Browsing by Author "Moyo, Thandazile"

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    Open Access
    A study of alternative techniques to mercury amalgamation for gold extraction in artisanal and small-scale gold mining
    (2022) Manzila, Archippe Ngwey; Petersen, Jochen; Moyo, Thandazile
    Artisanal and small-scale gold mining (ASGM) has many definitions depending on the context. However, the common theme that characterises gold mining operations that fall within this category is that they make use of rudimentary methods to mine and process gold. The ASGM sector is a source of livelihood for millions of people worldwide and continues to grow due to the ever-rising demand for gold, and high unemployment rates which have been exacerbated by the Covid-19 pandemic, particularly in developing countries. Mercury amalgamation is the method of choice to recover gold in ASGM. This method consists of contacting the gold found within an ore with mercury to form an alloy i.e., the mercury-gold amalgam and subsequently burning off mercury to recover the gold in a form known as sponge gold. The popularity of this method has to do with its simplicity of application, low cost, and quick returns. However, mercury is a highly toxic substance; therefore, its use presents serious health risks for artisanal miners and their communities, and environmental risks for the ecosystems surrounding their operations. These risks arise primarily from the amalgam burning stage, whereby mercury is vapourised, and the dumping of mercury-rich tailings into local rivers. This mercury release affects human health by causing serious diseases that may lead to death. From an environmental perspective, mercury has been reported to severely pollute river ecosystems, inevitably finding its way to food chains. Due to these issues, alternative technologies such as borax smelting, the Gemini table, thiosulphate, cyanide, chlorine, and urea leaching, to name a few, have been developed or adapted over the years to substitute mercury. However, most of these technologies have not been successfully implemented in artisanal mining operations. This lack of success is primarily due to their complexity and high cost, making them unattractive to artisanal miners. This study investigates the application of cyanide and thiosulphate leaching as alternatives to mercury amalgamation for the recovery of gold in ASGM operations. Although cyanidation is practiced in ASGM, in some regions, it is only employed to treat tailings from the mercury amalgamation process. This is undesirable due to the fact that exposing mercury to cyanide results in the mobilisation of elemental mercury found in the tailings as mercury cyanide. This project investigates gold extractions that can be achieved with cyanide and alkaline thiosulphate systems and compares the results to those of mercury amalgamation. This investigation was undertaken by conducting leach experiments using cyanide at 1 g/L, 3 g/L and 5 g/L, and ammonium thiosulphate at 0.1 M and 0.5 M, on 3 ore samples originating from artisanal mining locations. The experiments were conducted using batch stirred tanks reactors and the operating conditions (T= 26°C, solids loading: 30%, particle size: --300 +150 µm) were selected to mimic as closely as possible the conditions of artisanal mining processes. The findings of the study revealed that cyanide leaching was the better performing technology compared to thiosulphate leaching as it achieved gold extractions of 71.6%, 69.7% and 67.8% for the 3 ores samples (Sample 1, Sample 2, and Sample 3, respectively) while thiosulphate leaching achieved gold extractions of 54.1%, 35.6% and 38.0% for the 3 ores, respectively. Studying the minerology of the ores, using XRF, XRD, QEMSCAN, SEM-EDS and diagnostic leach, revealed the presence of sulphide minerals hosting refractory gold which contributed to the low gold extractions observed. Cyanide leaching proved to be a system that is easier to control compared to thiosulphate leaching, making it much more attractive to artisanal miners. It is recognised that cyanide is a toxic chemical, however, the method is already practiced in ASGM and cannot be simply wished away. Instead, steps must be taken for its safe and responsible use. Hence, this research makes recommendations on avenues that can be explored to reduce the risks associated with cyanide use. It was also found that cyanide leaching outperformed mercury amalgamation which typically achieves gold recoveries of 30-50%. Thiosulphate leaching may be capable of achieving better gold recoveries than mercury amalgamation as well, as one of the ore samples achieved a gold extraction of 54.1%. However, this would depend on the ore type and reagent conditions as it was found that the 3 ore samples responded differently to leaching.
