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Article

Hydrometallurgical Processing of a Low-Grade Sulfide Copper–Nickel Ore Containing Pt and Pd

by
Elena Latyuk
1,2,
Andrey Goryachev
2,
Vitaliy Melamud
1 and
Aleksandr Bulaev
1,*
1
Research Center of Biotechnology, Russian Academy of Sciences, 119071 Moscow, Russia
2
Institute of North Industrial Ecology Problems, Federal Research Center “Kola Science Center”, Russian Academy of Sciences, 184209 Apatity, Russia
*
Author to whom correspondence should be addressed.
Processes 2024, 12(6), 1213; https://doi.org/10.3390/pr12061213
Submission received: 10 May 2024 / Revised: 8 June 2024 / Accepted: 12 June 2024 / Published: 13 June 2024
(This article belongs to the Special Issue Recent Trends in Extractive Metallurgy)
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Abstract

:
The goal of the present work was to study the recovery of copper, nickel, and platinum group metals (PGMs) (Pt and Pd) from low-grade copper–nickel ore containing pyrrhotite, pentlandite, and chalcopyrite by column bioleaching followed by cyanidation. The ore sample contained the following: Ni—0.74%, Cu—0.23%, Fe—14.8%, Stotal—8.1%, and Ssulfide—7.8%. The Pt and Pd contents in the ore sample were 0.2535 and 0.515 g/t, respectively. Biological leaching in columns was carried out at 25, 35, and 45 °C for 140 days. A mixed culture of acidophilic microorganisms was used as an inoculum. Cu and Ni extraction depended on temperature, and at 45 °C, copper and nickel recovery was the highest, being 2.1 and 1.8 times higher than that at 25 °C, respectively. As a result, up to 35% of nickel and up to 10% of copper were recovered by bioleaching within 140 days. Bioleaching resulted in an increase in Pt and Pd recovery by cyanidation, but the effect on Pd recovery was insignificant. Pt recovery varied in the range of 3–40% depending on process conditions; Pd recovery was 44–55%.

