Saimm 202201 jan

VOLUME 122 NO. 1 JANUARY 2022

Palabora Mine Dillon Marsh, a Cape Town based artist, explores the mining industry of South Africa in his series called For What It’s Worth. Mines speak of a combination of sacrifice and gain, constructed to extract value from the earth but also exacting a price. This image combines photography and computer-generated elements in an effort to visualize the output of the Palabora Mine.

The CGI sphere represents a scale model of the 4.1 million tonnes of copper removed from the ground. For more information about this project visit www.dillonmarsh.com

Photographs by Dillon Marsh

The Southern African Institute of Mining and Metallurgy OFFICE BEARERS AND COUNCIL FOR THE 2020/2021 SESSION Honorary President

Nolitha Fakude President, Minerals Council South Africa

Honorary Vice Presidents

Gwede Mantashe Minister of Mineral Resources, South Africa Ebrahim Patel Minister of Trade, Industry and Competition, South Africa Blade Nzimande Minister of Higher Education, Science and Technology, South Africa

President

I.J. Geldenhuys

President Elect Z. Botha

Senior Vice President W.C. Joughin

Junior Vice President E. Matinde

Incoming Junior Vice President G.R. Lane

Immediate Past President V.G. Duke

Honorary Treasurer W.C. Joughin

Ordinary Members on Council

Z. Fakhraei B. Genc K.M. Letsoalo S.B. Madolo F.T. Manyanga T.M. Mmola G. Njowa

S.J. Ntsoelengoe S.M. Rupprecht A.J.S. Spearing M.H. Solomon S.J. Tose A.T. van Zyl E.J. Walls

Co-opted to Members M. Aziz M.I. van der Bank

Past Presidents Serving on Council

N.A. Barcza R.D. Beck J.R. Dixon R.T. Jones A.S. Macfarlane M.I. Mthenjane C. Musingwini

S. Ndlovu J.L. Porter S.J. Ramokgopa M.H. Rogers D.A.J. Ross-Watt G.L. Smith W.H. van Niekerk

G.R. Lane–TPC Mining Chairperson Z. Botha–TPC Metallurgy Chairperson A.T. Chinhava–YPC Chairperson M.A. Mello––YPC Vice Chairperson

Branch Chairpersons

Johannesburg Namibia Northern Cape Pretoria Western Cape Zambia Zimbabwe Zululand

D.F. Jensen N.M. Namate Jaco Mans Vacant A.B. Nesbitt Vacant C.P. Sadomba 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)

Honorary Legal Advisers M H Attorneys

Auditors

Genesis Chartered Accountants

Secretaries

The Southern African Institute of Mining and Metallurgy Fifth Floor, Minerals Council South Africa 5 Hollard Street, Johannesburg 2001 • P.O. Box 61127, Marshalltown 2107 E-mail: journal@saimm.co.za

Editorial Board S.O. Bada R.D. Beck P. den Hoed I.M. Dikgwatlhe R. Dimitrakopolous* M. Dworzanowski* L. Falcon B. Genc R.T. Jones W.C. Joughin A.J. Kinghorn D.E.P. Klenam H.M. Lodewijks D.F. Malan R. Mitra* H. Möller M. Mullins C. Musingwini S. Ndlovu P.N. Neingo M. Nicol* S.S. Nyoni P. Pistorius P. Radcliffe N. Rampersad Q.G. Reynolds I. Robinson S.M. Rupprecht K.C. Sole A.J.S. Spearing* T.R. Stacey E. Topal* D. Tudor* F.D.L. Uahengo D. Vogt* *International Advisory Board members

Editor /Chairman of the Editorial Board R.M.S. Falcon

Typeset and Published by The Southern African Institute of Mining and Metallurgy P.O. Box 61127 Marshalltown 2107 E-mail: journal@saimm.co.za

VOLUME 122 NO. 1 JANUARY 2022

Contents Journal Comment: Quality over Quantity

by Q.G. Reynolds . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

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President’s Corner: Prediction is difficult

by I.J. Geldenhuys . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

v

Obituary: Professor Huw Ronald Phillips

by Wits University VCO News . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

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PROFESSIONAL TECHNICAL AND SCIENTIFIC PAPERS Mine-impacted water: From waste to resource

by A.N. Clay, S. Joubert, and N.N. Moeketsi . . . . . . . . . . . . . . . . . . . . . . . . . .

Discharged mine-impacted water from old mine operations can be a renewable and a sustainable resource, provided that it is managed in a responsible and sensible manner. In that regard, mine water management and conservation must be addressed by all stakeholders involved. To date, the State does not appear to have meaningfully explored considered using this resource to contribute to the country’s economic activity. This paper explores the conversion of mine-impacted water from waste to a resource and how the inclusion of other stakeholders could benefit in the process.

Silicomanganese alloy from rich manganese slag produced from Egyptian low-grade manganese ore

by H. El-Faramawy, M. Eissa, S.N. Ghali, T. Mattar, A. Ahmed, M. El-Fawakhry, and E.M. Kotb . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

Owing to the depletion of high-grade manganese ore reserves, it has become necessary has led to examining consider the use of low-grade manganese ores. The low Mn/Fe ratio in low-grade manganese ore makes it unfit for the production of manganese ferroalloys. In this work, a blend of rich manganese slag with medium manganese ore in the presence of coke and a fluxing agent was smelted in a submerged electric arc furnace to produce silicomanganese alloy. The silicomanganese alloy obtained has a composition close to standard specifications.

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Directory of Open Access Journals

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THE INSTITUTE, AS A BODY, IS NOT RESPONSIBLE FOR THE STATEMENTS AND OPINIONS ADVANCED IN ANY OF ITS PUBLICATIONS. Copyright© 2022 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.

JANUARY 2022

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PROFESSIONAL TECHNICAL AND SCIENTIFIC PAPERS Safety aspects of large dragline-operated opencast mines – An overview

by A. Golder and I. Roy . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

15

The Jayant opencast project of the Northern Coalfield Limited in India experienced a number of dragline dump failures up to 2008. Several design, research, and academic institutions carried out slope stability studies on the internal dumps. This paper presents a brief note on recommendations regarding preventive measures taken up by the project authority to tackle the safety aspects of dragline dumps.

The concentration of rare earth elements from coal fly ash

by G.B. Abaka-Wood, J. Addai-Mensah, and W. Skinner . . . . . . . . . . . . . . . . . . . . . . . . .

21

The specific aim of this study was to assess the technical feasibility of recovering rare earth elements (REE) from coal fly ash using conventional preconcentration methods. This study confirmed that existing physical separation methods could be used to recover REE in coal fly ash prior to hydrometallurgical and pyrometallurgical processing, but significant variations in performance were found in the various beneficiation methods investigated, and some challenges persist.

A new preconcentration technique for the determination of PGMs and gold by fire assay and ICP-OES A. Masasire, F. Rwere P. Dzomba, and M. Mupa . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

29

The determination and quantification of platinum group metals (PGMs) and gold (Au) in platiniferous ores is difficult due to their low concentrations. In this paper investigation, iron (Fe), cobalt (Co), and copper (Cu) were used as co-collectors to enhance fire assay separation and preconcentration by fire assay. While the analyzed results indicated that Fe and Co are better co-collectors than Cu., cheaper and more readily available co-collectors are quicker and yield comparable results. This method is therefore considered a viable alternative for the practical quantification of PGMs platinum group metals and gold.

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al

Journ

Quality over Quantity

ent

Comm

M

uch has been made of the ongoing impact of the SARS-CoV2 pandemic on human activities worldwide, and after nearly two years we are not out of the woods yet, as the recent emergence of the omicron variant has demonstrated. The scientific evidence is however increasingly telling us that the evolution of the virus combined with robust immune responses from prior infection or vaccination is progressively reducing the disease’s severity; will 2022 see the ‘beginning of the end’ of the pandemic? With some cautious optimism, it seems likely.

On this topic, an interesting question was posed to the SAIMM Journal’s publications committee several months ago. A reader observed that the number of papers being published in each edition has decreased in recent history. Pulling the data from the last few years and performing some basic analysis and visualization of the results showed that there has indeed been a statistically significant drop in the number of papers reaching final publication in the Journal, from roughly eleven per edition in 2019 to between six and seven in 2021. Interestingly enough, the number of papers submitted to the Journal was relatively unaffected over the same period.

Due to the timing it is tempting to present pandemic effects as the sole cause, but there are other considerations. Firstly, the decision taken by the Journal to cease republishing of selected conference proceedings papers occurred at approximately the same time – historically this has contributed several papers to most editions. Secondly, the peer review process was extended to include a pre-screening stage which has increased its rigour and quality considerably. These changes aim to bring you, the members and readers, an improved and more focused Journal in the new year and beyond.

Q.G. Reynolds Pyrometallurgy Division, Mintek Process Engineering Department, Stellenbosch University

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Presi

er Corn

Prediction is difficult

E

ven at the best of times, it is human nature to want to predict and decode the future.

During times of uncertainty we are even more eager to predict what is to come.

At the start of 2022, some believe (read predict) that the pandemic will move into a new phase and will become endemic, perhaps allowing for some normalcy to return to travelling, attending events, professionally and personally. Others, such as the World Health Organization (WHO) caution us that it is too early to know whether the worst is over and that it is still statistically possible for another delta-type variant to emerge and put pressure on health systems globally. Of course, both views are possible, and represent a high-road and a low-road scenario. For most of us, the dream of ‘normal’, even if it is not exactly as it was before, is what we hope for in 2022. We yearn for a time before we all became overnight online virologists, vaccine experts, and followers of @rid1tweets for COVID-19 trends (@rid1tweets is the Twitter account of Dr Ridhwaan Suliman, applied mathematician). Dr Suliman showed us that maths can be useful to understand what is happening around us – I take that as a win for STEM, but we still would rather get to a point where we don’t have to know how many 7-day average cases were reported in our area.

Our days are filled with a continuous barrage of information via all portals, and during the past two years of the pandemic, we sought out information to help us cope, and make decision related to our health and wellbeing, because what is happening to us on a global scale is unprecedented. Never before was there a global pandemic during the era of digital and social media. And as we know, fake news, misinformation, and ‘lies travel halfway around the world while the truth is (still) putting on its shoes’. This popular quote is often attributed to the American author Mark Twain, perhaps because he was so outspoken on many things, but there is actually no evidence that he said it first. Ha! you say, fake news at its finest! Indeed. Because Mark Twain is attributed to having said this, it has more impact, so it perpetuates and becomes lore. However, this does not devalue the idea or the words but supports my next premise well. A good quote can be a thing of beauty and can inspire us, remind us of core principles, or just make us smile. My love of quotes increasingly changed into a ‘who actually said it’ hobby. The quest for the original story, and the ability to reference and cite as accurately as possible is in my opinion part of the learning journey.

The fact that the ‘lies travel’ quote is popularly attributed to Mark Twain illustrates really well the challenge we have in our daily lives. It has become almost a fulltime task to evaluate the veracity of information. The burden of truth is indeed a heavy one. While we cannot track the origin of the quote, the story does not end here. There are some who believe the Mark Twain attributed version can be traced to an adaptation of the writings of Johnathan Swift. Swift was an Anglo-Irish satirist, essayist, political pamphleteer, poet, and Anglican cleric and in 1710 he wrote in The Art of Political Lying: ‘ Besides, as the vilest Writer has his Readers, so the greatest Liar has his Believers; and it often happens, that if a Lie be believ’d only for an Hour, it has done its Work, and there is no farther occasion for it. Falsehood flies, and the Truth comes limping after it; so that when Men come to be undeceiv’d, it is too late; the Jest is over, and the Tale has had its Effect.’

Falsehood flies, and the truth comes limping after. Even before the term ‘fake news’ had been coined, Swift observed that lies are believed very quickly. Once fake news is believed, and even if it only for a short time, the damage has been done. A lie travels halfway around the world before the truth puts on its shoes. By the time the limping truth catches up, it is too late. The lie has become lore.

‘Prediction is difficult – particularly when it involves the future.’ (Yet another common Mark Twain misquote, which is actually attributed to the Danish parliament, and a story for another day). But as hard as predicting the future might be, I predict that the lines between lore and reality will remain blurred for the foreseeable future. It is incumbent on us as the receivers of information to dig a bit deeper into the sources, to understand the backstory and the context, and to validate the data. And even with the best of intentions, remember: ‘We don’t see things as they are; we see them as we are.’ Not Mark Twain. I leave this one for you all to Google the origin.

I.J. Geldenhuys President, SAIMM

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Obituary: Professor Huw Ronald Phillips Passing of Professor Emeritus Huw Phillips 10 July 1947 – 26 January 2022 Dear Colleagues and Members We learnt recently of the passing of Professor Emeritus Huw Phillips, a Wits stalwart renowned for his teaching, research, and leadership in mining engineering the world over. An integral part of the Wits community, Professor Phillips left an indelible mark in his field, and a legacy which will undoubtedly influence future generations of scholars.

Professor Phillips served for over 27 years as a Full Professor in the School of Mining Engineering at Wits, during which time he supervised more than 20 PhDs and over 40 Masters students. A formidable administrator and leader, he served as the Head of Mining Huw Phillips, recipient of the 2013 Brigadier Engineering from 1986 for a period spanning over 25 years. He also Stokes Award, receiving his Platinum Medal from served as the Chair of Mining Engineering until his retirement in 2012. Gordon Smith He was appointed as an Emeritus Professor in 2013 and continued to supervise postgraduate students and to serve the University in various roles.

In 2013, Professor Phillips was announced as the winner of the Southern African Institute of Mining and Metallurgy’s prestigious Brigadier Stokes Memorial Award. He was also a Retired Fellow of the Southern African Institute of Mining and Metallurgy. He was lauded by the mining industry for his unique input and was honoured by the Institute of Mine Surveyors of South Africa, the Mine Ventilation Society of South Africa, and the International Society of Mining Professors.

A leading researcher in mine safety and health, Professor Phillips documented a formidable career’s worth of mining research achievements, which earned him a Doctor of Engineering degree from Wits in 2019.

This Senior Doctorate was awarded to Professor Phillips on the strength of his original work which was described as a record of engineering development of major technological, economic or social significance and a distinguished contribution to the practice of engineering. His research at Wits spanned five key areas: mechanised mining systems; spontaneous combustion; mine ventilation in deep-level gold mines – including software tools for designing cooling strategies; monitoring and controlling respirable dust in coal mines; and preventing methane ignitions and coal dust explosions. His work will prove to be an invaluable resource for students of mining engineering and mining professionals for many years to come.

Born in Wales, Professor Phillips studied electrical engineering at the University of Bristol and took his first job with the National Coal Board in the UK. His interest in mining led him to complete an MSc and PhD in mining engineering at the University of Newcastle-upon-Tyne. His work and research in the coal sector continued as he relocated to the University of New South Wales in Australia in 1977, where he served for eight years. After spending his sabbatical leave in South Africa in 1981 with the Chamber of Mines Research Organisation, he returned to South Africa in 1985 as the Chamber of Mines Professor of Mining Engineering at Wits. During his time as Head of Department and subsequently Head of School, he dramatically increased undergraduate and postgraduate enrolments, and secured industry funding for research.

Professor Phillips dedicated much of his life to teaching, research, and making mines safer for thousands of miners. His work will live on for decades to come, and serves as an inspiration for future scholars in mining engineering. Please join us in extending our heartfelt sympathy to the family, friends, colleagues, students and industry partners who knew Professor Phillips well. We are with you during this difficult time but we also know that his legacy will outlast our grief. May he rest in peace.

Published with kind permission of Wits University VCO News

Mine-impacted water: From waste to resource

by A.N. Clay1, S. Joubert1, and N.N. Moeketsi1

Affiliation:

EY Minvest.

1

Correspondence to: N.N. Moekets

Email:

naledi.n.moeketsi@za.ey.com

Dates:

Received: 3 Mar. 2021 Revised: 23 Nov. 2021 Accepted: 1 Dec. 2021 Published: January 2022

How to cite:

Clay, A.N., Joubert, S., and Moeketsi, N.N. 2022 Mine-impacted water: From waste to resource.

Journal of the Southern African Institute of Mining and Metallurgy, vol. 122, no. 1, pp. 1-4

Synopsis For many years, mine-impacted water has been regarded as a problem and linked to long-term environmental liabilities. However, this water can be a renewable and a sustainable resource, provided that it is managed in a responsible and sensible manner.

South Africa’s National Water Resource Strategy (NWRS, 2013) considers water that can be used to contribute to economic activity to be a water resource. Although all water resources are considered as belonging to the State, government does not appear to have meaningfully explored the use of mineimpacted water to contribute to the country’s economic activity.

Africa is blessed with more sunshine than anywhere else, apart from Australia, yet we see no rollout of vast quantities of solar panels to ensure every African family has access to affordable power. This suggests that governments are incapable of managing such a free resource. At the same time, apart from the equatorial areas, water is a seriously constrained resource and yet we expect the same entities to manage a commodity none of us can live without.

This paper explores the conversion of mine-impacted water from waste to a resource and how the inclusion of other stakeholders (such as water users, landowners, and ordinary South Africans) could benefit the process.

Keywords mine-impacted water, water balance, water resources.

DOI ID: http://dx.doi.org/10.17159/24119717/1556/2022

Introduction Mining activities have been accompanied by a legacy of environmental issues resulting from the discharge of mine-impacted water from old mine operations. For many years, this mine-impacted water has been regarded as a residual problem and linked to long term environmental liabilities. Reclamation activities are governed by widespread national and international regulatory structures. In this regard, assuring compliance with national legislation and related regulations and following best practice international guidelines on mine closure and rehabilitation is vital.

Waste water and its treatment account for the biggest portion of the closure and rehabilitation liability of a mine. The National Environmental Management Act No. 107 of 1998 (NEMA) requires that this liability, including the responsibility for extraneous or polluted water, continues after closure. The inclusion of this requirement, setting out how to calculate the financial provision, is a clear indication that the quantum of the provision will necessarily increase over time. As a consequence, this will reflect as a significant liability on the balance sheet.

Mining as a polluter The historical management of mine rehabilitation in South Africa by the main mining companies has been perfunctory to say the least, hence the current situation of many abandoned sites. The mining industry has developed a bad reputation in this regard, with few examples of successfully rehabilitated sites. The creation of ‘rehabilitation trust funds’ was a concept and practical approach to ensuring companies contributed to the financial provisioning of the closure costs and was incorporated into Section 37A of the Income Tax Act No. 62 of 1968 (the ITA). This meant that contributions to the fund should be provided evenly by taking the terminal liability and dividing it by the life of mine.

During times when metal prices were high and companies were generating profits, the preferred option was to contribute more, and when profits were low, to decrease contributions. This was The Journal of the Southern African Institute of Mining and Metallurgy

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Mine-impacted water: From waste to resource counterintuitive, since during times of high prices the pay limits should go down and the life of mine would increase, meaning that the denominator increased and so contributions should also decrease. In some instances, companies would make provisions which were not ‘cash’, and so when a mining company was no longer profitable the provision disappeared as well. The accounting process also posed challenges in that, in a high inflationary economic climate, the rehabilitation liabilities were escalated at inflation and then discounted at the company’s cost of capital, which was generally more than the inflation rate. The resulting present value of the liability was less than the original estimate.

The government has since stepped in to ensure that this situation doesn’t persist, but the intervention is clearly too late.

Consequently, it is nigh impossible for any mining company to obtain a closure certificate. The obvious result is that mine owners will simply go into insolvency and then we are back to square one.

It should be remembered that during the process of State capture and takeover of Optimum Coal Mine the then Minister of Mineral Resources, Mosebenzi Zwane, authorized the transfer of the R1.4 billion rehabilitation trust fund (all trust funds need ministerial permission for expenditure) to an offshore Gupta-based business, never to be seen again.

The identification and quantification of rehabilitation liabilities is a fairly reliable process, and most companies take this matter seriously as part of their commitments to environmental, social, and corporate governance (ESG). These amounts are embedded in the financial statements under IAS 37 (Provisions, Contingent Liabilities and Contingent Assets) and scrutinized by activist bodies.

