copper–cobalt oxide ore in the DRC, achieving cobalt recovery exceeding 87% [29]. Clotilde et al. [30] observed that without ferrous ions, the cobalt leaching efficiency was only 36% in hydrochloric acid leaching; however, when the molar ratio of Fe²⁺ to Co exceeded 2:1, cobalt recovery increased to 96%. Cao et al. compared direct acid leaching with reductive leaching and found that the addition of sodium sulfite improved the leaching efficiencies of copper and cobalt to 95.88% and 91.88%, respectively [20]. Although these reductants effectively promote cobalt leaching, they significantly increase reagent costs and complicate subsequent separation processes. Zuo et al. [21] efficiently extracted copper and cobalt by the redox reaction between CuCo₂S₄ and CoOOH phases at 120 °C in 2.5 mol/L sulfuric acid for 0.5 h, where the leaching efficiencies of Cu and Co reached 99.67% and 98.20%, respectively. Although a relatively high temperature is needed, this process effectively decomposes oxidized ores by reductive sulfide ores without additional reductants, highlighting the role of redox reactions during leaching [31,32]. However, sulfide ore resources in the DRC are gradually depleting, making it challenging to rely solely on sulfide ores for sustainable cobalt recovery.

In this work, based on the slurry potential differences of two copper–cobalt oxide ores with similar mineralogical characteristics, a synergistic ore-blending strategy was developed to improve cobalt leaching. The influences of the blending ratio, endpoint acidity, slurry potential, and reductant dosage on cobalt leaching efficiency were investigated. Through regulating the slurry potential of mixed ores, cobalt recovery was significantly enhanced, reducing or even eliminating the need for chemical reductants.

2. Experimental Methods

2.1. Materials and Reagents

Two typical copper–cobalt oxide ores (denoted by Re-ore and Ox-ore) were obtained from Huagang Mining Company (Kolwezi, DRC). Analytical-grade sulfuric acid and sodium metabisulfite were purchased from Sinopharm Group Chemical Reagent Co., LTD (Beijing, China). Deionized water was used throughout the experiments.

2.2. Leaching Process

Leaching tests were conducted in sulfuric acid medium at a solid/liquid ratio of 1:3 and 25 °C for 3 h. A slurry containing 100 g ore and 300 mL sulfuric acid solution was agitated at 400 rpm using a top-mounted mixer (IKA, Staufen, Germany). The slurry potential and pH were monitored in real time. After leaching and filtration, the concentrations of Cu, Co, Mn, Mg, Al, and Fe were analyzed. The average value of three analyses was used. In the reductive leaching experiments, sodium metabisulfite was added after 1 h of acidic leaching to avoid rapid decomposition by strong acid. The blending ratio of Re-ore to Ox-ore was controlled according to the actual production situation, where the total copper grade was 2%–5% and the total cobalt grade was 0.2%–0.5%.

2.3. Analysis and Characterization Methods

Raw ores and leaching residues were digested using a four-acid method for elemental analysis. The contents of Cu, Co, Fe, Mg, Al, Ca, and Mn were analyzed by using an inductively coupled plasma optical emission spectrometer (iCAP PRO XPS, Thermo Fisher Scientific, Waltham, MA, USA) and atomic absorption spectrometry (iCE 3300, Thermo Fisher Scientific, Waltham, MA, USA). Industrial HY-R200T instruments (Henyi, Hangzhou, China) were used to monitor the slurry pH and potential in real time. Redox potential values (vs. Ag/AgCl) were collected using an ORP electrode (Henyi, Hangzhou, China) adjusted based on the standard hydrogen electrode (SHE). Phase analysis was performed using a D8 ENDEAVOR X-ray diffractometer (XRD, Bruker, Hamburg, Germany) operated at 40 kV and 40 mA with Cu Kα radiation. The morphology and elemental distribution were examined using a Regulus8230 scanning electron microscope (SEM, Hitachi, Tokyo, Japan). The valence states and species compositions of Co, Mn, and Fe were analyzed by an X-ray photoelectron spectrometer (XPS, Thermo Scientific K-Alpha, Thermo Fisher Scientific, Waltham, MA, USA).

