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▲

ii APRIL 2014 The Journal of The Southern African Institute of Mining and Metallurgy

OFFICE BEARERS AND COUNCIL FOR THE2013/2014 SESSION

Honorary President

Mark CutifaniPresident, Chamber of Mines of South Africa

Honorary Vice-Presidents

Susan ShabanguMinister of Mineral Resources, South AfricaRob DaviesMinister of Trade and Industry, South AfricaDerek HanekomMinister of Science and Technology, South Africa

PresidentM. Dworzanowski

President Elect

J.L. Porter

Vice-Presidents

R.T. JonesC. Musingwini

Immediate Past PresidentG.L. Smith

Honorary Treasurer

J.L. Porter

Ordinary Members on Council

H. Bartlett S. NdlovuN.G.C. Blackham G. NjowaV.G. Duke S. RupprechtM.F. Handley A.G. SmithW. Joughin M.H. SolomonA.S. Macfarlane D. TudorD.D. Munro D.J. van Niekerk

Past Presidents Serving on Council

N.A. Barcza R.P. Mohring R.D. Beck J.C. Ngoma J.A. Cruise R.G.B. Pickering J.R. Dixon S.J. Ramokgopa F.M.G. Egerton M.H. Rogers A.M. Garbers-Craig J.N. van der MerweG.V.R. Landman W.H. van Niekerk

Branch ChairmenDRC S. MalebaJohannesburg I. AshmoleNamibia G. OckhuizenPretoria N. NaudeWestern Cape T. OjumuZambia H. ZimbaZimbabwe S.A. GaihaiZululand C. Mienie

Corresponding Members of Council

Australia: I.J. Corrans, R.J. Dippenaar, A. Croll, C. Workman-Davies

Austria: H. Wagner

Botswana: S.D. Williams

Brazil: F.M.C. da Cruz Vieira

China: R. Oppermann

United Kingdom: J.J.L. Cilliers, N.A. Barcza, H. Potgieter

USA: J-M.M. Rendu, P.C. Pistorius

Zambia: J.A. van Huyssteen

The Southern African Institute of Mining and Metallurgy

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)

Honorary Legal AdvisersVan Hulsteyns Attorneys

AuditorsMessrs R.H. Kitching

Secretaries

The Southern African Institute of Mining and MetallurgyFifth Floor, Chamber of Mines Building5 Hollard Street, Johannesburg 2001P.O. Box 61127, Marshalltown 2107Telephone (011) 834-1273/7Fax (011) 838-5923 or (011) 833-8156E-mail: [email protected]

ContentsJournal Commentby H. Phillips . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . vPresident’s Corner by M. Dworzanowski . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . vii

Special ArticlesBook review—Digging Deepby N. Mayer . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . viiiMQA gives Wits University over R20 millionby K. Foss . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . ixManager: Regional Development . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . xThe SAIMM Library . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 346

PAPERSA critical evaluation of haul truck tyre performance and management system at Rössing Uranium Mineby T.S. Kagogo . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 293A comparative study between shuttle cars and battery haulersby W.H. Holtzhausen . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 299Investigation of cavity formation in lump coal in the context of underground coal gasificationby C. Hsu, P.T. Davies, N.J. Wagner, and S. Kauchali . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 305Arnot’s readiness to prevent a Pike River disasterby R. Weber. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 311Laser cladding AA2014 with a Al-Cu-Si compound for increased wear resistanceby K.J. Kruger and M. du Toit . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 317Development of a method for evaluating raw materials for use in iron ore sinter in terms of lime assimilationby W. Ferreira, R. Cromarty, and J. de Villiers . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 325The recovery of manganese products from ferromanganese slag using a hydrometallurgical routeby S.J. Baumgartner and D.R. Groot . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 331Wear of magnesia-chrome refractory bricks as a function of matte temperatureby M. Lange, A.M. Garbers-Craig, and R. Cromarty. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 341Comparing the extent of the dissolution of copper-cobalt ores from the DRC Region using sulphuric acid in the presence of hydrogen peroxide and in tartaric acidby S. Stuurman, S. Ndlovu, and V. Sibanda . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 347

International Advisory Board

R. Dimitrakopoulos, McGill University, CanadaD. Dreisinger, University of British Columbia, CanadaE. Esterhuizen, NIOSH Research Organization, USAH. Mitri, McGill University, CanadaM.J. Nicol, Murdoch University, AustraliaH. Potgieter, Manchester Metropolitan University, United KingdomE. Topal, Curtin University, Australia

The Journal of The Southern African Institute of Mining and Metallurgy APRIL 2014


▲iii

Editorial BoardR.D. BeckJ. Beukes

P. den HoedM. Dworzanowski

M.F. HandleyR.T. Jones

W.C. JoughinJ.A. LuckmannC. MusingwiniR.E. Robinson

T.R. Stacey

Editorial ConsultantD. Tudor

Typeset and Published byThe Southern African Instituteof Mining and MetallurgyP.O. Box 61127Marshalltown 2107Telephone (011) 834-1273/7Fax (011) 838-5923E-mail: [email protected]

Printed by Camera Press, Johannesburg

AdvertisingRepresentativeBarbara SpenceAvenue AdvertisingTelephone (011) 463-7940E-mail: [email protected] SecretariatThe Southern AfricanInstitute of Mining andMetallurgyISSN 2225-6253

THE INSTITUTE, AS A BODY, ISNOT RESPONSIBLE FOR THESTATEMENTS AND OPINIONSADVANCED IN ANY OF ITSPUBLICATIONS.Copyright© 1978 by The Southern AfricanInstitute of Mining and Metallurgy. Allrights reserved. Multiple copying of thecontents of this publication or partsthereof without permission is in breach ofcopyright, but permission is hereby givenfor the copying of titles and abstracts ofpapers and names of authors. Permissionto copy illustrations and short extractsfrom the text of individual contributions isusually given upon written application tothe Institute, provided that the source (andwhere appropriate, the copyright) isacknowledged. Apart from any fair dealingfor the purposes of review or criticismunder The Copyright Act no. 98, 1978,Section 12, of the Republic of SouthAfrica, a single copy of an article may besupplied by a library for the purposes ofresearch or private study. No part of thispublication may be reproduced, stored ina retrieval system, or transmitted in anyform or by any means without the priorpermission of the publishers. Multiplecopying of the contents of the publicationwithout permission is always illegal.

U.S. Copyright Law applicable to users Inthe U.S.A.The appearance of the statement ofcopyright at the bottom of the first page ofan article appearing in this journalindicates that the copyright holderconsents to the making of copies of thearticle for personal or internal use. Thisconsent is given on condition that thecopier pays the stated fee for each copy ofa paper beyond that permitted by Section107 or 108 of the U.S. Copyright Law. Thefee is to be paid through the CopyrightClearance Center, Inc., Operations Center,P.O. Box 765, Schenectady, New York12301, U.S.A. This consent does notextend to other kinds of copying, such ascopying for general distribution, foradvertising or promotional purposes, forcreating new collective works, or forresale.

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The Journal of The Southern African Institute of Mining and Metallurgy APRIL 2014 ▲v

The papers in this edition of the Journal are authored or co-authored by recent graduates in mining and metallurgy.They are based on final year undergraduate projects and werepresented at the annual Southern African Institute of Miningand Metallurgy’s Student Colloquium in November 2013. Thiswas held at the University of Johannesburg, and for the firsttime a student from Namibia presented a paper. While thetone of the presentations ranged from outrageouslyflamboyant to virtual stage fright, it was abundantly clear thatthe research being undertaken by these students was of aremarkably high standard. The diversity of topics, togetherwith the differences in approach by the different disciplines,was catered for by parallel sessions but delegates wereunanimous in their praise of the quality of the presentations.The choice as to which ten papers should be published in theJournal must have been extremely difficult.

Rather than attempting to summarize the papers orcomment on them individually, I would prefer to turn to thevexing question of mining research in South Africa. I excludemetallurgy, metallurgical, and chemical engineering not dueto any prejudice, but due to my ignorance of those subjectsand the feeling I have that research in these areas has faredbetter in recent years than has mining research. This feelingis reinforced by the very recent news that the CSIR hasdecided to disaggregate (their word, not mine) their Centre forMining Innovation and to reassign its remaining researchersto areas where similar competencies exist but which servicemany different industrial sectors and clients. This terminates50 years of mining research on the Auckland Park site, sinceit was in 1964 that the Transvaal and Orange Free StateChamber of Mines formed the Chamber of Mines ResearchOrganisation (COMRO) and established it in Carlow Road.

The need for a mining research organization wasrecognized following the inquiry into the Coalbrook disaster,which found there was no scientific basis for the design ofcoal pillars and highlighted the need for systematic research.Three existing laboratories (Dust and Ventilation, AppliedPhysiology, and Biological and Chemical Research) wereincorporated into the newly established COMRO and both aMining Research Laboratory and a Physical SciencesLaboratory were created. An Environmental Services Divisionwas added and in 1966 a Colliery Research Laboratorycompleted the Organization. During the 1970s and 1980sSouth Africa was undoubtedly the world leader in mostaspects of mining research and the research output wasprodigious. When I had the privilege of spending six monthsof sabbatical leave working at COMRO in 1981 there werenearly 700 employees, with three quarters of them beingactive researchers. The early 1990s saw structural changesand a reduction in size and in 1993 COMRO was taken overby the CSIR and became Miningtek. The staffing complimentwas significantly reduced over the years and Miningtek

eventually lost its status as a division of CSIR, and SouthAfrica lost its status as the leading country for miningresearch. Only the coal mining sector has sustained coherentresearch activities, through its collaborative researchprogramme, Coaltech.

It must be acknowledged that individual miningcompanies do undertake and sponsor research but thisconfidential research is often piecemeal, with contractsawarded worldwide to institutions and individuals who areexperts in particular research areas. The result of this hasbeen a dearth of opportunities for young people in SouthAfrica to develop their skills in mining research under thementorship of experienced researchers.

Over the past 15 years or so the focus of most miningcompanies has been on meeting the targets for transformationand sustainable development. Health and safety of theworkforce, care for the environment, and engagement withcommunities are undoubtedly a vital part of mining. However,the need to develop new mining methods and technology tomine safely and economically requires hard-core miningengineering research. Since the completion of majorprogrammes such as Deep-mine and Future-mine little hasbeen done to consolidate the findings of these programmesand put them into practice. The future of many mines is tomine deeper and this, coupled with a significant change in thecost of employment of the underground workforce, cries outfor efficient and effective mechanization, as the forerunner ofautomation of many mining operations. At present there is noorganization or institution commanding sufficient respectfrom the mining industry to be the leader or custodian of thenecessary research. Should it be a government departmentthat initiates a revival of mining engineering research?Should it be the Chamber of Mines, or the CSIR, or indeed aconsortium of universities? Time will tell, but time is alsorunning out and the store of knowledge from previousresearch is dissipating fast.

What lends optimism to the current situation is the qualityof the research undertaken by the students whose workappears in this edition of the journal. The two mining papers,three mineral processing papers, and three material sciencepapers are of a high standard and give a clear indication ofthe talent available to our industry. Let’s hope these youngpeople find careers that utilize their talents in meeting thechallenges faced in the mining and beneficiation of ourmineral resources.

H. Phillips

Journal Comment

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The Journal of The Southern African Institute of Mining and Metallurgy APRIL 2014 ▲vii

One of the key drivers for a successful South African miningindustry is a pipeline of mining and metallurgy graduates.Without these individuals the sustainability of the industry will

be in jeopardy. Therefore, a key aspect of SAIMM activities is thesupport of mining and metallurgy students and the tertiary institutionsthat provide their education. While there is some financial support

available via the SAIMM scholarship trust fund, the type of support thatthe SAIMM provides is to motivate students with their studies and improve the

probability of them being employed within the industry.The annual SAIMM Student Colloquium has proved to be a great success in bringing students

together from different tertiary institutions. The presentations given on the mining and metallurgyprojects being undertaken clearly show the quality of students being graduated. The 2013 subject matterlooked at different commodities, with some of the work conducted directly with a mining operation. Thetopics, which included different aspects of ore transport for underground and open pit operations,material wear properties, sintering for iron production, hydrometallurgical treatment of ferromanganeseslag, wear rate of refractory bricks in furnaces, and hydrometallurgical treatment of copper/cobalt ore,highlight the variety of work being conducted and its clear applicability to the mining industry. TheSAIMM Student Colloquium has started to broaden the involvement of students from our southernAfrican branches, which is an important initiative going forward. Setting a date for the colloquium thatcan encompass as many of the tertiary institutions as possible has always been a challenge and willcontinue to be, but our organizing committee always manages to get the best possible results each yearand all credit to them.

