Saimm 202404 apr

The beginning of 2024 has been mostly focused on rejuvenating the student chapters at universities. These play a vital role in creating a pipeline of future leaders in the mining and minerals industry.

the YPC is able to pick up on themes impacting the young professionals and strategically working with the greater SAIMM and industry to find initiatives that will inform, educate and best address ongoing challenges and create awareness. Some of these themes are high unemployment rates, university funding, retaining of young professionals in the mining and minerals industry,

professionals, especially females, in the mining and minerals industry. As the YPC we aim to assist through our mentorship programme, which allows young professionals to access diverse skilled and experienced mentors locally and globally, who are SAIMM members and part of our Company Affiliates, thereby providing assistance to our young professionals in addressing challenges they may encounter in the industry, as well as in their personal lives.

As the SAIMM is a voluntary organization, it is often difficult to find self-driven and committed volunteers to keep making a positive impact in our industry, and with young professionals this is also evident. We can only strive and ensure the industry grows and is sustainable through active collaboration between corporate organizations, universities and individuals. As such this is a call to arms – let us all come to the table to do our part for the greater good of the mining and minerals industry and generations of leaders yet to

Signing out, The Great Banda.

KONDWANI BANDA

The Southern African Institute of Mining and Metallurgy

OFFICE BEARERS AND COUNCIL FOR THE 2023/2024 SESSION

Honorary President

Nolitha Fakude

President, Minerals Council South Africa

Honorary Vice Presidents

Gwede Mantashe

Minister of Mineral Resources and Energy, South Africa

Ebrahim Patel

Minister of Trade, Industry and Competition, South Africa

Blade Nzimande

Minister of Higher Education, Science and Technology, South Africa

President

W.C. Joughin

President Elect

E. Matinde

Senior Vice President

G.R. Lane

Junior Vice President

T.M. Mmola

Incoming Junior Vice President

M.H. Solomon

Immediate Past President

Z. Botha

Honorary Treasurer

E. Matinde

Ordinary Members on Council

W. Broodryk M.C. Munroe

Z. Fakhraei S. Naik

R.M.S. Falcon (by invitation) G. Njowa

B. Genc

S.J. Ntsoelengoe

K.M. Letsoalo S.M. Rupprecht

S.B. Madolo

A.T. van Zyl

F.T. Manyanga E.J. Walls

K. Mosebi

Co-opted Council Members

M.A. Mello

Past Presidents Serving on Council

N.A. Barcza C. Musingwini

R.D. Beck S. Ndlovu

J.R. Dixon J.L. Porter

V.G. Duke M.H. Rogers

I.J. Geldenhuys D.A.J. Ross-Watt

R.T. Jones G.L. Smith

A.S. Macfarlane W.H. van Niekerk

G.R. Lane – TP Mining Chairperson

Z. Botha – TP Metallurgy Chairperson

K.W. Banda – YPC Chairperson

S. Nyoni – YPC Vice Chairperson

Branch Chairpersons

Botswana Vacant

DRC Not active

Johannesburg N. Rampersad

Limpopo S. Zulu

Namibia Vacant

Northern Cape I. Tlhapi

North West I. Tshabalala

Pretoria Vacant

Western Cape A.B. Nesbitt

Zambia J.P.C. Mutambo (Interim Chairperson)

Zimbabwe Vacant

Zululand C.W. Mienie

PAST PRESIDENTS

*Deceased

* W. Bettel (1894–1895)

* A.F. Crosse (1895–1896)

* W.R. Feldtmann (1896–1897)

* C. Butters (1897–1898)

* J. Loevy (1898–1899)

* J.R. Williams (1899–1903)

* S.H. Pearce (1903–1904)

* W.A. Caldecott (1904–1905)

* W. Cullen (1905–1906)

* E.H. Johnson (1906–1907)

* J. Yates (1907–1908)

* R.G. Bevington (1908–1909)

* A. McA. Johnston (1909–1910)

* J. Moir (1910–1911)

* C.B. Saner (1911–1912)

* W.R. Dowling (1912–1913)

* A. Richardson (1913–1914)

* G.H. Stanley (1914–1915)

* J.E. Thomas (1915–1916)

* J.A. Wilkinson (1916–1917)

* G. Hildick-Smith (1917–1918)

* H.S. Meyer (1918–1919)

* J. Gray (1919–1920)

* J. Chilton (1920–1921)

* F. Wartenweiler (1921–1922)

* G.A. Watermeyer (1922–1923)

* F.W. Watson (1923–1924)

* C.J. Gray (1924–1925)

* H.A. White (1925–1926)

* H.R. Adam (1926–1927)

* Sir Robert Kotze (1927–1928)

* J.A. Woodburn (1928–1929)

* H. Pirow (1929–1930)

* J. Henderson (1930–1931)

* A. King (1931–1932)

* V. Nimmo-Dewar (1932–1933)

* P.N. Lategan (1933–1934)

* E.C. Ranson (1934–1935)

* R.A. Flugge-De-Smidt (1935–1936)

* T.K. Prentice (1936–1937)

* R.S.G. Stokes (1937–1938)

* P.E. Hall (1938–1939)

* E.H.A. Joseph (1939–1940)

* J.H. Dobson (1940–1941)

* Theo Meyer (1941–1942)

* John V. Muller (1942–1943)

* C. Biccard Jeppe (1943–1944)

* P.J. Louis Bok (1944–1945)

* J.T. McIntyre (1945–1946)

* M. Falcon (1946–1947)

* A. Clemens (1947–1948)

* F.G. Hill (1948–1949)

* O.A.E. Jackson (1949–1950)

* W.E. Gooday (1950–1951)

* C.J. Irving (1951–1952)

* D.D. Stitt (1952–1953)

* M.C.G. Meyer (1953–1954)

* L.A. Bushell (1954–1955)

* H. Britten (1955–1956)

* Wm. Bleloch (1956–1957)

* H. Simon (1957–1958)

* M. Barcza (1958–1959)

* R.J. Adamson (1959–1960)

* W.S. Findlay (1960–1961)

* D.G. Maxwell (1961–1962)

* J. de V. Lambrechts (1962–1963)

* J.F. Reid (1963–1964)

* D.M. Jamieson (1964–1965)

* H.E. Cross (1965–1966)

* D. Gordon Jones (1966–1967)

* P. Lambooy (1967–1968)

* R.C.J. Goode (1968–1969)

* J.K.E. Douglas (1969–1970)

* V.C. Robinson (1970–1971)

* D.D. Howat (1971–1972)

* J.P. Hugo (1972–1973)

* P.W.J. van Rensburg (1973–1974)

* R.P. Plewman (1974–1975)

* R.E. Robinson (1975–1976)

* M.D.G. Salamon (1976–1977)

* P.A. Von Wielligh (1977–1978)

* M.G. Atmore (1978–1979)

* D.A. Viljoen (1979–1980)

* P.R. Jochens (1980–1981)

* G.Y. Nisbet (1981–1982)

A.N. Brown (1982–1983)

* R.P. King (1983–1984)

J.D. Austin (1984–1985)

* H.E. James (1985–1986)

H. Wagner (1986–1987)

* B.C. Alberts (1987–1988)

* C.E. Fivaz (1988–1989)

* O.K.H. Steffen (1989–1990)

* H.G. Mosenthal (1990–1991)

R.D. Beck (1991–1992)

* J.P. Hoffman (1992–1993)

* H. Scott-Russell (1993–1994)

J.A. Cruise (1994–1995)

D.A.J. Ross-Watt (1995–1996)

N.A. Barcza (1996–1997)

* R.P. Mohring (1997–1998)

J.R. Dixon (1998–1999)

M.H. Rogers (1999–2000)

L.A. Cramer (2000–2001)

* A.A.B. Douglas (2001–2002)

S.J. Ramokgopa (2002-2003)

T.R. Stacey (2003–2004)

F.M.G. Egerton (2004–2005)

W.H. van Niekerk (2005–2006)

R.P.H. Willis (2006–2007)

R.G.B. Pickering (2007–2008)

A.M. Garbers-Craig (2008–2009)

J.C. Ngoma (2009–2010)

G.V.R. Landman (2010–2011)

J.N. van der Merwe (2011–2012)

G.L. Smith (2012–2013)

M. Dworzanowski (2013–2014)

J.L. Porter (2014–2015)

R.T. Jones (2015–2016)

C. Musingwini (2016–2017)

S. Ndlovu (2017–2018)

A.S. Macfarlane (2018–2019)

M.I. Mthenjane (2019–2020)

V.G. Duke (2020–2021)

I.J. Geldenhuys (2021–2022)

Z. Botha (2022-2023)

Editorial Board

S.O. Bada

R.D. Beck

P. den Hoed

I.M. Dikgwatlhe

M. Erwee

B. Genc

R Hassanalizadeh

R.T. Jones

W.C. Joughin

A.J. Kinghorn

D.E.P. Klenam

J. Lake

H.M. Lodewijks

D.F. Malan

C. Musingwini

S. Ndlovu

P.N. Neingo

S.S. Nyoni

M. Phasha

P. Pistorius

P. Radcliffe

N. Rampersad

Q.G. Reynolds

I. Robinson

S.M. Rupprecht

K.C. Sole

T.R. Stacey

D. Vogt

F. Uahengo

International Advisory Board members

R. Dimitrakopolous

R. Mitra

A.J.S. Spearing

E. Topal

D. Tudor

Editor /Chairperson of the Editorial Board

R.M.S. Falcon

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Contents

Journal Comment: From the Desktop to the Coal Face by J.N. van der Merwe

Presidential Address: Quest for zero harm in South African deep gold mines by W.C. Joughin

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Copyright© 2024 by The Southern African Institute of Mining and Metallurgy. All rights reserved. Multiple copying of the contents of this publication or parts thereof without permission is in breach of copyright, but permission is hereby given for the copying of titles and abstracts of papers and names of authors. Permission to copy illustrations and short extracts from the text of individual contributions is usually given upon written application to the Institute, provided that the source (and where appropriate, the copyright) is acknowledged. Apart from any fair dealing for the purposes of review or criticism under The Copyright Act no. 98, 1978, Section 12, of the Republic of South Africa, a single copy of an article may be supplied by a library for the purposes of research or private study. No part of this publication may be reproduced, stored in a retrieval system, or transmitted in any form or by any means without the prior permission of the publishers. Multiple copying of the contents of the publication without permission is always illegal. U.S. Copyright Law applicable to users In the U.S.A. The appearance of the statement of copyright at the bottom of the first page of an article appearing in this journal indicates that the copyright holder consents to the making of copies of the article for personal or internal use. This consent is given on condition that the copier pays the stated fee for each copy of a paper beyond that permitted by Section 107 or 108 of the U.S. Copyright Law. The fee is to be paid through the Copyright Clearance Center, Inc., Operations Center, P.O. Box 765, Schenectady, New York 12301, U.S.A. This consent does not extend to other kinds of copying, such as copying for general distribution, for advertising or promotional purposes, for creating new collective works, or for resale.

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ISSN 2225-6253 (print) . ISSN 2411-9717 (online)

PROFESSIONAL TECHNICAL AND SCIENTIFIC PAPERS

Technology and growth in the South African mining industry: An assessment of critical success factors and challenges by V.W. Lumadi and S. Nyasha

This study determines the factors that influence acceptance of technology solutions in the mining industry in South Africa. The study encourages mining companies to implement usage of technology through regular engagement with the Minerals Council South Africa.

Evaluation of different surface characteristics and mineral grain size in the estimation of rock strength using the Schmidt hammer by K. Karaman

This study investigated the effect of surface roughness on Schmidt rebound hardness (RL) and showed that estimated unconfined compressive strength values are not statistically significant if Schmidt rebound tests are not performed on similar surfaces. Hence, roughness of the surface should be eliminated if variations are shown in the surface rebound hardness.

Communication constraints in the safety system on South African mines and implications for the exercise of the Right to Refuse Dangerous Work by N. Coulson and P.F. Stewart ......................................................................................

This follow-up article presents production supervisor perspectives on the Right to Refuse Dangerous Work (RRDW). The study demonstrates how, for mineworkers, the two distinct communication subsystems constrain, rather than facilitate, implementation of the RRDW.

Triboelectric characteristics and separation of magnesite and quartz by Z. Zhang, Y. Xu, H. Wang, J. Shi, J. Niu, and Z. Zhang

Tribo-electrostatic separation is a promising method for effectively utilising low-grade magnesite resources by removing quartz and improving the grade of magnesite. To determine the charge-to-mass ratios of pure magnesite and quartz, a triboelectric measurement system was employed, and a laboratory tribo-electrostatic separation system was used to separate low-grade magnesite. These experimental results demonstrate the effectiveness and potential of tribo-electrostatic separation in removing quartz and upgrading magnesite.

Can preconcentration of cassiterite from its pegmatite ore reduce processing costs and improve operational sustainability? by H. Simonsen, J.H. Potgieter, K.J. Nyembwe, and A. Chuma

Different concentration techniques were evaluated for preconcentration of a mineral ore at a coarser size to avoid energy and resource wastage. Three concentration techniques: dense media, shaking table, and flotation, coupled with characterization analysis, were used to assess the concentration response. The results suggest that use of dense media separation as a rough preconcentration method prior to further grinding, and the utilization of a more advanced concentration technique for mineral recovery and upgrade, constitute a successful approach to improve process economics.

Open-pit post-blast dust cloud lightning by N.C. Steenkamp

Lightning has been observed in dust clouds following open-pit blasting. The occurrence of this phenomenon is related to the physical composition and characterization of the mineral fines that comprise the bulk of the dust cloud. Silicate minerals, which are susceptible to fine fracturing during blasting, generate the initial charge. The ideal prevailing meteorological conditions need to be windless to minimize the effect of particle dissipation and have sufficient moisture to enhance the potential of generating lightning.

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Leaching of gold ore using creatine monohydrate by S.M.S. Ganji and K. Barani ...................................................................................... 213

Creatine was used as a lixiviant for gold dissolution from gold ore. The effects of creatine concentration, temperature, leach time, and pH were examined for their influence on extent of gold dissolution. A comparison between creatine and cyanide leaching showed that 90% gold dissolution was achieved at a cyanide concentration of 300 g/t, which is three times higher than the creatine concentration under optimal conditions.

Investigation of glycine leaching for gold extraction from Witwatersrand gold mine tailings with permanganate pre-treatment by A. Tapfuma, G. Akdogan, M. Tadie

In this study, the feasibility of extraction of gold from a Witwatersrand gold tailings using glycine was investigated using a two-level full factorial design. Although the solid–liquid ratio used in this study was below the industrial norm, it provides a starting point for investigating the applicability of this technology. The work demonstrates that potassium permanganate pre-treatment prior to glycine leaching of low-grade secondary gold resources, such as tailings, can be beneficial.

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Copper Cobalt

Journal Comment From the Desktop to the Coal Face

Ithought to write this note as a researcher, author, and collaborator, and who, over a few decades has seen and participated in the development of several new equations, procedures, and processes on a number of topics: in my case, obviously in the field of rock engineering. These are invariably published in established journals or presented at conferences of some or other form, some more formal than others.

One of my worst nightmares is that practitioners in the field will see new equations or procedures and simply apply them without having gone through the proper preparatory steps and investigations.

What is often not realized is that a journal or conference paper is no more than the briefest of summaries of work that was often done over a long time period, sometimes by numerous researchers. The background preconditions and limitations are often implied—not even always, but not explicitly stated in paper publications.

A paper can only be of a certain length and there is simply not space to include everything. Just a simple example is a paper based on a doctoral thesis. The thesis itself may vary in length from 200 to 400 pages, but the published paper based on it, not more than 10 or 12 pages. It is totally impractical for everything that was stated or found in an investigation report to make its way into a paper for general consumption: only the highlights of the outcomes can be published in the popular domain.

There may be errors in the paper, in the form of misprints or other errors, that were simply not identified during the refereeing process. Referees, as well as authors, are just people and mistakes can happen.

It is even possible, albeit not likely, for fundamental errors in thinking in the original reporting or experimental phase that, at the time of writing or refereeing, were overshadowed by other more convincingsounding arguments. There could be errors in the data and, often, the data may be incomplete.

Always remember that any new equation derived from experimentation or data analysis for anything is no more than a theory. In essence, it is a description of what the developer saw—or believed they saw—in the data, be it in the form of numbers or any other type of observation. But as long as it is on paper, it is no more than a theory.

It is against this background that I want to stress the point that any new process or design equation cannot simply be put into practice. No. This should never be done without careful consideration and evaluation of the local context.

Theory should not just blindly be put into practice. It is not a step from theory to practice: it is a process, and the process should be carefully planned and managed.

We have to progress and new processes have to be applied. Without incremental improvement, nothing will change. If we prefer to wait for perfection, it will never come. We would still have been mining with hammers and chisels.

Just one example of an error in the application from theory to practise is a lesson from a tragic minecollapse disaster half a century ago that we may have missed or simply not heeded, and which can serve as illustration. At the time, the mine was in need of increased coal production, but the predevelopment to unlock new reserves was not yet in place. Consideration was then given to mining the coal in the roof in areas that had previously been mined at a nominal height of 3 m: this in a 6 m thick coal seam. The coal was already exposed, the infrastructure was in place. The solution made perfect sense.

Copper Cobalt

Journal Comment (continued)

There was no formal design procedure for coal pillars at the time, but warnings were issued that increasing the pillar height would weaken the pillars and could possibly result in failure. It was not known at the time by how much the pillars would be weakened, if at all.

The mine did not simply go ahead and increase the pillar height. They took the responsible precaution of embarking on an experiment. The height was increased in an experimental area and visually observed. After about three months, nothing had changed and no collapse occurred. The experiment was regarded as successful and so the coal roof was mined in several areas. Then the collapse occurred.

So what went wrong?

Firstly, the effects of time on pillar scaling and subsequent reduction in pillar width were not known or appreciated. In retrospect, and only in retrospect, three months was too short a time for the experiment, particularly as no formal measurements were reported, if at all performed.

Secondly, the experimental area was small and surrounded by solid coal on all sides. The effect of scale and increased pillar load due to the greater expanse of mining was not known at the time.

Thirdly, no follow-up monitoring or continued observation of the experimental area was done.

This is just a very brief description of hasty application of what, at the time, could be considered as a theory: Increased height would not substantially reduce pillar strength. This was the theory, believed to have been backed up by experiment.

Several other contributary events took place before the major collapse occurred. The brief description here is just to illustrate the point that any change in an existing situation has to be carefully implemented and properly analysed.

It also serves to illustrate another vitally important omission in our current mining operations, a point that has been stressed so often by researchers like Prof Francois Malan of the University of Pretoria; namely, that we do not do anything remotely close to enough measurements in our mining operations.

So, my plea is this: don’t just scan a paper until you find an equation that suits your needs and go ahead with implementation. Study the paper, get more information, consider the background against which the paper was written, and then plan implementation very carefully and slowly.

Perform measurements. Have a suitable control area for comparison, measure and continue monitoring, and especially continue with the observations in the control area. Adapt if necessary.

Also bear in mind that equations reflect the ideal situation. In the case of bord-and-pillar mining, for instance, this means assuming a constant mining height and perfectly straight roadways. No off-line development, equipment in perfection condition, so perfectly constant and equal traction on cat tracks, etc. No errors in the placement or observation of survey pegs. No simple human error.

But we all know that reality is different.

Therefore, there also has to be some form of allowance for real mining practice. There cannot be universal guidelines for this because the degree of deviation will be different for different mines, and even different sections on a particular mine. The practical allowance for error will depend on the on-site extent of deviations.

Build up confidence, think, observe, think again before you do final implementation—and only then go ahead. And even then, continue monitoring. We have to progress, but we have to do so very carefully.

J.N. van der Merwe

President’s Corner

OQuest for zero harm in South African deep gold mines

n 5 April, 2024, I had the privilege of attending the Fall-of-Ground Action Plan (FOGAP) Day of Learning — an event that marked a significant milestone in the pursuit of mine safety. Hosted by the Minerals Council - South Africa (MCSA) in partnership with the Association of Mine Managers of South Africa (AMMSA), the South African Colliery Managers’ Association (SACMA), and the South African National Institute of Rock Engineering (SANIRE), the event showcased groundbreaking strategies to combat fall-of-ground incidents, which remains one of the most significant hazards in mining operations.

FOGAP was developed through the collaborative efforts of the MCSA Rock Engineering Technical Committee (RETC) and SANIRE and is aimed at eliminating fall-of-ground fatalities. The focal points of the day’s discussions were on innovative support systems and improved workplace conditions.

I was particularly impressed by the adoption of permanent mesh support in narrow tabular stopes. This innovation involves securing high tensile steel mesh to the hanging wall using an array of supports including rockbolts, timber props, timber packs, and mechanical props (Figure 1). While rockburst resistant supports, such as rapid yielding hydraulic props and engineered timber props, have been employed for many years, they have not entirely mitigated the risk of fatal injuries during rockbursts, as the rock tends to fail between these supports. Integrating the high tensile mesh into the support system could significantly reduce the risk of injuries and fatalities.

The challenge of installing mesh in narrow tabular stopes has been a topic of concern since I began my career in the 1990s. Due to the manual labour required, often in hot and humid conditions, and the complexity of the installation process, the task was deemed too difficult. Mesh installation is common in tunnels where stress and rockburst damage is a concern; where the mesh is fastened to the rock using rockbolts and sometimes high tensile steel cable lacing.

The landscape began to change around 2012 when in-stope rockbolting became widely adopted, allowing for support to be placed closer to the stope face and increasing the density of support. Coupled with the development of high tensile mesh, the installation process has become slightly more manageable. However, it’s important to acknowledge that poorly installed mesh can exacerbate safety risks, making the development and trial of installation methods critical. Once proven effective, it is imperative to train all stope teams in these methods. Not surprisingly, convincing labourers of the benefits of this additional effort is no small feat. It requires effective communication and demonstration—rolling out such significant changes is a considerable achievement that could substantially contribute to the goal of zero harm.

Another noteworthy advancement is the introduction of LED lighting, which has drastically improved the illumination of underground stopes. These low-power lighting systems can be installed swiftly at the start of a shift and are removed at the end to avoid damage from blasting and scraper cleaning. The enhanced visibility allows for easier identification and remediation of rockfall hazards, further bolstering workplace safety.

Figure 2 illustrates the annual number of seismic-related fatalities in the industry since 1984, revealing a pronounced decline from 147 fatalities in 1990 to none in 2022, which at first glance indicates a dramatic enhancement in workplace safety. However, a more nuanced view is warranted. A substantial portion of rockburst incidents occur in gold mines, often deeper than 2000 metres. Concurrently, there has been a notable decrease in both gold production and workforce within the gold mining sector (Figures 3 and 4). South Africa’s gold production peaked in 1970 at one million kilograms of gold, the highest recorded by any country in a single year. This figure has since steadily declined to just 100 000 kilograms by the end of 2023. Similarly, the workforce in gold mines has diminished from nearly 400 000 individuals in 1995 to 93 589 by the end of 2023.

The question then arises: Does the decline in fatalities accurately reflect enhanced safety measures, or is it merely a consequence of the contraction of South Africa’s gold mining industry? For a more reliable evaluation of safety performance,

Figure 1-High tensile steel mesh installed in stopes (left – courtesy: Sibanye Stillwater; right – courtesy: Harmony)

President’s Corner (continued)

(Seismic)

(https://www.ceicdata.com/en/indicator/south-africa/gold-production)

Yearly Employment Numbers in Gold Mines

it is imperative to control for certain variables. Historically, the MCSA reported fatality rates per 1000 workers, but this metric was later changed to reporting the absolute number of fatalities to emphasize the human cost and foster greater empathy. While the latter approach has its merits, it complicates the task of objectively assessing improvements in safety. In Figure 5, I have normalized the fatality figures to account for every 100 000 workers and per 100 tons of gold produced. These reference points are chosen for their approximation to current gold production and employment figures. Furthermore, this method aligns with the World Health Organization’s practice of standardising death rates per 100 000 individuals, enabling comparisons with other mortality causes. Although this normalization could benefit from more specific data, it offers a more reliable preliminary analysis of safety advancements. The data indicates a gradual decline in fatality rates from 1990 to 2008, followed by a noticeable shift between 2008 and 2010. Presently, the rate hovers around five fatalities per 100 000 workers, with 2017 and 2020 being particularly challenging years, and 2022 being remarkably safe with no reported fatalities.

Figure 2-Fatalities caused by seismic events in South African Gold Mines from 1980 to 2023 (from Minerals Council and SAMRASS database)

Figure 4-Yearly employment numbers in South African Gold Mines (Stats SA and DMRE)

Figure 3-Gold production in kilograms

President’s Corner (continued)

This analysis also reveals that year-on-year comparisons may not yield substantive insights, whereas long-term trends are more telling. The observed improvements are likely the result of various interventions, both technological and managerial. The MCSA, in collaboration with the mining industry, has launched several safety initiatives through the Mine Occupational Safety and Health (MOSH) programmes and, more recently, the FOGAP initiative. The future will determine if these efforts will lead to further advancements in safety.

In light of the reduction of fatality numbers, both due to contraction of the industry and safety improvements, it is necessary to shift focus towards tracking high-potential incidents to monitor safety performance. Large seismic events and major rockburst damage continue to be prevalent in deep-level gold mining operations, posing a persistent challenge in addressing the risk. Rockburst risk management requires a comprehensive strategy that encompasses multiple components. Each of these components makes a small contribution to reducing the risk, but when all are implemented, it results in a significant reduction in risk. These include optimising mining layouts and sequences to reduce the frequency of large seismic events and to redirect them away from active mining areas, modifying the characteristics of the rock (preconditioning) to reduce the likelihood of bursting, improving support to reduce the likelihood of damage caused by large seismic events and reducing the exposure of workers to rockburst damage.

Guidelines for optimising mining layouts and sequences have been developed over many years, together with numerical modelling tools to analyse stress changes caused by mining. However, the geological structures (faults and dykes) and rock mass characteristics are complex, and therefore the rock mass response will vary in different parts of the mine and between mines. It is important to gather as much information on the geology as possible and to develop structural models to better anticipate the rock mass response. Seismic monitoring systems are essential tools for measuring the rock mass response and evaluating the implementation of rockburst risk management strategies throughout the mine.

Unfortunately, the technology to predict the location and time of a large seismic event has not been developed, despite significant efforts in this regard. If it were possible to do this reliably, it would enable the removal of workers from the danger area before a significant rockburst. At present seismic monitoring systems can reliably measure changes in seismic behaviour, which may provide an indication of an increase in seismic hazard in an area of the mine. However, seismic monitoring systems have contributed immensely to understanding rock mass behaviour and reducing uncertainty, and continue to play a key role in rockburst risk management.

Through the FOGAP programme, the MCSA has initiated a seismic research project. The first phase of the project is in progress, which entails the review of current rockburst risk management practice, locally and internationally. The objective of this phase is to identify best practices and gaps which need to be addressed. While rockburst risk management practices were first developed in South Africa, international mines are going deeper and do experience high stress and high levels of seismic hazard, so there is an opportunity to learn from different approaches applied elsewhere. The second phase will involve the investigation of alternative methods for short term seismic hazard analysis using machine learning and Bayesian statistical methods applied to seismic monitoring data. Ultimately this will lead towards much improved guidelines on rockburst risk management, and training of rock engineers to implement these guidelines.

It is imperative to strive for zero harm and to implement practicable solutions to reduce the risk. Risk is the product of probability and consequence. Since probabilities are expressed as fractions, a probability of zero equates to one divided by infinity, which is mathematically unattainable. However, safety interventions may make the probability of occurrence so low that the resulting injuries and fatalities within the population of mine workers would be so infrequent as to result in effectively zero harm.

W.C. Joughin President, SAIMM

Figure 5-Normalized fatality rates

Affiliation:

1University of South Africa – Graduate School of Business Leadership, South Africa

Correspondence to:

V.W. Lumadi

S. Nyasha

Email: walter.lumadi@gmail.com sheillanyasha@gmail.com

Dates:

Received: 24 Aug. 2022

Revised: 14 Oct. 2023

Accepted: 14 Jan. 2024

Published: April 2024

How to cite:

Lumadi, V.W. and Nyasha, S. 2024. Technology and growth in the South African mining industry: An assessment of critical success factors and challenges. Journal of the Southern African Institute of Mining and Metallurgy, vol. 124, no. 4, pp. 163–172

DOI ID:

http://dx.doi.org/10.17159/24119717/2287/2024

ORCID:

S. Nyasha

http://orcid.org/0000-0003-2930-094X

Technology and growth in the South African mining industry: An assessment of critical success factors and challenges

by V.W. Lumadi1 and S. Nyasha1

Abstract

The mining industry has yet to fully accept and embrace the strategic role of technology and innovation in successful business planning and execution. In this study, we empirically determine the factors that influence acceptance of technology solutions in the mining industry in South Africa, assess critical success factors influencing adoption of advanced technology in the South African mining industry, and explore challenges faced by the mining industry in South Africa in adopting advanced technology in its operations. Thematic content analysis is employed in the data analysis. Results show that critical success factors to the adoption of mining technology include learning culture, knowledge sharing, high labour costs, and behaviour of other firms in the industry; commonly identified challenges include inadequate engagement with external stakeholders, uncertainties, the cyclical nature of the sector, and high risk related to the adoption of unproven technology and performance systems focused on volumetric production. The study encourages mining companies to implement usage of technology. This can be properly done through regular engagement with the Minerals Council South Africa.

Keywords technology, mining, South Africa

Introduction

Historically, South Africa’s mining industry has been at the heart of the economy’s development – given the country’s competitive position as one of the most naturally resource-rich nations in the world (Antin, 2013). The South African mining industry and its sustainability are threatened by many global and local challenges. These challenges vary from high production costs, low profitability, labour unrest, to increasing demands by government (Lane et al., 2015).