3. Results and Discussion

3.1. Chemical Analysis and Phase Composition

After grinding, Re-ore and Ox-ore were wet-sieved to determine their particle size distributions. As shown in Figure 1a, the particle size ranges mainly between 150 and 74 μm, with over 80% of particles below 100 μm. Figure 1b shows that Re-ore contains 6.40% Cu and 0.081% Co, while Ox-ore contains 3.15% Cu and 0.44% Co. In both Re-ore and Ox-ore, the copper-bearing minerals mainly include malachite, chrysocolla, libethenite, pseudo-malachite, chalcocite, covellite, chalcopyrite, etc. (Figure 1c). Cobalt is primarily present as stainierite associated with limonite. The gangue phases consist mainly of quartz, dolomite, chlorite, feldspar, etc. (Figure 1d). Among them, quartz accounts for more than 60%, and the feldspar content in Ox-ore is significantly higher than that in Re-ore. The X-ray diffraction (XRD) pattern in Figure 2a shows similar intercalation characteristics of copper-bearing minerals in both ores. The optical images (Figure 2b–f) show that malachite is mainly associated with limonite, chrysocolla, chalcocite, and gangue minerals, while chrysocolla occurs as an irregular granular form intergrown with malachite, limonite, and stainierite [33]. Notably, Re-ore contains secondary copper sulfide minerals such as chalcocite and covellite, which account for 11.89% of the total copper and are encapsulated in gangue or malachite as irregular or fine grains.

However, it constitutes only 1.23% of the total copper in Ox-ore. Stainierite, the main cobalt oxide mineral, appears as irregular grains associated with malachite, limonite, copper–cobalt–manganese oxides, and gangue minerals. Some copper–cobalt limonite occurs as irregular grains intergrown with chalcocite, blue chalcocite, malachite, chrysocolla, and stainierite. Except for differences in their cobalt content and secondary copper sulfide abundance, both Re-ore and Ox-ore exhibit typical characteristics of oxidized ores with highly similar chemical and mineralogical compositions.

3.2. Leaching Behaviors

The influence of reaction time on copper and cobalt leaching was evaluated at a liquid–solid ratio of 3:1 and endpoint pH of 1.5. Leachates were taken and analyzed at 10, 20, 30, 40, 60, 90, 120, and 180 min. Figure 3a,b show that copper leaching is rapid in both ores, basically stabilizing within 30 min, and the leaching efficiencies reach about 90%. The leaching efficiency of Cu in Ox-ore is slightly higher than that in Re-ore. The leaching of Co, Mn, and Ca in Re-ore increases rapidly within 60 min and then stabilizes, with cobalt recovery exceeding 75%. The Co leaching of Ox-ore reaches stability after 90 min but reaches only 22%, significantly lower than that of Re-ore. Moreover, manganese and calcium dissolution continue to increase. The rapid copper dissolution (within 30 min) indicates surface reaction control, consistent with direct acid leaching of copper oxide ores. In contrast, the slower cobalt leaching rate, especially in Ox-ore, indicates diffusion-controlled or mixed kinetics [34]. The influence of the endpoint pH on leaching was further evaluated at a liquid–solid ratio of 3:1 by gradually supplementing sulfuric acid during the 3 h leaching process. As shown in Figure 3c,d, increasing the slurry pH from 1.0 to 2.0 slightly decreases the recovery of Cu and Co from Re-ore but significantly reduces leaching of Fe and Ca. For Ox-ore, the copper leaching remains basically unchanged, while the Co leaching efficiency decreases from 25.6% to 21.1%. Moreover, the leaching of impurities such as Fe, Mg, Al, Ca, and Mn shows little change. Gradual acid addition results in slightly higher cobalt recovery compared to single-step acid addition, likely because a high initial acid concentration promotes rapid gangue reactions, reducing acid availability for valuable metals. Figure 3e,f show that as the endpoint pH increases, the acid consumption of Re-ore decreases from 229.2 to 172.6 kg/t, and the corresponding slurry potential decreases from 354 to 322 mV, indicating inherent reducibility. For Ox-ore, the acid consumption decreases from 148.5 to 91.3 kg/t, and the slurry potential decreases from 821 to 605 mV, reflecting an oxidative environment.

