Unlocking Refractory Gold: Synergistic Pretreatment Strategies for High-Efficiency Thiosulfate Leaching

Abstract

This study evaluates four physicochemical pretreatments—ultra-fine grinding, roasting, alkaline pressure oxidation (POX), and oxidative ammoniacal pre-leaching—for improving gold extraction from a refractory sulfide concentrate produced trough flotation. The gold extraction by direct cyanidation is only ~48.6%, mainly due to the encapsulation of gold by associated minerals. Ultra-fine grinding increased the BET surface area eight-fold but depressed gold dissolution from 74% to 18% due to accelerated thiosulfate decomposition and copper (I) passivation in the presence of a bigger surface area. Oxidative roasting at 750 °C converted pyrite–pyrrhotite to hematite without liberating additional gold, indicating limited benefit from thermal treatment. POX was conducted at 190 °C and 10 bar O₂ dissolved 33% of the solids and yielded only 26% of gold in a thiosulfate leaching step with 50% of the thiosulfate consumption. In contrast, a two-step oxidative ammoniacal conditioning (0.4 M NH₃ + 10 mM Cu²⁺ for 42 h) followed by thiosulfate leaching boosted gold extraction from 71% to 85% while cutting thiosulfate consumption from 48.4 to 29.0 kg t⁻¹. The results demonstrate that among the pretreatments investigated, oxidative ammoniacal pre-leaching provides the most effective and environmentally benign route to unlock encapsulated gold and enhance reagent efficiency for thiosulfate processing of refractory gold ore.

1. Introduction

Gold (Au) remains one of the world’s most valuable metals, and alkaline cyanidation has been used industrially for centuries because of its selectivity, and comparatively low-cost [1]. Today, approximately 85–90% of primary Au is still produced via cyanide leaching, whose rate-controlling step often is oxygen diffusion and mass-transfer [2]. The continued reliance on cyanide leaching is increasingly questioned by regulators and communities due to spills in Romania and Argentina, highlighting the ecological and social risks associated with cyanide-bearing tailings and process solutions [3]. Also, recent analyses indicate that about 24% of known Au reserves are of the refractory category, meaning that sub-microscopic gold is encapsulated within sulfide minerals and/or locked up by organic carbon and are, therefore, not available to conventional leaching without prior liberation. Sulfide minerals also consume oxygen and cyanide during cyanidation, resulting in gold recovery rates typically below 50%.

Thiosulfate (S₂O₃²⁻) has emerged as a possible candidate for replacing cyanide lixiviant due to low toxicity and benign degradation products. In ammoniacal thiosulfate leaching, the thiosulfate ligands complex is generated by (Au(I)) when cupric-ammine oxidizes metallic gold, forming the stable soluble ion [Au(S₂O₃)₂]³⁻. Ammonia stabilizes both Cu(II) and ammine complexes, while dissolved O₂ re-oxidizes Cu(I) back to Cu(II), closing the catalytic loop and sustaining dissolution [4,5,6]. The chemistry of ammoniacal thiosulfate leaching of gold is illustrated by Equations (1)–(3):

Au (s) + [Cu(NH₃)₄]²⁺ (aq) + 3 S₂O₃⁻² (aq) → [Au(NH₃)₂]⁺ (aq) + [Cu(S₂O₃)₃]⁵⁻ (aq) + 2 NH₃ (aq);

(1)

[Au(NH₃)₂]⁺ (aq) + 2 S₂O₃⁻² (aq) → [Au(S₂O₃)₂]³⁻ (aq) + 2 NH₃ (aq);

(2)

4 [Cu(S₂O₃)₃]⁵⁻ (aq) + 16 NH₃ (aq) + O₂ (g) + 2 H₂O (aq) → 4 [Cu(NH₃)₄]²⁺ (aq) + 12 S₂O₃⁻² (aq) + 4 OH⁻ (aq).

(3)

Nevertheless, thiosulfate systems are sensitive to mineralogy, copper catalysis, and polythionate formation, which influences the reagent consumption [4,6,7,8]. Research indicates that thiosulfate leaching is particularly effective for oxide ores, with high gold recovery rates [4,9]. Many researchers have also confirmed the high gold leaching efficiencies by thiosulfate from refractory ores [10,11,12]. In contrast, sulfide ores present a significant challenge, as their reactive surfaces catalyze the thiosulfate decomposition to polythionates. Therefore, higher thiosulfate dosages are required for these refractory materials [13,14].

To improve Au recovery from sulfide-rich ores and minimize the decomposition of thiosulfate by sulfide minerals during leaching, various pre-treatment methods have been studied to alter the mineralogical structure and expose the encapsulated gold [15]. Pre-treatment techniques such as chemical oxidation [16], roasting [17], ultra-fine grinding [18], bio-oxidation [19,20,21], and acidic leaching [22] have been shown to enhance gold liberation and increase thiosulfate leaching efficiency.