Today, most deep-level underground mines are aging, with travel times to the face sometimes reaching an hour or more (Minerals Council South Africa (Minerals Council), 2021). Consequently, with increasing depth and distance from the shaft, actual drill time at the workface has contracted, accounting for greater health and safety challenges, shrinking production, and contributing to burgeoning costs (Minerals Council, 2021). Modernization of mines by improving on current technology will help to improve safety and health, facilitating the quest for zero harm. It will also contribute to increased skills development, employment, exports and revenue, and the knock-on effect on local communities (Minerals Council, 2021). The South African mining industry continues to battle economically, even though a minor recovery of commodity prices and higher production numbers have been seen. The last ten years have seen a continued sustained rise of operational expenses and capital development costs, and rapid decline in productivity (Bryant, 2015). This trend is not sustainable, especially against other key structural challenges, such as declining grades and more stranded assets.

Financial indicators are evidence of the ongoing weak trends; however, investors still expect higher returns and profitability margins. The need to focus on innovations is essential to deal with current operational challenges and convert them into opportunities (Guzek, 2015). Urgent adoption is thus a necessity to curb these deteriorating and unsustainable trends for the industry and for the economies that are thereby affected. Mining companies are required to provide a safe working environment for the health and safety of the employees, so it is of great importance that these factors are not viewed in isolation. Rising productivity, alongside exploration, is the principal means by which mining can combat resource depletion (Humphreys, 2018). Modernization of mining has emerged as an enabler in redressing the decline of this critical industry, as it is said to have transformative effects in achieving safer work environments and improved efficiencies and productivity (Mavroudis, 2017). Organized labour unions do not see things in the same light, as they express concerns regarding the possibility of job losses as a result of new innovations.

Technology and growth in the South African mining industry

The National Union of Mineworkers has been at the forefront of labour representation in the mining sector and has often used mass strikes to reach its goals (Antin, 2013). Contrary to this belief, modernization will contribute to increased skills development, employment, exports, and revenue (Minerals Council, 2021).

South African mining companies are, by definition, innovative, but there is significant room for the industry to more readily embrace the fourth industrial revolution (4IR) and innovation (PWC, 2010). Modernization necessitates mining, and the South African mining industry necessitates modernization as well. Over the last decade, South Africa's multi-factor productivity has fallen by 7.6%, and mining cost inflation has been 2% to 3% higher than general inflation, resulting in two-thirds of our output being in the upper half of the global mining cost curve (Department of Mineral Resources and Energy (DMRE), 2021). Mining output fell by 10% and mineral sales fell by 11%, highlighting the importance of modernization, not only for the mining industry's longevity, but also for social good (DMRE, 2021).

Given the highlighted importance of the mining sector in the South African economy and current challenges faced by the sector, on the one hand, and the slow uptake of technological advancement in South African mines, on the other hand, it has become critical that a study be undertaken to explore determinants of the adoption of technology in this sector. Although the study acknowledges previous studies on technology in the South African mining sector (Dayo-Olupona, 2020; Ntsoelengoe; 2019), it also notes gaps that were not filled: the former looked at emerging technology selection for only surface mines, without giving coverage to all mine types; the latter dwelled on factors necessary for effective adoption of modernization in the South African mining industry only, without evaluating the critical success factors and challenges. This study aimed to cover this lacuna.

Against this backdrop, the objectives of the study are, therefore, to empirically determine factors that influence acceptance of technology solutions in the mining industry in South Africa; assess critical success factors influencing the adoption of advanced technology in the South African mining industry; and explore challenges faced by the mining industry in South Africa in adopting advanced technology in its operations. The primary contribution of this study comes from the exploration of factors that influence the adoption of new technologies in South Africa, paying specific attention to the mining industry. Knowledge of such information allows the country to navigate through factors impeding the adoption of advanced technology in the sector and increase its rate of adoption. It also allows for formulation and implementation of critical policies in the sector, which will ultimately benefit the economy. As a result, it is hoped that the study's findings will have a positive impact on the development of a sustainable mining sector in South Africa and shed light on mining's significant contribution to the country's gross domestic product (GDP).

The rest of this paper is organized as follows: a review of the literature is followed by discussion of estimation techniques, results, and conclusions of the study.

Literature review

Mining dynamics in South Africa

Work to date indicates that modernization significantly extends mine life, preserves mining employment, improves safety and health, and allows mining of lower-grade ore bodies and deeper resources, which creates an environment conducive to 24/7 operations until 2045 and beyond (Minerals Council, 2021). A

low-grade mine with a current conventionally mined life expectancy of some four years, using semi-mechanized methods, could extend operations to 15 years and, with full mechanization and 24/7 operations, to as much as 25 years (Minerals Council, 2021). Ultimately, without a shift in mining methodology, the industry will fail to profitably mine South Africa’s deep-level complex orebodies. This could result in sterilization of resources, accelerated and premature mine closures, and job losses. Research suggests that 200,000 job losses by 2025 could indirectly affect 2 million people (Minerals Council, 2021).

South Africa remains one of the world’s most significant mining destinations in terms of quantity and variety of mineral products produced (Technology Innovation Agency (TIA), 2012). It harbours the world’s largest reserves of platinum-group metals (PGM), gold, and coal, and is ranked first in the production of PGM, manganese, vanadium, and chrome (Statistics South Africa (StatisticsSA), 2015). South Africa’s rich endowment of mineral resources has played a vital role in the evolution of the economy. The mining industry was reported as the second-largest economic contributor in South Africa in 1980, with a contribution of 21% of GDP (StatisticsSA, 2017; 2022). According to a TIA report (2012), economic growth in South Africa will continue to be closely linked to proceeds from the mining sector.

While the role played by mining varies within different economies, the mining industry in many economies remains a key player in national growth, and South Africa is no exception. The presence of mineral resources in South Africa provides enormous potential for sustained economic growth and development. Generally, areas of major contributions of mining to the economy include foreign direct investment (FDI), revenue generation in the form of taxes and royalties, contribution to GDP, formal and informal employment, and sales revenue from exports (Bernard, 2018). Overall, mining contributes 1.5 million jobs and renders direct and indirect support to about 15 million people (Chamber of Mines, 2016).

The South African mining sector has, for more than 100 years, been considered a labour-intensive industry, using physically demanding manual drilling methods with blasting and cleaning on a stop–start basis, predominantly in narrow-reef hard-rock mining for gold, platinum, and chrome (Minerals Council, 2021). Although mining is still largely a labour-intensive process, the mining industry makes use of a wide range of technologies to reduce and prevent incidents related to health and safety. Central to curbing underground accidents is, as far as possible, the removal of miners from working-face dangers and in-stope health hazards. Where that is not possible, technology is directed at protecting employees.

Between 1984 and 2005, more than 11,000 mine workers died in South Africa, whilst in 2003, the death toll from mining accidents was approximately 270 fatalities. Consequently, an agreement was reached to reduce mining fatalities by 20% per annum (DMRE, 2017). The technology currently used represents incremental improvements, but offers significant contributions to making mines safer places to work. Furthermore, South African mining companies are collaborating with each other and equipment producers to develop better and safer working methods and technology (Minerals Council, 2021).

There are some 160 Mt of high-grade ore locked in underground support pillars that are accessible from current infrastructure, but at least double that could be mined below current infrastructure using appropriate technologies (Minerals Council, 2021). Once the accessible deposits are depleted, the challenge is to decide whether

Technology and growth in the South African mining industry

to go deeper or go somewhere else. In the past, going deeper was not possible due the cost and limitations in the technology. Today is different, with new advancements in technology (like robotics) and high commodity prices, so going further underground makes more economic sense, instead of transporting an entire fleet and infrastructure to a new location (Bravo, 2012).

Mineral sales represent a large export share in many economies, and South Africa is no exception. Through international trade of minerals, mining brings in huge foreign earnings from mineral sales and FDI into the South African economy. As stipulated by StatisticsSA (2017), revenue realized from both local and foreign mineral sales amounted to R 387 billion in 2015. Of this, coal contributed a significant 27%, followed by PGM with 25%, gold at 16%, iron ore at 10%, and the remainder by other minerals (Berman, 2017). Commodities with highest sales revenue contribution include PGM, coal, and gold. These three key minerals contributed 71% of the total mineral sales recorded in 2009, valued at R 171,876 billion (StatisticsSA, 2017; 2021). These minerals culminated in a 63% total contribution to minerals’ GDP in that same year.

Owing to the small domestic demand for most of South Africa’s mineral commodities, the mining industry is largely exportoriented (Bernard, 2018). Mining exports contributed between 40% and 50% of South Africa’s total exports between 2001 and 2015, and are a major source of foreign exchange earnings (StatisticsSA, 2015). As stipulated in the Chamber of Mines (2016) report, mining exports amounted to R 320 billion in 2015. Nearly a third (27%) was accounted for by PGM, 21% by gold, 16% each by coal and iron ore, 10% was oil products, approximately 5% each for chromium and manganese ores and concentrates, and the balance by a few others (Bernard, 2018). Moreover, if beneficiated mining products are taken into account, then 60% of export revenue emanated from this broader category (Berman, 2017).

In 2022, South Africa’s mineral production achieved a record high of R 1.18 trillion, up from R 1.1 trillion in 2021, which was the first time the industry topped the trillion rand mark (Minerals Council, 2023). This performance was driven by strong commodity

prices, providing the domestic economy with a vital injection of higher taxes to bolster the fiscus, increase employment, and improve wages (Minerals Council, 2023).

According to the latest statistics released by the Minerals Council (2023), the mining industry is one of very few sectors in South Africa that is adding jobs in the prevailing economic climate. The sector created 15,500 more jobs in 2022, lifting total employment to 475,560 in the sector (Minerals Council, 2023).

Determinants of the adoption of advanced technology in mining: A review of theoretical literature

Literature reveals interchangeable use of the terms ‘adoption’ and ‘diffusion’, although these terms are quite distinct from each other. Adoption refers to "the stage in which a technology is selected for use by an individual or an organization" (Carr, 1999); diffusion refers to "the stage in which the technology spreads to general use and application" (Rogers, 1983). Therefore, while the term adoption is used at individual level, diffusion can be thought of as adoption by the masses. Adoption at individual and organizational levels leads to mass adoption, which is termed as the diffusion of technology. Hence, while looking into the evolution of research of technology adoption, we considered diffusion studies as well as adoption studies. Theories and models that have evolved for explaining adoption of technology are diffusion-based theories and the technology organization environment framework.

Diffusion of Innovation Theory (Rogers, 1983)

Research in diffusion can be traced back to the epic work by Everett Rogers in 1983, known as the Diffusion of Innovation Theory, which has been widely applied by researchers over the years. The main idea of the theory is that four elements influence the spread of a new idea: innovation, communication channels, time, and social system. The process of diffusion consists of five stages; namely, knowledge, persuasion, decision, implementation, and confirmation. It results in six categories of users: innovators, early adopters, early majority, late majority, laggards, and leap-froggers (Rajesh and Rajhans, 2014). The theory can be depicted as shown in Figure 1.

(1995)

Source: Extracted from Rogers (1983)

Figure 1—The Diffusion of Innovation Theory (extracted from Rogers, 1983)

Technology and growth in the South African mining industry

100%

Number of percentage of adopters

Perceived benefits

Perceived

Source: Extracted from Oliveira et.al. (2011, p.117)

Figure 3—Iacovou, Benbasat, and Dexter model (extracted from Oliveira et al., 2011, p. 117)

The diffusion innovation theory provided the concept of S-shaped curve of adoption, also called the epidemic model of adoption. According to this curve, spread of infections among a population can be held as an analogy to the pattern of spread of a new technique or idea. According to this analogy, the rate of spread is initially slow; in the mid-range of the graph, the rate of spread accelerates, and finally the rate of spread tapers off, resulting in an S-shaped curve, as depicted in Figure 2.

The reasoning for such S-shape curve is that innovation has to initially come from outside the boundaries of the social system prevalent at that time. Eventually, the innovation is accepted by most members of the social system and the rate of spread declines. The S-shaped curve depicted in Figure 2 illustrates that there is a critical "take-off point", at which the slope of the growth curve becomes positive and the number of members who have adopted the innovation becomes so large that there are hardly any new members left to adopt it. According to Rogers (1983), this point occurs when 10% to 20% of the members of the social system have adopted the innovation (Rajesh and Rajhans, 2014). This S-shaped adoption curve applies to most innovations that arise from time to time. The phenomenal growth of the Internet over last 15 years is often interpreted by this law (Rajesh and Rajhans, 2014). Value of the innovation is enhanced for existing users as more and more people adopt the innovation.

Technology–organization–environment framework

The technology–organization–environment (TOE) framework was developed to identify three aspects of an enterprise’s context that influence the process by which it adopts and implements a technological innovation; namely, technological context, organizational context, and environmental context (Oliveira and Martins, 2010). It is important to understand these different contexts and how they influence technology adoption. According

S-shaped diffusion curve

to Oliveira and Martins (2010), technology context is described as technologies that are relevant to the organization, which can be found internally within the organization and those external to the organization. Technology can refer to equipment, processes, and information systems (Aizstrauta et al., 2015). The technology context can be applied to the mining industry. This because companies such as Rio Tinto mention being leaders in terms of technology in mining and having to develop technology internally, but also adopting technologies from suppliers (Rio Tinto, 2019).

According to Oliveira and Martins (2010), the context of the organization refers to measures that describe the organization, such as formal and informal processes, structures, communication processes, and size. These elements may differ for mining companies, depending on the product produced, their history and background, and how long the company has been in existence. All of these descriptive factors ultimately influence technology adoption. Environmental context describes the context in which an organization operates. This includes, for example, market structure, technology support, infrastructure, and government regulation. The theory of TOE is relevant to the mining industry because all context elements can be described to understand how these influence technology adoption.

Iacovou et al. (1995) described the characteristics that influence firms to adopt information technology (IT) innovations. According to Iacovou et al. (1995), there are three main circumstances that influence organizations in adopting IT innovation: the derived or perceived benefits that the company wishes to achieve, the organizational readiness, and pressures that are external to the organization. These three elements influence the decisions made by organizations. Figure 3 shows the association of the three elements. The perceived benefits can be described as the value drivers that influence company decisions to invest resources to achieve the desired goals. The perceived benefit for South African mining

Figure 2—S-shaped adoption curve (extracted from Rogers, 1983)

Technology and growth in the South African mining industry

companies is that they remain competitive when compared with their global mining counterparts; the benefit for the Minerals Council is for mining to remain a significant contributor to the economy of South Africa.

According to Ernst and Young (2015), the mining industry requires transformation, and a set of value drivers needs to be defined as part of the transformation strategy. The model refers to organizational readiness being one of the elements, and to infrastructure and funding specifically allocated for technology adoption. Similarly, this factor can be related to capital funding and information system backbones required to enable technology adoption for the mining industry. The last elements relate to external pressures faced by organizations, which refers to competitive pressure. For South African mining companies, this relates to the global and local pressures that are outlined by Lane et al. (2015).

Mavroudis and Pierburg (2017) argued that information systems form the basis for implementation of any technology solution because communications is an enabler and a requirement for all the mining technology themes outlined. The adoption models outlined above discuss specific elements or characteristics that are important for adoption of technology innovation, specifically for the Information Systems industry. The following factors can be concluded to be the same for all three adoption models: organizational factors, infrastructure and financial resources, and external factors. Incorporating all three models suggests that technology adoption is a complex, inherently social, developmental process; individuals construct unique, yet malleable, perceptions of technology that influence their adoption decisions. Thus, successfully facilitating technology adoption must address cognitive, emotional, and contextual concerns.

Determinants of the adoption of advanced technology in mining: A review of empirical literature

Using a qualitative approach, Sager (2021) investigated challenges to implementing autonomous mining operations. Results showed that operational changes have the most impact on people, before process and technology, and are often not given enough attention in projects. Additional obstacles are the cyclical nature of the mining business, coupled with uncertainty about the remaining life of the mine. This impacts the long-term financial planning needed for an autonomous mining project, which is resource-intensive and of several years’ duration (Sager, 2021).

Using data from 28 interviews with mining experts, Ediriweera and Wiewiora (2021) identified five environmental and four organizational barriers to technology adoption in mining. Inadequate engagement with external stakeholders, uncertainties, and cyclical nature of the sector are key barriers to innovation adoption. Other barriers identified include high risk related to adoption of unproven technology and performance systems focused on volumetric production. They also uncovered five enablers to overcome these barriers to achieve more successful technology adoption outcomes, including, among others, learning culture, knowledge sharing, and external stakeholder engagement.

Nasirov and Agostini (2018) studied the key issues (barriers and drivers) influencing the adoption of solar technologies in the Chilean mining industry from the perspective of mining actors. They found that implementation of automation may result in resistance from workforce and unions due to the fear that employees with lower skill sets may be made redundant at the 'mine of the future'. Furthermore, they found that mining managers may oppose

the adoption of alternative energy sources to meet operational energy demands due to their general resistance to change; policy makers may have a similarly negative impact due to not offering incentives for use of renewable energies for mining operations. Fujiono (2011) used agent-based modeling to demonstrate ways that innovations can be adopted in the mining industry. The agent-based modeling technique allows the modeller to design a system that consists of agents with unique characteristics (e.g., preferences, options, strategy, size). These agents behave and perform actions based on sets of rules that can be influenced by the aggregate behaviour of the system (Fujiono, 2011). The key finding from this study was that interaction between mining companies in the agent-based model can lead to emergent phenomena; in this case, the locked-in phenomenon. This finding suggests that interaction between diverse entities should receive more attention in the study of the collective behaviour of the mining industry towards an innovation, without ignoring the technological, economical, and environmental aspects of the innovation. In other words, the behaviour of other firms in the mining industry can be a barrier or enabler of adoption of technology. Hilson (2000) conducted a study on barriers to adopting cleaner technologies and cleaner production practices in the mining industry in the Americas. Using important regional examples, the barriers to adoption of the technologies were identified as legislative, technologic, and economic in nature.

Overall, of the various empirical studies reviewed, although they propose various critical success factors and challenges of embracing technological development as key to growth, they reveal the most common critical success factors as learning culture, knowledge sharing, high labour costs, and behaviour of other firms in the industry. Commonly identified challenges are inadequate engagement with external stakeholders, uncertainties, cyclical nature of the sector, and high risk related to adoption of unproven technology and performance systems focused on volumetric production.

Estimation techniques

This study adopted an exploratory research design to facilitate exploration of factors that enable or inhibit adoption of technology innovation solutions for the mining environment, particularly in the South African context. The exploratory research design was adopted because it helps in more efficiently understanding the problem.

The study took direction from Yin (2009) to employ the casestudy research strategy, which assists in building in-depth and contextual understanding of the case in question using multiple sources of evidence. According to Baxter and Jack (2008), the collective case-study strategy is intended for gaining insight and understanding of a situation or phenomenon when more than one case is being examined. The results from a collective case study are robust and reliable (Baxter and Jack, 2008). This research study used a collective case-study strategy, aligned with the need to identify key enablers and inhibitors in a mining environment that are necessary for effective technology adoption in the South African mining industry.

Data collection

The target population for this study comprised all mining companies, mining-related organizations, and mining associations in South Africa. The study focused on four of the eleven commodity clusters to collect data (diamond, coal, iron ore, and platinum), purposively selected because these constitute the greatest proportion of mining activity in the country. The criteria used in selecting the

Technology and growth in the South African mining industry

mining operations was that they needed to have implemented or be in the process of considering a technology or any modernization concept for at least one piece of equipment in their value chain. Using the purposive sampling technique, a sample of thirteen executives and senior managers from ten mining companies was drawn, from whom data were gathered using face-to-face or video-call semi-structured interviews, using an interview guide. The respondents held various positions, in 2021, responsible for implementing or introducing the advanced technology in the mining sector. All had experience ranging from 10 to over 20 years, and most possessed a degree at minimum. These statistics indicate that the respondents were highly educated and capable of implementing technology. Of the 13 respondents, five were females and eight were males; all were aged at least 18 years, with most being over 45 years of age. Following Saunders et al. (2018), who suggested that the sample size for qualitative analysis should comprise between five and 25 participants, the study’s sample size of 13 was considered sufficient to realize the study objectives in a defensible manner.

Before commencement of the data-gathering process, the study received ethical clearance, with adherence to the ethical considerations of informed consent, confidentiality, and beneficence.

Data analysis

Data analysis was undertaken using Thematic Content Analysis (TCA). The data collected through online recordings were converted to written format and analysed using Atlas-ti software. Braun et al. (2014) stated that TCA is an essential tool to evaluate textual data. According to Braun et al. (2014), TCA comprises six phases: data familiarization, coding, theme scan, reviewing themes, defining and naming themes, and then writing up the information. To ensure validity and reliability, the same questions were posed to the respondents, which was an effective way to minimize bias in the data collected.

Results

The primary objective of this study was to understand the critical organizational factors that support or inhibit acceptance of technology solutions for effective adoption of modernization in the mining industry of South Africa. In this section, data gathered is analysed and discussed in an attempt to realize the research objectives. The discussion of results is organized according to the sub-objectives of the study.

Objective 1: Factors that influence the acceptance of technology solutions in the mining industry in South Africa Information on the factors that lead to acceptance of technology by workers and the organization was gathered. These factors were found to be correlated to delivery of the organizations’ strategic business goals in delivering value for all stakeholders in the short, medium, and long terms. Respondents had the opportunity to rank the factors in order of importance. Figure 4 summarizes these factors.

As shown in Figure 4, 54% (7) of the respondents outlined that if technology is implemented in the mining industry, cost can be greatly reduced because less manpower would be required, while 61% (8) stated that mining companies might introduce technology in the mining sector due to legislative requirements. The respondents suggested that mining companies must comply with the Mining and Safety Act of 1996; meeting the requirements of the Mining Act and the implementation of technology would increase levels of safety in the mining sector because human error would be reduced. All respondents stated that implementing technology would ensure environmental sustainability because mines would access underground minerals, whilst reducing environmental degradation and pollution.

In addition, 100% (13) of the respondents seconded the idea that technology would improve efficiency in the mining sector and increase rates of productivity. Furthermore, 85% (11) of the respondents suggested that investing in technology would enable the mining industry to unlock future resources that are hidden deep underground.

In terms of value drivers, most respondents across all commodity clusters cited value drivers as the most important motivator for organizations to commit to adoption of technology and innovative solutions to enable value delivery to their shareholders. These findings are in line with the theory proposed by Iacovou et al. (1995), who claimed that perceived benefit is one of the most important factors influencing organizations' adoption of IT innovation. Aizstrauta et al. (2015), Bryant (2015), and Lane et al. (2015) also support this viewpoint. The primary value drivers, according to the respondents, are safety improvements, efficiency and productivity increases, operational cost reductions, and environmental sustainability. Bryant (2015), Lane et al. (2015), Mavroudis and Pierburg (2017), Stanway et al. (2017), and others corroborate these major value drivers. These motivations are recognized as major benefits that inspire mining firms to change their ways of doing things by implementing technology.

Technology and growth in the South African mining industry

Across the commodity clusters, all respondents from mining organizations agreed that there has been an emphasis on improving safety and health conditions in the industry. The iron-ore cluster's Iron Ore 2 respondent and the platinum commodities cluster's Platinum 2 respondent both emphasized the need for safety improvements and highlighted that it is a regulatory necessity. Most respondents from all mining commodity clusters agreed with Iron Ore 1 respondent that mining businesses are implementing technology-based solutions with the goal of boosting productivity targets. Experts 1 and 3 confirmed this.

Most respondents mentioned high labour and energy costs for mining operators as the two main contributors of high operational costs. These cost drivers are confirmed by Lane et al. (2015) and the Facts and Figures 2017 report released by the Minerals Council (2018). South African mining companies are driving efforts to reduce operating costs, in an attempt to improve profit margins and become competitive again. In addition, Iron Ore 1 respondent mentioned the unlocking of future resource value as a strategic value driver. This motivation for securing the future of mining comes as a unique finding when compared with the other respondents: most respondents were looking at surviving today, and not necessarily at securing future resources, given the depressed state of the economy. However, this contradicts this value driver, as highlighted by Lane et al. (2015), the Minerals Council (2016), and Jacobs and Webber-Youngman (2017), who noted that mining companies ought to secure future resources by investing in morerapid discovery of resources and investigating new beneficiation methods to mine low-grade ores.

Most platinum commodity-cluster respondents mentioned that environmental sustainability was one of the drivers for implementing technology solutions to minimize their carbon footprint and water utilization. Platinum 3 respondent commented that energy costs were on the rise, and the company was therefore looking for technologies with less-intensive energy requirements. This statement was confirmed by Bryant (2015), who noted that energy inefficiencies have been alarming: only 12% of the energy from equipment contributes to delivering production; the rest is dissipated as heat or friction. The Minerals Council (2016) confirmed environmental sustainability as one of the drivers for modernization, and defined the scope for modernization in six areas, one of which is environmental sustainability. Figure

Objective 2: Critical success factors influencing the adoption of advanced technology in the South African mining industry Figure 5 illustrates the critical success factors that are needed to ensure that workers and organizations accept technology in their mining processes.

As shown in Figure 5, 100% of respondents stated that workers need to be trained, and this would equip them with the skills necessary to operate advanced machinery. These results are consistent with Bryant (2015), who propounded that technical skills are an important factor because this leads to efficient use of technology in the mining sector. All respondents suggested that there is need for good employer–employee relationships and executive sponsorship to ensure successful implementation of technology in the sector. In addition, all respondents suggested that there was need for investment funding to purchase advanced machinery and for training of workers: this would positively influence the acceptance of technology. These findings were confirmed by Bryant (2015) and Mavroudis and Pierburg (2017), who mentioned the need to spend capital in order to implement required infrastructure and, in the case of automation, that capital is also required for procurement of hardware to be installed.

Research and development was cited by 90% of respondents as an important tool to ensure successful implementation of technology and 80% stated that communication is essential when the organization is implementing technology: without communication, the adoption of technology would not be a success. The importance of communication was confirmed and emphasized by Rogers (1983), who noted this as being an integral element of diffusion of the innovation process. Most respondents referred to the role of the sponsor in taking the lead to actively share information with other executive members, which then filters through to the rest of the organization to create a mutual understanding (Rogers, 1983). Finally, 70% of respondents suggested that managers must develop strategies that would manage resistance to change.

Objective 3: Challenges faced by the mining industry in South Africa in adopting advanced technology in its operations

The respondents identified four factors as potentially inhibiting

Technology and growth in the South African mining industry

and preventing technology solutions from being widely accepted, thus preventing mining companies from being effective in their modernization drive. These inhibiting factors are low technology readiness levels, capital funding, the lack of required digital skills necessary for executing technology functions, and behavioural change. The responses are displayed in Figure 6.

Technology readiness was cited by most respondents as the level of maturity of a technology innovation, either tested or implemented in a mining production environment.

Consistent with this objective, 70% of the respondents agreed that there was low impetus for companies to accept low readinesslevel technology, which is associated with high risk of introducing an unproven technology into an operational business value chain. This was described as an inhibitor for technology adoption because companies avoid accepting technologies that have a low readiness level.

Another inhibitor is capital: all respondents and industry experts suggested that lack of funding was a definite inhibitor for introducing a technology solution. Some respondents suggested that companies cannot fully implement technology if they do not invest capital in automating or replacing existing equipment with autonomous equipment, which requires additional capital. Lack of funds is a significant impediment to implementing technology solutions in the mining industry, which, according to Bryant (2015), has received limited support due to the cost-cutting efforts.

Respondents also identified lack of the necessary digital skills as an inhibitor. According to Laubscher (2018), given the country's low level of qualifications, these skills are difficult to obtain, which explains the difficulty in addressing the unemployment rate. Sirinanda (2019) echoed these findings, stating that mining companies should invest in professional dynamic technology capabilities to better understand the intricacies of the industry. Approximately 80% of respondents also suggested that adoption of technology requires the right skills, so a lack of skills for its implementation serves as an inhibitor to effective adoption of modernization technology.

Two-thirds of respondents suggested that behavioral change can also be an inhibitor to the effective adoption of technology because there could be resistance to change amongst employees, who are the main implementers of a technology solution. According to Koul and Eydgahi (2018), technological change influences social and

behavioural changes: if an individual's behavior does not change in response to these changes, acceptance will fail.

Conclusion

In this study, we empirically determined factors that influence acceptance of technology solutions in the mining industry in South Africa. We assessed critical success factors influencing the adoption of advanced technology in the South African mining industry and explored challenges faced by the industry in adopting advanced technology in its operations. TCA was employed in the data analysis. The results showed that critical success factors to the adoption of mining technology include learning culture, knowledge sharing, high labour costs, and behaviour of other firms in the industry; commonly identified challenges include inadequate engagement with external stakeholders, uncertainties, cyclical nature of the sector, and high risk related to the adoption of unproven technology and performance systems focused on volumetric production. Various factors influence implementation of technology in the mining sector, including the need to reduce costs, improve safety, discover new unexploited resources, and improve environmental sustainability.

Companies face various challenges in implementation of technology in the mining sector, so they need to mitigate this by appropriate strategies. From the results of the study, three key recommendations are proposed. Firstly, executives and senior managers in the mining sector should learn how other industries, such as agriculture, manufacturing, and healthcare, are adopting technology in their sectors. This could help the industry to overcome challenges they are facing by implementation of strategies used by other industries. Secondly, the Minerals Council is recommended to motivate mining companies to adopt technology in their activities and provide aid in the form of resources or information on how to best adopt technology. Thirdly, results of the study suggest that employees in the mining sector lack the necessary skills needed for successful adoption of technology. The government, alongside mining companies, is recommended to assist by offering training workshops to equip employees with the necessary skills. The government may also provide tax holidays or tax relief to firms that import advanced mining technology. This would reduce the cost of purchasing machinery, thereby incentivising mining companies to invest in advanced machinery.