1. Introduction

Platinum group metals (PGMs) are among the most demanded metals due to their application in industry and relatively low abundance. PGMs are widely used in many industrial sectors. This is due to their chemical inertness to various chemical reagents and their excellent catalytic properties. They are used in the chemical, petroleum refining, glass, dental, electronics, electrical, photovoltaic, fuel cell, medical, and pharmaceutical industries, as well as jewelry [1,2,3,4].
Current problems in PGM production are caused both by the current state of PGM production technologies and by the partial depletion of currently used PGM resources [4].
A significant part of PGM reserves is associated with massive and disseminated sulfide copper–nickel ores. Copper–nickel ores are currently processed mainly by obtaining copper–nickel concentrates using flotation and further pyrometallurgical processing. PGMs are concentrated in different products at different stages of processing: first, in nickel and copper concentrates, and then in electrolysis sludge, which is formed during the refining of nickel and copper. From the resulting sludge, PGM concentrates are obtained, which are then subjected to refining [1,2,4,5]. Thus, the main PGM source is nickel and copper–nickel sulfide ores, which are treated via beneficiation and pyrometallurgical processing.
Metallurgy is faced with the problem of depleting reserves of developed metal deposits, in particular nickel and copper–nickel ores, which can be processed using traditional beneficiation and metallurgy technologies. In this regard, the exploitation of small and technogenic deposits with low-grade and difficult-to-process ores is a relevant issue [2,4,6,7].
This problem is typical of the mining and metallurgical industry of the Arctic zone of the Russian Federation, for which an important technological, economic, and environmental task is to increase the recovery of non-ferrous and precious metals from both natural sulfide ores and man-made mineral formations, which include the accumulated and current waste of mining and metallurgical production. Processing of man-made mineral resources is an urgent goal, as it allows extracting additional valuable resources as well as reducing the load on the environment caused by long-term pollution, which is in turn critically important for sustainable metal production in regions with a long history of metal production where resources of high-grade conventional ores have been depleted due to mining during several decades [6,7,8]. The above local problem may have significance for global PGM supply, as the Russian Federation occupies one of the leading positions in the world in reserves, production, and exports of PGM metals. The main PGM reserves are concentrated in the Norilsk copper–nickel deposits of the Krasnoyarsk Territory (95.6% of Russian reserves). Relatively large reserves have been explored in the Murmansk region (3.7%), which is a part of the Arctic Zone of Russia [6,7,8].
The analysis of the state of the mineral resource base of PGMs leads to the conclusion that alternative technological pathways are needed that allow the cost-effective recovery of PGMs and non-ferrous metals from low-grade raw materials [4,5,9,10,11,12,13,14].
In recent years, it has been shown that PGM can be recovered using hydrometallurgical methods based on the leaching of non-ferrous metals and PGMs. Approaches based on bioleaching and further cyanidation have been studied [15,16,17] as alternative technologies for Ni and Cu recovery [18], as well as PGM ores and concentrates [4,9,10,11,12,13,14]. These biotechnologies are successfully used to recover non-ferrous metals and gold from sulfide ores and concentrates [15,16,17,18,19,20]. In the case of processing refractory gold ores and concentrates, bioleaching makes it possible to recover gold through the destruction of gold-bearing minerals (pyrite, arsenopyrite) due to the biooxidation of sulfides by acidophilic sulfur- and iron-oxidizing microorganisms. Biooxidation residue is then treated by cyanidation for gold recovery. Biooxidation of refractory sulfide ores and concentrates significantly increases gold leaching extent by cyanidation, which allows the wide use of this method on an industrial scale to obtain gold [16,19,20].
Despite the successful application of bioleaching and cyanidation for gold extraction, this method is not used in practice for PGM recovery. The possibility of bioleaching and further cyanidation for Ni/Cu and PGM recovery was shown in some studies, but this aspect of bioleaching application has been less studied in comparison to the biohydrometallurgical processing of gold-bearing ores [4]. Therefore, further development and application of biohydrometallurgical technologies for Ni/Cu and PGM production are urgent tasks.
In the present work, we studied the possibility of Ni, Cu, and PGM (Pt and Pd) recovery from a sample of technogenic ore from the Nud II deposit (Murmansk region, Russia) using hydrometallurgical pathways (bioleaching and cyanidation). This goal of this study is important both from the point of view of further development of biohydrometallurgical technologies for specific mineral raws and from the point of view of local issues of metallurgy in the Arctic zone of the Russian Federation.
To evaluate the dependence of Ni, Cu, Pt, and Pd recovery from the ore on temperature and duration of bioleaching, bioleaching of low-grade copper–nickel ore containing PGMs (Pt and Pd) in this work was performed under different conditions. Also, the composition of the microbial population formed under different temperatures was studied. In our previous works, we studied the possibility of bioleaching several mineral products from technogenic deposits of the Murmansk region (Russia) (low-grade ore, low-grade concentrate, as well as industrial sand) containing copper, nickel, and PGMs [10,11,21]. The possibility of Ni and Cu bioleaching from ores was studied at ambient temperature with indigenous mesophilic cultures, while the practice of bioleaching application suggests that temperature is one of the main factors affecting both bioleaching rate and microbial population activity [15,19]. Thus, in the present study, we have assessed the effect of different temperature conditions on the bioleaching of non-ferrous metals as well as on microbial population composition. Also, in our previous study, we demonstrated the possibility of Pt and Pd cyanide leaching from low-grade sulfide concentrate [11], while in the present study, we have studied the application of approaches based on bioleaching for the treatment of low-grade ore, which is more promising. Thus, the results of this work make a certain contribution to understanding patterns of Cu, Ni, and PGM recovery from low-grade sulfide ores, which is an urgent issue [4].

2. Materials and Methods

The object of the study was a sample of copper–nickel ore from the Nud II deposit (67°53′9″ N, 32°54′6″ E), Monchegorsk ore district, Murmansk region (the Kola Peninsula) (Figure 1). The ore sample contained the following: Ni—0.74%, Cu—0.23%, Fe—14.8%, Stotal—8.1%, and Ssulfide—7.8%. Element content determined using X-ray fluorescence (XRF) analysis is shown in Table 1. The Pt and Pd contents in the ore sample were 0.2535 and 0.515 g/t.
According to XRD analysis, the main ore minerals were pentlandite, chalcopyrite, pyrrhotite, and magnetite (Figure 2). In the work [22], it was shown that Pd and Pt in the deposit are mainly presented in the form of merenskyite ((Pd,Pt)(Te,Bi)2), michenerite (Pd0.75Pt0.25BiTe), and moncheite (Pt(Te,Bi)), while sperrylite (PtAs2) and platarsite (PtAsS) were also detected as minor PGM minerals.

