VOLUME 121
NO. 7
JULY 2021
CELEBRATING EXCELLENCE 1961–2021 UNIVERSITY OF PRETORIA
TOP 50
Department of Mining Engineering Faculty of Engineering, Built Environment and Information Technology Fakulteit Ingenieurswese, Bou-omgewing en Inligtingtegnologie / Lefapha la Boetšenere, Tikologo ya Kago le Theknolotši ya Tshedimošo
of universities ranked globally for MINERALS AND MINING ENGINEERING 2020/21 QS World University Subject Rankings
A reimagined future through excellence in mining engineering education As the University of Pretoria’s Department of Mining Engineering celebrates its diamond anniversary in 2021, it stands on the threshold of a reimagined mining industry.
The Department’s research efforts are concentrated on growing capacity in mechanisation and automation, rock-breaking and explosives engineering, management and leadership, and rock engineering. This research is not only aimed at ensuring the sustainability of the industry, but also at eradicating poverty, unemployment and inequality. Research related to mechanisation and automation will help restore the competitiveness of the mining industry and ensure that the bulk of the country’s mineral resources can be profitably extracted. This will require a substantial reinvestment in technology that will build on improvement and modernisation efforts towards fully autonomous, non-explosive, remote mining environments. It responds to an urgent need to develop nextgeneration mining systems, especially systems that will enable the mining of deeper narrow reef, hard-rock commodities, such as platinum and gold. Systems to make current mining operations safer, healthier, more productive and sustainable also need to be developed. This includes the digitisation of
mining operations to keep abreast with international developments in the areas of the Internet of Things and Industry 4.0-type applications. The future of mining in South Africa will see people placed on the foreground as a vital success factor alongside viable ore bodies, and well-developed and optimised operations. The Department’s approach is to make mining more sustainable by ensuring that the next generation of mining engineers are leaders who will boldly take the industry into the future. Leadership development has been on the Department’s agenda for the past decade, and its strategic importance will continue to be highlighted as the Department enters the next decade of mining education. The future impact of a graduate from the University of Pretoria thus lies in developing a shared vision of how technology can better benefit economies, societies and the human condition. The true impact of the 4IR lies in collaboration, and this is the foundation of sustaining the industry and equipping the next generation of mining engineers for an era beyond the 4IR.
Faculty of Engineering, Built Environment and Information Technology Fakulteit Ingenieurswese, Bou-omgewing en Inligtingtegnologie / Lefapha la Boetšenere, Tikologo ya Kago le Theknolotši ya Tshedimošo
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VOLUME 121 NO. 7 JULY 2021
Contents Journal Comment: Every crisis presents an opportunity K.M. Letsoalo. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . iv President’s Corner: Celebrating Ethical Leaders by V.G. Duke . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . v
News of General Interest Obituary . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . vi
STUDENT PAPERS Comparison of the mechanical properties of Grade 5 and Grade 23 Ti6Al4V for wire-arc additive manufacturing L. Mashigo, H. Möller, and C. Gassmann . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 325 A challenge of additive manufacturing is anisotropic differences in mechanical properties. The mechanical properties of Grade 5 and Grade 23 Ti6Al4V were compared for this application. Samples were extracted from WAAM-produced Ti6Al4V walls in different directions and at different positions. Differences in strength, hardness, and ductility were attributed to different microstructures resulting from thermal cycling during manufacturingwhich. Beneficiation of recycled process water at DRDGOLD’s ERGO plant, and its effect on gold recovery A. Narain, J. H. Potgieter, G. E. Rencken, and J. Smith. . . . . . . . . . . . . . . . . . . . 331 DRDGOLD, a South African gold producer re-treating surface tailings, has transitioned to a fully closed water circuit. Consequently, the accumulation of contaminants and reagents has led to changes in water composition, compromising leach performance and overall gold recovery. A study of the impact of different contaminants on gold recovery showed varying results. Significant differences were found in gold recoveries when using Rand Water and untreated process water. However, treated process water yielded similar gold recoveries to Rand Water.
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STUDENT PAPERS (CONTINUED) Effects processing parameters and building orientation on the microstructural and mechanical properties of AlSi10Mg parts printed by selective laser melting C. Phetolo, V. Matjeke, and J. van der Merwe . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 335 The mechanical properties of AlSi10Mg alloy were studied to determine the effect of processing parameters and building orientation on the mechanical properties. The microstructures and mechanical properties were characterized in the XY and Z building directions. The UTS and higher yield strength wre found to be higher in the Z orientation than in the XY orientation. Fractographic analysis revealed that crack initiation in both orientations started from the surface in a brittle manner due to surface flows, and then propagated via microvoid coalescence.
PROFESSIONAL TECHNICAL AND SCIENTIFIC PAPERS Sensitivity of shotcrete Brazilian discs coated with thin spray-on liners to environmental conditions M.J. Kanda and T.R. Stacey. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 341 Thin spray-on liners (TSLs) have been used as sealants and rock support in tunnels for over 25 years. Although laboratory tests have indicated satisfactory properties, this is not always demonstrated in actual mine environmental conditions. We investigated TSL performance in conditions similar to actual mine conditions, using Brazilian indirect tensile (BIT) specimens prepared from precast shotcrete. The results showed that environmental conditions have a significant influence on the tensile strength enhancement of shotcrete by TSLs. Water-based TSLs are most likely to be suitable for high humidity environments, although their performance decreases at higher temperatures. Numerical modelling confirmed the potential limitations of designing TSL support based only on laboratory testing under room conditions. Feasibility of tailings retreatment to unlock value and create environmental sustainability of the Louis Moore tailings dump near Giyani, South Africa N.K. Singo and J.D. Kramers. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 351 This study explores the possibility of reclaiming the tailings dump at the Louis Moore Mine in the Giyani Greenstone Belt, investigates potential hazards to communities in the vicinity, and identifies mitigation strategies. It was concluded that the residual Au reserve in the Louis Moore tailings dump is currently 0.20 t. Reworking the tailings would be viable, despite a potential environmental risk posed by elevated arsenic (As) concentrations. Further exploration is, however, required. Improving the environmental and economic aspects of blasting in surface mining by using stemming plugs A. Ur Rehman, M.Z. Emad, and M.U. Khan. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 359 The use of stemming plugs in surface blasting is gaining popularity, however, evaluations of the economic and ergonomic impacts are lacking thereby hindering their application in many mining operations. This paper evaluated the effectiveness of three types of stemming plugs. The results show that stemming plugs reduce the need for secondary blasting and increase blast performance. An economic analysis showed that the incorporation of stemming plugs can reduce blasting costs significantly.
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nal
Jour
ment
Com
Every crisis presents an opportunity
I
t is often said that every crisis presents an opportunity, and Covid-19 is yet another example. The South African mining sector has over recent years initiated discussion and conversation around technology and the fourth industrial revolution (4IR). This comes at a time where South Africa’s mining productivity has declined by over 7.6% in the last decade, with two-thirds of the mining output sitting in the upper half of the global mining cost curve. The discussions have always been a challenge, with various stakeholders highlighting the complexity of the South African mineral deposits, particularly the narrow reef deposits of gold and platinum. Another challenge that often arises is the interpretation of what a modern mine should look like in the South African context, as information in the public domain is focused on massive mines and trackless mobile machinery, which excludes a number of operations in the country. Covid-19 regulations in the South African mining sector have forced the industry to rethink mining as it was previously known, with a reduced number of employees now being permitted on site and constant monitoring of the possible spread of the virus. Mines were forced to think of creative ways around communications and interconnectivity across various points within the mining value chain. These efforts have stimulated the industry to further embrace elements of 4IR such as artificial intelligence (AI) through the mapping of Covid-19 hot-spots, the Internet of Things (IoT), cloud computing, and advanced wireless technologies through the integration of on-mine reporting systems and overall communications. PWC found that most South African mining companies invest in technology to increase efficiency while lowering costs and improving health and safety. The challenge, however, is that 69% of South African mining companies were considered digital followers, with only 6% being digital champions who have fully integrated their technology. A significant amount of work is still required from CEOs to inspire more confidence within mining companies to allow organizations to be more in the forefront of technology development. This is especially relevant given the nature and uniqueness of some of the gold and platinum mineral deposits. Although one could argue that some of the changes are not extreme, given where South African operations are with their technology journey, some are extreme – and this is a step in the right direction. Covid-19 has made it more evident that 4IR is more than a necessity: it will be a key enabler for the South African mining sector to take its place in the world of global competitiveness, not forgetting its key roles of safety and environmental responsibility.
K.M. Letsoalo
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President’s Corner
Celebrating Ethical Leaders Reminiscing on the Purpose of ‘The first step in the evolution of ethics is a sense of solidarity with otherAssociations human beings.’ – Mining Albert Schweitzer
A
year ago, I presented on the importance of ethical behaviour and ethical leadership at the Annual General Meeting. I believed this to be particularly relevant, given that South Africans often find themselves in positions of uncertainty and insecurity in our volatile political, economic and social environment. In light of the recent unrest, and prior to ending my term as President of the SAIMM, I feel it necessary to again emphasise the importance of ethical leadership in the South African context.
July witnessed a tragic development in the political arena. Stores were ransacked and destroyed by people displaying callous attitudes that confirmed the divisions within our political mix. This unrest also exposed the extent of the ever-increasing wealth gaps across South African society. For many hard-working individuals, the immediate effect was immense, with shop owners and assistants losing everything, and in some cases even their lives. None of us were completely unaffected by the fear and uncertainty that gripped the nation. Unemployment is reported at 32.6%, but the reality is that many more people are without work. Our GDP figures indicate little to no real growth since 2014 in 2010 terms (see graph). Our country was already in a precarious economic situation prior to the onset of the COVID-19 pandemic, and subsequently, approximately half a million more people lost their jobs as a consequence of the hard lockdown. This has worsened the degree of wealth inequality between social classes and there is evidence of a rise in domestic violence. The political, economic, and social situation in South Africa requires strong intervention if we are to strengthen our global standing, and more importantly, if we are to increase ethical awareness and solidarity among communities. It is within this context that I need to express how proud I am to be South African. Our country witnessed an inspiring coming-together of communities seeking to help one another, and where Real GDP (constant 2010 prices, seasonally adjusted). Source: Stats necessary, to protect their local SA - Gross Domestic Product (GDP) stores from looters. Many caring, honest, and respectful individuals simply connected with people to do what they truly believed to be right – opposing ethical violations. In addition, we have seen how leaders in our minerals industry, despite having to navigate numerous challenges over recent years, have remained steadfast in their commitment to promoting diversity and inclusivity in the workplace, while also supporting their employees, local communities, and our country. Decisive actions have resulted in positive and meaningful progress on important matters outside of the immediate business of making profits. The timely implementation of much-needed vaccination programmes is one good example. There are clearly many individuals in our country who have integrity and are able to display ethical leadership. They inspire a sense of community and team spirit within our businesses and our communities. The South African government can be comforted by the fact that it has the support of the people directing the fortunes of our minerals industry. I believe that we have been tested again over the past month, and that we will once more emerge as a much stronger and more resilient democracy on the other side. V.G. Duke President, SAIMM
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OBITUARY Navin Singh was born on 14 November 1971 in Durban, KwaZulu-Natal. He started his career at South Deep Gold Mine in 1996 as Strata Control Officer, progressing to the position of Chief Rock Engineer (Operations). In 2000 he relocated to Australia, and joined Western Mining Corporation as Senior Geotechnical Engineer at Olympic Dam Mine in South Australia. He returned to South Africa in 2001 to take up the position of Research Manager for Rock Engineering at the CSIR, and subsequently that of Programme Manager for Mining. He served for a time as New Technology Manager at Gold Fields Ltd, then moved to the Mine Health & Safety Council as Chief Research and Operations Officer, remaining in that position for the next 6 years. Navin returned to the CSIR in 2015 as Manager for Mining R&D. He was involved in establishing the Mandela Mining Precinct in partnership with the DST and the then Chamber of Mines, and became the Co-Director of the Precinct for the period 2016–2020. He held the position of Head of Technology, Kumba Iron Ore (Anglo American) when he passed on. Navin served as a Director on Coaltech Research Association (2015–2020), and also of the Mining Equipment Manufacturers of South Africa (2017–2020). Navin was a Fellow of the SAIMM and served as a Council member from 2018. He was also a Fellow of South African National Institute of Rock Engineering (SANIRE). When asked about his contribution to the SAIMM Council, Navin’s response was ’I am actively involved in driving improvements in mining through technology development, implementation, and adoption, and since SAIMM strives to bring technical leadership to the mining industry through knowledge dissemination, my involvement in this space is thus fully aligned’. His contribution to the SAIMM will be sorely missed. Navin’s passing on Monday, 12 July 2021 was a huge loss to all those that knew him. Our heartfelt condolences go to his family, friends, and colleagues.
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Comparison of the mechanical properties of Grade 5 and Grade 23 Ti6Al4V for wire-arc additive manufacturing L. Mashigo*1, H. Möller1, and C. Gassmann2 Affiliation: 1 Department of Materials Science and Metallurgical Engineering, University of Pretoria, South Africa. 2 GEFERTEC GmbH,Germany. Correspondence to: H. Möller
Email:
hein@heinmoller.co.za
Dates:
Received: 29 Jan. 2021 Revised: 15 Jul. 2021 Accepted: 21 Jul. 2021 Published: July 2021
How to cite:
Mashigo, L., Möller, H., and Gassmann, C. 2021 Comparison of the mechanical properties of Grade 5 and Grade 23 Ti6Al4V for wire-arc additive manufacturing. Journal of the Southern African Insitute of Mining and Metallurgy, vol. 121, no. 7, pp. 325–330 DOI ID: http://dx.doi.org/10.17159/24119717/1498/2021 ORCID: H. Möller https://orcid.org/0000-00016075-9965 L. Mashigo https://orcid.org/0000-00020391-2799 C. Gassmann https://orcid.org/0000-00033379-6618
Paper written on project work carried out in partial fulfilment of B.Eng (Metallurgical Engineering) degree *
Synopsis Wire-arc additive manufacturing (WAAM) is a directed-energy deposition technology that uses arc welding procedures to produce computer-aided designed parts, such as three-dimensional printed metal components. A challenge of additive manufacturing is the anisotropy. Interstitial elements play a significant role in the mechanical properties of Ti6Al4V of different grades. In this research, the mechanical properties of Grade 5 and Grade 23 Ti6Al4V were compared for this application. Samples were extracted from WAAM-produced Ti6Al4V walls in different directions (horizontal and vertical) and at different positions (top and bottom). The samples were subjected to optical microscopy and tensile and hardness tests. Grade 5 Ti6Al4V samples were found to have greater strength, greater hardness, and lower ductility, owing to the higher content of interstitial elements compared with Grade 23. The bottom samples had higher strength than the top samples, which is attributed to thermal cycling during manufacturing, resulting in different microstructures. Keywords Ti6Al4V, wire-arc additive manufacturing, anisotropy, heat accumulation, interstitial elements.
Introduction Additive manufacturing has been growing exponentially in the fabrication industry since its inception in the 1980s. Conventionally, subtractive fabrication has been more broadly utilized, in which raw material is subjected to a machining process to remove unwanted material to produce the designed component (Li et al., 2019). Additive manufacturing, in contrast, has the advantage of decreasing the cost and time of manufacturing because it directly produces the designed component without wasting any material (Antonysamy, 2012).
Wire-arc additive manufacturing (WAAM) Of the diverse additive manufacturing procedures, wire-arc additive manufacturing (WAAM) has the advantage of manufacturing complex shapes with greater material usage efficiency, and has one of the highest deposition rates. WAAM is a directed-energy deposition procedure that employs arc welding to produce components using a three-dimensional (3D) metal printer (AMFG, 2018). In the WAAM process, a component is designed using computer-aided design (CAD) software and the design is transferred to a 3D printing machine. A metal wire is melted onto a substrate using an electric arc as the heat source, and the component is manufactured in a layer-by-layer deposition process. After production, the printed component undergoes 3D geometrical scanning for quality control, followed by surface finishing (AMFG, 2018). WAAM-manufactured components have a high surface roughness and contain residual stresses and distortion because of the relatively high heat input (AMFG, 2018). Other techniques of additive manufacturing have been used to produce metal components. The laser powder bed fusion (L-PBF) process is more prone to form defects, such as porosity (Thuketana et al., 2020), which is less likely with WAAM (Biswal et al., 2019). Benson (2012) addressed the risks associated with metal powders and safety considerations when handling them. Owing to their high surface-to-volume ratio, powders are naturally more reactive than bulk materials; and incorrect handling can potentially lead to fires and explosions. The wire used in the WAAM process is safer to handle than powder.
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Comparison of the mechanical properties of Grade 5 and Grade 23 Ti6Al4V Ti6Al4V Ti6Al4V is used in various industries and is the most studied titanium alloy in the additive manufacturing sector. It exhibits a good balance of strength, ductility, and fatigue resistance. Ti6Al4V comprises an aluminium-stabilized α phase and vanadium-stabilized β phase (Carroll, Palmer, and Beese, 2015), and the transformation from β to α plays a significant role in this alloy and affects the final microstructure. This transformation is most influenced by the cooling rate and alloy composition (Antonysamy, 2012).
Microstructure The thermal history of a process determines the development of the phases in a structure, which in α+β Ti6Al4V can include primary α, lathlike α, colony α, hcp martensite α (α’), grain boundary α, acicular α, and prior β phases (Bintao et al., 2018). During WAAM, the titanium substrate is heated to its liquidus temperature as the Ti6Al4V wire melts at the fusion zone to form a weld bead. At the moment of heating, immediately below the fusion zone, α and β grains in the substrate heterogeneously nucleate, return to the prior completely β structure, and undergo faster grain growth. Grown β grains at the edge of the fusion zone serve as nucleation sites at which the solidification front epitaxially grows back through to the weld pool, where coarsening grain structures form as a continuation of the grains around the fusion zone. Thus, in the next layer, β grains keep growing from the coarse β grains formed in the previous layer (Wang et al., 2012). The solidification process during WAAM of Ti6Al4V promotes the epitaxial growth of columnar grains. These columnar grains influence the tensile strength, affect the mechanical properties, and play a role in the tensile anisotropy that develops in the vertical and horizontal directions of the same specimen (Bermingham, McDonald, and Dargusch, 2018). Bintao et al. (2018) found that, although factors such as the solidification microstructure, grain size, and morphology are dependent on the thermal history during WAAM, heat accumulation also has an influence, therefore understanding the impact of heat accumulation can improve process control and optimization.
Anisotropy Microstructural elements, such as β grain size, the thickness of α grain boundaries, size of primary α grains, and the presence of martensitic α grains, play a role in the mechanical properties of Ti6Al4V (Wang et al., 2012). A study of anisotropy in the corrosion behaviour of Ti6Al4V produced by WAAM showed that the vertical plane had higher corrosion resistance than the horizontal plane (Wu et al., 2018). These values are influenced by the grain sizes and phase orientation in the planes. Zhang et al. (2016) identified anisotropy in a study of fracture toughness and fatigue crack growth rate in Ti6Al4V produced by WAAM. The vertical plane (across the layers) had higher fracture toughness than the horizontal plane (along the layers). Grade 23 Ti6Al4V also had higher fracture toughness than Grade 5 (Zhang et al., 2016).
alloy (Azom, 2002). Interstitial elements, such as oxygen and nitrogen, play a role in microstructures because they are strong α-stabilizers and influence the α-to-β transition temperature (Antonysamy, 2012). Oxygen in titanium has a notable hardening effect (Bauristhene, Mutombo, and Stumpf, 2013; Oh, et al., 2011). Considering the role of interstitial elements in influencing anisotropy, the objective of this study was to investigate the mechanical properties of Grade 5 and Grade 23 Ti6Al4V printed walls produced using WAAM, in different directions relative to the building direction and at different positions in the walls.
Experimental methods Wire-arc additive manufacturing Grade 5 and Grade 23 Ti6Al4V wires with a diameter of 1.2 mm were deposited to produce 3D printed walls with dimensions of 195 mm × 125 mm × 31 mm using 3DMP® technology (Gefertec, Germany). The process parameters are provided in Table I. The wires were deposited by the oscillation strategy, as shown in Figure 1. The walls were stress-relief heat-treated in a Xerion vacuum furnace at 650°C for 2.5 hours, followed by furnace cooling. Standard specifications of Ti6Al4V wires and chemical compositions of the wires are shown in Table II; photographs of the walls are shown in Figure 2.
Metallographic preparation and microscopy Metallographic samples were prepared according to ASTM E3 requirements. Ti6Al4V samples cut from the walls were mounted and sequentially ground using P200, P600, P800, P1200, and P2500 SiC abrasive papers with water as the lubricant, polished using colloidal silica solution, and then etched using Kroll’s reagent. An Olympus BX51M optical microscope was used to examine the microstructures at a magnification of 50×s.
Hardness testing Sample hardness was measured using an HR-150A Vickers microhardness machine at a constant load of 2 kg force for 15 seconds.
Tensile testing Flat tensile samples were extracted as shown in Figure 3a and machined using a xenon wire-cutting machine. They were extracted in the vertical (ZX) and horizontal (XY) planes, as Table I
Process parameters for WAAM deposition Current Voltage Travel speed Wire feed speed
180 A 16 V 1.8 m/min 9.7 m/min
Grade 5 and Grade 23 Ti6Al4V The main difference between Grade 23 (also known as Ti6Al4V ELI) and Grade 5 Ti6Al4V is the reduced content of interstitial elements in the former. Low interstitial element content improves ductility and fracture toughness, but reduces the strength of the
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Figure 1—Schematic of the deposition strategy for WAAM Ti6Al4V walls The Journal of the Southern African Institute of Mining and Metallurgy
Comparison of the mechanical properties of Grade 5 and Grade 23 Ti6Al4V Table II
Standard specifications and chemical compositions of Ti6Al4V wires used in additive manufacturing (mass%) Grade 5 (ASTM B863 − 19) Grade 23 (ASTM B863 − 19) Grade 5 Grade 23
Ti Al V Fe C O N H Bal. Bal. Bal. Bal.
5.5–6.75 5.5–6.5 5.7 6.225
3.5–4.5 3.5–4.5 4.15 3.95
<0.25 <0.25 0.125 0.13
<0.08 <0.08 0.015 0.006
<0.20 <0.13 0.17 0.04
<0.05 <0.03 0.01 0.003
<0.015 <0.013 0.0012 0.0011
Figure 2—(a) Ti6Al4V Grade 5 and (b) Ti6Al4V Grade 23 walls
Figure 3—(a) Machined wall and (b) positions of machining of samples
shown in Figure 3b. The tensile samples had a gauge length of 30 mm and a cross-section of 6 mm × 4 mm. Tensile tests were conducted at a strain rate of 2 mm/min according to ASTM E8 requirements using an MTS Criterion Model 45 instrument.
