# Characterisation, concentration and acidic reductive leaching of limonitic laterites from New Caledonia

Agathe Hubau, Agnieszka Dybowska, Paul F Schofield, Diego Gianolio, Ana Santos, Richard J Herrington, Douglas Pino-Herrera, Hafida Tris, Marion Erard, Anne-Gwénaëlle Guezennec

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Agathe Hubau, Agnieszka Dybowska, Paul F Schofield, Diego Gianolio, Ana Santos, et al.. Characterisation, concentration and acidic reductive leaching of limonitic laterites from New Caledonia. Applied Geochemistry, 2025, 193, pp.106569. 10.1016/j.apgeochem.2025.106569. hal-05291578

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# Characterisation, concentration and acidic reductive leaching of limonitic laterites from New Caledonia

Agathe Hubau<sup>a,*</sup>, Agnieszka Dybowska<sup>b</sup>, Paul F. Schofield<sup>b</sup>, Diego Gianolio<sup>c</sup>, Ana Santos<sup>b</sup>, Richard J. Herrington<sup>b</sup>, Douglas Pino-Herrera<sup>a</sup>, Hafida Tris<sup>a</sup>, Marion Erard<sup>a</sup>, Anne-Gwénaëlle Guezennec<sup>a</sup>

## ARTICLE INFO

Editorial handling by: Elisa Sacchi

Keywords:

Laterite

Cobalt

Nickel

Reductive leaching

Limonites

Ore processing

Chromium speciation

## ABSTRACT

This study aimed at determining Ni and Co leaching kinetics from a New Caledonian laterite in an acidic medium (H₂SO₄ pH 1.5) and in a reductive environment (addition of SO₃²⁻ or Fe(II)) at 46 °C. The mineralogical study revealed that Co was mainly carried by Mn oxyhydroxides in the limonite sample. Conversely, Ni was hosted by both Fe and Mn oxyhydroxides. In the presence of a reductive reagent, Mn oxyhydroxides dissolved rapidly compared to goethite, the main Fe oxyhydroxide in the sample. Co, Mn and Ni reductive leaching yields reached 79 %, 83 % and 9 % respectively after 2 days. Based on these results, a Mn oxides concentrate was produced in order to efficiently leach Co while limiting Fe oxyhydroxide dissolution. This concentrate resulted from a combination of particle size and gravity separation steps. The volume/mass of sample was drastically decreased since the mass of the final sample was only 3.3 % of the initial one. Co content increased from 0.16 wt% in the limonite to 2.3 wt% in the concentrate, representing an enrichment factor of 13.8 and recovery yield of 60 %. Co. Mn and Ni leaching yields reached 87 %, 95 % and 80 % respectively in the Mn oxides concentrate leaching experiment. The difference in Ni behaviour was consistent with the mineralogical composition: Ni was mainly carried by the goethite in the laterite, while it was hosted mainly by the Mn oxyhydroxides in the Mn oxides concentrate. This study gives a proof of concept for the development a robust pre-concentration process to recover Co.

## 1. Introduction

Cobalt (Co) and nickel (Ni) have exceptional high-temperature strength, corrosion-resistance, catalytic and electrochemical properties which make them crucial elements for high performance applications. Due to the development of these applications, and in particular lithium-ion batteries, Co and Ni demands are continually growing. Co and Ni are thus classified by the European Commission as critical raw materials (European Commission, 2023).

During primary mining, Co is usually a by-product and more than 50 % of the world-wide Co production comes from Ni primary production. The majority of global Ni resources are hosted in laterite deposits and the remaining portion in sulfidic deposits (Mudd and Jowitt, 2014; Oxley et al., 2016). Owing to the exhaustion of sulfidic deposits as well

as to the increase in demand for Ni and Co, the production of these metals from laterites continues to increase, despite higher environmental and economic costs compared to sulfide processing (Fraser et al., 2021; Pell et al., 2021). In 2022, more than 70 % of the global Ni primary production was based on laterite treatment (MineralInfo, 2024). In particular, 6.2 % (200 kt) of the world-wide Ni production originates from New Caledonia, which is a local authority of the French Republic. Around 1.5 kt of Co were produced as a by-product from the New Caledonian laterites in 2019. The reserves of Co contained in the New Caledonian laterites were evaluated as 0.2 % of Co global reserves in 2019 (MineralInfo, 2020).

Laterite deposits are characterised by the presence of two distinct zones:

<footnote><sup>a</sup> BRGM, F-45060, Orléans, France</footnote>

<footnote><sup>b</sup> Resourcing the Green Economy, Natural History Museum, London, SW7 5BD, United Kingdom</footnote>

<footnote><sup>c</sup> Diamond Light Source Ltd, Harwell Science Campus, Chilton, OX11 0QX, United Kingdom</footnote>

<footnote>* Corresponding author.
E-mail address: a.hubau@brgm.fr (A. Hubau).</footnote>
* A limonite zone near the surface where Ni is mainly hosted in ferric iron oxides or oxyhydroxides, typically goethite (FeOOH) and to a lesser extent in Mn oxides, in which Mn occurs as Mn(III) and/or Mn (IV). Ni content in the limonite zone usually varies between 1 and 2 %.

* A lower saprolite zone typically dominated by hydrous magnesium silicates (MgO>20 %) (Freyssinet et al., 2005), with Ni content ranging between 1.5 and 2.5 %.

Co-bearing phases vary along the weathering profile, between Mn oxides and Fe oxyhydroxides (Dublet et al., 2017; Llorca and Monchoux, 1991), while its content reaches around 100 mg kg⁻¹ regardless of the zone (Bailly et al., 2021).

To obtain a Ni product from laterites that is suitable for the battery market, mixed hydroxide precipitate (MHP) is emerging as the most preferred material. Its production is based on High Pressure Acid Leaching (HPAL) of limonitic laterites. Historically, HPAL plants have often proven to be very costly and plagued by technical challenges (such as autoclaves scaling) and consequently limonites ore often remained stockpiled for future processing. Only the saprolite was treated by pyrometallurgical reduction to produce Ni. In recent years, changes in HPAL technology have occurred and many new HPAL plants are under construction, particularly in Indonesia. This technology is however a point of concern as its environmental footprint generates higher CO₂ emissions compared to the emissions of the European plants (Bartzas and Komnitsas, 2024). Moreover, its use in the New Caledonian context proved to be economically unviable (Requin et al., 2023). Even though the production of Ni from the three New Caledonian metallurgies (Société Le Nickel, Koniambo Nickel SAS and Prony Resources Nouvelle-Calédonie) could theoretically cover 14 % of Ni needs of the European Union by 2035, none of these operators are currently able to meet the demand.

Other hydrometallurgical techniques are thus being investigated in order to extract both Ni and Co from laterites (Oxley et al., 2016; He et al., 2023). The efficiency of atmospheric acidic reductive leaching for Co and/or Ni recovery has been demonstrated (Moro et al., 2023; Komnitsas et al., 2023). In these conditions, two distinct reactions occur. First, due to the presence of protons, goethite is in equilibrium with ferric iron, leading to its dissolution (Eq. (1), Senanayake et al., 2011):

$$FeOOH + 3H^+ \leftrightarrow Fe^{3+} + 2H_2O \quad \Delta G^{0}_{25^{\circ}C} = 13.14 \text{ kJ.mol}^{-1} \quad [1]$$

In reductive conditions, in the presence of an electron acceptor, Fe(III) is reduced into Fe(II) according to the half-reaction (Eq. (2)):

$$Fe^{3+} + e^{-} \rightarrow Fe^{2+} \quad [2]$$

Due to this reduction, the equilibrium shown in Eq. (1) is shifted towards goethite dissolution. The reduction (Eq. (2)) also produces the Fe(II) that will be involved in the dissolution of Mn oxides (Eq. (3) and Eq. (4), as a function of Mn oxidation state):

$$MnO_2 + 2Fe^{2+} + 4H^+ \rightarrow Mn^{2+} + 2Fe^{3+} + 2H_2O \quad \Delta G^{0}_{25^{\circ}C} = -88.68 \text{ kJ.mol}^{-1} \quad [3]$$

$$MnOOH + Fe^{2+} + 3H^+ \rightarrow Mn^{2+} + Fe^{3+} + 2H_2O \quad \Delta G^{0}_{25^{\circ}C} = -85.02 \text{ kJ.mol}^{-1} \quad [4]$$

Ni and Co dissolution rates are dependent of the type of oxide, type of ore and acidity (McDonald and Whittington, 2008; Senanayake et al., 2011). Acidity plays a key role as precipitates may be generated and can hinder the reactivity of ferric iron. Similarly, redox potential plays a crucial role as it regulates the ratio of ferrous to ferric iron concentrations, and thus the kinetics of Eqs. (1)–(4). Higher Co dissolution yields are often obtained, compared to highly variable Ni dissolution yields. In recent years, some studies focused on the in situ production of the reductive agent using microorganisms (Nasab et al., 2020; Newsome

et al., 2020; Johnson and Du Plessis, 2015). Due to its current use at the industrial scale, acidophilic bioleaching has particularly raised attention (Santos et al., 2020; Johnson et al., 2022; Stanković et al., 2022, 2024). This process requires specific operating conditions particularly in terms of pH (around 1–2) and temperature (between 20 and 60 °C).

The objective of this study was to investigate a novel hydrometallurgical process for the recovery of Ni and Co from New Caledonian limonites, as an alternative to HPAL. This process was based on acidic reductive leaching. This study also investigated the efficiency of a step to pre-concentrate Co- and Ni-bearing Mn oxides through mineral processing. Thanks to a detailed investigation of the mineralogy of the samples (bulk laterites and Mn oxide concentrates), the mechanisms involved in the dissolution of Ni and Co have been elucidated. The operating conditions were chosen with the consideration of applying a bio-based approach, using acidophilic microorganisms to produce the reductive agent.

## 2. Materials and methods

### 2.1. New Caledonia limonitic laterite

360 kg of limonitic laterites were sampled from the Penamax site by the geological survey of New Caledonia. Fig. S1 in the Supplementary Material shows the location of the sampling site and some photographs of the raw materials as received by BRGM before preparation. The bulk sample was dried, crushed to less than 20 mm (with Pinette) and quartered. This sample is called the ‘bulk coarse laterite’. In order to obtain a detailed characterization, 20 kg of the bulk coarse laterite were ground using a disk mill (Pulverisette 13 from Fritsch) and subsequently quartered. This sample is called the ‘bulk fine laterite’.

### 2.2. Production of the Mn oxide concentrate

An initial reductive leaching experiment highlighted the quick dissolution of Mn oxides along with the high levels of Co they host (detailed in section 3.4.1). It was thus decided to study the pre-concentration of Mn oxides before leaching. As the characterization of the bulk coarse laterite highlighted the predominance of Mn oxides in the largest size fractions (detailed in section 3.2.2), the concentrate was produced by particle size separation (Fig. 1). Gravity separation was additionally performed on an intermediate size fraction to remove chromite and thus further concentrate Mn oxides. It also aimed at removing Cr⁶⁺, as this is deleterious for micro-organisms that negatively affects acidophilic bioleaching (Santos et al., 2022).

72.5 kg of the bulk coarse laterite were added to a stirred tank to separate fine particles from larger ones. Wet sieving was performed at 2 mm, 800 µm, 400 µm, 100 µm and 50 µm. Each fraction was subsequently dried. Wet gravity separation (Mozley shaking table) was performed on the 100–400 µm fraction. For this operation, the washing water flow rate was maintained at 100 L h⁻¹ with 95 shakes per minute and 1° tilt angle. Three fractions were recovered (light, mixed and heavy fractions). After characterisation of the fractions, the concentrate was obtained by mixing the coarser size fractions (>2 mm; 800 µm - 2 mm and 400–800 µm) with the light fraction of the 100–400 µm.

From experimental results, several scenarios were assessed for the preparation of the Mn oxides concentrate:

- Scenario 1: Sieving at 50 µm, no density separation by Mozley table; the Mn oxides concentrate is composed of >50 µm particles.