There is a conundrum facing stakeholders – how to make sure the rehabilitation or decommissioning costs are clearly identified and defined as a ‘liability’ while at the same time creating a financing process to concert the supposed ‘liability’ into an ‘asset’ in an orderly and responsible manner. So, where does water fit into the problem? In general, water inflows and outflows can be readily identified for a mine in the water balance, which is crucial during operations in order to ensure the mine is dewatered and has enough pumping capacity to meet routine and exceptional circumstances such as a flood. Once mining operations have ceased, the mine is usually sealed up and the water in the workings left to decant via natural watercourses or aquifers flow.

That issue should be nothing more than restoring the situation to the pre-mining position, except that the mine void, which is now a new ‘reservoir’, connects to the natural aquifers. The water is often polluted as a result of exposure to minerals such as pyrite and other sulphides that generate acidic solutions with iron as a major pollutant. This is commonly called ‘acid mine drainage’ (AMD) (Simate and Ndlovu 2014).

An example of AMD is the decanting of polluted water from the coal mines in the Witbank coalfield and surroundings into the Crocodile River ecosystem, creating significant downstream problems. One could say that this was inevitable, but the question is what can be done about it now?

Legalities and regulatory issues Closure and remediation activities are governed by widespread legislative structures in various countries. Inclusive within the framework are international treaties and protocols, national acts, regulations, standards, and guidelines that address the management and financial accounting requirements. These requirements inform the input parameters that influence the Asset 2

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Retirement Obligation (ARO) as recognized in the International Accounting Standard (IAS) 37. The consideration of all aspects that have the potential to influence the ARO quantum is essential in understanding current ARO requirements, engineering solutions, and remediation liability.

In the South African context, the promulgation of the new Financial Provision Regulations (GNR1147) on 20 November 2015 in terms of NEMA resulted in a significant shift from Regulations 53 and 54 of the Mineral and Petroleum Resources Development Act No. 28 of 2002 (MPRDA). The purpose was to regulate the creation of financial provisions as contemplated in the Act for the costs associated with management, rehabilitation, and remediation of environmental impacts from prospecting, exploration, mining, or mineral production. Regulation 6 of GNR1147 requires that ‘An applicant must determine the financial provision through a detailed itemisation of all activities and costs, calculated based on the actual costs of implementation of the measures required for: ➤ Annual rehabilitation, as reflected in the annual rehabilitation plan ➤ Final rehabilitation, decommissioning and closure of the prospecting, exploration, mining or production operations at the end of the life of operations, as reflected in a final rehabilitation, decommissioning and mine closure plan and ➤ Remediation of latent or residual environmental impacts which may become known in the future, including the pumping and treatment of polluted or extraneous water, as reflected in an environmental risk assessment report.’

The current legislative requirements revolve around providing for water-related liabilities during operations and post closure, and no allowance is made for converting waste water into a resource that can ultimately be regarded as an asset. We believe that water can be a renewable and sustainable resource provided that it is managed in a responsible and sensible manner. If it isn’t, it is likely that water scarcity will lead to next round of global conflict if government is not on board. In recent years, many countries have been experiencing serious drought and water shortages, a good case in point being Cape Town in South Africa.

The authors have identified a specific need to classify water beyond the limits of that which would be required for national strategic planning. In this regard, the commercial imperative is to link the human right of access to water to the efficient utilization of various water resources. To this end, the very principle and philosophy of ownership is at the heart of the classification system, which in itself is intimately related to the cost of producing various types of water and the price at which it can be sold.

Many approaches have been taken to planning water management, primarily driven by government agencies. However, in many ways this is similar to other national planning issues including road, rail, health, retirement planning, farming, and air pollution. As a result, the governmental process is invariably in conflict with the realities of commerce, since the concept that ‘nothing is for free’ is an increasing global human population problem.

Tranformation from liability to asset The purpose of this paper is to outline a simple solution to the problem of mine-impacted water and the management of the long-term effects, technically and financially. To begin with, all mines must have a ‘water balance’ that accurately identifies The Journal of the Southern African Institute of Mining and Metallurgy

Mine-impacted water: From waste to resource and quantifies the inflows and outflows of water and a pumping capacity that maintains the appropriate water level for operating the mine.

An example of a water balance is shown in Figure 1. In general, a qualified person should be responsible for preparing this and a geohydrologist should be involved. At the centre of the process is the ability to create a stochastic link between the aquifers, which are generally recharged from rainfall, and the climatic conditions as well as the open volumes constituting the underground workings or open pits. Essentially the mine void is the ‘reservoir’.

It is the latter which forms the basis of the definition of an asset. The mine void has been created by the mining company, and is therefore an asset of the mine, but where do you ever see it booked as an asset? The nearest you may get is the accounting treatment of ‘deferred stripping’, whereby the cost of creating the pit is capitalized onto the balance sheet as an asset and then written off as the mine extracts ore (mineable reserves) in the production plan on an annual basis. When the void is exhausted it is seldom possible to completely backfill it, and so another ‘reservoir’ has been created as an asset.

The fact that the pit or underground workings can fill with water becomes a sustainability opportunity. It is quite ironic that the historical view is that a mine is a wasting asset, and once the mineral resources have been depleted there is nothing left of value. The approach suggested here challenges this philosophy, especially when the water balance demonstrates that a reliable, sustainable water volume is available for exploitation.

of information, namely the inflows, outflows, and reservoir volume, can be reconciled to a high level of accuracy. It is normal to estimate the water volumes in megalitres per day and then to annualise this quantity.

For recharge, annual rainfall statistics can be used from weather records and in this example, we used @RISK to calculate probabilities in the same manner as in the oil and gas industry.

The probability function can then be used to calculate the 10%, 50%, and 90% confidence limits for the best, mean, and least confident estimates of volumes, and a water resource declared.

In mineral resource estimation it is a fundamental principle to ensure that there are ‘reasonable and realistic prospects for eventual economic extraction’ as a prerequisite to signing off a resource. Therefore, understanding the pumping and water reticulation structure is crucial in the case of a water resource.

This information is generally available from the mine engineer, with records of pumps, pipe network, and power within the mine infrastructure.

After considering recharge, the ability to confidently define water volumes ,with a high level of confidence is relatively easy.

Financial engineering Once the water balance and the volumes have been estimated using probabilistic methods, the next question is how to obtain an asset value for the water. The first course of action is to seek a ‘value in use’ method where the sustainable volume of water can be pumped into a user network and distributed for sale, just like any other mineral. The quality of the water is a key factor governing the ability to sell the water but, in the case of polluted water, constructing a water purification plant of an appropriate nature is a simple engineering process.

For financial engineering, all that is needed is the cost of pumping, treatment, and distribution, and determination of the

Estimating volumes The water balance allows water volumes to be accurately quantified as inflows and outflows. This can be cross-referenced to the survey estimates of the volume of the mine working void from mine plans and production records. These three sources

Clean water

Fresh water Process

Precipitation and Drainage Re-circulated water

Process waste + water Dewatering water

Precipitation and Drainage

Waste rock piles

Tailings pond

Settling pond

Water treatment facilities

Back to process Household water use Directed to nature

Figure 1—Example of different minewater sources and flows (Punniken et al., 2016) The Journal of the Southern African Institute of Mining and Metallurgy

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Mine-impacted water: From waste to resource sales price. If these elements are incorporated into a cash flow model with escalation and taxation, the discounted cash flow method allows one to estimate the value of the project.

A key issue here is that the government regulates the mine’s responsibility for the rehabilitation and closure costs, and water has generally been regarded as a legacy issue that essentially never goes away. So, in fact, with the concept of ‘matching’, the use of the water resource in a sustainable financial solution should at least mitigate the monetary aspect of the rehabilitation liability, let’s say on a defined timescale such as ten years. If the net effect is to zero that out, then the sale price of water can be calculated.

Although not dealing with minewater, many countries licence private water utilities so that the efficient and sustainable use and distribution of water is financially sustainable.

Accounting recognition As many of the decommissioning and closure activities of mining operations have the potential to vary on a year-on-year basis, the International Financial Reporting Standards (IFRS) require that an Asset Retirement Obligation (ARO) is recognized within the auspices of IAS 37.

The scope of IAS 37 requires that a provision be recognized if: ➤ A present obligation (legal or constructive) has arisen as a result of a past event (the obligating event) ➤ Payment is probable (‘more likely than not’) ➤ The quantum can be estimated reliably.

When determining the ARO for a mining entity, the following aspects must therefore be considered: ➤ Legal obligations for decommissioning and demolition ➤ Legal obligations for closure and post-closure ➤ Legal obligations for meeting environmental and social rehabilitation commitments ➤ Constructive obligations for decommissioning and closure (for example, company policy, lender requirements, or similar obligations).

In light of the above, we believe that the Accounting Standards can be improved such that a water resource can be recognized on the balance sheet as an asset instead of a liability.

Definitions Aquifers—Underground rock strata that are saturated with water that can be brought to the surface through natural springs or by pumping (Oskin, 2018).

Asset—A resource controlled by [an] entity as a result of past events and from which future economic benefits are expected to flow to the entity (International Financial Reporting Standards, 2008).

Mine Closure—A period of time when the operational stage of a mine is ending or has ended, and the final decommissioning and mine rehabilitation is being undertaken (Australian Government: Department of Industry, Tourism and Resources, 2006).

Liability—The future sacrifices of economic benefits that the entity is obliged to make to other entities as a result of past transactions or other past events, the settlement of which may result in the transfer or use of assets, provision of services or other yielding of economic benefits in the future. (International Financial Reporting Standards, 2008).

Provision—A liability of uncertain timing or amount.

(International Financial Reporting Standards, 2008).

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Rehabilitation—The restoration of the post-mined landscape to the intended post-mining land use (Mine Closure and Completion, 2006).

Reservoir—An enclosed area for the storage of water to be used at a later date.

Water resource—Water that can be used to contribute to economic activity, including a watercourse, surface water, estuary, and groundwater in an aquifer (NWRS, 2013).

Water balance—The regulation or rationalization of human activity to match the sustainable local water supply, rather than base, or a process of balancing water supply and demand to ensure that water use does not exceed supply (NWRS, 2013).

Conclusion This paper is intended to serve as a transformation initiative for one of the most critical resources in Africa and the world in general. In the past, mine tailings were considered to be a pollutant and historical tailings with their associated dust and pollution issues were never viewed as an asset. However, that changed in the 1970s when commodity prices rose significantly and mining operations expanded, and with the recognition of retreatment liabilities the retreatment of tailings dams became possible, especially in South Africa.

Mining creates a number of underground water reservoirs.

This paper outlines an opportunity to transform reservoir volumes, through an acceptable water balance, into an asset that can be booked onto the balance sheet.

This is a work in progress for which the concept is clear and the mechanism of defining the quantities and qualities of water is relatively simple. A water resources and reserves standard is in preparation and will be reported upon as a follow-up to this paper in the near future.

In the meantime, creating financial engineering solutions that link the water quantities as assets with the rehabilitation liabilities still remains to be completed.

References Australian Government. 2006. Mine Closure and Completion. Department of Industry, Tourism and Resources. https://www.im4dc.org/wp-content/ uploads/2014/01/ Mine-closure-and-completion.pdf [accessed 11 November.2020].

Department of Water Affairs. 2013. National Water Resource Strategy. http:// www. dwa.gov.za/documents/Other/Strategic%20Plan/NWRS2-Final-emailversion. pdf. [accessed 11 November.2020].

International Financial Reporting Standards. 2008. https://www.ifrs.org/ issuedstandards/ list-of-standards/ [accessed 11 November.2020].

Responsible Jewellery Council. 2006. Leading Practice Sustainable Development Program for the Mining Industry (Australia). Mine Closure and Completion. https://www.responsiblejewellery.com/wp-content/uploads/ Mine-Rehabilitation-and-Closure-RJC-Guidance-draftv1.pdf [accessed 11 November.2020].

Oskin, B. 2018. Aquifers: Underground stores of freshwater. https://www. livescience. com/39625-aquifers.html [accessed 11 November.2020].

Punkkinen, H., Räsänen, L., Mroueh U., Korkealaakso, J., Luoma, S., Kaipainen.

Backnäs, S., Turune, K., Hentinen K., Pasanen, A., Kauppi, S., Vehviläinen, B., and Krogerus, K. 2016. Guidelines for mine water management. https:// www. vttresearch.com/sites/default/files/pdf/technology/2016/T266.pdf [accessed 11 November.2020].

Simate, G.S. and Ndlovu, S. 2014. Acid mine drainage: Challenges and opportunities. Journal of Environmental Chemical Engineering. doi: 10.1016/j. jece.2014.07.021 Votruba, L. and Broža, V. 1989. Basic function of water reservoirs. Developments in Water Science, vol. 33. pp. 19–60. u The Journal of the Southern African Institute of Mining and Metallurgy

Silicomanganese alloy from rich manganese slag produced from Egyptian low-grade manganese ore

by H. El-Faramawy1, M. Eissa1, S.N. Ghali1, T. Mattar1, A. Ahmed1, M. El-Fawakhry1, and E.M. Kotb1 Affiliation:

Central Metallurgical Research and Development Institute (CMRDI), Cairo, Egypt.

1

Correspondence to: S.N. Ghali

Email:

a3708052@gmail.com

Dates:

Received: 2 Jun. 2021 Revised: 23 Sep. 2021 Accepted: 29 Nov. 2021 Published: January 2022

How to cite:

El-Faramawy, H., Eissa, M., Ghali, S.N., Mattar, T., Ahmed, A., El-Fawakhry, M., and Kotb, E.M. 2022 Silicomanganese alloy from rich manganese slag produced from Egyptian low-grade manganese ore.

Journal of the Southern African Institute of Mining and Metallurgy, vol. 122, no. 1, pp. 5-14 DOI ID: http://dx.doi.org/10.17159/24119717/1648/2022 ORCID: H. El-Faramawy https://orcid.org/0000-00021828-0501 M. Eissa https://orcid.org/ 0000-00018688-0843 S.N. Ghali https://orcid.org/ 0000-00029195-0663 T. Mattar https://orcid.org/ 0000-00025659-208X A. Ahmed https://orcid.org/ 0000-00022264-4195 M. El-Fawakhry https://orcid.org/ 0000-00021258-8492 E.M. Kotb https://orcid.org/0000-00027488-6006

Synopsis As a result of the depletion of high-grade manganese reserves, it has become necessary to consider the exploitation of low-grade manganese ores. The main problem of low-grade manganese ore is the low Mn/Fe ratio, which makes it unfit for the production of manganese ferroalloys. In this work, a selective reduction technique was used for smelting low-manganese ore in the presence of coke and fluxing material in an electric submerged arc furnace to obtain a rich manganese slag and low-manganese pig iron. The product slag was blended with medium-manganese ore and smelted in an electric submerged arc furnace, in the presence of coke and the fluxing agent, to produce silicomanganese alloy. The influence of reducing agent ratio, charge basicity, and charge Mn/Si ratio on the smelting process was investigated.

The optimum conditions were found to be a coke ratio of, 1.35, Mn/Si ratio of 2.0, and charge basicity of 2.5. The silicomanganese alloy produced under these conditions satisfies the specifications for Si16Mn63 and Si17Mn65. The experimental results were applied on a pilot scale, producing a silicomanganese alloy with a chemical composition close to that of the standard specifications.

Keywords silicomanganese, low-grade manganese ore, medium-manganese ore, carbothermic reduction, submerged arc furnace.

Introduction Silicomanganese alloy is used as a source of silicon and manganese in steelmaking to produce different steel grades. It is also used as a deoxidizer in place of both ferrosilicon and ferromanganese. The consumption of silicomanganese is increasing as a result of growing steel production. Steelmakers prefer to utilize silicomanganese instead of both FeSi and FeMn due to several advantages, such as its low cost and greater effectiveness as a deoxidizer. Also, when used with aluminium to produce an effective complex manganese-silicon-aluminium deoxidizer, it causes less contamination from phosphorus, carbon, aluminium, and nitrogen in steel compared with the FeSi and FeMn mixture (Ahmed et al., 2014).

The Om Bogma manganese deposit in Egypt contains reserves of about 1.7 Mt, most of which are low-grade ores. Many studies have been carried out using ore dressing to separate the manganese oxides from the ore. The separation of manganese impurities by magnetizing roasting followed by low-intensity magnetic separation is not possible, owing to the mineralogical complexity of the ore. Unfortunately, ore dressing studies have proved that the elimination of manganese oxides present in iron ore is not successful.

In recent years, a programme of work has been initiated in the Steel and Ferroalloys Department, Central Metallurgical Research and Development Institute (CMRDI), aimed at the selective reduction of low-grade manganese ore in an electric submerged arc furnace to separate pig iron and produce a high-manganese slag for the production of manganese ferroalloys. The use of such slag in the production of silicomanganese alloys is very attractive from an economic point of view. High-manganese slag is characterized by a high manganese content, low excess oxygen, high Mn/Fe ratio, low fines content, low phosphorus content, and low cost (Eissa et al., 2004). Mn-rich slags produced from the injection of highmanganese pig iron were found to have levels of Mn >35%, (Mn)/(Fe) >7.65, (Mn)/(P) >285, and with both Mn- and Fe-oxides existing as lower oxides (MnO and FeO) (El-Faramawy et al., 2004).

Generally, silicomanganese is produced in an electric arc furnace by carbothermic reduction of manganese ores, high-manganese slag, and quartz in the presence of fluxing materials. The process in which high-carbon ferromanganese is produced using acidic rich-slag practice with subsequent

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Silicomanganese alloy from rich manganese slag silicomanganese production from the high-Mn slag is sometimes referred to as the duplex process. This process is different from the other practice, in which only Mn ore is utilized as source of manganese and the product slag is discarded (Steenkamp and Basson, 2013).

The most important reactions taking place during the process, the reduction of MnO and SiO2, are shown in Equations [1]–[3] (Jana, and Randhawa, 2009; Olsen, Tangstad, and Lindstad, 2007). The components such as MnO and SiO2 can be reduced simultaneously during the smelting process.

[1] [2]

[3]

Equations [1] and [2] are independent reactions and can be combined to form Equation [4].

[4]

At the process temperature of 1600°C, reduction of MnO to metallic manganese by solid carbon occurs according to Equation [2]. However, the reduction of SiO2 to metallic silicon by solid carbon (Equation [1]) cannot occur due to the high free energy requirement at that temperature (Olsen, 2016). The reduction of SiO2 to silicon metal by solid carbon at 1600°C can only take place in the presence of metallic Fe and Mn, which lower the activity of silicon in the alloy (Davidsen, 2011). According to Equation [3], SiO2 reduction occurs when the activity of SiO2 in the slag is suffiiciently high due to the decrease of MnO activity caused by the complete reduction of MnO (Kor, 1979). The other components, such as CaO, Al2O3, and MgO, remain stable during the smelting process and inflluence the activity of both SiO2 and MnO in the slag.

The formation of silicon carbide has to be considered in the silicomanganese process. The carbon concentration in the silicomanganese alloy is affected mainly by the silicon content.

With increasing silicon content in the alloy, carbon becomes less stable and carbides become the dominant stable phase, which leads to a decrease in the concentration of dissolved carbon in the alloy (Alex et al., 2007). The percentage of carbides increases with increasing total carbon in the alloy. In addition, temperature also has an important effect on the carbon concentration in two ways. A higher temperature increases the carbon solubility at constant silicon content, and also increases the silicon content, which therefore decreases the carbon concentration (Anacleto, Ostrovski, and Ganguly, 2006). Thermodynamically, when the silica activity in the slag exceeds 0.2 at 1600°C, silicon carbides start to form according to Equation [5] (Olsen, and Tangstad, 2004); [5]

Silicomanganese slag basicity (B) is expressed as the ratio between the basic and acid slag constituents, as illustrated in Equation [6]. According to equilibrium studies by earlier researchers, an alternative measurement of slag basicity, the R-ratio (Equation [7]) was used due to its strong inflluence on SiO2 activity in the slag and, therefore, its signifiicant effect on the alloy composition (Tang, and Olsen, 2004; Ding and Olsen 2000) 6

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.