As mentioned above, the low cobalt recovery from Ox-ore was attributed to the presence of high-valence cobalt phases requiring reductive dissolution [21,35]. To enhance cobalt recovery, sodium metabisulfite solution (0.5 kg/L) was added after 1 h of acid leaching to avoid rapid decomposition under initial strong acid conditions. Figure 4a shows that the leaching efficiencies of Cu and Fe in Ox-ore are unaffected by reductant addition, but Co recovery gradually increases as the reductant dosage increases. At 36.33 kg/t sodium metabisulfite consumption, the leaching efficiency of Co reaches 56.46%, and Mn leaching also significantly increases from 27.44% to 92%, indicating a coupled Co–Mn dissolution behavior. Although the slurry pH shows no significant change at a dosage of 14.27 kg/t, the potential rapidly decreases to 380 mV, but it increases to 650 mV within 20 min after reductant addition and stabilizes near 700 mV. Under the condition of 36.33 kg/t, excess sodium metabisulfite consumes acid, resulting in a significant increase in the slurry pH. Although the endpoint potential can be maintained at ca. 550 mV, the leaching efficiency of Co shows no effective increase compared with that for a 14.27 kg/t reductant dosage.

Thermodynamic analysis based on the established literature [13] indicates that due to the high reduction potential of the Co³⁺/Co²⁺ couple in acidic media, Co(III) phases are stable only under highly oxidizing conditions. The measured potential value of Ox-ore slurry (~800 mV) is insufficient to reduce Co(III), explaining its poor cobalt recovery without reductants. The addition of sodium metabisulfite enhances cobalt leaching, which can be attributed to the following redox reactions:

Na₂S₂O₅ + H₂SO₄ → 2SO₂↑ + H₂O + Na₂SO₄

(1)

SO₂ + H₂O → H₂SO₃

(2)

2CoO(OH) + H₂SO₃ + H₂SO₄ → 2CoSO₄ + 3H₂O

(3)

Fe₂O₃ + H₂SO₃ + H₂SO₄→ 2FeSO₄ + 2H₂O

(4)

MnO₂ + SO₂ → MnSO₄

(5)

MnO₂ + 2FeSO₄ + 2H₂SO₄ → MnSO₄ + Fe₂(SO₄)₃ + 2H₂O

(6)

Although sodium metabisulfite boosts cobalt leaching, the large amounts consumed would greatly increase production costs [16]. Given the inherent reducibility of Re-ore slurry, blending these two ores was explored to achieve synergistic cobalt leaching without external reductants. Figure 4d shows the effect of the Re-ore/Ox-ore blending ratio on leaching efficiency. The dotted lines represent the proportional weighted averages of individual ore leaching. The degree to which the experimental results deviate from the dotted lines can reflect the synergistic effect of ore blending. With increasing mass ratio of Re-ore, the leaching efficiency of Cu remains basically unchanged, but the cobalt leaching increases significantly. When the mass ratio of Re-ore increases from 0% to 50%, the leaching efficiency of Co increases from 22.5% to 76%, which obviously exceeds the proportional weighted leaching efficiency. These results indicate that blending with Re-ore can achieve synergistic cobalt recovery from Ox-ore without additional reductant consumption. Figure 4e shows that during the leaching process, the slurry pH remains stable but the slurry potential decreases significantly with an increase in the Re-ore ratio. At 20%, 30%, 40%, and 50% Re-ore, the endpoint potentials are 814, 776, 657, and 557 mV, respectively. When the blending proportion of Re-ore exceeds 40%, the slurry potential continuously decreases over time, confirming the promoting effect of Re-ore on cobalt leaching.

Further potential control of the blended ores was achieved by adding sodium metabisulfite after 1 h of acid leaching to adjust the potential to 420 mV. As shown in Figure 4g, when the mass ratio of Re-ore increases from 20% to 40%, the further decrease in slurry potential to 420 mV has little effect on Cu leaching, but the Co leaching efficiency increases to 74.1%, 75.4%, and 77.9%, providing improvements of 42.5%, 26.4%, and 13.9% compared to leaching without a reductant, respectively. However, even at 420 mV, approximately 22% of the cobalt remains unleached for the blended ore (Re-ore/Ox-ore = 0.4/0.6), likely due to encapsulation in gangue or insufficient reactivity at this potential [15,17]. Meanwhile, the variation in the slurry pH is similar under different ore-blending ratios. Slurry pH changes are similar across blending ratios, but potential profiles differ significantly before and after reductant addition, decreasing with higher Re-ore proportions. Correspondingly, iron leaching increases slightly, while manganese leaching is markedly improved. Given the higher grades of Fe (2.25%) and Mn (0.16%) in Ox-ore, the dissolved oxidizing species may elevate the slurry potential, making control challenging. The endpoint potentials increase to 600, 536, and 500 mV for 20%, 30%, and 40% Re-ore blends, respectively.