There is a tendency to forget that the SAIMM Western Cape branch encompasses the tertiaryinstitutions in the Western Cape which generate many high-quality metallurgy graduates. They also hostan annual event, a two day Mineral Processing conference, which highlights the significant amount ofwork which the undergraduates and postgraduates there produce. Last year I was asked by the Branchto give a keynote address at this annual conference, which allowed me to witness the significantcontribution these tertiary institutions make towards metallurgy research and the output of metallurgygraduates.

The SAIMM has also started organizing career days for mining and metallurgy students where thestudents have the opportunity to ask a panel of mining engineers and metallurgists questions aboutdifferent career paths within the mining and metallurgy disciplines. The first event, held at Sci Bono lastyear, proved to be a huge success. I was on the panel, so I can testify to this.

After students complete their studies and start a career in the mining industry, they become YoungProfessionals. This year the SAIMM hosted the first Young Professionals conference, where experienceswere shared across different commodities and disciplines by quite a diverse group of young miningengineers and metallurgists. This is another example of support to students as they leave their studiesand embark upon a career in the mining industry. Guidance given early in their careers will increase theprobability of them staying with a career in the mining industry.

The above points illustrate the extent to which the SAIMM supports mining and metallurgy studentsand their progression to Young Professionals. I cannot overemphasise how important this support is, andhow essential it is for the SAIMM to examine ways to expand this support.

M. DworzanowskiPresident, SAIMM

Presidentʼs

Corner

▲

viii APRIL 2014 The Journal of The Southern African Institute of Mining and Metallurgy

Jade Davenport, the author of Digging Deep, has written a very readable book onthe historic contribution that mining has made in developing South Africa into a

modern industrial state. It is not necessarily a must-read for people in the miningindustry, but the book is well written and should be seen as a South African miningbiography.

Digging Deep can be regarded as an anthology, each component of which can beread in isolation, but with a thread that links the parts into the complete story of thevital role that the mining industry played in this country’s economic and politicalhistory. I would, however, recommend that the chapters be read in sequence.

The book will be of interest to a broad spectrum of South African readers, as itis not a book on the technical aspects of the industry. There is, however, sufficientdetail that will allow mining professionals to find it a satisfying read. The history isclearly well researched and is detailed enough to help the reader get a goodperspective of the social, political, and economic impact on the development ofSouth Africa. In essence, it should have general appeal.

Although replete with the names of key players in the history of mining anddetails of production and costs, the book remains an easy read. However, theterminology may be unfamiliar to the general reader, and footnotes on the relevant pages would have been helpful.

The author has tried to use the relevant terms in the historical context of currency of the day and imperial/metricsystems, but has deviated occasionally in parts of the book. In an early chapter, an attempt is made to equate the valueof money then in use to the current rand.

The author is to be commended for writing the book in such a way that each chapter’s history is complete, yet linkedalong the historic path to the other chapters.

Digging Deep, by Jade Davenport, is published by Jonathan Ball.

The reviewer

Nap Mayer was involved in the mining industry from 1960 to 2000 with experience in copper, diamonds, coal, and gold.As Managing Director of Anglo American’s Gold and Uranium Division, he served on the Anglo American Board as anAlternate Director. He also served on the Atomic Energy Board as Vice- Chairman for a three-year period from 1996.

Nap is a retired, corporate member of the SAIMM.

N. Mayer

Book Review—Digging Deep

The Journal of The Southern African Institute of Mining and Metallurgy APRIL 2014 ▲ix

17 March 2014 - Johannesburg: The MiningQualifications Authority (MQA) handed over acheque for more than R20 million to Wits Universityon Friday, 14 March 2014. The money will gotowards support for seven lecturers in miningengineering and bursaries for 236 students in thefollowing disciplines: analytical, chemical, electrical,industrial, mechanical, metallurgical, and miningengineering, and geology.

The Head of the School of Mining Engineering,Professor Fred Cawood, said the long-standingpartnership between Wits and the MQA dated back to2005 and had strengthened to the point where it wasvalued at such a significant sum of money. ‘Thiscommitment speaks volumes about the MQA andsets an example for other SETAs’, he said.

Vice-Chancellor and Principal of Wits, ProfessorAdam Habib, echoed Cawood’s opinion that the MQAhad set an example for other industries, andreminded those present that the historicaldisenfranchisement of some South Africans hadcreated enormous levels of inequality that could onlybe addressed through collective action.

‘The VC can no longer say that hisresponsibilities end at the gates of the university.The CEO can no longer say that his responsibilitiesend with the company’s shareholders. How we beginto bridge institutional boundaries has becomeimportant. This partnership with the MQA istestimony to what can be done’, said Habib.

The total amount of the partnership is R23 592113.03. The total amount of support for lecturers is R4 624 113.03, and the total amount to be given inbursaries is R18 868 000.00.

Habib said the bursaries, to be given todisadvantaged students, would send a powerfulmessage of hope to the poor that talented peoplehave access to one of the best universities in thecountry, and that the support that would be given tolecturers was an investment in the creation of a newblack professoriate.

R100 000 will be used to support students whosestudies are being affected because they can’t affordnecessities such as spectacles. The MQA alsosupports the kitchen project, which feeds studentswho can’t afford lunch. Cawood said the School hadseen an almost 99% success rate in students whohad been assisted in this manner.

The CFO of the MQA, Yunus Omar, told studentand staff representatives from Wits that the MQAwas comprised of people who had been in theirshoes. ‘They know what the students and lecturersare going through’, he said.

K. Foss

Prof Frederick Cawood, head of the Wits School of MiningEngineering, Mr Yunus Omar, CFO of the MQA and Vice-chancellor and principle of Wits University, Prof Adam Habib

MQA gives Wits University over R20 million

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x APRIL 2014 The Journal of The Southern African Institute of Mining and Metallurgy

The Southern African Institute of Mining and Metallurgy (SAIMM)

MANAGER: REGIONAL DEVELOPMENT

Applications are invited for the newly created position of Manager: Regional Development with the Southern AfricanInstitute of Mining and Metallurgy. The primary purpose of this job is to sustainably expand non-South Africanmembership of the SAIMM by establishing and supporting National Branches in neighbouring countries.

Secondary activities are related to representation of the SAIMM on appropriate industry-level forums that supportmembership growth and Institute sustainability.

Requirements of Position

The incumbent will be responsible for, amongst other duties, the development and execution of a regional marketingand growth strategy that is aligned with SAIMM strategy and objectives; the development and support of new andexisting National Branches across Southern Africa; provision of direct strategic and tactical support to the NationalBranches; co-ordination of SAIMM support across National Branches to achieve an optimal Institute outcome; as wellas the development and implementation of National Branch events calendars that build stakeholder engagement andensure Branch sustainability. Other duties will include the establishment of annual operating plans and budgets withina three-year strategic plan for each National Branch.

The person appointed must be able to engage at a senior level in the minerals industry, and to communicate ontechnical matters in forums representing the Professional membership. Generic marketing skills around monitoringand analysing market trends, as well as studying competitors' products and services, are some of the marketing skillsrequired.

Other qualities include the ability to respond well to pressure; creative thinking; good presentation skills; and theability to motivate and lead teams.

Significant travel is required to build and maintain the National Branches.

Experience

The successful applicant should have a minerals industry background with 10+ years minerals industry experience;broad knowledge of the global mining industry from resource to customer; and be able to demonstrate peopleleadership experience i.e. has held a substantive position and/or project management experience.

It is recommended, but not a requirement, that the incumbent is a current Fellow/Member of the SAIMM or otherprofessional entity. Tertiary qualifications related to marketing and the minerals industry would be beneficial.

Enquiries: Sam Moolla; Tel: +27 11 834-1273/7; Email: [email protected] apply: Please submit a letter of motivation and a detailed CV with the names, contact numbers,

and e-mail addresses of three references.

Closing date: 30 June 2014.

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These papers will be available on the SAIMM websitehttp://www.saimm.co.za

Student PapersA critical evaluation of haul truck tyre performance and management system at Rössing Uranium Mineby T.S. Kagogo . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 293The factors affecting haul truck tyre performance and the effectiveness of the management system in the load-and-hauloperation are investigated with the aim of increasing tyre performance in terms of running time to failure

A comparative study between shuttle cars and battery haulersby W.H. Holtzhausen. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 299Two different underground batch coal haulers are compared to determine the more viable machine to implement. Running costs, capital costs, and maintenance costs are compared, and production rates and availability are used to determine which type of machine would be more reliable.

Investigation of cavity formation in lump coal in the context of underground coal gasificationby C. Hsu, P.T. Davies, N.J. Wagner, and S. Kauchali . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 305This project investigated, on a laboratory scale, cavity formation within a coal block due to the combustion reactions that take place during the coal gasification process in an underground environment. The results are compared with earlier work and a similar trend was observed, despite a slightly different methodology being employed.

Arnot’s readiness to prevent a Pike River disasterby R. Weber . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 311The effectiveness of Arnot Colliery’s measures to prevent conditions favourable for methane explosions is evaluated, with particular reference to the causes of the Pike River disaster in New Zealand.

Laser cladding AA2014 with a Al-Cu-Si compound for increased wear resistanceby K.J. Kruger and M. du Toit . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 317Aluminium alloy 2014 (AA2014) was coated with a 1.5 mm thick laser-deposited layer composed of silicon, copper, and aluminium with the aim of increasing the alloy’s wear resistance. It is shown that the Al-Cu system is very sensitive to silicon additions, and that wear resistance depends on solidification of the primary phase to as well as on the final phase distribution.

Development of a method for evaluating raw materials for use in iron ore sinter in terms of lime assimilationby W. Ferreira, R. Cromarty, and J. de Villiers . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 325A new test method is proposed that allows automatic evaluation of iron ores in terms of lime assimilation with increasing temperature. The new method, termed the Length Reducibility Test, is shown to be superior to the standard method in terms of precision and reproducibility of results, as well as ease of implementation.

The recovery of manganese products from ferromanganese slag using a hydrometallurgical routeby S.J. Baumgartner and D.R. Groot. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 331The recovery of manganese metal and other manganese products from ferromanganese slag by means of leaching, precipitation, and electrowinning is investigated. The various methods are compared in terms of selectivity, costs, and product quality.

Wear of magnesia-chrome refractory bricks as a function of matte temperatureby M. Lange, A.M. Garbers-Craig, and R. Cromarty . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 341The postulation that primary platinum group metal (PGM) matte will chemically react with magnesia-chrome bricks attemperatures above 1500°C was tested. Phase relations observed clearly indicate that chemical reactions take place between matte and the magnesia-chrome refractory under these conditions, and that these reactions are more complex than expected.Comparing the extent of the dissolution of copper-cobalt ores from the DRC Region using sulphuric acid in the presence of hydrogen peroxide and in tartaric acidby S. Stuurman, S. Ndlovu, and V. Sibanda . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 347A copper-cobalt oxide ore from the Central African Copperbelt was leached in two different environments, and the effects of acid and reducing agent concentration and leaching temperature determined. The results indicate the potential of tartaric acid to extract cobalt, rather than copper, from copper-cobalt ores..

Overview

Rössing Uranium Mine is an open pit miningoperation on the west coast of Namibia.Currently, there is a global off-road tyreshortage, which has a negative impact on largeoperations such as Rössing. According toCutler (2012), it has been estimated that mosttyre suppliers have about 25–30%undersupply in the market for the past 3 years.The demand has surpassed supply, as shownin Figure 1, and the tyre shortage crisis hasincreased tyre prices by as much as 425%since 2009.