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    A Study Of Cyanide-Glycine Synergistic Lixiviant And The Igoli Process As Suitable Replacements For Mercury Amalgamation In Artisanal And Small-Scale Gold Mining
    (2023) Masuku, Wilson; Moyo, Thandazile; Petersen Joachim
    Artisanal and Small-scale Gold Mining (ASGM) operations are characterized by the use of rudimentary tools and technologies owing to limited access to capital. ASGM is predominantly a poverty-driven exercise practiced as a source of livelihood, typically in rural communities where people lack other employable skills. Globally, ASGM accounts for 20-25% of gold production, while at local scales, this number varies and can be as high as 65% in countries such as Ghana and Zimbabwe. Mercury is used in ASGM to capture gold from free-milling ores, in a process called mercury amalgamation. This is the go-to technology in most ASGM operations owing to its availability and ease of operation. However, mercury amalgamation has low recoveries in the range of 30-33% of gold from the otherwise rich gold ores typically mined in ASGM. In the amalgamation process, about 70% of the mercury used is lost to the environment with the amalgamation tailings and during the roasting process. Mercury is a toxic heavy metal, and mercury poisoning can lead to neurological and behavioural disorders and has been a major concern globally, leading to the signing of the Minamata Convention treaty. Mercury-free gold concentration and extraction methods such as shaking tables and roasting with borax have been put forward over the years, but their uptake has been very limited. The reasons for this poor uptake have never been systematically studied but it is thought that, among other reasons, it has to do with that some technologies are too complex for the ASGM context. Beyond the mercury-free technologies proposed for the ASGM sector, gold extraction and recovery in the large-scale mining sector has attracted researchers' attention for years, with a plethora of technologies having been proposed and tested. Little effort has been made to establish if any of these technologies could be a good fit for ASGM. In this study, two mercury-free technologies (cyanide-glycine lixiviant and the iGoli process) were tested to establish their effectiveness in the leaching of gold from ores sourced from two ASGM sites. The ores were characterized using QEMSCAN, XRD and XRF to identify mineral phases, and quartz was found to be the most dominant mineral. Sulphide minerals in both ores host the largest percentage of gold. The cyanideglycine lixiviant uses a combination of cyanide and glycine to improve gold extraction. The results from this showed that dissolution rate increases with an increase in glycine concentration in non-agitated systems at 3g/l NaCN while the reverse was true in agitated systems at the same cyanide concentration and when it was varied. The percentage of gold extracted in the non-agitated system after 72 hours was 36% at 5 g/l glycine, 21% at 2 g/l glycine and 19% when no glycine was added. In agitated systems at 5g/l cyanide, the highest extraction after 24 hours was 81% at 2 g/l glycine. Increasing glycine concentration led to lower gold extractions with 5 g/l and 10 g/l glycine extracting 74% and 68% respectively. This trend of decreasing extraction with an increase in glycine concentration was observed at different fixed cyanide concentrations i.e., at 1 g/l, 3 g/l and 5 g/l. The iGoli process uses hydrochloric acid and sodium hypochlorite to leach gold. The extractions were very low and reported below the detection limit of the analytical instrument, and thus they cannot be reported with confidence. However, iron was analyzed and showed a 55% extraction of the total iron in the ore. Results from these two technologies were compared to those of mercury amalgamation and benchmarked against the conventional cyanide process. Beyond the purely technical, a case study of two ASGM sites was done with the objective to observe and understand the day-to-day operations in a typical ASGM site and identify limitations and opportunities for mercury-free technology adoption. Based on insights drawn from the case studies, it was concluded that the cyanide-glycine lixiviant is relatively easy to implement given the current process operation in the ASGM sector which makes use of vat tanks that do not agitate the slurry (lixiviant + ore). However, the observed poor recoveries associated with the technology in non-agitated systems would be a limitation. When more profits are realized, the ASGM practitioners can upgrade to agitated systems and add hydrogen peroxide as an oxidizing agent to improve extraction.
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    Open Access
    A Study on the Applicability of Agitated Cyanide Leaching and Thiosulphate Leaching for Gold Extraction in Artisanal and Small-Scale Gold Mining
    (Multidisciplinary Digital Publishing Institute, 2022-10-14) Manzila, Archippe Ngwey; Moyo, Thandazile; Petersen, Jochen
    Mercury amalgamation is the method of choice to recover gold in artisanal and small-scale gold mining (ASGM). However, despite the low cost and simplicity of this method, the use of mercury presents serious health and environmental risks, as well as low efficiency in gold extraction. This study investigates the application of cyanide and thiosulphate leaching as alternatives to mercury amalgamation. This investigation was undertaken by conducting leach experiments using cyanide at 1 g/L, 3 g/L, and 5 g/L, and ammonium thiosulphate at 0.1 M and 0.5 M, on three ore samples originating from an artisanal mining area in Zimbabwe. The operating conditions (T = 26 °C, solids loading: 30%, particle size: −300 + 150 µm) were selected to mimic as closely as possible the conditions of artisanal mining processes. It was found that cyanide leaching was the better performing technology compared to thiosulphate leaching, as it achieved gold extractions of 71.6%, 69.7%, and 67.8% for the three ore samples (Sample 1, Sample 2, and Sample 3, respectively), whereas thiosulphate leaching achieved gold extractions of 54.1%, 35.6%, and 38.0% for the three ores, respectively. Both methods outperformed mercury amalgamation, which typically achieves gold recoveries of 30%–50%. Studying the minerology of the ores, using XRF, XRD, QEMSCAN, SEM-EDS, and a diagnostic leach, revealed the presence of sulphide minerals hosting refractory gold which contributed to the low gold extractions observed. Besides achieving higher gold extraction, cyanide leaching proved to be a system that is easier to control compared to thiosulphate leaching, making it much more attractive to artisanal miners.