Figure 6—Challenges faced by the mining industry during the implementation of technology

Technology and growth in the South African mining industry

References

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Technology and growth in the South African mining industry

HEAD OF SCHOOL – SCHOOL OF MINING ENGINEERING FACULTY

OF ENGINEERING AND THE BUILT ENVIRONMENT

The University of the Witwatersrand, located in the heart of Johannesburg, one of Africa’s leading multi-cultural cities and the economic hub of South and Southern Africa is seeking to appoint a suitably qualified person to head the School of Mining Engineering which is recognised nationally and internationally for teaching, research, and industry partnerships.

Qualifications and requirements:

The successful appointee should possess high-level leadership and managerial skills; a PhD in Mining Engineering with evidence of significant professional and/or scholarly achievements with a strong research profile, national and international academic and/or professional standing, and expertise in any technical discipline in mining. Professional registration or eligibility for international applicants is recommended.

Key responsibilities:

• Provide visionary leadership, set strategic goals and objectives for the School of Mining Engineering.

• Develop and implement an industry-aligned research program by fostering research excellence, innovation and academic citizenship within the School.

• Lead and manage the School in meeting its strategic goals and participate in University-wide decision-making processes towards achieving the University’s strategic goals.

• Raise external funding for projects within the School’s vision and strategic goals.

• Develop and maintain collaborative relationships with industry partners, government agencies and other academic institutions globally.

• Manage and support staff recruitment, evaluation, and professional growth, promoting diversity and inclusion.

• Create an inclusive and supportive environment for students, promoting their academic success and well-being.

Remuneration:

A competitive university package is offered and will be commensurate with the level of appointment. The standard contractual period of employment is 5 years, after which the candidate may seek another term, or be integrated into an appropriate teaching and research position in the School.

To apply:

Please submit a cover letter of motivation, a detailed CV with names e-mail addresses and contact numbers of 3 referees, certified copies of qualifications, academic transcripts and South African ID, or a copy of your passport if not South African. Please state your research metrics in your cover letter. External applicants are invited to apply by registering on the Wits i-Recruitment platform at https://irec.wits.ac.za. Internal employees may apply on Oracle SelfService on the Wits Intranet by selecting “Apply for a Job”.

Enquiries:

Professor Thokozani Majozi Executive Dean: Faculty of the Built Environment – email Thokozani.Majozi@wits.ac.za Telephone +27 11 717 7012

Closing Date:

Please note that given the thrust of the University’s strategic plan on transformation, preference may be given to appointable applicants from the under-represented designated groups in terms of the relevant employment equity plans and policies of the University.

Affiliation:

1Karadeniz, Teknik Universitesi, Türkiye

Correspondence to:

K. Karaman

Email: kadirkaraman@ktu.edu.tr

Dates:

Received: 19 Sept. 2022

Revised: 20 Oct. 2023

Accepted: 16 Jan. 2024

Published: April 2024

How to cite: Karaman, K. 2024. Evaluation of different surface characteristics and mineral grain size in the estimation of rock strength using the Schmidt hammer. Journal of the Southern African Institute of Mining and Metallurgy, vol. 124, no. 4, pp. 173–184

DOI ID:

http://dx.doi.org/10.17159/24119717/2325/2024

ORCID: K. Karaman

http://orcid.org/0000-0002-3831-4465

Evaluation of different surface characteristics and mineral grain size in the estimation of rock strength using the Schmidt hammer

by K. Karaman1

Abstract

This study investigated the effect of surface roughness on Schmidt rebound hardness (RL). Four different test surfaces of rock samples were studied: natural, ground, cut surfaces, and core samples. There was significant variability of standard deviation based on the RL on the natural surface, which indicated high roughness of the rock surface, whereas surface polishing caused a significant decrease in standard deviation. ISRM and ASTM methods were compared to estimate unconfined compressive strength (UCS) for different testing surfaces. RL obtained from the cut surface was found to be more reliable than those obtained from other testing surfaces for the prediction of UCS; however, regression and ANOVA analyses revealed that the ISRM method gave a more accurate UCS estimation of rocks with highly rough surfaces. It was also shown that RL values obtained from a cut surface were significantly higher than those obtained from core samples. Therefore, a comparison between RL values obtained from core samples and cut surfaces was made based on previous studies. This study statistically showed that estimated UCS values are not statistically significant if Schmidt rebound tests are not performed on similar surfaces. In addition, the mineral grain sizes of the studied rocks, different testing surfaces compared with those in literature, and standard deviation from RL are evaluated and discussed. The Schmidt hammer technique is a rapid, inexpensive, and straightforward method for estimating UCS for preliminary assessment; however, roughness of the surface should be eliminated if variations are shown in the surface rebound hardness.

Keywords

Schmidt hammer technique, unconfined compressive strength estimation, surface roughness, mineral grain size

Introduction

The Schmidt hammer technique was initially developed in the late 1940s for testing the hardness of concrete (Schmidt, 1951) and, since the early 1960s, it has been used in rock mechanics practice (Deere and Miller, 1966). It has also been used for an increasing range of purposes, including the study of various weathering phenomena (Gokceoglu and Aksoy, 2000; Karpuz and Pasamehmetoglu, 1977), strength of joint walls (ISRM, 1978), rock discontinuity assessment (Young and Fowell, 1978), control of mine roof (Kidybinski, 1968), rock mass excavatability classification (Karpuz, 1990), performance of tunnel boring machine and roadheader (Bilgin et al., 1990; Poole and Farmer, 1978), penetration rate of drilling machines (Kahraman et al., 2003), determination of stabilization of glacially transported boulders (Wilson and Matthews, 2016), and saturation effect on strength and hardness (Karakul, 2017).

The use of the Schmidt hammer technique has been standardized by both the International Society for Rock Mechanics (ISRM, 2007) and the American Society for Testing and Materials (ASTM, 2013). Schmidt hammer models, such as L and N types, are designed with different impact energy levels. Orientation of the hammer, spacing between the impacts, surface roughness, weathering of the rock, size of the test sample, and the adopted test procedure are among significant parameters that influence the rebound values of rocks (Aydin, 2009; Goudie, 2006; ISRM, 1978; Karaman, 2020; Katz et al., 2000).

Hucka (1965) and Poole and Farmer (1980) indicated that the peak rebound values of repeated impacts at individual points are more reliable than first- or single-impact values; however, Shorey et al. (1984) used lower rebound values because they were more reliable for the estimation of unconfined compressive strength (UCS). Aydin (2009) stated that the density, distribution, and connectivity of its weak microstructural elements strongly affect the UCS values of a material; thus, high and low rebound values are equally necessary to reflect the nature of heterogeneity. Although different test procedures might be suitable for different applications, various researchers and institutions have suggested testing procedures that exhibit a wide variation in rebound values (Goktan and Gunes, 2005).

Evaluation of different surface characteristics and mineral grain size

Some authors mention surface roughness properties for the Schmidt rebound test. Williams and Robinson (1983) stated that rough surfaces yield lower hardness values than smooth surfaces for fresh gritstone blocks. Katz et al. (2000) carried out field measurements on three different surfaces: naturally weathered rock surfaces, surfaces manually polished with a grinding stone, and surfaces polished with an electrical grinder. They only published the standard deviation value changes from 1.93–5.57 (1.93 for surfaces polished with an electrical grinder, 3.80 for manually polished surfaces, and 5.57 for naturally weathered surfaces). They stated that high-quality polishing profoundly improved the quality of field measurements. Dabski (2009) studied limestone boulders that had early stages of weathering of glacially abraded surfaces and generally found higher rebound hardness values on polished surfaces than those obtained from non-polished surfaces. Cerna and Engel (2011) investigated variations of Schmidt hammer rebound value on the surface and sub-surface for a granite outcrop. They compared rebound hardness values obtained from natural and prepared surfaces, revealing that grinding before measurement provided more accurate data. Matthews et al. (2016) indicated that the first impact on surfaces tends to yield a relatively low RL value due to higher surface roughness, and such roughness effects were only removed after further impacts (usually less than five). Kogure (2019) proposed equations that distinguished two types of weathered surfaces, with higher rebound values at the surface of the indents than those at the surface of cliffs without indents. Karaman

Table I

(2020) investigated the effect of rock surface roughness on Schmidt hammer rebound number and confirmed that rebound values increased as the surface roughness decreased.

Many studies on the Schmidt hammer technique have been conducted on different surfaces (Table I). Some researchers performed the Schmidt hammer technique on surfaces without any polishing processes, while some used grinding stones before measurements. Surfaces were also prepared by researchers using an electric grinder and saw machine. Natural surfaces were widely used to determine weathering state and stabilization of glacially transported boulders. According to Table I, different researchers used various polishing methods before the Schmidt rebound measurements. Therefore, even if the rock types were the same, different rebound values were obtained because of variations in surface roughness properties.

According to the literature review, no study focused experimentally on the UCS estimation of rocks with different surface roughness properties. The main objective of this study was to investigate the effect of surface roughness on the relationship between UCS and RL. For this purpose, the UCS results for nine different rock types were correlated with the corresponding RL results. Many researchers compared their equations with previous studies, regardless of whether the test surface was the same. Therefore, this study aimed to investigate the usability of relevant test surfaces for comparison and how grain size and standard deviation affect RL measurements.

Summary of testing surface properties and conditions from the literature

Researchers Subject Rock/s

Type

Test surface processes

Kogure (2019) Mechanical characteristics of Pyroclastic N The test was performed without weathered pyroclastic rock rocks any polishing of the surfaces surfaces before the impact.

Wilson et al. (2019) Age determination of Granite N Boulder surfaces were not prepared glacially transported boulders boulders before measurement.

Yilmaz and Goktan Comparison of Schmidt hammer Masonry and L No treatment was necessary for (2019) technique and Equotip hardness building stones surface smoothening for the tester for rock strength evaluation core samples.

Han et al. (2019) A deep learning-based method Not given N Before measuring, the weathered for rock strength layers of the rocks were removed.

Goktan and Gunes Prediction of rock-cutting machine Mudstone, shale, N The rock surfaces were ground by (2005) performance sandstone hand with a carborundum wheel

Ozkan and Bilim Application of the Schmidt hammer Coal L The surfaces were manually polished (2008) technique in-situ on a coal face with a grinding stone.

Cerna and Engel Rebound value variation for strongly Granitic N The surfaces were prepared (2011) weathered and weakly outcrops using an electric grinder weathered granite outcrops

Buyuksagis and Effect of Schmidt hammer technique Different rock L/N The specimen surfaces were Goktan (2007) and test methods on UCS prediction types precision cut by diamondsegmented circular sawblades in the processing plant.

Vasconcelos Prediction of mechanical Granitic N The specimens were cut utilizing a et al. (2007) properties of granites rocks saw machine

Katz et al. Evaluation of mechanical Fine-grained quartz- N Naturally weathered rock surfaces. (2000) rock properties syenite

The surfaces were manually polished with a grinding stone

The surfaces were polished with an electrical grinder

Evaluation of different surface characteristics and mineral grain size

Materials and methods

Sampling and characterization of rocks

Rock samples were taken from the Black Sea region of Turkey (Trabzon and the surrounding area) (Figure 1). Their mineralogical and textural properties were determined using a trinocular polarizing research microscope. For the geological nomenclature of the claystone and quartzite samples, X-ray diffraction (Rietveld) analysis was performed (Karaman and Bakhytzhan, 2020). Petrographic thin-section analyses of the other samples are shown in Figure 2.

Each block sample was inspected for visible defects to provide standard testing samples free from cracks and fractures. Anisotropy is a significant parameter affecting the strength of the rocks. The studied rocks with no anisotropy (schistosity and foliation) showed no bedding planes, prismatic, pillow lava, or flow structure. The vicinity of the sampling locations was also checked to ensure that there were no faults or shear cracks.

Schmidt hammer measurements

Points were selected that avoided edge effects, cracks, and other visible structural weaknesses in the rock surface. Special attention was paid to ensuring that single impacts separated by at least the plunger diameter were made precisely on the rock surface. The hammer was periodically checked using the manufacturer’s test anvil. All RL tests were performed with the hammer held vertically downward. RL tests were performed using the L-type hammer in the laboratory on NX-size (54.7 mm diameter) core samples of five rock types belonging to different lithology definitions (granodiorite, basalt, diabase-1, diabase-2, and andesite). Hack et al. (1993) stated that no treatment is necessary for surface smoothening when the test specimens are obtained by coring. Test samples were rigidly supported using a core holder with a steel base during the testing. Aydin (2009) recommended sample sizes for the rebound hardness

test: NX size (54.7 mm) for core samples and at least 100 mm thickness for block samples.

RL measurements were also made on block samples with different surface properties, such as natural surface (without any polishing process), ground surface (using an electric grinder), and cut surface (using a saw machine). ISRM and ASTM methods that are widely used for rebound hardness determination (Buyuksagis and Goktan, 2007; Jamshidi et al., 2018; Karaman and Kesimal, 2015a) were used. These methods are described below.

Test Procedure 1 (ISRM, 2007): It is recommended to record twenty rebound values from single impacts separated by at least a plunger diameter and average the upper ten values.

Test Procedure 2 (ASTM, 2001): It is recommended to record ten rebound values from single impacts separated by at least the diameter of the piston, discarding readings that differ from the average of ten readings by more than seven units and determining the average of the remaining readings.

Roughness measurements

Barton and Choubey (1977) proposed ten standard joint roughness coefficient (JRC) profiles ranging from 0 to 20. The JRC value is the most commonly used measure for representing surface roughness (Hsiung et al., 1995). Each profile covers a range of two scales of JRC (i.e., 0–2). In the current study, a unique value (middle value of a range) was practically assigned to each profile, as in the study of Hsiung et al. (1995). The surface roughness was measured using a comb profilometer. The rock-surface roughness was quantified to assess its influence on rebound values and UCS estimation. The effects of weathering, instrument errors, and sample design were minimized to determine the surface roughness. The electric grinding tool was used stepwise to minimize dust formation and heating. Six roughness measurements, including impact points, were performed for each rock surface, and representative surface roughness was determined (Figure 3).

Figure 1—Major geological features of the study area (modified from Okay and Sahinturk, 1997)

Evaluation of different surface characteristics and mineral grain size

Density, porosity, and unconfined compressive strength tests

Core samples were prepared using a laboratory core drill and saw machines. For coring the rock blocks, a 54.7 mm diameter diamond coring bit was used. Trimmed core samples were used in the determination of density. The values of apparent porosity were obtained using saturation and caliper techniques. The UCS tests for fresh rocks were performed with a length-to-diameter ratio of 2.5, following the recommendations of ISRM (2007). All tests were carried out on intact rock samples. A machine with a 200 t capacity servo-control system was used for the UCS tests. The loading speed was applied within the limits of 0.5–1.0 MPa/s. The test was repeated five times for each rock type, and the average values are recorded as the UCS (Figure 4). The mechanical and physical properties of the samples are given in Tables II and III, respectively.

Results

Evaluation of standard deviation and roughness

The standard deviation of the Schmidt hammer rebound measurements was obtained in the range of 5.1–11.6 for the raw data. The average standard deviation of all rocks was 7.8 ± 2.4 for raw data, 4.3 ± 1.6 for natural surfaces (ASTM), 2.9 ± 0.8 for surfaces polished with an electrical grinder (ASTM), and 1.8 ± 0.5 for cut surfaces (ASTM). The average standard deviation of all rocks was 3.8 ± 1.3 for natural surfaces (ISRM), 2.0 ± 0.6 for surfaces polished with an electrical grinder (ISRM), and 1.1 ± 0.5 for cut surfaces (ISRM) (Table II). It was shown that the roughness of the test surface considerably affected the standard deviation. Figure 5 demonstrates a positive correlation between surface roughness and standard deviation, indicating that as the surface roughness

Figure 2—Microscopic images of the rock samples studied: (a) granodiorite, (b) diabase-1, (c) diabase-2, (d) andesite, (e) basalt, (f) lapilli tuff, and (g) limestone. Bt: biotite, Hbl: hornblende, Op: opaque mineral, Olv: olivine, Qtz: quartz, Pl: plagioclase, Prx: pyroxene

Figure 3—(a) Grinding process, (b) ground surface, test on (c) natural and (d) cut surfaces, (e) roughness measurements, and (f) evaluation

Evaluation of different surface characteristics and mineral grain size

Table III

Average test results and standard deviation values of samples

Rock type Density Apparent UCS (g/cm3) porosity (%) (MPa)

Granodiorite 2.65 ± 0.1 1.3 ± 0.3 170 ± 20

Diabase-1 2.79 ± 0.1 4.6 ± 0.3 116 ± 17

Diabase-2 2.83 ± 0.3 2.3 ± 0.5 183 ± 4

Andesite 2.55 ± 0.1 5.0 ± 0.3 86 ± 23

Basalt 2.56 ± 0.4 4.5 ± 0.9 163 ± 20

Lapilli tuff 1.86 ± 0.6 27.2 ± 1.2 12 ± 2

Clay stone 2.40 ± 0.5 8.0 ± 1.0 25 ± 10

Limestone 2.67 ± 0.1 0.7 ± 0.5 81 ± 12

Quartzite 2.39 ± 0.2 6.9 ± 0.6 60 ± 8

increased, the standard deviation also increased. A decrease in the standard deviation was shown for the natural surfaces compared with the raw data due to the rule of seven units (ASTM) and discarding the lowest 50% values (ISRM). According to the ISRM method, the standard deviation values were lower for all surfaces because the lowest ten of the twenty readings were not included in the calculations.

According to Figure 5, a high coefficient of determination (R2 = 0.82) was obtained between the standard deviation and roughness

Figure 5—Relationship between roughness and standard deviation

values of the test surfaces of all rocks except for Lapilli tuff. Large grains of various types and sizes were observed in a macroscopic view of the Lapilli tuff. Scattered data can be obtained from pyroclastic and breccia rocks.

Rock samples with rough surfaces were selected for this study to evaluate roughness effects, so natural roughness values of lapilli tuff, andesite, and granodiorite were found to be similar (Table IV). However, granodiorite contains quartz, which is harder than other minerals, and lapilli tuff contains rock fragments with different

Figure 4—(a–d) Some core samples, (e) Schmidt calibration anvil, (f) Schmidt hammer test, (g) before and (h) after the unconfined compressive strength test

Table II

Results of Schmidt rebound values for different surfaces

Evaluation of different surface characteristics and mineral grain size

Table IV

Average roughness values of the testing surfaces

Rock Natural Ground Cut types surface surface surface (JRC) (JRC) (JRC)

hardness. Therefore, the grinding process is affected by different hardness of particles (rock fragments, minerals, etc.) on the surfaces. Accordingly, the ground surfaces of granodiorite and lapilli tuff were rougher than those obtained from other rocks.

Relationship between rebound value and unconfined compressive strength

Predictive Analytics Software (PASW Statistics 18) confirmed the statistically derived equations. All variables (i.e., UCS and RL for all surfaces) were found to be normally distributed according to the Kolmogorov–Smirnov Z test and were then subjected to parametric statistical tests. Linear, power, exponential, logarithmic, and quadratic relationships between the variables were examined to obtain the most reliable equations. Analysis of variance (ANOVA) tables were also checked to determine whether regression models were significant. Similarly, the significance of coefficients in equations was examined.

Figure 6 shows the relationship between the UCS and rebound values obtained from different surfaces. Strong relationships (R2 ≥ 0.90) were obtained between the data pairs. As expected, the UCS values increased with increasing rebound values for all test surfaces. Statistically significant relations (exponential and power) within a 95% confidence level were obtained between UCS and RL for all test surfaces (Table IV). However, the natural surface and core samples had slightly lower coefficients of determination than those of the ground and cut surfaces. The study showed that the lapilli tuff and clay stone samples, which are weak rocks, affected the relationship type due to their low values of UCS and RL. Therefore, the relationship type for core samples is different (linear). A scatter plot of the Schmidt rebound hardness derived from the cut surfaces (JRC = 1) against the UCS values had the highest R2 value. The highest coefficient of determination could be related to the similar surfaces (cut and smooth) that were used in the UCS tests.

According to Figure 6 and Table V, the ISRM procedure gives a better prediction of UCS, for which the determination coefficients are within the range of 0.91–0.97. In contrast, the determination coefficients vary between 0.90 and 0.97 for the ASTM test procedure. These results agree with those of Buyuksagis and Goktan (2007) and Jamshidi et al. (2018), who obtained slightly higher coefficient of determination values with ISRM than ASTM. This study also revealed that the ISRM method provided a more accurate estimation of UCS for the natural test surfaces (JRC between 5 and 11); however, as the JRC decreased, the standard deviation values of

Natural surface

UCS = 3.95e0.071R

Ground surface

Cut surface

Core samples

UCS = 3.64e0.063R

UCS = 3.05e0.061R

UCS = 9.3R - 268

R2 = 0.92 R2

Natural surface

UCS = 5.6e0.07R

Ground surface

Cut surface

Core samples

UCS = 4.6e0.06R

UCS = 3.64e0.06R

R2 = 0.90

R2 = 0.93

R2 = 0.97

UCS = 0.0036R2.8

R2 = 0.90

Figure 6—Relationship between UCS and RL for different testing surfaces: (a) ISRM and (b) ASTM

the rebound readings also decreased, so the difference between the ISRM and ASTM methods became smaller (Figure 6).

The rebound values obtained from the cut surface were only close to those measured from the ground surface in the estimation of the UCS. It was observed that quite different UCS values could be predicted for the same rock type due to the effect of JRC (natural and cut surfaces). Increased JRC value leads to deviations not only in the standard deviation, but also in the empirical equations derived. A higher coefficient of determination was obtained as the JRC value decreased.

Comparison of RL values obtained from different rough surfaces

As seen in Table V, strong coefficients of determination were obtained between the UCS and RL data pairs of different rough surfaces within a 95% confidence level (R2 ≥ 0.90). The variations of RL values obtained from different rough surfaces were also tested using the one-way ANOVA for both methods (ISRM and ASTM). The Dunnett two-sided T-test was used to compare the RL values obtained in multiple tests to investigate the relationships between different surfaces: the RL values obtained from the cut surfaces were considered the control group. A significance level (SL) close to 1.00 indicates perfection of variance homogeneity (SL > 0.05). The variances of RL values were homogeneous (Levene statistic values = 1.397 and 1.416, and SL = 0.264 and 0.259 for ISRM and ASTM, respectively). When SL > 0.05, there was no difference between the mean values of the groups. According to the ANOVA results, no difference was obtained among the mean values of the groups (F = 1.163 and 1.717 and SL = 0.341 and 0.186 for ISRM and ASTM, respectively). The mean RL values from the ground surfaces were very close to the RL values obtained from the cut surfaces, with the lowest variation (Table VI and Figure 7).

Evaluation of different surface characteristics and mineral grain size

Table V

Results of the statistical evaluations

S. Le : Significance level

Table VI

Multiple comparisons of the RL

7—Comparison of mean values of RL obtained for different surfaces: (a) ISRM and (b) ASTM

The value of significance (SL > 0.05) for all groups revealed that there was no difference between the mean values of the groups (Table VI); however, SL (0.091) between the natural and cut surfaces was very close to the threshold value (0.05) for the ASTM method due to the high JRC values of the natural surface. In contrast, the study showed that RL values derived from the natural surface using the ISRM gave more similarity than ASTM, according to the SL value. It can be inferred from these findings that the ISRM method gives more accurate RL values for natural surfaces with high JRC values. Statistically, the ISRM method also provides better UCS prediction, especially on rough surfaces, because it excludes low values caused by surface roughness. Consequently, in terms of surfaces, using a ground or cut surface to find an RL value gives a more accurate prediction of UCS and lower JRC value than other surfaces.

Comparison of RL and estimated unconfined compressive strength values obtained from smooth surfaces

The rebound values obtained from core samples were lower than those obtained from cut surfaces, although both were smooth. For this reason, to verify the results obtained in this study, RL differences

were also examined by comparing surfaces (core and block samples) used in previous studies (Table VII) and extracting 125 RL–UCS values. The highest UCS–RL values are given in Figure 8. According to literature, the rebound values obtained from a cut surface are significantly higher than those obtained from core samples, although similar rock types were studied. This study confirmed that the differences between the rebound values obtained on smooth surfaces (core and block samples) were in close agreement with those obtained from previous studies. However, different Schmidt rebound values can also be obtained depending on the rock characteristics, test method, and application of the test parallel or perpendicular to the weakness planes (i.e., anisotropy and bedding planes, etc.).

Using the RL and UCS values from this study, the estimated UCS values were calculated and compared with the values from empirical equations proposed by different researchers (Buyuksagis and Goktan, 2007; Nazir et al., 2013) (Figure 9). The equation proposed by Nazir et al. (2013) for core samples was also used to estimate the UCS from RL of the block samples to compare the sample types (core and block). Similarly, the equation proposed by Buyuksagis and Goktan (2007) for block samples was also used in

Figure

Evaluation of different surface characteristics and mineral grain size

the estimation of UCS from RL of the core samples. As shown in Figure 9, the relationship between the UCS and RL derived from this study was very similar to that of Buyuksagis and Goktan (2007) at lower rock strength for the block samples.

According to the ANOVA results, no differences were detected between the mean values of this study and those reported by Buyuksagis and Goktan (2007) (SL = 0.248) for the block samples; however, estimated UCS values derived from the equation of Nazir et al. (2013) were significantly different from those of the current

study (SL = 0.039). The disparity could be related to the equation of Nazir et al. (2013) being developed using core samples (Figure 10(a)). Based on the ANOVA results, no difference was found between the mean values of this study and the study of Nazir et al. (2013) (SL = 0.184) for the core samples (Figure 10(b)); however, estimated UCS values derived from the equation of Buyuksagis and Goktan (2007) differed significantly from those of the current study (SL = 0.001) because their original equation was developed for block samples.

Effect of grain size on Schmidt rebound number

A trinocular research microscope was utilized to analyse the grain sizes of the samples. Although a grain exists in three-dimensional form in a sample, the grain size was expressed by area in this study. Three-dimensional grains were minimized by scanning the entire section with the help of a motorized table. Thin-section images related to grain sizes were examined using Clemex Image Analysis System software.

The largest and approximate average grain areas were calculated based on the grain sizes (Figure 11). While the average grain areas of the samples ranged from 0.23 to 4.8 mm2, the largest grain areas varied between 1.64 mm2 (andesite) and 58.33 mm2 (lapilli tuff). Furthermore, the plunger area of the Schmidt hammer (176.7 mm2) was about three times larger than the largest grain areas of lapilli tuff. However, larger grains (550–600 mm2) could be individually observed in macroscopic samples of the lapilli tuff (Figure 12).

This study confirmed that larger grains affected the rebound values, even if the surface of the lapilli tuff was smoothed. Aydin (2009) stated that when a surface contains grains with sizes comparable to the plunger tip diameter, the readings from these grains might significantly deviate from the average, depending on their strength relative to the matrix or dominant grain size. Findings that agreed with this suggestion of Aydin (2009) were obtained for the lapilli tuff samples (Figure 13(a)). In contrast, similar to the other rock samples, diabase-2 yielded consistent results regarding the effect of surface roughness and grain areas on Schmidt rebound values (Figure 13(b)). Scatter of the data increased as the JRC of the test surface of diabase-2 increased.

In such cases, impact points should be selected to obtain separate rebound readings from individual coarse grains and matrices (Aydin, 2009). On smooth surfaces where there were no large rock fragments, the readings were quite close to each other and showed a smooth line for cut surfaces (Figure 13(a)). However, the coarse grains of lapilli tuff had different rock fragments, which reflected different rebound values. Although all surfaces were evaluated together in this study, the findings derived from this rock

Figure 8—Effect of testing surfaces on peak rebound values

Figure 9—Comparison of estimated unconfined compressive strength values with previous studies

Figure

Nazir et al. (2013)

Evaluation of different surface characteristics and mineral grain size

Rock type

Not to scale R=15 mm

were not utilized in some graphs (Figure 5) to avoid affecting the results. In such rocks, the average hardness value is not affected and only the standard deviation values are variable.

According to the literature review, the effect of grain size on Schmidt rebound number has been addressed by a few researchers. Guney et al. (2005) stated that the estimation of engineering properties of rock materials (e.g., UCS via Schmidt rebound hardness) needs to be improved to account for more qualitative values that better represent the rock material, such as its origin, porosity, and grain shape. Atapour and Mortazavi (2018) produced artificial sandstones with different textural characteristics (median grain size: 0.31–1.63 mm and cement content: 15%–25%). They observed that the point load strength index and Schmidt rebound hardness of the samples increased as grain size increased.

Rock type

Andesite AD

mm2

1.64 mm2

Lapilli tuff

Limestone

AD

mm2 LD 1.71 mm2

Area of plunger diameter 176.7 mm2

Discussion

Aydin and Basu (2005) and Buyuksagis and Goktan (2007) studied the effect of hammer type on UCS estimation. It was reported that

Figure 11—Grain sizes of samples

Figure 12—Views of lapilli tuff samples

Figure 13—Effect of testing surface on variation of RL: (a) lapilli tuff and (b) diabase-2

AD: Average diameter, LD: Largest diameter

Evaluation of different surface characteristics and mineral grain size

the percentage difference (ISRM method) between L- and N-type hammers was about 15% on core samples in the study of Aydin and Basu (2005) and about 15%–20% on block samples in the study of Buyuksagis and Goktan (2007). Some researchers investigated the effect of the test procedure on Schmidt rebound hardness (Buyuksagis and Goktan, 2007; Jamshidi et al., 2018; Kahraman et al., 2002). Differences between the testing procedures similarly reached approximately 15% in the studies of Buyuksagis and Goktan (2007) and Jamshidi et al. (2018). In the current study, the average differences between the ISRM and ASTM methods were 12.30% for natural surfaces with high JRC values and 3.32% for cut surfaces with minimum JRC values. Some researchers investigated the effect of water content on Schmidt hammer rebound values and stated that water content reduces rebound values (Karakul, 2017; Sumner and Nel, 2002). In this study, the samples were kept in air-dried conditions at room temperature for five days to eliminate moisture in the samples and tests were carried out under the same conditions.