Results and discussion Tensile testing Figure 4 shows that Grade 5 had greater tensile (Rm) and yield The Journal of the Southern African Institute of Mining and Metallurgy
strength (Rp) than Grade 23 and lower elongation (A25) in the ZX plane, owing to its higher concentration of interstitial elements (Table II). The horizontal XY plane had higher tensile strength than the vertical ZX plane for both grades. This was influenced by the build direction (Z) of the walls, ZX being in the plane in which the walls were built and XY perpendicular to the build planes. The thermal cycle also played a role. The vertical direction was more ductile, and the difference between the two directions was greater for Grade 23. The standard deviation (SD) VOLUME 121
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Comparison of the mechanical properties of Grade 5 and Grade 23 Ti6Al4V
Figure 4—Ti6Al4V Grade 5 vs Grade 23 tensile test results at the horizontal XY and vertical ZX planes showing ultimate tensile strength (Rm), 0.2% yield strength (Rp), and % elongation (A25)
Figure 5—Grade 5 Ti6Al4V microstructures at (a) bottom and (b) top layers in the wall
Figure 6—Grade 23 Ti6Al4V microstructures at (a) bottom and (b) top layers in the wall
indicates the range of strengths. The tensile strength differences between the two directions were 8 MPa for Grade 5 and 31 MPa for Grade 23. Anisotropy was also seen in the elongation differences of the two planes of both grades of Ti6Al4V; that in the ZX plane being higher than in the XY plane.
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Microstructures Figures 5 and 6 show microstructures of samples extracted from the XY plane of the walls. Figures 5a and 5b show the lathlike matrix structure and lamellar structures, respectively, of Grade The Journal of the Southern African Institute of Mining and Metallurgy
Comparison of the mechanical properties of Grade 5 and Grade 23 Ti6Al4V 5 Ti6Al4V. When WAAM walls are built up from a substrate, the bottom layers are subjected to faster cooling rates, which result in martensitic plates formed into lathlike α structures. As additional layers are deposited on the wall, more heat accumulates, and the cooling rate is reduced (Bintao et al., 2018). Due to the increased heat accumulation, the upper layers have a fully lamellar α structure (Bintao et al., 2018). The same trend is witnessed in the Grade 23 microstructures shown in Figure 6; however, Grade 23 has a lower content of interstitial elements, which means that it has fewer α stabilizers. Oxygen and nitrogen are known to be strong α stabilizers (Kazantseva et al., 2017). This resulted in smaller α colonies, and the structures at the top layers of Grade 23 comprise smaller α colonies compared with those in the Grade 5 walls.
Positional differences Figure 7 shows the ultimate tensile strengths (Rm) of the walls at the bottom and top layers in the XY plane. The results show that samples taken from different positions in the walls exhibited different tensile strengths: the bottom layers had higher tensile strength than the top layers. This phenomenon is influenced by the heat accumulated by the wall and the cooling rate that it was subjected to, which in turn modified the grain sizes and microstructural evolution, as shown in Figures 5 and 6.
Hardness testing Hardness tests were carried out using the tensile samples taken in the XY plane. The results, presented in Figure 8, show that samples taken from the bottom of the walls had higher hardness than those taken from the top. This result supports the tensile data shown in Figure 7. This effect can be explained by the thermal cycle to which the walls were subjected. The hardness of Grade 5 was higher than that of Grade 23 due to its higher content of interstitial elements (Oh et al., 2011).
Correlation between tensile strength and hardness
Figure 9 shows that a linear relationship existed between the tensile strength and Vickers hardness of these WAAM Ti6Al4V alloys, with a correlation coefficient (R2) of 0.973. In this hardness range, the tensile strength of Ti6Al4V alloys produced using Gefertec’s 3DMP® WAAM process can be estimated by using Equation [1]:
Rm (MPa) = 4.51 VHN – 620
[1]
Conclusions The effect of interstitial impurities on the anisotropy of the physical properties of Grade 5 and Grade 23 Ti6Al4V alloy walls produced by WAAM was investigated. The results show that
Figure 7 – Positional difference of ultimate tensile strength (Rm) in the XY plane of Grade 5 and Grade 23 Ti6Al4V
Figure 8 – Comparison of Vickers hardness of Grade 5 and Grade 23 Ti6Al4V walls at different vertical positions The Journal of the Southern African Institute of Mining and Metallurgy
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Comparison of the mechanical properties of Grade 5 and Grade 23 Ti6Al4V
Figure 9 – Correlation between Vickers hardness and ultimate tensile strength of Ti6Al4V alloys
Grade 5, which has a high content of interstitial elements, had greater ultimate tensile strength (Rm), greater yield strength (Rp), and lower elongation (A25) than Grade 23. Tensile properties of both grades were slightly direction-dependent: the horizontal XY plane had higher tensile strength and lower elongation compared with the vertical ZX plane. Grade 23 displayed more marked directional differences in these properties than Grade 5 Ti6Al4V, owing to its lower content of interstitial elements. Vertical positional differences were also observed: the bottom layers of the wall had lathlike α structures due to the faster cooling rate at that position. As more layers were deposited and more heat accumulated in the wall, the cooling rate was reduced, which gives rise to lamellar α structures in the top layers. The heat accumulation and differences in cooling rates resulted in higher tensile strength and lower elongation in the bottom layers. A good correlation between the ultimate tensile strength and Vickers hardness was observed.
the Southern African Institute of Mining and Metallurgy, vol. 112, no. 7. pp. 563–575. Bermingham, M., McDonald, S., and Dargusch, M. 2018. Effect of trace lanthanum hexaboride and boron additions on microstructure, tensile properties and anisotropy of Ti-6Al-4V produced by additive manufacturing. Materials Science and Engineering, vol. 719. pp. 1–11. Bintao, W., Zengxi, P., Donghong, D., Dominic, C., and Huijun, L. 2018. Effects of heat accumulation on microstructure and mechanical properties of Ti6Al4V alloy deposited by wire arc additive manufacturing. Additive Manufacturing, vol. 23. pp. 151–160. Biswal, R., Zhang, X., Syed, A. K., Awd, M., Ding, J., Walther, F., and Williams, S. 2019. Criticality of porosity defects on the fatigue performance of wire + arc additive manufactured titanium alloy. International Journal of Fatigue, vol. 122. pp. 208–217. Carroll, B., Palmer, T., and Beese, A. 2015. Anisotropic tensile behavior of Ti-6Al-4V components fabricated with directed energy deposition additive manufacturing. Acta Materialia, vol. 87. pp. 309–320.
Acknowledgements Gefertec GmbH is thanked for supplying the Ti6Al4V walls used in this project. Mr Pierre Rossouw (CSIR) assisted with the heat treatment. Sincere gratitude is directed to Mr Sibusiso Mahlalela and Mr Kabelo Matea for assisting with the experimental work. Professor Kathy Sole is thanked for proofreading the article.
Gefertec. 2020. Gefertec GmbH. https://www.gefertec.de/en/3dmp-process/ [accessed 1 March 2020]. Kazantseva, N., Krakhmalev, P., Yadroitsev, I., Fefelov, A., Merkushev, A., Ilyinikh, M., and Kurennyk, T. 2017. Oxygen and nitrogen concentrations in the Ti-6Al-4V alloy manufactured by direct metal laser sintering (DMLS) process. Materials Letters, vol. 209. pp. 311–314. Li, J.L., Alkahari, M.R., Rosli, N.A., Hassan, R., Sudin, M.N., and Ramli, F.R. 2019.
Authors’ contributions Christof Gassmann: preparation and printing of Ti6Al4V walls, results interpretation, and supervision. Heinrich Möller: arranged samples, experimental tests, and supervision. Lesego Mashigo: sample preparation, optical analysis, tensile and hardness testing, and the first draft of this article.
References
Review of wire arc additive manufacturing. International Journal of Automation Technology, vol. 13, no. 3. pp. 346–353. Oh, J.M., Lee, B.G., Cho, S.W., Lee, S.W., Choi, G.S., and Lim, J.W. 2011. Oxygen effects on the mechanical properties and lattice strain of Ti and Ti-6Al-4V. Metals and Materials International, vol. 17, no. 5. pp. 733–736. Thuketana, S., Taute, C., Möller, H., and du Plessis, A. 2020. Characterization of surface roughness and subsurface pores and their effect on corrosion. Journal
AMFG. 2018. Autonomous manufacturing. https://amfg.ai/2018/05/17/anintroduction-to-wire-arc-additive-manufacturing/?cn-reloaded=1 [accessed 17 March 2020]. Antonysamy, A.A. 2012. Microstructure, texture and mechanical property evolution during additive manufacturing of Ti6Al4V alloy for aerospace applications. PhD thesis, Manchester University, UK. Azom. 2002. Azo Materials. https://www.azom.com/article.aspx?ArticleID=1547 [accessed 12 April 2020]. Bauristhene, A.M., Mutombo, K., and Stumpf, W.E. 2013. Alpha case formation mechanism in Ti-6Al-4V alloy investment castings using YFSZ shell moulds. Journal of the Southern African Institute of Mining and Metallurgy, vol. 113. pp. 357–361.
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of the Southern African Institute of Mining and Metallurgy, vol. 120. pp. 369–376. Wang, F., Williams, S., Colegrove, P., and Antonysamy, A.A. 2012. Microstructure and mechanical properties of wire and arc additive manufactured Ti-6Al-4V. Metallurgical and Materials Transactions A, vol. 44. pp. 968–977. Wu, B., Pana, Z., Lib, S., Cuiuria, D., Dingc, D., and Li, H. 2018. The anisotropic corrosion behaviour of wire arc additive manufactured Ti-6Al-4V alloy in 3.5% NaCl solution. Corrosion Science, vol. 137. pp. 176–183. Zhang, X., Martina, F., Ding, J., Wang, X., and Williams, S.W. 2016. Fracture toughness and fatigue crack growth rate properties in wire +arc additive manufactured Ti-6Al-4V. Fatigue and Fracture of Engineering Materials and Structures, vol. 40, no 5. pp. 790–803.
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The Journal of the Southern African Institute of Mining and Metallurgy
Beneficiation of recycled process water at DRDGOLD’s ERGO plant, and its effect on gold recovery A. Narain*1, J. H. Potgieter1,2, G. E. Rencken1, and J. Smith1 Paper written on project work carried out in partial fulfilment of M.Sc (Metallurgical Engineering) degree *
Affiliation: 1 School of Chemical and Metallurgical Engineering, University of the Witwatersrand, South Africa. 2 Department of Natural Science, Manchester Metropolitan University, UK. Correspondence to: A. Narain
Email:
arshir.narain@yahoo.com
Dates:
Received: 30 Nov. 2020 Revised: 13 May 2021 Accepted: 28 Jul. 2021 Published: July 2021
How to cite:
Narain, A. Potgieter, J.H., Rencken, G.E., and Smith, J. 2021 Beneficiation of recycled process water at DRDGOLD’s ERGO plant, and its effect on gold recovery. Journal of the Southern African Insitute of Mining and Metallurgy, vol. 121, no. 7, pp. 331–344 DOI ID: http://dx.doi.org/10.17159/24119717/1442/2020 ORCID: A. Narain https://orcid.org/0000-00023757-0571 J.H. Potgieter https://orcid.org/0000-00032833-7986 J. Smith https://orcid.org/0000-00022395-3109
Synopsis To conserve fresh water resources and comply with environmental regulations, DRDGOLD, a South African gold producer re-treating surface tailings, has transitioned to a fully closed water circulation system. Consequently, the accumulation of contaminants, as well as addition of reagents, has led to changes in water composition that have compromised leach performance and overall gold recovery. A two-sample t-test confirmed a significant difference in gold recoveries between the use of Rand Water, which was used as a benchmark, and untreated process water. Atomic absorption analysis of ERGO’s process water, confirmed the presence of iron, nickel, zinc, and manganese. A study of the effect of the identified contaminants on gold recovery showed that iron, nickel, and zinc have the largest negative effect on gold recovery, with iron and nickel being more detrimental than zinc. Sulphates were shown to have a possible passivation effect, which also influenced gold recoveries, although to a lesser extent than the heavy metals. Calcium, when present in excess, had a positive influence on gold recovery indicating the possible formation of a calcium aurocyanide complex. Lime softening successfully reduced the heavy metal and sulphate concentrations, and the gold recoveries obtained with the treated process water were similar to those as achieved with Rand Water. Keywords gold tailings, re-processing, water quality, gold recovery.
Introduction DRDGOLD is a South African gold producer and a world leader in the recovery of gold from the retreatment of surface tailings. DRDGOLD ERGO Mining (Pty) Ltd (ERGO) specialises in the re-treatment of gold mine tailings, with numerous operations and tailings reclamation sites spanning 163 km from the western to the central and eastern regions of the Witwatersrand. The current study focuses on the ERGO plant operations in Brakpan, which predominantly reclaims tailings dams in the eastern regions of the Witwatersrand. Approximately 64 000 t of tailings material are treated daily, requiring an estimated 60 ML/d of water. Since water is a strategic resource in water-scarce South Africa , several mining operations are at risk due to a limited supply of water. This has resulted in an increased utilization of recycled water, with the focus on recycling process water. ERGO’s metallurgical research department has, however, demonstrated improved gold recoveries when utilizing clean water compared with recycled process water in laboratory test work. This has led to a growing interest in research to determine the effect/s of recycled process water on gold recovery. While there is extensive literature available on the effect and interactions of metal-cyanide complexes with the aurocyanide complex, little is known about the effects of water quality on cyanide gold leaching (Rees, 2000). Cyanide has been used globally as a lixiviant for gold leaching since the process was first patented over 100 years ago. However, the processing and extraction of gold has become more complex as the simple, free-milling oxide ores have become depleted over the past decade (Rees, 2000). Due to the presence of various metal species in the ore, as well as in the process water, cyanide complexes of antimony (Sb), arsenic (As), cobalt (Co), copper (Cu), iron (Fe), nickel (Ni), thallium (Tl), and zinc (Zn) may form in the leach solution that is contacted with activated carbon (Sheya and Palmer, 1989). These complexes could result in lower gold recoveries. Rees (2000) showed that copper is the most problematic metal in gold leaching due to the rapid formation of copper cyanide complexes, which can consume a great deal of the available cyanide.
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Beneficiation of recycled process water at DRDGOLD’s ERGO plant, and its effect Boehme and Potter (1983) proved that silver has an adverse effect on the loading rate of gold, with ratios of silver to gold of 1:1 and 2:1. They also demonstrated that as the copper concentration in a slurry increased, the rate of gold loading onto activated carbon decreased. Hedley and Tabachnick (1968) showed that the zinc cyanide complex formed was deleterious to gold dissolution and subsequent gold recovery. Fleming and Nicol (1984) demonstrated that activated carbon adsorbs organic solvents, which affect the rate of gold loading. Activated carbon also adsorbs small quantities of cyanide, metal hydroxides, calcium (Ca), and iron sulphide, which contaminate the carbon, resulting in a decrease loading capacity. Fleming and Nicol (1984) concluded that activated carbon may have a stronger affinity for both Cu and Ni than for gold. Fink and Putnam (1950) found that the addition of trace amounts of sodium sulphide to the cyanide solution dramatically hindered gold leaching, and postulated that the sulphide ions passivate the surface of the gold by forming a layer of Au2S. Jeffrey and Breuer (2000) obtained similar results using an ore containing 5% silver. Most naturally occurring gold sulphide ores contain trace amounts of silver, which has been shown to increase the cyanide leaching reaction rate (Jeffrey and Breuer, 2000). As previous work focused mostly on the effects of heavy metals on gold loading on activated carbon, this investigation aims to extend current knowledge by considering the impact of typical water contaminants like Ca, magnesium (Mg), and sulphate on gold recovery. McDougall, et al. (1980) postulated that the mechanism of adsorption of aurocyanide onto activated carbon involves adsorption of aurocyanide as the less soluble Mn+[(Au(CN)2]n complex (M = K+, Ca+, Mg+) as the initial adsorption stage. Therefore, in this work we examined the quality of the recycled process water streams on a typical sulphide ore gold processing plant. Laboratory-scale gold flotation and cyanide leach test work were carried out to establish the effects of contaminants such as metal cyanide complexes, sulphates, Ca, and Mg, on gold recovery. An investigation of treatment options for process water at various temperatures, with the respective effect on gold recovery, was also undertaken.
Material and methods
Reagents In addition to using Rand Water and ERGO’s recycled process water for the experimental work, a number of reagents were required for leaching, flotation, and water treatment. These included calcium oxide (lime), sodium carbonate (soda ash), sodium cyanide, commercial flotation reagents (sodium normalpropyl xanthate (SNPX), 1,1,3 triethyloxybutane (Senfroth), alkyl dithiophosphate (Senkol)), nickel sulphate, iron sulphate, manganese sulphate, magnesium sulphate, and zinc sulphate. All chemicals used were analytical grade, and used as received, without further purification.
Reagent dosage for chemical precipitation The Minnesota Rural Water Association (2009) states that chemical precipitation is an effective and common method utilized for water purification. The addtion of chemicals such as lime (calcium hydroxide, Ca(OH)2) and soda ash (sodium carbonate, Na2CO3) increases the water’s pH and results in precipitation of the ions that cause hardness. If the total hardness of water is less than or equal to the total alkalinity, then hardness is attributed to carbonate hardness only, with no non-carbonate hardness. If the total hardness is greater than the total alkalinity, the carbonate hardness is equal to the total alkalinity concentration, and the non-carbonate hardness is calculated from the difference between total hardness and total alkalinity concentration. (Minnesota Rural Water Association, 2009). If total hardness is equal to or less than total alkalinity, then the lime dosage is calculated as shown in Equation [1] (Minnesota Rural Water Association, 2009): [1] where A = The carbon dioxide concentration in source water B = Bicarbonate alkalinity in source water C = Hydroxide alkalinity in source water D = Magnesium concentration in source water % excess = Amount of lime fed in excess to ensure pH > 10.5. When treating water that contains non-carbonate hardness, soda ash is required. The amount of soda ash is estimated using Equation [2]:
Sample preparation A low-grade, refractory sulphide gold tailings material containing approximately 0.20 g/t Au was used as the sample. The material, which had previously undergone milling, was sourced from the 4L50 reclamation site. The elemental composition was determined by X-ray fluorescence (XFR) analysis of a composite sample of the 4L50 feed. The contaminants present in the highest concentrations are summarized in Table I.
Table I
XRF analysis of ERGO’s 4L50 feed ore Fe mg/L
Ni mg/L
Cu mg/L
Mn mg/L
Ag mg/L
K mg/L
Zn mg/L
Suphur mg/L
28 300
108
54.4
240
0.11
6370
70.5
3328
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[2]
Analytical techniques Atomic absorption calibration Atomic absorption spectroscopy (AAS) was used for the quantitative determination of elements in solution. Standards of the various metallic ions were made up from pure chemicals and calibration curves constructed by preparing dilutions from the different stock solutions.
Fire assay The gold content of the solid samples was determined by fire assay at the metallurgical laboratory of Mechanical Analysis and Engineering Design (MAED) in Sandton. The Journal of the Southern African Institute of Mining and Metallurgy
Beneficiation of recycled process water at DRDGOLD’s ERGO plant, and its effect X-ray diffraction X-ray diffraction (XRD) analysis was used to characterize the crystalline materials formed during the chemical precipitation, using a Philips PW 1800 X-ray diffractometer. The XRD analysis was performed as a qualitative analysis to determine the mineral phases that had formed. The parameters used for the analysis are described in Table II.
Preparation of contaminant standards One of the aims of this study was to establish which ions, or combination of ions, present in the process water were responsible for the reduced gold recoveries compared to when Rand Water was used. To achieve this, Rand Water was spiked with individual chemical species to determine the effect on gold recovery. In this way it was possible to identify the individual ion or combination of ions that had the greatest effect on gold recovery. The contaminants present in the highest concentrations were identified from an initial ICP-MS analysis undertaken on a composite sample of the ERGO process water (refer to Table IX for the ICP-MS analysis). Standard solutions of the identified contaminants were prepared by dissolution of a known amount of the analyte in cyanide solution. Stock solutions of 1 g/L of Ni, Fe, Zn, Ca, and Mg were individually prepared. Fe, Zn, Ni and Mg were added as sulphate salts and Ni(CN)42-, Zn(CN)2, Fe(CN)64-, and Mg(CN)2 complexes were formed in solution by the addition of the respective molar requirement of sodium cyanide. The reagents used to prepare standard solutions and their respective stoichiometric dissociation reactions with sodium cyanide are summarized in Table III. Sulphate addition was controlled and varied by the addition of stock solutions of sulphuric acid. The stoichiometric dissociation reaction to produce the anionic sulphate is: provided in Table IV.
Experimental procedures Leach tests A dry composite sample was collected from the 4L50 reclamation
Table II
Parameters used for the XRD analysis Parameter
Value
Type of radiation Operating voltage Operating current 2θ range
Cobalt-K alpha (CoKa) 30 kV 10 mA 10–90°
site. The solid sample was slurried with each of the respective water sources to a density of 1.45 kg/L. A 1-litre head sample was collected, filtered, dried, and assayed for gold content. Three 1-litre samples were placed in separate 5-litre bottles. The lime, cyanide, and carbon dosages shown in Table V were added and a bottle roll leach was undertaken for 7 hours, corresponding to the leach circuit residence time at ERGO. After leaching, the carbon was screened and washed. The slurry was then filtered and the filtrate analysed for gold content using AAS. The filter cake was re-pulped and filtered twice before being dried in an oven at 110ºC. The dried residue and carbon were analysed for gold by fire assay. Figure 1 illustrates a block flow diagram of the leach test work.
Lime and soda ash softening A composite sample of process water was analysed to determine the initial calcium, sulphate, and heavy metal concentrations. The process water samples were treated with unslaked lime, as well as with a combination of unslaked lime and soda ash, at the required dosages. A one-hour waiting period was allowed for complete precipitation of contaminants. The resulting solution was filtered through a vacuum filter and the precipitate collected, dried at 90°C and weighed. A 250 mL volume of filtrate was used for the analysis of Ca, sulphate, and heavy metal concentration, and the remainder of the filtrate was used for the flotation and leach test work to observe the effect on gold recovery. The above procedure was repeated after heating the process water to temperatures of 20°C, 60°C, and 90°C. A diagrammatic representation of the lime-soda ash softening procedure is presented in Figure 2.
Flotation tests A composite 4L50 solid sample was slurried with each of the water sources to a density of 1.32 kg/L. A 5-litre portion of the slurry was transferred into a Denver laboratory-scale flotation cell. The initial pH of the slurry was measured and pH values outside of the optimal range of between pH 6–8 were adjusted using lime for pH values below 6, or sulphuric acid for pH values above 8. The flotation reagents and the required dosages are shown in Table VI. The concentrate was collected by scraping the froth at 10-second intervals after conditioning for 5 minutes. The concentrate was transferred into a labelled sample dish and dried in an oven at 110ºC. The dried concentrate was analysed for gold content. The tails were transferred into a bucket and allowed to settle before being prepared for the bottle roll leach test by Table IV
Sulphuric acid stoichiometric dissociation reaction Contaminant
Table III
Sulphur/ sulphate
R eagents used to prepare standard solutions and their stoichiometric dissociation reactions with sodium cyanide Contaminant Reagent Fe Zn Ni Ca Mg
Iron (II) sulphate Zinc sulphate Nickel sulphate Calcium oxide Magnesium sulphate
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Stoichiometric dissociation reaction
Sulphuric acid
H2SO4 +2H2O → 2H3O+ + SO42-
Table V
Lime, cyanide, and carbon dosages
Stoichiometric dissociation reaction Fe2+ + 6CN— → [Fe(CN)6]4Zn2+ +2CN— → Zn(CN)2 Ni2+ + 4CN— → [Ni(CN)4]2— Ca2+ + 2CN— → Ca(CN)2 Mg2+ + 2CN— → Mg(CN)2
Reagent
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Plant-scale dosage (g/t)
Lab-sale equivalent dosage (mg/L)
0.36 1.00 10
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Beneficiation of recycled process water at DRDGOLD’s ERGO plant, and its effect
Figure 1—Block flow diagram illustrating the leach test work
Figure 2—Diagrammatic representation of the lime-soda ash softening procedure
decanting the water and thickening the sample to a density of 1.45 kg/L. Figure 3 illustrates a block flow diagram summarizing the flotation and tails leach test work.