Oxidative roasting was commercialized in the early 20th century at the Gidji Gold Processing Plant (Australia), where fluidized-bed roasting of refractory concentrate achieved ~80–85% gold recovery at ~575 t d⁻¹ [1]. Roasting enhances gold recovery by converting iron sulfide minerals to iron oxides and sulfur dioxide, which can be captured for sulfuric acid production, and by releasing encapsulated gold from the sulfide matrix. It is typically carried out in the 600–1000 °C temperature interval, depending on process conditions and material properties [23,24,25]. Because most of the sulfur in the concentrate occurs as pyrite (FeS₂), the key reactions can be represented by the simplified equations given below; on the other hand, dominating the reactions depends on the roasting temperature and the oxidative/inert atmosphere [17,25]. The main reactions occurring during heat treatment are represented by the following equations:

4 FeS₂ (s) + 11 O₂ (g) → 2 Fe₂O₃ (s) + 8 SO₂ (g) ΔG = −775.75 kcal;

(4)

3 FeS₂ (s) + 8 O₂ (g) → Fe₃O₄ (s) + 6 SO₂ (g) ΔG = −557.46 kcal;

(5)

FeS₂ (s) + 3 O₂ (g) → FeSO₄ (s) + SO₂ (g) ΔG = −230.60 kcal;

(6)

FeS₂ (s) → FeS (s) + 1/2 S₂ (g) ΔG = 23.62 kcal;

(7)

4 FeS (s) + 7 O₂ (g) → 2 Fe₂O₃ (s) + 4 SO₂ (g) ΔG = −545.25 kcal;

(8)

2 FeS (s) → 2 Fe (s) + S₂ (g) ΔG = 67.41 kcal.

(9)

Similarly, pressure oxidation, as patented by Barrick Gold Corporation, has proven effective for treating complex ores containing refractory sulfide and carbonaceous material [26,27,28]. In these filings, the POX step is operated at 185–235 °C with PO₂ ≥ 20 psi for ~30–100 min, followed by ammoniacal thiosulfate leaching at pH 7–8.7 and 45–55 °C using 0.025–0.10 M thiosulfate, 5–50 ppm Cu²⁺, and ≥0.001 M sulfite. Under these conditions, 70–75% gold was extracted within 60–240 min [29]. Newmont Gold Co patented a heap-leach flowsheet that bio-oxidizes refractory ore using a dense culture of Thiobacillus ferrooxidans and Leptospirillum ferrooxidans and then switches to alkaline thiosulfate leaching to avoid carbonaceous preg robbing. Effluents are recirculated through the heap, and the biooxidation is tracked by pH, Eh, and Fe²⁺/Fe³⁺; after oxidation, the heap is washed and neutralized (lime or soda ash) before thiosulfate leaching. In the thiosulfate stage, the pregnant solution is recycled, and precious metals are recovered by precipitation, with the solution returned to the heap. Pilot and lab trials reported ~70% Au recovery versus ~20% by cyanidation, using pH 9.2–10.0, 0.1–0.2 M thiosulfate ((NH₄)₂S₂O₃ or Na₂S₂O₃), ≥0.1 M NH₃, and up to 60 ppm Cu(NH₃)₄²⁺ as the oxidation catalyst [30].

Ultrafine grinding (UFG) is an effective pretreatment for refractory gold ores where fine gold grains (1–20 μm) are physically locked within sulfide matrices. By producing micron- to submicron-sized particles, UFG promotes gold liberation without chemical reagents or SO₂ emissions associated with thermal treatments, making it environmentally favorable. Its industrial success has been demonstrated at operations such as Kalgoorlie in Australia [31], Kibali in South Africa [32], and Sukari in Egypt [15,33]. Ultrafine grinding performance depends strongly on parameters such as stirring speed, media size, and ball charge ratio, which vary with ore type [15]. Studies have shown that fine grinding markedly enhances gold recovery. For example, Celep et al. (2019) reported an increase from 40% to 66% extraction after grinding to 3.7 μm [34]. However, UFG mainly liberates physically locked gold and is ineffective for chemically bound forms like aurostibnite (AuSb₂). Additionally, its high energy demand, up to three times that of conventional milling, remains a key limitation [35].

Oxidative ammoniacal conditioning has been shown to be an effective pretreatment for enhancing gold extraction during thiosulfate leaching of sulfide ores [7]. Feng and van Deventer (2010) [7] reported that gold extraction increased from 69% (without pretreatment) to over 90% after 7–22 h of oxidative ammoniacal conditioning, accompanied by a substantial decline in thiosulfate usage.

Despite their effectiveness, many pre-treatment methods, such as fine milling and pressure oxidation, are associated with high operational costs, while others, such as bio-oxidation, require prolonged processing times. The optimum pretreatment for the pyrite–bearing Au ores is dependent upon the mineralogy of the ore, and conducting a comparative analysis of the processes is essential. The present study investigates the efficacy of physicochemical pretreatments, including grinding, roasting, alkaline pressure leaching, and oxidative ammoniacal conditioning, to improve gold recovery during thiosulfate leaching of a sulfide-rich flotation concentrate. By studying their impact on thiosulfate leaching kinetics, reagent efficiency, and overall gold recovery, the aim is to identify a route maximizing the Au extraction while minimizing reagent consumption and operating cost.