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    Open Access
    Assessment of a Shredding Technology of Waste Printed Circuit Boards in preparation for Ammonia-based Copper leaching
    (2020) Prestele, Marc Patrick; Petersen, Jochen; Moyo, Thandazile
    The electronic waste (e-waste) stream grows at a global annual rate of 3-5%, with an expected 50 Mt to be discarded worldwide in 2020 alone. These large amounts of e-waste pose considerable environmental and health problems while also presenting socio-economic opportunities to most nations, especially to developing countries such as South Africa. E-waste presents a specifically unique challenge to developing nations as they suffer the challenges associated with e-waste, but do not have sufficient waste volumes to adopt business models used in developed countries to harness the economic opportunities presented by the growth of this waste streams. Recycling of e-waste requires huge capital and operating costs to run integrated recycling facilities and developing countries generally lack this funding. Furthermore, developing countries suffer from inadequate infrastructure, absent legislation and lacking capital investment which are necessary for the processing of e-waste regardless of it being regarded as a secondary resource or waste. Printed circuit boards (PCBs) are a valuable fraction of e-waste, made up of tightly laminated metal-polymer composites containing several base and precious metals which makes them attractive to recyclers. Hydrometallurgy is a widely explored technology that allows for scalable operations for recovering metals from PCBs. However, for it to be effectively employed, the metals in PCBs need to be liberated or be accessible to leach agents. To date, this still heavily relies on energy-intensive pulverisation prior to the leaching and subsequent metal recovery stages. This paper explores the structure of the PCB, developing an understanding of how the structural design of the board translates to the difficulty in liberating or exposing the metals for leaching. The paper goes further to test and compare metal liberation techniques as well as compares energy consumption and costs associated with the techniques; with the view to identify a low energy and low capital investment method that would be suitable for adoption by small scale recyclers typical of those operating in South Africa. The structural design of the PCBs was explored through an intensive literature survey and conducting a case study of the PCB manufacturing process of a local company as well as running tensile tests, drop weight impact tests and three-point bending tests on a batch of custom-made PCBs supplied by the local company. The metal liberation methods tested included the use of an industrial grab shredder to size reduce and delaminate the PCBs, use of a planetary ball mill and some instances including precursors such as freezing the PCBs in liquid nitrogen or soaking the boards in NaOH to remove the upper- and lowermost epoxy layers. The effectiveness of each method was then evaluated using a diagnostic ammoniacal leach test in which the extent of copper dissolution from the PCB is used as an indicator of the performance of the liberation method. Results on the structural design of the PCBs showed that it would be suitable to use size reduction mechanisms that are based on impact stresses as the fibreglass and epoxy could absorb all other stresses at high intensity without failing. In general, all treated or untreated PCBs underwent a maximum of six shredding passes, with results generally producing poor recoveries, not exceeding 27.5%. “Untreated” PCBs, referring to PCBs that only have undergone shredding in the industrial grab shredder, showed increasingly iv higher copper recoveries with consecutively shredding cycles. The 6th cycle produced the highest copper recoveries of 6.80g (23.5%) after 72 hrs. PCBs that had been soaked in NaOH and undergone six passes through the industrial grab shredder recovered a maximum of 27.5%. Interestingly, using a similar process but only shredding the PCBs in four passes showed similar results at 26.14% Cu recovery. Shredding the PCBs in four passes and subsequently milling them for 60 min (without NaOH treatment) showed lower Cu recoveries at 13.29% and this was not improved by extending the milling time to 120 min. This showed that the NaOH treatment was more effective in exposing the outer layers of copper relative to the shredding and milling. It can be seen that apart from size reduction there is delamination of some of the shredded PCB pieces. However, this delamination is not always complete and Cu metal can still be seen covered by fibreglass and hence inaccessible to leach agents. It is concluded that the combination of the shredding and NaOH method has potential and it is recommended to incorporate a 2nd NaOH stage to further delaminate the inner layers of the PCB exposing the copper
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    An electrochemical and leach study of the oxidative dissolution of chalcopyrite in ammoniacal solutions
    (2016) Moyo, Thandazile; Petersen Joachim