Results and discussion Statistical analysis Ten leach tests were performed to determine if there were
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significant differences in gold recovery between using Rand Water and recycled, process water. The margins of experimental error and repeatability of the experimental results were calculated, and the averaged statistical analysis results are presented in Table VII. A two-sample t-test was undertaken using the experimentally determined washed residue values to determine if the differences in gold recoveries between using Rand Water and process water were significant. A two-sample t-test is an inferential test that determines if there is a significant difference between the means of two data-sets and whether the two data-sets are from the same or different populations (Central Virginia Governor’s School for Science and Technology, 2003). A summary of the t-test results is presented in Table VIII. The null hypothesis for the t-test states that the means of the two washed residues are the same (Central Virginia Governor’s School for Science and Technology, 2003). The calculated t-statistic value of 6.26 was greater than the t-critical value of 2.26, and the p-critical value obtained was less than 0.05; thus the null hypothesis was rejected, indicating that a significant difference exists between the gold recoveries when using Rand Water and process water.
Effect of individual contaminants on gold recovery The contaminants present in the recycled process water were identified from an initial ICP-MS analysis on a composite sample of the ERGO process water. The contaminants present in the highest concentrations are summarized in Table IX, as well as the corresponding typical Rand Water concentrations (Mohotsi, 2020). Rand Water was spiked with increasing amounts of each individual contaminant species to determine the effect on gold The Journal of the Southern African Institute of Mining and Metallurgy
Beneficiation of recycled process water at DRDGOLD’s ERGO plant, and its effect
Figure 3—Block flow diagram illustrating the flotation and tails leach test work
Table VI
Table VIII
Flotation reagent dosages
Reagent
Two-sample t-test on leach washed residue
Plant-scale dosage (g/t)
Lab-scale equivalent dosage (mL)
20 10 12
4.8 2.4 0.3
Sodium normal-propyl xanthate 1,1,3 triethyloxybutane Alkyl dithiophosphate
53.3 ± 3.3 47.9 ± 4.1
101.3 ± 3.0 101.9 ± 5.9
Process water Au (g/t)
0.094 <0.01 10.00
Pearson correlation Hypothesised mean difference df t stat. P(T<=t) one-tail t Critical one-tail P(T<=t) two-tail t Critical two-tail
Washed Gold Gold residue (g/t) dissolution % accountability %
Rand Water 0.202 ± 0.012 0.094 ± 0.009 Process water 0.202 ± 0.012 0.105 ± 0.005
Au (g/t)
Mean Variance Observations
Averaged statistical results for direct leach test work Head (g/t)
Rand Water
Table VII Sample ID
t-test: Paired two-sample for means
0.105 <0.01 10.00 –0.9 <0.01 9.00 6.26 0.00 1.83 <0.01 2.26
Table IX
ICP-MS analysis of ERGO recycled process water, with corresponding data for typical Rand Water Board Sample ID Process water Head sample Rand Water
Fe (mg/L)
Ni (mg/L)
Na (mg/L)
Mg (mg/L)
Ca (mg/L)
K (mg/L)
Zn (mg/L)
Sulphate (mg/L)
CNWAD mg/L
CNTotal (mg/L)
135
13.3
762
193
1135
29.7
24.5
3328
7.2
12.2
<0.05
5.0
10.3
7.3
18.5
2.8
<0.05
13.6
-
1.0
recovery. The main aim of the test work was to identify an individual ion or group of ions that had the greatest effect of gold recovery. This would facilitate the consideration of more comprehensive beneficiation options rather than following a ‘blanket’ overall treatment approach. The Journal of the Southern African Institute of Mining and Metallurgy
Effect of sulphur/sulphates on gold recovery The tailings processed at ERGO are sulphidic in nature. Jeffrey and Breuer (2000) showed that many sulphide minerals are to some extent soluble in cyanide solution; and thus some VOLUME 121
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Beneficiation of recycled process water at DRDGOLD’s ERGO plant, and its effect sulphur species were expected to be present in the leach solution. Elemental sulphur is one of the by-products of the breakdown of pyrite, and agglomerations of elemental sulphur were observed in filtered solutions prior to ICP analysis. This sulphur is thermodynamically stable as a sulphate ion under the cyanidation conditions (Rees, 2000). This study was aimed at establishing the effects of sulphates on the leaching efficiency of gold from a sulphide ore. The leach test results using Rand Water spiked with stock solutions of sulphuric acid to vary the sulphate concentration is presented in Figure 4. An approximate 10% reduction in gold recovery was observed after the addition of 50 mg/L of sulphate, and gold recoveries decreased further as the sulphate concentrations were increased. To prevent volatilization of toxic HCN gas, lime (CaO) was added to increase the leach pH to greater than 10.5. At this high pH, the calcium ions could have reacted with the sulphate ions to produce gypsum (CaSO4. 2H2O), which could cause fouling of the sites on the activated carbon and thus reduce gold recovery. The presence of hydrogen sulphide ions (HS-) has been found to be detrimental to the leaching of gold (Jeffrey and Breuer, 2000). This finding was pertinent in the present study because the gold would react directly with the sulphide as the pyrite was destroyed. At high pH values, there would be limited free hydrogen ions available to react with the sulphide. Furthermore, all metals are known to react readily with sulphide
to form insoluble metal sulphide complexes, which may passivate the surface of the gold and thus reduce gold dissolution.
Effect of calcium on gold recovery The Elsner equation [Equation 3] accurately describes the cyanidation of gold under controlled conditions. However, increased calcium concentrations resulting from lime addition may have a significant effect on subsequent gold dissolution and the solubility of the resulting aurocyanide complex. The effects of calcium on the leaching of gold from a sulphide ore were investigated by adding increasing concentrations of calcium (as CaO) to Rand Water and conducting laboratory-scale leach tests on ERGO’s 4L50 sulphide ore. The leach results are presented in Figure 5. [3] A 20% decrease in gold recovery occurred with the addition of 50 mg/L of calcium to Rand Water. This may have resulted from poor gold adsorption caused by passivation of the carbon surfaces by calcium foulants. As the initial calcium concentrations were increased above 50 mg/L, gold recoveries showed improvement with an increase in recovery of approximately 8.4% observed with a calcium addition of 500 mg/L compared to 50 mg/L. Davidson and Solet (2007) hypothesised the formation
Figure 4—Leach test results using sulphate-spiked Rand Water at pH 10.5
Figure 5—Effect of calcium-spiked Rand Water on gold recovery from tailings, and the mass of washed residue produced at pH 10.5
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Beneficiation of recycled process water at DRDGOLD’s ERGO plant, and its effect of a calcium aurocyanide complex that rapidly and strongly adsorbs onto activated carbon. These results would support the formation of a calcium aurocyanide complex as the gold recovery increased with an increased calcium concentration. Davidson and Solet (2007) further stated that the preferential adsorption of calcium aurocyanide onto activated carbon is due to the limited solubility of the calcium aurocyanide complex compared with the higher solubility of potassium and sodium aurocyanide (Davidson and Solet, 2007).
Effect of magnesium on gold recovery The effect of magnesium addition to Rand Water on gold recovery is presented in Figure 6. The addition of 50 mg/L of magnesium resulted in a 15.8% decrease in gold recovery compared to that without magnesium. With increases in magnesium concentration above 50 mg/L, gold recoveries improved and an increase in recovery of approximately 2.5% was observed at 500 mg/L of magnesium. Although excess concentrations of Mg aided gold recovery, recoveries remain more than 10% lower than when using clean Rand Water.
Effect of heavy metals on gold recovery Gold tailings are considered as refractory in nature as most of the free-milling gold has been leached previously, and generally contain sulphide minerals, organic carbon, and heavy metals such as iron and nickel (Rees, 2000). This leads to an increased cyanide consumption attributed to the formation of metal cyanide species that compete with the aurocyanide complex for sites on activated carbon. The stability constants of various metal cyanide species (Rees, 2000) are given in Table X. Aurocyanide is one of the most stable metal cyano-complexes. Iron (II) and iron (III) also form strong complexes, with the iron (III) complex marginally stronger, and the iron (II) complex marginally weaker than the gold cyanide complex (Rees, 2000). Nickel forms a strong cyanide complex in the +2 oxidation state, with Zn forming a weaker cyanide complex (Rees, 2000). Due to their similar stabilities, these complexes are all present to some extent in leach solutions. Their predominance is determined by the free cyanide concentration and is strongly pH-dependent (Rees, 2000). The stability constants are listed in terms of log K, where K is the equilibrium constant. The larger the equilibrium constant,
the greater the stability. Ferrocyanide is shown as an example in Equation [4]. [4]
Table X
tability constants for metal cyanide species (Rees, S 2000) Species
Stability (log K)
Au(CN)2 Ag(CN)2– Cu(CN)2– Cu(CN)32– Cu(CN)43–
39.3 20.5 16.3 21.7 23.1
–
Species
Stability (log K)
Fe(CN)6 35.4 Fe(CN)63– 43.6 Ni(CN)42– 30.2 Zn(CN)42– 19.6 Cu(CN)42– 9.1 4–
Effect of iron on gold recovery The effect of iron additions to Rand Water on leach recoveries are presented in Figure 7. An increase in iron concentration, together with the addition of a stoichiometric amount of cyanide, resulted in a decrease in gold recovery. Since the cyanide addition was stoichiometric based on the oxidative breakdown of pyritic ore, the iron is expected to be present in the ferrous form and in the presence of cyanide would predominantly form the Fe(CN)64complex (Rees, 2000). This is a very stable complex, and opinion in the literature is divided as to whether it would dissociate. As shown in Table X, the stability constants of aurocyanide and the Fe(CN)64- complex are similar (Kyle, 1997). The results show that as the initial iron concentration was increased from 50 mg/L to 500 mg/L, a 28% reduction of gold recovery occurred. These results are supported by Kyle (1997) who states that excess ferrous ions will readily displace gold from the aurocyanide complex.
Effect of nickel on gold recovery The leach test results using various nickel concentrations in Rand Water are presented in Figure 8. An increase in nickel concentration, together with the addition of the stoichiometric amount of cyanide, resulted in reduced gold recovery. The
Figure 6—Effect of magnesium-spiked Rand Water on gold recovery from tailings, and the mass of washed residue produced at pH 10.5 The Journal of the Southern African Institute of Mining and Metallurgy
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Beneficiation of recycled process water at DRDGOLD’s ERGO plant, and its effect
Figure 7—Effect of iron-spiked Rand Water leach tests on gold recovery from tailings, and the mass of washed residue produced at pH 10.5
Figure 8—The effect of nickel-spiked Rand Water on gold recovery from tailings, and the mass of washed residue produced at pH 10.5
stability constants in Table X show that nickel forms a medium strength cyanide complex. This complex is nine orders of magnitude less stable than the aurocyanide complex (Kyle, 1997). Additionally, the Ni(CN)42- complex contains four CN– ions compared to two in Au(CN)2- resulting in higher cyanide consumption. This allows a better dispersion in the solution without forming clusters, which enables Ni to be adsorbed onto activated carbon when present in high concentration (Sayiner and Acarkan, 2013). Similar behaviour was evident in the current study, resulting in a 23% decrease in gold recovery as the nickel concentration was increased from 50 mg/L to 500 mg/L. Addition of 100 mg/L of Ni resulted in an approximately 50% reduction in gold recovery compared to unspiked Rand Water.
Effect of zinc on gold recovery Zinc can be a major constituent of the process water due to the cyanide leaching of zinc minerals and the zinc cementation applied to precipitate gold and silver (Merrill-Crowe process), which is the process implemented at ERGO Brakpan. The leach results using various concentrations of zinc in Rand Water are presented in Figure 9. An increase in zinc concentration, together with the addition of the stoichiometric amount of cyanide,
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resulted in a depressed gold recovery. Kyle (1997) explains that zinc predominantly forms a tetracyano complex at a pH of 10. Due to the lower stability of the Zn(CN)42- complex compared to the Ni(CN)42- and Fe(CN)64- complexes, it is evident that the zinc cyanide complex adversely affects gold recovery but to a lesser degree than the nickel and iron cyanide complexes. It is postulated that the bulk of the zinc introduced would form zinc hydroxide, which rapidly precipitated out of solution due to the high leach pH (Kyle, 1997).
Lime and soda ash softening – chemical precipitation The leach test results showed that heavy metals such as Fe, Ni, and Zn have the largest negative effect on gold recovery. Sulphates also showed possible passivation effects, influencing gold recovery. Since lime addition is the general practice for increasing the leach pH at most gold leach plants, and given its easy availability, lime softening was considered a viable and economical beneficiation option. Furthermore, the solubilities of Ca, Mg, and silica (Si) are significantly reduced by increased temperature and they are therefore, more effectively removed by warm/hot softening than by cold/ambient softening (Suez, n.d.). The Journal of the Southern African Institute of Mining and Metallurgy
Beneficiation of recycled process water at DRDGOLD’s ERGO plant, and its effect
Figure 9 – Effect of zinc-spiked Rand Water on gold recovery from tailings, and the mass of washed residue produced at pH 10.5
The lime softening and soda ash softening test work parameters, final pH values, and resultant precipitate masses are summarized in Tables XI and XII. A one-hour waiting period was allowed for complete precipitation of contaminants. Figure 11 illustrates the settling of precipitants formed during the lime and soda ash treatment. The filtered precipitate from the lime plus soda ash softening test at ambient conditions is illustrated in Figure 12: Figure 13 shows a comparison between ERGO’s untreated recycled process water and samples that underwent lime and soda ash treatment/softening. The beaker on the far left contains untreated process water. The middle sample was treated with unslaked lime only, and that on the far right was treated with both unslaked lime and soda ash. Figure 10—Two 5 litre glass beakers of process water treated with unslaked lime (CaO)
Figure 10 shows two process water samples that were treated with unslaked lime and soda ash. The test was repeated at temperatures of 20°C, 60°C, and 90°C.
X-ray diffraction analysis of precipitates The precipitates formed from each water treatment procedure were subjected to X-ray diffraction (XRD) analysis. The main constituents were identified as quartz, magnesium silicate hydrate, calcium silicate, cordierite (identified as anthophyllite), bassanite, and gypsum. The formation of gypsum (CaSO4·2H2O)
Table XI
Lime softening test parameters Sample ID
Initial pH
Lime addition (mg/L)
Volume treated (L)
Temperature (°C)
pH after addition
Mass of precipitate (g)
Ca removal - Test 1 Ca2+ removal - Test 3 Ca2+ removal - Test 5
5.46 ± 0.05 4.30 ± 0.05 3.61 ± 0.05
756.0 ± 2.0 1170.0 ± 2.0 1170.0 ± 2.0
9.50 ± 0.01 10.0 ± 0.01 10.0 ± 0.01
19.1± 0.5 60.0 ± 0.5 90.0 ± 0.5
10.90 ± 0.05 10.92 ± 0.05 11.20 ± 0.05
7.7 ± 0.2 15.5 ± 0.2 28.9 ± 0.2
2+
Table XII
Lime and soda ash softening test parameters Sample ID Ca2+ + Mg2+ removal - Test 2 Ca2+ + Mg2+ removal - Test 4 Ca2+ + Mg2+ removal - Test 6
Initial pH
Lime addition (mg/L)
Soda ash addition (mg/L)
Volume treated (L)
Temperature (°C)
pH after addition
Mass of precipitate (g)
5.54 ± 0.05 4.67 ± 0.05 3.72 ± 0.05
756.0 ± 2.0 1170.0 ± 2.0 1170.0 ± 2.0
180.0 ± 2.0 180.0 ± 2.0 180.0 ± 2.0
9.50 ± 0.01 10.0 ± 0.01 10.0 ± 0.01
19.1 ± 0.5 60.0 ± 0.5 90.0 ± 0.5
11.02 ± 0.05 11.00 ± 0.05 11.30 ± 0.05
10.9 ± 0.2 17.9 ± 0.2 32.2 ± 0.2
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Beneficiation of recycled process water at DRDGOLD’s ERGO plant, and its effect XRD scans of the precipitates formed when the process water was treated with a combination of lime followed by soda ash indicated that lime treatment alone was more efficient. At ambient conditions, it was evident that only calcium-containing compounds (gypsum and calcite) , and quartz had precipitated. With the addition of soda ash, it was expected that magnesiumcontaining compounds would also be present in the precipitate. The addition of soda ash produced precipitates with fewer identifiable calcium- and magnesium-containing compounds than the tests where only lime treatment was used. This behaviour was observed at all temperatures investigated. Figure 11—Settling of precipitants formed during the lime and soda ash treatment. The sample on the left was treated with unslaked lime, and that on the right was treated with a combination of unslaked lime and soda ash
Figure 12—Filtered precipitate for lime plus soda ash softening at ambient conditions
Atomic absorption analysis
Tables XIII and XIV summarize the atomic absorption results from the various treatment techniques undertaken on the ERGO process water, at the respective treatment temperatures. The data shows that lime softening alone was sufficient to remove more than 99% of the heavy metals (Zn, Fe, and Ni) from ERGO process water. The apparent increase in potassium concentrations is attributed to impurities in the reagents and experimental error, since potassium concentration should not increase with lime and soda ash dosing. The increase in sodium concentration is attributed to the soda ash addition. Treatment of process water with a combination of lime and soda ash at ambient conditions was effective for the complete removal of Mg. However, a similar reduction in Mg concentration (> 99%) can be achieved using only lime treatment at elevated temperatures of 60°C and 90°C. A reduction of 87% Mg is, from a practical perspective, equally sufficient. It was also evident that the reduction in nickel concentration was negatively affected by soda ash addition as well as an increase in treatment temperature. Slatter et al. (2009) demonstrated that while sodium ions increase silver loading, the presence of sodium has a negligible effect on gold loading. From the XRD and atomic absorption results, it was concluded that lime softening/treatment on its own was sufficient to significantly reduce heavy metal concentrations. Furthermore, lime addition reduced the sulphate content due to the precipitation of gypsum.
Leach results for gold recovery Figure 13—Comparison of process water after softening treatment
and bassanite (CaSO4·0.5H2O) would consume free sulphates in solution, thus reducing the concentration of sulphates in the process water.
The lime and soda ash softened process water was utilized to perform laboratory-scale direct leach tests with the aim of determining the viability of the treatment options for gold recovery. Table XV summarizes the averaged data from triplicate leach tests undertaken using the differently treated waters at varying temperatures. Figure 14 gives a graphical representation of the results obtained.
Table XIII
Contaminant concentrations after respective treatments - heavy metals Sample ID Treatment Temperature (°C) Process water head sample Lime softening 20 ± 0.5 Lime - soda ash softening 20 ± 0.5 Lime softening 60 ± 0.5 Lime - soda ash softening 60 ± 0.5 Lime softening 90 ± 0.5 Lime - Soda ash softening 90 ± 0.5
mg/L
Zinc % Removal
24.5 0.02 99.9 0.02 99.9 0.04 99.8 0.02 99.9 <0.01 100.0 <0.01 99.9
mg/L
Iron % Removal
135 0.17 99.8 0.10 99.9 0.14 99.9 0.08 99.9 0.11 99.9 0.05 99.9
mg/L
Nickel % Removal
13.3 0.12 99.1 0.15 98.9 0.63 95.3 1.25 90.6 0.89 93.3 1.26 90.5
Sodium mg/L % Removal 762 776 920 784 1102 783 1036
-1.8 -20.7 -2.8 -44.6 -2.75 -35.9
(Note: Negative values indicate an increase in contaminant concentration)
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Beneficiation of recycled process water at DRDGOLD’s ERGO plant, and its effect Table XIV
Contaminant concentrations after respective treatments with inorganic ions
Sample ID
Treatment Temperature (°C)
Process water head sample Lime softening 20 ± 0.5 Lime - soda ash softening 20 ± 0.5 Lime softening 60 ± 0.5 Lime - soda ash softening 60 ± 0.5 Lime softening 90 ± 0.5 Lime - soda ash softening 90 ± 0.5
Magnesium mg/L % Removal 193 26.5 86.3 1.3 99.8 0.18 99.9 0.12 99.9 0.13 99.9 0.10 99.9
mg/L
Calcium % Removal
1135 498 56.2 565 50.2 871 23.3 896 21.1 599 47.2 697 38.5
Potassium mg/L % Removal 29.7 21.6 -6.6 31.4 -5.7 36.0 -21.5 30.5 -2.9 30.7 -3.6 31.2 -5.2
Sulphates mg/L % Removal 3328 2636 2795 2807 3094 1799 1989
20.8 16.0 15.7 7.0 45.9 40.2
(Note: Negative values indicate an increase in contaminant concentration)
Table XV
Direct leach results Water sample ID
Direct leach results Head grade (g/t) Washed residue (g/t)
Rand Water Process water Lime-softened process water – ambient temperature (20°C) Lime plus soda ash-softened process water – ambient temperature Lime-softened process water – 60°C Lime plus soda ash-softened process water – 60°C Lime-softened process water – 90°C Lime plus soda ash-softened process water – 90°C
0.200 0.200 0.200 0.200 0.200 0.200 0.200 0.200
0.094 0.105 0.094 0.104 0.085 0.111 0.107 0.110
Gold dissolution (%)
Gold accountability (%)
53.3 49.7 52.0 47.5 57.3 44.7 46.7 44.9
101.3 101.9 102.1 99.2 103.8 104.2 99.3 104.8
Figure 14—Graphical representation of direct leach results, indicating the gold dissolution achieved with each water type and the subsequent washed residue from the leach
The data in Table XV shows that there are differences in gold dissolution and thus gold recovery, when using Rand Water and ERGO process water for the direct leaching of the 4L50 reclaimed material. Rand Water produced a greater dissolution of gold from solids, which in turn resulted in a lower washed residue and a greater recovery of gold from the feed material. Treatment of ERGO process water with lime (lime softening only) produced a much lower washed residue than treatment The Journal of the Southern African Institute of Mining and Metallurgy
with a combination of lime and soda ash. Tests where only lime treatment was applied to ERGO process water produced a higher gold dissolution than the corresponding lime-soda ash treatment at all temperatures tested. The addition of soda ash as part of the proposed treatment of ERGO process water proved to be detrimental to gold recovery. Leach tests using process water that had been treated with soda ash yielded the lowest percentage gold dissolution with the VOLUME 121
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Beneficiation of recycled process water at DRDGOLD’s ERGO plant, and its effect highest washed residue concentrations. Furthermore, the results of tests with soda ash treated water were poorer than those using untreated process water. This could be a result of the carbonates from the soda ash addition reacting with calcium to produce CaCO3, a known foulant of activated carbon that occupies active sites, thus resulting in a reduction in gold recovery (Davidson and Solet, 2007). Process water treated with lime at 60°C provided the most significant improvement in gold recovery. However, heating 60 ML of process water per day from ambient (20°C) to 60°C will require approximately 2930 MWh of energy per day, and is thus not an economically viable option. At a unit cost of R1.00 per kilowatt-hour, the additional daily energy cost would be R2.93 million. Rand Water, at a unit cost of R30.92 per litre, is also not an economically viable option.