2. Experimental Work

2.1. Materials and Methods

The material investigated in this study is a sulfide-rich flotation concentrate supplied from a gold processing plant by Mandalay resources in northern Sweden. In routine operation, the plant first recovers free-milling gold by gravity separation, while gold finely disseminated in pyrite–pyrrhotite is upgraded using a conventional xanthate–frother flotation circuit. A bulk of flotation concentrate, with 80% of the particles passing a size of 190 µm, was collected, homogenized and riffle-split in the laboratory. The sub-samples were mineralogically characterized and used in pre-treatment and leaching experiments. The mineralogical composition of the ore was determined using X-ray diffraction (Malvern Panalytical Ltd., Almelo, The Netherlands), while the elemental composition was analyzed through ICP-OES (iCAP 7000 Series, Thermo Fisher, Waltham, MA, USA) after digestion. The results of the mineralogical and elemental analyses are presented in the discussion part. Laboratory-grade sodium thiosulfate, ammonia solution (25%), and cupric sulfate heptahydrate (99% purity) were sourced from Thermo Fisher Scientific. Milli-Q water was used for all experiments.

Elemental concentrations in the leach solutions were measured using the SPECTRO XEPOS energy dispersive X-ray fluorescence (ED-XRF) spectrometer (Spectro Ametek, Kleve, Germany). The concentration of thiosulfate was determined using an iodometric titration method. To minimize the interference of the cupric–ammonia complex during iodine titration, 10% acetic acid was added to the solution before the titration, using Vitex as an indicator. A platinum electrode was used to measure the solution’s redox potential with a double-junction Ag/AgCl reference electrode (3M KCl). All potentials were referenced against the Standard Hydrogen Electrode (SHE). The concentration of the blue cupric-ammonia complex was monitored at 609 nm using UV-Vis spectrophotometry (Du730, Beckman Coulter, Brea, CA, USA) on samples sealed in UV cells to prevent air contact.

To qualitatively characterize the morphology and chemistry of the surface of the leached mineral particles, scanning electron microscopy (SEM) combined with energy dispersive spectroscopy (EDS) (Zeiss Merlin, Oberkochen, Germany) was employed.

2.2. Leaching Experiments

Leaching experiments for the gold flotation concentrate were conducted in a 1 L glass reactor, equipped with an overhead stirrer fitted with a flat-bladed impeller. The reactor remained open to air through one sampling port. For each experiment, solid sample was added to 0.5 L of leach solution containing 0.2 M sodium thiosulfate, 0.4 M ammonia, and 10 Mm CuSO₄ reaching a pulp density of 50% (w/v). The pH was maintained at approximately 10–10.5 using ammonia, the stirring speed was kept constant at 200 rpm, and the temperature controlled at 25 °C, maintained using a temperature controller and heating mantle. The samples were collected at regular time intervals immediately analyzed by the iodine titration, UV-Vis analysis, and XRF analysis for measuring thiosulfate concentration (TS), Cu(NH₃)²⁺, and Au, respectively. All tests were performed in duplicate, and the average values are reported. The standard deviation of the results was consistently within 3%, indicating high experimental reproducibility. Au dissolution was calculated by following equation:

$$\mathrm{A}\mathrm{u} \mathrm{d}\mathrm{i}\mathrm{s}\mathrm{s}\mathrm{o}\mathrm{l}\mathrm{u}\mathrm{t}\mathrm{i}\mathrm{o}\mathrm{n} (\mathbf{\%})={\displaystyle \frac{C \ast V}{Wt \ast m}} \ast 100,$$

(10)

where C is the metal concentration in the leach solution (mg L⁻¹), V is the leachate volume (mL), Wₜ is the Au grade in the solid sample (%), and m is the mass of the solid sample (g).

2.3. Pre-Treatment Techniques

2.3.1. Grinding

To reduce the particles’ size and the number of liberate gold particles encapsulated within sulfide minerals like pyrite or pyrrhotite, the original sample was ground for 1, 3, or 6 min using a planetary ball mill (Fritsch Pulverisette 7, Idar-Oberstein, Germany). Subsequently, the particle size distribution was analyzed using a laser particle size analyzer (Malvern Mastersizer 3000, Malvern, UK), and the specific surface area was determined through BET analysis (Micromeritics Gemini VII, Norcross, GA, USA). The ground samples were subjected to thiosulfate leaching, as detailed in Section 2.2, and gold recovery to the liquid phase was determined.

2.3.2. Heat Treatment Experiments

Heat treatment of 200 g samples placed into a glass tube (Figure 1) were conducted using a gastight rotary furnace (Carbolite Gero, Hope, UK) at medium rotation speed with a gas flow of either 2 L/min of N₂ or 1.2 L/min of air. The furnace was set to a final temperature of 750 °C, with a heating rate of 5 °C/min, kept at target temperature for 2 h before cooling at a rate of 0.1 °C/min to prevent thermal shock. The furnace outgas was directed into a water solution to trap sulfur dioxide (SO₂). The resulting calcine was collected and kept in a sealed package for further analysis such as surface area, SEM, and XRD. Gold recovery from the calcine was subsequently evaluated through thiosulfate leaching.