Rössing currently has 32 KomatsuHaulpak 730E 2000 HP haul trucks which runon six tyres per truck, thus there are 192 tyresin operation at any given time. The downwardtrend in tyre life – 7371 hours in 2012compared to 10 119 hours in 2009 (Figure 2)has a negative impact on operations atRössing. Optimizing the tyre life is therefore anecessity in order to ensure that the operationdoes not run out of tyres, as obtaining tyres inthe current market is not easy.

According to Simulilo (2012), tyreperformance at Rössing over the past two

years has been below target, as illustrated inFigure 2. The mine has experienced prematurefailures in 49% of the tyres, with cutseparation being the most common cause. Theoutcomes of the project will help the minereduce the number of premature failures andincrease the average tyre life.

Rössing is currently under taking a cost-cutting exercise in the wake of poor productionand low uranium prices coupled withincreasing costs (Aluvilu, 2012). The projecthas the potential to assist in this, in line withthe company’s objective, which is to cut costswithout retrenching employees. Everyadditional hour of tyre life is money saved onthe procurement of new tyres.

Methodology and objectives

The objectives and methodology of the projectare summarized in Table I.

The methodology for the project has beenguided by the recommendations of Carter(2007) on areas of awareness for tyreimprovement as shown in Table II.

Results and analysis

Performance review

The operation uses two main brands of tyres;namely Michelin and Bridgestone, but doesuse other brands such as Goodyear andBelshina when the supply of the two mainbrands is short. The operation is currentlyrunning 35 Komatsu 730E 180 t dump trucks,which use tyres sizes of 37.00R57 (Michelin)and 42/90R57 (Bridgestone).

Figure 3 shows the frequency graph for thetyres in terms of the hours run to failure.Ideally the tyre life should fall in the

A critical evaluation of haul truck tyreperformance and management system atRössing Uranium Mineby T.S. Kagogo*Paper written on project work carried out in partial fulfilment of B. Eng. (Mining Engineering)

SynopsisThe factors affecting haul truck tyre performance and theeffectiveness of the management system at Rössing Uranium Minewere investigated with the aim of increasing tyre performance interms of running hours until failure. The main objectives were toidentify the types of tyre failure, their causes and cost implications,and evaluate the effectiveness of the management system. A siteseverity survey, weight study, and TKPH studies were conducted todetermine the pit conditions, and an analysis of failed tyres carriedout. The results showed that tyre performance at the mine hasdeclined from 2009 to date, and the increase in lost value amountedto R5.7 million in 2012 alone. The main cause of tyre failure is looserocks in the pit. The present management system in the load andhaul department is not effective enough due to operationalconstraints it is facing.

Keywordstyre failure, cost implications, types of failure, causes of failure,management system, TKPH, haul roads.

* Department of Mining Engineering, University ofPretoria, Pretoria, South Africa.

© The Southern African Institute of Mining andMetallurgy, 2014. ISSN 2225-6253. Paper receivedJan. 2014.

293The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 APRIL 2014 ▲

A critical evaluation of haul truck tyre performance and management system

10 000–12 000 hour interval to maximize tyre performance.However, in 2009 only 48% of tyres reached that interval,and only 18.7% and 2.5 % in 2011 and 2012, respectively. In2011 and 2012 no tyres reached a life of more than 12 000hours. In 2012 and 2011 most of the failures occurredbetween 6000 and 10 000 hours. 2009 has been chosen as abaseline year to compare 2011 and 2012 results because it isthe best performing year over a 7-year period.

Types of tyre failure

Figure 4 summarizes the different types of tyre failureexperienced at Rössing. Any of these modes of failure canoccur at any time in the tyre’s lifespan except worn out whichusually occurs at the back end of the service life of the tire.The biggest challenge the mine is facing is the fact that thereis no system to identify specific areas in the pit where aparticular tyre fails, hence the ‘hot spots’ where most tyrefailures occur cannot be identified. This situation has madeprioritizing specific areas in the pit for more attention andascertaining the exact causes of tyre failure a hard task.

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294 APRIL 2014 VOLUME 114 The Journal of The Southern African Institute of Mining and Metallurgy

Figure 1—Global tyre supply and demand (Cutler, 2012)

Figure 2—Tyre performance over the past 7 years

Table I

Objectives and methodologies

Objective Methodology

Conduct a market analysis on tyres. Literature review

Do a performance review • Data collection• Data analysis

Determine types of tyre • Data collectionfailure at Rössing • Data analysis

Identify and investigate possible • Road severity surveycauses of tyre failure • Loading area investigation

• Tipping area severity survey• Weight study• TKPH study• Interviews with the necessary

personnel• Observation of general pit

conditions

Determine cost implications • Data collectionof tyre failure • Data analysis

Evaluate the effectiveness of the • Operator questionnaireRössing tyre management system • Interview with necessary

personnel• Data collection and analysis

Identify areas of awareness Analysis and recommendations

Table II

Areas of awareness for tyre life improvement (Carter, 2007)

Awareness area Possible actions to Improve awareness area

Driver awareness • Make operators aware of the supplysituation

• Solicit input on areas of improvement • Provide incentives for improvements

Haul road design • Super-elevation in corners (if supersaren’t possible, reduce speed)

• Identify and remove soft spots in roads • Optimal road crown is 3%

Air pressure maintenance • Conduct regular pressure checks, withimmediate pressure corrections

• Daily preferred, weekly necessity • Install new O-rings and hardware when

mounting • Inspect/change/repair cracked wheels

and components • Inquire with dealer about temperature/

pressure monitoring • Analyse air pressure documents just like

any other

Mechanical maintenance • Check alignment • Check suspension components • Use ‘rock knockers’ • Rectify problems immediately

Tyre and rim inspection • Driver walk-around (train drivers what tolook for)

• Rim inspection for cracks or flangedamage

• Inspect valve hardware

TKPH management • Total GVW adherence (no overloading isacceptable)

• Adhere to speed limit

Support equipment • Proper and effective use of graders andrubber-tired dozers

• Equipment should be assigned toshovels

• Driver radio communication of spills androad damage

• Fix problem areas immediately

Analyse scrap tyres • Analyse history of scrap tires • List types of damage • List vehicles with multiple tire failures • Examine shift performance (individual

crews with problems; night vs. day)

Establish performance • Involve cross-section of mine incommittee joint efforts

• Plan consistent meeting schedule • Make assignments for change; follow up

for corrections

Communicate and report • Issue consistent, visible reports ofefforts

• Issue consistent, visible reports ofprogress

• Solicit suggestions.

Most of the tyres failed from wear, which is the idealsituation as tyres usually have a lot of hours on them at thetime of failure. In 2009 the fleet was a bit smaller than atpresent, but the fleet was the same for 2011 and 2012, whichshows that more tyres were lost in 2012 than in 2011,indicating a drop in tyre performance as the operating truckhours were virtually the same for both years.

Figure 5 shows the average tyre life for each failure modefor 2009, 2011, and 2012. The mine uses the operate-to-failure system whereby a tyre is replaced with a new tyreonly once it has failed. This system does not pose any safetyhazard as the radial tyres only deflate at failure and do notburst. Worn-out failure mode is the ideal type of tyre failureas it carries more hours and maximizes tyre life more thanthe other types of tyre failure. Operational failure andproduct-related failure remain the modes of failure with thelowest average hours across all three years, and much shouldbe done to limit those modes of failure.

Possible causes of failure

Driving over rocks is the general cause of most tyre failuresat Rössing. There is no reliable system to locate where thefailures take place in the pit, as tyres do not necessary failimmediately on coming into contact with the rock but mightfail a few hours or days later. A few areas have beenidentified as causing specific modes of tyre failure, and theresults from the observations and measurements arediscussed in detail below.

Tons-kilometres per hour (TKPH)

A tons-kilometres per hour (TKPH) study was done atRössing Uranium Mine to ascertain whether the mine is

using the right tyres for the site conditions. TKPH causes tyrefailure through heat separation and increased wear rate. TheTKPH study was done only on the Michelin tyre and not theBridgestone, but it was advised that the results for one brandare significant enough to determine the site conditions.

Results (real site TKPH):

➤ Front= 706➤ Rear= 661.

The tyre in use is a 37.00R57 XDR B4 with a TKPHrating of 848, which means that the tyres on site are the righttyres; therefore the premature tyre failures are not due to theTKPH rating being exceeded. The results of the TKPH studyare supported by the results of the weight study, whichyielded the loading field data used in calculating the TKPH(TKPH = Average tyre load x Average truck speed).

Road corners

Road corners can be a source of tyre damage but this was notcovered in the road severity survey, and an observationexercise was conducted to assess the conditions.

Figure 6 illustrates loose rocks on the shoulder of a bermat a traffic circle and in a turn, which are a source of sidewallcuts. Figure 7 shows partially buried rocks in a turn, whichare also a source of sidewall cuts and are not easily visible.The turning radius of the corners is within the minestandards, but the main concern is the loose rocks in thecorners as the mine lost six tyres in 2012 due to sidewallcuts. The corners need to be constantly dressed with sandand loose rocks removed, and operators should take care notto cut corners and expose the tyres to rocks.

Loading and tipping areas

Loading and tipping areas are among the areas that have thehighest potential to cause tyre damage as they are a source ofloose rocks that trucks can drive over. Rössing has two main



Figure 3—Frequency graph for average tyre hours

Figure 4—Failure modes

Figure 5—Average hours per failure

Figure 6—An example of loose rocks in a turn, with a truck about to cuta corner


loading areas; namely the stockpiles and the blastedmuckpiles in the pit. The two loading areas are serviced byshovels and front-end loaders with auxiliary equipment suchas dozers for clean-up.

Figure 8 shows a typical loading area condition atRössing with loose rocks. Loose rocks are inevitable in theloading area, but should be cleaned up regularly and truckoperators should not reverse into a loading area that containsloose rocks. Due to operational constraints, auxiliary clean-upequipment is not always available but effort should be putinto keeping the area clean. The shovel operator should alsopractice a culture of cleaning the area regularly when theclean-up dozers are not available.

Haul roads

A site severity survey was done to determine the pitconditions, including the haul road.

The severity ratings are from 1 to 5, with 5 being the bestcondition and 1 the worst. Not all the categories are of signif-icance to tyre failure, such as water and road width. Thecategories of high significance for tyre performance arespillage, gradient, banking, aggregates, undulations, andhammering, which all play a role in how the tyre fails and itswear rate.

Spillage has a rating of 2, which means that around 50%of the haul road on average is covered with spillage and therock size is in excess of 75 mm, which is large enough tocause tyre failure by modes such as rock cut penetration.Aggregates also have a rating of 2 and pose the same threatto tyres as spillage does. The gradient is rated 2 (about 10%both uphill and downhill), which can cause heat build-up inthe tyre leading to failure through heat separation.

Banking and hammering are of less concern as they haveratings of 3, which is good in the current state as tyres aresubjected to shock less than 10% of the time and less than50% of the roads have banking. Undulation is the biggestconcern, with a rating of 1 due to the presence of undulationsevery 3 m. Undulations on the haul road causes payloadspillages as well as a slight heat build-up in the tyre. Figure 9shows an undulating portion of an in-pit haul road, clearlyillustrating the severity of the situation.

Figure 10 shows an example of spillage on the in-pit haulroad, which also is a major problem as it is a source of looserocks that cause operational failures such as cut penetration.

Cost Implications

The cost implication of premature tyre failure is based onlyon the remaining value in the planned tyre life, whichconstitutes a direct cost incurred. There are other added costs

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Figure 7—Sidewall cut rocks in a turn

Figure 8—Loading area with loose rocks

Table III

Overview of in-pit haul road severity survey

Category Category Rating Commentcode

CL493 Spillage 2 Spillage over 50% of the haul road, size in excess of 75 mm

CL494 Gradient 2 Gradient over 10% either uphill/downhill

CL495 Banking 3 Less than 50% of roads have banking

CL496 Aggregates 2 Aggregates cover 50% of the width of the haul road (over 25 mm)

CL497 Firmness 3 Sinkage is not in main working areas and is not affecting the tyres

CL498 Water 3 Standing water does not cause damage to the tyres

CL499 Undulations 1 Undulations every 3 m

CL500 Hammering 3 Tyres are subjected to shockdamage less than 10% of the time

CL501 Road width 2 Equal to twice the width of a single truck

Figure 9—Undulation of in-pit haul roads at Rössing

incurred on top of the direct cost, such as the labour costinvolved in changing tyres more frequently. A single tyre cantake up to 8 hours to change, depending on which position itis in. A further indirect cost is the lost production timeincurred by tyre changes.