2.1. Column Bioleaching

Mixed cultures of acidophilic microorganisms, which were previously used for bioleaching of flotation tailings and sulfide concentrate, were used for bioleaching [23]. The culture was grown in a liquid medium where polymetallic ore flotation tailings containing pyrite were added as a substrate.
Cells of the mixed cultures were collected by centrifugation in sterile 500-mL tubes (9500 rpm, 15 min) using a Sigma 6K–15 centrifuge (Sigma, Osterode am Harz, Germany), resuspended in mineral medium, and inoculated in the columns in such a way that the initial cell number in the liquid phase was about 1 × 108 cells/mL. For the experiments, we used a liquid nutrient medium containing mineral salts (g/L): (NH4)2SO4—0.75, KCl—0.05, MgSO4 × 7H2O—0.125, K2HPO4—0.125, distilled water—1.0 L.
Ore samples (1 kg) with a particle size of −5 + 1 mm were placed in 1 L-polypropylene columns. To reach the required temperatures, columns were placed in TC-1/80 thermostats (SKTB, Smolensk, Russia). One liter of liquid medium was percolated through the ore layer using BT100-2J peristaltic pumps (Longer Precision Pump Co., Ltd., Baoding, China). The initial pH value of the leaching solution was 1.5. The experiment was carried out at different temperatures (25, 35, and 45 °C), since temperature has a significant effect on bioleaching processes. In industrial processes of heap bioleaching of sulfide ores, the ore heaps always heat up, and the temperature changes deep in the ore layer. This phenomenon is due to heat generation caused by sulfide mineral biooxidation and was observed during heap bioleaching of different sulfide ores, including polymetallic nickel ore [24]. Therefore, it is important to evaluate the dependence of the bioleaching rate of each ore studied on temperature and the possibility of carrying out the process at different temperatures.
The bioleaching process was carried out for 140 days. For the first 68 days (Stage 1) (Table 1), bioleaching was performed in parallel in 2 columns at each temperature, and then the solid phase from the 2 parallel columns was collected, mixed, and dried at a temperature of 80 °C and weighted. Samples of biooxidized ore (200 g) were grinded to a particle size of −100 μm, and the grinded samples were subjected to cyanidation (Section 2.4). 1 kg of each residue obtained at each temperature was then placed in the column, and bioleaching was continued. After 140 days of bioleaching (Stage 2) (Table 1), the obtained residues were subjected to cyanidation as described above. The separation of the bioleaching experiment into 2 stages made it possible to determine the effect of process duration on the bioleaching of copper and nickel, as well as to obtain solid residue samples to study Pt and Pd extraction by cyanidation.

2.2. Sampling and Analysis

When carrying out the ore bioleaching process, samples of the liquid phase were taken, and pH and redox potential (Eh vs. SHE) values of pregnant solutions were measured. The pH and Eh values were measured using a pH-150MI pH meter (Izmeritelnaya Tehnika, Moscow, Russia). Using titration with Trilon B, the concentrations of Fe3+ and Fe2+ ions in the liquid phase were determined. The concentrations of copper and nickel were determined using a Perkin Elmer 3100 flame atomic absorption spectrometer (Perkin Elmer, Waltham, MA, USA). The rates of copper and nickel leaching from concentrate residue were calculated by the concentration of Cu and Ni ions in the liquid phase. The measurement of these parameters made it possible to judge the activity of the microbial populations that carried out the bioleaching process. During the biooxidation of sulfide minerals, iron, copper, and nickel ion accumulation in the liquid medium, as well as the formation of sulfuric acid (leading to a decrease in pH value), occur. The high ratio of Fe3+/Fe2+ ion concentrations, which is characteristic of active bioleaching processes, leads to an increase in the Eh of the medium. The pH values during the bioleaching were adjusted by adding concentrated (98%) sulfuric acid to the pregnant solution to avoid increasing the pH of the medium above values unfavorable for microorganisms.
XRD analysis of the ore and bioleaching residue was carried out using an XRD 7000 X-ray diffractometer (Shimadzu, Kyoto, Japan).

2.3. Microbial Population Analysis

The analysis of the composition of microbial populations that formed under the experimental conditions was carried out by high-throughput sequencing on the MiSeq system (Illumina, San Diego, CA, USA) of the V3–V4 region of the 16S rRNA gene. The analysis was described in detail in our previous work [25]. In the present work, we analyzed samples of biomass from the inoculum as well as the microbial population formed in the column at each temperature after 140 days of bioleaching.