[6] [7]

where all constituents are expressed in terms of weight per cent.

Hence the composition and thermodynamic properties of slag are vital in determining the alloy composition, as they affect the distribution of Mn and Si across the slag-metal interface and thus the recovery of the metals in the desired ratio (Olsen, 2016; Tranell et al., 2007).

Nikolaev (1974) reported the optimum R-value to be 2.5, whereas Emlin et al., (1986) found the optimum R-value to be in the range of 1.2–2.2. These researchers also found that the basicity of slag was best adjusted through the addition of dolomite, which gives higher silicon- and manganese recoveries compared with the addition of limestone only or with dolomite. Other equilibrium studies (Ding and Olsen 2000; Olsen and Tangstad, 2004; Tang and Tangstad, 2007) concluded that the R-ratio strongly influences the SiO2 activity of the slag and thereby the Si content of the alloy.

As the basic oxides content is increased by the addition of dolomite, the interaction between Ca2+, Mg2+, and silicate ions becomes stronger than that between Mn2+ and silicate ions, resulting in the formation of stable Ca and Mg silicates (Kubaschewsky, Evans, and Alcock, 1979; Barin, 1989). This gives rise to free Mn2+ ions, which kinetically associate, with free O2ions in the slag and increase aMnO in the slag. The high MnO in the slag favours the transfer of Mn into the metal. In contrast, aSiO2 in the slag is diminished by the addition of basic oxides, which adversely affects the Si content of the alloy. Therefore, it is clear that maintaining the appropriate concentrations of both basic oxides and silica is required in order to yield the desired grade of SiMn alloy (Barin, 1989).

Silicomanganese smelting is carried out in an acidic slag with basicity lower than unity (B < 1) and defined as B = (CaO, wt%. + MgO, wt%)/(SiO2, wt%). Several studies have been done on the effect of slag basicity on Mn and Si distribution. The distribution ratio of an element is defined as the percentage of that element in the slag phase divided by the percentage in the metal phase.

Equilibrium studies on the distribution of Mn between Mn-SiFe-C alloys and MnO-CaO-MgO-SiO2-Al2O₃ slag at 1500°C under CO atmosphere revealed that an increase in the basicity ratio of the slag decreases the Mn distribution ratio (Cengizler, 2003).

In contrast, by increasing the silica concentration of the slag the Mn distribution ratio increases. The addition of alumina to the slag favours the transfer of Mn to the alloy, leading to a low MnO content in the slag. Cengizler, Eric, and Reuter (1997) modelled activity data for ferromanganese slags by applying neural nets at 1500°C for slag compositions in the range MnO 5–40%, CaO 4–35%, MgO 0.3–38%, SiO₂ 25–60%, and Al₂O₃ 2.5–7%. They found that the MnO activity coefficient in liquid slag varies on both sides of unity. Tang, and Olsen (2006) and Ding and Olsen (2000) concluded that the addition of alumina to acidic slag decreases the equilibrium MnO content in the slag. They also found that the activity coefficient (Ý) of MnO is less than unity in acidic slags, whereas ÝMnO > 1 in basic slags.

Viscosity is one of the most important properties of the slag and a critical parameter for many smelting processes. The viscosity of slag depends on slag composition, oxygen partial The Journal of the Southern African Institute of Mining and Metallurgy

Silicomanganese alloy from rich manganese slag pressure, and temperature. Viscosity, being the viscous resistance of the melt in the flow process, predominantly relies on the large, complex silica anions (e.g., SiO44-, Si₂O76-, and Si₇O108-).

Hence, a high SiO₂ content will increase the viscosity. Silica forms complex network crystals containing Si-O bonds, resulting in low recoveries of valuable metal from the slag. The network structure is destroyed if an appropriate amount of basic oxides is added to the molten slag (Olsen, Tangstad, and Lindstad, 2007), which promotes the further recovery of metals from the slag. In absence of a flux (i.e. dolomite) the Mn ore, and recycled slag are the sources of the CaO, MgO, and Al₂O₃, which form slag upon melting.

The difficulty and high cost of measuring slags viscosity have led to the development of numerous viscosity models. Viscosities and densities of typical HCFeMn and SiMn slags were calculated at a standard operating temperature of 1673 K for HCFeMn slags and 1873 K for SiMn slags. The calculations considered the percentage solids derived from the equilibrium phase composition determined with FactSage (Muller, Zietsman, and Pistorius, 2015). In this work, Weymann-Frenkel’s equation was used for determining the slag viscosity as a function in temperature and chemical composition (Ray and Pal, 2004) as given in Equation [8].

[8]

where  is viscosity, T is temperature, and A and B are independent parameters calculated according to the Urbain model (Rosypalová et al., 2014) This investigation aims at studying the possibility of using high-manganese slag obtained from Egyptian low-grade manganese ore in the production of standard silicomanganese alloy. Different parameters affecting the production process were studied using a bench-scale submerged arc furnace, and the

obtained optimum conditions were implemented on the pilot plant scale.

Experimental About 1 t of manganese-rich slag was produced on the pilot scale in the Steel and Ferroalloys Technology Department at the Central Metallurgical Research and Development Institute by selective reduction of low-grade manganese ore and used in the production of silicomanganese alloy.

Manganese ore, quartzite, dolomite, bauxite, and coke nut were suppled from Sinai Manganese Company. The raw materials were crushed in a jaw crusher to –55 mm (manganese-rich slag and medium-grade manganese ore), –35 mm (quartzite, dolomite), –10 mm (bauxite), and –25 mm (coke nut). All the raw materials were analysed by X-ray fluorescence. The chemical compositions of the raw materials are listed in Table I. Manganese minerals are complex and orebodies are also typically comples, hence the manganese ore composition is expressed in terms of Mn only, except where mineralogical information is available that will determine the manganese distribution between Mn/Mn2+, Mn3+, and Mn4+ (Steenkamp et al., 2020). The Mn-rich slag and mediumgrade Mn ore ware also subjected to XRD. Nineteen experimental heats were carried out – sixteen on the bench scale and three on the pilot plant scale – to investigate the influence of reducing agent ratio, charge basicity, and charge Mn/ Si ratio on metal yield and recoveries of Mn and Si. The benchscale SAF (Figure 1) is an open furnace operated with a reducing atmosphere. The power is supplied to the furnace through an AC stepwise transformer with primary electric power of 380 V and 220 A. The furnace transformer operates at a maximum current of 2000 A at different voltages ranging from 0 to 70 V passed through two 40 mm graphite electrodes that can be raised and lowered manually. The furnace shell is 340 mm in diameter and

Table I

Chemical composition of the raw materials (wt%) Constituents Mn-rich slag

Medium-grade Mn ore

Quartzite

Dolomite

Bauxite

Coke

MnO₂ 50.56 43.7 MnO 20.05 Mntot.

39.17 42.9 0.286 0.73 0.90 Fe₂O₃ 5.36 19.0 FeO Fetot.

4.17 13.3 98.338 9.43 1.50 SiO₂ 13.18 6.12 CaO 5.12 3.62 0.430 27.00 0.10 MgO 5.587 1.15 0.0450 16.65 0.05 Al₂O₃ 3.40 4.00 0.212 0.44 85.00 P 0.067 0.45 K₂O 2. 89 0.297 0.01 0.02 Na₂O 2.965 0.243 0.02 0.10 BaO 4.02 1.25 V₂O₅ 1.80 0.011 TiO₂ 3.50 0. 022 0.22 2.24 1.50 0.20 S 0.52 0.001 ZrO₂ 0.25 0.21 42.56 10.00 LOI Fixed carbon 85.47 Ash content 12.90 V.M 0.58 Moisture 0.11 0.80 Hum. 0.34 Total 99.219 99.914 99.851 99.42 99.17 99.95 Mn/Fe 9.39 3.23 The Journal of the Southern African Institute of Mining and Metallurgy

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Coke ash

0.68 15.26 46.18 2.97 0.27 32.87 0.030

0.98 0.132 0.030

99.40

7

Silicomanganese alloy from rich manganese slag

Figure 1—Schematic diagram of the bench-scale SAF

300 mm in height. The shell wall and bottom were lined with a thick rammed magnesite layer, using dead-burned magnesia with the chemical composition MgO 88%, CaO 7%, SiO₂ 1%, Al₂O₃ 1%, Fe₂O₃ 3%, and grain size 0-5 mm, supplied by Delta Steel Mill Company. The thickness of the magnesite layer was 60 mm for the wall and 100 mm for the bottom. The magnesite layer was lined with an inner lining of carbon paste, supplied by Sinai Manganese Company as unbaked paste with the composition fixed C 82%, volatiles 13%, ash 4.6%, P 0.02%, and S 0.38%. The furnace was operated at 45 V and 550 A, and the smelting temperature ranged from 1600 to 1650°C, measured optically using an infra-red pyrometer. The accuracy of the pyrometer readings was validated by limited measurements using an immersion thermocouple.

The 100 kVA pilot-scale SAF is an open furnace with a watercooled roof. The furnace transformer supplies a maximum current of 2080 A at voltages ranging from 35 to 100 V passed through two 75 mm diameter non-tilting graphite electrodes that can be raised or lowered manually. The pilot SAF was operated at the same operating power and temperature conditions as used in the benchscale work, and was lined with the same materials. The product alloy and slag were cast in metal moulds, and representative samples were taken for chemical analysis using a portable XRF instrument. The chemical analyses were validated using XRF. Using FactSage-based design calculations for the production of high-carbon ferromanganese on a pilot scale, Steenkamp (2020) found the recoveries of Mn and Fe at equilibrium conditions to be 76–97% and 100%, respectively. A comparison of Mn ferroalloy smelting in pilot-scale AC and DC submerged arc furnaces showed the Mn recovery to alloy to be 82.1 + 0.5% (AC) and 83.4 + 0.5% (DC) for FeMn smelting, and 69–70% and 70–80% for SiMn (Lagendijk et al., 2010).

By adjusting the slag composition and fluxing materials in the pilot-scale smelting of SiMn by using a Mn-rich slag in the charge, SiMn alloy containing 73% Mn, 15% Si, and 1.1% C has been produced with a metal yield of 65% and 70% Mn recovery (Eissa et al., 2004).

The Mn/Fe ratio for the most popular grades of SiMn alloys – Si16Mn63, Si19Mn63, and Si18Mn68 – ranges between 3.5 and 6. (Bureau of Indian Standards. 2013). Due to the lower Mn 8

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recovery compared to Fe recovery in the SiMn smelting process, the Mn/Fe ratio of the charge should be higher than the target Mn/Fe ratio of the product alloy The Mn/Fe ratio of manganeserich slag produced from low-grade manganese ore is about 9.39, as shown in Table I. This ratio is suitable to compensate for the low Mn/Fe ratio of the medium-grade manganese ore. The blend of manganese-rich slag and medium-grade manganese ore contains a relatively high Mn/Fe ratio and is suitable to be used for production of silicomanganese alloy.

At first, the quantities of coke required for the reduction of different blends, and the quantities of fllux material to attain different ratios of (CaO+MgO)/(Al₂O₃) were calculated.

Manganese ores and slag with the designed Mn/Fe ratios and the required quartzite were identified.

Manganese-rich slag was mixed with medium-grade manganese ore, quartzite, coke, and dolomite. The furnace was preheated, and the graphite electrodes were lowered to a position near the bottom lining of the furnace. A layer of coke bed was charged between the electrode tips. Then the current was switched on to strike an electric arc, which raises the furnace hearth temperature, and the mixed charge was divided into three parts and charged gradually. The furnace was operated at 45 V and a current of around 550 A from the start of charging until casting.

After the complete smelting of the blend, the molten metal and slag were left inside the furnace for enough time to ensure the complete reduction and settling of the silicomanganese alloy. The product silicomanganese alloy and discard slag were then cast into metal moulds of 100 mm inner diameter and 150mm height where the metal, covered by the slag, was left to cool to room temperature.

Based on the optimum conditions of coke ratio, basicity, and charge Mn/Si ratio in the blend that resulted in the highest yield and recoveries of Mn and Si, further heats of the silicomanganese alloy were performed in the 100 kVA pilot-scale submerged arc furnace. The product alloy and slag were cast into metal moulds of 300 mm diameter and 500 mm height and left to cool to room temperature. Representative samples of the metal and slag were analysed.

Results The Egyptian low-grade manganese ore with a high iron content was upgraded by smelting the ore in an open submerged arc furnace. This smelting was performed with lower power and a deficiency of carbon in order to reduce most of the higher iron oxides to metallic iron while reducing the higher manganese oxides to MnO, which partitions predominantly to the slag phase.

The smelting operation has to be strictly controlled, especially as regards the fixed carbon requirements, otherwise the pig iron will contain an unacceptable level of manganese. It is important to note that practically all the phosphorus will end up in the pig iron, thus leaving a P-free slag for silicomanganese production.

Several experimental heats were carried out to establish the optimum conditions for selective reduction of the iron to produce pig iron with the lowest manganese content and slag with the highest manganese content, lowest iron content, and highest Mn/ Fe ratio. At the optimum conditions of power, coke addition, and charge basicity, 93% iron recovery was obtained in the pig iron product with 96% metal yield, and 90% MnO recovery to the slag.

Adjusting the optimum smelting conditions produced pig iron with a low Mn content of 2–2.5% and a high-Mn slag containing 54% MnO and 3.2% FeO with a Mn/Fe ratio of 15.

The Journal of the Southern African Institute of Mining and Metallurgy

Silicomanganese alloy from rich manganese slag The input and output materials for the sixteen experimental heats that were carried out on the bench scale SAF are listed in Table II. These experiments were done to study the effect of different parameters on the SiMn production process.

Effect of coke ratio In the first series of experiments, seven heats (1-7) were carried out to investigate the influence of the coke ratio, which is the ratio of added coke to the stoichiometric requirement to reduce silicon, iron, and manganese oxides. In this series of tests, the blends comprised 2.5 kg of manganese-rich slag, 1.25 kg of medium-grade manganese ore, and 1.25 kg of quartzite, equivalent to 50%, 25%, and 25% by weight of the charge respectively. The coke weight ranged from 0.775 kg to 1.27 kg, corresponding to coke ratios from 0.95 to 1.55. All these tests were performed at a Mn/Fe ratio of 5.6. The metal yield and recoveries of manganese and silicon were calculated according to Equations [9]–[11]. The metal yield represents the weight of the product alloy divided by the theoretical alloy weight resulting from the complete reduction of the manganese, silicon, and iron in the charge and assuming a carbon content of 2% in the alloy. The manganese (and silicon) recovery represents the amount of manganese (silicon) in the product alloy divided by the total amount in the charge blend.

coke ratio of 1.35, then decrease slightly. Figure 3 shows that at a coke ratio of 1.0, manganese and silicon recoveries were low, and subsequently the manganese and silicon contents were below that in the standard silicomanganese alloy. With increasing coke ratio up to 1.35, both the manganese and silicon content in the product silicomanganese alloy increased as a result of the greater degree of reductions, leading to an increase in the metal yield.

The subsequent decline in the manganese and silicon recoveries with further increases in the coke ratio can be attributed to a few factors. Akil and Geveci (2008) stated that using carbon in excess of an optimum value may lead to the formation of an unreacted coke layer above the slag that probably hinders the settling of metal droplets from through the slag layer to the metal phase, thereby decreasing the quantity of alloy obtained. On the other hand, the coke ash contributes to slag formation. The ash content in coke is usually 8–13%.

Approximately 75–80% of the coke ash is Al₂O₃+SiO₂, with an A/S ratio of about 0.65. The contribution of coke ash to the final slag is estimated to be around 5%. With increasing coke ratio, the

[9]

[10]

[11] The effects of the coke ratio on the metal yield and recoveries of manganese and silicon are illustrated in Figures 2 and 3 respectively. The chemical compositions of the silicomanganese alloys and slags produces are given in Table III. Figures 2 and 3 indicate that the metal yield and recoveries of manganese and silicon increase with increasing coke ratio until a

Figure 2—Effect of coke ratio on the metal yield of SiMn

Table II

Input and output materials for the experimental laboratory heats Heat no. Input charge (kg)

Operating conditions

Mn-rich slag Mn ore Quartzite Coke Dolomite Bauxite Coke Charge Charge ratio basicity Mn/ Si ratio

1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16

2.5 2.5 2.5 2.5 2.5 2.5 2.5 2.5 2.5 2.5 2.5 2.5 2.5 2.85 2.65 2.35

1.25 1.25 1.25 1.25 1.25 1.25 1.25 1.25 1.25 1.25 1.25 1.25 1.25 1.4 1.35 1.15

1.25 1.25 1.25 1.25 1.25 1.25 1.25 1.25 1.25 1.25 1.25 1.25 1.25 0.75 1 1.5

0.775 0.86 0.94 1.02 1.1 1.19 1.27 1.1 0 0.1 1.1 0 0.06 1.1 0 0.03 1.1 0.135 0 1.1 0.215 0 1.1 0.295 0 1.07 0.215 1.09 0.215 1.12 0.215

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0.95 1.05 1.15 1.25 1.35 1.45 1.55 1.35 1.35 1.35 1.35 1.35 1.35 1.35 1.35 1.35

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1.95 1.95 1.95 1.95 1.95 1.95 1.95 1.3 1.5 1.7 2.3 2.5 2.7 2.5 2.5 2.5

1.99 1.99 1.99 1.99 1.99 1.99 1.99 1.99 1.99 1.99 1.99 1.99 1.99 3.07 2.45 1.63

Output Metal (kg)

Slag (kg)

0.875 1.25 1.1 1.45 1.72 1.675 1.64 1.3 1.45 1.5 1.72 1.8 1.62 1.7 1.77 1.77

1.6 1.3 0.4 1.15 0.3 1.1 1 0.45 2.6 1.5 1.2 0.95 1.22 0.9 1.9 0.85

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Silicomanganese alloy from rich manganese slag recoveries is given in Figure 6. The graphs indicate that the metal yield and recoveries of both manganese and silicon increased as the charge basicity increased up to an optimum value of 2.5, beyond which the metal yield and recoveries decreased.

During silicomanganese production the silica content in the slag changes continuously due to the reduction of SiO2. It is easier to control the basicity of the input materials than that of the produced slag. During smelting, both MnO and SiO2 are reduced from the slag. So, the alternative measurement of slag basicity, (R) given by Equation [7], containing only nonreducible oxides, is more relevant than that given by Equation [6].

The requirements for greater reduction of manganese oxides are different to those for reduction of silicon oxides, i.e. a basic slag is required for manganese reduction and an acidic slag for silicon reduction. Hence, the composition of the charge and the operating conditions need to be balanced to obtain good recoveries and the desired grade of SiMn. An increase in the basicity of the slag, i.e., increasing the proportion of basic oxides, leads to an increase in MnO activity and a decrease in SiO₂ activity in the slag. Therefore, more MnO and less SiO₂ is reduced by increasing the basicity of slag – in other words, increasing basicity decreases the equilibrium MnO and increases the equilibrium SiO₂ in the slag, as indicated in Figure 7 (Nokhrina, and Rozhikhina, 1998).

The influence of charge basicity on metal yield and manganese and silicon recoveries beyond the optimum can be discussed in terms of the effect of basicity on the viscosity of the slag. The

Figure 4—Photograph of metal droplets entrapped in discard SiMn slag Figure 3—Effect of coke ratio on Mn and Si recoveries in SiMn

increase in coke ash leads to greater amounts of SiO₂ and Al₂O₃ in the slag and hence a higher viscosity which hinders the movement of metal droplets through the slag layer as indicated in Figure 4. Thus a lower metal yield and lower manganese and silicon recoveries were attained (Olsen, Tangstad, and Lindstad, 2007).