- Scenario 2: Sieving at 100 µm, no density separation by Mozley table; the Mn oxides concentrate is composed of >100 µm particles.

- Scenario 3: Sieving at 400 µm, no density separation by Mozley table; the Mn oxides concentrate is composed of >400 µm particles.

- Scenario 4: Sieving at 100 µm and 400 µm, density separation of 100–400 µm fraction by Mozley table; the Mn oxides concentrate is
<figure><img src="image_2.png" /><figcaption>Fig. 1. Schematic diagram of the size fraction separation and the preparation of the bulk coarse laterite, the bulk fine laterite and the Mn oxides concentrate.</figcaption></figure>

obtained by combining the light fraction obtained after Mozley table and the >400 µm fraction;

- Scenario 5: Sieving at 50 µm and 400 µm, density separation of 50–400 µm fraction by Mozley table; the Mn oxides concentrate is obtained by combining the light fraction obtained after Mozley table and the >400 µm fraction.

For each scenario, recovery and purity yields were determined through calculations. For scenario 4, which was experimentally performed, calculated yields were compared to the experimental results.

## 2.3. Solid characterisation

To determine the particle size distribution of the bulk coarse and fine laterite samples, wet sieving was performed. For the bulk fine laterite sample, the size fractions were: >315 µm; 200–315 µm; 125–200 µm; 50–125 µm and <50 µm. For the bulk coarse laterite sample, the size fractions were: >2 mm; 800 µm - 2 mm; 400–800 µm; 100–400 µm; 50–100 µm and <50 µm. Size fractions were dried prior to any further analysis. Granulolaser (Mastersizer S, Malvern) was used to determine the fine particle size distribution (<50 µm). Solid samples were first suspended and agitated in water for 30 min–1 h using ultrasound, before the addition of 10 mL of DISPEX® (10 g.L⁻¹) and 10 mL of sodium hexametaphosphate (10 g.L⁻¹). The pulp was agitated again for 15 min with ultrasound. In the “sample presentation unit” of the analyser, 1 min of ultrasound was finally applied before the analysis (data were collected with the software Mastersizer-Sv2.18F).

Surface area measurements were determined with extended-pressure sorption analyser (ASAP 2050 from Micromeritics) using N₂ gas. Total surface area was calculated by the BET method, while the external surface area was identified with t-plot calculation (integrated in the software ASAP 2020 V1.01 used for data acquisition). Density was measured with gas displacement pycnometer (AccuPyc 1330 from Micromeritics).

Solid samples (bulk laterite, size fractions, Mn oxides concentrate

and residues) were analysed for a suite of 53 elements by Inductively Coupled Plasma Atomic Emission Spectroscopy (ICP-AES) and Inductively Coupled Plasma Atomic Mass Spectrometry (ICP-MS) at the OMAC Laboratories Ltd, Loughrea, Ireland. Elements Si, Al, Fe, Ca, Mg, Na, K, Cr, Ti, Mn, P, Ba, Ce, Cr, Cs, Dy, Er, Eu, Ga, Gd, Hf, Ho, La, Lu, Nb, Nd, Pr, Rb, Sm, Sn, Sr, Ta, Tb, Th, Tm, U, V, W, Y, Yb and Zr were analysed by ICP-MS after being prepared by lithium borate fusion. Elements Ag, As, Cd, Co, Cu, Li, Mo, Ni, Pb, Sc, Tl and Zn were analysed with ICP-AES after being prepared by 4 acid digestion (HF-HNO₃-HClO₄ HCl). Further analyses were performed for samples with a high Cr₂O₃ content (>5 wt%) in which they were prepared by sodium peroxide fusion and dissolution of the melt in HCl before ICP-AES analysis. For quality control purposes a set of standards, blanks and sample duplicates were analysed as independent laboratory checks.

X-ray powder diffraction (XRD) patterns were obtained with a PANalytical X'Pert Pro α1 MPD diffractometer equipped with an Xcelerator solid-state detector and using Co Kα radiation. Data were recorded in continuous mode over 4–90° 2θ with 0.0170 step size and 175 s counting time per step. The powders were packed into an aluminium sample deep well. Phase identifications were performed by pattern matching using the Powder Diffraction File (PDF) database of the International Centre for Diffraction Data (ICDD) and standard material from the mineral collection at the Natural History Museum.

Quantitative SEM-EDX was performed using the Oxford Instruments INCA XMax Energy Dispersive Spectrometer (EDS) on the Zeiss EVO 15LS scanning electron microscope (SEM). Objective lens to specimen working distance was kept constant at 10 mm (fixed focus). The electron beam accelerating voltage was 20 kV, and electron beam current 1.5 nA. Quant optimisation was performed on Co metal, typically every 3 h. The accuracy of EDX analysis was checked regularly at each session by collecting spectra and quantifying elemental concentrations in the reference sample of Kakanui augite (Jarosevich et al., 1980). Beam current was regularly checked throughout the analysis session. Point spectra and large area EDX maps were collected on the carbon coated polished block samples. Spectral interferences in EDX mapping, in particular Cr–Mn
and Fe–Co, were resolved by re-processing the raw EDX map data using the TruMap mapping mode function built in the Oxford Instruments AzTec® software used to process the EDX data. The TruMap function provides background and peak overlap corrections using spectral deconvolution techniques.

Electron microprobe analysis (EMPA) was performed using wavelength-dispersive X-ray spectrometry (WDX) on a Cameca SX100 electron microprobe, operating at 20 kV accelerating voltage, 20 nA current and a 1 µm spot size. Standards used were a combination of natural minerals for Na (jadeite), Mg (olivine), Al (corundum), Ca (wollastonite), Si (fayalite) and Ti (rutile); synthetic compounds for Ti and Mn (MnTiO₃), P (ScPO₄), K (KBr), Fe (FeO), Ni (NiO₂) and pure metals (Co). All data were matrix-corrected using the Cameca version of the PAP PhiRhoZ program (Pouchou and Pichoir, 1984a, 1984b). In this study, a detection limit of 0.03 wt% for Co was employed in EMPA analyses of goethite/hematite in order to reliably account for the impact of the Fe induced background (Dybowska et al., 2022).

Automated mineralogical analysis (AMA) was carried out using a Tescan Integrated Mineral Analyser (TIMA)-X system with the SEM TIMA3 GMU equipped with a field emission source (FEG) and SDD-EDS X-ray spectrometer (EDAX Element 30) combined with software TESCAN TIMA version 2.2.0. The operating conditions were as follows: beam energy 25 kV, working distance 15 mm, pixel size for both BSE imaging and EDS data acquisition was 5 µm, acquisition mode: high resolution mapping, measurement type: modal analysis, X-ray acquisition time per sample: 2 h, total area scanned per sample: 308 mm². All data were processed using a standard mineral spectra library customised with our own reference spectra.

X-ray absorption near edge structure (XANES) spectroscopy was performed using beamline B18 at Diamond Light Source Ltd, UK (Dent et al., 2009). The energy was selected using a Si(111) monochromator and the energy resolution and beam size were optimised using Pt coated collimating and focusing Si mirrors, an additional pair of flat mirrors was used to remove high energy harmonics. Samples were mixed with cellulose in proportions suitable to produce an edge step between 0.5 and 1.2. These mixtures were prepared as pressed pellets and mounted in transmission mode. XANES data were collected with the pellets mounted vertically either side of the Izero and Itrans ion chambers. Spectra were recorded in Qexafs mode from 5796 to 6795.5 eV at a step size of 0.25 eV with 3 spectra recorded for each sample and merged. The energy was calibrated by setting the position of the first maximum of the first derivative of the XANES spectrum of a Cr foil to 5989 eV. Spectra were background subtracted and normalised using Athena (Ravel and Newville, 2005) and the pre-edge features extracted by fitting the resulting background to a spline function. The pre-edge features were fitted using Gaussian components.

## 2.4. Leaching of limonitic laterites

### 2.4.1. Experimental set up

Leaching experiments were performed in batch mode in laboratory-scale glass reactors with a working capacity of 500 mL or 2 L. The stirred tank reactors (STR) were jacketed for warm water circulation to maintain a constant operating temperature of 46 °C. The temperature of the circulating water was controlled by a cryothermostat and thermocouples placed in the reactors. The reactors were equipped with four baffles mounted 90° apart and extended down to the base of the vessel to optimise the mixing of pulp. Agitation was performed using a dual impeller system (axial/radial) consisting of a standard 6-blade Rushton turbine combined with a 3-blade 45° axial flow impeller. The gas supply system was designed to accommodate N₂. The gas was injected beneath the turbine at the bottom of the reactors via a stainless steel pipe. The impellers and the gas injection pipe were positioned in order to respect the standard dimensions and thus to optimise gas mass transfer and mixing in the reactors. The top of the reactors was connected to a gas cooling system to prevent excessive evaporation.

### 2.4.2. Operating conditions

All leaching experiments were performed at 46 °C with a solid concentration of 5 % (w/v). Both bulk fine laterite and Mn oxides concentrate were used in these experiments. In order to have reducing conditions compatible with microbial activity that may occur during acidophilic bioleaching, nutrients were added to the leaching solution. The nutrient medium was made of Acidophilic Basal Salts (ABS) and Trace Elements (TE) (Nancucheo et al., 2016) with 100 µmol.L⁻¹ ferrous sulfate and a pH adjusted to 1.5.

First, the nutrient medium was added to the reactor and N₂ was bubbled at 0.5 L min⁻¹ to obtain oxygen-free conditions. Then, the solid was added and the gas injection was modified to only flush the top of the reactor with N₂ (248 mL min⁻¹) to avoid the presence of oxygen in the pulp. The stirring speed was closed to 500 rpm.

In Fe(II) leaching experiments, Fe(II) concentration was adjusted to the desired level by FeSO₄·7H₂O addition in the nutritive medium before the addition of the solid. In SO₃²⁻ leaching experiments, the redox potential was maintained lower than 600 mV vs Standard Hydrogen Electrode (SHE) by manual addition of a solution of 60 g.L⁻¹ Na₂SO₃. Leaching experiments lasted 2 days. Table 1 summarises the main operating conditions.

### 2.4.3. Leaching monitoring

pH values and redox potentials were measured with a frequency that was adjusted according to the kinetics (measurements were performed every 15 min during the first hour and every hour during the first 8 h). pH regulation was performed manually with H₂SO₄ 96 %. To determine the Fe(III) to Fe(II) concentration ratio from the redox potential measurement, the correlation from Yue et al. (2014) was used. Metal concentrations (Fe, Ni, Co, Mn, Cr and Al) in the liquid phase were determined by microwave plasma atomic emission spectroscopy (MP-AES) using a MP-AES 4210 from Agilent Technologies.

## 3. Results

### 3.1. Characterisation of bulk limonitic laterite

#### 3.1.1. Physical characterisation

Particle size distributions were determined (Fig. S2 in Supplementary Material) for both the bulk coarse and fine laterite samples and showed that a large majority of particles, around 95 %, were under 50 µm. The particle size distribution in both the bulk coarse and fine samples was similar with or without disk milling, except for the presence of large particles in the bulk coarse laterite.

The density and specific area were determined for the bulk fine laterite. The former was 3.8 g cm⁻³ and the latter reached 69.3 m² g⁻¹, which is typical for fine particle materials. There was no major difference between the specific area and the external specific area (61.3 m² g⁻¹), indicating that this material was not porous. Senanayake et al. (2011) reported similar specific areas for other laterite samples (between 39.8 and 55.1 m² g⁻¹ for synthetic goethite spiked with Co or Ni and 91.5 m² g⁻¹ for a laterite ore), despite the fact that they didn't report any particle size distribution.