A rebound value of a core sample can be determined as, for example, 40 using one of the many testing procedures and an L-type hammer; however, for the same rock type, the rebound value of a cut-block sample can be calculated as 70 using another testing procedure with an N-type hammer. Researchers agree that the Schmidt hammer is quick, easy to apply, and inexpensive for either a site or laboratory, and technical expertise is not needed to use it; however, experience is required to evaluate and interpret the data primarily for UCS estimation. Moreover, scatter of data for Schmidt rebound values can be decreased through meticulous attention in the tests by following the instructions.

According to the revised version of the Schmidt rebound test (Aydin, 2009), the test may be stopped when any ten subsequent readings differ only by 4 (corresponding to an RL repeatability range of ± 2), instead of twenty rebound values. Soiltest Inc. (1976) recorded fifteen rebound values from single impacts and averaged the highest ten, provided that the maximum deviation from the mean was less than 2.5. It is understood from literature that a low standard deviation or scatter of readings is desired for the Schmidt rebound test. Cut surfaces and core samples typically exhibit a small standard deviation. Karaman and Kesimal (2015b) investigated more practical Schmidt hammer tests by reducing the rebound readings, especially for estimating UCS. The standard deviation of their study was below 2.5. If the standard deviation is small, the average rebound values will be close to each other, even if different methods are used. According to the current study, standard deviation, which is an indirect parameter, is also a good indicator that reflects the roughness of the surface. Roughness of a surface where the Schmidt hammer is applied should be eliminated, and the test surface should be re-evaluated if the standard deviation is high.

Many researchers have found correlations between UCS and rebound values on different rock types (Aydin and Basu, 2005; Fener et al., 2005; Kahraman, 2001; Karaman and Kesimal, 2015a; Yagiz, 2009). They generally found strong correlations between RL and the UCS. Some researchers also compared their equations with those of various authors who correlated UCS values with rebound values (Andrade and Saraiva, 2010; Kong and Shang, 2018; Wang and Wan, 2019; Wang et al., 2017; Yagiz, 2009; Yilmaz and Sendir, 2002). However, considerable differences exist between the empirical relationships proposed for estimating UCS using the Schmidt rebound value. Yagiz (2009) mentioned that the differences might be related to variations in the type and characteristics of the rock studied, the range of the dataset used (density, UCS, and rebound value), and the test methods chosen by the different researchers.

He also stated that a comparison with previous research indicated that these should be used cautiously and only for the specified rock types. The current study also proposes that test surfaces, hammer types, and sample types (core and block) should be similar if the results are to be compared with previous research.

Conclusions and recommendations

The Schmidt hammer rebound test was conducted on different surfaces (natural, ground, cut surfaces, and core samples) to estimate the UCS of rock materials and to experimentally understand the roughness mechanism. Rebound values increased as the surface roughness decreased, and standard deviation values decreased. The highest rebound values and lowest standard deviations were obtained from cut surfaces, while the lowest rebound values and highest standard deviations were taken from natural surfaces. Furthermore, lapilli tuff, which had different rock fragments, led to variations in the rebound values due to coarse grains, even if the sample surface was smooth.

Statistical test results showed strong relationships (R2 ≥ 0.90) between RL and the UCS of the rocks for all test surfaces; however, according to the coefficient of determination, RL from the cut surface outperformed the other surfaces in the UCS estimation. This study also revealed that the ISRM method provides a more accurate estimation of UCS for rough test surfaces. The study recommends that RL measurements be conducted on smooth surfaces; at least, an electric grinder is highly suggested for block samples.

According to this study, the standard deviation from the impact readings is an indirect parameter that presents knowledge about a test surface. Therefore, the test surface should be re-evaluated if the standard deviation is high. This study also suggests that Schmidt hammer rebound values can be compared with those previous studies if similar surfaces (ground or cut) and test conditions (hammer type and test procedure) are evaluated.

Acknowledgments

The author is deeply grateful to Karadeniz Technical University Research Fund for the financial support (Project No: 7828).

Conflicts of interest

The author confirms that there are no known conflicts of interest associated with this publication and there has been no significant financial support for this study that could have influenced its outcomes.

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Affiliation:

1Visiting Senior Lecturer The Wits Mining Institute, The University of the Witwatersrand and Director, The Sarraounia Public Health Trust

2Independent Researcher, South Africa

Correspondence to:

P.F. Stewart

Email:

PaulFinlayStewartKZN@gmail.com.

Dates:

Received: 8 Nov. 2022

Revised: 10 Jan. 2024

Accepted: 16 Jan. 2024

Published: April 2024

How to cite: Coulson, N. and Stewart, P.F. 2024. Communication constraints in the safety system on South African mines and implications for the exercise of the Right to Refuse Dangerous Work. Journal of the Southern African Institute of Mining and Metallurgy, vol. 124, no. 4, pp. 185–192

DOI ID:

http://dx.doi.org/10.17159/24119717/2444/2024

ORCID:

P.F. Stewart

http://orcid.org/0000-0001-7148-0054

N. Coulson

http://orcid.org/0000-0001-7148-0054

Communication constraints in the safety system on South African mines and implications for the exercise of the Right to Refuse Dangerous Work

by N. Coulson1 and P.F. Stewart2

Abstract

The Mine Health and Safety Act No. 29 of 1996, as amended, embeds an occupational health and safety management system that facilitates communication between representatives of the state, capital, and labour. In underground mines, two communication loops between these role players serve to separate the production chain of command from the political tripartite relations. Worker-elected health and safety representatives are involved in both communication loops, but are severely constrained in escalating their occupational health and safety concerns to the legalpolitical level. This has a direct bearing on the exercise of the Right to Refuse Dangerous Work (RRDW). As previously reported, health and safety representatives were found to primarily consult their production supervisors in preference to their trade-union representatives. This follow-up article presents production supervisor perspectives on the RRDW. The study demonstrates how, for mineworkers, the two distinct communication subsystems constrain, rather than facilitate, implementation of the RRDW.

Keywords

South African mines, worker safety, health and safety representatives, safety system, Right to Refuse Dangerous Work

Introduction

South African mineworkers won the Right to Refuse Dangerous Work/Leave a Dangerous Workplace (RRDW/RLDW) after a century of grueling and hazardous large-scale mining practices (Republic of South Africa, 1996). From the inception of modern industrial mining in the late 1880s, relations between black mineworkers and white supervisors were racialized. The chain of command at the underground rockface for the first hundred years of mining comprised black African labourers and ‘hands-on’ supervisory team leaders (formerly referred to as ‘boss-boys’) who served under white miners, who were, in turn, answerable to their shift overseer (formerly shift boss) and mine overseer (formerly mine captain). Systematic physical assaults and violence drove gold mining production until the mid-1960s (Moodie, 2005). Only the white miner and white supervisors, as well as the sole occupational group of black rock-drill operators, received incentive bonuses. Later, more rational and ‘scientific’ job-grading systems (such as the Paterson system), along with production bonuses for all mineworkers, were introduced into the mine workplace (Moodie, 2005). However, the key workplace qualification, the miner’s ‘blasting certificate’, was only finally deracialized as late as 1988. Until then, the job of blasting, though effectively performed by the black team leader, had been reserved for white miners. Team leaders routinely deferred to the miners for both equipment and practical knowledge, despite possessing workplace experiential ‘tacit knowledge’ (Leger, 1985, 1992). The supervision of ‘gangs’ or teams of black labourers was a key role of both team leaders and the miners.

Despite liberal reforms in employment practice from the end of the 1960s, poor occupational health and safety (OHS) performance continued throughout the 1970s and 1980s under a command-and-control management style (Leger, 1986). At the end of the anti-apartheid struggle years (late 1980s), two types of worker safety representatives had emerged to address poor safety performance. Supervisor-selected ‘safety representatives’ were introduced under new mining regulations in 1988 (Department of Energy and Mineral Affairs, 1988), while ‘safety stewards’ were part of new safety agreements that were signed off with the then newly established National Union of Mineworkers (NUM) in some mining companies, such as Anglo American, but not in others, such as Goldfields. At the time, the new generation of trade union-appointed safety stewards stood in stark contrast to the employer-dominated arrangements of government regulations (Leger, 1985). It was these trade-union safety stewards who laid the basis for worker representation in the current OHS system on South African mines post democracy in 1994.

Communication constraints in the safety system on South African mines and implications

The promulgation of the Mine Health and Safety Act (MHSA) No. 29 of 1996 (Republic of South Africa, 1996) meant that, for the first time, there were legal OHS provisions in place that required consultation between representatives of the state, all trade unions, and the mining industry. These provisions finally complied with the International Labour Organisation (ILO) conventions, of which South Africa had been a signatory in 1919 (Hermanus et al., 2019) regarding Occupational Health and Safety (Convention 155) (ILO, 1981) and Safety in Mining (Convention 176) (ILO, 1995). At the time, this represented a shift from historical legacies and was a significant victory for the NUM. It was on the insistence of the NUM that new pluralist arrangements for worker–employer consultation under the MHSA (refer to Chapter 3 of the Act) on large mines found legal expression in a mandatory Health and Safety Collective Agreement signed between the employer and recognized trade unions at a mine site (Leon,1995). The intention was that worker-elected health and safety representatives (HSR) were either trade-union members, or at least buttressed by recognized trade unions. With high trade-union density remaining a feature of the South African mining sector today, the conditions for these OHS arrangements remain (Bhorat et al., 2014).

Section 23 of the MHSA or the expression of the RRDW/ RLDW (hereafter refered to as RRDW) is a particularly progressive aspect of the OHS regime on South African mines. Under the Act, there are no stipulations regarding either the conditions and/ or type of danger (such as imminent danger) limiting the right to refuse dangerous work. A South African mineworker can therefore subjectively believe the workplace is too dangerous to enter or work in and choose to withdraw, and in doing so, be protected from discrimination. However in the event of exercising this right, the MHSA does not stipulate the procedure to be followed. It was only as a direct result of research conducted for the Mine Health and Safety Council (Bid No: MHSC 5/15-16) in July 2016 that the Chief Inspector for Mines gazetted the Guideline for a Mandatory Code of Practice on the Right to Refuse Dangerous Work and Leave Dangerous Working Places (see Coulson et al., 2019). This guideline (Department of Mineral Resources, 2016) reinforces the role of worker Health and Safety Representatives (HSRs) and trade unions, and compels the employer to negotiate a site-specific procedure to be followed when a worker exercises the RRDW. This guideline states that the resolution of unsafe workplaces should be facilitated at the lowest possible level, preferably in the workplace.

This paper advances the analysis of two studies that we conducted that examined, firstly, the role and dominant communication loops involving worker HSRs (Coulson and Christofides, 2021) and, secondly, the implementation of the RRDW (Coulson et al., 2019; Stewart et al., 2013). Specifically, this paper traces the communication between workers and supervisors when the RRDW is formally activated. Coulson et al. (2019) had workers’ responses to the RRDW as its key focus; this companion article analyses the responses of supervisors. The evidence of supervisors sheds further light on how communication in the safety management system is constrained and that effectively serves to undermine implementation of the RRDW and hence overall mine health and safety.

A note on methodology

This paper reflects on case-study research of worker HSRs conducted on four underground mine sites in South Africa between 2015–2016, during which in-depth interviews (n = 82) were conducted with a purposive sample of worker HSRs to understand

their role. Other in-depth interviews (n = 17) were conducted with members of the employer OHS system and focus groups (n = 4) with production team members. The detailed methodology for this study can be found in Coulson and Christofides (2021) and Coulson (2018). The methodology and findings with respect to the study of how the RRDW had been implemented have been outlined (Coulson et al., 2019), including details of ethical protocols and permission to publish by the MHSC, and are not repeated here. In brief, a representative sector-wide quantitative survey, focus group discussions, and key informant interviews were used in a concurrent triangulation mixed-methods approach (Cresswell et al., 2003). This article presents new data specific to responses of supervisors directly engaged in the production process.

Results part 1: Published research findings from two studies

Healthy and safety representatives and two closed communication loops

The MHSA makes provision for a two-tier system of HSR within the workplace: workplace HSR and fulltime HSR (Figure 1). Workplace HSR fulfill and perform their safety role in designated work areas only and while engaged in the daily routines of their fulltime occupation. Fulltime HSR are elected and seconded for up to three years and work in a fulltime capacity on health and safety issues across the whole mine site. On a large mine, fulltime HSRs are often elected for a specific shaft, usually between two and four representatives per shaft. Fulltime representatives can co-ordinate or have oversight of the activities of the workplace HSR. The combined two-tier system on very large underground mines means that there can be hundreds of HSRs on site. Over a decade ago, the Chamber of Mines (now the Minerals Council South Africa) estimated that 40 000 HSRs in the industry needed training (CoM 2009–2010, cited in Tuchten, 2011).

The HSRs (see MHSA Section 30) have, inter alia, the right to raise any OHS issue with the employer, represent workers, and talk to inspectors. They have the right to withdraw workers from a dangerous workplace in accordance with Section 23 (RRDW) of the Act and can have at least equal, or even greater, numerical representation at the employer–employee mine health and safety committee (Sections 29, 26, Regulation 6.9.a, 25(2)) (Figure 1). These arrangements for worker representatives in South

Figure 1—The mine health and safety management system as found on South African mines, in compliance with the MHSA No. 29 of 1996 as amended [Note references to the specific sections of the Act]

Communication constraints in the safety system on South African mines and implications

African mines are generous when compared with other mining jurisdictions, such as Australia where there may only be one or two site-based HSR (see Walters et al., 2017).

On a large mine, a third type of HSR is also found; a tradeunion shop steward appointed under the Labour Relations Act (Republic of South Africa, 2002), who is a member of the union health and safety structure (UHSS). This type of worker representative has no specific powers under the MHSA, but is commonly a member of the mine health and safety committee. Despite robust provisions for worker representation under the MHSA, serious dysfunctions have arisen in the tripartite project in South Africa (Coulson and Christofides, 2021). Case-study research undertaken by Coulson demonstrated the presence of two dominant communication loops (Figure 2 and Figure 3) (Coulson and Christofides, 2021). These were found to be characterized by weak and strong communication channels between different role players. In practice, the arrangements for worker HSRs, although compliant with the MHSA, were characterized by inconsistent communication between fulltime HSRs and workplace HSRs, as well as weak communication with organized labour. The mine safety department, however, would generally be in constant contact with production management. This makes the production chain of command, from shift overseer to the frontline supervisor team leader and work teams, the dominant daily communication. Thus, the interface between the production supervisor and the workplace HSR dominates and overshadows the relationship between the workplace HSR and their fulltime HSR. Thus, the workplace HSR is locked into the day-to-day micro-politics of production, as opposed to escalating workers’ safety concerns with their fulltime HSR. Coulson and Christofides (2021) provided an explanation for this by drawing on the work of Luhmann (1989), who argued that closed socio-autopoietic (or self-regulating) loops of communication serve the purpose of simplifying complexity in the broader environment in which any system is located. Thus, the Workplace Safety and Production Loop reduces the complexity of the socio-technical processes of production into a series of standards, rules, and regulations of anticipated behaviour, practices, and workplace orders. The Compliance and Enforcement Loop simplifies the political capital–labour relations under the auspices of the statutory Mine Health and Safety Inspectorate (MHSI) of the Department of Mineral Resources and Energy (DMRE) and manages OHS issues tabled by any of the members of the tripartite arrangements.

As a consequence of these two self-regulating communication loops, the two-tier system of worker representatives – workplace HSRs and fulltime HSRs effectively operate as the “eyes and ears” of the employer. They do not, as intended by the legislation, primarily serve as the autonomous organizational trade-union voice of workers. Figure 2 shows how, within the two-tier structure, the worker HSRs are communicatively restricted to the employer OHS system and safety department, which dominate the Workplace Safety and Production Communication Loop 1.

The study further reported, as can be seen in Figure 3, that it is the UHSS, located in the Compliance and Enforcement Loop 2 –rather than fulltime HSRs, who had a strong relationship with the regulator/inspectorate. In fact, the ultimate compromise surfaces in the role of the fulltime HSRs (Coulson and Christofides, 2019). These representatives were found in the case-study research to report into the safety department, which is central to the Workplace Safety and Production Loop 1 (Figure 2). Although the fulltime HSRs who participated in the case-study research were all tradeunion members, not one of these representatives (on four casestudy sites) worked from the trade-union office on the mine shafts. Further, it was not the norm for fulltime representatives to even report regularly to the structures of organized labour. As explicitly noted in Figure 2, employer safety management actively discouraged the fulltime HSRs from interacting with the UHSS.

Worker perspectives on the Right to Refuse Dangerous Work

The research commissioned by the MHSC on the RRDW (Bid No MHSC/5/12-13) found that while there were high levels of awareness about the right, the exercise of the RRDW had not been fully realized (Stewart et al., 2013). Indeed, the concern was why workers did not act autonomously regarding their own safety, despite over 90% reporting awareness of possessing their legally entrenched right to refuse to do dangerous work (Stewart et al., 2013). Instead of individually claiming their legal right to exercise their RRDW, mineworkers would first consult widely before withdrawing from a dangerous workplace. Workers consulted, among others, their worker HSR (32%) and supervisors (71%) (Coulson et al., 2019). This finding, significantly, indicates a marked shift in trust relationships between black African mineworkers and supervisors on mines. Despite the historically tense, racialized relationship between African mineworkers and white supervisors, evidence pointed to greater co-operative, communicative interaction than the contextual, worker-oriented literature had previously thought existed (Coulson et al., 2019). For example,

Figure 2—Workplace Safety and Production Loop (from Coulson and Christofides, 2021) [FTHSR: fulltime HSR; WHSR: workplace HSR; UHSS: union health and safety structure]

Figure 3—Compliance and Enforcement Loop (from Coulson and Christofides, 2021) [FTHSR: fulltime HSR; UHSS: union health and safety structure; DMR MHSI: Department of Mineral Resources Mine Health and Safety Inspectorate]

Communication constraints in the safety system on South African mines and implications

the percentage of workers and supervisors who considered their relations as a ‘big problem’ when meeting production targets was relatively low across coal mines (11%) and gold mines (12%), though rising to 18% in platinum mines (Stewart et al., 2013). While subject to confirmation by future research, this suggests that de-racialization three decades ago has immeasurably improved the worker–supervisor relationship. Objectively speaking, this was a sound basis for implementation of the RRDW in the mining workplace.

In the article by Coulson et al. (2019) based on the MHSCcommissioned study of the RRDW (Stewart et al., 2013), we made several observations about how this right was experienced by workers. Using a typology of work refusals first developed by Gray (2002, 2009), we showed that the formal RRDW was, to reiterate, not the default position for workers, despite a very high level of awareness about the right. Obstacles to the formal practice of the right were found in the responses of workers to qualitative questions. The right was believed to apply to safety issues, rather than health, and procedures for the RRDW could not be distinguished from the general safety rules that apply in an underground mine workplace. Workers also described “feeling bad” if production was lost when they exercised the RRDW (Coulson et al., 2019).

HSRs were crucial to the formal exercise of the RRDW. Bezuidenhout et al. (2015) previously found workers responded positively to HSRs. We found more workers were asked to stop work by an HSR (56%) than had personal experience of the RRDW as an individual (45%) (Coulson et al., 2019). Despite this, workers were critical of the lack of capacity of the HSR to escalate an issue beyond the immediate workplace. Workers suggested that their HSR lacked power or misinterpreted their role. Informal expressions (both confrontational and non-confrontational) of worker resistance to a dangerous workplace were also commonplace (Gray, 2002). Up to a third of workers described going back into a workplace while still deeming it to be dangerous, despite having withdrawn (Coulson et al., 2019). The organization of production work around teams of workers made the exercise of the right more difficult, as well as fear: fear of repercussions, of supervisor threats, of vindictive behaviour, and even of peer pressure by those who did not want to sacrifice their production bonuses.

The employer, organized labour, and the Right to Refuse Dangerous Work

Both the case-study research and study of the RRDW found the employer and organized labour responded positively to the RRDW. For example, a general manager issued laminated ‘RLDW cards’, personalized with his signature to demonstrate his commitment to taking safety seriously, in which he gave workers the ‘authority to withdraw from any unsafe working environment and to refuse to carry out any instruction that will endanger you and your fellow workers’ (Stewart et al., 2013). However, this proactive stance also stepped over an invisible line: in another instance, a manager ‘demanded’ that workers withdraw from a dangerous workplace (Coulson et al., 2019). In our paper (Coulson et al., 2019), we argued that the RRDW is vulnerable to monopolization and integration into employer risk-management strategies at the expense of individual workers, who had yet to learn to claim the right in the interest of both their own health and safety. Compounding this, we observed that the Guideline for a Mandatory Code of Practice on the Right to Refuse Dangerous Work and Leave Dangerous Working Places is partial to the employer because the recommended dispute mechanism favours the employer in the final instance; it does not

recommend inspectors from the MHSI as the final arbitrators. Thus, contrary to the powers conferred on the HSRs and the mine health and safety committee under the MHSA, which enable worker representatives to make a direct plea to MHSI, the inspectorate in their guidance on the RRDW puts inspectors out of reach for HSRs (Coulson et al., 2019).

Nonetheless, it is not just the employer that has found a role for the RRDW beyond that for which it was intended. Our study found that workers (35%) agreed that the RRDW was abused and used for ulterior motives (Coulson et al., 2019). Other studies corroborate this: in 2013, in the months prior to the start of South Africa’s longest labour strike on platinum mines, the RRDW was used by rock-drill operators to slow down production and gain management attention for improved wages (Moodie, 2016; Stewart and Nite, 2017). Our case studies (Coulson and Christofides, 2021) found that organized labour sided with the employer in the event of the MHSI closing a workplace for violations of the MHSA: trade-union representatives wanted the workplace open and said that workers could use the RRDW to protect themselves in the event of danger. Thus, the burden was placed on workers to keep themselves safe, rather than the employer that is legally responsible under the Act.

Results part 2: New data from the study of the Right to Refuse Dangerous Work

As previously noted, the data collected from supervisors was part of the research conducted for the MHSC (Bid No: MHSC 5/12-13) (Stewart et al., 2013). This additional data are presented here under two headings: the first concerns supervisor response to the RRDW; the second concerns supervisor engagement with worker HSR.

The demographic profile and occupational descriptors of the supervisory respondents (n = 96) signals a largely experienced workforce (Table I). The supervisor informants were mainly drawn from the gold and platinum sector. Over half (55%) had more than 10 years of mining experience and four-fifths were black Africans, which confirms the extent of transformation of the historical racial division of labour that has occurred in frontline production management command post-apartheid (Table I). Thus, although racial discrimination and undercurrents remain characteristic of much of the post-apartheid mine workplace (see Shaw et al., 2010), the narrative is now more complex than the historical alignment of race and class: that of black workers versus white supervisors and managers.

The reported occupations of the sample of supervisors were as follows: production manager (3%), shift overseer (32%), team leader (23%), miners (21%), safety officers (6%), and engineering department (15%). This supervisory echelon manage both the dayto-day production targets and hazards in the workplace, i.e., they are responsible for both production and safety. The engineering team advise and determine controls for hazards, while the shift boss, team leader, and miner make up a chain of command tasked with meeting production targets and mining safely.

Right to Refuse Dangerous Work and supervisor responses

When supervisors were asked who they trusted to provide accurate information in the workplace, their answers reflected their line function. Immediate supervisor and/or mine management was the predominant source of accurate information reported. Colleagues or other stakeholders, such as trade unions and HSRs, were very poorly referenced by the management chain (Figure 4).

Supervisors can exercise the RRDW in their capacity as employees; however, unlike HSRs, who are protected from any form of liability when acting in their role as representatives, management

Communication constraints in the safety system on South African mines and implications

respective line supervisors – the miner and the shift overseer. Only at the level of the shift overseer did it appear that mine management became more routinely involved in discussions about a dangerous workplace (Figure 5). Thus, the findings show that the management chain of command remained largely intact during concerns about a dangerous workplace.

By comparison, Figure 6 shows that only 13% of supervisor respondents who responded to this question discussed their concerns with the HSR and/or mine health and safety committee, and similar percentage of supervisors discussed concerns with union representatives. From these findings, it can be inferred that few instances of a dangerous workplace were escalated to fulltime HSRs, to trade-union shop stewards/UHSS, or to the mine health and safety committee. The findings were also suggestive that, far from matters being resolved in the immediate workplace between supervisors, the production team, and the HSR, supervisors generally chose to escalate matters to their immediate superior because only 22% discussed matters with colleagues or workers.

representatives carry responsibility for safety performance at all times. When the supervisor respondents were asked if they had ever experienced the workplace as too dangerous for work, 45% responded positively and 55% said no. Of those who responded positively (as having encountered a workplace they deemed too dangerous to enter), when asked with whom they had discussed their concerns, nearly three-fifths (57%) of all supervisor respondents abstained from answering the question. Of those who did respond, both miners and team leaders largely relied on their

The significance of the high number of abstentions from supervisors responding to questions about the RRDW spoke to supervisor concerns and their structurally contradictory position in having to manage the competing demands of both production and safety. This is a long-standing tension in the position of underground line management. Although the reasons for this were not specifically explored in the research, more than 50% of supervisor respondents preferred not to implicate themselves in any kind of discussion about a dangerous workplace. Further, more than 65% of shift overseers, 60% of miners, and 41% of team leaders abstained from answering questions when asked if they had returned to a workplace from which they had withdrawn because they thought it was dangerous. Only a third of respondents (34%) thought procedures were adequate for the RRDW and 55% of respondents abstained from answering this question. The reasons behind these large numbers of abstentions warrant further examination. This is suggestive, however, of how supervisors cope with managing the contradictory demands of production and safety. Supervisors fear blame for something being sub-standard in the workplace. Workers in the same study reported supervisors turning a blind eye to risk taking, taking risks themselves, and that even when something was clearly sub-standard, they would coerce workers to continue work or victimize those who raised a concern (Coulson et al., 2019).

Health and safety representatives and supervisors’ responses

The valuable role of HSRs in the exercise of the RRDW was reinforced by supervisor respondents. According to supervisors,

Figure 4—Who supervisor respondents (n = 96) trust to provide accurate information in the workplace

Communication constraints in the safety system on South African mines and implications

the HSRs had most experience of the RRDW. This view aligned with 56% of workers who reported that they were requested by an HSR to withdraw from a workplace (Coulson et al., 2019). Figure 7 shows that the miner, team leader, and shift boss all reported that, in their experience, it was the HSR who was most likely to have asked workers to withdraw. This finding is in stark contrast to the finding that only 13% of supervisors consulted with HSRs about conditions in the workplace when concerned that it was too dangerous. As noted above, this diverges with regulatory guidance in the sector, which places the onus on the resolution of unsafe workplaces to be facilitated at the lowest possible level, preferably in the workplace. The insights from supervisors suggest that although HSRs may initiate the RRDW, they are unlikely to be found discussing the solution. This also correlates with workers’ experience that HSRs lack the power to ensure matters are adequately resolved (Coulson et al., 2019).

Discussion

The new data shared in this paper indicates that many supervisor respondents experienced chronic vulnerability. This arguably became acute when having to openly confront their structural position in response to researchers’ direct questions. Supervisory respondents chose to abstain from answering questions about their own behaviour in the event of having withdrawn from a dangerous workplace. Thus, the views of these underground supervisors remain under-researched, despite being the custodians of a collective wealth of organizational and experientially based mining knowledge.

Figure 8 summarizes the findings of this paper. This shows that the communication and interaction between parties, once the RRDW is triggered, did not precipitate a change to the dominant patterns of communication for the HSR (Coulson and

Figure 6—Who supervisor respondents (n = 96) discussed concerns about a dangerous workplace with [Note: 57% of supervisor respondents abstained from answering this question]

Figure 7—Supervisor respondents’ (n = 96) experience of the Right to Refuse Dangerous Work

Figure 5—With whom supervisor respondents (n = 96) discussed their concerns about a dangerous workplace [Note: 57% of supervisor respondents abstained from answering this question]

Communication constraints in the safety system on South African mines and implications

Christofides, 2021). Although the exercise of the RRDW was more likely to be triggered by the workplace HSR, the expectation that workers and supervisors may, as a consequence, come together to review the workplace, did not occur. Rather, supervisors, starting with the miner, escalated issues to their immediate superior, as determined by the production management chain. This upward communication, by supervisors serving to ‘cover their backs’ in the event of something going wrong, is an integral component of the Safety and Production Loop (Coulson and Christofides, 2021). Although research provided evidence of increased levels of trust between supervisors and production workers, once the RRDW had been formally triggered, the dominant Safety and Production Loop was enacted. The data provided very little evidence of direct engagement between HSRs and supervisors. The rapid escalation of issues up this historical management chain provides some insight why individual mineworkers may choose not to exercise the right, given that doing so could potentially involve the highest levels of production management. Both production workers and supervisors reported that they had more experience of the RRDW being triggered by the HSR; in this case, individual workers are protected from potential repercussions by the powers of the HSR. Individual workers experience strength in numbers, given that the HSR has the power to remove all workers from a dangerous workplace. Given that the RRDW will be escalated to senior production management, these precautionary moves by workers are cogent.