    Chalcopyrite is not only the most abundant of the copper sulphides, but also the most stable, making it recalcitrant to hydrometallurgical treatment processes especially in atmospheric leaching. Hence, pyrometallurgical processes are traditionally used to treat chalcopyrite concentrates. However, ore grades are falling and concentration processes are becoming increasingly costly, prompting need to revisit hydrometallurgical treatment processes (especially heap leaching), which are otherwise regarded as relatively economic and environmentally friendly. Key hydrometallurgical processes for chalcopyrite treatment are ferric sulphate, chloride and ammoniacal systems. The ferric sulphate system does not work well under atmospheric conditions, except in combination with thermophilic microorganisms, whereas the chloride system has only recently been evaluated more seriously for heap leach processes. The ammonia system remains relatively unexplored and most studies date back more than 40 years, but the system has considerable potential for further development. Ammonia systems can be effectively used to leach copper from chalcopyrite in the presence of an oxidant. The ammoniacal leaching system is heavily reliant on a good surface mass transfer system, hence it being widely studied in high pressure systems where oxygen was accepted to be the oxidant. Leach reactors were designed to use agitation systems which promote the abrasion of an iron based deposit layer thought to passivate the mineral surface. Most research on the ammonia leaching systems has previously been carried out in controlled or bulk leaching studies and only a few used electrochemical studies. A disconnect exits between the two approaches, resulting in different proposed fundamental reaction mechanisms and kinetic understanding. A fundamental electrochemical and controlled leach study of the oxidative leaching of chalcopyrite in ammoniacal solutions has been undertaken. The study covered the following aspects: a description of the mixed potentials, chemistry and kinetics of the anodic reaction, the cathodic reduction of the oxidants, the formation and effect of surface deposits and lastly a look at how results from electrochemical studies compare to those from the leaching of a similar mineral sample under similar solution conditions. A detailed study of the mixed potentials on a more or less pure chalcopyrite electrode has shown the redox reactions on the surface of the mineral to be controlled by the oxidation of chalcopyrite and reduction of copper(II). The presence of oxygen has been found to have no significant effect on mixed potentials in ammoniacal solutions in the presence of initial copper(II). Constant potential and potentiodynamic studies on the anodic reaction have shown the rate of the anodic reaction to increase with an increase in potential in a standard 1M ammonia/ammonium sulphate solution (which buffers at pH 9.6) in exponential fashion supporting conventional Butler-Volmer behaviour with a anodic transfer coefficient of 0.42 and a rate constant k* CuFeS2 of 0.0431 cms⁻¹. Increasing total ammonia increased the rate of reaction only at low concentrations; at higher concentrations increasing total ammonia had no effect on the anodic reaction. An increase of pH at fixed total ammonia concentration showed an increase in reaction rate, but the effect cannot clearly be discerned from the concomitant shift in relative proportion of free NH₃ and NH₄⁺. Coulometric studies have shown the oxidation reaction to proceed via the formation of a thiosulphate intermediate and this to be a 7-8 electron transfer reaction. A surface deposit layer consisting of iron, oxygen and small quantities of sulphur was formed and the sulphur component of this product layer was seen to be gradually depleted during leaching. Anodic currents were found to gradually decrease with time and this was linked to the growth of the surface deposit layer. However, the surface deposit layer did not passivate the anodic reaction; instead, it was proposed that the surface deposit layer adsorbed copper ions and displayed "ohmic" behaviour. The formation of the surface deposit layer was found to apparently promote the cathodic reduction of copper(II). While reduction of copper(II) was shown to be the primary reduction reaction, the presence of oxygen was seen to promote this reduction reaction through the regeneration of copper(II) in experiments that ran for longer time periods. An apparent accumulation of copper(I) on the mineral surface was seen to adversely affect the rate of the cathodic reaction and thus the overall rate of dissolution. The nature and morphology of the surface layer was found to be significantly influenced by the choice of cation in solution, which was thought to influence primarily the complexation/precipitation of ferric species forming near the surface. The degree of agitation during leach studies influences the rate of leaching due to the fragmentation of surface deposits, which are seen to slow the anodic reaction. A kinetic model has been developed for the anodic and cathodic reactions. This thesis presents significant new findings regarding the role of the copper(I)/copper(II) redox couple on the oxidative leaching of chalcopyrite. It also highlights the potentially limiting role of the cathodic reactions which have frequently been overshadowed by the focus on chalcopyrite oxidation reactions. Furthermore, the growth of a surface inhibiting layer which cannot be removed in heap leach systems due to the lack of mechanical agitation can now potentially be addressed by looking into the complexation and precipitation characteristics of cations in solution for ammoniacal leach systems.