<table><caption>**Table 1**

Operating conditions of the leaching experiments performed in this study.</caption><thead><tr><th>Solid</th><th>Condition</th><th>Reductive reagent</th></tr></thead><tbody><tr><td>Bulk fine laterite</td><td>Acid leaching<br>SO<sub>3</sub><sup>2-</sup><br>leaching<br>Fe(II)<br>leaching</td><td>–<br>Addition of a solution of 60 g.L<sup>-1</sup> Na<sub>2</sub>SO<sub>3</sub> to maintain the redox to less than 600 mV vs SHE<br>3 g.L<sup>-1</sup> added initially in the leach mixture</td></tr><tr><td>Mn oxides concentrate</td><td>Fe(II)<br>leaching</td><td>25 g.L<sup>-1</sup> added initially in the leach mixture</td></tr></tbody></table>
### 3.1.2. Chemical and mineralogical composition

The chemical composition of the bulk fine limonite was determined by ICP-AES and ICP-MS (Table 2). The sample was very rich in Fe (almost 70 wt% $Fe_2O_3$) and contained a significant concentration of Mn (>1 wt% MnO). Amongst trace elements Co and Ni were enriched with Ni concentrations (1.35 wt%) almost ten times higher than Co (1640 ppm). Very low Si concentrations ($SiO_2$ below 3 wt%) were detected indicating that the silicate (quartz, phyllosilicates, clay minerals) content of these samples was very low. The content of Mg and Ca was also low indicating a lack of carbonates, ferromagnesian silicates and serpentines.

The XRD pattern of the bulk fine laterite sample revealed it to be almost pure (poorly crystalline) goethite with only a trace of gibbsite detected (see Fig. S3 in the supplementary material). The bulk mineralogy of the sample agrees well with the bulk chemistry data which suggested no silicate or carbonate phases should be present in significant quantities.

SEM-EDX maps of the bulk fine laterite sample (Fig. 2) confirmed the sample was Fe rich. Fe rich particles, identified mainly as poorly crystalline goethite by XRD (Fig. S3), were ubiquitous in the sample and SEM imaging revealed this phase to have no distinctive morphology, often appearing as a porous mass, while EDX analysis indicated it contained traces of Si, Al and Ni. Manganese rich phases (identified as Mn oxyhydroxides) were observed as small inclusions or as very fine veins within the goethite. In addition, chromite grains were also identified with SEM-EDX. Two types of Mn oxyhydroxides were detected in the sample (Fig. 2): hydrated cryptomelane (K-bearing) and asbolane-lithiophorite intermediates (Al-bearing).

The Al-bearing Mn oxyhydroxides were Co enriched with an average Co to Ni ratio of 2.4 (Fig. S4 in Supplementary Material). Cobalt concentrations varied from 3.78 to 8.58 wt% (element) while Ni varied from 1.48 to 5.37 wt% (element). There was significant Fe also detected in this phase with concentrations ranging from 0.83 to 15.3 wt%. Cryptomelane was analysed with EMPA to measure the Co content which was found to vary from 0.13 wt% to 0.34 wt% with an average of 0.27 wt%. Iron-rich grains were also measured with EMPA and a very small proportion of these grains were identified as hematite, while the majority were Fe oxyhydroxide (goethite). Average concentration of Co in the Fe rich grains was found to be 0.05 wt% in goethite and 0.03 wt% in hematite. Nickel concentrations in goethite varied significantly from grain to grain and were generally much higher than for Co ranging from 0.03 wt% to 1.62 wt% while averaging 0.12 wt% in hematite. In addition, significant concentrations of Al (6 wt% $Al_2O_3$), Si (2 wt% $SiO_2$), Mn (0.8 wt% MnO) and Cr (0.8 wt% $Cr_2O_3$) were also detected in the goethite grains.

The abundance of different mineral phases in the bulk fine laterite sample was investigated using AMA. Fe oxyhydroxides were the most abundant phase, accounting for 94.6 mass% in the sample, Mn

oxyhydroxides accounted for 2.87 mass%, chromite 2.44 mass% and gibbsite 0.09 mass%. Within the Mn oxyhydroxide category, the K-bearing Mn oxyhydroxide was much less abundant than the Al-bearing one with the former accounting only 0.07 mass%. In the Fe oxyhydroxide category, a distinction was made between Ni-bearing and non-bearing grains, with the former accounting for 54.94 mass%.

## 3.2. Characterisation of the size fractions

### 3.2.1. Metal distribution

The metal distribution through the size fractions was determined for both the bulk fine and coarse laterite samples (Fig. 3A and B).

In the bulk fine laterite, Mn and Co contents increased (from 0.67 wt % to 14.5 wt% MnO and from 530 mg kg⁻¹ to 2.26 wt% Co) while Fe content decreased (from 71 wt% down to 21 wt% $Fe_2O_3$) with increasing particle size. Cr was mainly concentrated in the intermediate size fractions with concentrations increasing from 2.09 wt% to a maximum of 27.5 wt% $Cr_2O_3$ in the 100–200 µm fraction. Ni distribution was almost constant across all size fractions and varied from 0.9 to 1.4 wt%. Similar evolution of Fe content within the size fractions was observed in the bulk coarse laterite sample, with Fe content decreasing (from 71 wt% down to 24 wt% $Fe_2O_3$) with increasing particle size. Co, Ni and Mn contents were maximum in the 800 µm–2 mm fraction (3.11 wt% Co, 1.82 wt% Ni and 20.10 wt% as MnO respectively).

Other metals of strategic relevance are being considered significant within limonitic laterites around the world (e.g. Dybowska et al., 2022; Putzolu et al., 2019; Teitler et al., 2019; Chassé et al., 2017) and the distribution of these across the size fractions of the bulk fine laterite has also been monitored (Fig. 3C). Sc concentrations decreased as the size fraction increased, from 56 mg kg⁻¹ in the <50 µm fraction to 20 mg kg⁻¹ in the >315 µm fraction. Conversely, Li, Cu, Zn and La concentrations increased from <10 mg kg⁻¹ to 200 mg kg⁻¹ (Li), 52 mg kg⁻¹ to 187 mg kg⁻¹ (Cu), 414 mg kg⁻¹ to 806 mg kg⁻¹ (Zn) and 10 mg kg⁻¹ to 30 mg kg⁻¹ (La) as the particle size increased from <50 µm to >315 µm fraction. The Sc distribution varied linearly with the $Fe_2O_3$ content (Fig. S5, supplementary material), suggesting that its mineralogical deportment was almost entirely associated with the Fe-oxides (primarily goethite in this Penamax limonite) similar to previous studies (Teitler et al., 2019; Chassé et al., 2017). The concentration of Li increased linearly with MnO across the size fractions (Fig. S5) indicating that it was entirely hosted by the Mn-oxyhydroxides (asbolane-lithiophorite intermediates). Cu and La both decreased as a function of $Fe_2O_3$ and increased with MnO content (Fig. S5) suggesting that while they were likely hosted by both goethite and Mn-oxyhydroxides, these metals were more strongly partitioned into the Mn-oxyhydroxides. The partitioning of Zn was similar in that it was much more concentrated in the Mn-oxyhydroxides than in the goethite. However, there seems to be a large fraction of the Zn associated with $Cr_2O_3$ potentially being deported

<table><caption>**Table 2**

Chemical composition of the bulk fine laterite sample with major elements reported as wt% oxides and trace elements in mg.kg⁻¹ (ppm). LOI is the loss on ignition.</caption><thead><tr><th>Element/oxide</th><th>wt %</th><th>Element</th><th>ppm</th><th>Element</th><th>ppm</th><th>Element</th><th>ppm</th></tr></thead><tbody><tr><td>SiO<sub>2</sub></td><td>2.15</td><td>Co</td><td>1640</td><td>Ho</td><td>0.06</td><td>Tb</td><td>0.04</td></tr><tr><td>Al<sub>2</sub>O<sub>3</sub></td><td>5.5</td><td>Ni</td><td>13450</td><td>La</td><td>1.1</td><td>Th</td><td>0.05</td></tr><tr><td>Fe<sub>2</sub>O<sub>3</sub></td><td>69.6</td><td>Ag</td><td>1.6</td><td>Li</td><td>10</td><td>Tl</td><td>&lt;10</td></tr><tr><td>CaO</td><td>0.01</td><td>As</td><td>&lt;5</td><td>Lu</td><td>0.03</td><td>Tm</td><td>0.03</td></tr><tr><td>MgO</td><td>0.47</td><td>Ba</td><td>35.8</td><td>Nb</td><td>&lt;0.02</td><td>U</td><td>0.15</td></tr><tr><td>Cr<sub>2</sub>O<sub>3</sub></td><td>2.95</td><td>Cd</td><td>&lt;0.5</td><td>Nd</td><td>1.2</td><td>V</td><td>157</td></tr><tr><td>MnO</td><td>1.19</td><td>Ce</td><td>5.3</td><td>Mo</td><td>&lt;1</td><td>W</td><td>2</td></tr><tr><td>Na<sub>2</sub>O</td><td>0.02</td><td>Cs</td><td>&lt;0.01</td><td>Pb</td><td>&lt;2</td><td>Y</td><td>1.5</td></tr><tr><td>K<sub>2</sub>O</td><td>0.01</td><td>Cu</td><td>56</td><td>Pr</td><td>0.29</td><td>Yb</td><td>0.21</td></tr><tr><td>TiO<sub>2</sub></td><td>0.02</td><td>Dy</td><td>0.39</td><td>Rb</td><td>0.4</td><td>Zn</td><td>437</td></tr><tr><td>SrO</td><td>&lt;0.01</td><td>Er</td><td>0.19</td><td>Sc</td><td>59</td><td>Zr</td><td>2</td></tr><tr><td>BaO</td><td>&lt;0.01</td><td>Eu</td><td>0.1</td><td>Sm</td><td>0.34</td><td></td><td></td></tr><tr><td>P<sub>2</sub>O<sub>5</sub></td><td>0.01</td><td>Ga</td><td>4.3</td><td>Sn</td><td>1</td><td></td><td></td></tr><tr><td>LOI</td><td>16.1</td><td>Gd</td><td>0.4</td><td>Sr</td><td>1.4</td><td></td><td></td></tr><tr><td>Total</td><td>98.03</td><td>Hf</td><td>&lt;0.2</td><td>Ta</td><td>0.2</td><td></td><td></td></tr></tbody></table>
<figure><img src="image_3.png" /><figcaption>**Fig. 2.** SEM-EDX large area element maps showing the texture of the Mn rich, Fe rich and Cr rich grains in the bulk fine laterite sample (Red = Mn oxyhydroxide, Blue = chromite, Green = Fe oxyhydroxide – predominantly goethite) A and B highlights K-bearing Mn oxyhydroxide and Al-bearing Mn oxyhydroxide respectively with associated Co and Ni. Element distribution maps correspond to the whole of the SEM-EDX images A and B.</figcaption></figure>

in (ferrit)chromites similar to that found by Lambiv Dzemua et al. (2013).

The partitioning of metals in the bulk coarse laterite was calculated from the mass and the metal content of each size fraction (Fig. 3D). These results highlight that more than 60 % Co was contained in the coarser fractions (>50 µm) that represented around 6 % of the total mass of the sample. These fractions were also characterised by a high content of Cr. In particular, the 100–400 µm fraction concentrated around 20 % of the total Cr contained in the sample, and Cr concentration reached almost 20 wt% in this fraction. Cr and Mg partitioning are very similar, and different to Si, suggesting that most of the Mg budget is in chromite rather than in residual Mg-silicates. Additionally, the close similarity between Si and Fe partitioning supports the possibility that Si occurs as a trace component within goethite.

### 3.2.2. Mineral distribution

In addition to analysing bulk chemistry, XRD and AMA analyses were performed on the size fractions of both bulk fine and coarse laterites (see Fig. 4 for AMA and Fig. S6 in Supplementary Material for XRD). A significant reduction in the intensity of the goethite peaks was observed with an increase in particle size with goethite identified as the major phase in the smallest size fraction (<50 µm) but a minor phase in all the larger size fractions (Fig. S6). Peaks characteristic to Mn oxyhydroxide (lithiophorite) were not observed in the smallest size fraction but appeared in the XRD patterns of the larger size fractions (Fig. S6). Chromite was identified in the XRD patterns of the intermediate size fractions (50–500 µm).