Figure 8 shows in yellow the intended communication channels for worker representatives in the event of the exercise of the RRDW, as described in the recommended guidance issued by the MHSI. It shows how HSRs in the workplace are expected to escalate issues through the fulltime HSR, for eventual review at the mine health and safety committee. However, in cases where escalation of issues through the fulltime HSR may happen, the fulltime HSR could reasonably be expected to take concerns to the employer safety department, to whom they mostly report (Coulson and Christofides, 2021). This, in and of itself, need not present a problem, provided issues are resolved to the satisfaction of the worker representatives and their constituencies. Yet this is not the case. As we have previously reported, we found up to one-third of workers who had withdrawn from a dangerous workplace went back while believing it still to be dangerous (Coulson et al., 2019).

The manipulation of the RRDW that has emerged in the case of the employer and organized labour, and even workers themselves,

must mean that the second dominant communication loop on mines involving HSRs—the Compliance and Enforcement Loop (Figure 3)—is corrupted with respect to any advocacy efforts of HSR. The employer was found to hold interests in the right being exercised, as the very last control in risk management where other controls failed; organized labour was willing to use the right to advance other struggles, either to keep a mine open and producing (when closed by the MHSI for OHS violations), or to secure wage increases; workers report the right is abused. These deviations from the purpose of the right make the advancement of worker demands for OHS improvements immensely complex. Only the MHSI is in a position to enforce the intention of the MHSA, although, under the current guideline (Department of Mineral Resources, 2016), they too have stepped away from bolstering worker rights under the RRDW (Coulson et al., 2019). These concerns should be a central concern for the present study funded by the MHSC, Project CoE 200106 Impact of Implementing the Guide for the Mandatory Code of Practice on the Right to Refuse Dangerous Work (RRDW) and Leave Dangerous Working Place (RLDWP) on Occupational Health and Safety in the South African Mines, which was published in March 2024. This research expressly assesses the impact of the mandatory guideline (Department of Mineral Resources, 2016) that must include whether health and safety collective agreements signed between unions and management include an agreed procedure for the RRDW. Management has been shown to resist fulltime HSRs reporting independently to organized labour via the UHSS shop stewards (Coulson and Christofides, 2021); however, the introduction of the guideline in 2016 presented an opportunity for organized labour to reclaim the RRDW for individual workers.

Conclusion

The inclusion of supervisor data from a study on the implementation of the RRDW against the backdrop of closed communication loops, shown to severely constrain the effectiveness of worker HSRs, sheds light on why the RRDW is poorly adopted by individuals. It explains the ineffectiveness of HSRs as advocates for a safe working place. The data provided by supervisors in the RRDW Work Bid 65/12-13 study (Stewart et al., 2013) both prefigured and confirmed recent analysis (Coulson and Christofides, 2021); it justified concerns that HSRs were constrained from effectively performing in the interests of the RRDW. Workplace HSRs embodied the most practical experience of exercising the RRDW.

Figure 8—Dominant communication loops in the mine workplace and the Right to Refuse Dangerous Work [WHSR: workplace HSR]

Communication constraints in the safety system on South African mines and implications

These HSRs, however, were effectively communicatively disabled from addressing RRDW matters in the workplace and were hence not central to their resolution. In addition, the HSRs were effectively constrained and unable to escalate their concerns to trade-union safety structures. The reporting relations between workers and their workplace HSRs, and their fulltime HSRs and trade-union shop stewards—who sit on the mine safety committee and have access to the tripartite ‘Compliance and Enforcement’—is poorly institutionalized in comparison with the ‘Workplace Safety and Production’ system and its regular meetings along the managerial chain of command. Fulltime HSRs, as previously shown, were actively discouraged from reporting directly to their trade-union structures. The employer has held effective sway in dominating the OHS environment underground, notwithstanding the legislative intention of work representation in OHS. What has emerged here is that the ambivalence of supervisors, responsible for both production and safety, to worker representation contributes to the absence of systemic organizational support for the RRDW in the spirit of the MHSA. This needs urgent redress if the targets of zero harm are to be achieved in South African mining.

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Affiliation:

1Key Laboratory of Coal Processing and Efficient Utilization of Ministry of Education, China University of Mining & Technology, Xuzhou, Jiangsu, China

2School of Chemical Engineering, China University of Mining and Technology, Xuzhou, Jiangsu, China

3Department of Chemical and Biomolecular Engineering, National University of Singapore, Singapore

Correspondence to: H. Wang

Email: whfcumt@126.com

Dates:

Received: 30 Jun. 2023

Revised: 9 Jan. 2024

Accepted: 9 Jan. 2024

Published: April 2024

How to cite:

Zhang, Z., Xu, Y., Wang, H., Shi, J., Niu, J., and Zhang, Z. 2024.

Triboelectric characteristics and separation of magnesite and quartz.

Journal of the Southern African Institute of Mining and Metallurgy, vol. 124, no. 4, pp. 193–200

DOI ID:

http://dx.doi.org/10.17159/24119717/2884/2024

Triboelectric characteristics and separation of magnesite and quartz

by Z. Zhang1,3, Y. Xu1,2, H. Wang1,2, J. Shi1,2, J. Niu1,2, and Z. Zhang1,2

Abstract

Tribo-electrostatic separation is a promising method for effectively utiliing low-grade magnesite resources by removing quartz and improving the grade of magnesite. To determine the chargeto-mass ratios of pure magnesite and quartz, a triboelectric measurement system was employed, and a laboratory tribo-electrostatic separation system was used to separate low-grade magnesite. The results showed that the maximum difference in charge-to-mass ratio between pure magnesite and quartz particles was observed after friction with polyvinyl chloride. The grade and recovery of magnesite increased with temperature and initially increased with increasing voltage, feed rate, and air flowrate before gradually decreasing. The optimal conditions resulted in a concentrate with 46.91% MgO content and 77.36% recovery, while the quartz content decreased from 7.02% to 1.95%. These experimental results demonstrate the effectiveness and potential of tribo-electrostatic separation in removing quartz and upgrading magnesite.

Keywords magnesite, tribo-electrostatic separation, tribo-charging characteristics, charge-to-mass ratio

Introduction

Magnesite is a crucial raw material in the production of magnesium metal, refractory materials, magnesium chemicals, and magnesium oxide (MgO), which have extensive applications in metallurgy, aerospace, and other industries (El-Sayed, 2018; Sun et al., 2021a; Sun et al., 2023; Wang et al., 2022). As a strategic mineral, magnesite plays a vital role in the economic development of most nations, including China, where it is abundant (Al-Mallahi et al., 2020; Han et al., 2020; Hu et al., 2020; Zhu et al., 2023). Unfortunately, the over-exploitation of high-grade magnesite resources has led to depletion of highquality mines and the abandonment of numerous low-grade mines, causing both waste of magnesium resources and significant damage to the environment. This situation has become more pressing due to the increasing demand for magnesium ore (Shen et al., 2023; Sun et al., 2021b; Wang et al., 2022). Consequently, the development and utilization of low-grade magnesite resources is crucial, and it is imperative to employ advanced technology, equipment, and processing to fully utilize the available resources.

Flotation is currently considered the most effective way to separate magnesite and remove impurities (Brezáni et al., 2017; Sun et al., 2021c; Zhu et al., 2015). However, this method has strict requirements on particle size and is associated with high processing costs (Fan et al., 2021; Williams et al., 2022; Zhang et al., 2020; Zhang et al., 2023). Moreover, the wastewater and chemicals generated during the separation process lead to environmental pollution and resource waste, which is not in line with the government's green and low-carbon economic plan (Bentli et al., 2017; Wu et al., 2004). Additionally, abundant magnesite resources in the Xinjiang Uygur Autonomous Region of China cannot be utilized for flotation due to the scarcity of water (Wang et al., 2019a). In contrast, tribo-electrostatic separation technology is a dry beneficiation method that incurs low ecological costs, causes no pollution, and requires no water (Kianezhad and Raouf, 2020; Wang et al., 2017; Wang et al., 2018; Wang et al., 2019b). The resulting concentrate can be directly used for subsequent processing and has significant research and application potential in magnesite beneficiation.

Electrostatic separation is a solid sorting technique that relies on the variation in electrical properties between minerals (Dizdar et al., 2018). It encompasses a range of methods, such as dielectric separation, tribo-electrostatic separation, electrical classification, and high-gradient separation. Tribo-electrostatic separation is a type of electrostatic separation that exploits the difference in triboelectric charge between particles. The process generates charges of opposite polarity on the particles, which can be efficiently

Triboelectric

characteristics and separation of magnesite and quartz

separated by an electric field force. Tribo-electrostatic separation is a straightforward and cost-effective approach with enormous development potential and wide-ranging applications (Matsusaka et al., 2010; Mirkowska et al., 2014; Mirkowska et al., 2016). Bittner et al. (2014) described a triboelectric belt separator for beneficiation of fine minerals with an energy consumption of 1 kWh/t of materials processed.

This study investigated the triboelectric charging properties and sorting efficiency of magnesite by utilising a charge-tomass ratio (Q/m) experimental system and tribo-electrostatic separation system. The impacts of materials and temperature on the triboelectric charging properties and of operating parameters on the sorting efficiency during tribo-electrostatic separation were analyzed. These findings provide an effective dry separation method for low-grade magnesite, which holds significant importance for promoting industrial applications of triboelectric separation. Tribo-electrostatic separation also has great potential in the preparation of ultra-pure coal, desiliconization of calcium carbide slag, decarburization of fly ash, and the beneficiation of calcium carbonate, limestone, and barite. In addition, it can be used in feldspar, iron ore, bauxite, potassium fertilizer, phosphate ore sorting, and enrichment of food and feed.

Experimental

Charge-to-mass ratio test system

In this study, the Q/m is utilized to describe the charge generated after friction between the friction medium and material particles. The Q/m is defined as the ratio of charge to the mass of the charged body (nC/g). The measurement system is illustrated in Figure 1. It primarily consists of a constant-temperature and -humidity chamber, stirrer, stirring rod made of various materials, stirring cup, Faraday cylinder, and an electrometer.

Tribo-electrostatic separation system

The tribo-electrostatic separation system is illustrated in Figure 2. It consists of a Roots blower, gas storage tank, air heater, rotameter, screw feeder, bellows friction charger, high-voltage power supply, separation chamber, and an aggregator. The air-supply device, including the Roots blower and air storage tank, supplies the airflow for the entire system, which is heated to a predetermined temperature by the air heater. The material is transported to the pipeline by the screw feeder and delivered to the bellows friction charger under the influence of the airflow. The material particles are charged in the friction charger by continuous collisions between particles and between particles and the walls. The charged particles are then transferred to the separation chamber by the airflow. Under the effect of the electric force, the charged particles move to the separation chamber. The positively charged particles move to the side near the negative plate, while the negatively charged particles

move to the side near the positive plate. Finally, the product is collected by the aggregator.

Materials characterization

The magnesite sample utilized in this study was sourced from Liaoning province, China. X-ray diffraction (XRD) was used to analyze the mineral composition and determine the primary gangue minerals present. X-ray fluorescence (XRF) spectroscopy was utilized to determine the grade. Prior to the XRD and XRF analyses, the material was crushed to −74 µm. The results of the analyses are presented in Figure 3 and Table I. As per the XRF analysis, the MgO content was 45.36%, which indicates a low-grade magnesite. Based on the XRD results, the primary minerals present were magnesite and quartz, while the gangue contained a quartz content of 7.02%.

Scanning electron microscopy (SEM) and energy-dispersive spectra of the magnesite are presented in Figure 4. The SEM results indicate that magnesite and quartz were fully dissociated when the particle size was less than 74 μm. This dissociation provides a basis for the successful separation of magnesite and quartz using electrostatic separation.

X-ray fluorescence analysis results

Figure 1— Charge-to-mass ratio measurement system

Figure 2—Laboratory tribo-electrostatic separation system

Figure 3— X-ray diffraction spectrum of magnesite ore

Table I

Screw feeder

Gas storage tank

Bellows friction charger

High voltage power supply Rotameter

Roots blower

Air heater

High voltage power supply

Separation chamber and aggregator

Triboelectric characteristics and separation of magnesite and quartz

Results and Discussion

Tribo-charging characteristics of magnesite

Numerous theoretical and experimental studies have demonstrated that the triboelectric charging characteristics of particles play a crucial role in the tribo-electrostatic separation process (Dizdar et al., 2018; Wang et al., 2018; Wang et al., 2019a; Wang et al., 2019b). This study investigated the charge-transfer mechanism of magnesite and quartz during frictional charging at the microscopic level using atomic force microscopy. The results are presented in Figure 5, where Figure 5(a) illustrates that the surface potential of quartz was mostly green and relatively stable before charging, which enabled measurement of the surface potential after charging. From Figure 5(b), it can be observed that the surface potential of quartz turned blue after charging, with the blue area representing the charging area. The data in Figure 5(e) clearly show that the surface potential of quartz changed by −161 mV after charging. This confirms that charge transfer occurred on the surface of magnesite and quartz, where the surface of quartz was negatively charged due to the gain of electrons and the surface of magnesite was positively charged due

to the loss of electrons. The surface potential of magnesite is lower than that of quartz, and the surfaces of the two minerals carried opposite charges. This phenomenon is the basis for the frictional electrostatic sorting of magnesite and quartz.

The aim of the experiment was to investigate the effects of different tribocharging materials; namely, polyvinyl choride (PVC), polypropylene (PP), polyethylene (PE), and copper (Cu), on the surface potential and topography of quartz before and after charging. Figure 6 presents the results obtained. Before charging, the quartz surface potential diagram was green, indicating a stable potential. After friction and charging with the four materials, charged areas appear on the diagram. Specifically, after PP and PE particles were charged by friction with quartz, blue areas appeared on the surface potential diagram, indicating that the quartz surface gained electron charge, while the PP and PE particles lost electron charge. The surface potential of quartz decreased, and those of PP and PE particles were lower than that of quartz. After friction and charging with PVC and copper particles, red areas appeared on the quartz surface potential diagram, indicating that the surface of quartz lost electron charge, and the surfaces of PVC and copper particles gained electron charge. The surface potential of quartz rose, and the surface potentials of PVC and copper particles were higher than that of quartz. Comparison with the colour scale on the right confirmed the occurrence of charge transfer between the tribocharging materials and quartz surface.

The study also examined the Q/m variation of the particle population on a macroscopic scale. This involved investigating changes in the Q/m between magnesite and various friction materials, as well as the impact of temperature. The results provide important insights into selecting appropriate friction materials and temperatures during sorting tests. The study tested the effect of different temperatures (20–60oC) using PVC, PP, PE, and Cu as tribocharging materials.

In Figure 7, it is observed that magnesite particles became positively charged by losing electrons on their surface when triboelectrically charged with the four different friction materials. The surface potential of magnesite was lower than those of the four friction materials. At a temperature of 40°C, the Q/m of magnesite after friction with PVC, PP, PE, and Cu were 3.74 nC/g, 2.47 nC/g, 2.19 nC/g, and 1.23 nC/g, respectively. The Q/m of magnesite to PVC, PP, and PE exceeded 2.0 nC/g, whereas that to Cu was relatively low. These results suggest that PVC, PP, and PE are more suitable friction materials for magnesite separation, and the optimal temperature for the tribo-charging of magnesite is around 40°C.

Figure 4—(a) Scanning electron microscopy image of magnesite with particle size less than 74 μm. Energy-dispersive spectra of (b) Mg, (c) Si, and (d) Ca

Figure 5—Surface electric potentials of quartz and magnesite before and after charging

(a) SEM diagram (b) Mg

(c) Si (d) Ca

Triboelectric characteristics and separation of magnesite and quartz

The Q/m is affected by temperature, with the ratio gradually increasing as the temperature increases. However, the rate of change of the charge-to-mass ratio initially increased with temperature and then gradually decreased. It is important to note that the number of free electrons on the surfaces of both magnesite particles and the frictional materials increased with temperature. As the temperature increased, the amount of charge transfer between magnesite particles and frictional material increased due to enhanced frictional charging during their collision, leading to an increase in Q/m

Figure 8 shows that when quartz particles were triboelectrically charged with PVC and Cu, the surface potential became negative due to the release of electrons, and the surface potential of quartz

was higher than those of PVC and Cu. In contrast, when quartz particles were triboelectrically charged with PP and PE, their surface became positively charged due to transfer of electrons, and its surface potential became lower than those of PP and PE. At a temperature of 40°C, Q/m of quartz with PVC, PP, PE, and Cu after friction were −1.52 nC/g, 1.17 nC/g, 0.99 nC/g, and −0.91 nC/g, respectively. The change in Q/m with temperature after friction charging between quartz particles and different friction materials is similar to that observed for magnesite. As the temperature increased, Q/m continued to increase, and the rate of change of Q/m initially increased, followed by a gradual decrease.

The variation of Q/m of magnesite and quartz particles after friction with different materials is similar. However, Q/m of magnesite particles after friction was larger than that of quartz

Figure 6—Surface potentials and morphologies of quartz and four types of material particles before and after charging: (1) polyvinyl chloride, (2) polypropylene, (3) polyethylene, and (4) copper

Figure 7—Variation of charge-to-mass ratio of friction between magnesite and different materials at different temperatures

Figure 8—Variation of charge-to-mass ratio of friction between quartz and different materials at different temperatures

Triboelectric characteristics and separation of magnesite and quartz

particles. At 40oC, Q/m differences between magnesite and quartz with PVC, PP, PE, and Cu were 5.26 nC/g, 1.3 nC/g, 1.2 nC/g, and 3.2 nC/g, respectively. The order of the difference is PVC > PP > PE > Cu. Although Q/m of magnesite and quartz particles rubbed with PP and PE differed slightly, they had the same charge-toelectrode polarity after rubbing and were challenging to effectively separate under an electric field. Magnesite and quartz particles had opposite polarity to charged electrodes after friction with PVC and Cu materials, and they could move to opposite sides of the polar plate under the electric field, satisfying the requirements for friction materials. However, Q/m difference between magnesite and quartz particles and PVC after friction was significantly larger than that with Cu. Under the same electric field intensity, PVC is expected to have better separation efficiency than Cu. Despite the relatively small Q/m difference after Cu friction, its charge-to-electrode polarity meets the requirements. Additionally, Cu is stronger and more durable than PVC. Thus, Cu appears to be a promising choice for the frictional material.

Influence of temperature on tribo-electrostatic separation

We examined the impact of temperature on the tribo-electrostatic separation process by analysing the tribo-charging characteristics. The charging effect in a triboelectric charger is the key factor that influences material separation in this separation system. We thus investigated the effect of temperature on the separation of magnesite particles.

The contents of magnesia and quartz in magnesite products are used to grade them; thus, the contents of magnesia and quartz in the concentrate and recovery of magnesia were chosen as the primary evaluation criteria for the effect of magnesite electrification. Figure 9 displays that temperature has an impact on the contents of magnesium oxide and quartz in the positive and negative electrodes, as well as the recovery of magnesium oxide. Figure 9(a) reveals that as the temperature increased, the MgO content in the concentrate also increased. The growth rate initially rose with increasing temperature and then gradually declined. The temperature range of 40–50°C showed the fastest growth and the content tended to stabilize above this range. The highest MgO content of 46.56% was observed at 60°C. Conversely, the content of quartz in the concentrate continuously decreased as the temperature rose, with the minimum content of quartz being 2.33% at 60°C. The change in quartz content was opposite to the change in MgO content.

Figure 9(b) reveals that recovery of MgO increased with temperature, reaching a maximum of 66.2%. This finding, in

combination with analysis of the temperature-dependent Q/m, indicates that magnesite and quartz have relatively few free electrons on their surface at room temperature. Additionally, their surface charge barely changes after particle collision, resulting in little change to the grade and recovery of magnesite. However, an increase in temperature leads to an increase in the number of free electrons on the surface of magnesite and quartz, resulting in stronger electric field forces during the sorting process after particle collision. This enhanced sorting effect ultimately leads to an increase in the grade and recovery of magnesite in the concentrate. The grade and recovery of magnesite stabilized at 60°C.

Tribo-electrostatic separation of magnesite

Influence of sorting voltage on tribo-electrostatic separation

The separation of different electrified minerals is achieved by the change in particle movement trajectory due to their differences in particle charge and charge properties, which moves them to the two polar plates under the influence of the electric field force. The electric field force acting on the charged particles during their movement is determined by the sorting voltage, which is the distance between the two plates. In this study, we investigated the separation of magnesite under different separation voltages. Figure 10 illustrates that the MgO content in the concentrate initially increased with an increase in separation voltage, followed by a decrease, while the opposite effect was shown by quartz. The highest MgO content of 46.57% and the lowest quartz content of 2.33% were obtained at a sorting voltage of 15 kV. Additionally, Figure 10 shows that the recovery of MgO in the concentrate increased with the increase in separation voltage, up to 15 kV; however, when the voltage exceeded 15 kV, the recovery started to decrease. The maximum recovery of MgO of 72.88% was observed at 15 kV. In summary, the sorting voltage plays a crucial role in the separation of charged particles in the electric field. When the sorting voltage was less than 15 kV, the electric field force on the charged particles increases, resulting in longer lateral movement distance of particles within the electric field. This effectively separated the charged particles and improved the sorting effect. However, when the sorting voltage exceeded 15 kV, the particles with low or high charges polarized after contacting two sides of the electrode plates, leading to mixing of the charged particles and reduction in the sorting effect. Therefore, optimal sorting was achieved at a sorting voltage of 15 kV.

Figure 9—Influence of temperature on (a) grades of magnesite and quartz and (b) recovery of magnesite

Triboelectric characteristics and separation of magnesite and quartz

Influence of feed rate on tribo-electrostatic separation

The collision frequency in the friction charger, which determines the quality of sorting, is affected by the feed rate at certain airflow speeds. As depicted in Figure 11, when the feed rate was below 20 g/s, the MgO content in the concentrate continuously increased, while the quartz content decreased. Conversely, when the feed rate was above 20 g/s, the quartz content in the concentrate started to increase, while the MgO content decreased. The highest MgO content in the concentrate was 46.91% and the lowest quartz content was 1.9%. Figure 11 reveals that, when the feed rate was below 20 g/s, the recovery of MgO in the concentrate increased with the increase in feed rate, while the change in quartz content initially increased with the feed rate and then gradually decreased. When the feed rate was above 20 g/s, the recovery of MgO in the concentrate decreased with the increase in feed rate. The maximum recovery of MgO was 77.36%.

The feed rate plays a crucial role in determining the sorting effect by affecting the collision frequency in the friction charger, which is subject to a certain airflow speed. At low feed rates, the number of particles in the friction charger was small, resulting in low turbulence intensity and low frequency of particle-to-particle or particle-to-friction-charger sidewall collisions. This resulted in low tribo-charging of the particle surface, which reduced the effect of electric field force in the separation chamber, causing ineffective

separation and low concentrate grade and recovery. As the feed rate increased, the number of particles in the friction charger increased, and the gas flow rate fluctuated more, resulting in higher turbulence intensity and collision frequency. The resulting increase in particle charge and the effect of electric field force enhanced the separation process, leading to a higher concentrate grade and recovery. However, when the feed rate was too high, the collision frequency between particles and sidewalls decreased, leading to material accumulation, lower particle charge, weaker electric field force, and a decrease in the concentrate grade and recovery. Therefore, the optimal sorting effect occurred at a feed rate of 20 g/s.

Influence of airflow speed on tribo-electrostatic separation

The airflow speed plays a crucial role in determining the collision frequency, collision intensity, and movement time of charged particles in the sorting process, thereby impacting the sorting effect. Figure 12 illustrates that, for airflow velocities below 60 m3/h, the MgO content in the concentrate continuously increased, while the quartz content decreased. Conversely, when the airflow velocity was greater than 60 m3/h, the MgO content decreased and the quartz content increased. The concentrate displayed the highest MgO content (46.91%) and lowest quartz content (1.9%). Moreover, Figure 12 shows that, when the airflow velocity was less than 60 m3/h, the recovery of MgO in the concentrate increased with

Figure 10—Influence of sorting voltage on (a) grades of magnesium oxide and quartz and (b) recovery of magnesium oxide

Figure 11—Effect of feed rate on (a) grades of magnesium oxide and quartz and (b) recovery of magnesium oxide

Triboelectric characteristics and separation of magnesite and quartz

feed rate. Additionally, the MgO content increased with increasing airflow velocities. However, the recovery decreased with increasing feed rate. The maximum recovery of MgO was 77.36%.

The magnitude of the airflow velocity had a significant impact on the quality of sorting during electrostatic separation of magnesite, because it affected both the collision frequency and intensity of particles in the friction charger and the movement time of charged particles in the sorting chamber under the influence of the electric field. At low airflow velocities, the collision frequency and electrical charge of particles were low, resulting in ineffective separation due to the weak effect of the electric field force, resulting in low recovery of MgO. However, increasing the airflow velocity led to an increase in the movement speed and disorder of particles, as well as their electrical charge and the effectiveness of the electric field force, thereby enhancing the recovery of MgO. Nonetheless, if the airflow velocity exceeded a certain value, the movement time of charged particles in the sorting chamber reduced, thereby reducing the recovery of MgO. Consequently, it is essential to maintain a controlled airflow velocity to ensure optimal sorting of magnesite during electrostatic separation.

Conclusions

Based on the current results, several conclusions can be made:

(1) When PP and PE were used as the frictional materials, magnesite and quartz particles had the same charge polarity, which made it challenging to effectively separate them. However, when friction occurred between the particles and PVC or Cu, the magnesite and quartz particles acquired an opposite charge polarity to the electrode, which enabled them to move towards opposite sides of the electrode plate and achieve high-quality separation. Magnesite and quartz particles exhibited significantly different Q/m after friction with PVC, resulting in excellent separation efficiency.

(2) Increasing the temperature resulted in an increase in the number of free electrons on the surface of magnesite particles and friction material. This led to higher charge transfer, frictional charging, and Q/m during frictional collision between the particles and friction material. The change rate of Q/m initially increased with temperature and then gradually decreased.

(3) Raising the sorting voltage increased the electric field force

on the charged particles and lateral movement distance of particles in the electric field. This increase in force effectively separated the charged particles and improved the sorting effect. However, if the sorting voltage was too high, particles with a lower or higher electric charge became polarized upon contact with the plates on both sides. As a result, charged particles entered the positive and negative products without separation, leading to a reduced sorting effect.

(4) The grade and recovery of magnesite depend on the amount of charge in the friction charger and the movement time under the influence of the electric field in the sorting chamber. When the collision frequency of particles was low, the amount of charge and effect of electric field force were also low, and particles could not be effectively separated. In contrast, increasing the particle movement speed led to an increase in particle charge and strengthening by the electric field force, which enhanced the charging effect and recovery of magnesite. Therefore, to ensure effective separation of magnesite, operating parameters such as feed velocity and airflow velocity should be controlled.

Recommendations

Tribo-electrostatic separation offers various advantages, such as minimal maintenance, a completely dry process without the need for water and chemicals, a high specific feed rate, ease of operation, and environmental friendliness. The energy consumption is low, estimated to be no more than 4 kWh/t of processed materials according to the laboratory test unit.

In the next phase, we will investigate the micro-friction charging mechanism of magnesite and optimize the structure and materials of the bellows friction charger. Our aim is to enhance the separation effect by strengthening the friction charging process. Simultaneously, we are developing a new type of separation machine and accompanying devices to augment processing capacity while ensuring effective separation. The optimization efforts will progress from semi-industrialization to full industrial applications.

Acknowledgements

The authors acknowledge financial support by the Jiangsu Provincial Key R&D Plan (Social Development) Project (BE2022717) and the National Nature Science Foundation of China (51674257).

Figure 12—Effect of gas flow rate on (a) grades of magnesia and quartz and (b) recovery of magnesium oxide

Triboelectric characteristics and separation of magnesite and quartz

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Affiliation:

1Sustainable and Innovative Metals and Minerals Extraction Technology (SIMMET) Group, School of Chemical and Metallurgical Engineering, University of the Witwatersrand, Johannesburg, South Africa

2Department of Natural Science, Manchester Metropolitan University, Manchester, United Kingdom

Correspondence to: J.K. Nyembwe

Email: joseph.nyembwe@wits.ac.za

Dates:

Received: 3 May. 2023

Accepted: 15 Dec. 2023

Published: April 2024

How to cite:

Simonsen, H., Potgieter, J.H., Nyembwe, K.J., and Chuma, A. 2024. Can preconcentration of cassiterite from its pegmatite ore reduce processing costs and improve operational sustainability? Journal of the Southern African Institute of Mining and Metallurgy, vol. 124, no. 4, pp. 201–208

DOI ID:

http://dx.doi.org/10.17159/24119717/2791/2024

Can preconcentration

of cassiterite from its pegmatite ore reduce processing costs and improve operational sustainability?

by H. Simonsen1, J.H. Potgieter1,2, K.J. Nyembwe1, and A. Chuma1

Abstract

Different concentration techniques were evaluated for preconcentration of a mineral ore at a coarser size to avoid energy and resource wastage. Specifically, the aim was to reduce milling costs and energy required in the beneficiation of tin. In this study, cassiterite (0.17% Sn) was the mineral of interest in a pegmatite ore body associated with quartz (SiO₂ > 60%) and alumina (Al₂O3 > 20%).

Three concentration techniques, namely dense media, shaking table, and flotation, coupled with characterization analysis, were used to assess the concentration response. The results confirmed that particle size and mineral liberation impact the separation process. High recovery and grade were obtained with gravity concentration methods (dense media and shaking table) for coarser (+300 to +212 µm) and intermediate (+150 to +53 µm) particle sizes. Lower recovery and grade were identified for much finer sizes (−53 to −38 µm). Flotation produced a high-grade product at a relatively low recovery and appeared to be only applicable to finer grains. Separation efficiency based on Schulz’s equation measured a segregation performance of 74.8% for dense medium separation and 60.7% for the shaking table for the coarse and intermediate particle sizes. Flotation only achieved a separation of 30%–40%. The results suggest that use of dense media separation as a rough preconcentration method prior to further grinding, and the utilization of a more advanced concentration technique for mineral recovery and upgrade, constitute a successful approach to improve process economics.