2.4. Cyanidation Tests

The ore and bioleaching residues were subjected to sorption leaching by cyanidation to recover Pt and Pd. Leaching was carried out using a bottle agitator and technical grade sodium cyanide (Korund-Cyan CJSC) with the following parameters: solid phase content in the pulp—45%, sorbent content—5% (activated carbon Haycarb PLC), duration—24 h. The initial concentration of cyanide was 5 g/L. To determine the degree of PGM recovery, bioleaching and cyanidation residues, as well as sorbent, were analyzed for Pt and Pd content, and the recovery rate was calculated (analysis of PGM content in solid products was carried out in the analytical laboratory of JSC Regional Analytical Center Mekhanobr Engineering Analyte (Saint-Petersburg, Russia)).

3. Results

3.1. Bioleaching Experiments

The pH values (Figure 3) of the pregnant solution in the first 20–24 days of the bioleaching process were high (up to 3), and sulfuric acid had to be added to maintain a pH favorable for bioleaching. After that, the pH of the solutions stabilized at a level of 2–2.3.
The consumption of sulfuric acid was 27, 32, and 27 kg/t of ore at 25, 35, and 45 °C, respectively. The Eh values during the experiment varied from 540 to 930 mV. At the beginning of the process, there was a decrease in values below 600 mV, and it remained low until 14 days (at 25 and 35 °C), and then increased and exceeded 900 mV at 25 and 35 °C. At 45 °C, Eh values fluctuated throughout the observation period and were lower than those at 25 and 35 °C. This suggests that iron biooxidation was more active at lower temperatures than at 45 °C.
Changes in the Eh values of the solutions corresponded to changes in the concentration ratios of the iron ions Fe3+ and Fe2+. At 25 and 35 °C, ferrous iron was absent in the pregnant solution after 17 days. Therefore, after 17 days of bioleaching at 25 and 35 °C, total iron content corresponded to the concentration of Fe3+ ions (Figure 4a,b). At 45 °C, ferrous iron was present in the medium for a long time and was completely oxidized only on day 97 (Figure 4c).
The observed changes in the parameters of the liquid phase indicated that during bioleaching processes, there was a gradual increase in the activity of microorganisms. At the beginning of the experiments, there was a decrease in Eh and an accumulation of Fe2+ ions in the medium. That was due to the interaction of Fe3+ ions introduced into the liquid phase at the beginning of the process with sulfide minerals in the medium, as well as the dissolution of minerals in an acidic medium. Then, the oxidation of Fe2+ ions to Fe3+ occurred due to the activity of iron-oxidizing microorganisms.
The concentrations of copper and nickel gradually increased during the bioleaching process (Figure 5). It should be noted that the concentration of nickel ions exceeded the concentration of copper ions by an order of magnitude, which is primarily due to the fact that the chalcopyrite contained in the ore is more resistant to biooxidation than nickel sulfide minerals [18,26,27].
Also, this may be explained by the peculiarities of the mineral composition of the ore studied. The initial ore sample contained copper and nickel sulfides, such as pyrrhotite, pentlandite, and chalcopyrite [28] (Figure 2). Taking into account the previously studied features of the Nud II deposit ore, it can be argued that pentlandite is represented by monomineral grains and may contain an admixture of cobalt, which makes the ore suitable for nickel and cobalt recovery. At the same time, chalcopyrite in the ore is present in intergrowths, which significantly complicates the process of copper recovery into solution. Among the non-metallic minerals, feldspars and serpentine predominate. Serpentine is represented by antigorite. Serpentine may contain small amounts of pentlandite in the form of dissemination. The presence of magnetite was noted, which is present in intergrowths with silicate minerals, forming along cracks.
It should be noted that despite lower Eh values as well as a lower ferric iron concentration at 45 °C, copper and nickel recovery increased with increasing temperature. At 45 °C, copper and nickel recovery were 2.1 and 1.8 times higher than at 25 °C. It was previously shown that Cu recovery during bioleaching may increase with temperature increase and Eh decrease [26], which corresponds to the results of this work.
After interaction with a bacterial solution at 25 °C, the appearance of jarosite reflexes was noted (Figure 6a). In addition, the number of non-metallic mineral reflections increased with a general decrease in the peak intensity of copper- and nickel-containing sulfides. At a temperature of 35 °C, iron actively precipitated from the solution as jarosite and hydroxide-iron goethite (Figure 6b). After the ore interaction with the bacterial solution at a temperature of 45 °C, much more intense jarosite formation occurred, which was reflected in the appearance of new peaks in the diffraction pattern (Figure 6c).