Figure 11 shows the remaining tyre value at failure foreach type of tyre failure for 2011 and 2012. The highest costwas incurred due to rock cut penetration, which at R2 millionis a 233.3% increase from R807 000 in 2011. Impact fractureis the other type of failure with increased lost tyre value from2011 to 2012. Value lost due to failure by wear decreasedfrom 2011 to 2012. No tyre value was lost in 2011 due toheat separation, while some value was lost due to heatseparation in 2012. The highest costs are incurred throughfailure of tyres that have run the lowest number of hourswhich generally happens through premature failure.

The cost per hour is obtained by dividing the value of anew tyre by the life in hours. In 2011 cut separation incurredthe highest cost of R47.03 per hour, while failure due to wearhad the lowest cost of R15.24 per hour. In 2012 impactfracture had the highest cost per hour with R47.50 per hour,while wear was the lowest with R18.27 per hour. The costper hour results indicate that cut separation was the mostexpensive failure mode in 2011, while in 2012 impactfracture was the most expensive.

The high cost per hour of premature failure is due to thelow hours on the tyre at time of failure, while tyres that faildue to wear usually have had a long service, hence the lowcost per hour. The average cost per hour for 2012 for all thefailed tyres, irrespective of the failure mode, is R30.95 perhour, with the operational failures increasing the cost. Theaverage cost is still well above the target cost of R24.00 perhour which is guided by the price of the new tyre and thetarget lifespan.

Load and haul tyre management system

In the past, when the tyres were performing well, the loadand haul management system had a bonus incentive fortyres, which engendered a positive attitude among theoperators towards tyre preservation. Management also hadmore resources when it came to maintaining the roads, asDust Aside was used more, which kept the haul roads inexcellent condition. The load and haul tyre management isnot highly effective due to different constraints that the

department is facing. According to the superintendent of loadand haul at Rössing (Fotolela, 2012) the department is facingconstraints such as lack of operators and high productionpressure, which leads to tyres getting minimum priority.

Challenges facing the load and haul tyre managementinclude:

➤ Passive management➤ No road maintenance programme➤ No specific tyre preservation programme➤ Low utilization of auxiliary equipment➤ No system to identify point of failure in pit➤ No record keeping of road maintenance➤ Priority given to production at the expense of tyres.

According to Fotolela (2012) the department is workingon addressing the challenges as tyre preservation has beenidentified as a high-priority area for 2013 in an effort toimprove tyre performance.

Conclusion

The mining industry has been facing a tyre shortage for thepast 8 years, with a peak in 2008, and suppliers are currentlyrunning about 25–30% undersupply in the market for thepast 3 years, hence the shortage is not expected to decreaseuntil 2018. The mine experienced an upward trend in tyre lifefrom 2006 to 2009 in terms of average hours per tyre, with2009 having the highest average tyre life of 10 119 hours,while a downward trend has been evident since then with7371 hours in 2012. In 2012 operational failure was themajor failure category, accounting for 49% of all failures, andhad the lowest average hours. The biggest concerns in termsof possible tyre failures are the conditions of the haul roadscondition and loading areas, as they are the major sources ofloose rocks that cause premature failure. The mine lost R5.7million due to tyre failure in 2012, which is an increase fromR4 million in 2011. The operational failure category made upthe bulk of the tyre value lost, accounting for 73% of thetotal. The maintenance side of the tyre management systemis very effective, as tyre maintenance practice, storage, andmonitoring are within the Rio Tinto standards. The load andhaul side of the tyre management system is not as effective,as no proper road maintenance record is kept, and effectiveutilization of clean-up dozers is low. This is due partly to theconstraints, such as a shortage of operators, that thedepartment is facing.


The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 APRIL 2014 297 ▲

Figure 10—Example of spillage on the in-pit haul road

Figure 11—Lost tyre value for 2011–2012


Recommendations

1. A tyre campaign should be adopted to promote collectiveresponsibility and the spillage clean-up policy by allpersonnel reinstated. The campaign will not involveadditional resources and can be very beneficial in terms ofworker morale

2. A more aggressive approach should be taken to pitcondition maintenance where resources permit and toovercoming the operational constraints that the auxiliaryclean-up crews are facing in order to increase theireffective utilization. This will be a challenge as theoperation is facing an operator shortage, but effortsshould be made to work with the available resources

3. A system is needed to identify the exact areas in the pitwhere tyre failure is common and classify them as ‘redzones’ that need priority in terms of auxiliary equipment.

This will not be an easy system to come up with as mosttyre failures do not immediately follow the cause.Aggressive road maintenance can help in this regard

4. The feasibility of re-introducing the tyre incentiveprogramme should be investigated.

ReferencesALUVILU, P. 2012. Production Engineer, Rössing Uranium Mine. Personal

communication.

CARTER, R. 2007. Maximizing mining tyre life. Engineering and Mining Journal,(00958948), vol. 208, no. 6. July/August. 58 p.

CUTLER, T. 2012. EM tyre supply shortage – How it is affecting us’. AusIMMTechnical Meeting, 14 May.

FOTOLELA, D. 2013. Superintendent: Load & Haul, Rössing Uranium Mine.Personal communication

SUMULILO, F. 2012. Maintenance Engineer, Rössing Uranium Mine. Personalcommunication. ◆

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Introduction and mine background

During studies at the University of Pretoriaand field research at Exxaro’s Arnot coal mine,data was collected that could serve as aguideline for the selection of underground coalhauling machinery and what the presenttrends are related to these machines.

Arnot Colliery is wholly owned by ExxaroResources, the largest BEE (Black EconomicEmpowerment) contributing mining house inSouth Africa, and is situated in the WitbankCoal fields in the Highveld region. At presentthe mine is solely an underground operation,exploiting the No. 2 Lower seam with anaverage calorific value of 23.8 MJ/kg and ashcontent of 23%. Production is approximately2.5 Mt/a and is supplied to Eskom’s Arnotpower station on a cost-plus agreement(Exxaro, 2013).

Arnot employs a mechanized bord andpillar method with continuous coal-cuttingminers. At present there are two interlinkedshafts, 8 Shaft servicing five sections and 10 Shaft four sections. Currently there are foursections at 8 Shaft that utilize battery haulers,

with one using shuttle cars and one planningto change to shuttle cars. At 10 Shaft on theother hand, only shuttle cars are used.

Introduction to bord and pillar coalhaulage

In a typical underground mechanized bord andpillar mine section a Continuous Miner (CM) isused to cut the coal from the production facewith a rooftogether bolting machine thatinstalls permanent cemented roofbolts. Coal isloaded via the CM’s chain conveyor systemonto the batch hauling machines, either shuttlecars or battery haulers, which then transportthe coal load to an in-section crusher or feederbreaker where the coal is offloaded andcrushed to a more tolerable size. From thefeeder breaker, coal is transported out of themine via a series of conveyor belt systems.

Background of the project

Batch hauling systems in an underground coalmine are the most unreliable link in the chainof ore transport. The implications of costs,productivity, reliability, and safety of thesemachines was scrutinized by the managementof Arnot and it was found necessary toinvestigate these aspects further.

During discussions, different opinionsarose related to these machines and it was feltthat these criteria need to be fully investigatedand documented.

In particular, a critical comparison betweenbattery haulers and shuttle cars was requiredto determine the more feasible piece ofequipment to implement in terms of cost,safety, reliability, and productivity.

A comparative study between shuttle carsand battery haulersby W.H. Holtzhausen*Paper written on project work carried out in partial fulfilment of B. Eng. (Mining Engineering)

SynopsisThe purpose of this project was to compare two underground, batchcoal haulers – battery haulers and shuttle cars – in order todetermine the more viable machine to implement.The specific standards for battery haulers were investigated andcompared to the requirements of shuttle cars in order to identify theunnecessary expenses related to the legal requirements that areattached to the machines.Costs such as running costs, capital costs, and maintenance costswere researched and compared over a typical life of machine.Average production rates and breakdown times were obtained andused to determine which machine would be more reliable inachieving the required annual production.

Keywordscoal, bord and pillar, hauling, cost (running, capital, totalownership), productivity, reliability, safety, availability,constraints.

* Department of Mining Engineering, University ofPretoria.



A comparative study between shuttle cars and battery haulers

Objectives and methodologies

Battery haulers vs. shuttle cars

The major difference between the two systems is that batteryhaulers are battery powered and shuttle cars are cablepowered. This difference in itself entails advantages anddisadvantages.

Other factors are the coal load bearing methods, turningmechanisms, and flexibility in terms of use and transport.

Battery haulers are much more flexible as they can travelany route in order to load and offload coal, but this can entailsome drawbacks such as increased travelling distance anddecreased battery life. This is not the case with shuttle carssince they are confined to travelling a specific route due tothe trailing cable that supplies power to the drives. This,however, forces the section to do frequent section moves,where the section equipment is moved closer to the workingface.

Figure 1 depicts a typical battery hauler showing thearticulation joint. This joint is very useful for cornering andmanoeuverability in restricted conditions. The coal loadingportion is confined to the rear of the machine and offloadingis done with a hydraulic push-off system. This hydraulicmechanism is disadvantageous because it incurs spillages atthe feeder breaker and therefore side plates have to beattached to the feeder.

With the shuttle, car on the other hand, offloading can be synchronized with the feeder breaker since both usesimilar chain conveyor systems for loading and offloading(Figure 2). This reduces spillage and maintains a more

constant feed rate to the feeder breaker. Unlike batteryhaulers, there is no articulation, but steering is via all fourwheels, allowing the vehicle to corner easier.


Current trends

In the current South African coal mining industry Joy’sStamler battery haulers are the front runners in the marketand are used at most of the mines that use battery haulers.Joy is also the lead contracting company on the mine,supplying support and operations services to both the batteryhaulers and the shuttle cars.

According to Stewart (2013) in the past 5 years Joy sold15 battery haulers and 126 shuttle cars. During that time Joyhad a backlog of three battery haulers and almost 100 shuttlecars. This is a clear indication that the industry is movingaway from the use of battery haulers and is more prone tobuying shuttle cars.

The decrease in the use of battery haulers is resulting in adrop in the skills available for operation and maintenance, aswell as increased cost of such skills.

A key aspect to consider is the standardization of themine fleet in order to ensure better focus and skills for acertain machine, effectively increasing operational life andperformance.

Health and safety aspects

Incident studies have shown that shuttle car operators have areduced field of vision when the machine is loaded. In somecases operators lean out of the cab in order to see clearly, and

▲


Figure 1—Battery hauler (Joy Global, 2013)

Figure 2—Shuttle car (Joy Global, 2013)

Objectives Methodologies

Determine all factors that influence Collect information from differentthe advantages and disadvantages sources i.e. internet, management of battery haulers and shuttle cars and operators. Use questionnaires

as assistance

Determine safety and health issues Locate and consult articles about of both machines and evaluate them. the health and safety issues

involvedwith battery haulers and shuttle.