Keywords cassiterite, concentration, separation, efficiency

Introduction

Tin ore accounts for a small amount as a single mineral form, but a larger share in the form of key mineral or associated components. Tin (Sn) is listed as a critical and important metal. Critical metals are considered both important to society and vulnerable to supply disruption (Whitworth et al., 2022). Tin is mainly used as a plating agent for steel cans in the food industry, piping, and in glass manufacturing.

There are more than 50 types of tin-bearing minerals on earth. Of these, cassiterite (SnO₂) is the only natural mineral from which tin is usually economically extracted (Su et al., 2017). The mineral is identified in placer deposits, granite-related tin deposits, including vein type, and stanniferous pegmatites. The latter has been reported as economically favorable at a Sn grade of approximately 0.14% (Maritz and Uludag, 2019). Extraction and beneficiation of Sn from such deposits involves several stages to achieve production of metal as the final product. The beneficiation of SnO₂ usually uses its density because the mineral has a relatively higher density than that of co-existing minerals within the ore body assemblage. Traditionally, the mineral concentration flowsheets consist of gravity separation units, including jigs, spirals, and tables, for the primary concentration stage. The final stage typically utilizes enhanced gravity concentrators or flotation cells (Angadi et al., 2015). Thus, the process involves multi-stage grinding and multi-stage beneficiation.

Most minerals of interest are finely disseminated and intimately associated with gangue, so the need arises to unlock or liberate the mineral before separation can be undertaken. This is achieved by comminution, during which the particle size of the ore is progressively reduced until the liberated particles of mineral can be adequately separated (Wills and Napier-Munn, 2006). This is an energy-intensive process, which, in the mineral raw material sector, accounts for 50% or more of the energy consumption of a mine site. Cassiterite is characterized by extreme brittleness compared with other minerals usually present (Parapari, 2021). This attribute should be considered during size reduction prior to concentration. In production practice, −38 µm cassiterite exhibits very poor gravity separation and effective treatment of −10 µm cassiterite presents difficulties (Fu et al., 2023). It appears advantageous to recover cassiterite grains, where possible, at the earliest possible stage and at their largest size to avoid softness that is difficult to handle and recover (Ibrahim et al., 2022). Therefore, upgrading of mineral concentration of ores before fine comminution can decrease the capital expenditure and energy consumption of mining operations and

Can preconcentration of cassiterite from its pegmatite ore reduce processing costs

simplify environmental permitting (Leon et al., 2020). In this sense, targeted mineral preconcentration would, to some extent, discard gangue-related minerals at a coarse size to save on the energy required to further process barren waste material. In addition, mineral preconcentration could avoid generation of fines. This study explored the feasibility of preconcentration of SnO₂ from a pegmatite ore body at an early stage of the concentration process to discard/eliminate barren waste minerals, with the purpose of saving costs and improving operational sustainability of the process.

It is hypothesized that it will be possible to remove a large portion of the waste mineral at the front end of the process and that this will lead to substantial cost saving. The investigation employed characterization techniques, mainly bulk chemistry (X-ray fluorescence; XRF), to assess the liberation response of the mineral of interest (SnO₂) and the efficiency of different concentration methods, i.e., gravity separation (dense medium separation (DMS) and shaking table) and flotation.

Material and Methods

Materials and mineral validation

A 20 kg ore sample was obtained from the Uis deposit in the Erongo region of Namibia, which is currently operated by Andrada Mining. Homogenization was performed in accordance with the soil sampling protocol of the U.S. Environmental Protection Agency. The sample was run-of-mine material that had passed through an initial crushing operation and was collected from the silo feeding the mill for further size reduction at an estimated average grain size of > 300 µm (Figure 1(b)) and relative density of 2.662 g cm ³. The sample was characterized for its bulk chemistry and mineral composition using XRF and X-ray diffraction (XRD) analyses, respectively. Its composition revealed a Sn content of 0.17%, with

Si and Al as the major elements, at 30% and 12%, respectively (Figure 1(c) and (d)). Mineral quantification showed the presence of cassiterite (SnO₂), quartz (SiO₂), albite, and magnetite (Fe3O₄) at proportions of 0.27%, 75%, 24%, and 7.1%, respectively (Figure 1(e)). The tin content varied with screen size (Figure 1(f)). The as-received sample microscopically displayed the presence of free cassiterite grains and relatively coarse silicate mineral (Figure 1(g)), suggesting the possibility of tin-related minerals preconcentration, thus allowing silicate to be discarded at a coarse size. Attributes of the mineral phases are given in Table I.

Methods

A stepwise method was used during this investigation, with the aim of discarding barren material at a coarser size. A wet comminution (liberation) exercise was conducted at various resolved times (5, 10, 15, 20 min), using a laboratory rod mill at a volumetric charge of 40%, comprising rods and sample in a mass ratio of 0.25. The equipment was set at a rotation speed of 60 rpm. The as-received sample and each milled product were screened using sizes of 300, 212, 150, 106, 75, 53, and 38 µm. Three classes of particle sizes (coarse (+300 to +200 µm), intermediate (+150 to +53 µm), and fines (−53 to −38 µm)) were obtained. This group size segregation intended to process particles of relatively similar size and allowed the performance of the concentration process to be adequately assessed.

Differences in mineral properties of the ore assemblage were used to concentrate the SnO₂. Two gravity separation techniques were used: DMS and a Wilfley shaking table. Both methods make use of the relative density (RD) difference between the various

Figure 1—Sn ore initial characterization. (a) Particle size of as-received ore, (b) trace and (c) major elements,

Can preconcentration of cassiterite from its pegmatite ore reduce processing costs

Table I

Physical attributes and quantification of mineral phases

RD: relative density., Mhos hardness (%): percentage content

minerals in the ore. DMS was conducted using tetrabromoethane (C₂H₂Br₄ (2.97 g cm ³)) dissolved in acetone to obtain a final cut density of 2.75 g cm ³. The process isolated the sample into two fractioned products, referred to as sinks (heavy, RD > 3 g cm ³) and floats (light, RD < 2.7–2.65 g cm ³), using a 1 L conical separatory funnel at density cut of 2.97 g cm ³. The shaking table was set at an amplitude of 35 Hz with an inclination angle of 6° at a water flowrate of 500 L/h and feed rate of 50 kg/h. Three products were obtained from the shaking table: concentrates (heavy), tails (light), and middlings (mix of tails and residual concentrate).

Flotation was conducted using a laboratory cell (Denver) with oleic acid (collector, 600 g/t), sodium silicate (depressant, 300 g/t), Dewforth 40 (frother,t 20g/t), and H₂SO₄ as the pH regulator (maintained at pH 4.5–4.8). The device was agitated at 1126 rpm. The feed comprised a wide array of particle sizes (Figure 1(b) to highlight the importance of a specific size suitable to positively respond to the concentration process.

The efficiency of separation for all three techniques was evaluated. Bulk elemental chemistry was determined using an XRF powder method (Rigaku-ZSX Primus II, SQX analysis software, operating at 4 kW, 60 kV, and 150 mA). Mineral contents were determined from XRD analysis (Rigaku Ultima IV, PDXL analysis software), using a Cu Kα radiation source at 30 kV and 25 mA. Data were recorded over the range 5° ≤ 2θ ≤ 95°. The powdered samples were scanned at 0.5°/min with a step of 0.01°. Density was measured using a gas-displacement pycnometer (ACCUPY II 1345, Micromeritics).

Results and discussion

Effect of time on comminution and SnO2 liberation

Figure 2 summarizes the comminution characteristics of the SnO₂ sample at selected retention times. Figure 2(a) shows the particle size distribution; Figure 2(b) summarizes the comminution extent of the ore assemblage at 5, 10, 15, and 20 min. A decrease in content of larger particles (+300 µm) from 88% to 40% after 20 min milling was observed. The content of finer particles (< 38 µm) did not

significantly increase (Figure 2(b)). This could suggest that further grinding had little effect on comminution characteristics. However, the presence of fines particle (< 38 μm) had a significant effect on the redistribution of intermediate sizes quantities, including the 150, 106, 75, and 53 µm size fractions. Figure 2(a) indicates that the longest retained time (20 min) had the largest content of intermediate sizes, with the same relative content of fines. This result agreed with that of Yianatos et al. (2005), who reported a decrease in the content of the 212 µm size fraction by 3%, due to the presence of fines. The finer the particles produced in comminution, the greater is the amount of energy required to effect breakage (Stamboliadis, 2013).

Dense media concentrate characterization

Table II summarizes the separation results obtained from gravity concentration using the dense media with tetrabromoethane at a density cut of 2.75 g cm ³. The heavy portion (sink) is characterized with regards to recovery, Sn grade, and density of the various size fractions. In all cases, the results displayed the possibility for preconcentration. Although possible to preconcentrate the as-received sample, poor recovery (68%) and low Sn grade are reported, suggesting that rapid mineral liberation should be required (Table II). The comminution time-of-grind dictated the concentration achieved. A short grind time was characterized by coarse particles and high concentrate grade, accompanied by a low recovery. Grinding conducted for 5, 10, 15, and 20 min gave recoveries of 72%, 74%, 74%, and 73% at Sn grades of 0.63%, 0.79%, 0.81%, and 0.66%, respectively (Table II (+300 to +212 µm)). Improved SnO₂ concentration was identified for the intermediate size (+150 to +53 µm) with recoveries of 78%, 88%, 79%, 83%, and 80% at Sn grades of 1.6%, 2.9%, 5.3%, 6.1%, and 5%, for the asreceived sample and resolved grind times, respectively. This increase in Sn grade could be attributed to size reduction that resulted in SnO₂ liberation and redistribution among the particle sizes, ranging from −150 to +53 µm. Low concentrate recoveries are observed for relatively fine grains (−53 to −38 µm), which could be attributed to loss of SnO₂ fines due to the brittle nature of the mineral.

Figure 2—Tin ore comminution. (a) Particle size distribution and (b) coarse and fine particle size evolution as a function of time

Can preconcentration of cassiterite from its pegmatite ore reduce processing costs

Table II

Dense media preconcentration characteristics Size

Table III

Shaking table preconcentration characterization

These results agree with those of Angadi et al. (2017), who reported that Sn can be selectively upgraded using heavy-liquid separation of coarse particles at a density cut of 2.75 g cm ³ over silicate minerals, including quartz (2.4−2.65 g cm ³) and albite (2.61−2.64 g cm ³). For instance, for 15 min retention time, the Sn content of the coarse products (+300, +212 µm) was upgraded from 0.14% in the feed to 0.81%, at a recovery of 88.4%. The intermediate size fraction was upgraded to 6.1% SnO₂ at a recovery of 83%.

The efficiency of gravity separation appeared to decrease due to the presence of smaller particles, as reported by He et al. (2021). Lu et al. (2021) reported that gravity separation is a poor choice for

particle sizes < 40 µm, suggesting that all particles < 38 µm were not recovered and caused the low recoveries recorded for the fines. The observed upgrade further suggests that between 11.5% and 15.6% of the barren mineral can be removed at coarser and intermediate size distributions by selective size distribution. The fine particles would require a more advanced concentration technique for successful beneficiation.

Tabling concentrate characterization

Table III summarizes the concentration results for the various size fractions obtained during the tabling process and the respective Sn

Can preconcentration of cassiterite from its pegmatite ore reduce processing costs

grades. The tabling results from all feed fractions exhibited lower recovery than those of DMS using tetrabromoethane. This could be due to the presence and formation of a third “middling” product stream obtained from the shaking table concentration process.

Recovery appeared to decrease with decreasing particle size. At the coarser size, high recovery was observed for the short milling retention time (5 min); recovery decreased with an increase in milling time. The grade was inversely proportional to the recovery and increased with milling retention time. Although the tabling recoveries were lower than those of DMS, this technique produced a relatively higher-grade concentrate.

Lower recoveries were observed for the intermediate size fractions (+150 to +53 µm) and fines (−53 to −38 μm). In addition, these two size fractions contained lower SnO₂ grades than the respective values reported from the DMS process. The lower grade and recovery observed for the tabling process compared with DMS could be attributed to the wide array of particle sizes, mainly the intermediate group, which varied from +150 µm to +53 µm. The fine particle size fraction showed high losses due to the presence of a small ultrafine particle size fraction (−38 µm). In addition, the decrease could be attributed to the presence of the additional middling product stream.

Flotation concentrate characterization

Table IV summarizes the concentration results obtained for the flotation process in terms of the recovery and Sn grade for the two size fractions. The results show that particle size has an important role during the mineral upgrade. This method was ineffective at a coarse particle size, which is probably due to the resultant poor grain hydrophobicity. Literature further classifies oleic acid as a poor collector (Jin et al., 2021), which could have led to the low (< 40%) recovery.

Similar to the coarse size fraction, the intermediate and fine fractions were characterized by low recoveries accompanied by a fairly high Sn grade. However, the recovery and grade improvements were identified as related to the milling retention time, which was linked to the liberation of the targeted mineral (SnO₂) and material redistribution between particle sizes that were able to float easily. In this case, the poor recovery and grade reported for the sample retained for 5 min milling could be associated with the presence of relatively coarse grains (+300 µm) because most (68%) of the milled sample was within that particle size range (Figure 2(a). The improved recovery and grade observed for the longer retention time could indicate efficiency of flotation concentration. The obtained results support those of Zhang et

Preconcentration by flotation

Table IV

Figure 3—Mineral characterization comparison of feed and concentrate products for (a) dense media separation, (b) shaking table and (c) flotation.

Q: quartz, Al: albite, Cs: cassiterite, Mg: magnetite

Can preconcentration of cassiterite from its pegmatite ore reduce processing costs

Table V

Mineral content variation (15 min retained

by concentration process

Concentration process separation efficiency

Separation efficiency (%)

al. (2021), who reported that flotation was the most suitable concentration method for mineral sizes of −75+38 µm. Flotation process efficiency decreases for fine (−20 µm) and ultrafine (−10 µm) particle sizes (Leistner et al., 2016). These authors recommended an agglomeration operation prior to concentration using flotation.

Separation efficiency and X-ray diffraction characterization

Figure 3 summarizes the mineral contents of the concentrate obtained from each technique, and quantitatively and qualitatively compares the mineral composition of each concentrate product with its corresponding feed sample (Table V). Figure 3 and Table V display the mineral composition of the 15 min milled sample subjected to DMS, shaking table, and flotation separation, respectively. In all cases, the products showed an increase in Sn grade, indicated by XRF results (Tables II to IV) and a corresponding decrease in gangue minerals (silicates/quartz). These recovery and grade data can be used to further assess and evaluate the concentration process. Separation efficiency (Table VI), which is a measure of the metallurgical efficiency, is defined as the recovery difference between the valuable and gangue minerals to the concentrate (Can et al., 2019). This concept is best described by Equation (1) (Schulz, 1970):

where SE is separation efficiency (%), R is recovery of the valuable constituent/metal/mineral in the concentrate (%), cm is the assayed element grade in the mineral being concentrated (%), C is assayed

element grade in the concentrate (%), and f is assayed element grade in the feed (%).

At the coarser and intermediate sizes, both concentration techniques (DMS and shaking table) can successfully be applied for mineral preconcentration. DMS appeared to have far better efficiency than the tabling operation and is recommended as the main preconcentration step. For instance, the coarser size after 15 min milled retention time showed separation efficiencies of 74.13% and 59.89%, respectively, for DMS and the shaking table. For the intermediate sizes at the same retention time (15 min), separation efficiencies of 82.7% and 68.5%, respectively, were obtained. A decrease in separation was observed for the fines, with both techniques yielding < 40%, suggesting that particle size impedes the performance of these techniques.

Flotation, in contrast, showed the worst separation of only 37% for the intermediate fraction at a grade of 3% (15 min milled retention time); the fines had much higher grade (8%) at a separation efficiency of 40%. The poor efficiency of this method could suggest that an adequate size distribution for mineral liberation should be determined prior to the concentration process, because this technique recovered a concentrate grade of 8%. In addition, fines would require some sort of agglomeration prior to flotation to enhance the process and avoid Sn losses.

Conclusions

The preconcentration of SnO₂ from a pegmatite ore body was investigated. Three concentration techniques, namely dense media, shaking table, and flotation, were assessed and the results compared. The results indicated that particle size distribution

Table VI

Can preconcentration of cassiterite from its pegmatite ore reduce processing costs

dictates selection of the concentration process. DMS was effective at concentrating the coarse (+300 and +212 µm) and intermediate (+150 to +53 µm) particle sizes. The calculated Schultz separation efficiencies were 74.1% and 82.7% for the coarse and intermediate particle size fractions, respectively. By installing an additional DMS stage prior to further size reduction, the operation could allow discard of 10% to 15% of gangue-related minerals, which would contribute to energy saving and improved process sustainability. Flotation gave low recovery and could only be successfully applied to concentrate the fine sizes. The results suggest the use of DMS as the preconcentration technique for both coarse and intermediate particle sizes prior to further grinding, and application of flotation to the fine particle sizes to provide a mineral upgrade procedure to optimize the process in terms of comminution efficiency.

References

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Angadi, S.I., Eswaraiah, C., Ho-Seok, J., Mishra, B.K., and Miller, J.D. 2017. Selection of gravity separators for the beneficiation of the Uljin tin ore. Mineral Processing and Extractive Metallurgy Review, vol. 38, no. 1, pp. 54–61. doi: 10.1080/08827508.2016.1262856

Can, İ.B., Özsoy, B., and ErgÜn, Ş.L. 2019. Developing an optimum beneficiation route for a low-grade chromite ore. Physicochemical Problems of Mineral Processing, vol. 55, no. 4, pp. 865–878. doi: 10.5277/ppmp19006

Fu, L., Li, W., Pan, Z., Zhang, Z., Jiao, F., and Qin, W. 2023. Synthesis of modified polystyrene nanoparticles and their application in fine cassiterite flotation. Colloids and Surfaces A: Physicochemical and Engineering Aspects, vol. 681, p. 132608.

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5TH SCHOOL ON MANGANESE FERROALLOY PRODUCTION

THEME: DECARBONIZATION OF THE MANGANESE FERROALLOY INDUSTRY

3-4 JULY 2024 - CONFERENCE 5 JULY 2024 - TECHNICAL VISIT

Venue: Boardwalk ICC, Gqeberha, Eastern Cape, South Africa

The energy-intensive industry of manganese (Mn) ferroalloy production is one of the largest producers of direct carbon emissions. The demand for ferromanganese alloy, an additive of steel, follows the demand for steel and continues to increase. This means Mn ore and the production of Mn ferroalloys form an integral part in a future energy sector based on renewable energy technologies. Major structural features of wind turbines, solar panels, and energy storage devices are all made of steel components.

The current industrial practices and state-of-the-art in Mn ferroalloy production are heavily dependent on the use of fossil-based carbon. The 5th SAIMM school on Mn ferroalloy production thus aims to bring together industry and research in order to support smelters and foster collaborations between researchers towards adopting the transition of decarbonizing the ferroalloy industry. The conversation around the topic will shed light on some of the fundamentals and industrial integration of the various decarbonization strategies.

OBJECTIVE

• To create a platform to discuss the environmental impact of carbon emissions from the production of Mn ferroalloys.

• To provide the opportunity for industrialists and researchers to exchange views on the decarbonization of the Mn ferroalloy industry.

• To further enhance collaborations between parties.

TARGET AUDIENCE

• Local and international delegates from the Mn ferroalloy industry or those who support them.

• Existing and potential future industry role players.

• Engineering companies.

• Research/academic institutions.

• Companies providing funding for new Mn projects.

TOPICS

• Commercial production processes and overview of Mn production in South Africa and Europe, including potential new projects.

• Process fundamentals on thermodynamics, slag fundamentals, and reaction kinetics based on various decarbonization strategies.

• CO2 reduction programs in the industry, CO2 capture and energy recovery, Bio-carbon, H2 and, solar energy.

Contact: Gugu Charlie, Conferences and Events Coordinator

E-mail: gugu@saimm.co.za | Tel: +27 11 538-0238 Web: www.saimm.co.za

ECSA Validated CPD Activity, Credits = 0.1 points per hour attended.

Affiliation:

1Independent Consultant, Pretoria, South Africa

Correspondence to: N.C. Steenkamp

Email: ncs.contract@gmail.com

Dates:

Received: 30 Mar. 2022

Accepted: 10 Feb. 2024

Published: April 2024

How to cite:

N.C. Steenkamp. 2024.

Open-pit post-blast dust cloud lightning. Journal of the Southern African Institute of Mining and Metallurgy, vol. 124, no. 4, pp. 209–212

DOI ID:

http://dx.doi.org/10.17159/24119717/2071/2024

ORCID:

N.C. Steenkamp http://orcid.org/0000-0002-9912-6038

Open-pit post-blast dust cloud lightning

by N.C. Steenkamp1

Abstract

Lightning has been observed in dust clouds following open-pit blasting. It is proposed that the occurrence of this phenomenon is related to the physical composition and characterization of the mineral fines that comprise the bulk of the dust cloud. Silicate minerals, which are susceptible to fine fracturing during blasting, generate the initial charge. This is further enhanced by collisions and friction during the turbulent upcast in the post-blast dust cloud. Varying size fractions result in different rates at which these particles drop out of suspension and create a secondary temperature gradient. This results in the creation of two charged zones and increases the potential of developing a discharge in the form of lightning. The ideal prevailing meteorological conditions need to be windless to minimize the effect of particle dissipation and have sufficient moisture to enhance the potential of generating lightning.

Keywords open pit, blast, dust, cloud, lightning

Introduction

Open-pit blasting results in a dust cloud that is suspended over the area for several minutes. The phenomenon of lightning occurring within these dust clouds has been reported, but not studied or explained. The lightning is only visible for a short period after the initial blast until the dust cloud starts dissipating. Occurrences of lightning in volcanic pyroclastic clouds have been noted over a long historic period. Proposed mechanisms for the creation of lightning within these pyroclastic clouds have been more extensively studied and are used here as a proxy to interpret the formation of lightning in post-blast dust clouds. Lightning has also been observed in fire-induced pyro-cumulus clouds. The instances of post-blast dust cloud lightning are irregular. No instances of damage to mine infrastructure because of these lightning strikes have been reported to date.

Pyroclastic cloud lightning

During the eruption of several volcanoes, lightning was observed in the resulting ash cloud, referred to as a pyroclastic cloud. Dynamic interactions of ash particles within an eruptive plume result in electrical charging and charge separation within the plume, causing the formation of lightning strikes. The plume height plays a role in the potential for volcanic lightning: ash plumes between 7–12 km high concentrate water vapour, which may contribute to lightning activity; smaller ash plumes of 1–4 km have been shown to gain more electric charge from fragmentation of rocks near the vent of the volcano (McNutt, 2008). During eruptive volcanism, large amounts of ash, dust, rock, volatile gases, and lava are expelled in a very short period.

The conditions required for the formation of lightning in volcanic eruptive ash clouds have been summarized by Kaufman (2020). Volcanic ash develops an electric charge due to friction and collisions. This is the process of charge generation by break-up of rock particles. This may create a significant source of charge near the erupting vent. Ionic charges develop due to their difference in mass, resulting in different rates at which the particles drop out of suspension from the ash cloud.

These charges then need to be separated into different regions of the volcano ash plume. In a chaotic plume, this happens naturally as different sized ash particles fall at different velocities, and cause resultant development of a temperature gradient (Siegel, 2018). This creates different zones of charged particles, which can be positively or negatively charged. Two regions of oppositely charged particles develop. The

Open-pit post-blast dust cloud lightning

space between them becomes an electric field, which allows electricity to discharge through the air (Siegel, 2018). The maximum currents reached by volcanic lightning are 2 kA and they have an average appearance of 0.1–8 ms; temperatures as high as 12 000–28 000 K are reached by volcanic lightning in the discharge centre (Aizawa et al., 2016; Genareau et al., 2015; Wadsworth et al., 2017).

The electric charge distribution follows the Positive-NegativePositive (PNP) model proposed by Miura et al. (2002), in which the upper part of the ash cloud is dominated by volcanic gas and aerosols, and has a prevalent positive charge. In contrast, the middle part of the ash cloud mainly comprises negatively charged fine ash particles. The lower part of the ash cloud is dominated by gravitational settling of coarser and positively charged ash particles (Miura et al., 2002).

The fragments are further settled at different rates because of size. In the case of a bimodal size distribution, larger and positively charged particles are confined to the centre of the flow, while smaller and negatively charged particles follow the turbulence in the shear layer with the surrounding atmosphere (Miura et al., 2002). In the case of monodispersed particles, clustering of particles with either prevalent positive or negative charge generates transient electrical dipoles (Miura et al., 2002).

Other factors, such as eruption type and magma composition, do not seem to influence the occurrence of lightning (McNutt & Williams, 2010). It is also suggested that the atmospheric temperature plays a role in the formation of lightning: colder ambient temperatures promote freezing and ice charging within the plume, which leads to more electrical activity (Aplin et al., 2016).

Modification of ash particles by volcanic lightning strikes has been noted. These are referred to as lightning-induced volcanic spherules (Genareau et al., 2015). Four types of ash particles have been distinguished (Mueller et al., 2018):

Type I – particles that did not undergo lightning-interaction;

Type II – partially melted particles, comprising grains that did interact with the lightning strike, but not sufficiently to melt the entire grain, displayed by remains of a pristine surface or irregular shape of the particle;

Type III – Completely melted particles, which are entirely melted and typically have a spherical shape;

Type IV – Aggregates, which are clusters of up to four amalgamated particles of different types.

A pronounced loss of volatile elements in lightning-affected particles has been noted, showing clear depletion of Na < P and S < Cl < F (Keller, 2020).

Charge generation

The processes regulating the electrification of granular material flows are the composition, size, and kinetics of the solid particles, in conjunction with the ambient conditions (Cimarelli and Genareau, 2022). The main mechanism for generation of a charge during volcanism—and that can be applied to blasting—is fractoelectrification. This process has been observed during fracturing of crystals, rocks, glass, and materials such as metal and ice (Cimarelli and Genareau, 2022). Depending on the type of strained material, fracturing can promote the release of electrons, positive ions, neutral atoms, and electromagnetic radiation, promoting charging of the resulting fragmented particles (Cimarelli and Genareau, 2022).

This phenomenon has been explained in terms of the piezoelectric nature of certain substances, such as quartz, which enhance charge separation at the tip of a propagating fracture and

thereby generate large electric fields (Cimarelli and Genareau, 2022). Experiments conducted on pumice clasts that were forced to collide and fracture under vacuum conditions indicated that a net negative charge is held on the solid silicate particles and a net positive charge is either released as ions or carried on a very small proportion of fine particles generated upon collision (Cimarelli and Genareau, 2022). Tribo-electrification (or contact electrification) is the phenomenon of charging by collision and friction between bodies. Eruptive rock fragments that display a high level of heterogeneity in terms of chemical composition and physical characteristics, such as grain size, density, and grain shape, create a favourable environment for charging and redistribution of charge by the collision of these different particles (Cimarelli and Genareau, 2022).

The coexistence of supercooled water droplets and ice crystals in the expanding saturated plume suggests that the mechanism of hydrometeor interaction could be active in volcanic eruptions. Ash particles in the volcanic plume may act as ice nuclei once sufficient altitudes are achieved, promoting ice–ash charging. The eruption column must reach a height coinciding with the local −10°C or −20°C isotherm to achieve volcanogenic ice nucleation (Cimarelli and Genareau, 2022).

Case Study

Lightning has been observed in post-blast dust clouds above the open pits of the iron-ore mines between the towns of Kathu and Postmasburg, in the Northern Cape province of South Africa. The phenomena are noticed more often during the spring and early summer months.

Climate

The climate of the Kalahari in South Africa is described as semidesert. According to the Köppen and Geiger classification, the Kathu area is classified as BSh. Summer occurs from December to March and winter from June to September. Rainfall mainly occurs during the late summer and autumn months. The months with the highest and lowest relative humidities are April (49.82%) and October (25.41%), respectively (Climate-data.org). Cloud cover varies with the seasons, with least cloud cover during July and most cloud cover during October: overcast weather is present between September and April (weatherspark.com). Most wind is experienced in the period from July to December, dominated by northernly winds, with calmer conditions from January to June (weatherspark. com).

Geological setting and ore mineralogy

The iron-ore mines are developed in the Asbesheuwels Subgroup iron formation on the Maremane Dome, between Sishen and Postmasburg (Smith and Beukes, 2016). High-grade banded-iron formation (BIF)-hosted (> 60 mass%) iron hematite ore deposits developed mainly by supergene and/or hydrothermal leaching of silica from the iron formation host rock under oxidizing conditions. High-grade hematite iron ores only develop in areas where unconformity transects the BIF (Smith and Beukes, 2016). The BIF is a sedimentary deposit that consists of alternating thin layers of iron oxides, mainly hematite and magnetite with silicate chert and subordinate shale.

Open-pit blasting

The open-cast mining method is employed by the iron ore mines in this region. This requires regular drilling and blasting of waste lithologies and ore in the pit. This method fragments the ore and liberates large volumes of dust.

Open-pit post-blast dust cloud lightning

Huang et al. (2019) modelled the post-blast dust particles, showing that these settle at different rates depending on their size. The study found that gases rapidly propel the blast fragments into the air and the dust particles then settle from the dust cloud. The rate of settling of the dust cloud is also affected by the prevailing wind velocity, which causes diffusion over a large surface area. It was found that dust with a particle size of approximately 250 μm reached the ground after about 30 s. Particles with a size range of 60–100 μm settled slowly; those with a particle size below 40 μm settled with difficulty due to disturbance by air flow.