3.2. Microbial Population Analysis

Molecular biological analysis of the microbial population in the columns formed for 140 days showed that the species composition of microorganisms changed at different temperatures (Figure 7). In the inoculum obtained at 25 °C, representatives of the genera Ferroplasma, Cunuciliplasma, uncultivated group A-plasma (which were recently described as the iron-oxidizing archaea “Candidatus Carboxiplasma ferriphilum” based on metagenomic analysis) [25], and Ferrimicrobium were predominant. After long-term bioleaching of the ore at 25 °C, representatives of iron-oxidizing bacteria Leptospirillum replaced archaea and were predominant. At 35 °C, the genera Leptospirillum, Ferrimicrobium, and Acinetobacter and the archaea of the genus Cuniculiplasma and the uncultivated group A-plasma were predominant. In the population obtained at 45 °C, genera Acinetobacter, Staphylococcus, and Ralstonia dominated.
Thus, at 45 °C, typical acidophilic microorganisms were replaced by heterotrophic microorganisms, the role of which in bioleaching has not been described. Representatives of A-plasma, Acidiphilum, and Leptospirillum were detected in the population, but their proportion was comparatively low. It may explain the comparatively low iron-oxidizing activity of the population at 45 °C. It should be noted that the presence of Acinetobacter in the population of bioleach reactors was shown in our recent work, but its role in the bioleaching process has not yet been understood [25].

3.3. Cyanidation Tests

Bioleaching resulted in an increase in Pt and Pd recovery by cyanidation, but the effect on Pd recovery was insignificant (Figure 8). Pt recovery increased only after 140 days of bioleaching at 25 and 35 °C. In contrast to lower temperatures, bioleaching at 45 °C provides a higher Pt recovery after 68 days of bioleaching, while further bioleaching leads to a decrease in the recovery.