Flotation and tails leach results for gold recovery The lime and soda ash softened process water was used to perform laboratory-scale leach and flotation tests to assess the viability of the treatment options for gold recovery. Table XVI summarizes the averaged data from triplicate flotation and leach test work using the different water sources. Figures 15 and 16 give a graphical representation of these results. The data from Figures 15 and 16 shows that addition of soda ash to treat ERGO process water may prove detrimental to gold
recovery. The 4L50 material, with a gold head grade of 0.20 g/t, was slurried with the different water sources and subjected to flotation. The concentrate was collected, dried, and analysed for gold by fire assay, and the flotation tails underwent leach tests. It is important to note that the deviation in the tails head grades presented in Table XIV is attributed to efficiency of the flotation. With effective flotation, a lower tails head grade is expected due to the higher gold grades in the flotation concentrate. Therefore, a higher gold grade of the tails prior to leaching is attributed to the corresponding lower concentrate gold grade from ineffective flotation. The tails leach tests which utilized process water treated with soda ash resulted in the lowest gold dissolutions, as well as the highest washed residue values. The results of these tests were also inferior to those with ERGO process water. .
Conclusions ➤ Experimental work and statistical analysis indicated a significant difference in gold recoveries between using Rand Water and untreated process water. ➤ Analysis of a composite sample of ERGO process water confirmed the presence of appreciable concentrations of heavy metals. ➤ Leach tests conducted using Rand Water spiked with known concentrations of the identified contaminants showed that iron, nickel, and zinc had the largest adverse effect on gold
Figure 15—Flotation and tails leach results
Table XVI
Summarized flotation and tails leach results with different treated and untreated waters
Water sample ID
Rand Water Process water Lime softened process water – ambient Lime plus soda ash softened process water – ambient Lime softened process water – 60°C Lime plus soda ash softened process water – 60°C Lime softened process water – 90°C Lime plus soda ash softened process water – 90°C
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Flotation Flotation Tails leach Tails washed Flotation Tails leach Total gold Mass percentage head grade concentrate head grade residue recovery recovery recovery as concentrate (g/t) (g/t) (g/t) (g/t) (%) (%) (%) Mass % 0.20 0.20 0.20 0.20 0.20 0.20 0.20 0.20
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3.5 1.8 2.8 2.7 2.9 2.9 2.2 1.2
0.155 0.187 0.150 0.157 0.150 0.150 0.170 0.180
0.084 0.111 0.096 0.102 0.077 0.113 0.090 0.100
22.5 6.5 25.0 21.5 25.0 25.0 15.0 10.0
45.9 40.5 36.3 34.5 48.6 24.4 47.1 44.4
68.4 47.0 61.3 56.0 73.6 49.4 62.1 54.4
1.3 1.4 1.4 1.4 1.4 1.5 1.3 1.4
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Beneficiation of recycled process water at DRDGOLD’s ERGO plant, and its effect
Figure 16—Comparison of flotation concentrate and tails washed residue grades
recovery, with iron and nickel having a greater effect than zinc. ➤ The presence of sulphates indicated possible passivation effects, although these were less deleterious than the effects of heavy metal contamination. ➤ The increased gold recoveries observed as calcium concentrations increased could possibly be ascribed to the formation of a calcium aurocyanide complex. ➤ Due to the negligible differences in gold recoveries when using the lime treated process water and Rand Water, the use of lime was chosen as an attractive treatment option. ➤ Lime softening proved effective for the reduction of heavy metal concentrations by metal hydroxide precipitation, as well as reducing sulphate concentration by the precipitation of gypsum. ➤ While the leach results indicated that lime-treated process water at elevated temperatures produced higher gold recoveries; the heating requirement for 60 ML of process water per day makes this option economically unviable.
Recommendations It is not clear whether these negative effects of heavy metals on gold recovery were due to cyanide consumption by metalcyano complexes, fouling of the activated carbon, or precipitates passivating the gold surfaces before adsorption. This should be investigated in future work to establish the mechanism of this reduced gold recovery.
Acknowledgements DRDGold (Pty) Ltd is gratefully acknowledged for financial support of this investigation, and a bursary for the corresponding author. The authors owe a huge debt to Mr. Bruce Ebell from DRDGold for his insight and contribution to this work.
References Boehme, W. and Potter, G. 1983. Carbon adsorption of gold. Maximum loading and ionic contaminant effect on loading rates. Gold and Silver Heap and Dump Leaching Practice. Hiskey, J. ( ed.). American Institute of Mining, Metallurgy and Petroleum Engineers, Salt Lake City, NV. pp. 129–137. Central Virginia Governor’s School for Science and Technology. 2003. T Test. http://
Davidson, R. and Solet, M. 2007. The major role played by calcium in gold plant circuits. Journal of the Southern African Institute of Mining and Metallurgy, vol. 107, no. 7. pp. 463–468. Fink, C. and Putnam, G. 1950. The action of sulphie ion and of metal salts on the dissolution of gold in cyanide solutions. Mining Engineering, vol. 187. pp. 952–955. Fleming, C. and Nicol, M. 1984. The adsorption of gold onto activated carbon III. Factors influencing the rate of loading and equilibrium capacity. Journal of the South African Institute of Mining and Metallurgy, vol. 84, no. 4. pp. 85–93. Hedley, N. and Tabachnick, H. 1968. Chemistry of cyanidation. Mineral Dressing Notes no. 23. American Cyanamid Company, New Jersey. 54. pp. Jeffrey, M. and Breuer, P. 2000. The cyanide leaching of gold in solutions containing sulfide. Minerals Engineering, vol. 13, no. 10. pp. 1097–1106. Kyle, J.H. 1997. Stability of metal-cyanide and hydroxide complexes. Proceedings of the World Gold ‘97 Conference, Singapore. Australasian Institute of Mining and Metalurgy, Melbourne. pp. 163–169. McDougall, G., Hancock, R., Nicol, M., Wellington, O., and Copperthwaite, R. 1980. The mechanism of the adsorption of gold cyanide on activated carbon. Journal of the South African Institute of Mining and Metallurgy, vol. 80, no. 9. pp. 344–356. Minnesota Rural Water Association. 2009. Lime softening. Minnesota Water Works Operations Manual. 4th edn. Elbow Lake, MN. Chapter 16. Mohotsi, M. 2020. Rand Water - Reports for Johannesburg Metro. Water Quality Information Management, Johannesburg. Rees, K.L. 2000. Leaching and adsorption behaviour of gold ores, PhD thesis, Dearatment of Chemical Engineering, University of Melbourne. Sayiner, B. and Acarkan, N. 2013. Effect of silver, nickel and copper cyanides on gold adsorption on activated carbon in cyanide leach solutions. Physicochemical Problems of Mineral Processing, vol. 1, no. 50. pp. 277–287. Sheya, A. and Palmer, G. 1989. Effect of metal impurities on adsorption of gold by activated carbon. Cyanide Solutions. US Bureau of Mines. Slatter, K.A., Plint, N.D., Cole, M., Dilsook, V., and de Vaux, D., Palm. N., and Oostendorp, B. 2009. Water management in Anglo American process operations. Proceedings of the International Mine Water Conference, Pretoria. International Mine Water Association. pp. 46–55. Suez. Not dated. Suez water technologies & solutions. https://www.
www.cvgs.k12.va.us/DIGSTATS/main/inferant/d_tdist.htm [accessed 25 May
suezwatertechnologies.com/handbook/chapter-07-precipitation-softening
2020].
[accessed 1 October 2019].
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Effects processing parameters and building orientation on the microstructural and mechanical properties of AlSi10Mg parts printed by selective laser melting Affiliation: 1 School of Chemical and Metallurgical Engineering, Faculty of Engineering and the Built Environment, University of the Witwatersrand, Johannesburg, South Africa.
C. Phetolo*1, V. Matjeke1, and J. van der Merwe1 Paper written on project work carried out in partial fulfilment of BTech (Metallurgical Engineering) degree *
Correspondence to: J. van der Merwe
Email:
josias.vandermerwe@wits.ac.za
Dates:
Received: 12 Mar. 2021 Revised: 20 Jul. 2021 Accepted: 20 Jul. 2021 Published: July 2021
How to cite:
Phetolo, C., Matjeke, V., and van der Merwe, J. 2021 Effects processing parameters and building orientation on the microstructural and mechanical properties of AlSi10Mg parts printed by selective laser melting. Journal of the Southern African Insitute of Mining and Metallurgy, vol. 121, no. 7, pp. 345–350
Synopsis The mechanical properties and microstructure of AlSi10Mg alloy samples that were printed by selective laser melting (SLM) were studied to determine the effect of processing parameters and building orientation. After printing, the alloy was stress relieved at 250°C for 2 hours. The microstructures were analysed by optical microscopy and scanning electron microscopy (SEM) to determine the alloy phases and distribution. Phase transformation characteristics of the material were evaluated using differential scanning calorimetry (DSC). Mechanical properties were determined by subjecting the XY- and Z-built samples to tensile and nano-indentation testing. The samples from the tensile tests were then used to perform fractographic analysis by SEM. The microstructural properties in each orientation revealed a non-homogeneous microstructure which was characterized by a semi-elliptical tract and fine silicon precipitates, which were found to be softer along the fusion zone. The DSC thermograms revealed that the material underwent two phase transformations during the first heating cycle. The mechanical properties revealed a higher UTS, higher yield strength, and a lower percentage elongation in the Z orientation than in the XY orientation. Fractographic analysis showed that crack initiation in both orientations started from the surface in a brittle manner due to surface flows, and then propagated via microvoid coalescence. Keywords AlSi10Mg alloy, additive manufacturing, mechanical propeerties, microstructure.
DOI ID: http://dx.doi.org/10.17159/24119717/1575/2021 ORCID: J. van der Merwe https://orcid.org/0000-00034563-8078
Introduction AlSi10Mg is an aluminum alloy that is often used in the fabrication of objects with complex shapes and with thin walls. Its widespread use is due to its exceptional hardness, strength, and dynamic properties (EOS, 2014). These properties make AlSi10Mg ideal for applications in the aerospace, automobile, and electronics industries, where a superior strength to weight ratio and thermal properties are required. Due to these ideal properties, considerable research is being done on how to use the alloy in additive manufacturing (AM), which is a fabrication process used to print 3D objects. The process is carried out by printing successive layers of material until a complete 3D object, pre-defined in the computer aided device (CAD) file, is built (Ngo, 2018). The purpose of this study was to determine the effect of building orientation and processing parameters on the mechanical properties of AlSi10Mg samples derived by AM. The build orientation has been investigated by several authors; however, the combination of building direction and processing parameters has not been fully explored (Matjeke et al., 2020; Mfusi et al., 2018). AM is gaining considerable commercial acceptance because it enables the fabrication of complex shapes with good precision (Srivatsan, Manigandan, and Sudarshan, 2015). It is preferred over conventional fabrication processes like casting and deformation processes because it does not require new tools or moulds during the production of prototypes and customized parts, thus leading to lower production costs and lead times (Srivatsan, Manigandan, and Sudarshan, 2015).
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Materials and experimental procedure A 280 HL selective laser melting (SLM) platform with a bed size of 280 × 280 × 350 mm was used to produce the alloy test samples from AlSi10Mg alloy powder. Tensile specimens were produced ready for testing and no machining was performed prior to testing. Metal part geometry was fabricated directly from the computer aided design (CAD), using the the manufacturing parameters outlined in Table I. The widely used volume-based energy density E (J/mm3) is defined in Equation [1]: [1] where P is laser power (W), v is scan speed (mm/s), h is hatch spacing (mm), and l is the layer thickness. The samples were heat-treated at 250°C for 2 hours in a furnace in order to stress-relieve the as-built components after the SLM process.
Characterization A Microtrac laser particle size analyser and a Zeiss Merlin field emmission scanning electron microscope were used to determine the powder particle size distribution and the morphology respectively. The environment in the build chamber was kept inert by purging with argon gas at a flow rate of 0.3 l/min. Differential scanning calorimetry (DSC) was carried out to Table I
SLM process parameters
Laser power (W) Scanning speed (mm/s) Hatch spacing (mm) Scanning strategy Layer thickness (mm)
370 1000 180 Stripes 30
Results Particle size distribution and morphology The particle size distribution is summarized in Figure 1. The d50 of the distribution was found to be 43.89 µm, with the majority of the particles concentrated about the mean. Visual inspection confirmed the particle size distribution, the sample being dominated by large particles with a few smaller particles. Figure 2 shows that the combination of large particles and small particles resulted in a high degree of compaction. The majority of the particles were close to spherical. Some of the particles with irregular shapes were formed when small particles attached to larger spherical particles, and others were originally irregular shapes. The chemical composition of the as-built AlSi10Mg alloy is presented in Table III.
Metallography The microstructure of the samples built in the XY (horizontal) direction shows a semi-elliptical, or slightly parabolic, morphology as shown in Figure 4A. The dimensions of the
Mfusi et al., 2018
120
30
370 1300 100 Stripes 30
150 1000 50 Stripes 50
100
25
80
20
60
15
40
10
20
5
DSC processing parameters Parameter
Value
Heating rate (°C/min) Sample mass (mg) Argon gas flow rate (ml/min)
100 13.77 100
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All samples were prepared and tested in accordance with ASTM E8/E8M. Tensile tests were performed with an MTS Landmark servo-hydraulic test system using samples with a gauge length of 150 mm and diameter of 12.5 mm. The samples were were pulled at a rate of 0.8 mm/s at room temperature until fracture. An extensometer was used during the test, and the slope of the elastic region on the flow stress curve was measured and used to calculate the Young’s modulus. Hardness tests were carried out on a Brinell hardness tester at a load of 62.5 kg force using 1 mm diameter for 10 seconds. Nano-indentation tests were carried out on additional etched samples to examine the different phases in the microstructure, using a three-sided pyramidal Berkovich diamond indenter.
Tang and Pistorius (2017)
Table II
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Cumulative Passing (%)
Parameter This work
determine the exact critical transformation temperature and the enthalpy required to melt the powder. DSC testing and processing parameters are shown in Table II. Samples for metallographic examination were prepared by transversely sectioning the tensile test samples. The sectioned samples were mounted and polished using suspensive lubrication and 3 µm and 1 µm polishing cloths, with a 0.5 µm colloidal finish. The samples were then etched using a solution consisting of 1% HF, 1% HCl, and 2.5% HNO3 and examined by optical and electron microscopy.
0
0
50
100 150 Particle size (mm)
Volume Percent (%)
One of the main drawbacks of the casting process when compared to AM is the inability to produce repeatable results due to process-related defects such as porosity and inclusions (du Plessis et al., 2019; Girelli et al., 2019). This can be combated by optimizing process parameters like the building direction, which is essentially the orientation of material accumulation during fabrication (Qin et al., 2019). There is a vast pool of building directions from which to pick, including the Y, Z, XY, X, and 45° orientations. The building orientation direction chosen affects support structure production, building cost, tool path planning, and building time, as well as the microstructure and mechanical properties of the printed component (Li et al., 2016; Qin et al., 2019). As a result, it is critical to choose a building direction that will result in desirable mechanical properties, a homogeneous microstructure, and negligible flaws at a fair cost.
0 200
Figure 1—Particle size distribution of the powder used for SLM The Journal of the Southern African Institute of Mining and Metallurgy
Effects processing parameters and building orientation
Figure 3—AlSi10Mg powder at higher magnification, indicating particle morphology
Figure 2—AlSi10Mg powder at low magnification (100×)
Table III
C hemical composition of the AlSi10Mg alloy (wt%), as determined by EDS analysis Al
Si
Mg
O
89.3
9.8
0.6
0.3
semi-elliptical tracks were measured using the ImageJ analyser software. The measured semi-elliptical tracks presented in Figure 5 had an average width of 60 µm. The microstructure of the sample built in the Z (build) direction (Figure 6) also demonstrated a semi-elliptical morphology and spacing with an average width of 110 µm. However, it must be noted that because of overlap the track width cannot be determined from the interior of builds and the estimates of 60 and 110 µm are clearly too small, since the hatch spacing was 180 µm. The coarse silicon phase was also observed at higher magnifications, like in the XY (horizontal) direction shown in Figure 4D. The semi-elliptical
Figure 5—XY-direction microstructure showing grain size analyses
tract in the Z (or build) direction appears flatter than that in the XY orientation. The microstructure in both the XY and Z
Figure 4—Microstructures A and C represent the low and high magnifications of direction XY respectively, while, microstructures B and D represent the low and high magnifications of direction Z respectively The Journal of the Southern African Institute of Mining and Metallurgy
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Figure 6—Z-direction microstructure showing grain size analyses
direction comprises coarse and fine areas. The size of the of the precipitates decreases towards the centre of the semi-elliptical track. At higher magnifications, the focus is inside the core of the elliptical track of the microstructure. Figure 7 indicates a cellular pattern, which is a result of silicon particles forming an interconnected fine network surrounding the Si-rich aluminum matrix. The region highlighted within the black lines shows an area with a slightly coarser microstructure located at the melt pool boundary. At even higher magnification the microstructure shows second-phase precipitates as light grey particles within the darker grey matrix, as indicated in Figure 8.
Figure 7—Microstructure of AlSi10Mg printed by SLM
Calorimetry The calorimetric results are presented in Figure 9. The thermogram is characterized by two exothermic peaks at temperatures of 567.4°C and 687.3°C. These peaks represent the critical transformation temperature and melting point of AlSi10Mg. The first exothermic peak is followed by an endothermic peak at a maximum temperature of 604.6°C. The DSC measurements provided the exact properties that are associated with AlSi10Mg. This information can be useful during optimization of the processing parameters and during selection of a heat treatment process. However, the peaks occur at a considerably higher temperature than previously reported (Fiocchi et al., 2016). The thermogram also revealed an endothermic peak which starts at around the melting temperature of AlSi10Mg, which is 570°C, and therefore represents the melting of the sample.
Figure 8—Microstructure of AlSI10Mg indicating the presence of second-phase particles
Mechanical test results Tensile properties The tensile test results are summarized in Table V. The deformation characteristic and flow curves are presented in Figure 10. More than one samples was tested, but only the representative curves are shown, and although there was no slip in the jaws some of the samples showed a jagged edge during the initial stage of the test. The Z build orientation exhibited higher strength than the XY orientation. The modulus of elasticity for the XY orientation was 41 GPa, with the yield point in the region of 200 MPa at a strain of 0.006. The modulus of elasticity for the Z direction was 40 GPa, with the elastic limit being reached at 209 MPa and a strain of 0.008. The deformation behaviour in both orientations along the elastic region is comparable, and major difference start
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Figure 9—DSC thermogram of AlSi10M
to appear after the commencement of plastic deformation. The ultimate tensile strength (UTS) recorded for the XY orientation was 294 MPa, with the fracture point being reached after an elongation of 8.17%. The Z orientation recorded a UTS of 324 MPa, with fracture taking place after 5.9% elongation. The results of the tensile tests are summarized in Table V. The tests revealed that the XY orientation showed better ductility than the Z orientation. The higher ductility in the XY orientation came as The Journal of the Southern African Institute of Mining and Metallurgy
Effects processing parameters and building orientation Table VI
Vickers hardness results Orientation
Average hardness (HB)
XY 90+1 Z 94+1
Table VII
Microhardness and Young's modulus results
Figure 10—Engineering stress vs strain curves for Z and XY orientations
Fusion zone Semi-elliptical core Matrix Si particle Nano-ndentation) Young’s modulus (GPa)
a trade-off for strength, as the Z orientation showed higher yield strength and UTS. Nahmany et al. (2019) and Tang and Pistorius (2017) also reported a lower percentage elongation and a higher UTS for the Z orientation; however, the higher yield strength reported in this work for the Z orientation is contrary to previous findings.
104 80
144 85
113 68
608 166
rejected from the supersaturated aluminum matrix during heat treatment to form a Si-rich phase throughout the sample, but mostly concentrated along the cellular boundaries, and is which coarser on the melt pool boundaries rather than on the semielliptical core. The non-homogeneous microstructure was further investigated by performing nano-indentation tests to determine the nano-mechanical properties of different phases in the microstructure. The results from the nano-indentations seemed to contradict the findings of the microstructural analysis. The nano-indentation hardness results indicated that the fusion zone is harder than the semi-elliptical core, while the microstructural analysis showed that the semi-elliptical core has a finer microstructure and would thus be expected to be harder than the fusion zone. The Si particles had the highest hardness, and this would allow them to act as second-phase particles and strengthen the alloy via solid solution strengthening.
Hardness The average Brinell hardnesses and nano-indentation results are summarized in Tables VI and VII. The general measurement revealed that the harness is less in the XY orientation than the Z orientaton. Nano-indentation hardness was investigated to determine the properties of each phase, and it was found that the Al matrix was softer than the Si precipitates, as shown in Table VII. Si particles recorded the highest hardness at 608 HV, while the core of the semi-elliptical track recorded the second highest hardness of 144 HV. The fusion zone, which comprised the heataffected zone (HAZ) region (the semi-elliptical core) recorded 104 HV, while the matrix recorded the lowest hardness at 113 HV. The Young’s modulus results mirror the hardness results along the different phases, with the Si particles recording 166 GPa followed by the fusion zone, semi-elliptical core, and matrix at 85 GPa, 80 GPa, and 68 GPa respectively.
Conclusion It can be concluded that processing parameters and building direction should be considered together in order to optimize the mechanical properties of SLM-printed AlSi10Mg parts. Although the stress-relief process conducted in this investigation differed from the heat treatment used by Tang and Pistorius (2017) and Mfusi et al. (2018), an inference can be drawn from the pattern. The manufacturing process resulted in a non-homogenous microstructure in which different regions, such as the matrix, semi-elliptical core, and fusion zone were found, as well as Sirich second-phase particles in both the Z and XY orientations. The rejection of Si from the supersaturated matrix was confirmed by the exothermic peaks in the thermograms. The Z orientation showed better mechanical properties, with a higher UTS and yield strength but lower ductility compared to the XY orientation. Therefore, it can be concluded that for applications that require a material with high strength, the components should be printed in the Z orientation, and for cases where high ductility is required and high strength is not a primary requisite, the components can be built in the XY orientation.