Determine what the industry trend Contact manufacturers andis and what the future outlook is request current and future salesfor the use of these machines profiles

Determine costs such as running, Acquire information from maintenance, capital and others suppliers as well as minethat have a major influence on the employees on the costsuse of battery haulers and shuttle involved with the haulers.cars

Obtain production figures of all the Enquire production figures fromproduction sections for a certain the surveying departmenttime period and relate them to the of both 8 Shaft and 10 Shaftbatch haulers from that particular section

Determine and compare the Obtain downtime studies from thereliability and availability of engineering department and the machines calculate the relevant figures

according to a specific standard

Evaluate and analyse the results Study, in detail, all the resultsthat were obtained and draw conclusions on findings. Calculate costs, production figures, availabilities and other relevant information

this could result in injuries or fatalities. Another contributorto injuries is the cables that accompany the shuttle cars.These are tripping hazards, and when the haulers are inmotion or are cornering these cables can come under tension,and when this tension is released injury, damage, or fatalitiesmay also occur (Bezuidenhout, 2011)

The major safety and environmental concern of thebattery haulers is the fact that they are powered by lead-acidtype batteries that produce hydrogen gases and might incurleakage of these gases and the acid-based electrolyte. Thiscan lead to environmental contamination and exposure ofworkers to chemical-related injuries. The gases might alsopose an explosion risk and this is of great concern inunderground coal mines. (Van der Merwe, 2013)

Costs

Costs are one of the most important factors in thepresent–mining environment day. Several cost categorieshave to be considered. These include procurement, operatingor life–cycle costs, and the total ownership cost (TOC).

Using the 2014 projected procurement prices that weresupplied by Joy Global it was calculated that the difference ininitial capital for the two machines was approximately 35%,with battery haulers being the more expensive. This excludesthe additional requirement of installing battery bays as wellas the ancillary ventilation and safety equipment required forthe battery bay.

The cost difference between the major overhauling andreplacement bodies is 15% and 6% respectively, with batteryhaulers again being the more expensive.

Considering that Joy requires these machines to receive aminor overhaul on a 1.0–1.2 million ton period, and a majoroverhaul every 1.6–1.8 million ROM tons, the life–cycle costscan be calculated using Figure 3. The average total life–cyclecost of the battery hauler per ton hauled is R 5.89 more thanthat of the shuttle car.

With the projected annual requirement of 190 000 tonsper month per machine, the average operating costs would beR3.89 per ton and R2.98 per ton for battery haulers andshuttle cars respectively. Figure 3 shows the life-cycle costs.

The total machine life is approximately 2.4 Mt, whichwith the 190 000 ton per annum production target equates toapproximately 12 years. Using this knowledge, a 10%interest rate, and a 2013/14 electricity cost of 65.51 cents perkilowatt-hour (Eskom, 2013) the TOC was calculated,making the assumption that these machines are operating for18 hours a day and 26 days per month. Battery haulers havea total of 187 kW of motor power and shuttle cars 219 kW.

Over this 12-year period the battery haulers will costapproximately R3.6 million more to operate moving the samevolume of coal.

Productivity

The productivity of the different sections over a six-monthperiod was considered to try to identify whether shuttle car or battery hauler sections have a higher productivity. Severalother factors could also contribute to differences in produc-tivity, such as the fact that battery haulers on average carrymore tons.

After seam height corrections had been applied, it seenthat battery hauler sections produce on average 26 kt/monthwhereas shuttle car sections produce only an average of 21kt/month. This is a difference of 5000 t/month per miningsection. Also, battery haulers deviate from their monthlytargets by about -20% and shuttle cars -30%, clearlyindicating that there is a possibility that battery haulersections are more productive.

Reliability

Downtime data for all the machines was tabulated over a 1-year period to obtain an average engineering availability forthe two types of machines (Figure 4).

An average of 96.7% and 96.2% availability is achievedfor battery haulers and shuttle cars respectively with astandard deviation of 2% for shuttle cars and 1.3% forbattery haulers, indicating that both machines are very closein terms of reliability. Battery haulers had an average of 191hours per annum downtime and shuttle cars 165 hours.

Figure 5 shows the specific downtimes for the haulers.

➤ Battery haulers:– Average mechanical downtime is 28.65% of the

total– Electrical downtime is 54.9%– Hydraulic downtime totals 16.5%



Figure 3—Life-cycle costs

Figure 4—Engineering availability. Blue - battery haulers, grey - shuttlecars


➤ Shuttle cars:– Mechanical downtime is 48.1% of the total– Electrical downtime is 45.3%– Hydraulic downtime is 6.6%.

Despite the belief that battery haulers are simpler in termsof the electrical circuitry, they experience higher relativedowntimes due to electrical problems compared to shuttlecars. This may be due mainly to the Lionetics upgrades thatare done on the machines to transform the DC current fromthe batteries into an AC output. This must be done accordingto SANS 1654. Also, battery haulers suffer more hydraulicbreakdowns than shuttle cars and this raises concerns aboutthe efficacy of the articulation joint.

The relative mechanical downtimes of shuttle cars arehigher, and this is because some of these machines arenearly 30 years old. However, they still show very highavailabilities, similar to those of the battery haulers.

Another factor that needs to be considered is thatoperators, artisans, and technicians have much moreexperience with shuttle cars than on battery haulers.

Constraints

Another factor contributing to machine selection is compati-bility with other equipment such as feeder breakers andcontinuous miners. Through calculations and time studies itwas found that the average cycle time for battery shuttle is168 seconds and that of the shuttle cars is 190 seconds.Assuming that battery hauler payload is 18 t and the shuttlecar payload 16 t with 21 cycles per hour for battery haulersand 18 cycles per hour for shuttle cars, that battery haulershave a 378 t/h capacity and shuttle cars only 288 t/h.

With a maximum of three haulers in a section themaximum capacity in a section is 1134 t/h for battery haulersand 864 t/h for shuttle cars. Comparing these figures to the capacities of the feeder breaker (770 t/h) and the CM(840 t/h) shows that there will be some waiting time at theCM for both machines, but because of the increased flexibilityof the battery haulers they will thus be underutilized. Thisclearly shows that shuttle cars are more than capable oftransporting the required tons if they are maintained properlyand that battery haulers will not be used to their full capacity.(Figure 6).

Conclusions

Table I is a representation, according to level of significance,of which machine has superior performance in terms of thevarious criteria.

➤ In terms of safety and environmental concerns theshuttle car is a much better option

➤ Shuttle cars incur fewer costs in both running andprocuring

➤ Battery haulers offer better productivity➤ Both machines are very reliable➤ Shuttle cars are more adaptable and compatible with

the other equipment in the transport chain➤ Battery haulers are prone to be much more flexible➤ The coal mining industry seems to shy away from the

use of battery haulers.

Recommendations

According to the results obtained from this study, it isrecommended that the mine moves to the use of shuttle carsin order to reduce annual running costs, and also to reducethe funding required for battery bays, battery bay personnel,and refurbishing of batteries. In doing this the mine will alsostandardize the fleet and simplify ordering of spares andequipment.

However, use of battery haulers should not necessarily becompletely eliminated. In some cases, such as poor in-sectionand low seam conditions, battery hauler are the preferredchoice because of the better flexibility and manoeverability.Thus further study into the effects that bad section and seamconditions have on the haulers is required.

▲


Figure 5—Average downtimes of shuttle cars and battery haulers

Figure 6—Evaluation of capacities

Table I

Conclusive comparison

Criteria Weighting Shuttle car Product Battery hauler Product

Safety 10 8 80 6 60

Cost 9 8 72 6 54

Productivity 8 7 56 8 64

Reliabilty 7 9 63 9 63

Constraints 4 9 36 7 28

Flexibility 3 5 15 9 27

Trend 2 9 18 3 6

Total 340 302

Acknowledgements

➤ My mentor and supervisor, Mr Nico van der Merwe, forhis guidance and support throughout the course of thestudy

➤ The team at Exxaro Arnot for assistance with theresearch and the grant of this learning opportunity.

➤ The University of Pretoria’s Mining EngineeringDepartment for their guidance

➤ Special thanks to Wilma De Jager at Arnot for herassistance in collecting and organizing of the data.

References

ANGLO AMERICAN. 2013. Operations. www.angloamerican.co.za [Accessed 7Aug. 2013].

BEZUIDENHOUT, L.J.A. 2011. Operational Report Mpumalanga Region February2011. Department: Mineral Resources, Witbank.

Eskom. 2013. Tariffs and Charges 2013/14. www.eskom.co.za [Accessed 11Jul. 2013].

EXXARO RESOURCES. 2013. Arnot http://www.exxaro.com/content/ops/coal_arnot.asp [Accessed 17 Jan. 2013].

JOY GLOBAL. 2013. Haulage Systems. Product Overview. http://www.joy.com[Accessed 18 Jan. 2013].

KENTUCKY COAL EDUCATION. 2013. Glossary of Mining Terms. www.coaledu-cation.org [Accessed 25 Feb. 2013].

KLINKERT, D and MARAIS, S. 2012. Equipment Replacement. Mobile EquipmentReport. Arnot Coal.

LEES, D. 2009. Slope instability. Exxaro Arnot. Code of PracticeARNOT_SMCOP_09/1. Arnot Reliability Centre.

MTHETHWA, D. 2012. Battery Charging Bay Procedure. AR-BCB 01. Arnot CoalMine. 12 December 2012.

MYORS, A. 1998. Introduction of battery powered coal haulers into board andpillar panel production. Coal Operators’ Conference, University ofWollongong.

O’DONNELL, C. and SCHIEMANN, M. 2008. Hydrogen gas management for floodedlead acid batteries. Battcon Stationary Battery Conference 2008. MesaTechnical Assosiates, Inc.

SAMBO, V. 2013. Personal Communications.

STEWART, D.N. 2013. Equipment Sales trend at Joy for 5 Years. Personalcommunication.

VAN DER MERWE, N.J. 2013. Personal communications.

NKOSI, S. 2007. South African Coal Industry – challenges and opportunities.Coaltrans Conference, Rome, 21-24 October 2007.

UNIVERSAL COAL. 2012. Coal Mining In South Africa. http://www.universalcoal.com/projects/coal-mining-in-south-africa. [Accessed 8 Jul. 2013]. ◆



Introduction

Underground coal gasification (UCG) createsan opportunity to access coal that is deemedun-mineable due to its depth, the low qualityof the coal, and unsafe mining conditions. Inaddition, it is considered to be more environ-mentally friendly than conventional miningand the process is less capital- and labour-intensive (Self, Reddy, and Rosen, 2012). Theproduced syngas can be used by a host oftechnologies, but it is used mainly for theproduction of liquid fuels, power generation,industrial heating, or fertilizers (Seddon andClarke, 2011).

UCG is an in situ gasification process.There are many methods for the gasification ofcoal seams and the extraction of syngas; theutilization of an injection and production wellis investigated in this report. The injection wellallows for the feed flow of air/oxygen/oxygen-enriched air and steam, and the producedsyngas is extracted to the surface via theproduction well. However, a linkage is requiredbetween the two wells as the coal seams arenot sufficiently permeable to allow for thedispersion of air/steam across the horizontallength. There are several methods for creatingthis linkage path, namely: reverse combustionlinking (RCL), forward combustion linking(FCL), hydro-fracking, electro-linking, and in-seam linking (Abdul, Aqeel, and Gholamreza,2013). The experiment was based on in situgasification using the injection and productionwell with the FCL method.

FCL involves the use of directional drilling,which offers the least damage to the existingground structure (Directional Drill Pty Ltd,2011) and improves the feasibility, design,and operation of a UCG plant.

The purpose of this research was toinvestigate cavity formation in a coal block inthe context of UCG. The work of Daggupati etal. (2010) was used as a guide for the experi-mental investigation and X-ray tomographywas used to confirm and analyse the cavityformation.

Materials and methods

Experimental setup

The experimental setup was based on that ofDaggupati et al. (2010), with the main

Investigation of cavity formation in lump coalin the context of underground coalgasification

by C. Hsu*, P.T. Davies*, N.J. Wagner*, and S. Kauchali*Paper written on project work carried out in partial fulfilment of BSc. Eng (Mining)

SynopsisUnderground coal gasification (UCG) is becoming more popular asthe reserves of good quality, mineable coal are starting to diminish,and yet the global energy demand from coal is still increasing. Thepurpose of this research project was to investigate cavity formationwithin a coal block due to the combustion reactions in the context ofUCG. The cavity plays a pivotal role in the UCG process, as it isessentially the gasification reactor. Cavity formation in an in situgasification process using the forward combustion linking method(FCL) had been investigated, and a laboratory model was created tosimulate the process.

The experiment was performed by drilling a U-shaped tunnelinto a coal block, which was then combusted internally with air thatwas fed through an injection hole. A heating element (at approxi-mately 500°C) was used to supply the required heat for combustionat the base of the injection well. The coal blocks were analysedusing micro-focus X-ray tomography.