4. Discussion

It should be noted that in practice, unlike laboratory scale column leaching, heap leaching processes are carried out over a long period of time (up to several years) [29]. Therefore, the results obtained during 140 days of laboratory tests can only show the principal possibility of recovering metals. Achieving a high rate of metal recovery is possible by conducting very long tests. The results obtained in the present work showed that bioleaching of the studied ore can be carried out at a wide range of temperatures and allows the recovery of both copper and nickel, but at the same time, bioleaching of nickel occurs at a higher rate. Overall, the results showed that bioleaching, even for 140 days, made it possible to recover up to 35% of nickel, which indicates the promise of continuing work using the studied pathways.
Analysis of the microbial community showed that the population composition changes at different temperatures. At 45 °C, typical acidophilic microorganisms are replaced by heterotrophic microorganisms. As temperature increases in the ore layer are typical of heap bioleaching processes [24], further studies may be necessary to develop methods to control the composition and activity of the microbial population performing heap bioleaching. These approaches may include the inoculation of ore heaps with specific inoculates adapted to specific temperature conditions [30].
It was shown that bioleaching resulted in increased recovery of Pt and Pd by cyanidation. This may be due to the destruction of sulfide minerals during bioleaching, which leads to the release of noble metals for cyanide leaching, which is well known for gold concentrates. Copper and nickel can form complexes with cyanide ions, which can complicate the leaching of PGMs with cyanide and their further sorption on activated carbon. The removal of these metals using bioleaching helps to increase PGM recovery [31].
Our previous results [11] with low-grade concentrate showed that this type of raw material can be processed using hydrometallurgical methods (bioleaching followed by cyanidation) (Table 2). Reactor leaching at 30 and 40 °C was studied, and as a result, it was possible to recover about 70% of nickel and up to 34% of copper. It was shown that copper recovery was more dependent on temperature than nickel, since chalcopyrite is more resistant to biooxidation than nickel minerals. It was shown several times that PGM recovery from the bioleaching residue was greater than PGM recovery from the concentrate sample. Cyanide leaching made it possible to extract 5.5% Pt and 17.3% Pd from the concentrate and 37.8% Pt and 87.8% Pd from the bioleaching residue. Also, the present work shows that the duration of the bioleaching process may have a countervailing effect on the PGM recovery, depending on the temperature of bioleaching. Table 2 also shows the results of our previous studies on column and stirred bioleaching of Cu–Ni sulfide products from different deposits in the Murmansk region (Russia). Summarizing the results of these studies, as well as the present work, it can be concluded that Cu bioleaching is more temperature dependent in comparison to Ni bioleaching, while Ni recovery in all cases was higher. This may be explained by the properties of chalcopyrite and nickel sulfide minerals [18,26,27,32].
The experience of previous studies showed that bioleaching of low-grade ore in columns made it possible to recover 87% copper, 71% nickel, and 47% cobalt; further cyanidation of the solid bioleaching residue made it possible to recover 54% Pt, 90% Pd, and 86.7% Au [33]. Bioleaching of low-grade concentrate in columns allows for the recovery of 52% copper, 95% nickel, and 85% cobalt, and subsequent cyanidation of the bioleaching residue allows for the recovery of 20% Pt, 87% Pd, and 46% Rh [12]. A study of bioleaching in columns at various temperatures (65–80 °C) allowed the recovery of 69.9–91% copper, 93–98.5% nickel, and 76.8–86.1% cobalt, and cyanidation of the solid bioleaching residue allowed the recovery of 32.2–34.3% Pt, 92.5–96.5% Pd, and 63.4–97.5% Au [14].
Thus, these studies show the prospects for using hydrometallurgical methods for both low-grade ore and low-grade concentrate. The potential advantages and disadvantages of the application of bioleaching and further cyanidation for the treatment of Cu–Ni ores containing PGM are considered in Table 3. It was shown that for these materials, different values for the recovery of both non-ferrous metals and PGMs were obtained. In general, there was a trend towards increasing PGM recovery by cyanidation after bioleaching, which also allowed recovering copper and nickel. In addition, to increase the recovery of PGMs by cyanidation, a search for modifications to the cyanidation process is necessary, since PGMs have been shown to be more resistant to cyanide leaching compared to gold [12,13,14,16]. Also, further studies are required to solve the problem of the low rate of chalcopyrite bioleaching and copper recovery. For example, approaches based on chloride leaching may be studied for Cu/Ni ores, as it has been shown that chloride leaching may be used to improve the results of copper and nickel bioleaching [34,35].
The ore studied in the present work is low-grade ore characterized by a low content of both non-ferrous metals (Cu and Ni) and PGM (Pd and Pt). The Nud II deposit was exploited in the 1960s and 1970s, and reserves of high-grade ore suitable for beneficiation and further treatment via the pyrometallurgical route were exhausted [22]. Heap leaching (including bioleaching) is widely used for the treatment of low-grade copper ores [15,19], and the possibility of using it for Cu–Ni ore was also studied on an industrial scale [24]. Practical examples demonstrate that for the treatment of low-grade ores, which cannot be effectively subjected to beneficiation to obtain high-grade concentrate, heap leaching seems to be the only promising approach due to its simplicity and comparatively low CAPEX and OPEX [29]. Therefore, this method is promising for Nud II deposits and similar small deposits with low-grade ores. This study demonstrated the possibility of both Cu and Ni extraction from the ore studied by bioleaching and further Pt and Pd extraction by cyanidation.
At the same time, further research is needed to develop an economically promising approach, as cost–benefit analysis is a required step in the decision-making process for the selection of technological approaches for metal extraction. Practical experience suggests that real financial calculations cannot be obtained based on the results of laboratory tests since CAPEX and OPEX, as well as real revenue and recoupment, are determined by many factors, including the following:
  • Availability of energy, transport infrastructure, equipment, labor resources, etc. at the location of the deposit;
  • Metal prices, which continuously change due to the fluctuation of supply and demand caused by global economic and political trends;
  • Available resources of valuable metals in the deposit, which can be determined by a detailed geological survey;
  • Some factors that cannot be predicted before industrial-scale trials (possible environmental consequences, operating features of the selected industrial equipment, natural disasters, etc.).
Some examples of the application of biohydrometallurgical technologies [19,20] suggest that real cost–benefit analysis may be made based on long-term trials, including industrial-scale tests. Based on the results of our laboratory scale tests, it would be unjustifiable to perform a practical cost–benefit analysis. Our results may be used only as the basis for further scaling and pilot tests. At the same time, it should be noted that we studied column bioleaching, which is usually used on a laboratory scale for modeling heap leaching processes, which in turn are the least costly methods used in metallurgy. For example, this method is successfully used for low-grade copper ores with copper content comparable to that in the ore studied in the present work [29,35]. Therefore, the results obtained may be considered to demonstrate the principal possibility of the approach proposed for the studied sulfide ore, and these results should be used for further planning of pilot scale trials.