Discussion Based on the results of this study and other work, it seems that the combination of processing parameters and building direction affects the mechanical properties (Mfusi et al., 2018). Although Mfusi et al. did not conduct tensile tests on stressrelieved samples, an inference and pattern can be drawn from the as-built- tensile test results. Aboulkhair et al. (2016) reported that AlSi10Mg components printed by SLM owe their strength to three main factors: grain boundary strengthening, solid solution strengthening, and interaction of dislocations. The grain boundary strengthening in both the XY and Z orientation is due to the fine silicon microstructure, which is a result of rapid cooling during the printing process. The solid solution strengthening is provided by the silicon particles, which are Table V
Tensile test results YS (MPa) XY 200 Z 209
294 324
This work Tang and Pretorius (2017) Mfusi et al., (2018) UTS (MPa) %E YS (MPa) UTS (MPa) %E YS (MPa) UTS (MPa) %E 8.2 5.9
182 180
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Effects processing parameters and building orientation References Aboulkhair. N.T., Maskery. I., Tuck. C., Ashcroft. I., and Everitt. N.M. 2016. Improving the fatigue behaviour of a selectively laser melted aluminium alloy: Influence of heat treatment and surface quality. Materials & Design, vol. 104. pp. 174–182. Du Plessis, A. Glaser, D., Moller, H., and Mathe, N. 2019. Pore closure effect of laser shock peening of additively manufactured AlSi10Mg. 3D Printing and Additive Manufacturing, vol. 6, no. 5. pp. 245–252. doi: 10.1089/3dp.2019.0064 EOS. 2014. Material data sheet: Aluminium alloy AlSi10Mg. EOS GmbH. https:// www.eos.info/03_system-related-assets/material-related-contents/metalmaterials-and-examples/metal-material-datasheet/aluminium/material_ datasheet_eos_aluminium-alsi10mg_en_web.pdf [accessed 7 June 2021].
of additive-manufactured AlSi10Mg components. Progress in Additive Manufacturing, vol. 5. pp. 379–385. doi: 10.1007/s40964-020-00139-1 Mfusi, B.J., Tshabalala, L.C., Popoola, A.P.I., and Mathe, N.R. 2018. The effect of selective laser melting build orientation on the mechanical properties of AlSi10Mg parts. IOP Conference Series: Materials Science and Engineering, vol. 430. Conference of the South African Advanced Materials Initiative (CoSAAMI-2018), Vanderbijlpark, South Africa, 23–26 October 2018. doi: 10.1088/1757-899X/430/1/012028 Nahmany, M., Hadad, Y., Aghion, E., Stern, A., and Frage, N. 2019. Microstructural assessment and mechanical properties of electron beam welding of AlSi10Mg specimens fabricated by selective laser melting. Journal of Materials Processing Technology, vol. 270. pp. 228–240.
Fiocchi,. J., Tuissi,. A., Bassani,. P., and Biffi,. C.A. 2016. Low temperature annealing dedicated to AlSi10Mg selective laser melting products. Journal of Alloys and Compounds, vol. 695. doi: 10.1016/j.jallcom.2016.12.019
Ngo, T.D., Kashani, A., Imbalzano, G., Nguyen, K.T.Q., and Hui, D. 2018. Additive manufacturing (3D printing): A review of materials, methods, applications and challenges. Composites Part B: Engineering, vol. 143. pp. 172–196. https://doi. org/10.1016/j.compositesb.2018.02.012
Girelli, L., Tocci, M., Gelfi, M., and Pola, A. 2019. Study of heat treatment parameters for additively manufactured AlSi10Mg in comparison with corresponding cast alloy. Materials Science & Engineering A, vol. 739. pp. 317–328. doi: 10.1016/j.msea.2018.10.026
Qin, Y., Qi, Q., Scott, P.J., and Xiangqian, J. 2019. Determination of optimal build orientation for additive manufacturing using Muirhead mean and prioritised average operators. Journal of Intelligent Manufacturing, vol. 30. pp. 3015–3034. https://doi.org/10.1007/s10845-019-01497-6
Li, W., Li, S., Zhamg, A., Zhou, Y., Wei, Q., Yan, C., and Shi, Y. 2016. Effect of heat treatment on AlSi10Mg alloy fabricated by selective laser melting : Microstructure evolution, mechanical properties and fracture mechanism’, Materials Science & Engineering A, vol. 663. pp. 116–125. doi: 10.1016/j. msea.2016.03.088
Srivatsan, S.T., Manigandan, K., and Sudarshan, S.T. 20115. Additive manufacturing of materials: viable techniques. Additive Manufacturing. Innovations, Advantages, and Applications. Srivatsan, T.S. and Sudarshan, T.S. (eds). CRC Press, Boca Raton, FL. pp. 1–48. https://doi.org/10.1201/b19360
Matjeke, V., Moopanar, C., Bolokang, A.S., and van der Merwe, J.W. 2020. Effect of heat treatment on the microstructure and mechanical deformation behavior
Tang, M. and Pistorius, P.C. 2017. Oxides, porosity and fatigue performance of AlSi10Mg parts produced by selective laser melting. International Journal of Fatigue, vol. 94. pp. 192–201. u
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Sensitivity of shotcrete Brazilian discs coated with thin spray-on liners to environmental conditions M.J. Kanda1,2 and T.R. Stacey1 Affiliation: 1 University of the Witwatersrand, South Africa 2 Instituto Superior Politécnico de Tete, Mozambique. Correspondence to: M.J. Kanda
Email:
jacquesmpoyi@gmail.com
Dates:
Received: 26 Nov. 2020 Revised: 11 Apr. 2021 Accepted: 26 May 2021 Published: July 2021
How to cite:
Kanda, M.J. and Stacey, T.R. 2021 Sensitivity of shotcrete Brazilian discs coated with thin spray-on liners to environmental conditions. Journal of the Southern African Insitute of Mining and Metallurgy, vol. 121, no. 7, pp. 351–360 DOI ID: http://dx.doi.org/10.17159/24119717/1430/2021 ORCID: M.J. Kanda https://orcid.org/0000-00021220-855X
Synopsis Thin spray-on liners (TSLs) have been used as sealants and rock support in tunnels for over 25 years. Laboratory tests have indicated satisfactory properties that can provide adequate strength, adhesion, toughness, and elasticity as part of rock support systems in mining excavations. These characteristics are, however, not always demonstrated in practice, when mine environmental conditions do not correspond with laboratory environmental conditions. The shortcomings of TSLs experienced in the mining industry have raised concerns, resulting in scepticism from some operators regarding their utilization. The research described in this paper aims to investigate TSL performance in environmental conditions similar to those experienced in mines. Brazilian indirect tensile (BIT) specimens were prepared from precast shotcrete and coated with TSLs. The specimens were then exposed to different environmental conditions for up to 112 days before BIT testing under various conditions: laboratory room temperature and humidity; saturated room temperature; and combined saturated and high temperature (50°C). Additional uncoated shotcrete and TSL BIT specimens were prepared for comparison purposes. The results of the BIT tests showed that environmental conditions have a significant influence on the tensile strength enhancement of shotcrete by TSLs. Water-based TSLs are most likely to be suitable for high humidity environments, although their performance decreases at higher temperatures. Numerical modelling of TSL-coated BIT samples confirms the potential limitations of designing TSL support based only on laboratory testing carried out under room conditions. Keywords thin spray-on liner, Brazilian indirect tensile (BIT) test, TSL performance, environmental conditions, humidity, high temperature.
Introduction Accidents in underground mines are often rockfall related (Potvin, Stacey, and Hadjigeorgiou, 2004). To reduce accidents to a zero-harm target, various rock support techniques have been implemented and improved. Reinforcement supports such as frictional anchors, mechanically anchored rockbolts, and untensioned grouted dowels are common. Containment support such as shotcrete, wire mesh, and straps (Hoek, Kaiser, and Bawden, 2000; Stacey, 2001) has also been utilized to accommodate the rock mass displacement. These supports have shown advantages and limitations, depending on the rock mass qualities and mechanisms of rock mass failure. An emerging areal support type is the thin spray-on liner (TSL). TSLs are also referred to as thin sprayed membranes, multi-component polymeric (based) liners, thin coating materials, or just thin liners. Suppliers have indicated, supported by laboratory results, that TSLs possess notably higher tensile strength properties than materials commonly used for rock support, such as shotcrete, and are capable of spanning cracks in the rock mass, thus limiting the movements of mobilized rocks (Tannant, 2001). Laboratory indirect tensile strength testing (Masethe, 2015; Mpunzi et al., 2015) has confirmed this, indicating that a TSL coating can increase the tensile strength of the rock and shotcrete by some 30-40%. A significant increase in the pull-off strength of TSL-coated shotcrete was reported in Canada (Simser and Pritchard, 2012; Boeg-Jensen, and Swan, 2014). With regard to practical implementation in mines, TSLs also provide rapid curing (Yilmaz, 2010) and simplicity of utilization and transportation, owing to the compact equipment size and the small volumes of TSL materials required. All of these benefits could have been expected to lead to widespread adoption of TSLs in underground mines, but this has not transpired. The reason appears to be scepticism of mining operators regarding the actual performance of TSLs. Some mining companies are sceptical of new technology, particularly where safety is concerned, and have found that benefits claimed for TSLs were not observed in the field. Research has been carried out to understand the reasons behind
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Sensitivity of shotcrete Brazilian discs coated with thin spray-on liners to environmental conditions these discrepancies, and possibly to find solutions for effective utilization (Kanda and Stacey, 2019; Guner and Ozturk, 2018, 2016; Boeg-Jensen and Swan, 2014). Based on industry reaction regarding TSL performance (Kanda and Stacey, 2019), it is believed that laboratory tests on TSLs should be carried out under environmental conditions similar to those in which TSL areal support is expected to be used, therefore enabling selection of the most appropriate TSL. To date, this has not been done, since laboratory tests are generally performed under room temperature and humidity conditions. A programme of tensile tests on TSL materials under a range of environmental conditions demonstrated that mine-like conditions could severely degrade the performance of TSLs, and therefore lead to misjudgement of their real performance (Kanda and Stacey, 2019). However, in practice, a TSL does not act as a stand-alone material, but operates as a coating on the rock mass or on shotcrete. The research described in this paper aims to assess TSL behaviour in humid and high-temperature conditions when used as a coating. To achieve this, Brazilian indirect tensile (BIT) tests were carried out using shotcrete discs as a substrate, coated with the TSL materials, thus extending the research carried out by Mpunzi et al. (2015). Another important parameter to be investigated in the trials was the period of exposure to the environmental conditions after initial specimen preparation. This period should be longer than the commonly-used curing period of 28 days. The results for different environmental conditions will be compared with those carried out under the room conditions in which TSLs have been tested to date.
Effect of atmospheric conditions on coated BIT specimens Research published by Kanda and Stacey (2019) on this topic suggests that TSL uniaxial tensile strength (UTS) is sensitive to humidity and a combination of high humidity and temperature conditions. That research was motivated by the logical observation that the testing environment should match the field conditions in which the TSL is to be applied. One of the fundamental bases for this research relied on the suggestions from Osswald (2015) that polymeric material behaviour over time is temperature-related and could be explained through the Boltzmann superposition principle. This principle stipulates that in the linear viscoelastic regime, the strain (stress) responses to successive stress (strain) actions are cumulative. Kanda and Stacey (2020) described the results of a programme of UTS testing of TSLs in a range of environmental conditions. As in that programme, the current work will assess the effect of different environmental conditions on the tensile strength behaviour of TSL-coated shotcrete discs. In addition, the liners’ adherence behaviour on the substrates will be evaluated. TSLs are expected to contribute to resisting or smoothing crack expansion through their bonding characteristics and high tensile strength properties (Boeg-Jensen and Swan, 2014).
Laboratory testing and results TSL characteristics Three different TSLs were assessed through the Brazilian indirect tensile (BIT) test. For reasons of confidentiality they are referred to as TSL1, TSL2, and TSL3. Their characteristics are provided in Table I. TSL1 is a mixture of cement, polymer, and sand. The technical sheet suggests that the final product can prevent unravelling and oxidation, and can afford lateral constraint to the rock mass. This latter benefit should be an advantage regarding the mechanical properties being researched. The UTS indicated by the manufacturers is 9.7 MPa after 28 days of room curing. TSL2 is supplied in a bag as a single component powder, to which clean tap water is added to form the mixture. This cementitious product is claimed by the manufacturer to have a compressive strength of 44.7 MPa and a tensile strength of 3.1 MPa after 28 days. The TSL3 technical sheet presents this liner as a plasticizerfree aqueous copolymer emulsion, consisting of acrylic acid esters. It also contains wetting agents mixed with a cementitious binder. Among the claimed benefits are non-flammability and non-toxicity, and the ability to bond to dusty, wet, and fatty surfaces. The expected tensile strength after 28 days is 9.75 MPa.
Specimen preparation
The TSL materials were prepared using an electrically operated mixer under room conditions. The electric mixer is a Bosch drill driver with a rated power of 400 W and a vibration level of 13/6. Discs of shotcrete were prepared for use as substrates in conformance with International Society of Rock Mechanics and Rock Engineering (ISRM) recommendations (Ulusay and Hudson, 2007). They were cored from fibre-reinforced shotcrete (Wetcrete) panels 45 × 45 cm2 and 55 mm thick (see Figure 1a) provided by Concrete Lining Products (CLP), Carltonville, South Africa. These panels were cured in water containers for at least 30 days before supply, and had a characteristic strength of 40 MPa after 28 curing days. All cored discs were 42 mm in diameter and were trimmed to a length of 21 mm using a diamond saw, resulting in a length to diameter ratio of approximatively 0.5, as illustrated in Figure 1b. The coating of the shotcrete discs is illustrated in Figure 1c, the TSL thickness being 4 to 5 mm, as adopted by Mpunzi et al. (2015). The BIT specimens were prepared as follows: ➤ A thin film of release agent was applied on all the circumferential parts of the mould pieces to avoid the formation of air bubbles. ➤ The mould pieces were placed on a flat working surface, previously covered with a plastic film to avoid direct contact between the liner material and the working surface, to avoid damaging the liner during specimen removal. ➤ The TSL materials were mixed as recommended in the manufacturers’ technical specifications.
Table I
TSL characteristics Name TSL1 TSL2 TSL3
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Category
Components
Proportions of constituents
Cement-based (brittle) Cement-based (brittle) Cement-based (brittle)
Cement, sand, and polymer Mixed product of TSL2 and tap water Mixed product of TSL3 and polymer
3.85:1.41:1 - cement, sand, and polymer by mass 4.63:1 - ratio of mixed solid product and water by mass 5.2:1 - ratio of mixed solid product and polymer by mass
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Sensitivity of shotcrete Brazilian discs coated with thin spray-on liners to environmental conditions BIT testing was planned after exposure to these conditions for periods of 15, 28, 56, and 112 days. The focus was, however, on the periods longer that the usual TSL 28-day curing time.
Test set-up The BIT tests were performed on specimens coated with TSL1, TSL2, and TSL3. The aim was to assess the influence of environmental conditions on the indirect tensile strength of coated BIT shotcrete. TSL discs and uncoated shotcrete discs were also tested separately, in all cases following the ISRM suggested methods (Ulusay and Hudson 2007) The specimens were exposed to R, H, and HT conditions for periods up to 112 days, before testing in a MTS testing machine, shown in Figures 2a and 2b. An example of the output BIT test graphs is shown in Figure 2c. The tensile strength was calculated as follows: [1] where F is the compressive force applied along the disc diameter (N) t is the disc thickness (mm) d is the disc diameter (mm). In determining the disc thickness, the TSL thickness was not taken into account, since its effect has been shown to be negligible (Mpunzi et al., 2015). The TSL helps to strengthen the substrate by its resistance to tensile failure. Figure 1—(a) Panel of shotcrete, (b) cored disc of shotcrete, and (c) coating of shotcrete discs (Kanda, in press)
➤ The mixture was poured into the moulds, and spread using a spatula to ensure that it filled the moulds. ➤ A small quantity of the mixed TSL was applied to the cleaned shotcrete disc surface to ensure subsequent TSL adherence. ➤ The shotcrete disc was placed on the TSL in the mould and gently levelled. ➤ One day of setting time was allowed before removing the coated shotcrete. ➤ The procedure was repeated for the other surface of the shotcrete disc. ➤ Using a diamond saw, the TSL coating diameter was reduced to about a millimetre smaller than the shotcrete disc diameter, to ensure full contact between the steel platens and the disc during BIT testing. A total of 300 BIT specimenswere prepared: 50 shotcrete discs were coated for each of the three TSLs; and 50 other discs were prepared from each of the TSLs. In addition, 40 uncoated shotcrete discs were also prepared. From the 50 coated shotcrete discs and 50 TSL discs obtained from each TSL, three groups of 16 to 17 specimens were allocated to three environmental conditions. These were: room conditions (R), the humid condition, defined as storage in a container full of tap water at room temperature (H); and the humid and high-temperature condition, involving storage of specimens in a container of water which was placed in an oven at 50°C (HT). The uncoated shotcrete discs were split into three groups of 13 to 14 specimens each and allocated to each of the three environmental conditions. The Journal of the Southern African Institute of Mining and Metallurgy
Laboratory results After testing, only the specimens that failed along the loaded diameter were considered. Those that had circumferential failures were discarded. A sample of the tested BIT specimens is presented in Figure 3. Uncoated shotcrete discs, cored from the supplied panels of reinforced shotcrete, exhibited a near-constant BIT strength of about 5.2 MPa under room conditions. These results, and those for discs exposed to the different environmental conditions, are shown in Figure 4. As can be seen from Figure 4, H and HT conditions adversely influence shotcrete BIT strength, contrary to the expectation that the H condition would strengthen the shotcrete, since it was cured in saturated conditions. A possible explanation is that the discs were prepared using a diamond coring machine that employs water for lubrication and cooling during cutting, producing a significant quantity of fines. This could be one of the reasons for the subsequent strength loss. The BIT strength trends for the discs coated with TSL1, TSL2, and TSL3, and exposed to various environmental conditions are shown in Figure 5. The graphs in Figure 5 show that TSL1 and TSL3 are the best-performing coatings under room conditions, while water-based TSL2 performs better under saturated conditions, retaining the BIT strength just below the room strength. In HT conditions, none of the three TSLs was able to improve the overall BIT strengths over time. TSL3 showed the greatest strengthening capability in HT conditions. This TSL has better combined bonding and tensile characteristics than TSL1 and TSL2 in HT conditions. However, the evaluation of strengthening capability must take into consideration the substrate’s loss of strength under the same environmental conditions. To do this, representative strength curves of specimens were chosen based VOLUME 121
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Sensitivity of shotcrete Brazilian discs coated with thin spray-on liners to environmental conditions
Figure 2—(a) MTS model 643.15A-03 for BIT tests, (b) coated BIT specimen under test, and (c) example of BIT stress-displacement curves
Figure 3—Samples of tested discs of (a) TSLs, (b) uncoated shotcrete, and (c) coated shotcrete
on their average strengths. Table II displays a sample of these averages and the corresponding coefficients of variation (CoV) after exposure for 15 and 112 days. The TSL strengthening results obtained are summarized in Figure 6. Figures 6 shows the strengthening trends of TSLs under room conditions. All three TSLs enhanced the strength of the discs, thus demonstrating their potential to improve the resistance of shotcrete to crack expansion. In saturated conditions, the TSLs imparted a somewhat smaller strength improvement, and under HT conditions, the effect of all three TSLs was reduced. These results complement those from Figure 5, which displays the overall strength trends for coated discs without considering the
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influence of environmental conditions on substrate behaviour. Therefore, for shotcreted walls that are susceptible to tensile cracks, TSL1 would be recommended when the environment is close to room conditions. In saturated conditions, TSL2 should be recommended, while TSL3 would be the most appropriate of the three for walls in environments that combine both high humidity and temperature conditions. Since a TSL coating is expected to enhance the BIT strength of the substrate, the elongation of TSLs and uncoated shotcrete under BIT loadings was assessed. Elongation is proportional to the vertical displacement recorded at the BIT strength. Referring to the two cited properties, the behaviours of BIT discs of the TSLs and the uncoated shotcrete were compared. The results The Journal of the Southern African Institute of Mining and Metallurgy
Sensitivity of shotcrete Brazilian discs coated with thin spray-on liners to environmental conditions
Figure 4—Sensitivity of uncoated shotcrete discs to environmental conditions
Figure 5—Sensitivity of shotcrete discs with (a) TSL1, (b) TSL2, and (c) TSL3 coatings to atmospheric conditions The Journal of the Southern African Institute of Mining and Metallurgy
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Sensitivity of shotcrete Brazilian discs coated with thin spray-on liners to environmental conditions Table II
tatistics for representative BIT specimen behaviour under room (R), saturated (H), and high humidity and temperature S (HT) conditions. CoV: Coefficient of variation; d: Days ) Type of specimens
Statistics
15 days
R
112 days
15 days
H
112 days
15 days
HT
112 days
Uncoated BIT
Mean (MPa) CoV
5.20 10%
5.20 4.76 10% 8%
4.26 4.77 4.22 7% 5% 7%
TSL1 coated BIT
Mean (MPa) CoV
5.38 6%
6.30 5%
5.38 7%
5.00 5.40 4.38 6% 8% 8%
TSL2 coated BIT
Mean (MPa) CoV
5.28 7%
5.66 8%
5.35 5%
5.25 5.07 4.60 7% 3% 10%
TSL3 coated BIT
Mean (MPa) CoV
5.59 8%
6.07 6%
5.43 6%
5.03 5.40 4.98 8% 6% 6%
Figure 6—Effects of environmental conditions on shotcrete BIT discs coated with (a) TSL1, (b) TSL2, and (c) TSL3
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Sensitivity of shotcrete Brazilian discs coated with thin spray-on liners to environmental conditions are displayed in Figures 9, 10, and 11. These results support resorting to TSL1 under room conditions, while TSL2 should be used in saturated conditions, although having slightly lower elongation capabilities than shotcrete over longer periods of exposure. TSL3 has almost the same elongation ability with shotcrete in HT conditions and should be used in such conditions.
Modelling of a shotcrete disc coated with TSL under R and HT conditions The influence of a TSL coating on the tensile strength of rock and shotcrete was evaluated by Mpunzi et al. (2015) by means of BIT testing in the laboratory. This work indicated that the tensile strength of the shotcrete disc was increased by some 30 to 40% when coated with a TSL. As part of this evaluation, they carried out numerical modelling of a shotcrete disc coated with TSL, but that investigation was limited to TSL properties under room environmental conditions. To extend this work in the current research, and with the aim of simulating TSL performances when coated on underground walls subject to extension cracks, modelling of coated discs was carried out for two environmental conditions, R and HT. In these analyses, to focus on the impact of the coating on the substrate BIT strength, the material properties of the substrate in HT conditions were maintained constant and equal to those for room conditions. Input strength and deformation parameters for the analyses were adopted from the laboratory results, and TSL1 properties were used. The numerical modelling was carried out using the Rocscience RS3 three-dimensional finite element stress analysis package. The factor of safety (FoS) was assessed for BIT coated specimens exposed to varied atmospheric conditions for 112 days. To evaluate risk associated with the adoption of TSL properties, it is necessary to determine the probability of failure (PoF). Since a probabilistic analysis is not incorporated in RS3, the response surface methodology (RSM) (Chiwaye and Stacey, 2010;
Kanda and Stacey, 2016) was used to assist in generating FoS probability distributions with the help of an Excel spreadsheet add-in simulation package, Oracle Crystal Ball. From the distribution of FoS, the PoF can be obtained and the associated risk determined. In this analysis, only two strength properties of the TSL were considered, namely the uniaxial compressive strength (UCS) and the uniaxial tensile strength (UTS).