Effect of charge basicity The second series of tests was designed to illustrate the influence of charge basicity on the smelting process. Seven heats (5, 8–13) were designed to investigate the influence of the charge basicity as (MgO + CaO) / Al₂O₃ on the metal yield and Mn and Si recoveries through the addition of either bauxite or dolomite. In this series, the blends were composed of 2.5 kg manganese-rich slag, 1.25 kg medium-grade manganese ore, and 1.25 kg quartzite, representing 50%, 25%, and 25% of the total charge weight respectively, with a coke ratio of 1.35. The charge basicity varied from 1.3 to 2.7, as given in Table II, and the Mn/Fe of all blends was 5.6.

The chemical composition of product alloy and slag is shown in Table III. The influence of charge basicity on metal yield is illustrated in Figure 5, while the effect on manganese and silicon 10

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Figure 5—Effect of basicity on metal yield of SiMn The Journal of the Southern African Institute of Mining and Metallurgy

Silicomanganese alloy from rich manganese slag Table III

Chemical composition of the produced SiMn and discard slag from the bench-scale SAF Heat no. SiMn (wt%) SiMn slag (wt%)

1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16

C

Mn

Si

S

P

MnO

SiO₂

Al₂O₃

CaO

MgO

BaO

S

2.399 2.07 2.08 1.72 0.91 1.08 1.122 1.22 2.02 1.79 0.8 0.85 1.2 1.7 1.2 0.8

57.5 56.1 62.4 54.17 56.5 55.23 56.8 52.16 61.8 59.1 60.5 61.4 59.1 69 65.75 56.84

9.26 13.6 12.65 23.73 23.46 24 24.9 19 17.83 20.89 21.7 20 21.1 14 17.1 24.22

0.018 0.016 0.013 0.02 0.0066 0.006 0.0058 0.006 0.004 0.005 0.0059 0.005 0.006 0.004 0.013 0.006

0.057 0.069 0.06 0.04 0.098 0.091 0.1 0.112 0.072 0.07 0.055 0.083 0.097 0.037 0.084 0.116

32.31 27.14 22.91 15.9 14 13.37 13.46 15.81 13.71 14.22 16.44 12.98 12.1 13.55 12.89 12.71

45.64 45 44.09 42.56 41.07 39.66 39.14 45.43 42.07 43.29 40.72 40.11 41.55 38.28 39.79 47.7

5.9 7.18 9.42 11.8 13.4 14.3 14.11 13.5 14.2 12.12 10.4 10.76 10.65 13.07 10.62 10.26

6.36 8.37 9.65 10.95 12.94 12.8 13.23 10.55 11.9 10.79 11.88 15.36 16.2 13.98 14.54 12.14

5.87 6.94 8.48 12.4 14.64 14.56 15.1 9.65 13.1 13.73 14.5 16.73 16.53 15.85 16.1 11.82

1.7 2.75 2.8 3.91 2.16 2.9 2.46 2.7 3.22 3.61 3.2 2.46 1.01 3.12 3.8 3.23

0.412 0.392 0.4 0.332 0.12 0.336 0.42 0.52 0.16 0.212 0.468 0.204 0.088 0.172 0.116 0.284

viscosities of the silicomanganese slag in this series of tests were calculated at 1600°C for different slag basicities using Equation [8] (Riss and Khodo, 1967; Rosypalová et al., 2014) and plotted

Figure 6—Effect of basicity on Mn and Si recoveries in SiMn The Journal of the Southern African Institute of Mining and Metallurgy

versus the final slag basicity. Figure 8 shows that the lowest viscosity, 6.3 poise, was attained for a slag basicity ratio CaO/ SiO₂ of 0.37 and for (CaO + MgO)/Al₂O₃ at 2.5. The retardation of the reduction process at higher slag basicities beyond optimum conditions is the effect of the high viscosity of the slag, which seems to overcome the positive effect of increasing the MnO activity, with the net result of supressing the reduction process at

Figure 7—Effect of charge basicity on Mn and Si distribution ratio VOLUME 122

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Silicomanganese alloy from rich manganese slag

Figure 9—Effect of Mn/Si ratio on metal yield of SiMn

Figure 8—Effect of final slag basicity on the slag viscosity

high basicity. As indicated in Figure 4, this is due to the hindering of the sinking of metal droplets through the slag layer, which leads to entrapment of the metal droplets and thereby causes the lower metal yield and recoveries of manganese and silicon.

Effect of charge Mn/Si ratio Four experimental heats (12, 14–16), were designed to investigate the effect of the Mn/Si ratio of the charge on manganese and silicon recoveries and metal yield. In this series, the blends were composed of manganese-rich slag (50%) and medium-grade manganese ore (25%) with a coke ratio equal to 1.35. These heats had nearly the same Mn/Fe ratio of about 5.6 and the quartzite addition was varied from 15 to 30% of the charge blend. The input raw materials of the four experimental heats were given in Table II and the chemical compositions of the product alloy and slag are shown in Table III. The effects of the Mn/Si ratio of the charge on the metallic yield and recoveries of manganese and silicon are shown in Figures 9 and 10. The metal yield and Mn recovery increased with increasing charge Mn/Si ratio up to an optimum ratio equal to 2. A further increase was accompanied by a general decrease in the metal yield and manganese recovery.

At a low Mn/Si ratio, manganese and silicon were not reduced mainly from their free oxides, but from their oxides combined as silicates as given in Equation [4]. More free SiO₂ increases the silica activity, resulting in greater silica reduction with increasing silicon recovery. On the other hand, an enhanced silica activity decreases the MnO activity by forming stable ß-manganese silicates, as indicated in Equation [10], which coat MnO particles thus minimizing the contact area between the reducing agent and MnO, which in turn decreases the manganese recovery (Coetsee, 12

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Zietsman, and Pistorius, 2014; Yastreboff, Ostrovski, and Ganguly, 1998; Cengizler, 1993). With increasing Mn/Si ratio, the activity of MnO increases owing to the decreasing silica content, with the result that the metal yield and recovery of manganese increase up to a Mn/Si ratio of 2. The decrease in recoveries and metal yield beyond this point can be explained by the effect of the Mn/Si ratio on the distribution of manganese and silicon between the alloy and slag phase, as shown in Figure 11. However, with the addition of silica to the slag, an MnO.SiO₂ type complex is formed, which limits the activity of free MnO and hence affects the reduction of MnO to Mn. Thus, with an increase in silica the silicon content increases and manganese recovery decreases. The addition of silica also increases the slag liquidus, which in turn leads to a higher bath temperature favourable for Si recovery (Olsen, Tangstad, and Lindstad, 2007; Kor, 1979; Cengizler, and Eric, 2016; Swinbourne, Rankin, and Eric, 1995).

Therefore, a Mn/Si ratio of 2 in the blend is considered the most suitable to meet the Si and Mn requirement in the silicomanganese standard (Swinbourne, Rankin and Eric, 1995).

The higher silica content of the slag raises the slag viscosity and thus retards the slag/metal reactions (Kor, 1979; Ding and Olsen, 2000).

[12]

Production of silicomanganese alloys on a 100 KVA pilot-scale SAF The bench-scale tests indicated that the optimum conditions for producing a standard SiMn alloy were a coke ratio of 1.35, charge basicity 2.5, Mn/Si ratio 2, and Mn/Fe ratio 5.6. One experimental heat was carried out in triplicate in a 100 kVA pilot submerged arc furnace using these conditions. The charge composition and chemical analysis of the SiMn products are given in Table IV. The results indicate that that high-manganese slag produced by selective reduction of low-grade Egyptian manganese ore was successfully smelted on a pilot scale to produce a silicomanganese alloy containing about 65% Mn and 16–17% Si, with high manganese recovery of 68–75% and silicon recovery of 36–39%.

The Journal of the Southern African Institute of Mining and Metallurgy

Silicomanganese alloy from rich manganese slag

Figure 11—Effect of Mn/Si ratio on Mn and Si distribution ratios

manganese ore and local medium-Mn ore, in a blend consisting of 50% slag, 25% medium-manganese ore, and 25% quartzite with Mn/Fe equal 5.6, to produce SiMn alloy.

Increasing the coke ratio (the ratio of added coke to the stoichiometric amount) to 1.35 leads to an increase in the metal yield and manganese and silicon recovery to the product alloy to maximum values of 66.16% , 64.13%, and 52.88%, respectively. The optimum charge basicity, expressed by (CaO + MgO / Al₂O₃), that gives the higher metal yield and manganese recovery is 2.5.

The optimum Mn/Si charge ratio of 2.0 results in a metal yield (69.23%) and manganese recovery (72.93%), and a silicon recovery of 47.18%.

Figure 10—Effect of Mn/Si ratio on Mn and Si recoveries in SiMn

Conlusions This work investigated the possibility of using high-manganese slag produced by selective reduction of low-grade Egyptian

Table IV

Experimental heats for production of SiMn in a 100 kVA pilot plant SAF Heat no.

Charge (kg) Metal wt. Slag wt.

SiMn composition (wt%) Metal

1 2 3

Mn-rich slag

Med. ore Quartzite

Coke

Dolomite

(kg)

(kg)

22.5 22.5 22.5

15 12.5 15 12.5 15 12.5

12.5 12.5 12.5

3.27 3.27 3.27

17.5 16 18

0 0 1

The Journal of the Southern African Institute of Mining and Metallurgy

C

Mn

2.1 64.66 0.8 66.37 1.1 64.8

Mn Si recovery recovery

Si

Fe

P

yield

(%)

(%)

16.3 15.7 16.93

Bal. Bal. Bal.

0.084 0.046 0.091

65.76 60.13 67.64

72.82 68.34 75.07

36.38 32.04 38.87

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Silicomanganese alloy from rich manganese slag Based on the results of the bench-scale tests, pilot-scale tests in a 100 kVA submerged arc furnace were carried out successfully, producing a standard silicomanganese (Si16Mn63 and Si17Mn65) with high yield and recoveries.

Acknowledgements This work was carried out a part of an applied project in a technological alliance for extending and maximizing the local beneficiation of mineral resources in Sinai, financed by the Academy of Science and Technology, Egypt. The authors would like to acknowledge the Academy of Science and Technology for financial support and the use of their facilities. Special thanks and gratitude are due to the management and colleagues in Sinai Manganese Company for the supply of raw materials, data, information, and fruitful discussions.

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Tang, K. and Olsen, S.E. 2004. Thermodynamics of the MnO-containing slags and equilibrium relations associated with Mn ferroalloy productions. Proceedings of the International Conference on Molten Slags Fluxes and Salts, South African Institute of Mining and Metallurgy, Johannesburg. pp.19-23. https://www. pyrometallurgy.co.za/MoltenSlags2004/019-Tang.pdf Tang, K. and Olsen, S.E. 2006. Computer simulation of equilibrium relations in manganese ferroalloy production. Metallurgical and Materials Transactions B, vol. 37 no. 4. pp. 599-606.

Tang, K. and Tangstad, M. 2007. Modeling viscosities of ferromanganese slags.

Proceedings of INFACON XI, New Delhi, India, 18-21 February 2007. Indian Ferro-Alloy Producers Association. pp. 344–357. https://www.pyrometallurgy. co.za/InfaconXI/344-Tang.pdf Tranell, G., Gaal, S., Lu, D., Tangstad, M., and Safarian, J. 2007. Reduction kinetics of manganese oxide from HC FeMn slags. Proceedings of INFACON XI, New Delhi, India, 18-21 February 2007. Indian Ferro-Alloy Producers Association. pp. 231-240. https://www.pyrometallurgy.co.za/InfaconXI/231Tranell.pdf Yastreboff, M., Ostrovski, O., and Ganguly, S. 1998. Carbothermic reduction of manganese from manganese ores and ferromanganese slag. Proceedings of the 8th International Ferroalloys Congress, Beijing, China, 7-10 June 1998. China Science & Technology Press. pp. 263–270. https://www.pyrometallurgy.co.za/ InfaconVIII/263-Yastreboff.pdf u The Journal of the Southern African Institute of Mining and Metallurgy

Safety aspects of large draglineoperated opencast mines – An overview

by A. Golder1 and I. Roy1

Affiliation:

epartment of Civil & D Environmental Engineering, Birla Institute of Technology, Mesra, Jharkhand, India.

1

Correspondence to: A. Golder

Email:

amitgolder91@gmail.com

Dates:

Received: 8 Dec. 2020 Revised: 1 Dec. 2021 Accepted: 8 Dec. 2021 Published: January 2022

How to cite:

Golder, A. and Roy, I. 2022.

Safety aspects of large draglineoperated opencast mines – An overview.

Journal of the Southern African Institute of Mining and Metallurgy, vol. 122, no. 1, pp. 15-20 DOI ID: http://dx.doi.org/10.17159/24119717/1452/2022 ORCID: A. Golder https://orcid.org/0000-00025078-1613 I. Roy https://orcid.org/0000-00032895-5135

Synopsis The Jayant opencast operation is one of the largest opencast coal mines in India. Prior to 2008 the mine experienced a number of dragline dump failures, which was a major hindrance in sustaining production.

Northern Coalfields Limited (NCL) and the mine management engaged several design, research, and academic institutions to carry out dump slope stability studies, particularly of dragline dumps. Birla Institute of Technology prepared a report on the investigations in May 2009. In this paper we review the findings of the report and the measures taken to tackle the safety aspects of dragline dumps.

Keywords slope stability, circular failure, opencast mining, shear strength, overburden dump.

Study area The Jayant Project of Northern Coalfields Limited (NCL) is situated in the Madhya Pradesh district of Sidhi in the Singrauli Coalfields. The location of the project is latitude 24° 05’45” to 24° 11’ 25” N and kongitude 82° 38’ 21” to 82° 40’ 45” E, as per Indian Survey toposheet no. 63 L/12. The Shaktinagar rail station on the Chopan-Katni line of the East Central Railway is approximately 5 km from the project (CMPDI, 2007).

The area of the Jayant Block in the northeastern section of the Singrauli Coalfield is 11.10 km2. The Jayant opencast operation of the project is located on a hilly plateau with the RL varying from 390-430 m.

The total net geological reserve is 305.50 Mt, while the mineable reserve is 282.71 Mt (as at 31 March 2018) and thus the overall volume of overburden with a common stripping ratio of 2.60 m3/t is 907.20 million m3. There are three different seams present in the Jayant Block, i.e., Turra Seam, Purewa Bottom Seam, and Purewa Top Seam as shown in Figure 1 (Singh et al., 2014; Sharma and Roy, 2015).

In this area, most of the overburden is medium-grained to coarse-grained sandstone, carbonaceous shale, and sandy shale. The the dragline dump is situated on shale and sandy shale that provides a competent foundation. The floor of the dump is covered with a thin layer comprising a wet mixture of coal dust, carbonaceous shale, and sandstone (Singh et al., 2012), and fragments of waste rock, which is referred to as interface material. Two types of circular failure surfaces are envisaged as shown in Figure 2. ➤ Failure within the material of dump ➤ Failure within both the dump material and interface material.

Hydrogeological factors The hydrogeological parameters that control the stability of the dump are determined as follows. ➤ An attempt was made earlier to delineate or establish the water table/phreatic surface within the dump by installing piezometers. However, the piezometers could not be installed due to difficulty experienced in drilling through the loose, fragmented, and heterogeneous dump material. ➤ In the absence of sufficient hydrogeological data, the position and curvature of the phreatic surface inside the dump, as well as the seepage height above the dump toe as shown in Figure 3, was observed visually and reported by mine officials during rainy seasons. ➤ It is not feasible to evaluate the phreatic surface in the dragline dump using piezometers because the soil is heterogeneous. The water table height (Hw) is estimated by Casagrande's equation (Murthy, 2002; Sengupta and Roy, 2015; (Moosavi, Shirinabadi, and Gholinejad, 2016).

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Safety aspects of large dragline-operated opencast mines – An overview

Figure 1—Geological cross-section of the Jayant opencast mine (Sharma and Roy, 2015)

Figure 2—Schematic diagram of dragline working (Sharma and Roy, 2015)

[1] where Pa = Height of seepage at the toe of the dump.

L or β = Overall angle of slope of dump.

Dp = Behind the toe of dump i.e., 60 m.

By applying Casagrande’s equation, at a distance of 300 m from the toe of the dump the height of the phreatic surface within the dump is calculated as tabulated in Table I. With this height of water table, the seepage and hydrostatic forces are calculated and considered in the stability calculation.

The phreatic surface (DPY, DPX) (Figure 2) is also evaluated through relevant condition: [2] The height of the water table is calculated as 25 and 36 m (Table I).

Table I

Height of water table corresponding to different seepage heights Seepage height (Pa) (m)

Height of water table (Hw) (m)

3 6

25 36

The height of seepage face is shown in Figure 2 16

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The upward thrust of the water can be defined as the product of the unit weight of the water and the volume of the dragline dump submerged under the water table within the failure mass (Roy, 2016) The seepage force is calculated as the product of the upward thrust and the sine of the gradient of the horizontal phreatic surface (Murthy, 2002; Sengupta and Roy, 2015) (Figure 4).

Seismicity and blast vibrations Seismic forces are regarded according to the Indian Standard criteria for earthquake-resistant structural design (5th edn) IS 1893:2002 (IS-1893 (part 1), 2002). The horizontal seismic coefficient (Ah) design for the Jayant dragline dump is determined by the following expression (Sengupta and Roy, 2015): [3] where Z = Zone factor (study area is located in zone III according to IS 1893:2000) I = Importance factor R = Response reduction factor Sa/g = Average response acceleration coefficient of dump mass.

According to the Indian seismic map, the project is located in zone III, with the horizontal seismic coefficient of 0.02 m/s2, as per the IS code considered here. The blast vibration coefficient on the dump mass due to ongoing blasting was estimated such that the horizontal coefficient of 0.04 m/s2 will include both seismicity and blasting (Mosinets and Shemyakin, 1974).

The Journal of the Southern African Institute of Mining and Metallurgy

Safety aspects of large dragline-operated opencast mines – An overview

Figure 3—Circular planar failure surface at Jayant mine

Dragline dump height The dragline dump height, which varies between 60 to 100 m, and the surcharge load of the shovel dump on the dragline dump are also considered in the stability analysis (Mosinets and Shemyakin, 1974; Zaitseva and Zaitsev, 2009; Sengupta, Sharma, and Roy, 2014).

Coal rib According to existing practice in this mine, a coal rib of 7 m base width and 3 m top width with average Turra seam thickness of 19 m, as shown in Figure 3, is considered as a resisting force against dump failure (Roy, 2003). The coal rib left at the toe of the dump acts as a retaining wall and reduces dragline dump rehandling to some extent (Colwell and Mark, 2003; Besimbaeva et al., 2018).

Laboratory tests for the generation of geotechnical information Figure 4—Free body diagram of individual slices. W - dead load of slice, i - sine gradient, S - seismicity factor

Samples of the dump material as well as the interface material were collected and transported to BIT Mesra for determination of the strength parameters (Ranjan et al., 2017) (Table II).

Recommendations Dump floor inclination The mine floor inclination varies from 2° to 4° (CMPDI, 2018). For stability calculations a dump floor inclination of 3° is considered here (Sengupta and Roy, 2015).

Considering the above parameters and by applying both Fellenius and Bishop's simplified method (Moosavi, Shirinabadi, and Gholinejad, 2016), the slope angles for the dragline dump are calculated (Table III) and recommended for a minimum factor of safety of 1.20 for different heights of the seepage face (Besimbaeva et al., 2018).