The significant and sequential decrease of the Fe content from the smallest (<50 µm) to the largest size fractions (>2 mm) was correlated to a decrease of goethite content observed in the XRD patterns and further confirmed in the AMA analysis (Fig. S6 and Fig. 4, respectively). Goethite abundance in the bulk fine laterite was calculated at 86 wt% in the smallest size fraction but only 22 wt% in the largest size fraction.

This is consistent with the chemical analysis given in Fig. 3A.

The significant and sequential increase of Mn and Co content from the smallest to the largest size fractions was correlated to the increase of Mn oxyhydroxide content with the particle size, as identified from XRD and AMA analysis (Fig. S6 and Fig. 4 respectively). From the AMA analysis of the size fraction samples (from <50 µm to >315 µm), the abundance of Mn oxyhydroxides was calculated at <1 mass% in the smallest size fraction and 27 mass% in the largest size fraction. Co distribution closely followed that of Mn indicating that the majority of Co is likely associated with Mn. In contrast, Ni concentrations did not systematically rise in line with an increase of Mn oxyhydroxide content in the larger size fractions. This observation suggests that a significant portion of Ni is likely to be associated with the Fe oxyhydroxides in the bulk fine laterite sample.

As observed in the bulk sample, the Mn oxyhydroxides detected in the size fractions were predominantly lithiophorite-asbolane intermediates with minor cryptomelane. The lithiophorite-asbolane intermediates were found to constitute 67–86 % of all Mn oxides/oxyhydroxides identified in the various size fractions.

### 3.2.3. Chromium speciation

The presence of hexavalent chromium in limonitic laterites has been reported from the Philippines (Delina et al., 2020) and New Caledonia, Koniambo deposit (Fandeur et al., 2009) and Penamax deposit (Santos et al., 2022). Its presence was investigated in the different size fractions. Fig. 5 shows the pre-edge region of the Cr K-edge XANES spectra for the size fraction samples. XANES analysis revealed that these features comprised two distinct components, with maxima at ~5990 eV and ~5993 eV. The normalised intensity of the peak at ~5993 eV is known to be an excellent indicator of the presence of chromium (VI) (Peterson et al., 1997; Huggins et al., 1999). The relative intensities of these two peaks varies, to a small degree, depending on the mineralogical deportment of chromium (III) between chromite, goethite and hematite,
<figure><img src="image_4.png" /><figcaption>**Fig. 3.** Fe, Al, Mn and Cr content (left) and Ni, Co, Si and Mg content (right) in the size fractions of bulk fine laterite (A) and coarse laterite (B), together with Cu, Sc, Zn, Li and La content in the bulk fine laterite (C) and metal partitioning in the bulk coarse laterite (D).</figcaption></figure>

with the latter two producing pre-edge features comprising two peaks of equal intensity (Fandeur et al., 2009). From Fig. 5, it can be seen that the peak intensity at 5993 eV of the pre-edge features was significantly greater in the <50 µm sample than for all the remaining size fraction samples. This strongly indicates that hexavalent chromium was concentrated into the <50 µm sample and its content was substantially lower in the larger size fractions.

Using beamline specific methods similar to those described by Peterson et al. (1997) and Fandeur et al. (2009), the proportion of chromium that is present as Cr (VI) in the size fractions was calculated. The percentage of Cr present as Cr⁶⁺ approached 10 % for the <50 µm sample and was 6 times less for the 50–100 µm sample (~1.5 %). The proportion of Cr⁶⁺ continually decreased as the size of the fractions increased.

### 3.3. Mn oxides concentrate

#### 3.3.1. Production of the concentrate

(Bio)processing tanks are one of the main investment costs of the STR (bio)leaching process. Beneficiation methods enable to increase ore grade and to decrease the volume of materials to be treated, which reduces the size of the tanks and consequently their costs. In this study, Mn oxides, that host the majority of Co, demonstrated high leaching

performances (see Section 3.4.1). Thus, by increasing the quantity of Mn oxides in the feed, the quantity of solid that needs to be treated is consequently decreased, driving the investment cost down. In this study, a Mn oxides concentrate was produced. Attrition was first performed to liberate the fine particles agglomerated on the coarser ones and to improve the subsequent sieving operations. The coarser fractions (>50 µm) obtained after sieving represented around 6 % of the total mass of the sample and contained around 50 % of the Mn and more than 60 % of the Co. However, these coarser fractions were also characterised by a high content of Cr hosted in chromite. In particular the fraction 100–400 µm has a high Cr content (19.2 %wt). As chromite has a high density, the 100–400 µm fraction was subjected to density separation with a Mozley table in order to remove the Cr phases.

Table 3 gives the chemical composition of the different fractions obtained after Mozley table experiments (light, mixed and heavy fractions). The Mozley table concentration step enabled the production of a light fraction containing less Cr and more Co than in the initial 100–400 µm fraction. The light fraction was also enriched in Mn and Ni compared to the initial one.

Table 4 gives the Mn, Co, Cr and Ni contents of the Mn oxides concentrate with respect to each scenario (outlined in section 2.2), as well as the theoretical enrichment factors (ratio between the content in the concentrate and the initial content) and the recovery yield (ratio
<figure><img src="image_5.png" /><figcaption>**Fig. 4.** Distribution of Fe oxides/oxyhydroxides and Mn oxyhydroxides in various size fractions of the bulk fine laterite in randomly selected fields from the whole sample mosaic maps. On the BSE images, grains identified as Fe oxides/oxyhydroxides are coloured in orange and grains identified as Mn oxyhydroxides are coloured in pink. Note that in the <50 µm sample the small particles have agglomerated together to give the initial appearance of larger grains. Scale bar 1 mm.</figcaption></figure>

between the mass of metal recovered in the concentrate and the mass of metal in the initial sample). Table S1 in the supplementary material provides the theoretical composition of the concentrate in each scenario, calculated from the data of Fig. 3 and Table 3. The final selection of the process design was based on a compromise between the enrichment factor and the recovery yield. The best options for Mn oxides concentration were Scenarios 3 and 4 with Mn enrichment factors of 14.0 and 12.5 respectively, and recovery yields of 29 % and 42 % respectively. Scenario 4, that combines the particles larger than 400 µm with the light fraction obtained from the Mozley table of the 100–400 µm fraction, was chosen as the best option as it avoided losing too much Mn compared to scenario 3.

The procedure was applied to a 72.5 kg bulk coarse laterite sample,

leading to the production of 2.4 kg of Mn oxides concentrate. The composition of the Mn oxides concentrate is given in Table 5 and is very close to the calculated theoretical composition shown in Table 4. Enrichment factors reached 12.8 and 13.8 and recovery yields 43 % and 60 % for Mn and Co, respectively, and 4.4 and 1.2 for Cr and Ni. Moreover, the percentage of Cr as Cr⁶⁺ was 0.72 % ± 0.05 (Fig. 5), which shows that the concentration processing enabled a huge decrease in the Cr⁶⁺ content.

### 3.3.2. Mineralogy of the Mn oxides concentrate

The XRD analysis of the Mn oxides concentrate indicated the following phases present: Mn oxyhydroxide (lithiophorite and lithiophorite-asbolane intermediates), gibbsite, aluminian chromite,
<figure><img src="image_6.png" /><figcaption>**Fig. 5.** Comparison of the extracted pre-edge features of the normalised Cr K-edge XANES spectra for the limonitic laterite size fractions (left) and of the Cr⁶⁺ percentages calculated from the Cr XANES pre-edge features across the size fraction samples of bulk fine laterite (right). Error was calculated as the uncertainty from the least squares peak fitting converted to %Cr⁶⁺.</figcaption></figure>

<table><caption>**Table 3**

Mass fraction (in wt%) and metal content (in wt%) from the Mozley table fractions.</caption><thead><tr><th rowspan="2"></th><th rowspan="2">Mass fraction (wt%)</th><th colspan="8">Content (in wt%)</th></tr><tr><th>Fe</th><th>Al</th><th>Mn</th><th>Cr</th><th>Mg</th><th>Ni</th><th>Co</th><th>Si</th></tr></thead><tbody><tr><td>100–400 µm (before Mozley)</td><td>–</td><td>24.6</td><td>7.3</td><td>5.4</td><td>19.2</td><td>2.7</td><td>1.0</td><td>1.0</td><td>0.5</td></tr><tr><td>Light fraction from 100 to 400 µm</td><td>54.4</td><td>24.3</td><td>7.9</td><td>9.3</td><td>12.7</td><td>1.8</td><td>1.6</td><td>1.7</td><td>0.7</td></tr><tr><td>Mixed fraction from 100 to 400 µm</td><td>21.2</td><td>25.1</td><td>6.3</td><td>1.0</td><td>26.7</td><td>3.6</td><td>0.2</td><td>0.2</td><td>0.4</td></tr><tr><td>Heavy fraction from 100 to 400 µm</td><td>24.4</td><td>27.8</td><td>6.1</td><td>0.5</td><td>26.3</td><td>3.6</td><td>0.1</td><td>0.1</td><td>0.2</td></tr></tbody></table>

<table><caption>**Table 4**

Calculated metal content (in wt%) of the enriched material, enrichment factor and recovery yield obtained for each concentration scenario.</caption><thead><tr><th rowspan="2"></th><th rowspan="2">Mass fraction (wt%)</th><th colspan="4">Content (in wt%)</th><th colspan="4">Enrichment factor</th><th colspan="2">Recovery yield (%)</th></tr><tr><th>Mn</th><th>Co</th><th>Cr</th><th>Ni</th><th>Mn</th><th>Co</th><th>Cr</th><th>Ni</th><th>Mn</th><th>Co</th></tr></thead><tbody><tr><td>Scenario 1</td><td>5.5</td><td>7.9</td><td>1.5</td><td>13.5</td><td>1.2</td><td>8.6</td><td>9.3</td><td>6.7</td><td>0.9</td><td>47</td><td>52</td></tr><tr><td>Scenario 2</td><td>4.4</td><td>8.9</td><td>1.7</td><td>12.5</td><td>1.3</td><td>9.7</td><td>10.6</td><td>6.2</td><td>0.9</td><td>43</td><td>47</td></tr><tr><td>Scenario 3</td><td>2.1</td><td>12.9</td><td>2.6</td><td>5.0</td><td>1.5</td><td>14.0</td><td>15.7</td><td>2.5</td><td>1.1</td><td>29</td><td>33</td></tr><tr><td>Scenario 4</td><td>3.3</td><td>11.5</td><td>2.2</td><td>7.9</td><td>1.6</td><td>12.5</td><td>13.7</td><td>3.9</td><td>1.2</td><td>42</td><td>46</td></tr><tr><td>Scenario 5</td><td>3.8</td><td>11.5</td><td>2.1</td><td>7.4</td><td>1.6</td><td>12.5</td><td>13.1</td><td>3.7</td><td>1.2</td><td>47</td><td>49</td></tr></tbody></table>

<table><caption>**Table 5**

Chemical composition of the Mn oxides concentrate obtained from sieving and density separation compared to the initial bulk coarse laterite.</caption><thead><tr><th rowspan="2"></th><th rowspan="2">Mass fraction (wt%)</th><th colspan="7">Content (in wt%)</th></tr><tr><th>Fe</th><th>Al</th><th>Mn</th><th>Cr</th><th>Mg</th><th>Ni</th><th>Co</th><th>Si</th></tr></thead><tbody><tr><td>Bulk coarse laterite</td><td>100</td><td>48.7</td><td>2.9</td><td>0.9</td><td>2.0</td><td>0.3</td><td>1.3</td><td>0.2</td><td>1.0</td></tr><tr><td>Mn oxides concentrate</td><td>3.3</td><td>20.3</td><td>9.2</td><td>11.8</td><td>8.8</td><td>1.2</td><td>1.7</td><td>2.3</td><td>0.4</td></tr></tbody></table>

goethite and hematite. Mn oxyhydroxide, gibbsite and chromite were the most abundant. The mineral composition of the sample is in agreement with the chemical analysis and confirms concentration of Mn (in the form of Mn oxyhydroxide), Cr (in the form of chromite) and Al (in the form of gibbsite and chromite) in the sample, with a substantial depletion of Fe. By contrast to the bulk limonite, Fe oxides/oxyhydroxides were identified as minor phases.