Model discretization A disc of 42 mm diameter, 21 mm thick and coated with a 5 mm TSL was modelled, as illustrated in Figure 7. The circular section was subdivided into 40 equal parts. Loading was applied to four of these parts, and the four diametrically opposite parts were restrained; these represented the platen conditions. Loading of the model was chosen to produce an FoS between 1.50 and 2 in R condition, an assumption made to create stability. Table III shows the input parameters for shotcrete, and the TSL room and HT conditions.
Modelling results The base cases of the FoS obtained from modelling of the coated BIT disc from BIT models related to R and HT conditions were
Table III
Input parameters of coated BIT model for R and HT conditions (E: Young’s modulus) Material inputs E (MPa) Compressive strength (MPa) Poisson’s ratio Tensile strength (MPa) Unit weight (MN/m3) Thickness (mm)
Substrate
TSL R
TSL HT
48 650 52 0.276 N/A 0.021 21
13 740 32.78 0.186 4.60 N/A 5
10 010 20.10 0.180 1.90 N/A 5
Figure 7—Coated BIT model The Journal of the Southern African Institute of Mining and Metallurgy
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Sensitivity of shotcrete Brazilian discs coated with thin spray-on liners to environmental conditions 1.61 and 1.15, respectively. The RSM dictated the determination of a total of 2n+1 FoS values, corresponding with five values for each model type, as illustrated in Tables IV. The FoS distributions generated through RSM are presented in Figures 8a and 8b for R and HT conditions, respectively. The base cases represent the averages of the FoS, while the PoF values are presented as the certainty percentages of the FoS that are less or equal to unity. These PoFs are 14.98% and 36.44% for R and HT, respectively. Risk is defined by the relationship: Risk = PoF *Consequence The consequence was assumed to be identical in R and HT conditions. Therefore, the comparative study of the risks of using
TSLs in R and HT conditions relied solely on a comparison of the PoFs. Based on the coated BIT results from room and HT conditions, the risk of employing TSL1 in HT conditions is more than twice that for R conditions. It may therefore be concluded that it is inappropriate to design TSL rock support for deep underground HT conditions using parameters determined in normal room conditions in the laboratory.
Discussion Laboratory tests The choice of 50°C for high-temperature conditions was motivated by the fact that, although rock temperatures in deep
Figure 8—PoF charts for a coated BIT disc created using Oracle Crystal Ball for (a) R conditions and (b) HT conditions
Table IV
ombination of basic FoS values required for RSM application for (a) R condition and (b) HT condition (after Chiwaye and C Stacey, 2010). (R: Room condition; HT Humid and high temperature conditions; Sig t: tensile strength, b: reliability index) Nº Type Distribuition Mean stdev ‘-’ case a 1 2
UCSTSL R Sig tTSL R
Normal Normal
0.5 0.5
0.07 0.02
0.435 0.485
Nº Type Distribuition Mean stdev ‘-’ case b 1 2
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UCSTSL R. Sig tTSL R.
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0.5 0.5
0.07 0.02
Input variables FoS From RS3 b Best estimate ‘+’ case ‘-’ case Best estimate ‘+’ case ‘-’ case Best estimate ‘+’ case 0.500 0.500
0.565 0.515
1.520 1.560
1.610 1.610
1.710 1.640
0.944 0.969
1.000 1.000
1.062 1.019
Input variables FOS From RS3 b Best estimate ‘+’ case ‘-’ case Best estimate ‘+’ case ‘-’ case Best estimate ‘+’ case
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0.500 0.500
0.570 0.520
1.040 1.120
1.150 1.150
1.230 1.190
0.904 0.974
1.000 1.000
1.070 1.035
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Sensitivity of shotcrete Brazilian discs coated with thin spray-on liners to environmental conditions and ultra-deep mines range between 50 and 140°C (Oniyide, 2015), TSLs will usually be applied in openings that are ventilated. Therefore, the interface between the liner and the rock mass is assumed to be at temperatures between the high rock temperature and the cool ventilation temperature. H and HT BIT specimens were coated one face after the other, under room conditions. This process required three to four days for a secure setting of the coatings before the specimens were exposed to H and HT conditions for the required periods. In practice, however, there is no ‘waiting time’ between the application of liners and their exposure to the environmental conditions; and TSLs may be applied directly onto hot and moist or wet rock. Additional studies on less brittle TSLs that can soon be handled are therefore encouraged, to obtain possibly more representative strength enhancement behaviour in various environmental conditions. Other factors such as the type of tests used to assess TSL performance, ground preparation, and scale effect. may be investigated in the suggested studies. In the assessment of coated BIT strength enhancement, the decline in substrate strength that occurs in H and HT conditions was taken into account. This decline may be due to the friability of the substrate’s surface when exposed to those environmental conditions. Therefore, a coated shotcrete disc should not be fully affected by the weakening effect of H and HT conditions. This assumption is supported by the fact that the shotcrete surface is covered with TSL before it is exposed to more aggressive environmental conditions. However, the strengths of the coated BIT specimens tested after long exposures in H and HT conditions could decline to the extent of becoming lower than the uncoated substrate strengths. This proves that there has been a loss of the substrate strength. Otherwise, if the substrate did not lose strength, the recorded BIT strength could have remained, in the worst scenario, not less than that of the uncoated shotcrete. Thus, the shotcrete worst case was preferred for analyses. Although wallrock deformation behaviour in underground conditions may be more complex than in the BIT specimens, the BIT tests have provided a reasonable basis for the evaluation of TSL comparative performance under room and HT conditions. A testing method that could better simulate the wallrock deformation than the BIT tests might bring more understanding and knowledge in the consideration of rock support using TSLs.
Numerical modelling In the modelling, it was assumed that the material properties of the shotcrete substrate in HT conditions are equivalent to those under R conditions. This assumption is not borne out by the results of the laboratory tests, which showed that HT conditions affected the tensile strength of the shotcrete. However, the assumption was incorporated in the modelling to focus attention on the enhancement effect of TSLs in varied environmental conditions. Moreover, the declining strength of the substrate followed a negative logarithmic curve, indicating that the strength remained constant after a longer exposure period. In practice, the rock mass with shotcrete surface support in deep mines will have been subjected to such environmental conditions, and thus have already reached its long-term strength. Only TSL1 properties were modelled and the resulting risk in using this TSL in room and HT conditions applies only for this liner. Further analyses need to be undertaken for other TSL types – for example, TSL3 showed better strength resistance in HT conditions. In addition, single values of the TSLs’ UCS and UTS The Journal of the Southern African Institute of Mining and Metallurgy
were adopted, whereas the laboratory tests showed that these properties were normally distributed, and this variability needs to be taken into account. Furthermore, the bonding properties of TSL should also be included in such studies. However, the omission of these parameters in the current study is justified by the assumption of their constancy and the fact that the analysis has been a comparative one. The sensitivity of the FoS evaluations to the change in TSL material properties due to the environmental effects suggests that the combined numerical analysis/FoS approach could have potential for design calculations. However, further research will be required to confirm this.
Conclusions Previous publications have indicated that TSLs can provide significant benefit in the support of underground excavations in rock masses. Laboratory testing of TSL materials and indirect tensile testing of TSL-coated rock and shotcrete discs have confirmed this supposition. However, the implementation of TSLs as support in mines has been limited, possibly due to scepticism from mining operators regarding their effectiveness in the mining environment. Most TSL test data reported in the literature has been obtained in ’standard’ room temperature and humidity environments, which are not typical of conditions in deep mines. In the research described in this paper, results of intensive and reproducible tests carried out on TSL materials exposed to high humidity and 50°C conditions have been described. Indirect tensile testing of TSL-coated shotcrete discs was carried out to determine whether TSL performance under deep mine conditions matched that in normal laboratory conditions. The results showed that a TSL coating can enhance the tensile strength of the shotcrete under the deep mine conditions, but not to the same extent as indicated for normal laboratory conditions. Whereas in the laboratory the tensile strength of the tested TSLs was enhanced by between 9% and 21%, in the deep mine conditions the improvement varied between 4% and 18%. The degree of enhancement depends on the duration of exposure to the more severe conditions, and on the type of TSL. In addition to the laboratory testing, numerical modelling of coated BIT test specimens was carried out. These tests showed that, based on the input parameters from laboratory testing, the factor of safety could be as high as 1.61 after 112 days of exposure to room (R) conditions. However, when exposed for the same time to high temperature and humidity (HT) conditions, the factor of safety may be closer to unity. Extending these studies to probabilistic analyses showed that the resulting probabilities of failure of the coated BIT specimens for R and HT conditions would be about 15% and 36%, respectively. This indicates that to ensure that the rock support will provide a safe and stable working environment, it is essential to determine the properties and performance of TSLs under environmental conditions that are typical of the mining environments in which they will be applied.
Acknowledgements The first author would like to thank the High Polytechnic Institute of Tete, Instituto Superior Politécnico de Tete (ISPT), Mozambique, for financial support.
References
Boeg-Jensen, P. and Swan, G. 2014. The operational and laboratory aspects of a thin spray-on liner. Proceedings of the 7th International Conference on Deep and High Stress Mining. Australian Centre for Geomechanics, Perth. pp. 241–251. VOLUME 121
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Sensitivity of shotcrete Brazilian discs coated with thin spray-on liners to environmental conditions Chiwaye, H. and Stacey, T.R. 2010. A comparison of limit equilibrium and numerical modelling approaches to risk analysis for open pit mining. Journal of the Southern African Institute of Mining and Metallurgy, vol. 110. pp. 571-580. Guner, D. and Ozturk, H. 2018. Creep behaviour investigation of a thin spray-on liner. International Journal of Rock Mechanics and Mining Sciences, vol. 108. pp. 58–66. Guner, D. and Ozturk, H. 2016. Experimental and numerical analysis of the effects of curing time on tensile mechanical properties of thin spray-on liners. Rock Mechanics and Rock Engineering, vol. 49. pp. 3205–3222. Hoek, E., Kaiser, P.K., and Bawden, W.F. 2000. Support of Underground Excavations in Hard Rock. CRC Press, Boca Raton, FL. Kanda, M.J. (In press). Long term performance of thin spray-on liners for permanent support in underground mines. PhD thesis, University of the Witwatersrand. 359 pp. Kanda, M.J. and Stacey, T.R. 2020. Tensile strength sensitivity of thin spray-on liners to changes in environmental conditions. Journal of the Southern African Institute of Mining and Metallurgy, vol. 120, no. 4. pp. 251–259. Kanda, M.J. and Stacey, T.R. 2019. Review of the practical effectiveness of thin spray-on liners based on information from suppliers and observations from the mining industry. Proceedings of the First International Conference on Mining Geomechanical Risk. Australian Centre for Geomechanics, Perth. pp. 443–458. Kanda, M.J. and Stacey, T.R. 2016. The influence of various factors on the results of stability analysis of rock slopes and on the evaluation of risk. Journal of the Southern African Institute of Mining and Metallurgy, vol. 116. pp. 1075–1081. Masethe, R. 2015. Quantification of the effects of thin spray-on liner application on shotcrete tensile strength. MSc dissertation, University of the Witwatersrand. Mpunzi, P., Masethe, R., Rizwan, M., and Stacey, T.R. 2015. Enhancement of the tensile strengths of rock and shotcrete by thin spray-on liners. Tunnelling and Underground Space Technology, vol. 49. pp. 369–375. Oniyide, G. 2015. Thermo-mechanical behaviour of rocks from the Bushveld Igneous Complex with relevance to deeper mining. PhD thesis, University of the Witwatersrand. 363 pp. Osswald, T.A. 2015. Understanding polymer processing: processes and governing equations. Carl Hanser Verlag GmbH Co KG, Munich.
Potvin, Y., Stacey, T.R., and Hadjigeorgiou, J. (eds). 2004. Surface Support in Mining. Australian Centre for Geomechanics, Perth. Simser, B. and Pritchard, C. 2012. Innovative ground control at Xstrata's Nickel Rim South Mine, Sudbury, Ontario. Presentation: Work Place Safety North, Sudbury, April 2012. https://www.workplacesafetynorth.ca/sites/default/files/resources/ Innovative%20Ground%20Control%20at%20Xstrata.pdf Stacey, T.R. 2001. Review of membrane support mechanisms, loading mechanisms, desired membrane performance, and appropriate test methods. Journal of the South African Institute of Mining and Metallurgy, vol. 101. pp. 343–351. Tannant, D. 2001. Thin spray-on liners for underground rock support—Testing and design issues. Proceeding of the International Seminar and Field Trials on Surface Support Liners: membrane, shotcrete and mesh. Australian Centre for Geomechanics, Perth. pp. 1–18. Ulusay, R. and Hudson, J.A. (eds). 2007. The Complete ISRM Suggested Methods for Rock Characterization, Testing and Monitoring: 1974-2006, Compilation arranged by the ISRM. Turkish National Group, Ankara, Turkey. 628 pp. Yilmaz, H. 2010. Tensile strength testing of thin spray-on liner products (TSLs) and shotcrete. Journal of the Southern African Institute of Mining and Metallurgy, vol. 110. pp. 559–569.
Appendix Declarations Financial support: The research was supported by the University of the Witwatersrand, which availed both logistics and financial support whenever needed. The High Polytechnic Institute of Tete (Instituto Superior Politécnico de Tete (ISPT)) also contributed through financial support to the first author to complete his studies. Conflicts of interest/Competing interests: The authors declare that they have no conflict of interest
Figure 9—Comparisons of shotcrete and TSL BIT vertical displacements after 15 days of exposure in (a) room, (b) saturated, and (c) HT conditions
Figure 10—Comparisons of shotcrete and TSLs BIT vertical displacements after 28 days of exposure in (a) room, (b) saturated, and (c) HT conditions
Figure 11—Comparisons of shotcrete and TSLs BIT vertical displacements after 112 days of exposure in (a) room, (b) saturated, and (c) HT conditions
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Feasibility of tailings retreatment to unlock value and create environmental sustainability of the Louis Moore tailings dump near Giyani, South Africa N.K. Singo1 and J.D. Kramers1 Affiliation: 1 Department of Geology, University of Johannesburg, South Africa. Correspondence to: N.K.Singo
Email:
kenneth@singoconsulting.co.za singo.kenneth@gmail.com
Dates:
Received: 20 Feb. 2020 Revised: 17 Apr. 2021 Accepted: 28 Apr. 2021 Published: July 2021
How to cite:
Singo, N.K. and Kramers, J.D. 2021 Feasibility of tailings retreatment to unlock value and create environmental sustainability of the Louis Moore tailings dump near Giyani, South Africa. Journal of the Southern African Insitute of Mining and Metallurgy, vol. 121, no. 7, pp. 361–368 DOI ID: http://dx.doi.org/10.17159/24119717/1138/2021 ORCID: J.D. Kramers https://orcid.org/0000-00026223-9530
Synopsis The reprocessing of tailings resources to extract gold on an industrial scale has become common practice. While these projects are common in the Witwatersrand basin, similar low-technology processes are not operational in smaller goldfields. This study explores the possibility of reclaiming the tailings dump of the Louis Moore Mine in the Giyani Greenstone Belt, investigates potential hazards to communities in the vicinity, and identifies mitigation strategies. Auger samples were taken from the Louis Moore tailings at depths of up to 5 m. Aqua regia leach analyses show Au concentrations of up to 1 g/t. Inferred estimations based on ordinary kriging (OK) and inverse distance weighted (IDW) methods put the residual Au resource in the Louis Moore tailings dump at 0.20 t. Reworking the tailings is viable, although a potential environmental risk exists in the form of elevated arsenic concentrations. Further exploration is required to determine the mineralogical associations of Au and As. Tailings reworking would assist in raising funds for mine rehabilitation. The secondary tailings could potentially be repurposed, which would provide employment and facilitate community development, as well as deliver environmental benefits. Keywords Louis Moore, tailings, gold, safety and health.
Introduction The tailings retreatment sector in South Africa has been receiving a lot of attention due to the economic, environmental, financial, and political imperatives impacting new and existing tailings facilities. Blake (2013) asserts that some mining houses are exploring the unlocking of revenue in old tailings facilities because of the increasing cost of deep level mining, including operations with low recoveries (Sabbagha, 1982). The increase in commodity prices (Daniel and Downing, 2011) has led to the realization of retreating tailings. In most cases, extraction (>90%) has been by means of by bioleaching (Carmen et al., 2017; Ahmadi et al., 2015). Bio-processing these waste materials has economic as well as environmental benefits (Bryan et al., 2006). The Louis Moore Mine is an abandoned underground mine, situated in the Ka-Mavalane village, about 10 km north of the town of Giyani along the road leading to Malamulele in Limpopo Province. Tailings constitute the only significant mine residue at the site. This study assesses the environmental hazards associated with these tailings and explores the economic potential of reclaiming gold. Environmental hazards associated with this type of deposit are potentially high levels of arsenic (As) from arsenopyrite, which is normally associated with Au in shear zone-hosted Au deposits in Archean greenstone belts (Ward, 1999), as well as acid mine drainage (AMD) and its associated metal content (Alpers et al., 1994). The economic potential is strongly dependent on the Au content. To assess risk and economic potential, the Louis Moore tailings were systematically sampled by auger, and geochemical analysis, mineralogical studies, and resource modelling undertaken to determine the potential of reworking the tailings.
Geological setting The Giyani Greenstone Belt (GGB, formerly known as the Sutherland Greenstone belt) is situated at the northern edge of the Kaapvaal Craton. Its northwestern boundary defines the contact with the high-grade metamorphic Limpopo Belt (McCourt and van Reenen, 1992; Carranza, Sadeghi, and Billay, The Journal of the Southern African Institute of Mining and Metallurgy
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Feasibility of tailings retreatment to unlock value and create environmental sustainability 2014; Kramers, Henzen, and Steidle, 2014), with multiple shear zones developed along this contact. The now defunct Klein Letaba, Franke, Gemsbok, and Fumani mines are examples of Au deposits developed along such shear zones (McCourt and van Reenen, 1992; Gan and van Reenen, 1996). The Louis Moore Mine is not situated in the main greenstone belt, but in a small greenstone belt remnant in the high-grade metamorphic gneisses to the north. Owing to poor outcrop conditions, no accurate geological map exists (Billay, Sadeghi, and Carranza, 2014). In the immediate surroundings of the Louis Moore Mine, banded gneissic rocks and micaceous schists occur, and foliation has a general east-west strike and a northerly dip. The Louis Moore orebody is hosted in tremolite schist, amphibolite, and granitoid gneiss, and is associated with quartz veins (Lombard, 1956). However, in comparison with other Au mineralization in the region and from the sparse outcrop mapping, it is likely that the orebody is associated with shear zones (Gan and van Reenen, 1996).
which was stored. For analysis of major and trace elements by ICP-OES and ICP-MS, a further 1/10 dissolution was done and samples were analysed from a 0.3 M nitric acid solution. The instruments used included a SPECTRO® analytical instrument, ARCOS ICP-OES, and a Perkin Elmer® NexIon® 300 series quadrupole ICP-MS. The pH values of slurries of sample material in deionized water were measured after being left to equilibrate overnight. Since the probes of pH meters are very sensitive and can easily be contaminated by particulates, comprehensive and user-friendly pH (pH–Fix 0-14) colour indicator strips (also called dip-in-readoff) were used to measure pH.
Statistical analysis and modelling
Material and methods
In order to understand element distribution patterns in the tailings dump, the analytical data was statistically evaluated. Inverse distance weighted (IDW) and ordinary kriging (OK) methods were trialled. VulcanTM 3D software version 9.1.7 was used to model dump solids and surfaces, followed by block model estimation.
Drilling and sampling
Results and discussion
Exploratory drilling using a traditional hand auger was carried out on the dump (which measures about 100 × 200 m) in November and December 2015. Holes were drilled to a maximum depth of 5 m, and material from every metre depth interval was split to yield a representative sample. This resulted in a total of 47 samples (less than 50, as some holes reached the base of the tailings at less than 5 m depth).
pH of tailings
Sample analysis The main analytical work was conducted at the University of Johannesburg (UJ) using ICP-OES as part of a multi-element analysis on sample leaches. Cross-checks were carried out by the Sociéte Générale de Surveillance (SGS) Johannesburg laboratory; X-ray fluorescence (XRF) was used on pressed powders, while fire assay was used for Au. Twenty samples were further analysed using quadrupole ICP-MS at UJ for Au and uranium (U) for verification. X-ray diffraction (XRD) analysis was carried out on four samples by the Council for Geoscience, Pretoria, to assess the mineral content. For the ICP-OES and ICP-MS analyses at UJ, aqua regia extraction was used, as per ISO 11466 of 1995. Aqua regia (20 ml) was added to 3.0 g of a dry tailings sample in a glass beaker. After being stored overnight at room temperature (allowing for slow oxidation of organic matter), the samples were boiled under reflux for two hours at a maximum temperature of 125°C. After cooling, the solution was filtered, evaporated to dryness, and re-dissolved to yield 100 ml of 1 M nitric acid,
The Louis Moore Mine tailings have pH values ranging from 7 to 9, averaging 8. The neutral to weak basic conditions throughout the tailings dump can be attributed to the presence of dolomite (primarily associated with the mineralization) in the tailings. Residual lime used in pH adjustment during cyanidation may contribute to these values as well. While this means that no AMD currently emanates from these tailings, the overall alkaline chemistry is not optimal for plant growth as the availability of nutrients for uptake could be limited.
Mineralogical investigations on tailings The XRD analysis results of four tailings samples, which were selected randomly, are shown in Table I. Mineralogical analyses on tailings by means of XRD are expected to show primary and secondary mineral occurrences. Primary minerals are ore and gangue minerals that were processed and deposited as tailings without any changes or alteration (Jambor, 1994). Secondary mineral phases are those that have been produced by processes that can lead to precipitation (Alpers et al., 1994; Novhe et al., 2014) and therefore formed after the tailings were deposited. Tailings from the Louis Moore Mine have a variable, mostly high, quartz content. There are significant to abundant concentrations of dolomite, amphibole, serpentine, mica, and smectite minerals in samples KSG1D, KSG2D, and KSG3B. These minerals are absent in KSG5D, which contains plagioclase
Table I
XRD analysis of selected Louis Moore Mine tailings samples (wt. %) Sample
Dolomite
KSG1D KSG2D KSG3B KSG5D
20 5 43 –
Gypsum Talc/pyrophyllite Clinopyroxene Ilmenite Amphibole Serpentine Plagioclase Quartz Clinochlore Mica Smectite – 2 – –
– 2 2 –
7 – – –
3 – 4 –
35 31 23 tc
3 2 6 –
4 – – 13
15 30 8 51
– 5 – 33
5 16 5 2
7 7 8 –
Notes: (-) Not detected (tc) Trace
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The Journal of the Southern African Institute of Mining and Metallurgy
Feasibility of tailings retreatment to unlock value and create environmental sustainability and clinochlore (Table I). It appears that different gangue assemblages and/or wallrocks were encountered at different times during mining. Generally, the mineral content reflects a mixture of country rocks typical of greenstone belt lithologies and granodiorite intrusions. The significant dolomite component is somewhat surprising and may reflect a carbonatization process related to ore emplacement, as found elsewhere in GGB Au deposits (Gan and van Reenen, 1996). Pyrrhotite or pyrite were not detected by XRD in the Louis Moore tailings but may be present in small amounts. The absence of jarosite, which is commonly associated with AMD, also attests to a low sulphide content in the past. The mineralogy is in accordance with the neutral to mildly alkaline pH values observed.