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    Exploration of the thiosulphate process for the dissolution of gold from electronic waste and its recovery through ion-exchange
    (2022) Maharaj, Dasmi; Petersen, Jochen; Moyo, Thandazile
    Electrical and electronic equipment (EEE) has substantially grown over the past few years due to vast technological advancements and consequently so has electronic waste (e-waste). This growth has shown cause for concern with a generation of 53.6 million metric tonnes of e-waste globally in 2019 alone. Part of this concern may be due to the slow adoption of formal collection and recycling practices in developed countries whilst developing countries bear the brunt of the e-waste produced within the country and the e-waste imported from other countries. Furthermore, the burden of the increasing levels of e-waste has detrimental effects on the environment. In such cases, e-waste landfills have been known to leach metals, such as lead, into soil and groundwater of nearby regions thereby affecting human and animal life in the area. Despite the evident hazardous materials associated with e-waste, there is still value in this waste. The value is attributed to the metals such as gold, silver, copper and palladium found in e-waste. Printed circuit boards (PCBs) are a small but nonetheless crucial fraction of global e-waste making up 6% by mass. The gold content in this small fraction is of much higher concentrations than in typical primary gold ores thus PCBs represent the most economically attractive portion of e-waste. Hydrometallurgical processing has previously been applied for the recovery of gold and copper from primary ores possibly due to it being considered environmentally friendly. Therefore, this thesis investigates a hydrometallurgical process for recovering gold from PCBs. Literature studies show ammonium thiosulphate to be a viable option in comparison to the more widely used cyanide leaching route (Aylmore & Staunton, 2014). Thus, the ammonium-thiosulphate system containing ammonia, ammonium thiosulphate and copper (II) sulphate pentahydrate was incorporated as a hydrometallurgical option for the recovery of gold. This study explores the formulation of synthetic gold solutions using gold powder as well as its application to PCB gold leaching in the same ammonium-thiosulphate system. Gold recovery using ion-exchange processes was investigated with the use of a medium-base (AuRIX®100) and two strong-base anion exchange resins (Purogold™ MTA5013SO4 and MTA5011SO4). The AuRIX®100 resin was specifically developed for and is currently used in the gold-cyanide system. The MTA5011SO4 resin is currently used in the Barrick Goldstrike operation for thiosulphate systems and the MTA5013SO4 resin was produced by Purolite® for thiosulphate-copper systems. In addition, two eluants (ammonium nitrate and ammonia) were tested to effectively elute gold from the resin. Furthermore, copper was monitored throughout the leaching and ion exchange experimentation due to its catalytic effect in the ammonium-thiosulphate system. Determining the most appropriate dilution of the ammonium-thiosulphate solution after leaching to ensure solution stability in the periods between sampling time and analysis was explored as part of the investigation. In addition, gold powder dissolution from a 99.99% pure gold powder and gold leaching from PCBs were investigated. PCBs were cut by means of a bandsaw for the purposes of fitting them into a reactor whilst limiting copper liberation. Various additional background copper concentrations were introduced into the gold powder dissolution and leaching system to determine its effect on gold and copper extractions. For the ion exchange processes, capacity tests on all three resins were conducted to establish the resin operating capacities before loading and elution of synthetic gold solutions. Loading and elution tests were conducted at three flowrates (10 mL/min, 25 mL/min and 50 mL/min) and with two eluants; ammonium nitrate and ammonia. Kinetic and equilibrium experimental work was investigated on the MTA5013 resin with the addition of chloride ions as a competing anion to loaded gold on the active sites of the resin. The MTA5011 resin was introduced into the experimental work for confirmatory results of the MTA5013 resin. This was necessary as both the Purogold™ resins are similar with the exception of particle size and thus were expected to behave in the same manner. Gold concentrations of 141 ppm (representing 100% extraction) and 4.1 ppm (representing 91% extraction) after 24 hours was extracted for gold powder dissolution and PCB leaching experiments respectively. Background copper concentrations of 0.045 M and 0.1 M resulted in the highest gold and copper extraction values for gold powder and PCB respectively. In addition, the strong-base anion exchange resin (MTA5013) proved to be more suitable than the medium-base anion exchange resin for the ammonium-thiosulphate system. However, its low gold loading values of 8.06 meq/L (milliequivalents per litre of resin) and elution of 5.80 meq/L proved that it was ineffective in removing large gold amounts from the synthetic solution at a resin volume of 5 mL. Loading at low flowrates of 10 mL/min whilst eluting at 25 mL/min resulted in the highest loading and elution concentrations of both gold and copper. Copper loading and elution concentrations of 150 meq/L and 38.4 meq/L respectively were measured. This translated to 3.78% recovered in loading and 67.5% recovered in elution. High recoveries of gold and copper representing 75% and 67.5% respectively were achieved with ammonium nitrate as an eluant. Only 21.7% of the loaded gold was eluted when using ammonia as an eluant. Final kinetic and equilibrium test work revealed that the MTA5013 resin has an affinity for the aurothiosulphate ion over the weak chloride ion. This is attributed to low concentrations of chloride in solution therefore being ineffective in displacing the aurothiosulphate ion from the resin. Results from this study suggested that anions such as thiosulphate and tetrathionate competed strongly for the resin active sites and this was in agreement with Nicol & O'Malley (2002) who made the same postulation. Moreover, final gold loading concentrations were low (8.06 meq/L) given the operating capacities of the resin (0.77 eq/L). It was proposed that this