5. Conclusions

The possibility of the recovery of copper, nickel, and PGMs from low-grade copper–nickel ore containing pyrrhotite, pentlandite, and chalcopyrite by column bioleaching at 25, 35, and 45 °C for 140 days followed by cyanidation was studied. As a result, up to 35% of nickel and up to 10% of copper were recovered by bioleaching within 140 days. It was shown that both Ni and Cu recovery depended on temperature and were higher at 45 °C. Bioleaching resulted in an increase in Pt and Pd recovery by cyanidation, but the effect on Pd recovery was insignificant. The recovery of Pt varied in the range of 3–40% depending on process conditions; the recovery of Pd was 44–55%.

Author Contributions

Conceptualization, E.L. and A.B.; methodology, A.B.; investigation, E.L., V.M. and A.B.; writing—original draft preparation, E.L., A.G. and A.B.; writing—review and editing, E.L., A.G. and A.B.; supervision, A.B. All authors have read and agreed to the published version of the manuscript.

Funding

The reported study was funded by RFBR according to the research project 19-35-50073 and by the Ministry of Science and Higher Education of the Russian Federation.

Data Availability Statement

The data that support the findings of this study are available from the corresponding author upon reasonable request.

Conflicts of Interest

The authors declare no conflicts of interest.

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Figure 1. Nud II deposit in the Monchegorsk ore district.
Figure 1. Nud II deposit in the Monchegorsk ore district.
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Figure 2. X-ray diffraction pattern of the ore sample. Amp: amphibole (No. 41-1366); S: serpentine (No. 50-1606); Po: pyrrhotite (No. 24-220); Pn: pentlandite (No. 8-90); Ccp: chalcopyrite (No. 37-471); Mt: magnetite (No. 19-629); F: feldspar (No. 41-1480); T: talc (No. 13-558); g: goethite (No. 29-713); Q: quartz (No. 46-1045); Ja: jarosite (No. 36-427).
Figure 2. X-ray diffraction pattern of the ore sample. Amp: amphibole (No. 41-1366); S: serpentine (No. 50-1606); Po: pyrrhotite (No. 24-220); Pn: pentlandite (No. 8-90); Ccp: chalcopyrite (No. 37-471); Mt: magnetite (No. 19-629); F: feldspar (No. 41-1480); T: talc (No. 13-558); g: goethite (No. 29-713); Q: quartz (No. 46-1045); Ja: jarosite (No. 36-427).
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Figure 3. Changes in liquid phase pH and Eh values during the bioleaching at 25 (a), 35 (b), and 45 (c) °C.
Figure 3. Changes in liquid phase pH and Eh values during the bioleaching at 25 (a), 35 (b), and 45 (c) °C.
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Figure 4. Fe3+ and Fe2+ ion concentrations (g/L) in the liquid phase during the bioleaching at 25 (a), 35 (b), and 45 (c) °C.
Figure 4. Fe3+ and Fe2+ ion concentrations (g/L) in the liquid phase during the bioleaching at 25 (a), 35 (b), and 45 (c) °C.
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Figure 5. Recovery dynamics Cu (a) and Ni (b) at 25, 35, and 45 °C.
Figure 5. Recovery dynamics Cu (a) and Ni (b) at 25, 35, and 45 °C.
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Figure 6. X-ray diffraction pattern of the samples: (a) bioleaching residue after 25 °C, (b) bioleaching residue after 35 °C, and (c) bioleaching residue after 45 °C. Amp: amphibole (No. 41-1366); S: serpentine (No. 50-1606); Po: pyrrhotite (No. 24-220); Pn: pentlandite (No. 8-90); Ccp: chalcopyrite (No. 37-471); Mt: magnetite (No. 19-629); F: feldspar (No. 41-1480); T: talc (No. 13-558); g: goethite (No. 29-713); Q: quartz (No. 46-1045); Ja: jarosite (No. 36-427).
Figure 6. X-ray diffraction pattern of the samples: (a) bioleaching residue after 25 °C, (b) bioleaching residue after 35 °C, and (c) bioleaching residue after 45 °C. Amp: amphibole (No. 41-1366); S: serpentine (No. 50-1606); Po: pyrrhotite (No. 24-220); Pn: pentlandite (No. 8-90); Ccp: chalcopyrite (No. 37-471); Mt: magnetite (No. 19-629); F: feldspar (No. 41-1480); T: talc (No. 13-558); g: goethite (No. 29-713); Q: quartz (No. 46-1045); Ja: jarosite (No. 36-427).