The tomography results showed that the coal tended to crackalong the bedding plane after a short duration of combustion, due toeither the formation of clinker or the expansion of swelling vitrinitealong the horizontal tunnel. The deposit was thicker at the base ofthe injection well compared to the base of the production well; thismay have been caused by the turbulence of the air flow and therelatively high oxygen concentration at the base of the injectionwell. A comparison of the results with work by Daggupati et al.(2010) showed the same trend, despite the slightly differentmethodology applied.

Keywordsunderground coal gasification, cavity formation, gasificationreactor, forward combustion linking method (FCL), clinker,vitrinite, micro-focus X-ray tomography.

* University of the Witwatersrand, Johannesburg,South Africa.



Investigation of cavity formation in lump coal in the context of underground coal gasification

difference being the use of a heating element as the heatsource in place of the LPG used by Daggupati et al. A U-shaped tunnel was drilled into a 15 cm × 11 cm × 12 cm coalblock; the vertical injection and production wells were 5 mmin diameter and the horizontal tunnel connecting the two was8 mm diameter. Stainless steel pipes were inserted into thevertical wells, one of which was connected to a gas rotameterand the air cylinder, and the other allowed the venting of theproduct gases into the fume hood. A Type K thermocouplewas used inserted into the coal block to monitor thetemperature throughout the experiment. Cutile refractorybricks were clamped together and surrounded the experimentin order to minimize heat loss and to keep the high-temperature zone separated from the external surroundings.The heat source was provided by a heating element made of 2m of Kanthal D wire with 8.25 Ω/m resistance; the wire waswound into a spiral around a 3 mm rod. This wire wasconnected to a Variac (POWERSTAT® VariableAutotransformer) to control the voltage applied to theelement and hence the temperature. The experimental setupis shown in Figure 1.

Experimental procedure

An inlet air flow was set and maintained at 25 mL/s; this wasfed into the injection well and monitored by a rotameter. Theheating element was switched on using the Variac and thetemperature was measured using a thermocouple. Theexperiment was allowed to run for 3 hours and the resultingcoal block was analysed using the Nikon XTH 225 ST micro-focus X-ray tomography system at South African NuclearEnergy Corporation (NECSA) (Hoffman and De Beer, 2013).This produced 3D images of the coal samples and allowed themapping of the cavity based on density differences.

Results and discussion

A total of three experiments were performed, two of whichyielded successful results, and which are referred to asExperiment 1 and Experiment 2.

Coal characterization

The characteristics of the coal that was used are summarizedin Table I. The ash content was18.8%, and from the maceralanalysis it was found that the coal was composed of 80.8%

vitrinite by volume. It is important to note that all of theanalyses were done using crushed coal samples, whereas theexperimental work utilized large blocks of coal. However, theanalysis can be assumed to be representative of the coalblock as a whole.

Experiment 1

Expansion due to either the formation of clinker or theexpansion of vitrinite caused a large crack to form along thelength of the coal block during the first experiment. Clinker isa porous substance that forms during the combustion of coalthat has a high ash content, non-combustible minerals, and alow gross calorific value (Mitra, 2011). The clinker is createdby the fusion of non-combustible minerals (such as iron,calcium, and sodium) upon exposure to high temperatures(Afri Coal Investments Pty Ltd, 2013). Vitrinite is one of theorganic substances found within coal and it is known to swellupon heating. Figures 2 and 3 show areas of darker greywhich represent regions with a lower density material thancoal. The low-density region is indicative of either clinker orplasticized vitrinite. For the purpose of this report, theseregions will be referred to as ‘clinker’ as the material has notyet been analysed.

More clinker formed around the inlet gas well (the leftwell) than the production well (right well); this is also shownin Figures 4 and Figure 5, where the shape of the clinkerzone is shown. A higher concentration of oxygen wasavailable at the inlet well, and the extent of the combustionreactions as well as the temperature was therefore higher

▲


Figure 1—Diagrammatic representation of experimental setup (side view)

Table I

Coal characteristics

Component Standard Composition (%)

Carbon ASTM D 5373 71.05Hydrogen ASTM D 5373 4.29Nitrogen ASTM D 5373 1.78Oxygen ASTM D 5373 1.69Sulphur ASTM D 4239-05 0.99Ash ISO 1171:2010 18.8Moisture ISO 11722:1999 1.4

Figure 2—Cross-sectional front view of combusted coal

than at the production well. The permeability of the clinkerallowed oxygen to infiltrate deeper into and react with thecoal (Wakatsuki, et al., 2009); this was notable particularlyat the base of the injection well where the air flow wasturbulent. The height difference between the clinker layers ateach of the two wells is 8 mm and the width difference is 26mm due to the large crack that formed where the surface areaof coal is in contact with air (see Figures 2 and 3). Theteardrop shape of the clinker gives an indication of the airflow pattern in the coal.

Heat was provided to the coal for 30 minutes, but the airflow continued for a longer period. It was observed that thetemperature continued to rise for a further 30 minutes untilthe self-heating system halted due to the eventual cooling ofthe coal surface by the air flow, which entered the coal at23°C. When coal reaches temperatures of approximately600°C, a self-sustaining heating will be established(Hoffman, 2008).

Experiment 2

The coal used for Experiment 2 was cemented into therefractory chamber, whereas the coal in Experiment 1 wasonly surrounded by refractory sand; this prevented the coalfrom cracking to such an extent. The shape of the clinker asshown in Figures 6 and 7 is similar to that seen inExperiment 1, however it is more cylindrical. Upon closerinspection (see Figures 8 and 9) it can be seen that theclinker layer is in fact slightly thicker at the base of theinjection well (the well on the left of each photograph) thanthe production well (on the right).



Figure 3—Cross-sectional top view of combusted coal

Figure 4—3D cross-sectional model of combusted coal with clinkerformation

Figure 5—3D model of clinker formed in the coal after the duration ofthe experiment

Figure 6—Cutaway section of the coal showing the clinker formation

Figure 7—The shape of the clinker formation


Comparison to the research by Daggupati et al

The current research was roughly based on the work done byDaggupati et al. (2010), the major differences between thetwo investigations being the type and size of the coal used,the heat source adopted, and the experimental duration.However, the general trends of cavity formation can still becompared.

Figures 10 and 11 compare the top views of the cavityformation in the experiment by Daggupati and Experiment 1respectively. The injection well is indicated by the white circleon the left of each image and the production well is on theright.

It is clear that in both cases the cavity at the base of theinjection well is larger than at the production well. This shapeis not as obvious in Figure 11 due to the fact that a heatingelement was used across the entire length of the horizontaltunnel, causing clinker to form. It is also noteworthy that ashwas formed in Daggupati’s experiment, whereas clinker wasformed in the current investigation. This can be attributed tothe different coal characteristics and operating conditions.

Conclusions and recommendations

A laboratory-scale method was used successfully to simulatecavity formation during underground coal gasification. It wasfound that cracking tends to occur along the bedding plane ofthe coal due to the formation of clinker or expanded vitrinite,which forms in a thicker layer at the base of the injection wellthan at the base of the production well. This can be attributedto the turbulence of the air flow as well as the relatively highoxygen concentration at the base of the injection well. The airflow pattern has a great effect on cavity formation. Thecombustion products were not analysed, but should be infuture work. It is recommended that further work be done onthe heating source, such as using an alternative method. Theuse of Liquefied Petroleum Gas (LPG) is recommended as thismethod allows the experiment to mimic more of an industrialUCG process, in addition, if the temparature of the cavitycould be controlled more effectively the clinker formationshould be able to be minimized. Further investigationsshould take changes in the process variables into account;these variables include varying the injection air flow rate, thetype of coal, duration of heating and the temperature of theheating source. Using steam together with the air wouldallow gasification reactions to occur in addition tocombustion reactions, and thus mimic actual UCG processesmore accurately.

▲


Figure 8—Cross-sectional front view of combusted coal withdimensions

Figure 9—Cross-sectional top view of combusted coal

Figure 11—Tomogram showing the plan view of the cavity produced inExperiment 1

Figure 10—Plan view of the cavity obtained by Daggupati et al. (2010)

Acknowledgements

The authors would like to extent their appreciation to theirsupervisor, Professor Wagner; to the workshop manager, MrSamuel-McRae; to Doctor Kauchali and to Mr Hoffman andMr de Beer at the micro-focus X-ray facility at Necsa.

References

ABDUL, W.B., AQEEL A.B., and GHOLAMREZA, Z. 2013. Underground coal gasifi-cation: From fundamentals to applications. Progress in Energy andCombustion Science, vol. 39. pp. 197–199.

AFRI COAL INVESTMENTS PTY LTD. 2013. What causes clinkers in coal fired boilers.http://www.africoal.co.za/what-causes-clinkers-in-coal-fired-boilers[Accessed 18 Oct. 2013].

DAGGUPATI, S., GANESH, A.,MANDAPATI, R.N., MAHAJANI, S.M., MATHUR, D.K.,SHARMA, R.K., and AGHALAYAM, P. 2010. Laboratory studies on combustioncavity growth in lignite coal blocks in the context of underground coalgasification. Energy, vol. 35. pp. 2374–2386.

DIRECTIONAL DRILL PTY LTD. 2011. What is horizontal directional drilling.

http://www.directional-drill.com/horizontal-directional-drilling.html[Accessed 26 Oct. 2013].

HOFFMAN, G. 2008. Natural clinker. http://geoinfo.nmt.edu/staff/hoffman/clinker.html [Accessed 18 Oct. 2013].

HOFFMAN, J.W. and DE BEER, F. 2013. Micro-focus X-ray tomography facility(MIXRAD) at NECSA. Personal communication.

MITRA, S.K. 2011. Coal testing and analysis. http://www.mitrask.com/coaltesting-analysis/index.html [Accessed 24 Oct. 2013].

SEDDON, D. and CLARKE, M. 2011. Underground coal gasification (UCG), itspotential prospects and its challenges. http://www.duncanseddon.com/underground-coal-gasification-ucg-potential-prospects-and-challenges[Accessed 25 Apr. 2013].

SELF, S.J., REDDY, B.V., and ROSEN, M.A. 2012. Review of underground coalgasification technologies and carbon capture. International Journal ofEnergy and Environmental Engineering, vol. 3, no. 16. pp. 1–8

WAKATSUKI, Y., HYODO, M., YOSHIMOTO, N., NAKASHITA, A., and IKEDA, T. 2009.Material characteristics of clinker ash and examination of applicability forembankment. Powders and grains 2009: Proceedings of The 6thInternational Conference on Micromechanics of Granular Media, Golden,CO, 13–17 July 2009. http://scitation.aip.org/content/aip/proceeding/aipcp/10.1063/1.3179872 [Accessed 24 Oct. 2013]. ◆



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Project background

What is coal seam methane?

Methane (CH4) is formed as part of the processof coal formation. When coal is mined methaneis eventually released from the freshly brokencoal face. Methane can also be released as aresult of natural erosion or faulting. The depthof the seam predicts the amount of methanecontent present. Methane is directly exposed tofresh air when mining takes place on surfaceand is confined when mining takes placeunderground..

Methane gas in coal mines

Methane gas in underground coal mining is abig concern. Methane explosions aredevastating, causing significant loss of life anddamage to property, and there is a significantindustry effort to prevent these accidents fromoccurring.

Methane becomes explosive only if it isdiluted to between 5%–15% by volume in air.Failure to provide adequate ventilation to

dilute the methane to less than 5% by volumeincreases the threat of an explosion.

Following the Pike River disaster in NewZealand in November 2010, it became a majorconcern at Arnot 10 Shaft to manage itsmethane levels so as to avoid a similarincident.