may be due to other anions such as polythionates in the system occupying the active sites including the small resin volume. High concentrations of polythionates in solution compete for resin active sites despite the resin affinity for the aurothiosulphate ion over polythionates in the system. It is concluded from this study that the ammonium-thiosulphate was efficient in leaching gold, obtaining almost 100% extraction in PCB leached solutions and 100% in gold powder synthetic solutions. Strong-base anion resins were proven to be ineffective in obtaining high gold recoveries for the system at low resin volumes and the resin indicating a high concentration of polythionate attachment relative to gold. However, the resin did demonstrate a higher affinity for the aurothiosulphate ion relative to copper species in the solution and given a larger resin volume: solution volume ratio, high gold recoveries are possible. Furthermore, an ammonium nitrate eluant is considered appropriate in removing high concentrations of gold from the resin.
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    Exploring the potential for local end-processing of e-waste in South Africa
    (2019) Sadan, Zaynab; Petersen, Jochen; Moyo, Thandazile
    E-waste is one of the fastest-growing waste streams in the world, and South Africa (SA) is no exception. This is driven by increased consumer demand and access to electrical and electronic equipment, in addition to perceived equipment obsolescence, and rapid advancements in technology. E-waste recycling presents an opportunity in providing a source of secondary resources such as metals, plastics and glass, as well as employment and economic opportunities in both developed and developing countries. Furthermore, the diversion of this waste stream from landfills or dumps prevents additional land use and the potential negative impacts on human health and the environment. E-waste collection and upgrading is a relatively small-scale but growing industry in SA. Only 12% of e-waste generated was estimated to be recycled in 2015. Most of SA’s ewaste volumes are inaccessible due to lack of formal take-back schemes, lack of consumer awareness, as well as e-waste being kept in storage or disposed of in landfills. E-waste recyclers in SA generally carry out collection, dismantling and sorting, refurbishing, as well as pre-processing of value fractions. There is currently limited local end-processing capacity, therefore partially upgraded value fractions are prepared for export, while non-viable fractions are stockpiled or disposed of in local landfills. The business case for local end-processing of e-waste value fractions, particularly metals, does not seem feasible due to the inconsistent and insufficient volumes available. Furthermore, SA faces unique socio-economic challenges such as an unregulated yet well-established informal sector. Additionally, the legal framework presents many inhibitors for e-waste recyclers. This research study builds upon the knowledge that there is a limited understanding of the feasibility of existing e-waste end-processing technologies for implementation in the South African socio-economic and legislative context. Therefore, this research intends to find out what are the key barriers and enablers to implementing e-waste end-processing technologies in SA. Qualitative research methods were used to uncover the current challenges faced by local recyclers and other stakeholders in the value chain. The data collection thus took the form of interviews, site visits and desktop research. The findings show that the e-waste recycling industry in SA is undergoing many new developments in terms of research and investment interest, as well as the anticipated outcomes from the recent submissions of Industry Waste Management Plans (IndWMP). The industry shows potential as an emerging secondary resource economy, however, the extent to which it will mature is dependent on the organisation of its collection network as well as the development of local end-processing and manufacturing capacity. The collection network and infrastructure are currently supported by both informal and formal recyclers who provide a diversity of collection strategies and a wide network of ewaste sources. However, efforts to increase recycling rates by accessing volumes in storage and increasing consumer awareness and engagement is necessary. Besides the economies of scale required to support the development of local end-processing, alternative technologies to large-scale smelting should be considered for the SA context. While this is seen through initiatives by SA Precious Metals, end-processing technologies is still inaccessible to small and medium recyclers due to cash flow issues as well as cherry-picking of high-grade materials. Therefore, recyclers require further support in terms of dealing with non-viable fractions. This includes research and investments into technologies and business models for the recycling of low-value materials including plastics, as well as subsidies for the cost of safe disposal or treatment of these fractions. Additionally, acquisition of product markets and an increase in manufacturing capacity is necessary to accelerate industry development. The legislative framework also poses limitations on recyclers in the e-waste value chain, stemming from the legal definition of e-waste as a liability as opposed to a resource. While the legislation is unlikely to change, provisions to relieve any legal barriers should be implemented. This includes permissions for pilot projects to test new technologies, as well as legal support for smaller recyclers in the form of consultancy as well as guidelines for sustainable waste management practices should be provided. Finally, while there are many challenges present in the e-waste recycling industry today, the IndWMP offers an opportunity for collaboration between key stakeholders, including the relevant government bodies. Plans have been submitted and the outcomes of approved plans will be revealed at the beginning of 2019. The plans offer solutions for recycling subsidies, increasing the collection and recycling rates, as well as investment into technology, research and enterprise development. However, successful implementation of these plans will only occur if integration and collaboration of the local e-waste community prevail over greed and the struggle for power.