2.3.3. Pressure Oxidative Leaching

The pressure leaching experiment was conducted in a 2 L titanium autoclave equipped with a stirrer and temperature control system (Figure 2), with the possibility to supply oxygen at the desired pressure. After adding 150 g of gold concentrate and 1 L of 90 g/L sulfuric acid (H₂SO₄) solution containing 50 g/L of Fe₂(SO₄)₃ as oxidant, oxygen was added and the system heated to 80 °C; afterwards, during a dwell time of 8 min, the oxygen pressure (pO₂) was increased to 10 bars. The stirring speed was fixed at 400 rpm. In the second step, the temperature was increased with 2 °C/min to 190 ± 3 °C, and while keeping the temperature constant for 4 h, oxygen was fed to maintain the pressure at 10 bars and to ensure thorough oxidation. After completion, the oxygen feed and stirring were stopped and the reactor was cooled to 35 °C before separating leachate and leach residue through filtration. The leach residue was dried for later analysis and thiosulfate leaching tests, and the liquid was collected for further analysis to assess the extent of leaching and gold recovery.

2.3.4. Ammoniacal Oxidative Leaching

A solid sample of 250 g of was mixed with 500 mL of solution containing 0.4 M ammonia and 10 mM CuSO₄ to catalyze the oxidation of sulfur species and increase the overall leaching kinetics. The mixture was placed in a bottle roller rotating at 15 rpm and maintained at room temperature (20 °C) for a total duration of 24–50 h. After this initial pre-treatment, 0.2 M thiosulfate (TS) was added to the same solution. The pH of the solution was adjusted and stabilized at pH = 10, and leaching was conducted for an additional 24 h, as described in Section 2.2. Finally, samples were collected for analysis using both ICP-OES (iCAP 7000, Thermo Fisher, Waltham, MA, USA) and EDXRF techniques to determine the concentration of metals in the liquid and solid phases.

3. Results and Discussion

3.1. Characterization of Sample

Based on chemical analysis (see

Table 1. Chemical composition of gold flotation concentrate.

Element	Al	Si	S	Ca	Fe	Cu	Zn	Mg	Ti	P
Unit	W%	W%	W%	W%	W%	mg/kg	mg/kg	mg/kg	mg/kg	mg/kg
Feed	1.5	4.2	13.1	1.4	23.8	3721	1438	4850	695.7	103

), it is evident that iron, sulfur, and silicon are the predominant elements, while calcium and aluminum can be considered minor elements in the sample. The mineralogical study via XRD, as depicted in Figure 3a, indicates that feed is predominantly composed of pyrite, pyrrhotite, quartz, and gypsum, while micas, calcite, and clay minerals are in minority.

The SEM analysis shown in Figure 3b confirms textural heterogeneity. The brightest grains correspond to high atomic number phases, and nearly identical Fe- and S-rich areas correspond to pyrite–pyrrhotite aggregates. Fragments with strong Si and O identify as quartz or feldspathic silicates. K-, Al-, and Mg-enriched areas correspond to sheet silicates such as biotite, and isolated Ca spots denote sparse calcite. The colored image at the bottom left assembles these signals: purple domains are iron sulfides, olive green areas are aluminosilicates (Sp. 1), dark green areas are quartz (Sp. 14), and bright green spots (Sp. are) is related to Ca-rich minerals. Hence, the image shows that coarse iron-sulfide grains are locked in a matrix of quartz and mica. The SEM-EDS analysis, including point analysis (Figure 3b), indicates the presence of iron sulfides, quartz, aluminosilicates, and calcite as dominant species and confirms the XRD evaluation.

Diagnostic leaching using thiosulfate confirmed that 81.3% of gold was recoverable through initial thiosulfate leaching, indicating that most of the gold was in an accessible form (

Table 2. Diagnostic leaching result for gold flotation concentrates as feed material.

Phases	Exposed Gold	Encapsulated in Sulfides	Encapsulated in Oxides	Encapsulated in Silicates	Total
Content g/t	26.0	3.9	1.0	1.0	32.0
Distribution%	81.3	12.3	3.3	3.1	100.0

). Subsequent treatment with 20% HCl at 50 °C for 5 h released an additional 3.3%, likely from gold associated with carbonate or oxide minerals. Leaching with 5 M nitric acid at 70–80 °C for 7 h dissolved refractory sulfide phases, releasing 12.3% of the gold that had been encapsulated in minerals like pyrite. At each stage, the leachates were filtered and both filtrate and residue were analyzed by ED-XRD for Au content to ensure mass balance closure (>99%). Excess reagent concentrations and extended contact times were used to prevent kinetic limitations. Despite these steps, 3.1% of the gold remained trapped in the solid residue, potentially within silicate phases such as quartz, which is resistant to chemical dissolution in the used media. This deportment falls within the refractory class and highlights the importance of selecting appropriate pretreatment techniques for optimizing gold leaching.