Figure 5 shows that the Re-ore residue particles have a clear interface, where the distribution of Cu is locally aggregated and cobalt is difficult to distinguish, indicating that the unleached copper and cobalt may be encapsulated. The Ox-ore residue particles are finer, with a more distinct cobalt distribution, suggesting limited dissociation of cobalt-containing phases in oxidative environments. Compared with those of Re-ore and Ox-ore, the leaching residue particles of blended ore are significantly finer and amorphous. Moreover, the distribution of residual Cu and Co is more dispersed. This phenomenon may be related to the particle refinement in the leaching process because the acid consumption of Re-ore is higher than that of Ox-ore. When sulfuric acid was added according to the blended ore ratio, the initial acidity was enhanced for Ox-ore, thereby altering the morphology of the leaching residue. However, the synergistic leaching of Co should mainly result from the reduced slurry potential rather than particle refinement.

XRD analysis (Figure 6a) shows that due to their low grades, copper and cobalt phases are difficult to identify in the three ores and residues. The main gangue phases include KMg₃Si₃AlO₁₀(OH)₂ (phlogopite), (Mg,Fe)₆(Si,Al)₄O₁₀(OH)₈ (chlorite), CaMg(CO₃)₂ (dolomite), Fe₂O₃, and SiO₂. The content of chlorite in the residue significantly reduces after leaching, indicating significant decomposition during the acid leaching process. The layer structure of phlogopite and chlorite is conducive to penetration of leaching agents and valuable metal dissolution, thus ensuring high copper recovery. FT-IR spectra (Figure 6b) show that there is no obvious difference in the bonding forms of the main phases in the three samples. The peaks near 900 cm⁻¹ correspond to the Si-O-Si stretching vibration, and the peaks near 3600 cm⁻¹ are attributed to the O-H stretching vibration. Moreover, the O-H peaks near 3420 cm⁻¹ are significantly weakened for the three residues, indicating that the hydroxyl-containing phases are significantly dissolved. Raman spectra (Figure 6c) show that the peaks near 454 cm⁻¹ and 3624 cm⁻¹ correspond to the Si-O-Si bending vibration and O-H stretching vibration, respectively. No new peaks appear after leaching for Ox-ore, and the peak intensity and position show no significant change. However, for Re-ore, the peak near 1104 cm⁻¹ disappears and a new peak appears at 1020 cm⁻¹ after leaching, which might be the vibration peak of sulfate ions in the precipitated calcium sulfate, indicating that some Ca-containing phases are transformed into calcium sulfate during the leaching process. The characteristic peaks of the blending ore residue near 454 and 1015 cm⁻¹ exhibit the features of both Ox-ore and Re-ore residues, and the -OH peak near 3350 cm⁻¹ is significantly weakened.

The leaching of Co is affected by the potential and acidity of the feed solution, but the valence-state influence of multi-valent elements is also particularly important. Figure 7 shows the XPS spectra of Fe 2p, Mn 2p, and Co 2p in the residues of Ox-ore, Re-ore, and blended ore. The Fe2p peaks of Ox-ore and Re-ore have similar features. The characteristic peaks at 710.8, 723.5, 726.4, 713.9, 719.5, and 732.3 eV correspond to the 2p_3/2/2p_1/2 orbits, satellite peaks of Fe²⁺ and 2p_3/2/2p_1/2 orbits, and satellite peaks of Fe³⁺, respectively. After leaching of the blended ore, the peak at 710.8 eV remains unchanged but the relative intensity increases, suggesting increased relative abundance or surface exposure of Fe²⁺ species. The peak shifts to 724.4, 727.9, 714.3, 718.5, and 731.7 eV indicate that Fe ions react with components such as OH⁻ and SO₄²⁻ in the leachate, thereby changing their binding energy. The Co 2p spectra of Ox-ore are located at 781.2, 796.8, 787.1, and 803 eV, which corresponds to the 2p_3/2/2p_1/2 orbits of Co²⁺ and 2p_3/2/2p_1/2 orbits of Co³⁺, indicating mixed Co³⁺/Co²⁺ species. Although the Co 2p peaks of Re-ore are similar to those of Ox-ore, the 2p_1/2 peak of both Co²⁺ and Co³⁺ shifts towards the low-energy direction, indicating that there are more low-valent cobalt species that are prone to leaching at a low slurry potential. The Co 2p peak of the blended ore residue is more similar to that of Ox-ore, indicating that some insoluble cobalt sulfide (Co₉S₈) and CoOOH particles still remain. Mn 2p peaks are mainly located at 641.5 and 649.9 eV for Ox-ore and 641.4 and 649.7 eV for Re-ore, indicating that the manganese species result from MnCO₃ or MnO₂ phases. Due to their high leaching efficiency, no Mn species could be detected in the blended ore residue. By comparing the leaching efficiencies of Fe and Mn, it can be inferred that the low slurry potential also benefits the formation of Fe²⁺ and Mn²⁺, further promoting cobalt recovery.