Contaminative elements assessment summary Concentration data for elements of interest in terms of value or environmental hazard is listed in Table II. This section focuses mainly on the negative impacts of elements. Arsenic can cause cancer, nervous system damage, as well as skin, lung, liver, and heart damage. Chromium (Cr) is carcinogenic in the soluble hexavalent state. Copper (Cu) and nickel (Ni) are non-cumulative when ingested, but large doses of Cu can cause digestion and kidney problems, while Ni inhalation can cause respiratory problems and even cancer. Lead (Pb) is a cumulative toxin and particularly affects brain and nervous system development in children. High levels of U emit radioactivity, which is a serious problem with tailings from the Witwatersrand gold mines. If inhaled, U causes damage to lungs, and after entering the bloodstream, the absorbed uranium tends to bio-accumulate and remain for many years in bone tissues (Glaser, Hippel, and Frank, 2006). Very high uranium intakes (50 to 150 mg) can cause acute kidney failure and death. At lower intake levels (25 to 40 mg), damage can be seen by the presence of protein and dead cells in the urine. It is important to note that kidneys repair themselves over a period of 8 weeks after uranium exposure has ceased, but only if the intake was at low levels (WHO, 2000). The Louis Moore tailings do not show U concentrations above normal uncontaminated soil levels (1–2 µg/g). They contain significant concentrations of the metalloid As (mean 277 µg/g) and contaminative metals Cr, Cu, Ni, Pb, and zinc (Zn). The average values for Cd and Pb are within allowable limits for soils. Those for Cu and Zn exceed target values for soils only
slightly, while concentrations of Cr, Ni, and particularly As, are well above South African and international soil target values. The concentration of As in natural soil ranges from 1 to 40 µg/g. While in most of the boreholes As concentrations are below detection levels from bottom to top, three boreholes have high As levels over their complete depth, and one has high levels at 1–2 m only. Thus, while there does not seem to be a vertical zonation of As, there is a very irregular horizontal distribution. Arsenic is fairly well correlated with P, and high P (526.95 µg/g) levels are found in the same boreholes that also have high As contents. Under oxidizing conditions, As and P are present in solution as the arsenate (AsO43-) and phosphide (PO43-) oxyanions, respectively, which are both adsorbed on, and co-precipitated with, Fe oxides and hydroxides, although this tendency is stronger for arsenate than for phosphate (Strawn, 2018). Under moderately reducing conditions, As is present as the arsenate (AsO33-) ion, which is not adsorbed on Fe oxides, while P remains as the phosphate ion (Strawn, 2018). Thus, the correlation of P with As indicates that (a) conditions were oxidizing, and (b) both elements were in solution during sedimentation of the tailings. This could have led to localized concentrations of both elements in the tailings. Borehole KSG5 (one of the sites with high As and P concentrations over the whole depth) contained stagnant water during sampling. If this was a low point during deposition of the tailings, it might have led to local concentration of dissolved As and P which then became fixed by adsorption. Arsenic is immobile in soils under neutral and slightly acidic conditions (Blowes et al., 2003), but can be mobilized at pH<2 or >9 (Gersztyn, Karczewska, and Gałka, 2013). Although the pH values up to 9 found in the tailings are very marginal to the conditions for As mobilization, potentially high levels of As in the groundwater around Louis Moore Mine are a concern. Communities are discouraged from residing or working on or adjacent to these tailings, since the dust blown from them could still be rich in As. Cu, Ni, Pb, and Zn would be released as cations under natural leaching conditions. The leaching of these elements from the tailings and their mobility in groundwater is expected to be very low at neutral pH, increasing somewhat at pH values >9 as Cu(OH)2, Pb(OH)2, Ni(OH)2 and Zn(OH)2 complexes are formed. Furthermore, Cr when oxidized to Cr6+ is highly mobile at high
Table II
Summary of contents of elements of interest (µg/g)
As(1) No. of samples Minimum Maximum Range Mean Median Standard deviation Detection limit WHO limits (soil)
47 0.00 2151 2151 277 0.00 456 8.3 40
Au(3) Cd(2) Cr Fe 47 0.00 0.94 0.94 0.37 0.39 0.25 0.08 n/a
47 0.00 0.8 0.8 0.4 0.00 0.3 0.8 0.8
47 69.0 727 658 359 389 190 5.4 100
47 0.00 134.5 134.5 40.47 16.60 39.70 3.3 36
P(1)
47 23710 82661 58950 51191 53062 13736 8.3 n/a
P 47 0.00 526.9 526.9 113.8 0.00 161.1 11.7 n/a
Ni 47 43.16 459.74 416.58 224.39 204.64 116.41 8.3 35
Pb
U(3)
47 0.00 87.52 87.52 26.88 29.00 25.00 10.0 85
15 1.09 1.33 0.24 1.15 1.14 0.06 0.01 n/a
Zn 47 25.23 128.03 102.80 69.79 68.81 24.53 8.3 50
Notes: 1. Arsenic was detected in only 16 samples from just 4 boreholes (see Figure. 1). The mean As concentration for the samples in which As was detected is 724 ± 456 µg/g (1 SD). Phosphorus (P) was detected in all 18 samples from these boreholes; the mean P concentration being 255 ± 148 µg/g (1 SD). 2. Cadmium (Cd) was not observed above the detection limit of 0.8 µg/g, which is listed as the maximum concentration. 3. Au analyses were checked by ICP-MS (15 samples selected randomly) which confirmed the ICP-OES data. U could not be analysed by ICP-OES due to interference of an iron (Fe) line, and 15 samples were analysed by ICP-MS. The Journal of the Southern African Institute of Mining and Metallurgy
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Figure 1 – Strip log showing As concentrations and distribution in all boreholes
Figure 2—Histogram of Au distribution
pH, which could possibly be a serious environmental concern in the surroundings of the Louis Moore tailings (van der Sloot and van Zomeren, 2012). However, the Cr in the tailings is most likely hosted in silicates and oxides (notably chromite). Although this Cr was oxidized and dissolved in the aqua regia leach, chromite would be stable in natural weathering processes.
Distribution of Au in auger holes During drilling a general homogeneity and lack of distinct layering was observed in the tailings. Generally, in all ten holes, the effects of dumping season and fluctuating mining operations are evident. Given the apparent lack of mobility for Au (and other trace elements), these factors are the main cause of variation in concentrations. In all drilled holes, Au concentrations are high (up to 1 ppm), which is attributed to poor plant recovery efficiency during mining. These concentrations require a formal geometallurgical study to help assess the economic feasibility of
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extracting Au, with a view to re-treating the tailings, which, in turn, will assist in cleaning up the environment and allowing the land to be further developed. The histogram for Au data shown in Figure 2 presents an even and random (non-skewed and nonnormal) distribution of concentrations. A tailings dump is not a natural deposit geologically formed in situ, and the deposition of Au in it is not controlled by geological processes of ore deposit formation. To draw a solid conclusion when working with tailings is therefore very difficult. Although this is not the subject of this section, it is worth noting that no significant systematic enrichment or depletion with depth has been noted for most trace and major elements analysed in this study. Generally, the random variation of Au values indicates the effects of dumping season and fluctuating mining operations, as well as variations in the ore grade encountered during mining. This presents a positive outlook for further exploration and exploitation. The Journal of the Southern African Institute of Mining and Metallurgy
Feasibility of tailings retreatment to unlock value and create environmental sustainability Modelling (block model development) and grade estimation Survey points (shown in red in Figure 3) were used to model the top surface of the dump together with drill-hole collars (shown in green). Bottom strings digitized from end-of-hole points were used to model the floor or base. The top and bottom surfaces of the Louis Moore Mine dump were appended to form a 3D closed solid on which the block model design was based, and which serves to estimate the volume of the tailings dump for an estimate of the total gold content.
Block model design Two block models were created for the entire tailings dump area, which included the upper and bottom surfaces as constraints. Air blocks outside dump domains were deleted to optimize block model size and ensure that no air blocks outside the dump triangulations were estimated. Block models were validated by visual inspection and creating diagonal sections across dump domains. For visualization purposes, the tailings dump was exaggerated ten times in the Z dimension. This was done because the 5 m drill-holes resulted in dense Z extents on the block model. This exaggeration does not change the extents of the block and triangulation models; it is solely for the purpose of visualization.
Visual validation Visual validation was performed by viewing cross-sections and comparisons of the grade in the block model to the sample
database. The correlation between the sample grades and the block model is good. In order to check for potential biases in the block model grade estimate, the IDW block grades were compared to the OK estimation at a zero cut-off grade. OK estimation provides one of the best estimates of the mean grade of a tailings because the drill-hole sample data is declustered. In Figure 5, the two grade models are compared, and it can be seen that they provide a very similar prognosis for tonnage to be mined as a function of the cut-off grade.
Estimation of the total Au resource within the tailings dump No density measurements were conducted in the UJ laboratory on tailings samples. The tonnage calculated is based on densities obtained from the previous studies and laboratory tests. South Africa’s Witwatersrand gold ores have densities ranging from 2.74 kg/m3 to 2.77 kg/m3, based on calculations for rocks containing 94% silica with accessory pyrite. The density of greenstone schists and gneissic rocks, on the other hand, varies between 2.6 and 2.9 kg/m3. The banded iron formation (BIF) records 3.6 kg/m3 (Hacto Corp, 2011). As such, a density of 1.4 kg/m3 for the Louis Moore tailings dump was assumed. The total tonnage of tailings material was used to deduce the total amounts of the various elements. The total area for the Louis Moore Mine tailings dump is 4.3 ha (43 000 m2) and the volume is 387 000 m3. The average density is 1.4 kg/m3. No geological losses (dykes, sills or even washouts) were included
Figure 3—A starting basin showing solid modelled surfaces for the topography of the Louis Moore tailings dump (grey surface indicates tailings)
Figure 4—Barren (un-estimated) Louis Moore Mine tailings dump block model section before grade estimation, exaggerated in the Z dimension by 10× to view the blocks section The Journal of the Southern African Institute of Mining and Metallurgy
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Figure 5—Typical block model cross-section of Au grades. W-E section showing Au sample data and block model estimates of the Louis Moore Mine tailings dump
Figure 6—The Louis Moore Mine tailings dump grade/tonnage report for gold, comparing OK (au_ok) and IVD (au_ivd) estimation
in the mineral resource estimation, as there should not be any structure present in the dump that will cause losses and a 100% reclamation of the dump is assumed (thus corresponding to a cutoff grade of zero in Figure 5 and complete removal of the dump). Taking the mean Au concentration from Table II (0.000037%), the total metal tonnage (T, in metric tons) is derived from the volume (V, in m3), the density (r, in t/m3) and the grade (C, in %) as T = V × r × C/100 = 387 000 × 1.4 × 0.000037/100 = 0.20 t (assuming 100% recovery).
Conclusion Conclusions from the data reported in this study relate to (a) environmental and health risk, and (b) potential Au exploitability. There is no indication of AMD emanating from the tailings (pH values range from 7 to 9), but a potential health risk is presented by As, which could be mobilized under mildly alkaline conditions and contaminate soils and groundwater. It is unlikely that groundwater has been contaminated with heavy metals; the pH of the tailings is insufficiently high to mobilize Cu, Ni, Pb, and Zn as OH complexes. While Cr6+ could be mobile, the stability of its host mineral phases probably prevented its liberation. The main risk is associated with the possible inhalation or ingestion of dust by residents of the nearby community. Any remedial measures
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should include stabilizing the dump surface (slope stabilization), e.g., by vegetation to prevent loose soil from being blown or washed away) or engineered covers comprising a water-saturated clay liner to minimize oxygen ingress, thus reducing oxidation rates. Communities living in affected areas could be relocated, but intervention measures requiring a relatively small budget, such as fencing to limit open access to the tailings by humans and animals, are recommended. The Au resource enhances the economic viability of an environmentally beneficial surface retreatment project. Further exploitation of this dump using simple technology is proposed; for example, sluice guns could be used to mine the dump and limit dust generation. The site is located 200 m away from a perennial river, which could (assuming a water use licence is granted) serve as a source of water. Reclamation of the mined dump can then follow to mitigate further environmental challenges. Ultimately, all feasibility studies on future land use must consider the possibilities for proper rehabilitation of these areas, following the removal of the physical obstacles to development, and see to it that they are developed to their full potential. The main aim is to extract the residual minerals. Following extraction of the valuable metals, the following potential land uses should be considered. Power generation or urban development: The tailings dump covers a large area of land and is elevated, and hence would be a suitable site for wind turbines or a solar (photovoltaic) energy farm. Being in the tropics, solar radiation is intense. The tailings are less than 500 m from the village, which makes it inexpensive and easy to connect power lines. Alternatively, if the dump can be reclaimed and all toxic metals extracted, the area could be used for village extension. Heritage site: The Louis Moore mines are over 100 years old and qualify to be declared heritage sites in terms of the National Heritage Resource Act, 1999 (Act No. 25 of 1999) under Section 32 (5) (b) (i). Academic and industrial research potential: Due to the complexity of the tailings, geometallurgical studies must be undertaken before deciding on the viability of reworking, presenting opportunities for further research. Geology and mining engineering students could use these sites as case studies. This study has revealed a number of geological and mineralogical issues that can be further investigated in detail. Community development projects: Tailings materials have similar mechanical properties to the clay used in the brick and tile The Journal of the Southern African Institute of Mining and Metallurgy
Feasibility of tailings retreatment to unlock value and create environmental sustainability industry. The option of making bricks and tiles will lead to jobs being created in the re-mining/retreatment process, along with training/skills transfer to the local community. It is important to use materials with that low heavy metal concentrations to avoid contamination of land and water or possible adverse health effects for occupants of dwellings built using contaminated material.
Acknowledgements Analytical work external to UJ was funded jointly by the Geological Society of South Africa Research Education and Investment Fund (GSSA REI) and the DST-NRF Centre of Excellence for Integrated Mineral and Energy Resource Analysis (CIMERA).
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Council for Geoscience, Pretoria. Ronaldson, J.H. 1909. Report on the Louis Moore Gold Mine, Klein Letaba Gold Field. Rosner, T., Boer, R., Reyneke, R., Aucamp, P., and Vermaar, J. 2001. A preliminary assessment of pollution in the unsaturated and saturated zone beneath reclaimed gold-mine residue deposits. Report 797/1/01. Water Research Commission, Pretoria, South Africa. pp. 1–2. Ryan, B. 2014. Giyani gold hopes it is a case of third time lucky. Business Day, 4 July 2014. https://www.businesslive.co.za/bd/companies/mining/2013-07-04news-analysis-giyani-gold-hopes-it-is-a-case-of-third-time-lucky/ Sabbagha, C.M. 1982. Practical experiences in pumping slurries at ERGO. Proceedings of the 8th International Conference on the Hydraulic Transport of Solids in Pipes. BHRA Fluid Engineering, Wharley End, Bedfordshire, UK. Strawn, D.G. 2018. Review of interactions between phosphorus and arsenic in soils from four case studies. Geochemical Transactions, vol. 19. doi: 10.1186/ s12932-018-0055-6 Taylor, C.D., Schulz, K.J., Doebrich, J.L., Oris, G.J., Denning, P.D., and Kirschbaum, M.J. 2005. Geology and non-fuel mineral deposits of Africa and the Middle East. Open File Report 2005-1294-E. US Geological Survey. Van der Sloot, H.A. and van Zomeren, A. 2012. Characterisation leaching tests and associated geochemical speciation modeling to assess long term release behavior from extractive wastes. Mine Water and the Environment, vol. 31. pp. 92–103 Wang, L.C., Liu, Q., Huang, X.S., Liu, Y.M., Cao, Y., and Fan, N. 2009. Gold nanoparticles supported on manganese oxide for low-temperature Co oxidisation. Applied Catalysis B: Environmental, vol. 88, no. 1-2. pp. 204–212. Ward, J.H.W. 1999. The metallogeny of the Barberton Greenstone Belt, South Africa and Swaziland. Memoir 86. Council for Geoscience, Pretoria, South Africa. 116 pp. Wedepohl, H. 1995. The composition of the continental crust. Geochimica et Cosmochimica Acta, vol. 59. pp. 1217–1232 Wedepohl, K.K. 1978. Handbook of Geochemistry. Volume 2, Part 5. SpringerVerlag., Berlin, Heidelberg, New York. Wilkins, B. 2013. The changing face of tailings retreatment. Modern Mining, August 2016. Windley, B.F. 1997. The Evolving Continents. Wiley, New York. WHO. 2000: The chemical toxicity of uranium. World Health Organization, Geneva. VOLUME 121
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Improving the environmental and economic aspects of blasting in surface mining by using stemming plugs A. Ur Rehman1, M.Z. Emad2, and M.U. Khan2 Affiliation: 1 Missouri University of Science & Technology, Rolla, MO, USA. 2 University of Engineering and Technology Lahore, Pakistan. Correspondence to: M.U. Khan
Email:
usman@uet.edu.pk
Dates:
Received: 11 Mar. 2021 Revised: 25 Jul. 2021 Accepted: 26 Jul. 2021 Published: July 2021
How to cite:
Rehman, A. Ur, Emad, M.Z., and Khan, M.U. 2021 Improving the environmental and economic aspects of blasting in surface mining by using stemming plugs. Journal of the Southern African Insitute of Mining and Metallurgy, vol. 121, no. 7, pp. 369–378 DOI ID: http://dx.doi.org/10.17159/24119717/1573/2021 ORCID: A. Ur Rehman https://orcid.org/0000-00032299-6251 M.Z. Emad https://orcid.org/0000-00018537-8026 M.U. Khan https://orcid.org/0000-00026260-6679
Synopsis The use of stemming plugs in surface blasting is gaining popularity because of the environmental, technical, and economic advantages. Although previous studies have established the operational effectiveness of different types of stemming plugs, evaluations of the economic and ergonomic impacts are lacking, thereby hindering their application in many mining operations. Consequently, a significant proportion of mining operations end up performing less efficient blasts. We evaluated the effectiveness of three types of stemming plugs by conducting multiple full-scale production blasts. The results show that stemming plugs reduce the need for secondary blasting and increase blast performance. An economic analysis showed that the incorporation of stemming plugs can reduce blasting costs significantly. Keywords stemming plugs, economic analysis, blast performance, surface mining, confinement.
Background Optimum utilization of explosive energy is a prime consideration in blasting due to rising costs, production pressures, financial challenges, environmental constraints, and other factors. Blast performance is usually measured in terms of rock fragmentation, the shape of the muckpile, and resulting blasting fumes. The explosive output affects the entire mining operation from post-blast scaling to the overall efficiency of the ore processing plant. Thus, minimization of the blasting cost, safety risks, and environmental impacts is extremely desirable in surface mining operations. All of these factors can be positively impacted by improved rock fragmentation (Mohamad et al., 2013; ur Rehman, 2017; ur Rehman, Emad, and Khan, 2019). Increased mineral production demands have put immense pressure on mines and quarries. Surface mines use drilling and blasting as this is historically the most economical way of rock breakage. Existing operations primarily scale up production by increasing the blast size. This situation creates an imbalance between rock fragmentation and other blasting factors like ground vibration, flyrock, air blast, and post-blast fumes. Various blasting parameters affect blast efficiency and the optimization of optimizing these parameters to achieve an efficient blast is a complex process (Khandelwal and Singh, 2006; Singh and Singh, 2005; Yang and Liu, 2010; Yang et al., 2010). Zhu et al. (2008) state that blast performance depends on the blast geometry, rock mass properties, and explosive specifications. Mohamad et al. (2013) established that blast-hole diameter, stemming length, burden priming, spacing, and delays are important for blasting dynamics and hence the rock fragmentation. On surface mines, the tops of the blast-holes are filled with an inert material, called stemming, after the hole has been charged,. This practice helps in confining explosives gases, improves rock fragmentation, and reduces flyrock. The stemming provides confinement for the gases created by detonation of the explosive. This helps improve rock fragmentation, prevents ejection of the collar rock, and increases safety. Stemming material redirects the explosives energy back into the rock to facilitate its breakage ( Cevizci, 2012).