3.2. Direct Cyanidation Test for Feed Material

A direct cyanidation test was carried out to evaluate the leachability of gold under standard conditions. A 1000 g of the homogenized concentrate was leached with a 0.25 wt% NaCN solution for 24 h. At the end of leaching, leachate was assayed for dissolved gold by A.A.S. Despite the high gold content, the overall recovery in the solution was poor (ca. 48.6%). The cyanide leach was yielding only 5.7%, even with grinding the feed to a particle size distribution of 95% < 75 µm, aiming to liberate fine gold inclusions. This confirms that the material is refractory: no matter how small the particles are, the gold remains locked within matrices and thus remains inaccessible to direct cyanidation. Moreover, the use of forged-steel grinding media exacerbates the problem. Galvanic coupling between steel balls (Fe⁰) and the sulfide mineral surfaces (e.g., pyrite, and pyrrhotite) accelerates the surface precipitation of iron hydroxides [36]. The steel acts as the anode, oxidizing to Fe²⁺/Fe³⁺, while the sulfide mineral surface acts as the cathode, facilitating oxygen reduction. The main reactions can be represented as follows:

• Anodic (on steel surface):

Fe⁰ → Fe²⁺ + 2e⁻,

(11)

Fe⁰ → Fe³⁺ + 3e⁻.

(12)

• Cathodic (on pyrite surface):

½ O₂ + H₂O + 2e⁻ → 2 OH⁻.

(13)

These galvanic reactions lead to localized increases in Fe²⁺ and hydroxide ions that subsequently form iron hydroxides upon exposure to air, which scavenge free cyanide by transforming up to 75% into ferrocyanide complexes that cannot dissolve gold. This also deplete dissolved oxygen in the pulp, further hindering the oxidative dissolution process [36]. Accordingly, an effective physicochemical pretreatment step is required to unlock encapsulated gold and achieve high extraction efficiencies in subsequent leaching.

3.3. Pretreatment Strategies for Enhanced Gold Liberation and Recovery

3.3.1. Pre-Treatment with Ultrafine Grinding (UFG)

The original sample was ground using a planetary ball mill for 1, 3, and 6 min with the aim of liberating submicron gold inclusions that remain encapsulated at coarser sizes. The particle size distribution for both the original feed and the ground materials have been shown in Figure 4a. Grinding resulted in a substantial reduction in particle size, with D₉₀ decreasing from 302 µm in the original feed to 61 µm, 44 µm, and 38 µm after 1, 3, and 6 min of grinding, respectively. The BET surface area analysis in

Table 3. Effect of grinding time on particle size distribution and BET surface area.

	No-Grinding	Grinding Time
		1 min	3 min	6 min
D10	16.5	2.36	1.60	1.37
D50	107	15.6	9.16	9.02
D90	302	61.3	44.2	38.6
BET, m2/g	0.70	1.64	3.22	5.65

confirmed that grinding led to a significant increase in surface area, with the values rising by factors of 2.34, 4.60, and 8.07, respectively. Although an increase in surface area is typically expected to enhance leaching efficiency by exposing more reactive surfaces, the leaching tests showed a decline in performance (Figure 4b). After 6 min of grinding, gold dissolution dropped to 18.5%, indicating a detrimental effect of excessive grinding on leaching efficiency. The redox potential and pH were significantly lowered (Figure 4c) leading to the precipitation of copper sulfide accompanied by the solution turning colorless; this suggests that the oxidation of Cu(I) to Cu(II) was inhibited and, consequently, there was not enough oxidant in the solution for gold extraction. Additionally, thiosulfate consumption increased by 10% during the same leaching period due to the larger surface area of pyrite/pyrrhotite, which consume thiosulfate and form polythionates/elemental sulfur. These products can passivate gold surfaces and depress leach efficiency [14,36,37].

Moreover, as it has been mentioned in Section 3.2, the galvanic effect and iron hydroxides’ formation may also be attributed to changes in the solution’s chemistry during leaching and reducing the gold dissolution as reported by other researchers [36]. Our preliminary experiments, in which small amounts of iron hydroxide were deliberately added to the leaching system, confirmed that even very low concentrations (≈2 wt%) of Fe(OH)₃ can significantly reduce the gold leaching rate (≈27 wt%), supporting the inhibitory role of iron hydroxides observed after extensive grinding. The findings suggest that while ultra-fine grinding increases surface area, it can also destabilize the leaching environment by altering pH, redox potential, and Fe/Cu speciation, therefore hindering the efficiency of thiosulfate leaching.

3.3.2. Pre-Treatment with Heat Treatment

The roasting route subjects the gold flotation concentrate to a 750 °C oxidative heat treatment for 2 h, volatilizing part of the sulfur and converting pyrite–pyrrhotite into porous iron oxide that expose sub-micron gold. The calcined residue is then cooled and transferred to an ambient-pressure thiosulfate leach (25 °C, 24 h), where the liberated gold is complexed as Au(S₂O₃)₂³⁻ for downstream recovery (Figure 5).

The temperature profile for the heat treatment is shown in Figure 6. From room temperature, the feed is ramp-heated to 750 °C in roughly 140 min, and the maximum temperature is maintained for ~120 min (Yellow shading) before the furnace is allowed to cool to ambient temperature. During the heating ramp, up to 300 °C the surface moisture and crystalline water are expelled and all carbonates (e.g., calcite, dolomite) decompose. Between about 350 °C and 600 °C, pyrite and pyrrhotite are oxidized if oxygen is available and pyrolyse in inert atmosphere forms SO₂ or elemental sulfur. At the isothermal hold at 750 °C in air, most of the residual sulfide is oxidized and magnetite is further oxidized to porous hematite (Fe₂O₃). These reactions may destroy the encapsulating sulfide matrix and coarsen liberated gold particles, but they can also generate fresh iron-oxide that re-coats the gold if the atmosphere is overly oxidizing.