Table II

Shear strength parameters determined by laboratory testing (large shear box apparatus) Parameter

Cohesion (kN/m2) Angle of internal friction (°) Bulk density (kN/m3)

Dump material at natural moisture condition

Interface material in submerged condition at the base of the dragline dump

Interface layer separating the coal rib/barrier

75 25 20.6

40 21 Not required in calculation

155 35 16

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Safety aspects of large dragline-operated opencast mines – An overview Table III

Prediction of dragline dump geometry Height of dragline dump (H) (m)

Overall slope angle of the dragline dump (°)as shown in Figure 3 Seepage height of water (Pa) = 3 m, and height of water table (Hw) = 25 m (Figure 3)

Seepage height of water (Pa) = 6 m, and height) of water table (Hw) = 36 m (Figure 3

33 32 31 30 28

31 30 29 28 26

60 70 80 90 100

The results are documented for different dragline dump geometries in Table III. The angle of repose of the dragline dump is the overall angle with respect to the horizontal plane over which it is standing. The dragline dump is considered to be cohesionless for the purpose of determining the angle of repose, although in reality the dump has an emerging water table, an inclined floor, pore pressure within the dump material, and is affected by blast vibrations and pore pressure within the dump material. Also, cohesion is generated due to the compaction of the dump material under its own weight.

Hence the angle of repose of the dragline dump is 37° in the ideal case, but in actual site conditions owing to the above prevailing geo-engineering considerations the overall slope of the dump will differ.

The above recommended slopes of the dragline dump are maintained by optimizing the following parameters of the dump profile as shown in Figure 3. ➤ At the mining level of the dragline – berm width ➤ Berm width present at coal rib/barrier ➤ Angle of slope below mining level of the dragline.

It is recommended that truck and shovel dumps overlying the dragline dump are formed 120 to 150 m away from the toe of the dragline dump (Sharma and Roy, 2015), (i.e. the interval between the toe of the dump formed by the shovel dump and dragline dump should be at least two cut widths 120–150 m). In this case, the dragline dump will act as a foundation for the shovel dump.

Hence, the geotechnical properties are considered to be the same for both the shovel dump and its foundation. Accordingly, the following combinations of shovel dump are calculated and recommended (Government of India, 2017; Directorate General of Mine Safety, 2008) in Table IV and shown in Figure 3.

The recommended overall slope angle of the shovel dump can be maintained by adjusting the berm width at the coal rib roof level.

Precautionary measures In addition to managing the slope, several proposals are suggested and implemented to ensure the stability of internal dumps as well as dragline dumps (Sharma and Roy, 2015): Table IV

Prediction of shovel dump geometry Shovel dump height (m)

60 70 80 18

Angle of overall slope (°)

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i. Topsoil is dumped separately away from the existing internal overburden dump. ii. To form a foundation for the dragline dump, no surfacesoil may be dumped at the level from where coal has been extracted. iii. By ensuring normal gravitational seepage of water in the direction of the sump area, nominal collection of water takes place where coal has been extracted. iv. The dragline dump receives sufficient time to settle, followed by supplementary truck dumping, therefore the distance between the toe of shovel dump and the dragline dump is between 120 and 150 m, i.e. two cuts beyond the toe of the dragline dump (Sharma and Roy, 2015). v. The voids in the dragline dumps are consolidated with the help of dozers. vi. Some coal is left at the toe of the dump to act as a barrier (coal rib). It is designed in such a way that the overburden dump should cover up the coal rib/barrier as much as possible, and that the coal rib/barrier is likely to burn naturally. vii. Efforts are made to extract coal from the coal rib without blasting, at regular intervals of 200 to 250 m along the strike length of the pit, so that there is no accumulation of water against the coal rib. viii. Before dumping by dragline, the interface layer is cleaned from where coal has been extracted to as great an extent as possible (Singh et al., 2012). If possible, fragmented overburden rock is dumped to cover the slushy floor at the base of the dragline dump to increase the coefficleint of friction at the dump floor. ix. As per the recommendations of BIT Mesra, Ranchi (Roy, 2016): a) If possible, the mine floor (foundation of internal dump) may be ripped or blasted at intervals to a depth of 1 to 2 m, thereby increasing the coefficient of friction prior to dumping by dragline (Government of India, 2019). It is also recommended that minor blasts to promote the flow of water to the sandstone beds below the open pit floor should be carried out to limiting water retention at the base of the dump. b) Regular monitoring of the dumps through a non-contact survey using a laser profiler or laser scanner is proposed to detect any movement of overburden dumps or dump faces that will indicate a potential dump failure. As the dragline dumps are inaccessible, a reflectorless instrument based on laser technology is recommended for surveying the The Journal of the Southern African Institute of Mining and Metallurgy

Safety aspects of large dragline-operated opencast mines – An overview displacement of the dump face between the crest and the toe of the dump. c) It is proposed that monitoring of the dump should be carried out and recorded at seven-day intervals during the monsoon and post-monsoon seasons (July to November), whereas in the dry season (December to June) monitoring should be done at 15-day intervals. In the case of any movement of the overburden dump, the de-coaled floor near the toe of the dump is declared as a hazard zone with removal of men and machinery from the hazard zone. d) The seepage of water from the face of the dragline dump is to be monitored when the coal rib has been breached at 7-day intervals from July to November and at 15-day intervals for rest of the period.

Conclusion Based on the recommendation of the Birla Institute of Technology Mesra, the Jayant opencast project has maintained the dump profiles by adjusting the following parameters as shown in Figure 3a) Berm width at the coal rib roof level b) Angle between the coal rib roof and the dragline mining level c) Berm width at the dragline mining level d) Slope angle above the dragline mining level.

The abovementioned measures have successfully prevented any major failure of backfilled dumps in spite of the huge volumes of waste rock (around 40 million m3 per year loose volume) handled and dumped within the de-coaled area.

Acknowledgement The authors are grateful to the administration of Birla Institute of Technology Mesra, Ranchi (India) for their permission to present this paper. This research work was financially supported by Coal India Limited (CIL), Kolkata - 700156, India (Project Code No.: CIL/R&D/01/68/2018). The opinions expressed in this paper are those of the authors and not to the institution to which they belong.

Funding Agency: Coal India Limited, Kolkata, India (P. Code No.: CIL/R&D/01/68/2018).

The authors declare that they have no conflict of interest.

References Besimbaeva, O.G., Khmyrova, E.N., Nizametdinov, F.K., and Oleinikova, E.A. 2018. Assessment and prediction of slope stability in the Kentobe open pit mine. Journal of Mining Science, vol. 54. pp. 988–994. https://doi.org/10.1134/ S1062739118065143 Bureau of Indian Standards. 2002. IS-1893 (part 1). Criteria for earthquake resistant design of structures. New Delhi. pp. 14–23.

CMPDI. 2018. Revised mine closure plan of Jayant opencast project.

CMPDI. 2007. Environmental impact assessment (EIA).

Colwell, M. and Mark, C. 2003. Analysis and design of rib support (ADRS) A rib support design methodology for Australian collieries. Proceedings of the 24th International Conference on Ground Control in Mining. West Virginia University, Morgantown, WV. https://www.cdc.gov/niosh/mining/UserFiles/works/pdfs/ aador.pdf The Journal of the Southern African Institute of Mining and Metallurgy

Dash, A.K. 2019. Analysis of accidents due to slope failure in Indian opencast coal mines. Current Science, vol. 117. pp. 304–308.

Directorate General of Mine Safety (DGMS). 2008. Report of high powered committee on accident in the West Coal Section of Jayant, NCL. Dhanbad, Jharkhand.

Government of India. 2019. Ministry of Coal. https://coal.nic.in/ Government of India. 2017. Coal mines regulations. New Delhi.

Moosavi, E., Shirinabadi, R., and Gholinejad, M. 2016. Prediction of seepage water pressure for slope stability at the Gol-E-Gohar open pit mine.

Journal of Mining Science, vol. 52. pp. 1069–1079. https://doi.org/10.1134/ S1062739116061601 Mosinets, V.N. and Shemyakin, E.I. 1974. Research on the seismic effects of blasting in the mining industry. Journal of Mining Science, vol. 10. pp. 442–447. https://doi.org/https://doi.org/10.1007/BF02501433 Murthy, V.N.S. 2002. Geotechnical Engineering: Principles and Practices of Soil Mechanics and Foundation Engineering. Jurnal Ekonomi Malaysia. https://doi. org/10.1017/CBO9781107415324.004 Singh, P.K., Singh, M.P., Volkmann, N., Naik, A.S., and Börner, K. 2014.

Petrological characteristics of lower Gondwana coal from Singrauli coalfield, Madhya Pradesh, India. International Journal of Oil, Gas and Coal Technology, vol. 8. pp. 194–220. https://doi.org/10.1504/IJOGCT.2014.064849 Ranjan, V., Sen, P., Kumar, D., and Sarsawat, A. 2017. Enhancement of mechanical stability of waste dump slope through establishing vegetation in a surface iron ore mine. Journal of Mining Science, vol. 53. pp. 377–388. https://doi.org/10.1134/ S106273911702226X Roy, I. 2016. A report on CIL R&D Project: Development of guidelines to predict distance between the toe of the shovel dumps and that of the dragline dumps with consideration of safety and economical design of both shovel dumps and dragline dumps of Coal India. Ranchi, Jharkhand.

Roy, I. 2003. The role of interface material at the base of internal dumps and effectiveness of coal rib in the safe working of opencast coal mines. Journal of Rock Mechanics and Tunnlling Technology, vol.. 9. pp. 155–172.

Sengupta, S. and Roy, I. 2015. Study of internal dump stability of Dudhichua open cast project, Northern Coalfields Limited, India. Journal of The Institution of Engineers (India): Series D, vol. 96. pp. 67–75. https://doi.org/10.1007/s40033014-0061-5 Sengupta, S., Sharma, S., and Roy, I. 2014. Investigation of shear strength parameters of highwall rock slopes and overburden dump mass in opencast coal mines. International Journal of Engineering, Management, Humanities and Social Sciences Paradigms (IJEMHS), vol. 7. pp. 1–6.

Sharma, S. and Roy, I. 2015. Slope failure of waste rock dump at Jayant opencast mine, India: A case study. International Journal of Applied Engineering Research, vol. 10. pp. 33006–33012.

Singh, P.K., Roy, M.P., Paswan, R.K., Singh, V.K., Sinha, A., Shastri, L.B., Singh, C.P., and Roy, M.P. 2012. Effect of production blasts on waste dump stability.

Rock Fragmentation by Blasting. Singh, P.K. and Sinha, A. CRC Press, London.

Zaitseva, A.A. and Zaitsev, G.D. 2009. Influence of geological and technological factors on the internal dump capacity in flat deposits. Journal of Mining Science, vol. 45. pp. 380–389. https://doi.org/https://doi.org/10.1007/s10913-0090048-z u VOLUME 122

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Online Short Course

MINE TO MILL Safety aspects of large dragline-operated opencast mines – An overview RECONCILIATION: FUNDAMENTALS AND TOOLS FOR PRODUCTIVITY IMPROVEMENT SHORT COURSE DATE: 21-24 JUNE 2022 ONLINE FROM 9AM TO 13PM SAST

Mining is an integrated operation that includes planning, mining (excavation), processing, and metallurgy where metals are extracted. A miningplan of operation forecasts the amount of ore to be mined and/or metal to be produced at the beginning of a year. The model-mine-mill (M3) reconciliation is a set of activities carried out by mining operations in order to measure success and predict success of mine plan. Success in annual production of a mine depends on many factors such as a good mine plan based on a nearly robust resource/ reserve model, execution of mine-plan in the operation supported by reliable laboratory services and disciplined work-culture led by a competent workforce.

A model-mine-mill (M3) reconciliation process helps to gain knowledge on the actual (or, likely) reasons for achieving the planned production or, not. When properly implemented, a detailed model-mine-mill (M3) reconciliation process leads to finding out the areas for further improvements leading to substantial enhancements in productivity. A successful mining operation often conducts the reconciliation at regular time intervals.

T H E TA R G E T AU D I E N C E • • • • • •

Mining engineers Geologists Process engineers associated with mining operations.

Financial institution professionals working on mining projects Consultants Researchers.

Dr Abani Samal –Program lead

Dr Abani R Samal holds PhD from SIU Carbondale, DIC & MS from Imperial College, London and M Tech (Mineral Exploration) degree from IIT (ISM) Dhanbad, India. He is in the mining industry since 1996. Currently he is the principal and owner of GeoGlobal, LLC and providing consulting services to exploration and mining companies, government agencies and major financial institutions worldwide. Mineral deposit evaluation, applied geostatistics and mine-mill reconciliation are his technical specializations.

Dr Samal offers professional development programs to industry professionals and researchers worldwide. Major professional societies including SME, SEG, AIG (online), CIM-IUGS and MEAI have hosted his professional development programs in USA, Canada, Guinea, India and Colombia.

Dr Samal also serves in the editorial boards of two international journals: Mining, Metallurgy & Exploration journal (MMEX) and Natural Resources Research Journal.

THE PROGRAMME The program will be offered virtually in 4 sessions each session with three contact hours of virtual meetings and two take home assignments / exercises. The sessions will have multiple Q&A opportunities.

Outline of the program: Session 1: Introduction of the program and participants Reconciliation: An introduction to the concept Session 2: Various types and nodes of reconciliation in a mining project: LT Resource/reserve, ST Resource/Reserve, Production, Stock piles and Mill Session 3: Factors in reconciliation Session 4: Error measurements & Q&A

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FOR FURTHER INFORMATION, CONTACT: Camielah Jardine E-mail: camielah@saimm.co.za JANUARY 2022 VOLUME 122 The Journal of the Southern African Institute of Mining and Metallurgy Head of Conferencing Tel: +27 11 834-1273/7 SAIMM Web: www.saimm.co.za

The concentration of rare earth elements from coal fly ash

by G.B. Abaka-Wood1, J. Addai-Mensah1,2, and W. Skinner2

Affiliation:

Future Industries Institute, University of South Australia, Mawson Lakes Campus, Adelaide.

2 Department of Mining and Process Engineering, Namibia University of Science and Technology, Windhoek, Namibia.

1

Correspondence to: G.B. Abaka-Wood

Email:

george.abaka-wood@unisa.edu.au

Dates:

Received: 15 Jun. 2021 Revised: 1 Sep. 2021 Accepted: 3 Sep. 2021 Published: January 2022

Synopsis Recently, coal fly ash has become a potential candidate as a secondary resource of rare earth elements (REE). In this investigation, we studied the recovery of REE from fly ash from a commercial power plant. The specific aim was to assess the technical feasibility of recovering REE from the coal fly ash using conventional preconcentration methods, including gravity separation, magnetic separation, and froth flotation. The experimental results revealed that flotation achieved major gains in REE recovery and upgrading. However, during gravity and wet magnetic separation tests, the bulk of REE reported to the tailings. The results showed significant variations in the performance of the various beneficiation methods investigated. This study has confirmed that existing physical separation methods could be used to recover REE from coal fly ash prior to hydrometallurgical and pyrometallurgical processing, although some challenges persist.

Keywords rare earth elements, coal fly ash, physical separation, flotation.

How to cite:

Abaka-Wood, G.B., Addai-Mensah, J., and Skinner, W. 2022 The concentration of rare earth elements from coal fly ash.

Journal of the Southern African Institute of Mining and Metallurgy, vol. 122, no. 1, pp. 21-28 DOI ID: http://dx.doi.org/10.17159/24119717/1654/2022 ORCID: G.B. Abaka-Wood https://orcid.org/ 0000-00022815-0022 J. Addai-Mensah https://orcid.org/ 0000-00018381-2108 W. Skinner https://orcid.org/0000-00029606-023X

Introduction Over the past decade, rare earth elements (REE) have gained significant attention in the global market due to their industrial and technological applications. Although the demand for REE has increased tremendously, commercial or large-scale production is limited to a few countries. REE have been classified as critical and strategic materials due to their high supply risk and increased global demand, especially in the defence, energy, electronics, and automotive industries (Abaka-Wood et al., 2019a; Blissett, Smalley, and Rowson, 2014; Liu, Huang, and Tang, 2019). With the current depletion of highgrade resources globally, there has been a corresponding increase in attempts to exploit unconventional or secondary resources such coal fines, fly ash, permanent magnets, mining tailings, and phosphorusbased products as sources of REE (Abaka-Wood et al., 2019b, Liu, Huang, and Tang, 2019; Seredin et al., 2013).

Over the past two decades, coal deposits have been considered as alternative resource for REE (Hower et al., 2020; Pan et al., 2020; Sahoo et al., 2016; Seredin and Dai, 2012; Sis, Ozbayoglu, and Sarikaya, 2004; Zhang et al., 2015). On average, the REE content in coal is 68 ppm, and 404 ppm in coal ash (Ketris and Yudovich, 2009; Seredin et al., 2013; Zhang et al., 2015). Recent studies suggest that coal products contain sub-economic REE concentrations, which may be recovered and upgraded (Seredin et al., 2013).

Table I shows the total REE production in different countries along with estimated reserves (US Geological Survey, 2021). China controls the majority of REE resources and production, accounting for about 58–60% of documented global production in 2019–2020. The USA and Australia produce about 15– 18% and 7–15% of the global rare earth oxides (REO), respectively. The data also shows that China holds 41 Mt, representing approximately 37% of the economic demonstrated REE resources, followed by Brazil and Vietnam, with 21 Mt (18%) each, Australia (4.1 Mt, approximately 3%), and the USA (1.5 Mt, 1.2%).

Zhang et al. (2015) suggest that the total amount of REE in coal could be about 50 Mt (Table II, which equates to about 42% of the total global traditional REE reserves. This implies that coal may be considered as an unconventional REE resource, subject to market circumstances and the development of efficacious beneficiation methods to recover the REE. However, Zhang et al. (2015) suggested that the complexity of the composition and distribution of REE in coal and its by-products have limited studies aimed at recovering REE from coal products.

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The concentration of rare earth elements from coal fly ash Table I

Table II

World REE reserves and production (US Geological Survey, 2021)

he recoverable reserve (CURR) in coal for different T scenarios (adapted from Zhang et al., 2015)

Country

Source

Production, t

USA Australia Brazil Myanmar Burundi Canada China Greenland India Madagascar Russia South Africa Tanzania Thailand Vietnam Other countries Total (rounded)

Reserves, t

2019

2020

28 000 20 000 710 25 000 200 – 132 000 – 2 900 4 000 2 700 – – 1 900 1 300 66 220 000

38 000 17 000 1 000 30 000 500 140 000 3 000 8 000 2 700 – – 2 000 1 000 – 240 000

1 500 000 4 100 000 21 000 000 – – 830 000 44 000 000 1 500 000 6 900 000 12 000 000 790 000 890 000 – 22 000 000 310 000 120 000 000

It is well documented that REE are mainly concentrated in bastnasite, allanite, monazite, xenotime, florencite, kimuraite, lanthanite, and zircon, and in some cases form composite particles with clay minerals such as kaolinite and illite (Dai et al., 2012; Gupta et al., 2017; Hood et al., 2017; Seredin and Dai, 2012; Sun et al., 2007; Zhang et al., 2015). Depending on the source and composition of the coal, the gangue components of coal fly ash include substantial amounts of quartz, calcite, siderite, rutile, and illite. The differences between physical and physicochemical characteristics (density, magnetic susceptibility, conductivity, surface hydrophobicity) of REE and gangue minerals may be exploited to recover and upgrade REE through gravity concentration, magnetic separation, froth flotation, and electrostatic separation (Zhang et al., 2015).

Lin et al. (2017) studied the enrichment of REE from coal and coal by-products using particle size, magnetic, and density separation methods. Particle size separation was conducted using 20–150 μm sieves mounted on a 3-inch shaker equipped with an electromagnetic vibrator. The experimental results indicated that the REE were enriched in the nonmagnetic products after separation. Elsewhere, Blissett, Smalley, and Rowson (2014) investigated coal fly ashes from the UK and Poland to evaluate any chance of recovering their REE content. The work demonstrated marginal REE upgrading after classification of the nonmagnetic inorganic content of the samples.

As a result of the increased use of biomass as a co-firing fuel in power plants, there is a corresponding increase in production of fly ash, with varying chemical properties which may not be suitable for traditional application in the cement/ concrete industry (Franus, Wiatros-Motyka, and Wdowin, 2014; Wdowin et al., 2014). Coal fly ash forms the bulk of the fly ash generated globally at an annual production rate of 750 Mt (Perämäki, Tiihonen, and Väisänen, 2019). Hence, it is important to investigate the recovery of REE as an alternative application for coal fly ashes. The aim of this work is to exploit the differences in the physical and physicochemical properties of REE and gangue minerals to recover and upgrade REE from a typical coal fly ash (CFA) produced from a commercial power plant.