The sample was further analysed using SEM-EDX where EDX maps and point spectra were collected to probe the chemical composition of the phases identified. Mn, Fe and Cr were the most abundant elements detected. Fig. 6A reveals abundant grains of chromites, Fe oxides/oxyhydroxides and Mn oxyhydroxides.

In Fig. 6B, a large grain of Mn oxyhydroxide is shown with corresponding distribution maps of Mn, Fe, Co, Al and Ni within the grain. Fe and Si distribution overlaps indicating that Fe was likely to be present as Fe oxyhydroxide inclusions in the Mn oxyhydroxide grain. Co-localisation of Co, Ni and Mn further confirmed the association of

these elements within Mn oxyhydroxide.

All Mn oxides analysed were hydrated. K-containing Mn oxyhydroxide was identified alongside Al-containing Mn oxyhydroxide (Fig. S7 in Supplementary Material). Aluminium concentrations in the latter varied from 3.4 wt% to 13.4 wt%. The major elements detected in this phase (in addition to Mn and Al) were Co (1.1–9.6 wt%), Ni (1.1–14.3 wt%) and Fe (0.4–16 wt%). Co/Ni ratio varied from 0.1 to 4.8. Other elements detected included Ca, K, and Si but their concentrations were all below <0.5 wt%. For the K-bearing Mn oxyhydroxides, elements detected in addition to Mn (43–53 wt%) and K (1.8–3.5 wt%) were Al (0.45–2.4 wt%), Fe (0.44–4.33 wt%), Co (0–2.4 wt%) and Ni (0.31–6.81 wt%). Based on the chemistry data, the K-bearing phase can be identified as hydrated cryptomelane while the Al-bearing phases can be identified as primarily lithiophorite-asbolane intermediates, with a very few grains having Al content high enough for the phase to be classified as lithiophorite. The K-bearing phase was found to be in a low abundance with the majority of grains probed classified as lithiophorite.
<figure><img src="image_7.png" /><figcaption>**Fig. 6.** SEM-EDX montaged map (A) of the sample showing the distribution of Mn (blue), Fe (red) and Cr (yellow) over the BSE image (left) and individual element distribution maps for K, Co and Ni (right). The potassium distribution map relates to the K-bearing Mn oxyhydroxide. Co and Ni are co-localised with Mn in Al-bearing Mn oxyhydroxide in the sample; and (B) of an individual Mn oxyhydroxide grain showing the distribution of Fe (red), Mn (blue) and Al (green) (left) and elemental maps of Al, Mn, Fe, Co, Ni and Si (right).</figcaption></figure>

asbolane intermediates. Average Co and Ni concentrations were signif-
icantly lower in the K-bearing Mn oxyhydroxide (0.72 wt% Co and 2.57
wt% Ni) compared with the Al-bearing phase (5.77 wt% Co and 4.1 wt%
Ni).

AMA was used to investigate the modal mineralogy and abundance
of the mineral phases in the concentrate sample. In agreement with the
XRD and SEM-EDX data, the AMA analysis revealed the presence of the
following phases: Mn oxyhydroxides of 2 types: Al- and K-bearing with
the latter being much less abundant (38.09 mass% Al-bearing and 0.94
mass% K-bearing Mn oxyhydroxides), chromite, gibbsite and Fe oxides/
oxyhydroxides (Fig. S8 in Supplementary Material). Chromite and Al-
bearing Mn oxyhydroxides were the most abundant phases consti-
tuting 38.4 and 39.0 mass% respectively, followed by Fe oxides/oxy-
hydroxides at 17.3 mass% and gibbsite at 5.16 mass%.

**3.4. Leaching experiments**

Leaching experiments were carried out on both the bulk fine laterite
and the Mn oxides concentrate.

3.4.1. Bulk fine laterite leaching

3.4.1.1. Metal dissolution. Three leaching experiments were performed on the bulk fine laterite at 5 % (w/v) solids, at 46 °C and pH around 1.5: in the presence of sulfuric acid only, with Fe(II) added initially or in the presence of continuously added SO₃²⁻.

In the presence of the reductive reagents, the redox potential was
maintained lower than 700 mV vs SHE while it reached around 1000 mV
vs SHE in the presence of acid only (data not shown). This means that Fe
was only Fe(III) in the acidic medium, while it was mainly Fe(II) in the
reductive experiments.

In the absence of a reductive reagent, dissolution rates were

<figure><img src="image_8.png" /><figcaption>**Fig. 7.** Metal dissolution kinetics during leaching of the bulk fine laterite at pH 1.5. Experiments were performed with Na₂SO₃ (added to maintain a redox of 600 mV vs SHE), initial Fe(II) (3 g.L⁻¹) or acid only (error bars correspond to the relative standard deviation between a triplicate).</figcaption></figure>
relatively low for all metals (Fig. 7 and Table 6). A large difference between acid only and reductive conditions was observed for Co and Mn, as dissolution yields reached around 80 % after 2 days in the presence of a reductive reagent, irrespective of its nature, against less than 5 % without the reductive reagent. The dissolution rates reached 7.57 mg kg⁻¹·h⁻¹ and 1.25 mg kg⁻¹·h⁻¹ for Mn and Co respectively during the first hour of the Fe(II) experiment. The Fe(II) reductive leaching experiment was performed in triplicate and gave a high reproducibility for metal dissolution kinetics, as seen by the very small error bars.

Cr dissolution yields reached around 3 % in the presence of a reductive reagent, while Al dissolution yields reached around 15 % (see Fig. S9 in supplementary material). Ni dissolution was also higher in the presence of a reductive reagent compared to only an acid environment. Two dissolution phases can be distinguished: a very fast increase of Ni dissolution yield in the first 6 h, followed by a slower increase.

At pH 1.5 (acid only), the Fe dissolution rate was low and the increase of total Fe concentration reached around 40 mg·L⁻¹ after 2 days (Fig. S9 in Supplementary Material). In the presence of Fe(II) in the initial medium, the total Fe concentration did not change over time. As the initial concentration of Fe is around 3 g·L⁻¹, the low increase of Fe concentration is within the range of uncertainty of the analysis and might be limited due to Fe precipitation. In the presence of SO₃²⁻, the Fe concentration increased over time up to 200 mg·L⁻¹. The final yield of Fe dissolution however remained very low, reaching 0.9 % after 2 days. Pregnant leach solution (PLS) compositions are given in Supplementary Material Table S2.

### 3.4.1.2. Reagent consumption.
Table 6 gives the cumulative proton and reductive reagent consumptions in the three leaching experiments. It revealed that acid leaching consumes less protons compared to reductive leaching. However, metal dissolution yields were very low. SO₃²⁻ leaching consumed much more protons and reductive reagent than Fe (II) leaching. This is mainly due to the fact that SO₃²⁻ reacts with acid to produce SO₂, which is involved in the metal dissolution. It should be noted that SO₃²⁻ consumption may be overestimated due to the equilibrium between SO₃²⁻ and SO₂ in acidic environments followed by SO₂ release from the reactor due to the injection of N₂.

By contrast to dissolution kinetics, there were large variations in reagent consumptions in the triplicate of Fe(II) leaching, even though the pH and redox potential were almost the same in the triplicate (data not shown – less than 4 % of relative standard deviation). A relative deviation of 15 % and 18 % was observed for Fe(II) and proton consumption respectively. The mean Fe(II) requirement was evaluated to be 0.30 mol Fe(II)·kg⁻¹ ore. Characterisation revealed a content of 2.87 wt % of Mn oxyhydroxides in fine laterite bulk sample (see paragraph 2.1.2) which amounts to around 0.18 mol of Mn oxyhydroxides per kg of ore (mean value of molar mass of lithiophorite and asbolane). These experiments thus approached the theoretical stoichiometry of the reaction between Fe(II) and Mn oxides (2 mol of Fe are required for the dissolution of 1 mol of Mn).

<caption>**Table 6**</caption>

Final metal dissolution yields, cumulative proton and reductive reagent consumptions during leaching experiments of bulk fine laterite.

<table><thead><tr><th></th><th>Acid leaching</th><th>Fe(II) leaching</th><th>SO<sub>3</sub><sup>2-</sup> leaching</th></tr></thead><tbody><tr><td>Final Fe dissolution yield (%)</td><td>0.2</td><td>n.d.</td><td>0.92</td></tr><tr><td>Final Mn dissolution yield (%)</td><td>1.9</td><td>83.9</td><td>88.1</td></tr><tr><td>Final Co dissolution yield (%)</td><td>1.9</td><td>79.1</td><td>82.4</td></tr><tr><td>Final Ni dissolution yield (%)</td><td>0.4</td><td>8.8</td><td>9.4</td></tr><tr><td>Proton consumption (mol H<sup>+</sup>·kg<sup>-1</sup> ore)</td><td>0.69</td><td>1.39±0.25</td><td>8.12</td></tr><tr><td>Reductive reagent consumption (mol reagent·kg<sup>-1</sup> ore)</td><td>–</td><td>0.30±0.05</td><td>2.79</td></tr></tbody></table>

### 3.4.1.3. Residues characterisation.
The residue of the experiment with SO₃²⁻ was analysed with XRD, ICP-AES and SEM-EDX. As for the initial bulk fine laterite sample, the XRD pattern of the leached residue was dominated by goethite and minimal differences were observed between the two patterns (Fig. S3) other than a lower background for the leached residue. This lower background is likely associated with the removal of Mn (Mn fluoresces under Co radiation) that was confirmed with SEM-EDX and ICP-AES analysis (see below and Fig. S10).

The bulk chemical composition of this residue was Fe₂O₃: 68.30 wt %, Al₂O₃: 4.30 wt%, SiO₂: 2.25 wt%, Cr₂O₃: 2.89 wt%, MnO: 0.19 wt% and Ni 1.26 wt%. At the trace level, Co: 86 mg kg⁻¹, Sc: 44 mg kg⁻¹, Li: <10 mg kg⁻¹, Cu: 53 mg kg⁻¹, Zn: 446 mg kg⁻¹ and La: 10 mg kg⁻¹. The main elements detected in the residue with SEM-EDX were Fe, Si, Cr, Mg and Al (Fig. S10). EDX spectra collected from the Fe and Cr rich particles indicated these were Fe oxyhydroxides with associated Si, Al and chromites. Fe oxyhydroxide grains were analysed in more detail by extracting EDX spectra from a number of particles with a particular focus on the content of trace elements detectable by EDX. 51 grains were analysed and Ni was detected in 90 % of these grains with concentrations ranging from 0.82 wt% to 1.53 wt%.

These chemical and mineralogical analyses of the residue imply a minimal goethite dissolution, with no major alteration of its structure as evidenced from the XRD pattern, and to the removal of the Mn phases during reductive leaching. This is consistent with the dissolution yields calculated from the liquid phase. The low removal of Ni and Sc supports the association of both elements with goethite as proposed in section 3.2.1. The Cu, La and Zn concentrations in the residue are very similar to those in the <50 µm fraction which is overwhelmingly dominated by goethite (>99 % Fe- and Fe-Cr oxides) as described in section 3.2.2. The Ni content and goethite dominated mineralogy of this residue suggests that while having the vast majority of the Co and Mn removed it can still be considered a Ni-rich limonitic resource and processed/leached accordingly to extract the 1.26 wt% Ni and associated bi-products such as Sc (e.g. du Plessis et al., 2011; Hallberg et al., 2011).

### 3.4.2. Mn oxides concentrate leaching

Leaching was performed on the Mn oxides concentrate to evaluate the composition of the leachate as well as the leaching kinetics and the Fe(II) requirement. Results were compared to those obtained from the bulk fine laterite. As the Mn oxides concentrate was substantially richer in Mn oxyhydroxides (around 12-fold), the initial Fe(II) concentration was increased to 25 g·L⁻¹ rather than 3 g·L⁻¹ as used with the bulk fine laterite to ensure an excess of Fe(II).