As illustrated in Figure 6b, in the presence of oxygen, the glass tube’s color turned red, likely due to the formation of iron oxides. Conversely, in the absence of oxygen, a yellow powder was deposited inside the furnace, which was identified as elemental sulfur, indicating sulfur release during thermal decomposition. Additionally, the final product had a dark gray color in the absence of oxygen, while under oxygenated conditions, the product color turned dark red, confirming the formation of hematite. These observations were validated through XRD analysis (see Figure 7a), where FeS was identified as the major phase in the nitrogen (N₂) atmosphere, while hematite (Fe₂O₃) and anhydrite (CaSO₄) were dominant under oxygen conditions. Minor phases included silicate minerals such as quartz, biotite, and zeolite.

Elemental analysis conducted by XRF showed that the bulk chemistry of the calcines changes with the roasting atmosphere. The N₂-treated calcine contains 18.0% sulfur versus 2.17% under oxidizing conditions. In the air-oxidized product, iron (46.1%) and calcium (1.7%) are enriched relative to the N₂-treated sample (Fe 34.3%, Ca 1.4%), consistent with partial sulfur removal and relative enrichment of the remaining elements. The gold grade, measured by fire assay, remains essentially constant at ~45 ppm in both products; however, subsequent leach recovery will depend on how accessible it is within the very different mineral matrices produced by inert versus oxidizing roasting.

Further confirmation was provided by SEM-EDS analysis, which revealed differences in surface morphology and particle structure between the two treatments. In Figure 7b, the image looks melted and grain edges are blurred and rough. Such high roughness is typical when pyrite and pyrrhotite are fully oxidized to porous magnetite/hematite in the presence of oxygen. Diffusion of SO₂ and inward diffusion of O₂ leave a sponge-like iron-oxide skeleton. By contrast, the N₂-treated sample (Figure 7c) retains sharp, angular fragments with relatively smooth facets and clean interparticle boundaries. The particles stay separate and show clean edges because they only lose some sulfur instead of fully burning. As a result, the air-roasted material is more porous and may let leach chemicals in more easily, while the nitrogen-roasted material is tougher and may still trap some gold inside.

EDS elemental analysis for both samples has been performed, and elemental mapping has been provided in Figure 8. In the N₂-roasted sample, the elemental maps show that iron and sulfur are still coupled: where the Fe signal is the strongest, it coincides almost perfectly with the S map, indicating that the major Fe-bearing phase remains an Fe–S mineral (residual pyrite/pyrrhotite). Oxygen is confined mainly to the fine matrix and to discrete particles rich in Si, Al, and K, indicating that O is hosted largely by quartz and aluminosilicate gangue. The overlapping Al–Si–K–Mg spots correspond to liberated mica/chlorite fragments. Although the Ca-rich spots are attributed to calcium sulfide, they can partly be calcium oxide and remaining calcium carbonate grains not fully dissociated at 750 °C.

In the oxidized sample in Figure 8 (right), iron remains abundant and is now accompanied everywhere by a strong oxygen signal, while the sulfur is significantly less. This Fe/O pairing indicates that most of the original pyrite/pyrrhotite has been converted to iron oxides (magnetite/hematite), as it is confirmed by the phase diagram in Figure 9. Like the previous image, clusters rich in Al, Si, K, and Mg are mica-chlorite minerals, while calcium is present at least as CaS, in which sulfur may resist oxidation.

Figure 10 compares thiosulfate (TS) leaching behavior for three feeds: the untreated flotation concentrate as original feed, its oxidized calcine (air-treated), and its pyrolyzed calcine (N₂-treated,). The oxidized calcine used virtually all of the added thiosulfate (≈100%), far more than the original concentrate (≈75%) and over twice the demand of the inert roast (≈40%). Even though so much reagent is consumed, gold recovery from the oxidized calcine is the worst (~37%). Extensive iron-oxide surfaces catalyze TS oxidation to polythionates, leading to higher reagent loss. Also, iron-oxide skin formed during roasting blocks the gold surface and stops further dissolution. Moreover, when thiosulfate is used up during leaching, the normal redox cycle between Cu(I) and Cu(II) collapses because copper becomes locked up as stable copper-ammine complexes such as Cu(NH₃)₄²⁺ and Cu(NH₃)₂⁺. Therefore, the electrochemical reactions that drive gold dissolution can no longer take place [2,7]. The leach liquor’s intense dark blue color is diagnostic of tetraamminecopper(II), [Cu(NH₃)₄]²⁺, confirming that virtually all dissolved copper has been converted to stable ammine complexes. The pyrolyzed sample sits in the middle: it keeps most of its sulfide structure, wastes much less thiosulfate, but recovers only half of the gold.

In summary, the untreated flotation concentrate delivered the highest gold extraction in thiosulfate leaching, whereas both thermal pretreatments, particularly the oxidative roast, lowered recovery and, in the case of the oxidized calcine, sharply increased reagent consumption and passivation risk.