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Total REE, t

Mohr and Evans (2009) Rutledge (2011) World Energy Council (2007)

49 211 600 50 666 237 61 333 575

Coal ultimate recoverable reserves, Gt

680 700.1 847.5

Materials and methods Materials CFA samples from a coal-fired commercial power plant in Australia were used in this investigation. The samples were dried and split into representative subsamples and stored in vacuum in plastic bags for subsequent characterization and beneficiation studies. The elemental composition of the CFA was obtained using inductively coupled plasma mass spectroscopy (ICP-MS).

At the end of each beneficiation test, the separation products were weighed, dried, and their chemical content determined by ICP-MS. Particle size analysis was conducted using a Malvern Mastersizer 2000 laser-diffraction instrument with a measurement range of 0.02–2000 μm. A representative sample of CFA was prepared into a pulp (25 wt.%) and stirred at 800 r/min for 30 minutes, to de-agglomerate the particles. Subsamples were drawn and analysed using the Malvern Mastersizer 2000. The results are presented in Figure 1. The data obtained was used to estimate the 10th, 50th, and 90th percentile particle sizes, d10, d50, and d90, respectively, which are presented in Table III.

Particle size separation The distribution of REE in the CFA as a function of particle size was investigated by wet screening tests. A set of test sieves was selected and arranged in sequence 150, 106, 75, and 38 µm to screen approximately 150 g of representative CFA sample.

The particles retained on the respective sieves were dried in an electric oven at 70°C for 12 hours, weighed, and analysed for their chemical content.

Gravity separation Gravity separation tests were conducted using a Knelson concentrator (KC). The KC was operated at a feed flow rate of 150 g/min at a pulp density of 25 wt.% solids. During the tests, the pulped material was diluted with fluidization water at flow rate from 2.5 to 7.5 L/min, while the bowl rotation speed was set at 73 G. The concentrate and tailing fractions obtained were submitted for chemical analysis.

Magnetic separation A wet high-intensity magnetic separator (Model L-4) was used to investigate the recovery of REE at different magnetic field intensities. For each test, 50 g of CFA sample was mixed with deionized water to yield a 17wt.% pulp (Abaka-Wood, AddaiMensah, and Skinner, 2016, Abaka-Wood et al., 2019a). The pulp was thoroughly mixed and transferred with additional water at a rate of 1 L/min to minimize entrainment of nonmagnetic fractions on the matrix during operation. The magnetic field intensity was The Journal of the Southern African Institute of Mining and Metallurgy

The concentration of rare earth elements from coal fly ash

Figure 1—Frequency–particle size distribution of the CFA material obtained using the Malvern Mastersizer

Table III

Table IV

CFA particle size data obtained by laser diffraction

hemical composition of CFA sample determined by C ICP-MS

Percentile

d10 d50 d90 Volume-weighted mean, µm

Value µm

3 16 119 48

adjusted (0.11–1.74 T) by increasing the current from 1 to 21 A. The wet high-intensity magnetic separator (WHIMS) was fitted with a medium expanded metal (MEX) matrix.

Flotation separation A 250 mL Polish microflotation cell (Instytut Metali Niezelaznych, Gliwice) was used in the flotation of CFA samples. During the tests, 50 g of dried CFA was transferred into the flotation cell along with distilled water. The impeller speed and air flow rate were set at 720 r/min and 1.5 L/min, respectively. The initial pH of the pulp was noted and adjusted to the desired value using HCl or NaOH. Oleic acid was used a collector. During the flotation tests, 1000 g/t oleic acid was added with 5 minutes conditioning time. Froths were collected every 15 seconds over 10 minutes. The flotation products were dried, weighed, and analysed by ICP-MS.

Results and discussion Concentrations and recovery potential The ICP-MS analysis (Table IV) revealed that REE in the CFA material were dominated by cerium (Ce, 167 ppm), lanthanum (La, 87.4 ppm), neodymium (Nd, 65.6 ppm), and yttrium (Y, 58 ppm), which made up about 85% of the total REE (TREE) content. Aluminium (Al), silicon (Si), calcium (Ca), and iron (Fe) were the major gangue elements in the CFA sample. The outlook coefficient, Koutl, which is the ratio of critical (Nd, Eu, Tb, Dy, Y, and Er) to excessive (Ce, Ho, Tm, Yb, and Lu) REE was proposed by Seredin (2010) to assess the market quality of REE deposits. REE resources with a high Koutl are regarded as profitable.

The Koutl of the CFA material was calculated to be 0.82, which makes the CFA promising for potential economic extraction of REE, especially since no mining and comminution costs will be incurred. Specifically, the CFA sample contained approximately The Journal of the Southern African Institute of Mining and Metallurgy

REE

Concentration, Non-REE ppm

La Ce Pr Nd Sm Eu Gd Tb Dy Ho Er Tm Yb Lu Yb Y TREE

87.4 Al 167 Ca 18.2 Fe 65.6 Mg 12.7 P 2.5 Si 11.6 K 1.68 U 10.1 Th 1.88 Na 5.55 0.7 4.2 0.7 4.2 58 447.8

Concentration, %

10.9 3.91 5.34 0.92 0.08 23.1 1.54 0.001 0.003 0.37

32% critical REE. The critical REE content and Koutlare more comparable to those obtained for Finnish peat and biomass combustion fly ash, which contain 28–32% critical REE with Koutl values >0.7 (Perämäki, Tiihonen, and Väisänen, 2019). The TREE concentration in CFA (447.81 ppm) is similar to a typical coal fly ash, which has an average concentration of 404 ppm (Perämäki, Tiihonen, and Väisänen, 2019; Seredin and Dai, 2012). According to the data obtained, the CFA can be considered promising for recovery of REE, based on the outlook coefficient and critical REE content. However, the TREE concentration is less than 800 ppm, which has been suggested as the cut-off grade for coal seams thicker than 5 m (Seredin, (2010). It is worth noting that the CFA used in the present study is already of a suitable particle size for subsequent beneficiation.

Particle size separation Wet sieve analysis was conducted to study REE distribution in the various size fractions. The results are summarized in Figure 2, which shows the REE concentration, REE distribution, and mass yield of the respective particle size fractions. The results indicate variable REE concentrations in the respective size fractions, with VOLUME 122

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The concentration of rare earth elements from coal fly ash

Figure 2—Mass yield, REE concentration, and REE distribution within the respective particle size fractions of CFA

Figure 3—REE recovery and grade as a function of fluidization water flow rate

+150 µm having the least TREE concentration of 441 ppm, whereas the highest TREE concentration was measured in the –75 +53 µm and –150 +106 µm size fractions. However, the largest proportion of REE in the CFA was contained within the –38 µm size fraction, which accounted for 67% and 69% of the TREE and feed mass, respectively.

Gravity separation Gravity separation tests were conducted using a laboratory-scale KC. During operation, the concentrating beds formed within the KC are prevented from consolidating by means of a water current which flows through the fluidization holes, leading to the formation of fluidized beds within each ring (Abaka-Wood et al, 2019c). The fluidization water flow rate was varied from 2.5– 7.5 L/ min in this investigation.

The results obtained have been summarized in Figure 3 and Table V. It can be seen that REE recovery and upgrade is significantly affected by changes in the fluidization water flow rate. An increase in the water flow rate from 2.5 L/min to 7.5 L/ min saw a decrease in REE recovery from 34% to 22%, with a corresponding increase in REE grade from 352 ppm to 407 ppm.

REE minerals are known to have specific gravities typically >4.0 24

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(Subbarao, 1980), hence they were expected to report with the KC concentrate. As shown in Table V, the bulk of REE reported with the KC tailings along with Si particles. Notably, about 53% of the Fe was recovered into the KC concentrate at 2.5 L/min water flow rate. This decreased significantly as the fluidization water flow rate was increased. The results suggest that the recoveries of both Fe and Si followed the same trend as the REE.

Magnetic separation The influence of magnetic field intensity on REE recovery was studied using a laboratory-scale WHIMS. In this test, the magnetic field intensity was increased from 0.11 T to 1.74 T. The cumulative recoveries of REE, Fe, and Si at the applied field intensities are shown in Figure 4. REE recovery increased from 11% to 37% with an increase in applied field intensity from 0.11 T to 1.74 T, with a corresponding increase in grade from 328 ppm to 516 ppm.

The bulk of the REE in the CFA reported to the nonmagnetic fraction, which accounted for 63% and 60% of the TREE and mass, respectively, with an REE grade of 490 ppm.

The high REE content in the nonmagnetic fraction can be attributed to the fact that most of the REE particles may be associated with the nonmagnetic minerals such as quartz and The Journal of the Southern African Institute of Mining and Metallurgy

The concentration of rare earth elements from coal fly ash Table V

Results of KC tests on CFA sample Fluidization water flow, L/min

2.5 5.0 7.5

REE Grade, ppm

Fe recovery, %

Si recovery, %

Recovery, %

352 418 407

34 31 22

53 33 32

42 31 2

Figure 4—Cumulative REE, Si, and Fe recoveries, REE grade, and mass yield in different magnetic separation products

Figure 5—Effects of pulp pH on REE recovery and upgrade from CFA using oleic acid as a collector

other silicates. This is shown by the high Si content (64%) in the nonmagnetic fraction. Furthermore, REE recovery follows both the Si recovery and mass yield trends. Another significant observation was the high Fe content (56%) in the concentrate obtained at an applied field intensity of 0.11 T, and the low Fe content (28%) in the nonmagnetic fraction. The results suggest that half of Fe content in the CFA could be rejected at low magnetic field intensity, while upgrading REE in the corresponding nonmagnetic fraction along with silicate minerals.

Flotation The effects of changes in pulp pH on REE recovery and upgrade were investigated during the flotation experiments. The pulp pH The Journal of the Southern African Institute of Mining and Metallurgy

was decreased from the pristine CFA pulp value of 11 ± 0.5 to 7.

Dilute solutions of HCl and NaOH were used as pH modifiers. The results presented in Figure 5 show that a decrease in pulp pH from 11 to 9 caused a significant decrease in REE recovery from 65% to 13%, and a corresponding increase in REE grade from 498 ppm to 678 ppm. A further decrease in pH from 9 to 7 resulted in an increase in recovery to 37% at a concentrate grade of 633 ppm. The results suggest that the recovery and upgrade of REE from CFA is pulp pH-dependent. In terms of REE recovery, the best result was achieved at pulp pH 11. This is consistent with flotation tests carried out by Satur et al. (2016), where fatty acids were used to achieve the highest REE recovery from a silicate–haematite ore at pH 11.

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Figure 6—Ratio of REE/Fe and REE/Si in flotation concentrates obtained at pulp pH 7–11

Figure 6 presents the ratios of REE content to Si and Fe in the flotation concentrates obtained at the respective pulp pH levels tested. The results suggest significant proportions of both Fe and Si in the froths recovered. However, lower quantities of Fe and Si reported to the concentrates obtained at pH 9, which explains the highest REE concentrate grade achieved. It can be noticed that the recoveries of Fe and Si follow those of mass yield and REE recovery, respectively. It has been demonstrated that Fe can be recovered using oleic acid at mild to strong alkaline pulp pH values (Abaka-Wood, Addai-Mensah, and Skinner, 2017a, 2017b; Abaka-Wood et al., 2019b; Joseph-Soly, Quast, and Connor., 2015; Quast, 2000, 2006). The high Si recoveries could be attributed to entrainment due to the presence of a high level of ‘fine’ particles in the feed (Duarte and Grano, 2007; Leistner Peuker,and Rudolph, 2017).

Implications and future work The CFA used in the present study had a TREE grade of 447.8 ppm, with about 32% critical REE and an outlook coefficient of 0.82. This places the CFA within a potentially economic range if feasible beneficiation methods could be developed to selectively recover and upgrade REE values. The present study indicates that significant REE recovery could be achieved by flotation. However, during both magnetic and gravity separation, the bulk of the REE reported with the tailings fractions, which suggests that both these physical separation methods may be combined with flotation as an appropriate beneficiation strategy in future investigations.

For instance, either gravity or magnetic separation may be optimized to remove Fe-bearing particles prior to concentrating REE via froth flotation. Depressants could be used to minimize the recovery of silicate gangue minerals during REE beneficiation.

For example, sodium silicate, disodium sulphide, starch, ammonium lignosulfonate, and sodium oxalate have been shown to be effective depressants of various gangue minerals in REE flotation (Abaka-Wood, Addai-Mensah, and Skinner, 2017b; Chelgani et al., 2015; Satur et al., 2016; Zhang and Anderson, 2017).

A combination of the respective separation methods will help achieve a concentrate recovery and grade for subsequent REE extraction by hydrometallurgical or pyrometallurgical processes.

Elsewhere, Abaka-Wood et al. (2019b), Jordens et al. (2016), 26

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Yang et al. (2015), and Xiong et al. (2018) have demonstrated the feasibility of achieving significant REE recovery and upgrade through the combination of selected physical separation methods and flotation. These previous investigations will serve as a guide in future studies which will target producing richer REE concentrates from the CFA. However, prior to any further studies with the view of improving the results obtained in the present work, a detailed mineralogical study of the CFA to determine the true chemical and physical deportment of the REE and gangue species will be carried out. Mineralogical studies including quantitative X-ray diffraction (QXRD) and Quantitative Evaluation of Minerals

by Scanning Electron Microscopy (QEMSCAN) analyses have been demonstrated to be crucial in establishing REE and gangue minerals deportment in different ores (Edahbi et al., 2018; Smythe et al., 2011; Van Rythoven et al., 2020). Smythe, Lombard, and Coetzee (2013) pointed out that QEMSCAN analysis provides useful information in estimating elemental recoveries and selecting potential beneficiation techniques for recovering minerals of interest. Furthermore, it was suggested that QEMSCAN analysis gives a good indication of the textural relationships, mineral association, and liberation characteristics between the mineral of interest and gangue minerals, which is crucial in selecting concentration processes for recovering REE.

Conlusions The current study highlights the potential for CFA from a commercial power plant to contain significant REE reserves for subsequent beneficiation. Chemical characterization of representative CFA samples showed a concentration of 447.8 ppm TREE. The bulk of the REE in the CFA sample was concentrated in the fine fraction (< 38 µm), which contained more than half of the TREE. Furthermore, the varying distribution of REE between the respective size fractions indicates that both the coarse and fine fractions of CFA can be considered for REE beneficiation. The high outlook coefficient, coupled with the fact that CFA carries no mining and comminution costs, suggest that the recovery of REE could present a potential economic advantage.

Based on the results obtained, the KC does not appear to be a suitable gravity concentration method for concentrating REE from the CFA used in the present study. This is shown by the low recoveries and poor upgrades in the KC concentrates produced.

However, WHIMS tests on CFA showed appreciable REE upgrade in magnetic concentrates produced at 1.08–1.74 T, although the corresponding recoveries were low. The results point out that the bulk of the REE were concentrated in the nonmagnetic fraction.

Flotation using oleic acid resulted in higher REE concentrations in the froths produced than in the tailings. The pulp pH affected the recovery and upgrade of REE, with the highest recovery achieved at pH 11, whereas the best REE upgrade occurred at pH 9, where the lowest recovery was observed. Overall, the study suggests that beneficiation processes combining the methods employed in the present study may achieve enhanced REE recoveries and upgrades. This will be investigated in future work, along with hydrometallurgical or pyrometallurgical separation processes.

Acknowledgements This work was supported by the Australian Government Research Training Program Scholarship and Future Industries Institute of the University of South Australia (Adelaide, Australia).

The Journal of the Southern African Institute of Mining and Metallurgy

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ANNUAL MEETING AND CONFERENCE

NAMIBIA UNIVERSITY OF SCIENCE AND TECHNOLOGY

The concentration of rare 8-11 earth elements fly ash VISITS SEPTEMBER 2022from TOURScoal AND TECHNICAL

12-14 SEPTEMBER 2022 CONFERENCE

INVITATION FROM THE ORGANIZING COMMITTEE

It

is a great honour, that I hereby and accountability to local extend this invitation to you, communities is more pronounced, dear esteemed members, to as voice continues to be added the 32nd Society of Mining Professors to expectation, with unparalleled Annual Meeting and Conference clarity. Our contribution to the to be hosted by the Namibian mining industry discourse, in the University of Science and Technology form of relevant research and welland at the Windhoek Country Club trained graduates, remains a top and Resort, Windhoek, Namibia, priority. from 8–14 September 2022. I wish to Namibia is richly endowed with extend my fraternal greetings to the mineral resources, and the delegates, SOMP members, spouses mining industry significantly and partners who will travel from contributes to the national economy. around the globe to be part of this Namibia also boasts a strategic Prof.

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A new preconcentration technique for the determination of PGMs and gold by fire assay and ICP-OES

by A. Masasire1,3, F. Rwere2, P. Dzomba3, and M. Mupa3 Affiliation:

S iyanda Union Mine,, Limpopo, RSA. 2 Chinhoyi University of Technology, Department of Chemistry, Chinhoyi, Zimbabwe.

3 Bindura University of Science Education, Department of Chemistry, Bindura, Zimbabwe.

1

Correspondence to: M. Mupa

Email: mathewmupa@buse.ac.zw

Dates:

Received: 23 May 2021 Revised: 23 Nov. 2021 Accepted: 29 Nov. 2021 Published: January 2022

How to cite:

Masasire, A., Rwere, F., Dzomba, P., and Mupa. M. 2022.

A new preconcentration technique for the determination of PGMs and gold by fire assay and ICP-OES. Journal of the Southern African Institute of Mining and Metallurgy, vol. 122, no. 1, pp. 29-36 DOI ID: http://dx.doi.org/10.17159/24119717/1638/2022 ORCID: A. Masasire https://orcid.org/ 0000-00027403-4962 F. Rwere https://orcid.org/ 0000-00027037-6850 P. Dzomba https://orcid.org/ 0000 0016821 2606 M. Mupa https://orcid.org/ 0000-00032185-9386

Synopsis Accurate determination of platinum group metals (PGMs) and gold (Au) has always proven to be a difficult task, due to their low concentrations in platiniferous ores. The most common preconcentration technique used in analysis of these metals is fire assay with a flux containing nickel or lead. This technique can be improved by using co-collectors. Here we hypothesise that Fe, Co, and Cu can be used as co-collectors to enhance separation and preconcentration of PMGs and gold by fire assay. To test this hypothesis, geological exploration samples from Siyanda Union Mine (Northam, South Africa) were analysed by an inductively coupled plasma-optical emission spectrometer (ICP-OES) for PGMs (Pt, Pd, and Rh) and Au. A control sample (certified reference material AMIS 0426) was also analysed by the same technique. The PGM and Au recoveries from the control sample ranged from 83% to 105% for all three co-collectors, with relative standard deviations less than 10% for the control sample and 7% for the geological samples. The PGM and Au recoveries by Fe and Co co-collectors were modestly higher than that of the Cu-collector. These results indicate that Fe and Co are better co-collectors than Cu, presumably due to the loss of analyte when using Cu. Taken together, Fe and Co co-collectors can be viable alternatives for analysis of PGMs and gold using the fire assay method.

Keywords Platinum group metals, ICP-OES, co-collection, fire assay.

Introduction South Africa is a prominent global supplier of platinum group metals (PGMs). The main reserves of PGMs are the Bushveld Complex (Tanner et al., 2019), where platiniferous ores are obtained from the Merensky Reef (Creech et al., 2014). The PGMs are found in low concentrations, typically less than 10 g/t (Serbin, Bazel, and Ru, 2017). The low PGM grade of these ores contributes to their high market value.