Co and Mn dissolution reached higher yields with the Mn oxides concentrate compared to the bulk fine laterite (Fig. 8A and C). The final Co and Mn concentrations in the liquid phase reached 1 g·L⁻¹ and 6 g·L⁻¹ respectively (compared to 66 mg·L⁻¹ and 392 mg·L⁻¹ respectively with the bulk fine laterite, see Table S2 in Supplementary Material). In a similar manner, Ni dissolution yield reached up to 80 % with the concentrate material, while it was limited to less than 10 % with the bulk fine laterite (Fig. 8B and C). Al dissolution yield was also enhanced to reach around 40 % during Mn oxides concentrate leaching. Dissolution kinetics were similar between both solids, even if the Mn oxides concentrate particles were larger than the bulk fine laterite.

An increase in Fe concentration due to Fe dissolution was not observed during these experiments (Fig. S11 in Supplementary Material). Indeed, Fe concentration actually decreased during Mn oxides concentrate leaching, probably because of Fe precipitation. In this experiment, Fe(II) consumption reached 8 g·L⁻¹ after 2 days, thus 10 times higher than the required Fe(II) for bulk fine laterite (0.8 g·L⁻¹). The cumulative acid consumption reached 14.4 mol H⁺·kg⁻¹ concentrate, which is also around 10 times higher than the consumption noticed during the bulk fine laterite leaching experiment. It should be highlighted that, even if this high reagent consumption increases the cost of the leaching, it is counterbalanced by the fact that the resulting Mn oxides concentration in the leachate is more than 10 times higher
<figure><img src="image_9.png" /><figcaption>**Fig. 8.** (A) Co, Mn, (B) Al and Ni dissolution yields over time during Fe(II) leaching experiments on the Mn oxides concentrate and bulk fine laterite, together with (C) final dissolution yields.</figcaption></figure>

than the leachate of the bulk fine laterite.

## 4. Discussion

Bulk fine laterites were leached in the presence of sulfuric acid. The kinetics of Fe oxyhydroxide dissolution was slow. The analysis of the mineralogical composition of this sample revealed that Fe oxyhydroxides were mainly goethite. The dissolution of goethite under acid leach with $H_2SO_4$ has been studied by several authors (e.g. Landers et al., 2007, Landers et al., 2009; Liu et al., 2009; Senanayake et al., 2011) who showed that goethite solubilisation was dependent on the operating conditions such as acid strength and temperature but also the properties of the mineral itself and in particular the substitution of Cr and/or Al in the goethite structure. Liu et al. (2009) studied dissolution of various minerals in the lateritic ore with $H_2SO_4$ acid concentrations ranging from 0.5 mol.L$^{-1}$ to 7.5 mol.L$^{-1}$ and temperature ranging from 50 to 105 °C. Enhanced dissolution of goethite was observed with increasing acid concentration and temperature, with goethite dissolving completely with 2.5 mol.L$^{-1}$ acid and at 80 °C. Senanayake et al. (2011) also reported similar results from a temperature range of 25–80 °C. Under the mild leaching conditions used in this study i.e. low temperature and acid strength, dissolution of goethite is expected to be low. In addition, the substitution of Al and/or Cr for Fe in the goethite structure has been shown to stabilise it against proton attack i.e. acid dissolution (Schwertmann, 1991). This is because the dissolution behaviour of a mineral can be related to the metal oxide bond strength with Al–O and Cr–O bond strengths higher than that of Fe–O (Landers et al., 2009). Our characterisation data indicate Cr associated with goethite in the New Caledonia laterite studied. The Cr-enrichment of goethite might thus hinder its acid dissolution, as previously observed by Landers et al. (2009) in $H_2SO_4$ leaching experiments at 40 °C.

The presence of a reducing agent ($SO_3^{2-}$) enabled a shift in the equilibrium of acid dissolution of goethite towards its dissolution, due to the reduction of Fe(III) into Fe(II), but only to a small extent. The final yield of Fe dissolution remained very low, reaching 0.9 % after 2 days, compared to 0.2 % in acid leaching experiments.

The characterization of the laterite identified the other main phases as predominantly Al-bearing Mn oxyhydroxides (asbolane-lithiophorite intermediates) together with a K-bearing Mn oxyhydroxides (cryptomelane). During the bulk fine laterite leaching, these Mn oxyhydroxides dissolved very quickly in the presence of reductive reagents, which is consistent with a previous study by Li et al. (2011). Co dissolution kinetics are correlated with Mn ones.

Mineralogical analysis of the size fractions highlighted that Fe oxyhydroxide content increased when the particle size decreased, while Mn oxyhydroxide content decreased. Co content followed Mn content in the size fractions, suggesting that Co was mainly hosted by Mn oxyhydroxides. In contrast, Ni concentrations did not systematically rise in line with an increase of Mn oxyhydroxides content in the larger size fractions. These observations suggested that a significant portion of Ni was likely to be associated with the Fe oxyhydroxides in the bulk fine laterite sample.

The low extraction yield of Ni in leaching experiments and high content of Ni in the leached residue confirmed that a significant portion of Ni was associated with goethite. The quantity of dissolved Ni was compared to the quantity of Fe and Mn dissolved in the reductive leaching experiments (Fig. 9). Two distinct reaction periods are visible: the first one lasted few hours and corresponds to a very quick release of Ni and Fe with a molar ratio of 1. The second one is also linear but with a molar ratio of 0.0141. The relation between released Ni and Mn was also linear, with a molar ratio of 0.15. In the first period the quantity of Ni dissolved is associated with quickly dissolving Mn oxyhydroxides, while the Ni-bearing Fe oxyhydroxides dissolved at a much slower rate. This is consistent with the study from Luo et al. (2015).

The slopes of the curves of metal dissolution enable determination of the molar content of Ni in goethite, together with the molar content of Fe, Co and Ni in Mn oxyhydroxides. These concentrations were denoted $x^{\text{leaching}}$ and were compared to the molar content $x^{\text{EMPA}}$ calculated from the mass content $w^{\text{EMPA}}$ determined during the bulk fine laterite characterisation (see section 3.1.2). As an example, the molar content of Ni in goethite was estimated with the following Eq. (5):

$$x_{\text{Ni in goethite}}^{\text{EMPA}} = w_{\text{Ni in goethite}}^{\text{EMPA}} \cdot \frac{M_{\text{goethite}}}{M_{\text{Ni}}} \quad [5]$$

Where $w_{\text{Ni in goethite}}$ refers to the mass content of Ni in goethite grains (determined with EMPA analysis to be between 0.03 wt% and 1.62 wt %), $M_{\text{goethite}}$ and $M_{\text{Ni}}$ to the molar mass of goethite and Ni (in g.mol$^{-1}$).

The molar content of Ni in goethite calculated from the slope of the curve of Ni vs Fe released (Fig. 9) is 1.41 mol%, within the range from EMPA analyses (0.045–2.45 mol%). This content is also consistent with the characterization of the leaching residue (section 3.4.1.3): the EMPA analysis detected 0.82 to 1.53 wt% Ni in the goethite (i.e. 1.24–2.32 mol % Ni in goethite) while the bulk chemistry revealed a content of 1.26 wt % Ni in the residue (i.e. 1.91 mol% Ni if the residue is only composed with goethite).

In the case of Ni and Fe in Mn oxyhydroxides, a mean molar mass of
<figure><img src="image_10.png" /><figcaption>Fig. 9. Dissolved Ni as a function of dissolved Mn (A) and Fe (B), and dissolved Co as a function of dissolved Mn (C) during leaching experiments in the reductive conditions.</figcaption></figure>

lithiophorite and asbolane was used as the Mn oxyhydroxides molar mass. Approximation based on the hypothesis that Ni was contained mainly in Al-bearing Mn oxyhydroxides (as the K-bearing Mn oxyhydroxide was much less abundant than the Al-bearing one, see section 3.1.2) enabled an estimation of the molar content of Ni in Mn oxyhydroxides. As Ni mass content varied from 1.48 wt% to 5.37 wt%, Ni molar content was estimated between 4.0 mol% and 14.7 mol%. This value is consistent with the leaching experiments results (slope of the curves of Ni vs Mn released) where the values of 15.2 mol% and 14.5 mol% were obtained for $SO_3^{2-}$ and Fe(II) experiments respectively (Fig. 9).

For Fe, the molar content in Mn oxyhydroxides was estimated from mineralogical characterisation to be between 2.38 mol% and 43.9 mol% (mass contents ranging from 0.83 to 15.3 wt%, see section 3.1.2). The value determined from leaching experiments (15 mol%) is consistent with this range. For Co, the molar content in Mn oxyhydroxides was estimated from mineralogical characterisation to be around 16.3 mol% (6 wt%), while the values determined from leaching experiments are 16.6 mol% (in the case of Fe(II) leaching) and 17.9 mol% (in the case of $SO_3^{2-}$ leaching).

Generally, the results from the leaching experiments and mineralogical analysis were in the same order of magnitude. The differences observed between both methods likely arise from the difficulty of obtaining ideal values with EMPA analysis considering the characterisation challenges posed by the bulk material and uncertainties surrounding the nature of the deportment of Fe within the Mn oxyhydroxides.

Co and Mn contents are consistent with those presented by Dublet et al. (2017) on the regolith from New Caledonia. The Co content of 1640 ppm, as well as that of Mn of 0.92 wt%, corresponds, according to their study, to a regolith sampling depth of approximately 30 m. Sampling depth was unknown in our study. At that depth, they describe the presence of chromite, goethite, asbolane/lithiophorite and gibbsite, which is perfectly consistent with our results. Dublet et al. (2017) used X-ray Absorption Near Edge Structure spectroscopy (XANES) and

deduced that about 15 % of the Mn was contained in the goethite, as well as 30 % of the Co. In our study, the distribution of these elements, deducted from both analysis and leaching tests, is consistent with their results. The combination of mineralogical analytical techniques and leaching tests therefore provided a comprehensive view of the phases that is equivalent to that of Dublet et al. (2017), without requiring the use of XANES (synchrotron), which is an expensive and difficult-to-access equipment. Our results also allowed to further characterize the sample, by determining Ni content and its distribution.

As the leaching experiment revealed quick dissolution of Mn oxyhydroxides compared to goethite, the separation of Mn phases before leaching was assessed. Due to the presence of very large particles (>800 µm) of Mn oxyhydroxides in the bulk coarse laterite, sieving was performed. Density separation was added to the process, in order to further concentrate Mn oxides by removing part of the chromite. Both methods are cheap, robust and easy to implement. From a mineralogical point of view, the production of a Mn oxides concentrate significantly reduced the content of Fe oxyhydroxides in the sample. It enabled an increase in the Co content from 0.16 % to 2.3 %. The volume/mass of sample for leaching was also drastically decreased since the mass of the final sample was only 3.3 % of the initial one. Both features might lead to a significant reduction of the process costs (both CAPEX and OPEX): on one hand, less volume of materials to be treated means smaller leaching reactors and lower energy consumption. On the other hand, higher Mn oxides content in the materials to be treated means more concentrated leach liquor and, subsequently a reduction of costs associated with downstream processing facilitating increased efficiency in terms of recovery yield. Mozley table is a suitable device to assess density separation with a sample below 100 kg but is less efficient than other devices that are used at larger scales. Consequently, for larger amounts of material better separation efficiency might be reached using other equipment for density separation (such as shaking tables). Moreover, any such process, performed at the pilot scale, should be capable of operating in continuous mode to ensure that its application at an industrial scale would be compatible with the properties of the pulp (pumping, piping systems).
The leaching of the Mn oxides concentrate demonstrated high Co, Ni and Mn recovery yields. As the content of Fe oxyhydroxides was significantly reduced, most of the Ni in the concentrate was associated with Mn oxyhydroxides. The final Ni dissolution yield in Mn oxides concentrate leaching experiment was thus highly improved compared to the bulk fine laterite. Co and Mn dissolution yields from the Mn oxides concentrate were also improved. Such approach supports the conclusions of Stanković et al. (2022), revealing that goethite-poor ores are amenable to reductive (bio)processing but goethite-rich ores treatment still require some improvements. The PLS resulting from Mn oxides concentrate leaching has significant metal concentrations (>1 g.L⁻¹ in Co, around 700 mg.L⁻¹ in Ni and 6 g.L⁻¹ in Mn). In their work, Zheng and Cheng (2007) concluded that an effluent containing 1–5 g.L⁻¹ Mn could offer a growing low cost resource of manganese from Ni laterite. Such concentrations also offer opportunities for Co and Ni purification and recovery, even if current extraction technologies are generally dedicated for PLS where Ni concentrations are higher than Co ones (Cheng et al., 2010).