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Improving the environmental and economic aspects of blasting in surface mining Drill cuttings are most commonly used as a stemming material due to their availability at the hole collar (Halim Cevizci, and Ozkahraman, 2012). Thus, the stemming material for the blast-hole is mainly comprised of loose soil, dirt, and rock cuttings. An inherent problem with using drill cuttings as a stemming material is non-uniformity in terms of particle size and compaction. This sometimes leads to blowouts from the stemming section. Due to poor stemming practice, gas venting through the stemming material occurs and causes significant loss of explosive energy. Poor explosives energy distribution results in a less efficient blast in terms of rock fragmentation and increases the requirement for secondary blasting. The sudden liberation of gases due to poor stemming can also cause flyrock. Flyrock is a major safety hazard as it can injure personnel, and damage structures and equipment. Another prime blasting consideration, in addition to the type of stemming material, is the length (depth) of stemming material. A decrease in stemming length will increase the quantity of the explosive in a blast-hole, hence increasing the explosives cost, the charge per delay, and environmental and safety hazards. A transition from drill chipping towards aggregate as stemming material has been observed, as more and more mines are using aggregate as stemming material with excellent (Gomes-Sebastiao, G.L. and de Graaf, W.W. 2017). The significance of stemming in blasting is an established topic that has been studied by various researchers (Cevizci, 2012, 2013, 2014; Cevizci and Ozkahraman, 2012; Choudhary and Rai, 2013a; Eloranta, 1994; Kojovic, 2005; Kopp, 1987; Mohamad et al., 2012; Rai and Imperial, 2005; Sazid, Saharan, and Singh, 2016; Sharma and Rai, 2015; Singh and Sastry, 1988; Trivedi, Singh, and Raina, 2014). The incorporation of stemming plugs as opposed to the conventional stemming material is gaining popularity in surface blasting operations. Stemming plugs are a well-known accessory used in blasting along with conventional stemming material. Various surface mines have been incorporating stemming plugs for the last three decades and numerous varieties of commercial plugs are currently available from several manufacturers. Different researchers have studied the efficacy of stemming plug compared to conventional stemming ( Cevizci, 2012; Choudhary and Rai, 2013a; Kojovic, 2005; Kopp, 1987; Mohamad et al., 2012; Rai and Imperial, 2005; Rai, Ranjan, and Choudhary, 2008; Sharma and Rai, 2015; Singh and Sastry, 1988). Various types of plugs, including air plugs, stemming plugs, small inflatable balloons, plaster of Paris packings, cement concrete, and grouting are used as stemming material. Various researchers have compared the effects of stemming plugs and conventional stemming systems on rock fragmentation, and the shape of the muckpile. Laboratory experiments to test the effectiveness of stemming on rock fragmentation showed a 25% loss of energy when blast-holes were not stemmed (Zhang et al., 2020). Brinkmann (1990) established that inadequate stemming results in improper explosives confinement, causing about 50% loss in explosive energy due to premature venting through the collar of the blast-hole. Bartley (2002) reported that the incorporation of stemming plugs can increase gas retention for almost 50 milliseconds as compared to conventional stemming Eloranta ( 1994) used coarser stemming material with poured precast concrete, with a resulting reduction in the stemming ejection velocity. Brinkmann (1990) showed that the length of stemming plays an important role in blast design in the context of blowouts, as stemming blowouts result in excessive
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vibration and loss of confinement. He further concluded that a shorter stemming length and greater fines content are the most common reasons for stemming blowouts. The length of stemming was deemed critical by Chiapetta and Wyciskalla (2004), as a shorter length can result in blowouts and other collar-related problems. Jenkins and Floyd (2000) found that inadequate stemming results in more blowouts and boulders. Kojovic (2005) achieved an increased mill throughput and reduction in powder factor using fine aggregate stemming. Sazid, Saharan, and Singh (2012) concluded that stemming plugs control flyrock, improve fragmentation, achieve higher gas retention, eliminate boulder formation, and result in better utilization of explosive energy. Different stemming plugs apart from conventional drill cuttings are used in the industry. Yang et al., (2018) reported that the use of stemming plugs increased gas retention and reduced explosive consumption. Cevizci and Ozkahraman (2012) reported that plaster of Paris stemming plugs improved fragmentation. Numerical simulation for stemming analysis showed an increase in gas pressure with plaster stemming ( Cevizci, 2019). Ramulu (2012) reported an improvement in tunnel advance by incorporating stemming plugs. Choudhary and Rai (2013b) found that poor fragmentation, an irregular final wall, oversized boulders, high dozing cost, and loose muckpile were improved by using stemming plugs. Sharma and Rai (2015) reported that stemming plugs improved fragmentation, machine utilization, and muckpile throw. A study in Portugal (Neves et al., 2016) found improved blast performance with the application of stemming plugs. Stemming plug application resulted in better machinery utilization and fewer loads lost (Gomes-Sebastiao and de Graaf, 2017). In another application of stemming plugs, it was found that funnel-shaped stemming plugs controlled flyrock and increased the velocity of detonation (Bhaskar et al., 2019). Recent research has suggested further evaluation of stemming plugs based on human factors and economics (ur Rehman, Emad, and Khan, 2020). Even though stemming plugs have been utilized for almost three decades, their application is still new to many operations. The issue of technology adoption and consumer response is not limited to the mining industry, but has been observed and studied for different industries (Stavrakas, Papadelis, and Flamos, 2019). The literature review comprehensively establishes the effectiveness of stemming plugs in different settings and concludes that stemming plugs are effective in improving blast performance. The engineering aspects of using stemming plugs are widely reported (Ipmawati et al., Nainggolan, Wiyono, and Sunarya, 2018; Kabwe, 2018; Kabwe and Banda, 2018; Marinho et al., 2017; Saharan, Sazid, and Singh, 2017). An assessment of the impact of blast-hole design parameters on the total cost of opencast operations showed the significant importance of economic analysis (Bilim, Çelik, and Kekeç, 2020). However, no study has evaluated the economic impact of incorporating stemming plugs in blasting. The current work constitutes a novel contribution to the literature as it evaluates the economic impact of using stemming plugs at a full-scale quarry blasting operation. We evaluated the application of stemming at a large full-scale surface quarry in Punjab, Pakistan and carried out a critical economic analysis and estimate of the cost effects. The study also discusses the acceptability of various stemming plugs among workers.
Research methodology The project was carried out at the DG Khan Cement Company The Journal of the Southern African Institute of Mining and Metallurgy
Improving the environmental and economic aspects of blasting in surface mining Limited quarry located in District Chakwal, Pakistan. The quarry is located at 32°42’53.9”N 72°49’32.3”E (Google Earth, accessed 18 August 2015). The highest point at the quarry is 846 m above sea level with the lowest bench being developed at 777 m above sea level. The average limestone production from the quarry is about 16 000 t/day, and the plant production about 6700 t of cement per day. The blasting engineer designs the blast pattern which is followed by the drillers. The holes are charged using an interval charging technique and filled with stemming material. The blasting connections are made according to the design and the blast is initiated. Successful blasts yield fragmentation within the crusher’s size range, with flyrock and blast-induced vibrations under the permissible limits. Bench heights range from 11 to 14 m based on the strata and geology of the deposit. Figure 1 shows the plan view of the quarry. Two rows of holes are drilled in a staggered blast pattern. Experiments were conducted by adding stemming plugs to the existing blasting system. No other amendment to the blast design was made. The angle of the drill-holes was kept at 15 degrees relative to the vertical face. The blast-hole burden was 4 m and the spacing 5 m. Blast-hole depth was equal to bench height and sub-drill was 1 m for toe removal and smoother floor conditions. As per the blast design, 3 m of stemming was added in the top of the hole loaded with explosives. A combination of high explosives and blasting agent was used in each hole. The primer was dynamite (local name Wabox). After loading the primer the bottom charge was filled with boosters (local name Boosters) followed by ANFO blasting agent prepared at the site. All calculations were carried out following existing practice. When comparing different stemming devices with conventional stemming, all other blast design configurations were kept constant.
Figure 2—Stemming plug 1. (a) Plastic mould, (b) installation in blast-hole
Figure 3—Stemming plug 2. (a) borehole with air line; (b) inflated plug
Field evaluation The first type of stemming plug is the plastic moulded plug, as shown in Figure 2a and referred to as stemming plug 1. The application of stemming plug 1 required it to be inserted in the blast-hole after the explosives were loaded (Figure 2b). Once the plug had been inserted the rest of the hole was filled with conventional stemming material. The second type of stemming plug comprised inflatable rubber balls, which are representative of air plugs as discussed in the literature. The air plug is referred to as stemming plug 2. Conventional stemming material of 0.5 m length was added before stemming plug 2 was inserted in the blast-hole. The upper
Figure 1—Plan view of DG Khan cement quarry (Google Maps, accessed 18 August 2015) The Journal of the Southern African Institute of Mining and Metallurgy
Figure 4—Stemming plug 3 (mortar preparation)
portion of the blast-hole was again filled with conventional stemming. Figure 3a shows stemming plug 2 being inserted, with an air line for inflating the ball, and Figure 3b shows its final shape after installation in the blast-hole. For the third type of stemming plug, cement mortar was used (stemming plug 3). Similar to the installation of stemming plug 2, 0.5 m of conventional stemming was first added, followed by stemming plug 3 in the form of a slurry (Figure 4). The rest of the blast-hole was filled with conventional stemming material. The purpose of adding 0.5 m of conventional stemming before inserting stemming plug 2 was to avoid the formation of an air deck at top of the charged hole, whereas for stemming plug 3 0.5 m of conventional stemming was added to avoid contact of the mortar with the explosives. Both of these situations could have altered the explosive performance. Holes for conventional and experimental stemming plugs were initiated with a single initiator to avoid changes in rock mass conditions such as crack initiation, back-break, or anything emanating due to the adjacent blasting. VOLUME 121
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Improving the environmental and economic aspects of blasting in surface mining The fragmentation of the muckpile obtained after blasting was assessed by image analysis using Split Desktop software. This software processes images based on visible fragmentation. The most critical consideration was capturing the images of the muckpile. It was suggested that at least three images should be obtained with a high-resolution camera from different angles for a single analysis. The software allows the user to delineate the boundaries of the pile and particle sizes in the image for better processing. Sampling was carried out based on the directions accompanying the software. Three reference scales were placed at different distance for the best delineation. The images were then processed to determine the particle size distribution and ascertain the percentage of blast fragmentation. This work used three images per blast immediately after each blast; however, for better clarity and representation of the fragmentation size multiple images should be taken at regular intervals during muckpile loading. In this way, a comprehensive three-dimensional fragmentation result can be obtained. Future work should follow the suggested technique. A blast was categorized as a poor blast if it resulted in oversized boulders that required secondary blasting. Figure 5 shows the conventional stemming material.
Figure 5—Conventional stemming material
Table I
ummary of worker feedback regarding stemming plug S installation Aspect
Plug 1
Plug 2
Plug 3
Extra time required (minutes) Adaptability Failures in application Recovery time/failure No. of accessories Worker evaluation
1 Low 0 0 0 Positive
7 Moderate 30 8 1 Neutral
5 High 0 0 3 Negative
A total of six full-scale production blasts were performed. The ground was marked with spray paint to differentiate between two types of blasts while capturing images for fragmentation analysis. The reference mark was corrected using GPS. A muckpile image can be seen in Figure 6a. A standard football of 9-inch diameter was used as a reference. The fines cut was kept at 6.5 to 12 inches based on the type of geological formation. The fines cut was higher for loose strata as during loading and maneuvering it was expected to be reduced to 6 inches, the crusher passing size. The blasting crew was not familiar with the usage of stemming plugs, so on-site training and demonstrations were done. Worker feedback was also collected for seven different aspects to gauge human factors related to the stemming plugs application. A summary of the feedback is given in Table I. The feedback was gathered based on the relative perceptions of workers for each stemming plug.
Results and discussion Six full-scale blasts were performed – blasts 1, 3, and 5 with stemming plugs (types 1, 2, and 3 respectively), blasts 2, 4, and 6 using conventional stemming material. The results for all six blasts are summarized in Table II. The results were evaluated based on the percentage of material below the fines cut, the percentage above the passing size of the crusher, and the top size after the blast. Blasts 1 and 2 were performed at the same bench. Stemming plug 1 (plastic mould) was used in blast 1, and conventional stemming in blast 2. Blast 1 resulted in a D50 of 12 inches and D80 of 22 inches. The percentage of material below the fines cut was 27%, with no material greater than the crusher’s permissible limit. The largest boulder was 43.5 inches. Figure 6 shows the post-blast image and screenshot of the Split Desktop analysis. For blast-2 the D50 and D80 were 2 and 10 inches, with 69% below the fines cut. The percentage of the material above the acceptable crusher limit was 5.25%, and boulders requiring secondary blasting were also observed with a peak size of 143 inches. Figure 7 shows the post-blast image and screenshot from Split Desktop. The fragmentation results of blast 3 were compared with those of blast 4. Both these blasts were conducted at the same bench, with stemming plug 2 (inflated plug) employed in blast 3 and conventional stemming in blast 4. The D50 and D80 for blast 3 were 11.5 and 40 inches respectively, with 15% of the material below the fines cut. No material above the crusher’s input limit was observed, with the largest size being 63.41 inches. Figure 8 shows the post-blast image and Split Desktop screenshot. Blast 4 resulted in a D50 of 20 inches and D80 of 50 inches. The percentage of material below the fines cut was 20%, with 8% larger than the size of the crusher’s intake. The maximum particle size observed was 101 inches. Figure 9 shows the post-blast image and Split Desktop screenshot.
Table II
Summary of blast results from Split Desktop Blast no. 1 2 3 4 5 6
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Stemming
D50 (inches) D80 (inches) Fines cut size (inches) Percentage below fines cut Percentage above passing size Maximum size (inches)
Plug 1 No plug Plug 2 No plug Plug 3 No plug
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22 10 40 50 35 75
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27% 69% 15% 20% 28% 22%
0% 5.25% 0% 8.03% 0% 10.40%
43.5 243 63.41 101.2 46.21 92.05
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Figure 6—Blast 1. (a) Muckpile after the blast and (b) cumulative size distribution
Figure 7—Blast-2. (a) Muckpile after the blast and (b) cumulative size distribution
Figure 8—Blast 3. (a) Muckpile after the blast and (b) cumulative size distribution
Figure 9—Blast 4. (a) Muckpile after the blast and (b) cumulative size distribution The Journal of the Southern African Institute of Mining and Metallurgy
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Improving the environmental and economic aspects of blasting in surface mining Lastly, the blast 5 results were compared with those of blast 6. Stemming plug 3 (cement mortar/slurry) was used in blast 5, while blast 6 employed conventional stemming. The results for blast 5 show a D50 of 20 inches and D80 of 50 inches. The percentage of material b of 15 inches with D80 at 35 inches and 28% of the material less than the fines cut. There were no fragments beyond the permissible limit of the crusher, and the maximum size of the boulders was 46.21 inches. Figure 10 shows the post-blast image and screenshot from Split Desktop. For blast 6, the D50 and D80 were 25 and 75 inches, with 22% of the material below the fines cut and 10.4% larger than the required size limit of the crusher. The maximum size of the boulder was 92 inches. Figure 11 shows the post-blast image and screenshot from Split Desktop. The results for blasts 1 and 2 showed that the incorporation of stemming plugs improved gas retention, resulting in a more homogeneous particle size distribution. For a conventional blast (blast 2), the particle size distribution indicated excessive crushing and huge boulders. Although finer particle sizes are easy to haul and require less power during comminution, the accumulation of boulders is a problem as to their manoeuvrability, and secondary blasting is an economic and operational burden with higher safety risks. Comparison of both blasts suggested that blast 1 had a uniform explosives energy utilization and did not require secondary blasting. Stemming plug 1 appears therefore to be a useful product that should be incorporated in the blasting operation.
The worker feedback (Table I) suggests that the use of stemming plug 1 requires one extra minute, with minimal learning needed for its inclusion in blasting practice. Troubleshooting of stemming plug 1 was simple, as in case of a blockage no extra handling accessory was needed at the blast location. The incorporation of stemming plug 2 required 7 extra minutes for each blast-hole and extended the loading time significantly. In case of installation failures (which occurred 30 times during the tests), the total time of application is greatly increased, thereby delaying the overall blasting operation. The workers also had to make sure they carried the required inflation set-up to the blast site and used it correctly to avoid overfilling. However, with more usage and practice, total application time is expected to reduce. Thus, stemming plug 2 was less user-friendly than stemming plugs 1 and 3, requiring longer operational time and adding cognitive load on the blasting crew. Stemming plug 3 took five extra minutes to load each blast-hole and workers were required to master mortar mixing skills. Due to the quick setting time of cement, the mixture had to be prepared in small batches, which caused workers to dislike stemming plug 3. Of all the tested plugs, the operational efficiency of stemming plug 1 was the best. Another type of stemming plug (the self-inflating plug) might be a good alternative to stemming plug 1 as the installation procedure is similar to that for stemming plug 1. However selfinflating stemming plugs were not available for this work.
Figure 10—Blast 5. (a) Muckpile after the blast and (b) cumulative size distribution
Figure 11—Blast 6. (a) Muckpile after the blast and (b) cumulative size distribution
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Improving the environmental and economic aspects of blasting in surface mining A modification of stemming plug 3 exists in the form of ready-made cartridges that expand while being lowered down the blast-hole. No work on these types of stemming plug could be located in the literature. Such stemming plugs are available, although not in Pakistan. The authors therefore relied on the best available source for information. The performance of all stemming plugs tested was found to be acceptable and their incorporation improved the blast fragmentation. The comparison based on rock fragmentation suggested that stemming plug 3 resulted in the best rock breakage, stemming plug 1 the second-best blast results, whereas stemming plug 2 resulted in the least promising rock fragmentation. On the other hand, based on worker feedback, stemming plug 1 was most favoured and stemming plug 3 the least preferred.
Table V
Detailed costing of secondary blasting Machinery
Cost
Hours (or no. of units per kg) Total cost
Drilling machine $28.64 Haul truck $17.69 Front shovel $32.15 Dozer $25.30 Nonel $1.75 per unit Safety fuse $0.09 per metre Plain detonator $0.16 per unit Ammonium nitrate $0.47 per kg Blaster $2.48 per kg Grand total (US$) $493.11
6 4 4 2 15 2 1 20 14.28
$171.84 $70.76 $128.60 $50.60 $26.21 $0.19 $0.16 $9.40 $35.35
Table VI
Economic evaluation As indicated in Table II, all blasts that were performed without stemming plugs required secondary blasting.This material required additional hauling and loading. Based on the accounts of labour, explosives, and machinery cost, an economic analysis was conducted comparing blasts with stemming plugs and with conventional stemming. Table III indicates the cost of explosives and other accessories. Along with the added explosives cost for secondary blasting, hauling, and manoeuvring cost is also increased, as after the second blast all such operations need to be repeated. As per information received from the management of DG Cement, the operational cost of machinery was calculated on an hourly basis. For boulder blasting, drilling operations must be dedicated, therefore one drilling machine must spend six drilling hours. One complete shift of the blasting team was employed for boulder blasting. Expenses incurred for lubricants, filters, and spare parts were incorporated under the maintenance head. Fuel consumption per hour is given in Table IV, along with fuel rates received from the refinery. These rates are lower than the prevalent market price due to bulk contracting and the absence of a third party.
Table III
Explosives prices Product
Price (US$)
Nonel Detonating cord Safety fuse Plain detonator Ammonium nitrate Blaster Dynamite (Wabox)
$1.75 per unit $0.24 per metre $0.09 per metre $0.16 per unit $0.47 per kg $2.48 per kg $4.14 per kg
Cost of each plug Type of plug Plug 1 (plastic moulded) Plug 2 (air-plug) Plug 3 (cement mortar)
Unit price (US$)
Total no. of units used
Total cost (US$)
$2.6 $1.06 $3.19
47 80 56
$122.20 $ 84.80 $ 178.64
As the blasting crew and machine operators were on monthly wages and their extra hours did not have any effect on costs, that expense is not included in the economic analysis. The rest of the extra operational time is quantified in the form of US dollars. Table V indicates the time utilized by each item of machinery along with the cost and amount of explosives used in secondary blasting. Blast 2 resulted in only 5.25% oversized material. Applying the normalizing method, blasts 4 and 6, will cost around an additional $754.31 and $976.81 respectively. The cost of the stemming plug was an additional cost in the conventional system, but that helped in reducing the cost for secondary blasting. The costs per hour (or any other parameter) given to gauge the economics of operation are site-specific. Table VI lists the cost of each stemming plug and the number used. As all other parameters remained the same and no extra shift was needed for operators and labour when utilizing the stemming plugs, thus no extra installation cost was added. Furthermore, the shipping cost of plugs was considered. Stemming plug 1 is a rigid plastic material with no installation issues, therefore the numbers are the same as what was actually used. Stemming plug 2 suffers from bursting problems, and approximately 30 air-plugs were wasted so their cost was also added to the total number. This is an inherent issue with air-plugs. Similarly, based on the site conditions, six additional units were considered for cement mortar due to wastage during mixing, pouring, and sticking of mortar in the container.
Table IV
Fuel consumption and maintenance Machine Drilling machine Haul truck Front shovel Dozer
Fuel (l/h)
High-speed diesel (US$/h)
Maintenance cost/hour (US$/h)
Total cost per hour (US$)
40 25 45 35
$25.90 $16.19 $29.14 $22.67
$2.74 $1.5 $3.01 $2.63
$28.64 $17.69 $32.15 $25.30
The Journal of the Southern African Institute of Mining and Metallurgy
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Improving the environmental and economic aspects of blasting in surface mining Table VII
Cost comparison: stemming plugs vs secondary blasting Type of plug
Total cost (US$)
Total cost of secondary blasting (no plug) (US$)
Savings per blast (US$)
Annual savings (280 blasts)
Plug 1
$122.20
$493.11
$370.91
Plug 2
$84.80
$754.31
$669.51
$187 462.80
Plug 3
$178.64
$976.81
$798.17
$223 487.60
Table VII indicates the main crux of the economic analysis. The cost comparison shows that the use of plugs brings significant annual cost savings in all three cases. The greatest saving was achieved with stemming plug 3. However, the comparative difficulty of installation remains a challenge wirh this plug.
References Bartley, D. 2002. Blast fragmentation enhancement using MOCAP Vari-Stem holes plugs. http://www.varistem.com/PDF/DBAstudy1.pdf Bhaskar, A., Baranwal, A., Ranjan, P., Jena, T., Shkhar, M., and Chakraborty, D. 2019. Application of plastic funnel in blast hole to improve blasting efficiency of opencast coal mine at West Bokaro. Proceedings of the 2019 Coal Operators’ Conference, University of Wollongong, 18-20 February 2019 Aziz, N. and
Conclusions
Kininmonth, R. (eds), https://ro.uow.edu.au/coal/752
Three different blast-hole stemming systems were tested to investigate the effect on blasting efficiency (fragmentation) and economics. The fragmentation analysis was carried out based on the largest boulder size, D50, D80, cumulative size distribution, and percentage of fragmentation greater than the limit of 75 inches. Incorporation of stemming plugs in the blasting operation resulted in better rock fragmentation and production of fewer boulders. The economic benefits of using stemming plugs are considerable. Mineworkers’ perceptions were also considered in the comparison of different stemming plugs. The plastic moulded plug yielded the best fragmentation, and the cement mortar plug was found to be the most economical. The performance of the air plug was reasonable as regards both blast outcome and economics. A conventional stemming system results in 5–11% oversized boulders. Stemming plugs can reduce costs by US$0.13–0.2 million for a quarry conducting 280 annual blasts of 16 000 t/ day. This amounts to savings of US$ 0.03–0.045 per ton of limestone. Similar studies could be carried out in different geological environments to further evaluate the impact of stemming plugs on the economics of drilling and blasting at other operations. Aggregates are widely used as stemming material and their comparison with commercial stemming systems (as tested in this work) could also be considered for future work.
Acknowledgements
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Conflict of interest statement
and muckpile shape parameters. International Journal of Mining and
This paper might positively influence Vari-stem stemming plugs or Pagel Chemical. Stemming plug 2 do not have a trademark name or specific brand that can be influenced. Other than that, on behalf of the other author, the corresponding author states that there is no conflict of interest of which we are aware at this point. JULY 2021
Bilim, N., Çelik, A., and Kekeç, B. 2020. Assessment of the effect of blasthole design
Cevizci, H. and Ozkahraman, H.T. 2012. The effect of blast hole stemming length to
The authors would like to acknowledge the Vari-Stem company for providing free stemming plugs. Special thanks to Pagel Chemicals for providing chemicals/additives for the cement mortar plugs, and Mr Tanzeel Jabbar, and Ahsan Saleem for providing air-plugs. This work would not have been possible without the support of DG. Khan Cement Company, Mr Rafi Ur Rehman, Mr Hidayat Shah, Mr Reehan Shoukat, Mr Sajid Ur Rehman, and their blasting team. Special thanks to the reviewers for their useful comments on improving our work.
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SAIMM Advertising Opportunities Committed to the minerals and metals economy in Southern Africa
The Southern African Institute of Mining and Metallurgy
VOLUME 120
NO. 1
JANUARY 2020
VOLUME 120
NO. 2
FEBRUARY 2020
VOLUME 120
NO. 5
MAY 2020
Minerals
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