The Pike River disaster

The Pike River disaster shocked the world. On19 November 2010 at 3:45 pm there was anunderground methane explosion at the PikeRiver coal mine which resulted in the loss of29 lives. Daniel Rockhouse and Russel Smithwere the only two people underground thatsurvived the explosion. The emergencyresponse was led by the New Zealand police. Arescue attempt was prevented by a lack ofinformation regarding the conditionsunderground. On 24 November a secondexplosion occurred. and all hopes of findingthe 29 miners underground alive wereabandoned. The focus moved to the recoveringof the bodies. However, conditionsunderground made this impossible. Twofurther explosions occurred, the second ofwhich ignited the coal underground. The mineentrances were sealed in January 2011. Thisevent raised concerns throughout the coalmining industry to prevent methaneexplosions. (Royal Commission of the PikeRiver Coal Mine Tragedydegy, October 2012)

Scope of studyThe aim of the study at Arnot was theprevention of methane explosions and notpreventing a methane explosion from leadingto a coal dust explosion. Thus the project’smain emphasis was on what happened at PikeRiver, how the tragedy could have beenprevented; what measures Arnot already has

Arnot’s readiness to prevent a Pike Riverdisasterby R. Weber*Paper written on project work carried out in partial fulfilment of B. Eng. (Mining Engineering)

SynopsisMethane explosions in underground coal mines are a major concernacross the mining industry. After the Pike River disaster, Arnot Coalbecame even more aware of the explosion risk. To determinewhether Arnot has adequate precautions against a methaneexplosion, such as the one that led to the Pike River disaster, aliterature survey covering the causes of the incident and whatpreventative measures should have been in place was conducted.Many different methane explosion prevention methods wereevaluated, as well as codes of practice. Based on the findings,Arnot’s measures to prevent conditions favourable for a methaneexplosion were evaluated.

Keywordsmethane explosion, code of practice, dominoes, ventilation,LTR(last through road).

* Department of Mining Engineering, University ofPretoria.



in place, and what further measures it needs to take in orderto prevent a similar disaster. There was no emphasis onactions to be taken in the event of a methane explosion, onlyon prevention. The project also did not consider rescueprocedures or assistance to the bereaved families.

MethodologyA survey was conducted and a hierarchy of information onmethane explosions built up by using methods such asconsulting with experts in the field and contacting othermines that had suffered methane explosions. From thehierarchy the most relevant information was selected.Underground visits were conducted during December 2012and June 2013. The visits were scheduled so as to ensurethat enough time was spend at each installation to ensure theeffectiveness of Arnot’s methane management and explosionpreventative measures. Following the evaluation, conclusionswere drawn and the necessary measures put into place tomake Arnot ‘Pike -River ready’.


Ventilation requirements➤ COP requirement—Barrier pillars should be spaced

15 m apart.➤ Actual—From Figure 1 it can be seen that the barrier

pillar spacing separating panels averages 23.35 m, thuscomplying with the COP

➤ COP requirement—Each section should have its ownventilation district with a separate set of ventilationcontrols

➤ Actual—It can be seen from Figure 2 that Section 4 andSection 12 have separate ventilation districts, whichmeans each section has separate intakes as well asreturn airways. The air flows are kept separate bymaking use of walls and air crossings, thus complyingwith the COP

➤ COP requirement—Under normal conditions 0.25 m3/sof fresh air should be supplied in the last through-road,and 0.3 m3/s under high-risk conditions

➤ Actual—Taking the average bord height and width as 3m and 6.5 m respectively, the following air flows werecalculated:Normal conditions = 7 faces * 3 m * 6.5 m * 0.25 m3/s

= 43.2 m3/sHigh- risk conditions = 7 faces * 3 m * 6.5 m * 0.3 m3/s

= 41 m3/sFrom Figure 3 it can be seen that that sections 4 and 12

comply with the COP. The air flow in Section 3 was measuredas 46.9 m3/s, thus also complying with the COP.

Abandoned panel ventilation

➤ COP requirement—Abandoned panels should be sealedoff or ventilated until they are sealed off. Where theyare still ventilated, the air flow velocity in the LTRshould be greater than 0.5 m/s

➤ Actual—From Figure 4 it can be seen that panel N31 isabandoned and thus sealed off as required by the COP.Figure 5 shows an abandoned panel that is still beingventilated prior to sealing off. The air flow velocity is1.1 m/s, thus complying with COP.

Arnot’s readiness to prevent a Pike River disaster

▲


Table I

Depth vs. methane content (Van der Merwe, 2013)

Depth interval Mean methane content (metres) (cubic metres per ton of coal)

100 0.02500 0.991000 3.731500 4.892000 7.09

Figure 1—Mining layout showing barrier pillars

Figure 2—Air-crossing separating ventilation districts

Figure 3—Section 4 and Section 12 layout and air flow rates

Intakes and returns roads minimum velocities

Table II and Table III show the average intake and returnvelocities for each section. All of the readings can be seen tocomply with the COP requirement of 0.5 m/s.

Schedule of checks

Table IV shows the inspection intervals prescribed by theCOP, together with the actual inspection intervals notedduring underground visits. It must be noted that theinspection intervals are those prescribed under normalconditions, although there could be exceptions.

Analysis of results

‘Swiss cheese’ model of causation

Each layer in a ‘Swiss cheese model’ (Figure 6) represents adefensive system labelled by type (at the top). The holes ineach layer represent gaps in the defensive system. Thesegaps can be created by active failures, human error,violations etc. Once these gaps line up there is no defenceand an accident such as the Pike River disaster is likely tooccur.

For example, if the ventilation at Pike River had beenadequate and if there had been no ignition sources, then theaccident would not have occurred. However, both theventilation and the spark prevention measures wereinadequate. The more defensive systems in place, the betterthe chances that not all of the holes in the model will line up(the probability of an incident decreases).

It is thus of great importance to have as many defensivesystems as possible. If the critical systems fail, the secondaryor ancillary systems must kick.

Dominoes at Pike vs. Arnot mandatory Code of Practice

From Table V it can be seen that 24 dominoes were identifiedthat led to the Pike River disaster. Visual inspections,calculations, and interviews with employees indicated thatArnot is able to prevent all 24 of these dominoes.

Conclusion

Failure to control the methane levels resulted in led to theinevitable explosion at Pike River during 19 November 2010,which resulted in the unnecessary loss of 29 lives. By



Table II

Intake velocities

Section COP intake requirement (m/s) Section minimum measurement (m/s) Average intake velocity (m/s)

3 0.5 1 1.34 0.5 1 1.212 0.5 1 1.1

Figure 4—Example of sealed and unsealed abandoned panels

Figure 5—Air velocity in last through-road of abandoned panel

Table III

Return velocities

Section COP return requirement (m/s) Section minimum measurement (m/s) Average return velocity (m/s)

3 0.5 1.2 1.74 0.5 1.3 1.912 0.5 1.3 1.7


thoroughly working through the Pike River disaster, 24 maindominoes contributing to the Pike River disaster wincident aswere identifiedlisted. It can be seen that ventilation practisceswas the main issue involved.

Following up on a review of Arnot’s methane controlsystems, which were obtained and data of which includedobtaining from the ventilation readings, conducting visualunderground inspections, and exercising to communicatewith the interviews with relevant persons at the mine itself,Arnot’s practices were compared to those of Pike River. Fromthese comparisons it can be seen that Arnot can prevent all ofthe twenty four 24 main dominoes that played a role at PikeRiver.

The fact that Pike River was 150 m deep, compared withonly 60 m at Arnot, could also have played a role. The mean

methane content measured in cubic metres per ton of coalincreases with increasing depth of the mine. The methaneemission rate would therefore have been higher at Pike Riverthan at Arnot. (This can be seen from Table I).

Recommendations

Increase the scoop/line brattice efficiency

The scoop/line brattice efficiency can be calculated bycomparing the quantity of air entering the section (point 1,Figure 7) to the quantity of air leaving the section (point 2).As an example, if the amount of air entering is 1.6 m/s andthe amount leaving is 1.6 m/s, the efficiency will amount to100 per cent.

▲


Table IV

Inspection intervals

Inspected According to COP Inspections interval measured Compliance with COP?

Main ventilation to section Start of shift and then every three hours Start of shift and then every three ✓

hours. Deviations were less than 10 minutes.

Last through-road velocity Start of shift and then every three hours Start of shift and then every three ✓


Positive ventilation of faces Start of shift and then every three hours Start of shift and then every three ✓


Scrubber screen and fan on the Start of shift and when changing picks Start of shift and every time picks ✓

mechanical miner are changed

Ventilation brattices and section walls Start of shift Start of shift ✓

installed according to standard

Trailing cables Start of shift and at least once during shift Start of shift and mostly once, ✓

sometimes twice.

Flame proofing Start of shift visual inspection Start of shift ✓

Operating conditions of spray nozzles Start of shift and when changing picks Start of shift and mostly once, ✓

sometimes twice.

Mechanical miner onboard flammable Start of shift Start of shift ✓

gas monitor tested

Test for flammable gasin each heading Start of shift and then every three hours Start of shift and then every three ✓

up to second-last row of support hours. Deviations were less than 10 minutes

Figure 6—Swiss cheese model (Bredel, 2013)

It is recommended that the efficiency be maintained equalto or greater than 70 per cent.

From Table VI it can be recommended that the scoop/linebrattice efficiency on Section 4 must be drastically increased,since this low efficiency results in less air being supplied inthe last through-road. Section 12 can also consider improvingtheir efficiency. Section 3 has the highest efficiency, but it isstill nevertheless recommended that more effort be put intominimizing the number of readings below 70 per centefficiency.

Reporting format

It is recommended that I.O (in order) and O.O.O (out oforder) should not be used to report on ventilation readings inthe shift overseer’s daily logbook. The quantity of thereadings should rather be recorded so as to build up a recordof ventilation readings underground.

Suggestions for further work

As we have learned, by placing the main fan at the PR in theunderground vicinity was ‘a major error’. It appears that thesafety measures similar in Australia (but not legalrequirements in NZ) were not enforced nor instituted to beginwith. It is thus suggested that the aforementioned and vitallegislation in different countries regarding the prevention ofmethane explosions be investigated. In addition, investi-gations should be extended to the prevention of coal dustexplosions, and not only methane explosions. As with mainfans that are not banned underground within the New-Zealand laws it is suggest that there could be further lookedinto the role the laws of the different countries play when itcomes to the prevention of methane explosions. It can also besuggested that it must be further looked at the prevention ofa coal dust explosion and not only that of a methaneexplosion

References

BREDEL, P. 2013. Risk [Interview] (14 Apr. 2013). Senior lecturer, University of

Pretoria

MINING-TECHNOLOGY. 2012. mining-technology. http://www.mining

technology.com/projects/pikeriverminenewzeal/pikeriverminenewzeal3.

html [Accessed 23 Jan. 2013].

Royal Commision of the Pike River Coal Mine Tradegy, October 2012. Report

Volume 1, Wellington, New-Zealand.

VAN DER MERWE, N. 2013. Mineral recource manager, Exxaro, Arnot.

WIKIPEDIA. 2013. Wikipedia. http://en.wikipedia.org/wiki/Pike_River_Mine

[Accessed 23 Jan. 2013]. ◆



Table VI

Efficiency table

Section Above 70% Below 70% Average%

3 36 18 754 8 46 6412 24 30 68

Table V

Dominoes

Domino Complies with COP

Pike River Arnot

Proper ventilation of goaf area x ✓

Sealing of the hydro panel x ✓

Unrealistic drive towards coal production x ✓

Main fan placed underground x ✓

In-bye ventilation fragile x ✓

Ventilation engineer x ✓

Second intake x ✓

Assessing of Health and Safety reports x ✓

Free venting of drained methane x ✓

Tripping of machines x ✓

Sensor bypassing x ✓

Experience of mine management/ x ✓

constant change

Sensor at bottom of vent shaft x ✓

Placement of the sensors and reporting x ✓

Sufficient sensors x ✓

Voltage cables near pipes transporting x ✓

methane creating a hazard

Stoppings separating x ✓

intake air from return air

Diverting air away from face and shutting x ✓

down of air while work continuous

Risk assessment conducted when fixing x ✓

restricted and nonrestricted areas.