Reductive leaching of Mn oxides has been studied extensively but has never found industrial application, particularly due to the cost of reagents and the difficulty in treating solutions when using Fe(II), due to the presence of high concentrations of Fe in the PLS. Hence, there is a growing interest in using bacteria that avoid the addition of Fe(II) to the system and allow the control of Fe oxidation state. Reductive bioleaching is one of the metallurgical options that are currently studied to extract Co and Ni from limonite (Johnson et al., 2022). While the dissolution reaction of goethite and manganese oxide are chemical phenomena, the role of the microorganisms is to provide Fe(II) by catalysing the reduction of Fe(III), as described in Eq. (6).

$$6Fe^{3+} + S^{0} + 4H_{2}O \xrightarrow{\text{bacteria}} 6Fe^{2+} + SO_{4}^{2-} + 8H^{+} \quad \Delta G_{25^{\circ}\text{C}} = -241.4 \text{ kJ.mol}^{-1} \quad [6]$$

The concept of reductive bioleaching was first introduced by Hallberg et al., in 2011 and since then several studies dealing with the proof of concept of reductive bioleaching have been released (e.g. Johnson et al., 2022) and show varying Co and Ni dissolution yields. The results obtained in our study, in reductive conditions that are compatible with bioleaching operating conditions, may help to better understand the Co and Ni recovery yields. It demonstrates that bioleaching could be an option for the efficient recovery of Co from Mn oxides in New Caledonian laterites.

In order to use bioleaching, it is however important to consider Cr. Chromate ions have a significant deleterious impact upon bioreductive leaching of limonites (Johnson et al., 2017; Santos et al., 2022). XANES analysis demonstrated that the Cr(VI) was not contained in the Mn oxides concentrate. As the main phase removed during the production of the concentrate is goethite (which resides mostly in the <50 µm fraction), it strongly suggests that the Cr(VI) is associated with the goethite

in the original laterite. This is consistent with the study from Santos et al. (2022), which proposed that the chromate within the limonitic laterite from Penamax was primarily associated with goethite particle surfaces. The preparation of a Mn oxides concentrate has removed the majority (or all) of the chromate, which is an important factor when considering bioprocessing.

To conclude, in the past, mining operators from New Caledonia mainly focused on Ni recovery from saprolites only. Co recovery, which was little considered, may benefit from the development of new processes. This study demonstrates, with small-scale tests, high Co recovery through the use of mineral processing and reductive atmospheric leaching. This could be considered as a first step for the treatment of limonites, before Ni recovery (Fig. 10). This process benefits from the very limited dissolution of Fe oxyhydroxides, thus minimising reagent consumption. The use of mineral processing also enables removal of part of the goethite before leaching, thus limiting the quantity of solid residues.

Such process may be of interest in a context broader than New Caledonia limonites. In particular, a large number of limonites deposits have similar structure and mineralogy, Ni being associated to ferric iron minerals whereas Co is hosted by Mn minerals (Santos and Schippers, 2023).

This article provides a proof-of-concept for a new process dedicated to Co and Mn recovery. Further development of this process could focus in particular on increasing the solid concentration during leaching and on enhanced metal recovery from the PLS. The leaching was performed with a limited solid concentration (5 % w/v). It is assumed that increasing the solid concentration to 20 % (w/v), as often performed in leaching operations, will not affect the dissolution kinetics if reagent concentrations are linearly increased. Such results were already observed for other (bio)leaching case studies (Joulian et al., 2024). It is also necessary to study the scale-up of the process (both mineral processing and leaching) and reagent regeneration could also be tested. A life-cycle assessment of the developed process would also be necessary to fully evaluate the benefits of this solution compared to HPAL and other Ni and Co production process.

## 5. Conclusion

Reductive leaching is one of the metallurgical options that are studied to extract Co and Ni from limonites. In this view, a limonitic ore from New-Caledonia was finely characterised and Co and Ni leaching experiments were performed.

Leaching in the presence of sulfuric acid revealed a low goethite dissolution rate. It was shown to be higher by reduction compared to acid dissolution with other conditions being equal, but the final yield of Fe dissolution remained very low, reaching 1.9 % after 2 days. Ni was partly hosted by Fe oxyhydroxides, thus leading to limited Ni dissolution (9 %). However, a very fast increase of Ni dissolution yield was observed

<figure><img src="image_11.png" /><figcaption>Fig. 10. Proposed integrated processing for Co and Ni recovery from New Caledonian limonites.</figcaption></figure>
in the first 6 h of leaching, related to the small proportion of Ni hosted by Mn oxyhydroxides that were quickly dissolved in the presence of a reductive reagent, irrespective of its nature. Mn and Co dissolution rates were very fast with dissolution yields reached around 80 % after 2 days.

From these observations, the acidic reductive leaching process could be considered for Co recovery as a first step for the treatment of limonites, before recovery of the bulk of the Ni. It benefits from very limited dissolution of Fe oxyhydroxides, thus limiting reagent consumption.

In order to further improve the process, a Mn oxides concentrate was produced by attrition, sieving and density separation (using a Mozley table). It enabled an increase in the Co content from 0.16 wt% to 2.3 wt %. The volume/mass of sample for leaching was drastically decreased since the mass of the final sample was only 3.3 % of the initial one. Enrichment factors reached 13.8 and 12.8 and recovery yields 60 % and 43 % for Co and Mn, respectively. It showed that even if Cr was enriched in the concentrate, less than 0.75 % of the Cr was hexavalent. The leaching of the Mn oxides concentrate showed high dissolution rate for Co and Mn, reaching dissolution yields of 87 % and 95 % after 2 days respectively. Ni dissolution yield reached up to 80 % with the concentrate material, while it was limited to less than 10 % with the bulk laterite. As the Mn oxides concentrate was substantially richer in Mn oxyhydroxides, Ni in the concentrate was mainly associated with these minerals and was much more amenable to dissolution than Ni from the bulk laterite. The production of a Mn oxides concentrate proved highly effective, enabling the production of a Co-rich leachate.

### CRediT authorship contribution statement

**Agathe Hubau:** Writing – original draft, Visualization, Validation, Methodology, Investigation. **Agnieszka Dybowska:** Writing – review & editing, Visualization, Investigation. **Paul F. Schofield:** Writing – review & editing, Visualization, Supervision, Investigation, Funding acquisition. **Diego Gianolio:** Writing – review & editing, Investigation.

**Ana Santos:** Writing – review & editing. **Richard J. Herrington:** Writing – review & editing. **Douglas Pino-Herrera:** Writing – review & editing. **Hafida Tris:** Investigation. **Marion Erard:** Investigation. **Anne-Gwénaëlle Guezennec:** Writing – review & editing, Supervision, Project administration, Funding acquisition.

### Funding

This work is part of the European collaborative research project CROCODILE (First of a kind commercial compact system for the efficient recovery of cobalt designed with novel integrated leading technologies), funded by the European Union’s Horizon2020 research and innovation programme under grant agreement No 776473.

### Declaration of competing interest

The authors declare the following financial interests/personal relationships which may be considered as potential competing interests: Agathe Hubau reports financial support was provided by Horizon (2020). Diego Gianolio reports equipment, drugs, or supplies was provided by Diamond Light Source Ltd. If there are other authors, they declare that they have no known competing financial interests or personal relationships that could have appeared to influence the work reported in this paper.

### Acknowledgments

We are indebted to colleagues at the Service Géologique de Nouvelle Calédonie (S. Lesimple) for preparing and providing the limonite samples.

This work was carried out with the support of Diamond Light Source, instrument B18 (proposal SP32381).

## Appendix A. Supplementary data

Supplementary data to this article can be found online at https://doi.org/10.1016/j.apgeochem.2025.106569.

### Data availability

Data will be made available on request.

### References

Bailly, L., Picault, M., Juillot, F., Lesimple, S., 2021. Les Minéralisations Cobaltiferes de Nouvelle Calédonie, Rapport Intermédiaire V3. BRGM/RP-71261-FR, p. 44. In French.

Bartzas, G., Komnitas, K., 2024. Cradle to gate life-cycle assessment of battery grade nickel sulphate production through high-pressure acid leaching. Sci. Total Environ. 952, 175902. https://doi.org/10.1016/j.scitotenv.2024.175902.

Chassé, M., Griffin, W.L., O'Reilly, S.Y., Calas, G., 2017. Scandium speciation in a world-class lateritic deposit. Geochem. Perspect. Lett. 3, 105–114.

Cheng, C.Y., Boddy, G., Zhang, W., Godfrey, M., Robinson, D.J., Pranolo, Y., Zhu, Z., Wang, W., 2010. Recovery of nickel and cobalt from laterite leach solutions using direct solvent extraction: part 1 — selection of a synergistic SX system. Hydrometallurgy 104, 45–52. https://doi.org/10.1016/j.hydromet.2010.04.009.

Dent, A.J., Cibin, G., Ramos, S., Smith, A.D., Scott, S.M., Varandas, L., Pearson, M.R., Krumpa, N.A., Jones, C.P., Robbins, P.E., 2009. B18: a core XAS spectroscopy beamline for diamond. J. Phys. Conf. Ser. 190, 012039. https://doi.org/10.1088/1742-6596/190/1/012039.

Delina, R.E., Arcilla, C., Otake, T., Garcia, J.J., Tan, M., Ito, A., 2020. Chromium occurrence in a nickel laterite profile and its implications to surrounding surface waters. Chem. Geol. 558, 119863. https://doi.org/10.1016/j.chemgeo.2020.119863.

du Plessis, C.A., Slabbert, W., Hallberg, K.B., Johnson, D.B., 2011. Ferredox: a biohydrometallurgical processing concept for limonitic nickel laterites. Hydrometallurgy 109, 221–229. https://doi.org/10.1016/j.hydromet.2011.07.005.

Dublet, G., Juillot, F., Brest, J., Noël, V., Fritsch, E., Proux, O., Olivi, L., Ploquin, F., Morin, G., 2017. Vertical changes of the Co and Mn speciation along a lateritic regolith developed on peridotites (New Caledonia). Geochem. Cosmochim. Acta 217, 1–15. https://doi.org/10.1016/j.gca.2017.07.010.

Dybowska, A., Schofield, P.F., Newsome, L., Herrington, R.J., Mosselmans, J.F.W., Kaulich, B., Kazemian, M., Araki, T., Skiggs, T.J., Kruger, J., Oxley, A., Norman, R.A., Lloyd, J.R., 2022. Evolution of the Piaui laterite, Brazil: mineralogical, geochemical and geomicrobiological mechanisms for cobalt and nickel enrichment. Minerals 12, 1298. https://doi.org/10.3390/min12101298.

European Commission, 2023. Study on the Eu's Fifth List of Critical Raw Materials – Final Report.

Fandeur, D., Juillot, F., Morin, G., Olivi, L., Cognigni, A., Webb, S.M., Ambrosi, J.-P., Fritsch, E., Guyot, F., Brown Jr., J.E., 2009. XANES evidence for oxidation of Cr(III) to Cr(VI) by Mn-Oxides in a lateritic regolith developed on serpentinized ultramafic rocks of New Caledonia. Environ. Sci. Technol. 43, 7384–7390. https://doi.org/10.1021/es900498r.