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Figure 7. Analysis of microbial populations performing bioleaching (proportion of the 16S rRNA gene fragment sequences, %).
Figure 7. Analysis of microbial populations performing bioleaching (proportion of the 16S rRNA gene fragment sequences, %).
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Figure 8. Pt and Pd recovery (% ± SD based on 2 measurements) by cyanidation from the ore and bioleaching residues.
Figure 8. Pt and Pd recovery (% ± SD based on 2 measurements) by cyanidation from the ore and bioleaching residues.
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Table 1. Element content in the ore sample and biooxidation residue according to X-ray fluorescence (XRF) conducted using the X-550 XRF Analyzer (SciAps, Inc., Andover, MA, USA).
Table 1. Element content in the ore sample and biooxidation residue according to X-ray fluorescence (XRF) conducted using the X-550 XRF Analyzer (SciAps, Inc., Andover, MA, USA).
SampleSiKCaMgAlFeCuNi
Ore16.930.193.996.185.3514.800.230.74
Bioleaching residueStage 1
(68 days)		25 °C15.440.203.685.544.9312.260.220.52
35 °C17.090.183.696.185.2812.670.190.50
45 °C13.660.173.085.054.2611.480.170.44
Stage 2
(140 days)		25 °C11.630.142.324.703.928.790.110.29
35 °C13.250.183.054.974.3111.630.160.32
45 °C14.630.182.984.994.109.390.090.20
SampleCoZnPSMnCrTi
Ore0.030.010.068.100.090.070.18
Bioleaching residueStage 1
(68 days)		25 °C0.020.010.066.370.090.060.18
35 °C0.020.010.046.430.090.060.17
45 °C0.020.010.055.160.080.070.20
Stage 2
(140 days)		25 °C0.020.010.036.020.060.040.14
35 °C0.020.010.056.740.070.050.18
45 °C0.000.010.056.010.080.060.15
Table 2. Results of experiments on the bioleaching of Cu–Ni sulfide products.
Table 2. Results of experiments on the bioleaching of Cu–Ni sulfide products.
ProductTreatmentExtraction, %Reference
NiCuPtPd
Sulfide copper–nickel ore
(4.25% Ni, 3.90% Cu)		Column bioleaching for 80 d
at 19 °C 		16.57.5n.s. 1n.s.[10]
Sulfide copper–nickel ore
(1.88% Ni, 1.76% Cu)		22.512.7n.s.n.s.
Low-grade sulfide copper–nickel concentrate (0.70% Cu, 2.30% Ni, 0.1 g/t Pt, 1.35 g/t PdStirred-tank reactor bioleaching for 40 d at 30 °C 70.014.0n.s.n.s.[11]
Stirred-tank reactor bioleaching for 40 d
at 40 °C and cyanidation		72.034.03888
Industrial sands
(0.32% Ni, 0.22% Cu)		Stirred-tank reactor bioleaching for 28 d at 30 °C 49.139.3n.s.n.s.[22]
1 Not studied.
Table 3. Potential advantages and disadvantages of the application of bioleaching + cyanidation for the treatment of Cu–Ni ores containing PGM.
Table 3. Potential advantages and disadvantages of the application of bioleaching + cyanidation for the treatment of Cu–Ni ores containing PGM.
AdvantageDisadvantage
Possibility of Cu/Ni and PGM recoveryComparatively low Cu and Pt recovery
Absence of the ore dressing stage and smeltingUse of cyanide solution for PGM leaching; potential leakage of bioleachate and cyanide solution
Comparatively low environmental impact (absence of gas emissions and low energy consumption)
Comparatively simple process; high CAPEX is not required; small deposits with low-grade ores/wastes may be treatedComparatively high duration of bioleaching
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Latyuk, E.; Goryachev, A.; Melamud, V.; Bulaev, A. Hydrometallurgical Processing of a Low-Grade Sulfide Copper–Nickel Ore Containing Pt and Pd. Processes 2024, 12, 1213. https://doi.org/10.3390/pr12061213

AMA Style

Latyuk E, Goryachev A, Melamud V, Bulaev A. Hydrometallurgical Processing of a Low-Grade Sulfide Copper–Nickel Ore Containing Pt and Pd. Processes. 2024; 12(6):1213. https://doi.org/10.3390/pr12061213

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Latyuk, Elena, Andrey Goryachev, Vitaliy Melamud, and Aleksandr Bulaev. 2024. "Hydrometallurgical Processing of a Low-Grade Sulfide Copper–Nickel Ore Containing Pt and Pd" Processes 12, no. 6: 1213. https://doi.org/10.3390/pr12061213

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