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    Investigating Early-Stage Process Flow and Reactor Sequencing to Maximise Gold Extraction ln the Thiosulphate Leaching of Waste Printed Circuit Boards
    (2023) Gonte, Melissa; Moyo, Thandazile
    Fast growing volumes of waste electronic and electrical equipment (WEEE) present a rising environmental challenge while offering opportunities for creating a circular economy. Once dismantled and sorted, waste printed circuit boards (PCBs) are one of the value-bearing fractions of WEEE. Waste PCBs contain valuable metals like copper (Cu) and gold (Au) that can be profitably recovered using metallurgical methods like hydrometallurgy, pyrometallurgy, or a combination of the two. Successful implementation of pyrometallurgical technologies have been demonstrated in large-scale integrated recycling operations. These require high capital investment, large volumes of waste, and advanced scrubber/filter equipment to combat toxic flue gas pollution. Hydrometallurgical technologies present a viable processing route for the African context, where waste volumes are relatively low and energy supply is uncertain and expensive. Many lixiviants, such as cyanide, halides and aqua regia can be used to extract Au from waste PCBs. Due to its great selectivity and relatively low toxicity, this research focuses on the ammonium thiosulphate chemistry for the extraction of Au from waste PCBs. While ammonium thiosulphate solutions are effective in dissolving Au, there are challenges regarding reagent consumption when the system is used on materials that contain significant quantities of Cu. This is because Cu dissolves preferentially and plays a role in catalysing the breakdown of the thiosulphate ion. The method of pre-treatment and the sequence in which metals are leached in a multi-stage leach process determine the presence of co-existing metals, especially Cu for dissolution from waste PCBs in an ammonia thiosulphate system. In this work we hypothesized that the extraction of Au prior to delamination and size reduction will reduce the loss of Au, owing to Au being situated on the topmost layer of discarded PCBs. It would also result in less exposure of Cu to the leaching system which would ultimately limit reagent consuming side reactions. Four potential leaching sequences were examined to evaluate Au extraction, Au loss and base metal (BM) (Cu and Ni) co-extraction. Sequences A and B involved Au leaching of cut (A) and shredded (B) PCBs using a 0.5M ammonium thiosulphate ((NH4 )2S2O3) in the presence of 0.04M copper sulphate (CuSO4) and 1M ammonia (NH3), at a solid/liquid ratio of 100g/L. Sequences C and D involved BM extraction from cut (C) and shredded (D) PCBs prior to Au leaching, under the same conditions. iv The results showed that sequence A had the highest Au extraction of 97% Au. However, the coextraction of Cu and Ni, was also high at 21% Cu and 96% Ni, respectively. Nevertheless, in this study, this was preferable due to the considerable Au loss that other sequences experienced. Sequence D, B, and C, each suffered an overall Au loss of 53% , 26% and 20%, respectively. Leaching of Au leaf (93.1% Au) in the presence of predetermined amounts of Cu and Ni in 0.5M (NH4)2S2O3 and 0.5M sodium cyanide (NaCN) was used to simulate the leaching of Au from waste PCBs. Within the first hour of leaching, a high Au extraction of 99% and 88% was achieved in (NH4)2S2O3 and NaCN solutions, respectively, in the absence of background Cu and Ni. Addition of predetermined quantities of Cu (11500 mg) and Ni (519 mg) in both lixiviants resulted in a decline in Au extraction to 42% and 37%, respectively. The actual leaching of PCB in the same concentrations of (NH4)2S2O3 and NaCN gave 97% and 38% Au extraction, respectively. The optimisation tests (of sequence A) showed that a 0.5M(NH4)2S2O3 and 1M NH3 lixiviant concentration and a solid to liquid (S/L) ratio of 100g/L was optimal for leaching. Under these conditions 99% Au was extracted within 7.5 hours. In comparison, 36% Au extraction was achieved when NaCN was used for leaching under the same conditions. This showed that it is difficult to implement the optimal sequence, obtained for (NH4)2S2O3 leaching, in the leaching of waste PCBs using NaCN. NaCN suffers from the interference of foreign ions as side reactions compete for the reagent. Other than NaCN's level of toxicity, the results suggest that this system also requires implementation of the prior removal of BMs before leaching of Au. In conclusion, the highest Au extraction occurs when Au is extracted from PCBs using sequence A (before delamination, aggressive size reduction, and BMs extraction). Sequence A's low coextraction of Cu demonstrates that most of the Cu remains interlocked within the boards during Au extraction.