Fraser, J., Anderson, J., Lazuen, J., Lu, Y., Heathman, O., Brewster, N., Bedder, J., Masson, O., 2021. Study on Future Demand and Supply Security of Nickel for Electric Vehicle Batteries. Publications Office of the European Union, Luxembourg. https://doi.org/10.2760/212807, 2021, ISBN 978-92-76-29139-8, JRC123439.

Freyssinet, P., Butt, C.R.M., Morris, R.C., Piantone, P., 2005. Ore-forming processes related to lateritic weathering. In: Hedenquist, Jeffrey W., Thompson, John F.H., Goldfarb, Richard J., Richards, Jeremy P. (Eds.), One Hundredth Anniversary Volume. Society of Economic Geologists. https://doi.org/10.5382/AV100.21.

Hallberg, K.B., Grail, B.M., du Plessis, C., Johnson, D.B., 2011. Reductive dissolution of ferric iron minerals: a new approach for bioprocessing nickel laterites. Miner. Eng. 24, 620–624. https://doi.org/10.1016/j.mineng.2010.09.005.

He, F., Ma, B., Qiu, Z., Wang, C., Chen, Y., Hu, X., 2023. Enhanced extraction of nickel from limonitic laterite via improved nitric acid pressure leaching process. Miner. Eng. 201, 108170. https://doi.org/10.1016/j.mineng.2023.108170.

Huggins, F.E., Najih, M., Huffman, G.P., 1999. Direct speciation of chromium in coal combustion by-products by X-ray absorption fine-structure spectroscopy. Fuel 78 (2), 233–242. https://doi.org/10.1016/S0016-2361(98)00142-2.

Jarosevich, E., Nelen, J.A., Norberg, J.A., 1980. Reference samples for electron microprobe analysis. Geostand. Newsl. 4, 43–47. https://doi.org/10.1111/j.1751-908X.1980.tb00273.x.

Johnson, D.B., Bryan, C.B., Schlomann, M., Roberto, F.R., 2022. Biomining technologies: extracting and recovering metals from ores and wastes. https://doi.org/10.1007/978-3-031-05382-5.

Johnson, D.B., du Plessis, C., 2015. Biomining in reverse gear: using bacteria to extract metals from oxidized ores. Miner. Eng. 75, 2–5. https://doi.org/10.1016/j.mineng.2014.09.024.

Johnson, D.B., Hedrich, S., Pakostova, E., 2017. Indirect redox transformations of iron, copper and chromium catalyzed by extremely acidophilic bacteria. Front. Microbiol. 8, 211. https://doi.org/10.3389/fmicb.2017.00211.

Joulian, C., Hubau, A., Pino-Herrera, D., Guezennec, A.G., 2024. Bioleaching of polymetallic sulphidic mining residues: influence of increasing solid concentration on microbial community dynamics and metal dissolution. Res. J. Microbiol. 175 (1–2), 104112. https://doi.org/10.1016/j.resmic.2023.104112.
Kommitsas, K., Petrakis, E., Bartzas, G., 2023. A novel and greener sequential column leaching approach for the treatment of two different Greek laterites. Sci. Total Environ. 854, 158748. https://doi.org/10.1016/j.scitotenv.2022.158748.

Lambiv-Dzemua, G., Gleeson, S.A., Schofield, P.F., 2013. Mineralogical characterization of the Nkamouna Co-Mn laterite ore, southeast Cameroon. Miner. Depos. 48, 155-171.

Landers, M., Gilkes, R.J., 2007. Dehydroxylation and dissolution of nickeliferous goethite in New Caledonian lateritic Ni ore. Appl. Clay Sci. 35, 162-172. https://doi.org/10.1016/j.clay.2006.08.012.

Landers, M., Gilkes, R.J., Wells, M., 2009. Dissolution kinetics of dehydroxylated nickeliferous goethite from limonitic lateritic nickel ore. Appl. Clay Sci. 42, 615-624. https://doi.org/10.1016/j.clay.2008.05.002.

Li, G., Rao, M., Jiang, T., Huang, Q., Peng, Z., 2011. Leaching of limonitic laterite ore by acidic thiosulfate solution. Miner. Eng. 24, 859-863. https://doi.org/10.1016/j.mineng.2011.03.010.

Liu, K., Chen, Q., Hu, H., 2009. Comparative leaching of minerals by sulphuric acid in a Chinese ferruginous nickel laterite ore. Hydrometallurgy 98, 281-286. https://doi.org/10.1016/j.hydromet.2009.05.015.

Llorca, S., Monchoux, P., 1991. Supergene cobalt minerals from New Caledonia. Can. Mineral. 29, 149-161.

Luo, J., Li, G., Rao, M., Peng, Z., Zhang, Y., Jiang, T., 2015. Atmospheric leaching characteristics of nickel and iron in limonitic laterite with sulfuric acid in the presence of sodium sulfite. Miner. Eng. 78, 38-44. https://doi.org/10.1016/j.mineng.2015.03.030.

McDonald, R.G., Whittington, B.I., 2008. Atmospheric acid leaching of nickel laterites review. Part I. Sulphuric acid technologies. Hydrometallurgy 91, 35-55. https://doi.org/10.1016/j.hydromet.2007.11.009.

MineralInfo. https://www.mineralinfo.fr/fr/substances/cobalt-co, 2020-. (Accessed 2 July 2025).

MineralInfo. https://www.mineralinfo.fr/fr/substance/nickel-ni, 2024-. (Accessed 2 July 2025).

Moro, K., Haubrich, F., Martin, M., Grimmer, M., Hoth, N., Schippers, A., 2023. Reductive leaching behaviour of manganese and cobalt phases in laterite and manganese ores. Hydrometallurgy 220, 106101. https://doi.org/10.1016/j.hydromet.2023.106101.

Mudd, G.M., Jowitt, S.M., 2014. A detailed assessment of global nickel resource trends and endowments. Econ. Geol. 109, 1813-1841. https://doi.org/10.2113/econgeo.109.7.1813.

Nancucheo, I., Rowe, O.F., Hedrich, S., Johnson, D.B., 2016. Solid and liquid media for isolating and cultivating acidophilic and acid-tolerant sulfate-reducing bacteria. FEMS Microbiol. Lett. 363. https://doi.org/10.1093/femsle/fnw083.

Nasab, M.H., Noaparast, M., Abdollahi, H., Amoozegar, M.A., 2020. Kinetics of two-step bioleaching of Ni and Co from iron rich-laterite using supernatant metabolites produced by Salinivibrio kushneri as halophilic bacterium. Hydrometallurgy 195, 105387. https://doi.org/10.1016/j.hydromet.2020.105387.

Newsome, L., Arguedas, A.S., Coker, V.S., Boothman, C., Lloyd, J.R., 2020. Manganese and cobalt redox cycling in laterites; biogeochemical and bioprocessing implications. Chem. Geol. 531, 119330. https://doi.org/10.1016/j.chemgeo.2019.119330.

Oxley, A., Smith, M.E., Caceres, O.Y., 2016. Why heap leach nickel laterites? Miner. Eng. 88, 53-60. https://doi.org/10.1016/j.mineng.2015.09.018.

Pell, R., Tijsseling, L., Goodenough, K., Wall, F., Dehaine, Q., Grant, A., Deak, D., Yan, X., Whattoff, P., 2021. Towards sustainable extraction of technology materials through integrated approaches. Nat. Rev. Earth Environ. 2, 665-679. https://doi.org/10.1038/s43017-021-00211-6.

Peterson, M.L., Brown Jr., G.E., Parks, G.A., Stein, C.L., 1997. Differential redox and sorption of Cr(III/VI) on natural silicate and oxide minerals: EXAFS and XANES results. Geochem. Cosmochim. Acta 61, 3399-3412. https://doi.org/10.1016/S0016-7037(97)00165-8.

Pouchou, J.L., Pichoir, F., 1984a. Application to the analysis of homogeneous samples. Rech. Aerosp. 3, 13-38.

Pouchou, J.L., Pichoir, F., 1984b. A new model for quantitative X-ray microanalysis, part II: application to In-depth analysis of heterogeneous samples. Rech. Aerosp. 5, 47-65.

Putzolu, F., Boni, M., Mondillo, N., Maczurad, M., Pirajno, F., 2019. Ni-Co enrichment and high-tech metals geochemistry in the Wingellina Ni-Co oxide-type laterite deposit (Western Australia). J. Geochem. Explor. 196, 282-296.

Ravel, B., Newville, M., 2005. ATHENA, ARTEMIS, HEPHAESTUS: data analysis for X-ray absorption spectroscopy using IFEFFIT. J. Synchrotron Radiat. 12, 537-541. https://doi.org/10.1107/S0909049505012719.

Requin, A., Nguyen, A., Monteil, S., Le Moine, D., Desfossés, 2023. Avenir de la filière du nickel en Nouvelle-Calédonie, Inspection Générale des Finances. Conseil Général de l'Economie 72 in French.

Santos, A.L., Dybowska, A., Schofield, P.F., Herrington, R.J., Cibin, G., Johnson, D.B., 2022. Chromium (VI) inhibition of low pH bioleaching of limonitic nickel-cobalt ore. Front. Microbiol. 12, 802991. https://doi.org/10.3389/fmicb.2021.802991.

Santos, A.L., Dybowska, A., Schofield, P.F., Herrington, R.J., Johnson, D.B., 2020. Sulfur-enhanced reductive bioprocessing of cobalt-bearing materials for base metals recovery. Hydrometallurgy 195, 105396. https://doi.org/10.1016/j.hydromet.2021.105396.

Santos, A.L., Schippers, A., 2023. Reductive mineral bioprocessing. In: Johnson, D.B., Bryan, C.G., Schlömann, M., Roberto, F.F. (Eds.), Biomining Technologies. Springer, Cham. https://doi.org/10.1007/978-3-031-05382-5_15.

Schwertmann, U., 1991. Solubility and dissolution of iron oxides. Plant Soil 130, 1-25. https://doi.org/10.1007/BF00011851.

Senanayake, G., Childs, J., Akerstrom, B.D., Pugaev, D., 2011. Reductive acid leaching of laterite and metal oxides — a review with new data for Fe(Ni,Co)OOH and a limonitic ore. Hydrometallurgy 110 (1-4), 13-32. https://doi.org/10.1016/j.hydromet.2011.07.011.

Stanković, S., Goldmann, S., Kraemer, D., Ufer, K., Schippers, A., 2024. Bioleaching of a lateritic ore (Piauí, Brazil) in percolators. Hydrometallurgy 224, 106262. https://doi.org/10.1016/j.hydromet.2024.106262.

Stanković, S., Martin, M., Goldmann, S., Gäbler, H.-E., Ufer, K., Haubrich, F., Moutinho, V.F., Giese, E.C., Neumann, R., Stropper, J.L., Stummeyer, J., Kaufhold, S., Dohrmann, R., Oxley, A., Marbler, H., Schippers, A., 2022. Effect of mineralogy on Co and Ni extraction from Brazilian limonitic laterites via bioleaching and chemical leaching. Miner. Eng. 184, 107604. https://doi.org/10.1016/j.mineng.2022.107604.

Teitler, Y., Cathelineau, M., Ulrich, M., Ambrosi, J.P., Munoz, M., Sevin, B., 2019. Petrology and geochemistry of scandium in New Caledonian Ni-Co laterites. J. Geochem. Explor. 196, 131-155.

Yue, G., Zhao, L., Olvera, O.G., Asselin, E., 2014. Speciation of the H2SO4-Fe2(SO4) 3-FeSO4-H2O system and development of an expression to predict the redox potential of the Fe3+/Fe2+ couple up to 150 °C. Hydrometallurgy 147-148, 196-209. https://doi.org/10.1016/j.hydromet.2014.05.008.

Zheng, W., Cheng, C.Y., 2007. Manganese metallurgy review. Part I: leaching of ores/ secondary materials and recovery of electrolytic/chemical manganese dioxide. Hydrometallurgy 89 (3-4), 137-159. https://doi.org/10.1016/j.hydromet.2007.08.010.