Post-blast dust cloud lightning

It is suggested that the lightning observed in post-blast dust clouds results from a combination of the mineralogical composition of the ore, the fragment sizes and shapes of the suspended dust, and the prevailing weather conditions at the time of the blast.

These mines mainly comprise contrasting ore and gangue minerals, i.e., iron oxides and silicate oxides. Blast fragmentation of the ore produces a dust consisting of the ore mineral fines, driven turbulently upwards by hot blast gases. The resulting fines tend to produce angular fragments, which collide and rub against each other during the upward turbulent movement, resulting in fracto-electrification of the dust particles. The mass difference between the silicate and iron oxide minerals, along with their size difference, results in these particles dropping out of suspension at different rates from the dust cloud. The lighter silicate (chert) tends to develop piezoelectric charges along the sharp edges of the fragments.

It is suggested that the heavier iron oxide and large particles will fall out of suspension faster along the centre of the blast cloud, while the finer material and preferential chert fines fragments will dissipate towards the top and edges of the dust cloud. It is also reasonable to suggest that the post-blast heat-retention capacity difference between the iron oxides and silicates will be sufficient to create a notable temperature gradient.

The more-frequent observation of post-blast dust cloud lightning in the spring and summer months suggests that a minimum moisture content is required and the ambient air temperature needs to be moderate (relative to the extreme cold of winter and heat of summer in the Kalahari) to increase the potential for lightning to develop. The short-lived nature of post-blast dust clouds also preferentially requires a wind-free day to ensure that the dust cloud does not dissipate too quickly.

If all these conditions are favourably combined, a large enough opposite ionic charge difference may develop between the top, centre, and bottom of the post-blast dust cloud, resulting in the development of localized lightning strikes (Figure 1).

Conclusion

The occurrence of post-blast dust cloud lightning is suggested to occur as the result of a combination of three factors. The first is the composition of the mineral species liberated by the blast: the more susceptible the mineral is to generating a charge due to friction, the more likely it is to develop a sufficiently high charge to induce a lightning strike. The second is the size and shape of the fines that are suspended in the air. The third factor is the prevailing meteorological conditions, specifically relating to the presence of moisture in the atmosphere at the altitude of the dust cloud and a lack of wind.

Future work

Future work will focus on collecting settling dust samples from post-blast dust clouds when lightning was observed. The collected samples will be studied to determine the presence or absence of lightning-induced spherules using scanning electron microscopy. The prevailing meteorological conditions on the days that post-blast dust cloud lightning is observed need to be recorded in more detail.

Acknowledgement

The editorial inputs of an anonymous review of the first draft are acknowledged.

References

Aizawa, K., Cimarelli, C., Alatorre-Ibargüengoitia, M.A., Yokoo, A., Dingwell, D.B., and Iguchi, M. 2016. Physical properties of volcanic lightning: Constraints from magnetotelluric and video observations at Sakurajima volcano, Japan. Earth and Planetary Science Letters, vol. 444, pp. 45–55.

Aplin , K.l., Bennett, A.J., Harrison, R.G., and Houghton, I.M.P. 2016. Electrostatics and in situ sampling of volcanic plumes. Volcanic Ash, Elsevier. pp. 99–113.

Cimarelli, C. and Genareau, K. 2022. A review of volcanic electrification of the atmosphere and volcanic lightning. Journal of Volcanology and Geothermal Research, vol. 422, pp. 9–17. Climate-data.org. 2022. https://en.climate-data.org/africa/southafrica/northern-cape/kathu-27075/ [accessed March 2022]

Genareau, K., Wardman, J.B., Wilson, T.M., Mcnutt, S.R., and Izbejov, P. 2015. Lightning-induced volcanic spherules. Geology, vol. 43, no. 4, pp. 319–322.

Huang , Z., Ge, S., Ling, D., and Yang, L. 2019. Numerical simulation of blasting dust pollution in open-pit mines. Applied Ecology and Environmental Research, vol. 17, pp. 10313–10333.

Figure 1—Turbulent post-blast cloud. (a) Heavy fines fragments drop from suspension in the centre of dust cloud and lighter fines fragments dissipate to the edges. (b) Development of positive centre and negative edge charge zones (modified after Cimarelli and Genareau, 2022) (a)

Open-pit post-blast dust cloud lightning

Kaufmann, M. 2020. Why the Taal volcano’s eruption created so much lightning. https://mashable.com/article/volcanolightning-why-taal-eruption [accessed March 2022]

Keller, F. 2020. Volcanic lightning: Impacts on plume-suspended ash particles. https://blogs.egu.eu/divisions/gmpv/2020/06/18/ volcanic-lightning-impacts-on-plume-suspended-ash-particles/ [accessed March 2022]

McNutt, S.R. 2008. Volcanic lightning: global observations and constraints on source mechanisms. Bulletin of Volcanology, vol. 72, no. 10, pp. 1153–1167.

McNutt, S.R. and Williams, E.R. 2010. Volcanic lightning: global observations and constraints on source mechanisms. Bulletin of Volcanology, vol. 72, no. 10, pp. 1153–1167.

Mueller, S.P., Helo, C., Keller, F., Taddeucci, J., and Astro, J.M. 2018. First experimental observations on melting and chemical modification of volcanic ash during lightning interaction. Scientific Reports, vol. 8, no. 1, pp. 1–9.

Miura, T., Koyaguchi, T., and Tanaka, Y. 2002. Measurements of electric charge distribution in volcanic plumes at Sakurajima volcano, Japan. Bulletin of Volcanology, vol. 64, no. 2, pp. 75–93.

Siegel, E. 2018. How do volcanoes make lightning? https://www. forbes.com/sites/startswithabang/2018/02/09/how-dovolcanoes-make-lightning/?sh=3a2e6df34cac [accessed March 2022]

Smith, A.J.B. and Beukes, N.J. 2016. Palaeoproterozoic banded iron formation hosted high-grade hematite iron ore deposits of the Transvaal Supergroup, South Africa. Episodes, vol. 39, pp. 269–284.

Wadsworth, F.B., Vasseur, J., Llewellin, E.W., Genareau, K., Ciamrelli, C., and Dingwell, D.B. 2017. Size limits for rounding of volcanic ash particles heated by lightning. Journal of Geophysical Research: Solid Earth, vol. 122, no. 3, pp. 1977–1989. Weatherspark.com. https://weatherspark.com/y/89144/AverageWeather-in-Kathu-South-Africa-Year-Round [accessed March 2022] u

E-mail: camielah@saimm.co.za

Tel: +27 538 0237

Web: www.saimm.co.za

Affiliation:

1Department of Mining Engineering, Lorestan University, Khorramabad, Iran

Correspondence to: K. Barani

Email: Barani.k@lu.ac.ir

Dates:

Received: 28 Jun. 2022

Accepted: 10 Feb. 2024

Published: April 2024

How to cite:

Ganji, S.M.S. and Barani, K. 2024. Leaching of gold ore using creatine monohydrate. Journal of the Southern African Institute of Mining and Metallurgy, vol. 124, no. 4, pp. 213–218

DOI ID:

http://dx.doi.org/10.17159/24119717/2179/2024

Leaching of gold ore using creatine monohydrate

by S.M.S. Ganji1 and K. Barani1

Abstract

Creatine was used as a lixiviant for gold dissolution from gold ore. The effects of creatine concentration, temperature, leach time, and pH were examined for their influence on extent of gold dissolution. The results show that at a temperature of 75°C and pH 11.5, gold dissolution increased from 51% to 88% on increasing creatine concentration from 25 to 50 g/t. Further increase in creatine concentration showed a negative effect on the gold dissolution. Gold dissolution increased from 47% to 87% by increasing the temperature from 25°C to 55°C, and from 61% to 89% by increasing pH from 10 to 11.5. Under the optimum conditions of solids’ content of 30%, pH 11.5, creatine concentration of 100 g/t, temperature of 75°C, and hydrogen peroxide addition of 2%, about 90% of the gold could be dissolved from the ore after 24 h. A comparison between creatine and cyanide leaching showed that 90% gold dissolution was achieved at a cyanide concentration of 300 g/t, which is three times higher than the creatine concentration under optimal conditions.

Keywords creatine, cyanide, gold dissolution, leaching

Introduction

Amino acids are organic compounds that are the building blocks of protein. They contain amino (–NH2) and carboxyl (–COOH) functional groups, along with a side chain (R group) that is specific to each amino acid. About five hundred naturally occurring amino acids are known. According to the number of amino and carboxyl groups, amino acids are classified into three groups: (1) neutral (one amino and one carboxyl), such as glycine, alanine, valine, leucine; (2) acidic (one amino and two carboxyl groups), such as aspartic acid, glutamic acid; and (3) basic (two amino and one carboxyl group), such as lysine and histidine (Nelson et al., 2005; Oraby and Eksteen, 2015a).

Geochemists have identified more than twenty kinds of amino acid in sediments and sedimentary rocks (Rashid and Leonard, 1973). They found that organic acids play an important role in the transport and concentration of gold in rock- and ore-forming processes.

Jingrong et al. (1996) studied the capability of organic acids (DL-aspartic amide, DL-alanine, glycine, L-galacystine, DL-aspartate, and histidine) to extract gold and identified suitable conditions. The study showed that gold is readily soluble in amino acids, with temperature, concentration, acidity, and type of amino acid being important factors affecting gold solubility. Using histidine, the highest solubility occurred at 80°C and at pH 6–8.

Leaching of gold from pure gold and gold ores in solutions of amino acid and cyanide solutions containing amino acids has recently been investigated (Azadi et al., 2019; Barani et al., 2021; Eksteen, 2014; Eksteen et al., 2017; Eksteen and Oraby, 2014, 2015; Oraby and Eksteen, 2014, 2015a, 2015b). Oraby and Eksteen (2014) investigated the selective leaching of copper from a gold–copper concentrate in glycine solutions. The results showed that 98% of total copper dissolved in a lixiviant system containing glycine and H2O2, in 48 h at room temperature and a pH of 10.5–11. The copper dissolution rate increased with increasing glycine and H2O2 concentrations and increasing the pulp density decreased the copper dissolution rate. Oraby and Eksteen (2015a) also studied the leaching of gold, silver, and gold–silver alloys in glycine solution in the presence of hydrogen peroxide. Results showed that at neutral and alkaline conditions the solution of glycine and H2O2 was capable of dissolving gold and silver. The gold dissolution rate increased with increased temperature, glycine concentration, silver content, and pH. Eksteen and Oraby (2015) studied the leaching of gold from gold foils in the different amino acid solutions in the presence of hydrogen peroxide. The results showed that glycine (C2H5NO2), histidine (C6H9N3O2), and

Leaching of gold ore using creatine monohydrate

alanine (C3H7NO2) dissolved gold even at low concentrations (0.1 M). The dissolution rate of gold increased with increasing amino acid concentration, peroxide concentration, and pH. Glycine dissolved more gold than histidine and alanine. Oraby et al. (2017) investigated the leaching of gold and copper from gold–copper ores and concentrates in cyanide and glycine–cyanide solutions. The results showed that the gold dissolution rate in the glycine–cyanide system was almost three times greater than that of conventional cyanidation. Oraby et al. (2020) also used glycine as a lixiviant in an alkaline environment for base and precious metals recovery from printed circuit boards. The results showed that in the first stage, alkaline glycine solutions selectively dissolved copper, zinc, and lead over precious metals; in the second stage, gold and silver were recovered using glycine and small amounts of cyanide. Barani et al. (2021) explored the leaching of a polymetal gold using cyanide–glycine solutions. They found that at 75% of gold dissolved using 1500 mg/L cyanide (without glycine), 80% gold dissolution was achieved using 200 mg/L cyanide with 0.5 mg/L glycine, indicating a reduction in cyanide consumption of more than 80%.

In most previous studies, glycine is preferred over the other amino acids, because of its relatively low cost and bulk availability. In this study, creatine (C4H9N3O2) was used for gold recovery from a gold ore. To the authors' knowledge, creatine has not been reported as a lixivant for gold recovery.

Creatine is a popular sports supplement used to increase muscle mass, boost strength, and enhance exercise performance. It is a peptide. Peptides are generally formed when short chains known as peptide bonds form between two or more amino acids. The bonds, covalent in nature, form when the carboxyl group of one amino acid reacts with the amino group of another amino acid. Peptides can bond together, making long molecules known as polypeptides. The shortest polypeptides are dipeptides, which consist of two amino acids bonded by a single peptide bond. Creatine is formed from three amino acids: glycine, methionine, and arginine. Creatine supplements are available on the market in four forms: ethyl ester, gluconate, monohydrate, and nitrate (Stryer et al., 1995; Strumia et al., 2012). In this study, creatine monohydrate was used as a gold lixivant. Figure 1 shows chemical structure of creatine and creatine monohydrate.

Experimental procedure

Materials

The sample used in this research was taken from a run-of-mine (ROM) ore pile of the Agh-Darreh gold mine, which is one of two known gold deposits within the Cenozoic–Recent hydrothermal system of the Takab region in northwestern Iran. The host rock

of this deposit is mainly carbonate. Silver, mercury, iron, and manganese compounds occur in high amounts. The most important mineral containing gold in the sulfide zone of this deposit is pyrite (FeS2). An ore sample weighing 40 kg was crushed and passed through a 270 μm sieve. Sub-samples, each weighing 200 g, were prepared from the main sample by riffling. Figure 2 shows the particle size distribution of the representative sample. Table I shows the chemical compositions of the representative sample, determined by X-ray fluorescence spectroscopy (XRF; Philips X Unique II). The sample analysed mainly SiO2, CaO, Al2O3, and Fe2O3. The gold content was determined by fire assay followed by atomic absorption spectrophotometry (AAS). The ore contained 2 ppm Au. X-ray diffractometry (XRD) showed, in order of abundance, that quartz, calcite, smectite, illite, dolomite, barite, goethite, kaolinite, iron arsenate dihydrate, jarosite, and pyrite occurred in the sample.

Mineralogy examination

A sample was divided into six size fractions (+50, −50+40, −40+30, −30+20, −20+10, and −10 μm); polished sections were prepared from each fraction. The sections were examined by scanning electron microscopy (SEM). Figure 3 shows backscattered-electron images. Image analysis showed that the gold particles were locked with other minerals in the coarse size fractions (+50 and −50+40 μm) (Figure 3-A). In the −10 µm size fraction, gold particles were free—they were not locked with other minerals (Figure 3-B).

Leaching tests

The leaching tests were carried out in a stainless-steel laboratory reactor (1.2 L). The reactor was immersed in a water bath that controlled the temperature. In all experiments, analytical-grade reagents and deionized water were used. The ground ore sample (200 g) and 466 mL deionized water (30% solids by mass) were placed in the reactor. The pH was adjusted to the desired value by adding NaOH or H2SO4. The desired amount of creatine was then added and the suspension was agitated (1000 rpm) for 24 h. Thereafter, the suspension was filtered and the Au in the dried cake and solution measured with AAS.

Results

Effect of creatine concentration

Figure 4 shows the effect of creatine concentration on the gold dissolution at solids’ content of 30%, pH 11.5, 2% H2O2, temperature of 75°C, and leach time of 24 h. The gold dissolution increased from 51% to 88% with increasing creatine concentration from 25 to 50 g/t. Further increase in creatine concentration showed a negative effect on gold dissolution, decreasing from 89% to 84%

Figure 1—Chemical structures of (a) creatine and (b) creatine monohydrate Figure 2—Particle size distribution of the sample

Leaching of gold ore using creatine monohydrate

4—Effect of creatine concentration on gold dissolution at leach conditions of 30% solids content, pH 11.5, 2%

Table I

*Loss on ignition

with increasing creatine concentration from 100 to 1000 g/t. As the creatine concentration increased beyond 100 g/t, the creatine molecules formed peptide bonds and the free creatine concentration decreased, which led to the decrease in the dissolution of gold.

Effect of temperature

Figure 5 shows the effect of temperature on gold dissolution at a creatine concentration of 100 g/t, solids content of 30%, 2% H2O2, leach time of 24 h, and pH 11. The gold dissolution at

leach time of 24 h, and temperature of 75°C

Figure 5—Effect of temperature on gold dissolution at leach conditions of 100 g/t creatine, 30% solids content, pH 11.5, 2% H2O2, and leach time of 24 h

room temperature (25°C) was low (47%), but increased to 87% on increasing the temperature to 55°C. Further increase in temperature to 65°C and 75°C had no significant effect on the dissolution of gold. The results show that temperature is an effective parameter in the range of 25–55°C.

Effect of pH

Figure 6 shows the effect of pH on gold dissolution at a creatine concentration of 100 g/t, solids content of 30%, 2% H2O2, leach time of 24 h, and temperature of 55°C. The gold dissolution increased from 61% to 89% by increasing pH from 10 to 11.5. The gold dissolution decreased by about 6% points on increasing pH from 11.5 to 12. Gold dissolution in creatine–peroxide solutions is very sensitive to the leaching pH. Increasing the pH promotes

Figure 3—Backscattered-electron images of (A) −50+40 μm (arrow points to a locked particle of gold) and (B) −10 μm fractions (yellow particle is liberated gold)

Figure

H2O2,

Leaching of gold ore using creatine monohydrate

6—Effect of pH on gold dissolution at leach conditions of 100 g/t

decomposition of peroxide and generates hydroxide ions and oxygen (Equations [1] and [2]) or produces hydroxide radicals (Equation [3]) (Süss et al., 1998), which can accelerate gold dissolution (Nowicka et al., 2010). These results are consistent with those obtained for glycine (Eksteen and Oraby, 2015; Oraby and Eksteen, 2015b).

Effect of leaching time

Figure 7 shows the effect of time on gold dissolution at a creatine concentration of 100 g/t, solids content of 30%, 2% H2O2, pH 11.5, and temperature of 55°C. Most of the gold (84%) dissolved in the first 12 h; dissolution only increased by a further 4% points from 12–24 h. Further increases in leach time to 32 and 48 h had no significant effect on gold dissolution. This can be explained by the reduction of gold concentration in solution and passivation of the gold surface by a layer of AuOH adsorbed to the gold surface (Oraby and Eksteen, 2015b). The dissolution also released sulfide ions, leading to the formation of a passive layer of Au2S on the surface of gold, which reduces its dissolution (Dai and Jeffrey, 2006; Lorenzen and van Deventer, 1992).

Cyanide leaching

For comparison between creatine leaching and cyanide leaching, four leaching experiments at different cyanide concentrations were carried out at solids contents of 30%, pH 10.5, and room temperature. Figure 8 shows the percentage of gold dissolution after 24 h. The results show that the maximum gold dissolution (90%) was obtained at 300 g/t cyanide. This gold dissolution value (90%) was obtained by a 100 g/t cyanide solution.

Free cyanide in final filtrate solutions was immediately analysed by titration with AgNO3 using rhodamine as colorimetric endpoint indicator (from yellow to salmon), following the routine recommended by the International Cyanide Code (2012), and described in the Standard Methods (APHA-AWW A-WEF, 1998).

Figure 9 shows the free cyanide concentration in the final filtrate solutions. At a cyanide concentration of 300 g/t, 50% of the cyanide was consumed and 50% remained in the solution as free cyanide, which could be recycled or neutralized at a cost.

Figure 7—Effect of leaching time on gold dissolution at leach conditions of 100 g/t creatine, 30% solids content, 2%

Figure 8—Effect of cyanide concentration on gold dissolution at leach conditions of 30% solids, 2% H2O2, pH 11.5, and temperature of 25°C

9—Free cyanide concentration in cyanide leach solutions

Conclusion

Creatine was used as a lixiviant for gold dissolution from gold ore. The effects of creatine concentration, temperature, leach time, and pH were investigated for their influence on gold dissolution. Gold dissolution increased from 51% to 88% with increasing creatine concentration from 25 to 50 g/t. Further increase in creatine concentration showed a negative effect. The results show that temperature is an effective parameter in the range of 25−55°C. Gold dissolution increased from 47% to 87% with an increase in temperature from 25 to 55°C. Gold dissolution in creatine–H2O2 solutions is very sensitive to the leaching pH, increasing from 61% to 89% with increase in pH from 10 to 11.5. Most gold dissolved within 12 h, and longer leach times had no significant effect on the

Figure

creatine, 30% solids content, 2% H2O2, leach time of 24 h, and temperature of 55°C

H2O2, pH 11.5, and temperature of 55°C

Figure

Leaching of gold ore using creatine monohydrate

dissolution. The optimum conditions were defined as solids content of 30%, pH 11.5, creatine concentration of 100 g/t, temperature of 75°C, and 2% H2O2, at which almost 90% of the gold was dissolved from the ore within 24 h.

Using cyanide leaching, 90% gold dissolution was achieved at a cyanide concentration of 300 g/t and ambient temperature, which is three times higher than the creatine concentration under the optimal conditions. Approximately 50% of the cyanide remained in the solution as free cyanide, which can have safety, cost, and environmental consequences.

Bulk cyanide and creatine monohydrate prices are USD 1800−2200/t and USD 5000−10,000/t, respectively (Alibaba. com). Creatine monohydrate is 2.7−4.5 times more expensive than cyanide. To achieve a gold recovery of 90% from the investigated ore, cyanide consumption was almost three times that of creatine. However, leaching with creatine requires a higher temperature than cyanide, so its economic advantage remains unknown. A technoeconomic study is required. As creatine is safer to work with and less harmful to the environment, we recommend that its application be explored further.

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Leaching of gold ore using creatine monohydrate

ESG IN THE MINERALS INDUSTRY CHALLENGES AND OPPORTUNITIES

DATE: 16-17 OCTOBER 2024

VENUE: FOCUS ROOMS, MODDERFONTEIN, JOHANNESBURG

BACKGROUND

Environmental, Social, and Governance (ESG) considerations have become increasingly important in the business world, as it contributes to long-term sustainability and responsible corporate behaviour.

In the minerals industry, ESG considerations are particularly important due to the sector’s significant environmental and social impacts. Mining operations involve land use, energy and water consumption, and waste generation, which have lasting effects on ecosystems. Additionally, the industry faces challenges related to labour practices, community engagement, and the impact on indigenous populations.

An ESG-driven strategy is not only a responsible approach to business but also a strategic imperative for long-term success. It can contribute to risk mitigation, enhances reputation, attracts capital, and fosters innovation, making it a competitive advantage in today’s business landscape. In the mining industry, ESG considerations are crucial for addressing environmental and social challenges and ensuring the industry’s sustainable development.

The role of the Southern African Institute of Mining and Metallurgy (SAIMM) in the promotion of ESG is based on the premise that sustainability, and the contribution of the mining and minerals industry to society, is dependent on the professional and ethical conduct of minerals industry professionals – our members.

On this basis, the purpose and focus of this conference is to influence professional behaviour, and foster industry dialogue on sustainability and responsible mining through Environmental, Social, Governance, and Sustainability-related matters.

We invite you to share your knowledge and experience with an audience of like-minded individuals to inspire growth and change.

Camielah Jardine: Head of Conferences

ECSA Validated CPD Activity, Credits = 0.1 points per hour attended.

Affiliation:

1Stellenbosch University, South Africa

Correspondence to: A. Tapfuma

Email: anthonytapfuma@gmail.com

Dates:

Received: 17 Jul. 2023

Accepted: 16 Feb. 2024

Published: April 2024

How to cite:

Tapfuma, A., Akdogan, G. and Tadie, M. 2024. Investigation of glycine leaching for gold extraction from Witwatersrand gold mine tailings with permanganate pre-treatment. Journal of the Southern African Institute of Mining and Metallurgy, vol. 124, no. 4, pp. 219–230

DOI ID:

http://dx.doi.org/10.17159/24119717/2967/2024

Investigation of glycine leaching for gold extraction from Witwatersrand gold mine tailings with permanganate pre-treatment

by A. Tapfuma1, G. Akdogan1, M. Tadie1

Abstract

The Southern African region contains abundant gold tailing heaps that are an environmental concern due to dust pollution and chemical contamination of nearby water bodies. Nevertheless, these heaps contain gold with potential to be extracted for financial gain. In this study, the feasibility of extraction of gold from a Witwatersrand gold tailings using glycine was investigated using a two-level full factorial design. Effects of variables such as glycine concentration at levels of 0.5–1.5 M with a solid–liquid ratio of 50–200 g/L were investigated. Although the solid–liquid ratio used in this study was below the industrial norm, it provides a starting point for investigating the applicability of this technology. To enhance the gold extraction, potassium permanganate pretreatment and copper ion addition were adopted. The results revealed that 91.4% gold extraction was achieved after 24 h pre-treatment of the tailings using 2 g/L potassium permanganate followed by subsequent leaching using 1.5 M glycine at pH 12, with 50 g/L solid–liquid ratio and 2 g/L copper ions. Statistically, the solid–liquid ratio and potassium permanganate pre-treatment of the tailings were found to be significant at 95% confidence interval, whilst interactions of copper ions and glycine concentration were significant factors. The work demonstrates that potassium permanganate pre-treatment prior to glycine leaching of low-grade secondary gold resources, such as tailings, can be beneficial. Furthermore, the methodology developed in this work provides an opportunity for further investigation of the recovery of gold hosted in complex mineralogical matrices.

Keywords tailings, glycine, gold, leaching, potassium permanganate, Cu ions

Introduction

Owing to strong complexation with gold and technical simplicity, cyanidation has been used as the conventional process in the extraction of gold (Fleming, 1992; Nicol et al., 1992). However, the process is reported to impose environmental, health, and safety challenges due to the high toxicity of cyanide (Hilson and Monhemius, 2006). Alternative reagents for gold leaching, such as thiourea, thiosulfate, halide, and thiocyanate that are more environmentally benign have been investigated over the past few decades and have been reported in several reviews (Aylmore, 2016; Gökelma et al., 2016; Hilson and Monhemius, 2006; Prasad et al., 1991). The thiosulfate process is the most developed alternative method for gold leaching, albeit having its technical challenges. These include the high reagent consumption, generation of polythionates, which readily absorb on anion-exchange resins and impair the recovery of gold from solution, complex chemistry, and sensitivity to process conditions like pH and Eh (Aylmore, 2005, 2016).

Glycine was proposed by Eksteen and Oraby (2014) as a potential alternative to the cyanidation process, and has attracted significant attention as a lixiviant in the leaching of metals from ores. It is a cheap, benign, and chemically stable lixiviant for gold dissolution. The process of using glycine is similar to the cyanidation concept, in which oxygen is used to oxidize gold from the metallic state to gold(I), as represented by Reaction 1 (Eksteen and Oraby, 2015; Oraby and Eksteen, 2015b):

[1]

Oxygen or oxidation is a key process step for the reaction, having significant implications for the kinetics, and therefore the design of the process. It has been noted that, in comparison with cyanide and operated under similar conditions, glycine extraction is significantly slower (Oraby et al., 2019). Slow kinetics prompted the consideration of heap leaching as a feasible process route for gold extraction using this process (Eksteen et al., 2018). To combat limitations that may relate to mass transfer of oxygen in the system, stronger oxidizing agents have been applied, resulting in significant improvements in extraction. Hydrogen peroxide and potassium permanganate are commonly investigated oxidants (Oraby et al., 2020; Perea and Restrepo, 2018). Reactions 2 (Eksteen and Oraby, 2015b) and 3 (Oraby et al., 2020) are proposed

Investigation of glycine leaching for gold extraction from Witwatersrand gold mine tailings

for hydrogen peroxide and potassium permanganate, respectively. Up to 85% gold extraction from an oxide ore within 48 h was reported using 2 g/L of potassium permanganate by Oraby et al. (2020). After extraction of the gold glycine complex from the ore, carbon adsorption can effectively recover the gold from the pregnant leach solution, as reported by Eksteen and Oraby (2014) and Tauetsile et al. (2018).

Given the reported high adsorption recovery, most research efforts have concentrated on exploring alternative approaches to enhance gold leaching within the glycine system. One proposed strategy involves the synergistic use of cyanide, as suggested by Oraby et al. (2017). However, the use of cyanide alongside glycine is an environmental trade-off. Elevated temperatures have been shown to improve extraction of gold, provided the supply of oxygen is sufficient, as shown by Altinkaya (2020) and Broeksma (2018). Various other ways of improving the leaching extent include increasing glycine concentration (Altinkaya et al., 2020) and the inclusion of Cu2+ ions as a catalyst for the leaching process (Eksteen and Oraby, 2015a). The body of work conducted with this method has focused on primary gold sources that have economic grades (Altinkaya et al., 2020; Eksteen et al., 2018; Oraby et al., 2020) compared with the tailings grades in the current study. Therefore, leaching of tailings with alternative lixiviants is certainly a potential area for development.

In this paper, historical tailings samples originating from the processing of Witwatersrand ore body were used as a case study for gold recovery using glycine. To investigate the feasibility of glycine as a lixiviant, parameters such as solid–liquid ratio and glycine concentration were investigated, together with use of permanganate to pre-oxidize gold and copper ions to provide a possible catalyst effect. This paper presents a base-case scenario for further investigation of the potential for the application of glycine leaching for tailings reprocessing.

Experimental

Materials and sample preparation

The sample used for the investigation was collected from the West Rand of the Witwatersrand mining district. Prior to sample characterization, the tailings was prepared to represent the bulk tailings sample. The sample was mixed using a shovel split using the cone-and-quartering technique. Two quarters were taken to a

Table I Summary

riffle splitter to further obtain a sample for rotary splitting. Four kilograms of samples from this process were taken to a rotary splitter to divide them into ten samples, each of 400 g. These samples were packed into different sample bags for characterization and leaching.

Chemicals and equipment

All experiments were conducted using solutions prepared from analytical-grade reagents and deionized water. The chemicals are listed in Table I with their respective purities. Gold and other elements in solutions were determined by Thermo ICap 6200 inductively coupled plasma optical emission mass spectrometry (ICP-MS) conducted at Stellenbosch University, Central Analysis Facility (CAF).