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    Investigation of a hydrometallurgical process route to recover metals from waste printed circuit boards
    (2019) Chirume, Blessing Hellen; Petersen, Jochen; Moyo, Thandazile
    The loss of valuable materials such as base and precious metals is increasing due to the increase in waste electronic and electric waste (WEEE). Most of these metals in WEEE are on the printed circuit boards (PCBs). This study aims to compare different pre-treatment methods to recycle copper from PCBs using a hydrometallurgical process. In order to obtain a uniform/consistent sample across all the tests done, similar custom-made PCBs with 55.45% wt copper were used to compare different parameters. Pre-treating the PCBs is the first stage of the process and it is done to liberate metals which are then dissolved in subsequent leaching stages. Eight different pre-treatment methods were explored. The pre-treated PCBs were then leached under similar conditions in a diagnostic leach test in order to get an indication of the effectiveness of the pre-treatment. Copper recoveries corresponding to each of the pretreatment methods were compared. In addition to recovery, other factors such as time taken for copper recovery, material losses incurred, practicability, environmental impact, health and safety were used to compare the pre-treatment methods. A score was given for each factor and the average was used to choose the optimal pre-treatment method. A method where the PCBs were cut into 1.5 cm x 2 cm pieces and then soaked in 2 M NaOH at 40 °C for 24 hours had the highest average score. This pre-treatment method was then used to prepare PCBs that were used for test work done with the aim to optimise copper leaching. The influence of total ammonia concentration, liquid to solid ratio and choice of ammonium salt used in the buffer system, were investigated in the copper leaching optimisation stage of this thesis. Using ammonium carbonate resulted in lower recoveries compared to ammonium sulphate in the diagnostic leach test. Increasing the ammonia concentration to 7M did not have a significant effect on the copper recovery. Decreasing the liquid to solid ratio from 20ml/g to 10ml/g resulted in a slower rate of recovery. The optimal leaching conditions were found to be; 750 ml mixture of 4 M NH3, 2 M (NH4)2SO4, 100ppm CuSO4 at 25 °C and 500 rpm using the optimal pre-treatment method for the PCBs.
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    Study of the dissolution of chalcopyrite in solutions of different ammonium salts
    (SAIMM, 2017-09-11) Moyo, Thandazile; Petersen, Jochen
    The oxidative leaching of chalcopyrite in ammoniacal solutions has been evaluated using electro-analytical techniques. The anodic dissolution process has been established to be a seven-electron transfer process under nitrogen in ammonia–ammonium sulphate solutions and ammonia–ammonium carbonate solutions. The number of electrons transferred in the carbonate and sulphate salts suggests that the sulphur is oxidised to a thiosulphate intermediate and the copper and iron released as Cu+ and Fe2+. The Cu+ and Fe2+ is subsequently oxidised in solution in non-Faradaic reactions. The deportment of Fe2+ and S2O32- is affected by choice of the ammonium salt used in the leaching process. In the perchlorate salt, only five electrons are transferred, however observations made on the mineral surface after leaching do not support the formation of elemental sulphur. Scanning electron microscopy and energy-dispersive spectroscopy analysis of the mineral surface suggest presence of an iron–sulphur surface layer completely free of copper under the ammonia–ammonium sulphate conditions, an iron-rich surface layer under ammonia–ammonium perchlorate solutions and absence of surface layer build up under ammonia–ammonium carbonate solutions.