3.3.3. Pre-Treatment with Pressure Oxidative Leaching (POX)

Pressure oxidation (POX) technology has been applied to treat refractory gold concentrates to enhance gold recovery. The primary objective of the pressure oxidation experiments was to break down pyrite and other sulfide inclusions that encapsulate gold, thereby improving gold accessibility during leaching. In the Fe-S-H₂O Pourbaix diagram at 25 °C (Figure 11), the stability fields of pyrite (FeS₂) and pyrrhotite (Fe_1−xS) lie below about 0.3 V (SHE) at acidic pH. Once the solution potential is raised above ≈+0.3 V at pH < 1—conditions typical of our pressure leach (Eh ≈ +0.65–0.75 V, pH ≈ 0.1)—both sulfides move well outside their stability domains and are thermodynamically driven to oxidize.

Key steps of the process include pressure oxidation followed by thiosulfate leaching. The results show that 33% of the solids were dissolved during the PL process, leaving a residue still rich in pyrite and pyrrhotite. However, 50% of the thiosulfate reagent was consumed during subsequent leaching and the gold dissolution reached only 40%. This confirmed that pre-oxidation had a negative impact on gold leaching with thiosulfate. Several factors could have contributed to this outcome, such as the formation of passivating layers (e.g., jarosite or sulfur species) on gold surfaces. However, XRD spectra of PL residue (see Figure 12a) showed that the partial oxidation of pyrite (FeS₂) and pyrrhotite (Fe_1−xS) did not form insoluble jarosite or sulfur and may have led to soluble ferrous sulfate (Equation (14)). The presence of anhydrite in XRD spectra indicates that reaction happens between sulfuric acid and calcium-containing gangue minerals (Equation (15)). Also, the presence of white particles in the leach residue may be attributed to CaSO₄ formation. Quartz and other silicate minerals, such as chabazite, are generally stable under pressure oxidation conditions and resistant to both sulfuric acid and high-temperature oxidative reactions (

Table 4. Chemical analysis of leachate and residue after PL and TS leaching.

ELM.	Al	Si	S	Ca	Fe	K	Mg	Zn	As	Pb
	%	%	%	%	%	mg/kg	mg/kg	mg/kg	mg/kg	mg/kg
Feed	3.0	8.0	12.2	1.34	22.9	4738	8109	1718	87.2	510.3
PL leachate	0.24	0.14	3.18	0.15	22.1	406.8	1485	207.3	65.1	12.6
PL residue	2.04	14.01	23.4	4.10	30.9	4064	6632	221.6	1631	1140
TS residue	2.10	13.40	22.8	2.00	31.9	3970	5670	223.8	1161	725

). This result was also confirmed by the SEM-EDS results illustrated in Figure 12b. Therefore, they mainly remain in the residue; however, the leach liquor analysis (Table 4) shows that K-bearing aluminosilicates such as micas were partially attacked as Al, Si, and K were released into the solution. Thus, although base-metal sulfides were the primary targets, secondary reactions involving silicate breakdown and anhydrite re-encapsulation help explain the low gold dissolution and highlight the need for a pretreatment route that minimizes gangue dissolution while maximizing sulfide breakdown.

FeS₂ (s) + 14 Fe³⁺ (aq) + 8 H₂O (aq) → 15 Fe²⁺ (aq) + 2 SO₄²⁻ (aq) + 16 H⁺ (aq) ΔG = −89.03 kcal

(14)

CaCO₃ (s) + H₂SO₄ (aq) → CaSO₄ (anhydrite) (s) + CO₂ (g) + H₂O (aq) ΔG = −32.30 kcal

(15)

3.3.4. Pretreatment with Ammoniacal Oxidative Leaching

In conventional ammoniacal thiosulfate leaching, sulfide minerals oxidize and break down, releasing gold that was locked within their structures. At the same time, the sulfides deplete dissolved oxygen and promote thiosulfate decomposition, which accelerates reagent loss and generates passivating products that hinder further gold leaching [7,26]. Ammoniacal oxidative pre-treatment has also been studied as a potential solution to these challenges and significantly improved the efficiency of gold leaching from sulfide ores and reduced thiosulfate consumption [7,26,38]. Under mild oxidative conditions, ammonia selectively dissolves sulfide minerals, forming soluble species such as sulfate, thiosulfate, and thionates, while precipitating iron as ferric hydroxide in alkaline conditions. Based on Figure 13, the formation of stable ammine complexes with Cu (pH ≈ 9–10, Eh ≈ +200 to +400 mV) helps stabilize the solution, while cupric ammine complexes catalyze the oxidation of sulfur species, enhancing the overall leaching kinetics. Studies suggest that oxidative pre-treatment with ammoniacal solutions can reduce thiosulfate consumption and improve gold dissolution by partially oxidizing sulfides before the addition of thiosulfate [15,26].