The analysis of the platiniferous ores and flotation concentrate samples is challenging because of the low PGM concentrations and their heterogeneous distribution in the matrix. Therefore, their determination is usually preceded by isolation from the gangue material and preconcentration (Berezhnaya and Dubinin, 2016; Bayrak et al., 2017). In South African mining and metallurgical testing laboratories this is accomplished mainly by fire assay using nickel sulphide or lead collection followed by spectrometric determination (Vanhaecke et al., 2010).

The growing demand for PGMs and Au has led to concerns about their future supply. This has resulted in a renewed interest in the recycling of end-of-life materials. The supply of PGMs from recycling has doubled over the past decade. Makua et al. (2019) recovered PGMs from a pregnant leach solution by using solvent extraction and cloud-point extraction. They concluded that the efficiency of the cloud-point extraction method depends on the pH of the solution, the surfactant and the complexing agent, the hydrochloric acid concentration, and the presence of a reducing agent. Although the cloudpoint extraction method is more environmentally friendly than fire assay, it is time-consuming.

Carelse et al. (2020) assessed the distribution of gold and silver in alloys produced from the smelting of printed circuit boards, using SEM-EDS, EPMA, and LA-ICP-MS analyses. Gold and silver were found to be most enriched in the lead phase of the tap, which indicates that lead is a good collector of gold.

Different analytical techniques have been used to determine PGMs and Au in geological work because

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A new preconcentration technique for the determination of PGMs and gold by fire assay and ICP-OES of their economic importance (Hughes, McDonald, and Kerr, 2015; Jansen et al., 2016). Liipo et al. (2019) characterized the South Georgian complex copper-gold ores by fire assay and ICP-OES. Other authors have attempted to determine PGMs and Au in ores using X-ray fluorescence (XRF), but have encountered sensitivity problems (Hinds and Burgess, 2014; Díaz, Hahn, and Molina, 2017). In addition, a variety of techniques have been employed for sample decomposition and preconcentration of PGMs and gold prior to ICP-OES measurement. These studies have concentrated on fire assay with nickel sulphide (NiS) collection to characterize the metals in geological samples. However, low recoveries of the PGMs using NiS collection are reported in the literature (Morcelli et al., 2004). In other studies, co-precipitation with Te after HCl digestion of the NiS button has been shown to improve PGM recoveries (Morcelli et al., 2004). However, the results for Pd and Pt are reportedly lower than the certified values for reference materials (Morcelli et al., 2004).

It has also been shown in some studies that co-collectors can improve the recovery of PGMs, and allow accurate determination in a possibly cost-effective manner. Modification of the Pb fire assay procedure using Ag or Au as collector for the PGMs from rocks, minerals, and ores has been extensively reviewed (Balcerzak, 2002). Co-collectors normally used in the determination of PGMs include silver, platinum, palladium, and gold (Ndovorwi, 2014). Suominen, Kontas, and Niskavaara (2004) used Au and Ag as co-collectors in the fire assay analysis of Pd, Pt, and Rh in geological reference materials and showed variable recoveries with Ag and better recoveries with Au. Studies on iridium (Ir) and ruthenium (Ru) as co-collectors for PGMs and gold in ores and concentrates showed that Ir was a useful cocollector for concentrate material only at very low concentrations, and Ru was not useful at either low or high concentrations (Ndovorwi, 2014). However, limited research has been done to demonstrate the effectiveness of Fe, Cu, and Co as co-collectors for accurate analysis of PGMs and gold. Using Fe, Cu, and Co can be cost-effective compared to precious metals like silver, platinum, palladium, and gold. Accurate determination of PGMs and Au in geological samples provides important information for mineral exploration (Volzhenin et al., 2018). Therefore, the goal of this study was to develop a lead-based fire assay method for the preconcentration of PGMs and Au using iron, cobalt, and copper as co-collectors and quantitative determination using ICP-OES.

Experimental procedure Instrumentation Basic mining laboratory equipment was used for crushing, pulverizing, milling, and splitting of the samples. Industrial Analytical, 2018 model muffle furnaces (0-1300°C) were used for

fusion and cupellation. The ThermoFisher ICAP 7400 radial model ICP-OES instrument (Germany) was used to determine the PGMs and Au.

Reagents Analytical grade nitric acid (55%) and hydrochloric acid (32%) were obtained from Merck (Germany). The lead-based flux was obtained from Terranova (South Africa) and ICP standard reference solutions were obtained from De Bruyne (South Africa).

AMIS 0426, a UG2 ore-based certified reference material, was obtained from African Mineral Standards (South Africa).

Sampling and pre-processing Borehole core sampling was conducted at the Siyanda Union Mine Northam, South Africa as per their mine sampling standards. A total of 30 samples were collected for analysis. The samples were crushed, pulverized, and milled to minimum 95% passing the 75 μm sieve. The milled samples were homogenized in a blender for 48 hours. A map of the sampling points is shown in Figure 1.

Sample fusion The fluxing method was adapted from Rodríguez-Rodríguez and Miguel, (2018) with slight modifications. Briefly, 100 g of each sample was mixed with 300 g of the lead-based flux comprising Na2CO3 (39.1%), borax (22.0%), SiO2 (9.23%), starch from corn meal (3.0%), PbO (26.4%), CaF2 (2.3%), and paraffin (900 mL).

Then, 20 mL of 20 mg/L solution of Cu, Fe, and Co were added to 10 geological samples and ten certified reference material samples (AMIS 0426). The mixture was introduced into a clay crucible, the surface was covered, and the crucible was placed in a pre-heated muffle furnace. Three fusion conditions were employed, viz. 900 ± 50°C for 60 minutes, 1100 ± 50°C for 60 minutes, and 1200 ± 50°C for 60 minutes. Once the fusion was completed (no effervescence was observed in the melted sample), the crucible was removed from the furnace and its contents poured into an iron mould.

After solidification, the crystals that were formed by the slag were crushed with a hammer to release the lead button.

Cupellation The cupellation method was adapted from Rodríguez-Rodríguez and Miguel, (2018) with slight modifications. The cupellation temperature was increased from 900°C to 1000 ± 50°C. The lead button was put in a magnesite cupel (previously dried at 1000 ± 50°C for 60 minutes) and ignited in a furnace at 1000 ± 50°C until the lead melted. The furnace door was left slightly open so that the lead could be oxidized and most of the lead could be absorbed by the cupel. The temperature was kept constant until all the lead was removed. At the end of this process, a button of PGMs and Au (a prill) was obtained. Table I shows the fusion and cupellation

Figure 1—Map of the sampling points for geological exploration samples (location coordinates -24.97841395, 27.1433952) 30

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A new preconcentration technique for the determination of PGMs and gold by fire assay and ICP-OES conditions. Figure 2 illustrates the analytical steps involved in the of determination of the PGMs and Au by ICP-OES.

The ICP-OES parameters as shown in Table III were adopted from Tao et al. (2017) with minor modifications.

Dissolution of prill

Analytical performance

The prill dissolution and instrumental analysis methods were adapted from Tao et al., (2017) with slight modifications. Briefly, a PGM and Au prill was placed in a 10 mL volumetric flask containing 1 mL of concentrated HNO3.The mixture was boiled until there was no effervescence, followed by the addition of 3 mL of concentrated HCl, Then, 2 mL of 50 mg/L yttrium (10 mg/L final concentration) was added to the mixture as an internal standard. The solution was diluted to the 10 mL mark using deionized water prior to analysis.

Using the optimum conditions, the intensity of each of the platinum group metals and gold was determined and quantified by the Thermofisher ICP-OES ICAP 7400 radial with yttrium (10 mg/L) as an internal standard. The analytical performance of the proposed fire assay method was validated by the linear range, the limit of detection (LOD), the correlation coefficient, and the relative standard deviation, as illustrated in Table IV. The linear range was between zero and 120 mg/L, with correlation coefficients from 0.997 to 0.999. Based on the 3-sigma blank

Preparation of calibration standards The multi-element reference standards contained Pt (1000 mg/L), Pd (500 mg/L), Rh (200 mg/L), and Au (50 mg/L). Table II lists the concentrations of the standard solutions used to calibrate the ICP-OES instrument.

Yttrium was used as an internal standard (1 mL of 1000 mg/L stock solution). The working standards were made up to a 100 mL volume with deionized water. The control and geological exploration samples from each co-collector were analysed in 10 replicates.

Fluxing The sample is mixed with leadbased flux in heat resistant clay crucibles

Table II

Working standard solutions for ICP-OES calibration

Vol (L) Pt (mg/L) Pd (mg/L) Rh (mg/L) Au (mg/L)

Blank

Std 1

Std 2

Std 3

0 0 0 0 0

1 10 5 2 0.5

2 20 10 4 1

4 40 20 8 2

Std 4 Std 5 Std 6

6 60 30 12 3

Fusion furnace

The sample and flux in clay crucibles are melted at high temperature to separate the PGMs and Au from slag

Cupellation The lead button is heated at high temperature (1000°C ± 50°C) in a furnace to recover PGM and Au prill

ICP-OE S.

Prill dissolution.

The PGM and Au solution is analysed using Thermo Scientific ICAP 7400 ICP-OES.

8 80 40 16 4

12 120 60 24 6

The prill is dissolved in aqua regia and solution made to mark with de-ionized water

Figure 2—Analytical steps for the determination of PGMs and Au using ICP-OES

Table I

Experimental conditions for fusion and cupellation Condition

Parameters

Fusion

Cupellation

Co-collector (mg/L)

1

Temperature Time Temperature Time Temperature Time

900 ± 50°C 60 min 1100 ± 50°C 60 min 1200 ± 50°C 60 min

1000 ± 50°C 60 min 1000 ± 50°C 60 min 1000 ± 50°C 60 min

20

2 3

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A new preconcentration technique for the determination of PGMs and gold by fire assay and ICP-OES Table III

ICP-OES operating parameters ICP-OES instrument

ThermoFisher ICAP 7400

RF power (W) Auxiliary gas flow (L/min) Nebulizer gas flow (L/min) Coolant gas flow (L/min) Pump speed (r/min) Viewing height (mm) Plasma view Element

1200 0.5 0.5 12 50 12 Radial Emission line, nm

Pt Pd Rh Au

214.423 340.450 369.236 208.209

approach, as recommended by IUPAC for spectrochemical measurements (Tao et al., 2017), the limits of detection of the proposed method for the target platinum group metals and gold in a 100 g sample were in the range of 0.001 mg/L to 0.047 mg/L. Figure 3 show the calibration curves for Pt, Pd, Rh, and Au. The analytical figures of merit by fire assay ICP-OES are also shown in Table IV.

Results Table V show the average elemental compositions for the control

sample and the geological exploration samples obtained using Fe, Co, and Cu co-collectors according to the experimental procedures described in Table I. Table V also shows the percentage relative standard deviation (RSD) and percentage recovery for PGMs and Au. The recoveries obtained from the Fe co-collector were 97.4% for Pt, 87.6% for Pd, 95.8% for Rh, and 104.9% for Au. In case of Co co-collector, the recoveries were 99.5% for Pt, 88.7% for Pd, 90.2% for Rh, and 97.2% for Au. Finally, the recoveries obtained from the Cu co-collector were 92.2% for Pt, 88.1% for Pd, 83.1% for Rh, and 88.1% for Au. These results demonstrate that Fe and Co co-collectors can be used for the determination of PGMs and Au. However, there is slight loss of analyte using Cu as co-collector for Pt and Pd, as indicated by the relatively low recoveries.

The precision of PGMs and Au determinations was calculated as percentage relative standard deviation obtained from ten separate determinations of PGM concentrations in the reference material and geological exploration samples. The RSD values for the certified reference material were less than 7% for Pt and Pd using all three co-collectors, whereas for Rh the RSD was greater than 7% for Fe co-collector (10.97%) and less than 7% for Co and Cu co-collectors. In the case of Au, the RSD values for Fe and Co co-collectors were all greater than 7%. These values demonstrate that the determination was not highly precise for Au, but moderate for Rh and relatively precise for Pt and Pd using the three co-collectors. Also shown in Table V are the elemental compositions and RSD values of the geological samples. As shown in Table V, the RSDs for geological exploration samples obtained

Figure 3—ICP-OES calibration graphs for Pt, Pd, Rh, and Au

Table IV

Analytical figures of merit by fire assay and ICP-OES Element

Pt Pd Rh Au 32

Linear range

Linear equation

Correlation coefficient

LOD (mg/L)

0–120 mg/L 0–60 mg/L 0–24 mg/L 0–6 mg/L

Y = 13.66*X + 4.231 Y = 159.1*X + 61.83 Y = 76.77*X + 29.05 Y = 37.07*X - 7.991

0.999 0.999 0.999 0.997

0.047 0.025 0.080 0.001

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A new preconcentration technique for the determination of PGMs and gold by fire assay and ICP-OES using experimental condition 2 in Table I for Fe, Co, and Cu co-collectors is less than 7%, a result that demonstrates that the procedure employed is relatively precise.

Tables VI and VII show statistical accuracy testing results for control and geological exploration samples performed using a Student’s t-test. As shown in Table VI, the ρ-value for Pt (0.011) is less than the α-value (0.05) with the Fe co-collector, an indication that there is a significant difference between the mean Pt value and the certified reference value. Pd, Rh, and Au have ρ-values greater than α-value with the Fe collector, suggesting that there is no significant difference between the analysis results and the certified values. In case of Co, the ρ-value for Rh is less than the α-value (ρ value=0 vs α=0.05), whereas the ρ-values for Pt, Pd, and Au are greater than the α-value. These results demonstrate

that for the Co co-collector, there is a significant difference between the mean Rh value and the certified reference value of the control sample. With Cu co-collection the results for Pt, Rh, and Au were significantly different from the certified values (ρ-values < α =0.05), and there was no significant difference in the Pd results.

The statistical results for geological samples are shown in Table VII. The ρ-values for Pt and Pd are greater that the α-value for all the co-collectors, suggesting that Fe, Co, and Cu can be useful co-collectors for geological samples. For Au, the ρ-values are less than the α-value for all co-collectors, demonstrating that the mean values are significantly different from each other. The mean Rh results obtained from the Co and Cu co-collections showed no significant difference. However, the mean results from

Table V

Summary results for control and geological exploration samples using Fe, Co, and Cu as co-collectors Element Co-collector *Sample ID

Pt Fe

Co

Cu

Pd Fe

Co

Cu

Rh Fe

Co

Cu

Au Fe

Co

Cu

Certified reference material concentration (mg/L)

AMIS 0426 (10) 2.20 ± 0.17 Geo (10) AMIS 0426 (10) 2.20 ± 0.17 Geo (10) AMIS 0426 (10) 2.20 ± 0.17 Geo (10) AMIS 0426 (10) 1.07 ± 0.12 Geo (10) AMIS 0426 (10) 1.07 ± 0.12 Geo AMIS 0426 (10) 1.07 ± 0.12 Geo (10) AMIS 0426 (10) 0.41 ± 0.05 Geo (10) AMIS 0426 (10) 0.41 ± 0.05 Geo (10) AMIS 0426 (10) 0.41 ± 0.05 Geo AMIS 0426 (10) 0.041 ± 0.02 Geo (10) AMIS 0426 (10) 0.041 ± 0.02 Geo (10) AMIS 0426 (10) 0.041 ± 0.02 Geo (10)

Elemental composition (mg/L) and ±SD

RSD (%)

Recovery (%)

2.143 ± 0.054 2.157 ± 0.043 2.190 ± 0.054 2.137 ± 0.042 2.034 ± 0.044 2.122 ± 0.081 0.937 ± 0.056 0.848 ± 0.015 0.951 ± 0.025 0.858 ± 0.011 0.943 ± 0.022 0.857 ± 0.013 0.392 ± 0.043 0.364 ± 0.014 0.369 ± 0.012 0.332 ± 0.014 0.340 ± 0.015 0.346 ± 0.014 0.043 ± 0.004 0.016 ± 0.001 0.040 ± 0.003 0.017 ± 0.001 0.037 ± 0.003 0.015 ± 0.001

2.52 1.99 2.45 1.97 2.16 3.82 5.98 1.77 2.63 1.28 2.33 1.52 10.97 3.85 3.25 4.22 4.41 4.05 9.30 6.25 7.50 5.88 8.11 6.67

97.4 99.5 92.5 87.6 88.9 88.1 95.8 90.0 82.9 104.9 97.5 90.2

*n-value is in () Table VI

Statistical accuracy testing results for the control sample in relation to the certified values Co-collector Element

Fe

Co

Cu

Certified reference material concentrations (mg/L)

ρ value (significance level used was α = 0.05)

2.20 1.07 0.41 0.04 2.20 0.98 0.41 0.04 2.20 0.98 0.41 0.04

0.011 0.553 0.659 0.053 0.613 0.871 0.000 0.932 0.000 0.386 0.000 0.016

Pt Pd Rh Au Pt Pd Rh Au Pt Pd Rh Au

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A new preconcentration technique for the determination of PGMs and gold by fire assay and ICP-OES are not statistically different, indicating that the co-collecting capabilities of Pd are comparable. The Rh results from Co, Fe, and Co-Cu co-collectors show no difference. However, the Fe and Cu results are different. The Au results from all co-collectors were significantly different.

Table VIII also shows that the variances of Pt and Pd from all the co-collectors are statistically equal, but the Au results are statistically different since the ρ-values are all less than α. The Rh results from Co and Cu co-collection show no significant differences; however, the results of Fe, Co, and Fe-Cu cocollection indicate a significant difference.

Table VII

Statistical accuracy results for the geological exploration samples by comparison of the means ρ value (significance level used was α = 0.05)

Element Collector

Pt Pd Rh Au

Fe -Co Fe-Cu Co-Cu Fe -Co Fe-Cu Co-Cu Fe -Co Fe-Cu Co-Cu Fe -Co Fe-Cu Co-Cu

0.333 0.279 0.644 0.109 0.182 0.831 0.000 0.015 0.068 0.007 0.030 0.030

Discussion Accurate determination of the PGMs and Au is essential for geochemical and cosmochemical studies because these metals have significant economic values and can also provide important information about the origin, fractionation, and transportation of PGMs during geological processes (Qi et al., 2003). Generally, the determination of PGMs in geological materials is difficult because of their low crustal abundance (with background levels of a few nanograms per gram or less), heterogeneous distribution, and the complexity of sample preparation procedures. Accurate determination of these metals requires the use of highly sensitive analytical instruments. Because of its sensitivity and capability to measure traces, ICP-OES has been successfully employed to fully characterize PGMs and Au in geological materials and automotive catalysts (Senila et al., 2020). In addition, a variety

Fe/Co and Fe/Cu co-collection indicated a significant difference.

Table VIII summarizes the statistical analysis of variance for the certified reference material (control) and geological exploration samples. There is no significant difference in variance for the Pt results obtained from Fe and Co co-collectors for the control sample. The variances for the Pt results obtained from the Cu/Co and Cu/Fe co-collectors are statistically different. The variances of the Pd results obtained from all the co-collectors

Table VIII

Statistical analysis of variance Element

Variance 1

Variance 2

Certified reference material (CRM)

ρ value (significance level used was α = 0.05

Pt Fe Co Cu Co Fe Cu Cu Fe Co Pd Fe Co Cu Co Fe Co Cu Fe Co Rh Fe Co Cu Co Fe Cu Cu Fe Co Au Fe Co Cu Co Fe Cu Cu Fe Co 34

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0.137 0.000 0.137 0.000 0.000 0.000 0.703 0.933 0.703 0.895 0.933 0.895 0.201 0.001 0.201 0.073 0.001 0.073 0.167 0.050 0.167 0.281 0.005 0.281 VOLUME 122

Geological exploration sample ρ value (significance level used was α = 0.05

0.750 0.435 0.750 0.861 0.435 0.861 0.226 0.301 0.226 0.982 0.301 0.982 0.000 0.035 0.000 0.088 0.035 0.088 0.043 0.043 0.043 0.000 0.043 0.000 The Journal of the Southern African Institute of Mining and Metallurgy

A new preconcentration technique for the determination of PGMs and gold by fire assay and ICP-OES of techniques have been employed for sample decomposition and preconcentration of PGMs prior to ICP-OES measurement.