Sample characterization techniques

Material characterization included the determination of particle size analysis, bulk mineralogy, and qualitative gold imaging. Grain size analysis was carried out to determine the particle size distribution using a Micromeritics Saturn DigiSizer 5200 laser diffraction particle size analyzer. Fire assay was conducted on ten representative samples to determine the gold head grade of the tailings. Bulk mineralogy in the form of quantitative X-ray diffraction (QXRD) using a PANalytical Aeris diffractometer and X-ray fluorescence spectroscopy (XRF) using a PANalytical Axios wavelengthdispersive spectrometer was conducted for chemical analysis. A Zeiss Merlin field-emission gun scanning electron microscope (SEM) was used on a representative sample of the tailings to identify gold grains and their associations. The conditions used for imaging were 20 kV accelerating voltage, a beam current of 16 nA, with a working distance of 9 mm.

Experimental design

Towards understanding the leaching response of the tailings material, cyanidation tests were carried out as a form of diagnostic characterization at 2 kg/t NaCN, initial pH of 11, 50% solids content at ambient conditions. Three tests were conducted and the extent of gold extraction was evaluated. Three additional experimental runs were carried out using the glycine conditions used by Oraby et al. (2020). The tailings material was subsequently leached at 2 g/L glycine concentration (4.6 kg/t glycine), 30% solids content, pH 10.5, 2.0 g/L potassium permanganate, and ambient temperature.

In the main body of the study, a glycine concentration range of 0.5–0.5 M and Cu2+ ion range of 0–2 g/L with a solid–liquid ratio ranging from 50–200 g/L was used to understand their effects on gold leaching. Whilst the solid–liquid ratio is not reflective of an industrial process, which would have higher solid loading of up to 50% (m/v), the study can be used to inform further studies on higher solid–liquid ratios. Glycine concentrations were above

Investigation of glycine leaching for gold extraction from Witwatersrand gold mine tailings

stoichiometric requirements for the gold content in the solids to cater for any free glycine loss by oxidation and with the assumption that free glycine would be recycled in an industrial setting. Cu2+ ions were introduced as copper sulfate into the leach system for those experiments investigating its possible catalytic effect. The amount of copper used ensured the glycine-to-gold stoichiometric ratio maintained an excess of glycine. Permanganate was used for pre-treating the tailings to oxidize the material and as an oxidant in leaching step to enhance gold extraction in all experimental runs. Potassium permanganate addition in leaching was used due to its strong oxidizing nature compared with O2 or H2O2. Moreover, there is a claim that it produces manganese dioxide colloids that act as additional oxidants (Oraby et al., 2020; Pratesa et al., 2019).

Experimental procedure

A full factorial design was generated to investigate the effect of glycine concentration, solid–liquid ratio, potassium permanganate pre-treatment, and copper concentration at two levels, according to the high and low values of the ranges stated in the previous section. Table II shows the sixteen experiments conducted in the design of experiments (DOE).

All leaching experiments were conducted in a 2 L reactor open to atmosphere. The agitator was fitted with a two-radial-blade impeller and set at a constant speed of 300 rpm, which was sufficient to keep particles suspended. Although natural aeration was allowed, the use of an oxidant in all experiments was intended to further catalyse oxidation.

Two phases of experimentation were conducted. Phase 1 investigated the effect of potassium permanganate pre-treatment on gold extraction. The pre-treatment procedure summarized in Table III involved oxidizing the tailings (50% solid loading) in 2 g/L potassium permanganate solution for 24 h. The pre-treatment initial conditions were pH 5.10 ± 0.05 and Eh of 110 ± 3 mV, which is in the region for reduction of potassium permanganate (Pourbaix, 1974). In the subsequent leaching stage, alkaline glycine solution was added to the agitated slurry to make up the desired concentration of glycine and solid–liquid ratio required for leaching. Additional potassium permanganate was added to the leach slurry (2 g/L) and a few drops of NaOH were momentarily added to adjust the pH to 12.

Full factorial design table

Phase 2 involved a study of the effects of glycine concentration and solids loading on the extraction of gold without tailings pretreatment. The experimental procedure involved the addition of the tailings sample and deionized water to the leaching vessel, followed by the addition of a specific volume of alkaline glycine solution to achieve the required initial glycine concentration. A few drops of 250 g/L permanganate solution was added to obtain an initial 2 g/L permanganate concentration in the leaching solution. This was followed by adjustment of the pH to 12 and commencement of leaching time. For those experimental runs where the effect of copper ions was investigated, a calculated amount of copper sulfate was added to achieve the desired 2 g/L Cu2+ initial concentration in the leach solution. The pH was maintained at 12 ± 0.2 during the experiment. Table IV summarizes the leaching conditions applied. Variables are presented within specified ranges, while constant parameters are listed with precise values.

Aliquot samples (10 mL) were collected at different time intervals and filtered using 45 µm filters. At the end of the experiment, the entire pulp was vacuum filtered, followed by washing with deionized water to remove entrained leaching solution. The residue was then left in the oven for 24 h at 50°C to dry. Fire assay was used to determine the gold in the dried residue while ICP-MS was used to analyse the filtrate solutions.

Results and discussion

Tailings characterization results

The particle size distribution of the tailings (Figure 1) showed

Table III

Summary of pre-treatment conditions

Table II

Investigation of glycine leaching for gold extraction from Witwatersrand gold mine tailings

Table IV

that 80% of the sample had a diameter of less than 88 µm, which is close to the conventional value of 80% passing 75 µm used in the leaching of primary sources of gold (Stange, 1999). The gold content of the tailings (bulk head grade) was analysed to be 0.842 ± 0.04 g/t, which is similar to some reported Witwatersrand tailings grades of about 0.5 g/t (Janse van Rensburg, 2016).

XRD and chemical analysis of the tailings samples by XRF are presented in Figure 2 and Table V, respectively. As is typical of the

Witwatersrand ores, quartz was the dominant mineral, with a grade of 77.8%. Figure 3 shows an example of SEM imaging conducted on the tailings to identify association and grain size. Discrete particles in the sample were verified to be associated with quartz, with the largest identified grain size being 0.71 μm in diameter.

Cyanidation experiments conducted as described above resulted in 24 h gold extraction of 65.7 ± 2.95%, indicating that the sample can be identified as a moderately refractory material, according to classification by Asamoah et al. (2014). Preliminary tests conducted using glycine leaching conditions applied by Oraby et al. (2020) on this tailings material yielded 15.7 ± 0.82% gold extraction, and therefore a need for alternative measures to increase gold extraction in this material using glycine. Pre-oxidation of the tailings in potassium permanganate prior to leaching was therefore considered in the presence of Cu ions.

Leaching experiments and results

Effect of solid–liquid ratio and reagent dosage

The effect of solid–liquid ratio on gold leaching using alkaline glycine is shown from Figure 4 to Figure 7. Figure 4 shows the effect of solid–liquid ratio on leaching at a concentration of 0.5 M

Table V

Figure 1—Tailings sample particle size distribution

Figure 2—X-ray diffractogram of the tailings sample

Investigation of glycine leaching for gold extraction from Witwatersrand gold mine tailings

5—Effect of solid–liquid ratio on gold leaching at 0.5 M glycine after 24 h tailings pre-treatment in 2 g/L potassium permanganate at (a) 0 g/L copper ions (b) 2 g/L copper ions

glycine for a sample without pre-treatment at (a) 0 g/L copper ions, (b) 2 g/L copper ions. In the absence of copper ions, a maximum extraction of 51.5% was observed for the 50 g/L solid–liquid ratio compared with 24.8% for the 200 g/L solid–liquid ratio, representing an approximate 100% increase in gold extraction. In the presence of copper ions, however, 87.9% gold extraction was observed for 50 g/L compared with 26.2% for 200 g/L solid–liquid ratio in the presence of copper, which is an approximate 335% increase in extraction due to reduction of the solid–liquid ratio. A clear synergism is observed between reduction of solids ratio and presence of copper ions under the conditions of 0.5 M glycine and no pre-treatment. Figure 5 shows the effect of solid–liquid ratio

at 0.5 M glycine after 24 h tailings pre-treatment in potassium permanganate at (a) 0 g/L copper ions, (b) 2 g/L copper ions. In the absence of copper, the leached gold amounted to 61.6% for 50 g/L compared with 49.0% for 200 g/L solid–liquid ratio. In the presence of copper ions, extraction of 90.9% was achieved for the 50 g/L solid–liquid ratio and 53.1% for the 200 g/L solid–liquid ratio. It is evidently seen that, in the presence of copper ions, the clear effect of copper ions is noted by improved leaching performance, with extraction being approximately 71% higher at the lower solid–liquid ratio.

At 1.5 M glycine, Figure 6 shows the effect of solid–liquid ratio without tailing pre-treatment for leaching at (a) 0 g/L copper

Figure 3—Images produced by Zeiss Merlin field-emission gun electron microscope to show (a) association and (b) gold grain size in the tailings sample

Figure 4—Effect of solid–liquid ratio on gold leaching at 0.5 M glycine without tailings pre-treatment at (a) 0 g/L copper ions (b) 2 g/L copper ions

Figure

Investigation of glycine leaching for gold extraction from Witwatersrand gold mine tailings

ions (b) 2 g/L copper ions. In the absence of copper ions, gold extractions were 78.4% and 39.7% for 50 g/L and 200 g/L solids density, respectively. In the presence of copper, extractions were 57.8% and 21.5% for 50 g/L and 200 g/L pulp densities, respectively. Increasing solid–liquid ratio ultimately reduced the extraction of gold by about 100% for each of the solid–liquid ratios investigated.

Extraction results for tailings pre-treatment at 1.5 M glycine are shown in Figure 7, where the effect of solid–liquid ratio at (a) 0 g/L copper ions (b) 2 g/L copper ions is shown. Under these conditions, in the absence of copper ions, gold extractions were 71.0% and 59.2% for 50 g/L and 200 g/L solid–liquid ratios, respectively, which is less than 100% increase in extraction with reduced solid–liquid ratio. In the presence of copper, extractions were 91.4% and 37.1% for 50 g/L and 200 g/L solid–liquid ratios, respectively (248% increase in extraction with reduced solid–liquid ratio).

Previous work on the use of glycine for leaching of gold has shown that a reduction in solid–liquid ratio is directly proportional to an increase in extraction (Oraby et al., 2019). A 20% increase in gold extraction was observed when solid–liquid ratio was changed from 50% to 20% in the leaching of an oxide ore (100% passing 75 μm) using alkaline glycine at 50 g/L concentration. In general, most scholars have shown that decreasing the solid–liquid ratio in a leaching system, regardless of the lixiviant, will support high leaching of gold from the ore (Brittan and Plenge, 2015; Öncel et al., 2005; Tripathi et al., 2012). Notably, under such conditions, reagent dosage (kg/t of ore) increases with a reduction in the amount of solids present and positively affects the diffusion of reactants (glycine) and products (e.g., gold glycinate complex) from the surface. Increased concentration gradients ultimately drive faster

mass transfer. For the chemical reaction step, reduced solid–liquid ratio leads to higher excess lixiviant, which ultimately increases the rate of chemical reaction.

The results presented in this study in Figures from 4 to 7 and those from replication of work of Oraby et al. (2020) clearly indicate that a significant increase in reagent dosage (kgglycine/tore) is required to achieve high extraction efficiencies. Figure 8 shows a graph of gold extraction against glycine addition for leaching tests that had no additional treatments, i.e., no pre-treatment or addition of Cu2+ ions. In many instances in the literature, it is noted that an increase in glycine concentration results in an increase in gold leaching in the absence of copper ions (Eksteen and Oraby, 2015a; Oraby and Eksteen, 2015b; Oraby et al., 2019, 2020). Figure 8 shows that for

Figure 6—Effect of solid–liquid ratio on gold leaching at 1.5 M glycine without tailings pre-treatment at (a) 0 g/L copper ions (b) 2 g/L copper ions

Figure 7—Effect of solid–liquid ratio on gold leaching at 1.5 M glycine after 24 h tailings pre-treatment in 2 g/L potassium permanganate at (a) 0 g/L copper ions (b) 2 g/L copper ions

Figure 8—Gold extraction as a function of reagent dosage in the absence of pre-treatment and Cu2+ for glycine leaching

Investigation of glycine leaching for gold extraction from Witwatersrand gold mine tailings

this tailings material, an increase in reagent concentration from 4.6 kg/t (300 g/L, 0.026 M) to 174 kg/t (200 g/L, 0.5 M) resulted in a minor increase in extraction (less than 10% points).

It should be noted that the extent of the effect of increasing glycine concentration on extraction is not replicated amongst different leaching materials. For example, Oraby et al. (2019) observed an 18.2% increase in gold dissolution by increasing glycine concentration from 3.75 g/L (0.05 M) to 15 g/L (0.20 M) when leaching an oxide ore in the absence of copper. A similar trend observed by Oraby et al. (2020) was an over 30% increase in extraction when glycine concentration was increased from 0.5 g/L to 2.0 g/L at pH 10.5 and 1 g/L permanganate concentration for an oxide ore.

Within the context of these results, it is also noted that whilst solid–liquid ratio reduction has a notable impact on the recovery, other operating conditions, such as the presence of Cu2+ ions and pre-treatment of the material by potassium permanganate prior to alkaline leaching in glycine and permanganate as oxidant, are investigated. The effects of these parameters are expanded further in the following sections.

Effect of permanganate pre-treatment

Figure 9 shows the extraction values for all tests conducted under all conditions tested on the tailings material. All leaching tests were conducted under alkaline conditions, as described in the methods, and included 2 g/L potassium permanganate added to the leach to act as an oxidant. Results indicate high extraction efficiencies

under these conditions in comparison with conditions where the pre-treatment was not implemented, except for experiments done at 50 g/L solid–liquid, 1.5 M glycine in the absence of copper. The improved extraction values due to permanganate pretreatment are independent of the presence or absence of Cu2+ ions. Highest extraction of gold (91.4%) was observed at 50 g/L solid–liquid ratio and 1.5 M glycine, in the presence of copper ions after preoxidation. The second-highest extraction (90.9%) was also noted at 50 g/L solid–liquid ratio but 0.5 M glycine in the presence of copper ions after pre-treatment. These values are within 0.5% points of each other and considered to show negligible difference. What this indicates is that glycine reagent dosage does not control the extraction. Both results are in the presence of Cu2+, however, suggesting that the combination of the presence of Cu species, the action of pre-oxidation, and high reagent dosage has a notable effect on gold extraction.

Investigations conducted to inform the processes occurring during permanganate pre-treatment were carried out using ultraviolet–visible spectroscopy (UV–Vis) of the pre-treatment solution during and after pretreatment to test for any gold species in solution and thereby determine whether any gold species were potentially pre-oxidized. Results in Figure 10(a) show UV-Vis spectra for solutions from 1 h to 24 h of pre-treatment. Figure 10(b) shows UV-Vis spectra for Au(I) and Au(II) chloride in deionized water, which were used as reference spectra to infer the presence of gold species in the pre-treatment solution. Distinct peaks at wavelengths of 280 nm for all solutions collected during

Figure 9—Variation of gold extraction with different conditions employed in this study

Figure 10—Ultraviolet–visible spectra of (a) solutions from permanganate pre-treatment of tailings

(b) gold(I) and gold(III) chloride solutions at 1 g/L concentrations

Investigation of glycine leaching for gold extraction from Witwatersrand gold mine tailings

pre-treatment are shown, with intensity increasing in relation to pre-treatment time, which is an indicator of increased oxidation of gold available to the oxidant. Using Figure 10(b) as a reference, the Au(I) chloride and Au(III) chloride peaks lie closer to the observed 280 nm wavelength, with Au(I) chloride being the closest. It can be inferred that the peak at the 280 nm wavelength may be associated with already-leached gold during pre-treatment in the form of Au(I); however, Au(III) cannot be excluded based on this analysis alone.

Elemental analysis of the pre-treatment solution was also conducted to verify the presence of gold in solution. In addition, other elements found in solution were used as signatures of the dissolution effect of the oxidant on the host minerals in the ore. Figure 11 shows these results, in the form of the degree of extraction of various elements in the tailings during pre-treatment. In Figure 11(a), the results for gold extraction show that up to 5% of gold was leached during this pre-treatment stage. It is possible that the gold may have been readily accessible to the highly oxidising pretreatment solution, e.g., as the discrete micro-sized particles shown in Figure 3, and therefore readily oxidized into solution by the strong oxidising agent.

It was also hypothesized that pre-treatment may improve gold extraction using glycine by oxidation of grains of gangue minerals, resulting in increased lixiviant exposure to gold. It has been determined that up to 60% of gold is extractable by cyanide under typical cyanidation conditions. Whilst this is not severely refractory, it does indicate some gold to be non-accessible to conventional cyanidation. Pre-treatment/pre-leaching of ore is an ore-

conditioning step that has been well documented in the literature for refractory ores using reagents such as sodium hydroxide (Mutimutema et al., 2022; Snyders et al., 2018). It has also been reported that during this step, oxidation of refractory species results in the release of associated gold. In the pre-treatment step applied in this study, the pH was recorded to be pH 5 and Eh of 110 mV, which is a region in which MnO4 is reduced to form MnO2, which is a reduction process that might be responsible for dissolution of minerals during pre-treatment (Pourbaix, 1974). Figure 11(b)–(f) indicate the extraction of elements associated with the gangue mineralogy in the ore, such as Si from quartz and S from gypsum released during pre-oxidation. Elements such as P, Mg, Fe, and S, all contained in the gangue minerals, also increased with time, albeit the levels of extraction were low. Whilst potassium permanganate is a strong oxidising agent, it does not compare with the effect of preleaching agents such as caustic.

Effect of glycine and copper concentration

The role of Cu ions in glycine leaching systems is yet to be fully detailed. Previous work, however, has shown that there is a potential for increased extraction due to the presence of Cu ions. Eksteen and Oraby (2015) showed that copper ions accelerated the leaching process in tests that showed more gold extraction compared with leaching without copper at the same experimental conditions. The leaching conditions for that work were 0.1 M glycine, 0.1% H2O2, pH 11.9, and 4 mM Cu2+ (Eksteen and Oraby, 2015). By contrast, previous work by the same research group showed that increasing glycine concentration in the presence of copper ions led

Figure 11—Dissolution of (a) gold, (b) iron, (c) magnesium, (d) silicon, (e) sulfur, and (f) phosphorus during 24 h pre-treatment of the tailings material in 2 g/L potassium permanganate

Investigation of glycine leaching for gold extraction from Witwatersrand gold mine tailings

to reduced extraction of gold (Oraby and Eksteen, 2015a). Linear sweep voltammograms demonstrated that an overvoltage was required for the oxidation of gold in a gold–copper system with excess free glycine molecules compared with no free molecules. This demonstrated that the presence of free glycine has an impact on oxidation. The observation may suggest that the glycine–copper reaction may be strongly dependent on glycine concentration, in addition to yielding reaction products that inhibit gold oxidation. In Figure 9, the results obtained support the observations of previous studies, despite the subject leached material being different. The presence of Cu ions in the leach is observed to consistently produce greater extraction at glycine concentrations of 0.5 M compared with concentrations of 1.5 M at the same solid–liquid ratio. The only exception is the condition where 50 g/L solid–liquid ratio was leached in 1.5 M glycine after permanganate pre-treatment. Lastly, more work is needed in the glycine context to explain this observed trend and the interaction between copper ions and glycine concentration during gold dissolution.

Repeatability

Figure 12 shows the results for three experiments done at 1.5 M glycine concentration, 50 g/L solid–liquid ratio at 2 g/L copper ions after 24 h pre-treatment of the tailings in 2 g/L potassium permanganate. The experiments were conducted to verify the repeatability of the tests using the conditions that yielded the highest extraction. The 24 h gold extraction was 88.76 ± 2.33%, which represented a coefficient of variation of 0.026, showing that the experiment runs could be repeated without great deviation.

Statistical analysis

Table VI shows the regression analysis used to estimate the relationship (model) between gold extraction and the investigated parameters (effects of pre-treatment, solid–liquid ratio, glycine concentration, and copper concentration) within the investigated ranges. The p-values represent the probability value, which denotes the significance of a factor, which depends on a threshold. The common threshold is 0.05, where p < 0.05 represents significance of the factor at 95% confidence interval. The regression coefficient signifies the amount by which a change in the value of the factor must be multiplied to give the corresponding average change in gold extraction for this model. The factor effects and pair interactions were investigated using 24 h leach recoveries as the output response using Design-Expert software.

The results showed that the glycine concentration and copper ion concentrations investigated were insignificant at a 95%

Table VI

Statistical summary of p-values and coefficients

confidence interval because the p-values were greater than 0.05. This was experimentally indicated, where the effect of glycine concentration and copper concentration did not have a definite relationship with gold extraction.

Solid–liquid ratio and pre-treatment were significant as individual variables and so were the glycine–copper and S/L–Cu interactions. The effect of pre-treatment was also found to be significant at a 95% confidence interval. Notably, pre-treatment as a process variable had a positive model coefficient of 7.89, which reflects that gold extraction increased as pre-treatment time increased. In contrast, solid–liquid ratio had a negative coefficient, which is also in agreement with the experimental observation that gold leaching decreased as solid–liquid ratio increased.

As mentioned above, Table VI also demonstrates that there are factor interactions that are significant in this DOE. The two significant interactions are glycine–copper and copper–S/L, as shown in Figure 13 and Figure 14. The glycine–copper interaction chart seen in Figure 13 is in agreement with the qualitative analysis discussed above. Figure 14 exhibits the copper ion–solid–liquid interaction, which statistically shows high gold extraction in the presence of copper at a low solid–liquid ratio, whereas its presence at high solid–liquid ratio results in lower extractions.

Conclusions

A glycine leaching study for a Witwatersrand ore tailings was conducted to investigate the effect of different process variables

Figure 12—Gold leaching at 1.5 M glycine concentration, 50 g/L solid–liquid ratio at 2 g/L copper ions after 24 h pre-treatment of the tailings sample in 2 g/L potassium permanganate

Figure 13—Interaction chart for glycine–copper concentration

Investigation of glycine leaching for gold extraction from Witwatersrand gold mine tailings

on gold extraction. The work showed that high gold extraction was attainable with low solid–liquid ratio and permanganate pre-treatment. The results revealed that 91.4% gold was leached from the pre-oxidized tailings sample within 24 h by using 1.5 M glycine at pH 12 and solid–liquid ratio of 50 g/L with 2 g/L copper concentration. In addition, the absence of pre-treatment under the same conditions yielded 87.9 % gold extraction. Statistical analysis indicated that solid–liquid ratio and pre-treatment were significant variables for the gold leaching. Furthermore, results from the statistical analysis showed that there was significant interaction between glycine concentration and copper ions during the leaching process. High gold extraction was observed to accompany an increase in concentration in the absence of copper; however, the reverse was true and higher extraction accompanied low (0.5 M) glycine concentration and the presence of copper. Although the solid–liquid ratio was below that expected to be economic for a tailings operation, the study has shown that pre-treatment and copper catalysis of the leaching may be opportunities to explore conditions necessary for economic extraction at higher solid–liquid ratio and low glycine dosages. In conclusion, glycine can leach gold from the Witwatersrand gold ore tailings under alkaline conditions with permanganate pre-treatment, but the economic feasibility of such a process requires further study. A potential opportunity for improving extraction of gold from glycine in such a system may be analogous to the carbon-in-pulp (CIP) process applied in cyanidation, in which continuous extraction of the gold complex in-situ may drive the reaction forward.

Acknowledgements

The authors thank and acknowledge the Royal Society and the African Academy of Sciences for funding through the FLAIR fellowship grant number FLRR119154. The authors would also like to acknowledge Ms Esther Nwagboso.

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Investigation of glycine leaching for gold extraction from Witwatersrand gold mine tailings

2ND BATTERY MATERIALS CONFERENCE 2024

5 AUGUST 2024 - WORKSHOP

6 -7 AUGUST 2024 - CONFERENCE

The intensified search over the past decade for alternatives to fossil fuels as sources of energy, has led to the development of a number of renewable energy technologies. A major issue with renewable energy sources is its intermittency which requires energy storage. This requirement has led to an exponential growth in the demand for batteries and research into battery technologies. The largest application by far has been in transportation, followed by the balancing of electrical distribution grids.

Of the raw materials required for battery manufacture, metals such as cobalt, manganese, vanadium and to a lesser extent nickel are concentrated in southern Africa. The supply of lithium, on the other hand, is mainly concentrated in Australia, Chile and Argentina with only Zimbabwe boasting significant resources in Africa. These activities have created both opportunities and challenges. Opportunities such as new value chains for the associated raw materials, with several production companies with battery-material metals in their plant feedstocks undertaking research towards producing battery-grade products. Challenges such as the means for recycling these batteries once they reach the end of their (first) life.

The aim of this conference is to provide the opportunity for thought leaders in the global battery value chain to exchange ideas on recent developments in the fields of:

• Materials and high-purity intermediates for battery components

– Novel battery chemistries

• Flow-battery electrolytes

• Processes for the recycling of batteries

• Market outlook and legislative implications

– New projects and entrepreneurship in the battery industry

• Related case studies.

An even sharper focus can be provided by addressing the following questions and hypotheses:

• Will future battery developments and applications in southern Africa centre more around bulk energy storage by drawing on regional metal resources and addressing local bulk energy shortages?

• Will lithium-ion batteries continue to dominate most battery applications, with other battery technologies serving only niche applications?

• What (exactly) are the criteria and specifications for battery materials, intermediates and electrolytes required to achieve the envisaged performance and life of the batteries?

– What impact will South Africa’s electrical supply issues have on the local motor manufacturing industry and the market for EVs?

NATIONAL & INTERNATIONAL ACTIVITIES

20-23 May 2024 — The 11TH World Conference of Sampling and Blending 2024 Hybrid Conference

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19-20 June 2024 — European Conference Hydrogen & P2X

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Beijing, China

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Worley

MINE PLANNING AND DESIGN COLLOQUIUM

5-6 SEPTEMBER 2024

VENUE: ELECTRA MINING NASREC, JOHANNESBURG

BACKGROUND

The Southern African Institute of Mining and Metallurgy (SAIMM) Mine Planning colloquiums have consistently highlighted deficiencies in mine planning skills over the years. The colloquium held in 2012, 2014, 2017 and 2019 all emphasized the need for developing skill-sets with various mine planning tools within the context of multiple mining methods.

Newer tools and skills for the future of mining discussed in the 2019 colloquium suggests an acknowledgment of the evolving nature of the industry. This likely includes advancements in technology, such as automation, artificial intelligence, and data analytics, which are increasingly becoming integral to modern mine planning and operations.

By addressing these deficiencies and adapting to newer tools and skills, the mining industry can better meet the challenges and opportunities presented by evolving technologies, and changing market demands.

REVIEW TOPICS

· Understand strategic plans and their impact to tactical and operational planning

• The importance of mineral resource estimation process and its influence of mine planning and design

• Minerals Processing

• The planning process design criteria – the starting point to good mine planning

• Optimizing the mine plan

ECSA Validated CPD Activity, Credits = 0.1 points per hour attended.

• Equipment selection

• Mine access – haul roads and declines – the neglected area of mining

• Financial technical evaluations and its importance to the mine planner.

• New technologies – Are they appropriate in mine planning.

CALL FOR PRESENTATIONS

Submit an abstract to be considered as presentations only at the colloquium.

Prospective presenters are invited to submit titles and abstracts of their presentations in English.

Only in-person presentations will be considered for this colloquium.

Abstracts should be no longer than 500 words and should be submitted to: Camielah Jardine, Head of Conferencing and Events, E-mail: camielah@saimm.co.za

The complete Proceedings volume will be made available on the internet for public access after the colloquium.

KEY DATES

1 May 2024 - Submission of Abstracts

1 August 2024 - Submission of Presentation

No marketing presentations will be considered. All presentations must include case studies of actual work implementation.

Delegates attending the Mine Planning and Design Colloquium will also have complimentary access to the Electra Mining Expo Centre on the colloquium days

FOR FURTHER INFORMATION, CONTACT:

Camielah Jardine, Head of Conferences and Events

E-mail: camielah@saimm.co.za Tel: +27 011 538 0237 Web: www.saimm.co.za

BACKGROUND

SANCOT SYMPOSIUM 2024

28-29 OCTOBER 2024

LESOTHO HIGHLANDS, LESOTHO

Sharing the Lesotho Experience

With the continued pace of urbanisation, economic and population growth, the availability of space for necessary infrastructure in the urban environment is becoming a major challenge. This is occurring in conjunction with climate change and a focus on reducing impact on the environment. These are the key factors driving the necessity and relevance of tunnelling, as tunnels are increasingly seen as a means to provide sustainable, safe and reliable provision of transport, electricity, gas, water, sewage facilities and extraction of raw materials.

Whilst the public and private sectors come to terms with the high capital expenditure required for tunnel construction, we live in an age of continued technological development and the application of these technologies presents an opportunity to better and more cost-effectively design, construct, and monitor tunnels. Furthermore, it is imperative that tunnelling consultants and contractors keep up to date with rapidly changing tunnelling technologies in order to remain viable in a competitive industry.

THEME

This conference concentrates on advances in the tunnelling industry, current best practice and how technology has improved tunnelling design, construction, supervision and monitoring. The conference is being held in Lesotho, and the ongoing Polihali Transfer Tunnel works can be visited by interested delegates. This project is a great analogy for the advances in tunnelling, as the ongoing tunnelling works (Polihali Transfer Tunnel – Phase 2) can be contrasted with the Phase 1 Lesotho Highlands Tunnel Projects, which were completed almost 30 years ago.

Gugu Charlie Conferences and Events Coordinator