This method involved two-step leaching. In the first step, an ammonium–copper solution was used as oxidant in a 1 L closed reactor to prevent ammonia loss. The copper concentration was maintained at 10 mM, and the oxidation process was conducted for 24, 42, and 50 h. This helped the partial oxidization of sulfides prior to the addition of thiosulfate. After this step, 0.2 M thiosulfate (TS) was added, and leaching continued for another 24 h. The result is illustrated in Figure 14, indicating that gold extraction increased significantly from 71% without pre-treatment to 85% after 42 h of oxidation because the Cu²⁺/NH₃/O₂ mixture selectively oxidized the outer layers of pyrite and, more critically, the highly reactive pyrrhotite that otherwise catalyze thiosulfate breakdown [39]. The same chemistry stabilizes the Cu(I)/Cu(II)-thiosulfate redox couple; so far, less reagent is consumed in side reactions. Feng and van Deventer also reported that even a modest removal of sulfide by this treatment suppressed mineral-driven thiosulfate decomposition and raised subsequent gold dissolution by more than 20% [7]. Consistent with these findings, our tests showed that the same pretreatment cut thiosulfate usage from 48.4 kg t⁻¹ to 29.0 kg t⁻¹.

Feng and van Deventer (2010) [7] presented that in the ammoniacal thiosulfate leaching, sulfide minerals oxidatively decompose to partially or completely release gold from the sulfide matrices. However, in this process, the sulfide minerals will consume dissolved oxygen in the solution and catalyze the decomposition of thiosulfate. This will cause excessive thiosulfate consumption and leaching passivation due to thiosulfate decomposition in the ammoniacal thiosulfate leaching of sulfide ores.

A comparative summary of the four pretreatment methods is presented in

Table 5. Comparison of pretreatment methods and their impact on gold leaching efficiency and thiosulfate consumption.

Pretreatment Method	Key Conditions	Gold Leaching Efficiency (%)	Thiosulfate Consumption (kg t−1)	Main Observations
Ultra-fine grinding	6 min grinding; 8× increase in BET surface area	18.5	54.4	Excessive grinding accelerated thiosulfate decomposition and Cu(I) passivation; severe decline in gold recovery.
Roasting	750 °C, 2 h, air atmosphere 750 °C, 2 h, Inert atmosphere	37.0 56.0	64.0 25.6	Pyrite–pyrrhotite converted to hematite; no significant liberation of gold; limited benefit from thermal treatment.
Pressure Oxidative Leaching (POX)	190 °C, 10 bar O2	40.0	32.0	Dissolved ~33% of solids; high reagent and energy consumption and low gold recovery.
Oxidative ammoniacal pre-leaching	0.4 M NH3 + 10 mM Cu2+, 42 h	85.0	29.0	Enhanced gold liberation, lower reagent use, and improved leaching kinetics.

, illustrating their influence on both gold extraction and thiosulfate consumption. These results clearly demonstrate that, among the evaluated pretreatments, oxidative ammoniacal conditioning offers the most efficient and environmentally benign route by effectively unlocking encapsulated gold while minimizing reagent loss.

The proposed route begins with the flotation concentrate, which is subjected to oxidative pretreatment with copper amin complex, to liberate finely disseminated gold and suppress sulfide-catalyzed thiosulfate breakdown. The oxidized pulp then enters a thiosulfate leach, where gold dissolves as the stable [Au(S₂O₃)₂]³⁻ complex. The pregnant solution is purified in an ion-exchange (IX) process with a strong-base anion resin (Amberlite IRA-402) to strip residual copper and other competing anions while simultaneously loading gold. Gold is subsequently eluted from the loaded resin by a mixture of sodium chloride and sodium sulfite and is recovered by cementation with high-purity zinc dust, achieving >99% gold recovery in a form suitable for refining. This sequence offers an environmentally benign, cyanide-free route for treating refractory sulfide concentrates.

4. Conclusions

The sustainability challenge is intensifying as a growing share of global gold resources is refractory. Such ores generally require pretreatment to break down the host minerals and make gold accessible to leaching. Thiosulfate leaching is a promising cyanide-free route for refractory ores; however, it remains challenging due to high thiosulfate consumption, polythionate formation, and the need for tight copper–ammonia–oxygen control. Therefore, process economics focus on selecting a pretreatment that both liberates gold and suppresses side reactions. Grinding the sample to finer sizes (from 300 μm to 38 μm) to liberate more gold was not successful. Ultrafine comminution destabilized the thiosulfate system by shifting pH and Eh and altering Fe/Cu speciation, which ultimately suppressed gold dissolution. Pressure oxidation using Fe³⁺ at 10 bar and 190 °C, applied to break down the sulfide matrix, was also underperformed: passivating films formed on gold and re-encapsulation in secondary phases (e.g., anhydrite) limited extraction. High-temperature thermal pretreatments, particularly oxidative roasting, produced the same effect, increasing passivation and lowering recovery from ~60% to ~37%. This study shows that the key to unlocking this refractory ore does not lie in finer grinding or more aggressive acid pressure leaching but in choosing a pretreatment that balances liberation with chemical selectivity. By targeting the highly reactive pyrrhotite-rich matrix with a mild, oxidative ammoniacal conditioning step (Cu, 10 Mm + 0.4 M NH₃), gold dissolution reached ≈ 85% with a thiosulfate consumption of 45% compared to the one without pre-treatment. We attempted to suppress ferric-catalyzed thiosulfate degradation and prevent the formation of passivating Fe-oxy-hydroxysulfate skins. This strategy transforms a concentrate once deemed unsaleable into a feed that can deliver high gold extraction at modest reagent cost, providing a cleaner, more energy-efficient alternative to conventional cyanidation-based flowsheets.