Nickel sulphide collection and sodium peroxide fusion followed

by Te co-precipitation are a common method for PGM analysis (Balcerzak, 2002). The nickel sulphide fire assay method offers the advantage of accommodating a large sample mass, and all of the PGMs can be concentrated by this procedure. However, the disadvantages are the relatively large amounts of reagents used, often resulting in an analytical blank with a higher concentration of PGMs, introduction of Cu and Ni to the solutions, which often cause interference problems with the PGMs, and the requirement that the composition of the flux be changed according to the composition of the sample matrix (Qi et al., 2003).

In this work, we developed a new low-cost technique to accurately determine PGMs and Au using Fe, Co, and Cu cocollectors during sample decomposition and preconcentration prior to ICP-OES. This technique is relatively simple, with higher PGM and Au recoveries. As the results demonstrate, better recoveries for PGMs and Au are obtained using Fe and Co as co-collectors. However, with Cu as a co-collector, the recoveries for Pt and Pd were lower than those obtained with Fe and Co collectors. Since co-collectors are used in the preconcentration of PGMs and Au, the decrease in recovery with Cu co-collector is likely due to losses of PGMs and Au into the slag. Generally, concentrates from UG-2 ores contain much higher levels of chromium, which results in the formation of chromium-rich spinels during melting (Nell, 2004). During the converting process care is needed to avoid excessively oxidizing conditions, which result in cobalt, nickel, and copper losses to the slag (Nell, 2004).

This could also be the reason for the lower recoveries using Cu as a co-collector.

Amassé (1998) successfully developed a method for determination of PGMs by employing selenium and tellurium as carriers in the presence of a potassium iodide catalyst. The extraction yields obtained using this method were between 95% and 100% for PGMs and around 80% for gold (Amassé, 1998).

Jin and Zhu (2000) determined Pt, Pd, Rh, and Au in geological samples by nickel sulphide (NiS) fire assay and co-collecting using Te before ICP-MS analysis, and showed that the recovery of Au was enhanced from 65%. to 80%. In the present study the Au recovery was above 90%, indicating that co-collection with lead fire assay can be a reliable technique for Au recovery. Wang and Brindle (2014) digested 0.5–2.0 g of the sample in a solution of 1% (m/v) L-cysteine and 1% HCl (v/v), for 10–12 minutes, and recovered between 94% and 100% gold according to the the ICP-MS results. In the current work, the recoveries from the Fe and Co co-collectors were in a similar range. This indicates that the method can be an alternative for the determination and quantification of Au. Zhang et al. (2014) determined Pt, Pd, Ru, Rh, and Ir in ultrabasic rock from the Great Dyke of Zimbabwe using lead fire assay laser ablation and ICP-OES. The recoveries of Pt, Pd, Ru, Rh, and Ir were 92.5%, 91.25%, 91.25%, 92.5%, and 93.75%, respectively. Consequently, the method of using Fe, Co, and Cu as co-collectors can be a suitable alternative for the analysis of PGMs and Au in large sample volumes.

Conclusion In this work we developed a new preconcentration technique for the determination of Pt, Pd, Rh, and Au in control and geological samples by the addition of Fe, Co, and Cu co-collectors. Fire assay was used for the preconcentration process. The samples were The Journal of the Southern African Institute of Mining and Metallurgy

digested in aqua regia and quantification was done using ICP-OES. The method was validated using the AMIS 0426 certified reference material. Based on percentage recoveries, we showed that Cu is not a good co-collector for Au and Rh determination, presumably due to loss of analyte in the slag, whereas Fe and Co are better cocollectors for the determination of PGMs and Au. Together, this study and the methodology developed for the determination of PGMs and Au can present a novel means to quantify PGMs and Au in control and geological exploration samples.

Acknowledgements The authors thank the staff of the Geology Department of the Siyanda Union Mine for their assistance with the sampling. We also thank the Siyanda Union laboratory for sample preparation and analysis by ICP-OES.

References Amassé, J. 1998. Determination of platinum-group elements and gold in geological matrices by inductively coupled plasma-mass spectrometry (ICP-MS) after separation with selenium and tellurium carriers. Geostandards Newsletter, vol.

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A new preconcentration technique for the determination of PGMs and gold by fire assay and ICP-OES Makua, L., Langa, K., Saguru, C., and Ndlovu, S. 2019. PGM recovery from a pregnant leach solution using solvent extraction and cloud point extraction: A preliminary comparison. Journal of the Southern African Institute of Mining and Metallurgy, vol. 119, no. 5. pp. 453–458.

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Tanner, D., Mcdonald, I., Harmer, R.E.J., Muir, D.D., and Hughes, H.S.R. 2019.

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Ndovorwi, F. 2014. Determination of platinum , palladium , rhodium and gold in ores and concentrates using iridium and ruthenium as co-collectors by fire assay. MSc thesis, University of Zimbabwe.

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Qi, L., Gregoire, D.C., Zhou, M.F., and Malpas, J. 2003. Determination of Pt, Pd, Ru and Ir in geological samples by ID-ICP-MS using sodium peroxide fusion and Te co-precipitation. Geochemical Journal, vol. 37, no. 5. pp. 557–565. https:// doi.org/10.2343/geochemj.37.557 Rodríguez-Rodríguez, Y. and Miguel, O. 2018. Determination of gold in geological samples combining the fire assay and ultraviolet visible spectrophotometry techniques. Academia Journal of Scientific Research, vol. 6, no. 1. pp. 27–33. doi: 10.15413/ajsr.2017.0112 Senila, M., Cadar, O., Senila, L., Böringer, S., Seaudeau-Pirouley, K., Ruiu, A., and Lacroix-Desmazes, P. 2020. Performance parameters of inductively coupled plasma optical emission spectrometry and graphite furnace atomic absorption spectrometry techniques for Pd and Pt determination in automotive catalysts. Materials, vol. 13, no. 22. pp. 1–13. https://doi.org/10.3390/ ma13225136 Serbin, R., Bazel, Y., and Ru, S. 2017. Speciation of platinum by GFAAS using various possibilities of analytical signal enhancement. Talanta, vol. 175 (June). pp. 46–52. https://doi.org/10.1016/j.talanta.2017.06.078

The Platinum conference series has covered a range of themes since its inception in 2004, and traditionally addresses the opportunities and challenges facing the platinum industry.

This prestigious event attracts key role players and industry leaders through: • High quality technical papers and presentations • Facilitating industry networking • Having large, knowledgeable audiences • Global participation, and • Comprehensive support from industry role players.

Vanhaecke, F., Resano, M., Koch, J., McIntosh, K., and Günther, D. 2010.

Femtosecond laser ablation-ICP-mass spectrometry analysis of a heavy metallic matrix: Determination of platinum group metals and gold in lead fireassay buttons as a case study. Journal of Analytical Atomic Spectrometry, vol. 25, no. 8. pp. 1259–1267. https://doi.org/10.1039/c002746d Volzhenin, A.V, Petrova, N.I., Medvedev, N.S., and Saprykin, A.I. 2018. Multiple probe concentrating of Au and Pd in geological samples for atomic absorption determination with two-stage probe atomization. Microchemical Journal, vol. 138. pp. 390–394. https://doi.org/10.1016/j.microc.2018.01.037 Wang, Y. and Brindle, I.D. 2014. Rapid high-performance sample digestion for ICP determination by ColdBlockTM digestion. : Part 2: gold determination in geological samples with memory e ff ect elimination. Journal of Analytical Atomic Spectrometry, vol. 29. pp. 1904–1911. https://doi.org/10.1039/c4ja00189c Zhang, N., Ma, Y., Shen, Y., and Gao, X. 2014. Determination of platinum , palladium , ruthenium , rhodium , and iridium in ultrabasic rock from the Great Dyke of Zimbabwe by inductively coupled plasma – optical emission spectrometry. Analytical Letters, vol. 47, no. 12 pp. 2072-2079. https://doi.org/10 .1080/00032719.2014.893441 u

The 8th International PGM Conference will, under the guidance of the organising committee, structure a programme which covers critical aspects of this continually evolving and exciting industry.

The success and relevance of this event to the industry really depends on your participation and support. You can participate in this event as an organising committee member, author/presenter, delegate or sponsor.

For further information contact: Camielah Jardine Head of conferencing | email: camielah@saimm.co.za | web:www.saimm.co.za

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NATIONAL & INTERNATIONAL ACTIVITIES 2022 23–24 February 2022 — Drill and Blast Hybrid Short Course 2022-Online via Zoom Wits Club, Johannesburg Contact: Camielah Jardine E-mail: camielah@saimm.co.za Website: http://www.saimm.co.za 27 February–3 March 2022 — TMS Furnace Tapping 2022 Anaheim, California, USA https://www.tms.org/AnnualMeeting/TMS2022/ Programming/Furnace_Tapping_2022/AnnualMeeting/ TMS2022/Programming/furnaceTapping.aspx?hkey=718f6af71852-445c-be82-596102913416 23 March 2022 — SANCOT Online Webinar: Introduction to ´Practical Guide to Rock Tunnelling´ with a special focus on relevant tunnelling case studies Contact: Gugu Charlie E-mail: gugu@saimm.co.za Website: http://www.saimm.co.za 20–27 May 2022 — ALTA 2022 Nickel-Cobalt-Copper, Uranium-Rare Earths, Gold-PM, In-Situ Recovery, Lithium & Battery Technology Online Conference & Exhibition Perth, Australia, Tel: +61 8 9389 1488 E-mail: alta@encanta.com.au Website: www.encanta.com.au 23–26 May 2022 — Mine Planning and Design Online School Contact: Gugu Charlie E-mail: gugu@saimm.co.za Website: http://www.saimm.co.za 14–16 June 2022 — Water | Managing for the Future Online Conference Vancouver, BC, Canada https://www.mineconferences.com Website: http://www.saimm.co.za 20–23 June 2022 — International Exhibition of Technologies and Innovations for the Mining and Energy Industry Chile https://www.aia.cl/ 21–24 June 2022 — Mine to Mill Reconciliation: Fundamentals and tools for productivity improvement Online Short Course Contact: Camielah Jardine E-mail: camielah@saimm.co.za Website: http://www.saimm.co.za 21–23 August 2022 — International Mineral Processing Congress Asia-Pacific 2022 (IMPC) Melbourne, Brisbane + online https://impc2022.com/

The Journal of the Southern African Institute of Mining and Metallurgy

21–25 August 2022 — XXXI International Mineral Processing Congress 2022 Melbourne, Australia + Online www.impc2022.com 24–25 August 2022 — Battery Materials Conference 2022 Misty Hills Conference Centre, Muldersdrift, Johannesburg, South Africa Contact: Gugu Charlie E-mail: gugu@saimm.co.za Website: http://www.saimm.co.za 8–14 September 2022 — 32nd Society of Mining Professors Annual Meeting and Conference 2022 (SOMP) Windhoek Country Club & Resort, Windhoek, Namibia Contact: Camielah Jardine E-mail: camielah@saimm.co.za Website: http://www.saimm.co.za 15–20 September 2022 — Sustainable Development in the Minerals Industry 2022 10th Internationl Hybrid Conference (SDIMI) Windhoek Country Club & Resort, Windhoek, Namibia Contact: Gugu Charlie E-mail: gugu@saimm.co.za Website: http://www.saimm.co.za 28–29 September 2022 — Thermodynamic from Nanoscale to Operational Scale (THANOS) International Hybrid Conference 2022 on Enhanced use of Thermodynamic Data in Pyrometallurgy Teaching and Research Mintek, Randburg, South Africa Contact: Camielah Jardine E-mail: camielah@saimm.co.za Website: http://www.saimm.co.za 24–26 October 2022 — 8th Sulphur and Sulphuric Acid Conference 2022 The Vineyard Hotel, Newlands, Cape Town, South Africa Contact: Gugu Charlie E-mail: gugu@saimm.co.za Website: http://www.saimm.co.za 2–4 November 2022 — PGM The 8th International Conference 2022 Sun City, Rustenburg, South Africa Contact: Camielah Jardine E-mail: camielah@saimm.co.za Website: http://www.saimm.co.za 13–17 November 2022 — Copper 2022 Santiago, Chile https://copper2022.cl/

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Company affiliates The following organizations have been admitted to the Institute as Company Affiliates 3M South Africa (Pty) Limited AECOM SA (Pty) Ltd AEL Mining Services Limited African Pegmatite (Pty) Ltd Air Liquide (Pty) Ltd Alexander Proudfoot Africa (Pty) Ltd Allied Furnace Consultants AMEC Foster Wheeler AMIRA International Africa (Pty) Ltd ANDRITZ Delkor (Pty) Ltd Anglo Operations Proprietary Limited Anglogold Ashanti Ltd Arcus Gibb (Pty) Ltd ASPASA Aurecon South Africa (Pty) Ltd Aveng Engineering Aveng Mining Shafts and Underground Axiom Chemlab Supplies (Pty) Ltd Axis House Pty Ltd Bafokeng Rasimone Platinum Mine Barloworld Equipment -Mining BASF Holdings SA (Pty) Ltd BCL Limited Becker Mining (Pty) Ltd BedRock Mining Support Pty Ltd BHP Billiton Energy Coal SA Ltd Blue Cube Systems (Pty) Ltd Bluhm Burton Engineering Pty Ltd Bond Equipment (Pty) Ltd Bouygues Travaux Publics Castle Lead Works CDM Group CGG Services SA Coalmin Process Technologies CC Concor Opencast Mining Concor Technicrete Council for Geoscience Library CRONIMET Mining Processing SA Pty Ltd CSIR Natural Resources and the Environment (NRE) Data Mine SA Digby Wells and Associates DRA Mineral Projects (Pty) Ltd DTP Mining - Bouygues Construction Duraset Elbroc Mining Products (Pty) Ltd eThekwini Municipality Ex Mente Technologies (Pty) Ltd Expectra 2004 (Pty) Ltd

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Exxaro Coal (Pty) Ltd Exxaro Resources Limited Filtaquip (Pty) Ltd FLSmidth Minerals (Pty) Ltd Fluor Daniel SA ( Pty) Ltd Franki Africa (Pty) Ltd-JHB Fraser Alexander (Pty) Ltd G H H Mining Machines (Pty) Ltd Geobrugg Southern Africa (Pty) Ltd Glencore Gravitas Minerals (Pty) Ltd Hall Core Drilling (Pty) Ltd Hatch (Pty) Ltd Herrenknecht AG HPE Hydro Power Equipment (Pty) Ltd Huawei Technologies Africa (Pty) Ltd Immersive Technologies IMS Engineering (Pty) Ltd Ingwenya Mineral Processing (Pty) Ltd Ivanhoe Mines SA Joy Global Inc.(Africa) Kudumane Manganese Resources Leica Geosystems (Pty) Ltd Longyear South Africa (Pty) Ltd Lull Storm Trading (Pty) Ltd Maccaferri SA (Pty) Ltd Magnetech (Pty) Ltd Magotteaux (Pty) Ltd Malvern Panalytical (Pty) Ltd Maptek (Pty) Ltd Maxam Dantex (Pty) Ltd MBE Minerals SA Pty Ltd MCC Contracts (Pty) Ltd MD Mineral Technologies SA (Pty) Ltd MDM Technical Africa (Pty) Ltd Metalock Engineering RSA (Pty)Ltd Metorex Limited Metso Minerals (South Africa) Pty Ltd Micromine Africa (Pty) Ltd MineARC South Africa (Pty) Ltd Minerals Council of South Africa Minerals Operations Executive (Pty) Ltd MineRP Holding (Pty) Ltd Mining Projections Concepts Mintek MIP Process Technologies (Pty) Limited MLB Investment CC Modular Mining Systems Africa (Pty) Ltd

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MSA Group (Pty) Ltd Multotec (Pty) Ltd Murray and Roberts Cementation Nalco Africa (Pty) Ltd Namakwa Sands(Pty) Ltd Ncamiso Trading (Pty) Ltd New Concept Mining (Pty) Limited Northam Platinum Ltd - Zondereinde Opermin Operational Excellence OPTRON (Pty) Ltd Paterson & Cooke Consulting Engineers (Pty) Ltd Perkinelmer Polysius A Division of Thyssenkrupp Industrial Sol Precious Metals Refiners Rams Mining Technologies Rand Refinery Limited Redpath Mining (South Africa) (Pty) Ltd Rocbolt Technologies Rosond (Pty) Ltd Royal Bafokeng Platinum Roytec Global (Pty) Ltd RungePincockMinarco Limited Rustenburg Platinum Mines Limited Salene Mining (Pty) Ltd Sandvik Mining and Construction Delmas (Pty) Ltd Sandvik Mining and Construction RSA(Pty) Ltd SANIRE Schauenburg (Pty) Ltd Sebilo Resources (Pty) Ltd SENET (Pty) Ltd Senmin International (Pty) Ltd SISA Inspection (Pty) Ltd Smec South Africa Sound Mining Solution (Pty) Ltd SRK Consulting SA (Pty) Ltd Time Mining and Processing (Pty) Ltd Timrite Pty Ltd Tomra (Pty) Ltd Traka Africa (Pty) Ltd Ukwazi Mining Solutions (Pty) Ltd Umgeni Water Webber Wentzel Weir Minerals Africa Welding Alloys South Africa Worley

The Journal of the Southern African Institute of Mining and Metallurgy

2 DAY HYBRID CONFERENCE

THANOS PROJECT

THERMODYNAMICS FROM NANOSCALE TO OPERATIONAL SCALE

INTERNATIONAL CONFERENCE

ON ENHANCED USE OF THERMODYNAMIC DATA IN PYROMETALLURGY TEACHING AND RESEARCH 28-29 SEPTEMBER 2022 - CONFERENCE VENUE - JOHANNESBURG (MINTEK)

BACKGROUND Fundamental knowledge of thermodynamic principles and data is important in understanding and improving processes used in the production of metals as well as in the design and development of new processes. This is particularly so given the fact that the production of metals from ores and/or secondary resources using pyrometallurgical processes involve complex thermochemical phenomena as a result of high temperatures and application of energy to materials. In most cases, however, pyrometallurgists do not fully appreciate the immense potential of thermodynamics to the design and operation industrial processes. This trend is worrying so as the engineering society is moving towards competencies focusing on a wide area of knowledge. The shift towards “Wikipedia knowledge” is a natural consequence of availability of huge amounts of information, but invariably, tends to occur at the expense of fundamental knowledge which forms the backbone of high quality thermodynamics teaching and research. In some instances, students and researchers tend to regurgitate derivations of thermodynamic equations with no indication of how such thermodynamic principles and data are to be put to practical use. To keep the interest and the dedication to the teaching, learning and application of thermodynamics principles and data, new teaching methods must continuously be developed with emphasis on how the fundamental knowledge is used in the research, design and operation of pyrometallurgical processes.

CONFERENCE OBJECTIVES The broad objective of the International Conference on enhanced use of Thermodynamic Data in Pyrometallurgy Teaching and Research is to enhance the use of thermodynamic data in pyrometallurgy teaching and research. The ultimate goal is to increase competitiveness of the South African pyrometallurgical industry by demystifying thermodynamics and equipping the industry to use thermodynamic principles and data in metal production. Hosted by the Metallurgy Technical Programme Committee of the Southern African Institute of Mining and Metallurgy, this conference will focus on two main pillars: (a) the enhanced use of thermodynamic tools and data and the understanding the fundamental reaction mechanisms in metal production, and (b) developing methods for teaching thermodynamics and enhancing the teaching and learning and availability of thermodynamic methods and data. The project is funded by the Research Council of Norway through the Programme for International Partnerships (INTPART) under the project “Thermodynamic from Nanoscale to Operational Scale” (THANOS).

FOR FURTHER INFORMATION, CONTACT: Camielah Jardine, Head of Conferencing

E-mail: camielah@saimm.co.za Tel: +27 11 834-1273/7 Web: www.saimm.co.za

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