Based on the above analysis, the residual sulfide mineral in Re-ore plays a crucial role in establishing the reducing environment. The synergistic reductive leaching mechanism of cobalt via blending is shown in Figure 8. Due to the low slurry potential (~300 mV) of Re-ore, the blended ore can lower the slurry potential to ca. 500 mV, which is within the stability field of Co²⁺ according to established E-pH diagrams for the Co-SO₄^2--H₂O system [13,21]. In addition, the dissolved Fe²⁺ and Mn²⁺ ions further contribute to reducing capacity and facilitate electron transfer. These sulfides undergo oxidation, generating reducing species that promote the reduction of Co(III) to Co(II) in Ox-ore. The relevant reactions are as follows:

Cu₂S + Fe³⁺ → Cu²⁺ + CuS + Fe²⁺

(7)

Cu₂S + Mn⁴⁺ → Cu²⁺ + CuS + Mn²⁺

(8)

2CoO(OH) + 2FeSO₄ + 3H₂SO₄ → 2CoSO₄ + Fe₂(SO₄)₃ + 4H₂O

(9)

The proposed ore-blending strategy offers distinct advantages over conventional reductive leaching [4,16,17,18]. It achieves high cobalt recovery comparable to that with chemical reductants (76.0% without any added reductant), while significantly reducing reagent consumption and cost. The operating potential range (557–420 mV) is higher than that typically used in reductive leaching (350–400 mV), suggesting that the blending approach may be easier to control in industrial settings. Moreover, by minimizing or eliminating external reductants, the process reduces downstream complications and wastewater treatment requirements. Thus, the ore-blending strategy presents a promising approach for improving cobalt recovery from complex oxide ores in the DRC, potentially maximizing resource utilization and reducing dependence on external reagents.

4. Conclusions

This study investigated the leaching behaviors of two mineralogically similar copper–cobalt oxide ores (Re-ore and Ox-ore) and developed a synergistic blending strategy to enhance cobalt recovery. Chemical and mineralogical analyses revealed analogous gangue compositions and copper-bearing minerals in both ores. Re-ore contains significantly more cobalt (0.081% vs. 0.44%) and secondary copper sulfides (11.89% vs. 1.23% of total Cu), while Ox-ore contains more feldspar. Crucially, the Re-ore slurry demonstrates inherent reducibility, whereas Ox-ore slurry exhibits oxidative characteristics. Under sulfuric acid leaching at pH 1.5, both ores achieve high copper extraction (>90%), but cobalt recovery from Re-ore exceeds 75%, contrasting sharply with Ox-ore’s limited 22% recovery. This disparity is attributed to refractory high-valent cobalt phases in Ox-ore requiring reductive dissolution. Blending Re-ore with Ox-ore generates a pronounced synergistic effect, progressively lowering the slurry potential with increasing Re-ore ratio. At 50% Re-ore incorporation, the cobalt leaching efficiency surges from 22.5% to 76%, far exceeding proportional predictions. Residue characterization confirmed that cobalt phases in Ox-ore are solubilized in the reducing environment facilitated by Re-ore blending. Enhanced dissolution of Fe²⁺ and Mn²⁺ further correlates with potential-dependent cobalt recovery. This ore-blending strategy provides a cost-efficient alternative to chemical reductants by leveraging intrinsic ore properties to optimize cobalt extraction from challenging oxidized ores.