Variable-speed drives x ✓

Training of employees x ✓

Widening of the panel from 30m to 45m x ✓

High methane readings reported x ✓

Main fan flameproof x ✓

Figure 7—Layout of a basic section at Arnot

Organised byOrganised by

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Introduction

AA2014 is a lightweight and high-strengthmaterial that has proven invaluable inapplications that require a high strength-to-weight ratio. The alloy finds most of itsapplications in the aerospace industry,although new inroads into the drill pipingindustry have been established. Due to the lowalloying element content in AA2014, thematerial displays poor hardness and, therefore,poor wear resistance. There is therefore a needto increase the wear resistance of AA2014without altering the bulk mechanical propertiessuch as ductility and strength-to-weight ratio.

Scientific background

Due to the low alloying element content,AA2014 is not particularly hard and istherefore not abrasion resistant. This is thecase for most 2000 series aluminium alloys,

and it has become desirable for industrialapplication to increase the wear resistance ofthese alloys without compromising thematerial’s strength or strength-to-weight ratio.The alloy composition and relevant mechanicalproperties are shown in Tables I and II respec-tively.

A potential solution to the low wearresistance of AA2014 is to coat the materialwith a wear-resistant layer. This project willconsider the use of laser cladding an Al-Cu-Sicompound onto the AA2014 substrate with theintention of increasing the wear resistance ofthe surface while leaving the bulk mechanicalproperties of the material unchanged. The aimof this cladding process is to produce a hardermaterial by intermetallic phase formationbetween aluminium and copper in the form ofAl2Cu. This intermetallic phase formation aswell as the microstructure can be seen inFigure 1. This microstructure was achievedusing a 40 wt% Cu 60 wt% Al cladcomposition (Dubourg, 2002). It was foundthat the intermetallic phase was the first tosolidify, and the resulting hardness andabrasion resistance was the highest in thealloy range that was tested.

In the current investigation, silicon wasadded in an attempt to minimize porosity inthe cladding. If successful, this claddingprocess could be applied to many aluminiumalloys in which wear resistance is importantbut impractical to achieve by conventionalmethods. Tough alloys that do not necessarilydisplay good wear characteristics could besurfaced using this process to increase theresistance to abrasive wear while maintainingthe toughness of the bulk material.

Laser cladding AA2014 with a Al-Cu-Sicompound for increased wear resistanceby K.J. Kruger* and M. du Toit*Paper written on project work carried out in partial fulfilment of B. Eng. (Metallurgical Engineering)

SynopsisAluminium alloys have gained popularity in many industries due totheir high strength and low weight. One shortcoming of aluminiumalloys is their poor resistance to abrasion and erosion wearcompared to materials such as stainless steels. In this project,aluminium alloy 2014 (AA2014) was coated with a 1.5 mm thicklaser-deposited layer composed of silicon, copper, and aluminiumwith the aim of increasing the wear resistance. The amount ofsilicon, copper, and aluminium added to each sample wasdetermined by a mixtures model. It was discovered that the Al-Cusystem is very sensitive to silicon additions and that wearresistance depends on the primary phase to solidify as well as onthe final phase distribution. Two primary phases were identified;alpha aluminium and theta intermetallic. It was observed that theclad layer increases both the hardness and wear resistance ofAA2014, and that the material solidifying as primary alphaaluminium displayed a lower hardness but higher wear resistancethan the samples containing primary theta phase. All clad layersperformed better in terms of wear resistance than the uncladsamples. The knowledge gained and principles used in this projectcould be applied to many other aluminium alloys.

Keywordsaluminium alloys, laser cladding, wear resistance.

* Department of Materials Science and MetallurgicalEngineering, University of Pretoria, Pretoria, South Africa.



Laser cladding AA2014 with an Al-Cu-Si compound for increased wear resistance

Literature review

Laser cladding

Laser cladding is a low heat input, rapid solidificationprocess. The advantages of this process include high levels ofproductivity due to fast welding speeds, low levels of dilution,refined microstructures due to rapid solidification, minimaldistortion of the work piece, and small heat-affected zones(HAZs) (CSIR, 2012). There is a general consensus on theadvantages of laser cladding in these respects, regardless ofthe cladding feed material or substrate being clad (Liu, 1995;Hyatt, 1998; CSIR, 2012; Joining Technologies, 2012). Thisprocess makes it possible to bond a material with acompletely different microstructure and mechanical propertiesto a substrate. The process makes use of high-intensity lightemission that is focused using several lenses to a pointabove, below, or on the surface of the work piece dependingon the intended application (Joining Technologies, 2012). In

the case of laser cladding, the focal point of the beam isslightly above the work piece so as to melt the feed materialwithout melting much of the substrate. This causes minimaldilution and rapid solidification of the feed material onto thesubstrate while ensuring that a metallurgical bond isachieved. These properties are desirable for the above-mentioned application. This will result in little to no changein the substrate bulk properties while achieving the desiredresult of depositing a layer of feed material that will act as anabrasive-resistant coating.

A powder feeder was used to feed the material into theweld pool via a carrier gas. The powder feeder method waschosen over other methods, such as preplaced powders,because it is the only method that is industrially viable (Vilar,2001). Figure 2 illustrates the process of laser cladding usinga powder feed. It can be observed from the schematic that thepowder is fed into the weld pool via a powder injection nozzleand is transported in space by a carrier gas. Once the powderenters the weld pool, it melts and will rapidly solidify to forma coating as the laser moves along the track.

Cladding material

An important microstructural aspect to consider whenselecting materials suitable for wear resistance is thepresence of a finely distributed hard phase in a matrix of amore ductile phase. It is also important to ensure that the

▲


Table I

Typical chemical composition (in weight %) of aluminium alloy AA2014. Single values denote maximum limits.(Capalex, 2013)

%Cu %Fe %Si %Mn %Mg %Cr %Zn %Ti %Other %Al

3.8–4.9 0.5 0.5 0.3–0.9 1.2–1.8 0.1 0.25 0.15 0.15 Balance

Table II

Typical mechanical properties of AA2014 in the T4, T351 heat-treated condition (Kaufman, 2002)

Ultimate tensile strength (MPa) Yield strength (MPa) Brinell hardness number (500 kg/ 10 mm) Modulus of elasticity (GPa)

472 325 120 73

Figure 1—Microstructure of laser-clad 40%Cu-Al material formingprimary intermetallics and a secondary eutectic phase etched withKeller’s reagent (Dubourg, 2002)

Figure 2—Schematic of laser cladding using a powder feeder (Vilar,2001)

cladding material is compatible with the substrate and that nobrittle phases are formed along the fusion line that couldresult in poor metallurgical bonding.

When considering an abrasion-resistant layer, the mostimportant aspect is the cladding hardness. There is a generalconsensus in the literature that material hardness is directlyproportional to abrasion resistance (Kim, 2006; Jeong, 2003;Caldwell, 1988). This is due to the fact that harder materialsare more resistant to plastic deformation and will thereforeresist the process of wear more effectively than softermaterial. Figure 3 illustrates the effect of hardness on thewear resistance of pure metals and several steel alloys.

From Figure 3 it is clear that, in general, as the hardnessof a material increases, so does its abrasion wear resistance.

It can be observed from the results of the study byDubourg (2002) that the wear resistance of the layer isdirectly related to cladding hardness. Figure 4 is a graphicalrepresentation of the wear resistance of an aluminium sampleas the copper content is increased.

Figure 5 shows a graphical representation of the hardnessof the clad material measured at varying depths prior to anypost-cladding heat treatment (Dubourg, 2002). It is evidentthat an increase in copper content increases the hardness ofthe clad layer. The material displays a rapid drop in hardnessat a depth of approximately 0.9 mm, most likely due to thetransition into the virgin aluminium base material.

Figure 5 indicates that the hardness of the claddingincreases with copper content. This is due to an increase inthe amount and particle size of the intermetallic θ phase.

Figure 6 displays the equilibrium Al-Cu binary phasediagram. It can be seen that the eutectic composition is at 33wt% Cu; however, this true only for the binary system underequilibrium conditions.

Pseudo-binary Al-Cu-Si phase diagrams are considered inFigures 7 and 8 at 1% and 10% silicon respectively. It is clearthat as the silicon concentration of a sample is increased, the

formation of primary theta phase is favoured at lowerconcentrations of copper. The formation of either primaryalpha aluminium or primary theta phase could affect themechanical properties of the material and must therefore becarefully monitored.



Figure 3—The relationship between wear resistance and hardness inseveral pure metals and alloys (Tylczak, 1992)

Figure 4—Graphical representation of the decrease in wear rate underdifferent wear parameters with increasing copper content in thecladding material (Dubourg, 2002)

Figure 5—Graphical representation of hardness for various claddingcompositions as a function of depth (Dubourg, 2002)

Figure 6—Aluminium-copper phase diagram (Murray, 1992)


Methodology

Twelve samples were laser-clad with layers containing from30–54 wt% copper and 0–3.4 wt% silicon (Table III). Onesample remained unclad and was subjected to the sametesting procedure as a control:

The layer compositions were selected using a statisticalmethod known as the mixture lattice design. This design waschosen to maximize the useful data that is extractable fromthe experiments and allow for the mapping of materialcharacteristics based on chemistries using contour lines..

The centre point in the experiment (samples 1, 6, and 12)was performed in triplicate in order to test the consistency ofthe cladding process.

The samples were produced by gas-feeding three powders(pure Cu, pure Al, and 12%Si-Al alloy) from three hoppersinto the weld pool via a triaxial nozzle.

The laser parameters used to produce the samples arelisted in Table IV.

These parameters produced a smooth, well-bonded cladlayer, but have not been optimized for all the samples. Afterpolishing, some unmelted particles were observed in theinter-bead region of most samples.

The clad samples were subjected to chemical analysis,micro-Vickers hardness tests, and microstructural analysis

using optical emission spectrometry (OES), as opticalmicroscopy, and scanning electron microscopy usin energy-dispersive spectroscopy (SEM-EDS). Based on the results ofthese tests, two samples were selected for wear testing,together with the unclad control sample.


The results from the OES chemical analysis, micro-Vickershardness tests, and SEM-EDS phase analysis are displayed inTable V.

Chemical analysis

It is evident from the Table V that while the silicon contentwas acceptable in each case, the copper content in all sampleswas consistently lower than the designed copper content.This implies a problem with the copper delivery system. Apossible solution would be pre-mixing of the powders andusing only one hopper.

The system is very sensitive to silicon content and for thisreason, unexpected solidification patterns were observed.Several samples solidified as primary α-Al while otherssolidified as primary θ-intermetallic.

Localized EDS measurements indicated that neither theprimary nor the eutectic phase contained silicon values higherthan 1%. Microstructural examination of the cladding wastherefore carried out.

Microstructure

The addition of silicon promotes the formation of primary θ

▲


Figure 7—Vertical section of the Al-Cu-Si phase diagram containing 1%Si (Raghavan, 2007)

Figure 8—Vertical section of the Al-Cu-Si phase diagram containing10% Si (Raghavan, 2007)

Table III

Example of a table used to capture experimentaldata (weight %) (Balance is aluminium)

Sample no. Si content in clad (%) Cu content in clad (%)

1 Unclad control2 0 303 1.1 37.14 2.3 44.35 3.4 51.56 0 427 2.3 44.38 4.5 46.59 6.8 48.810 0 5411 1.1 49.212 2.3 44.313 3.4 39.5

Table IV

Final laser parameters

Laser parameter Final value

Power 2.5 kWSpot size 3 mmStep-over distance 0.8 mmResultant overlap 73.3%Welding speed 1.5 m/minGas flow rate (1.5 l/min per hopper) 4.5 l/min

phase, which was apparent in the microstructure of mostsamples (as seen in Figure 9). However, the primary α phaseformed in several samples, (Figure 10).

It can be seen from Figure 9 that there is a large amountof primary θ phase and very small amounts of eutectic phasein the microstructure of sample 9. This is considered to be thedesired microstructure due to the large amounts of hardintermetallic phase surrounded by a more ductile eutecticphase.

As shown in Figure 10, sample 12 displays primary phase solidification and this results in soft particles that aresurrounded by the harder (α-Al +θ) eutectic phase.

Samples containing higher amounts of silicon (3% andmore) contained a third phase, consisting of small greyparticles that were visible under an optical microscope. Anexample of these particles can be seen in Figure 11.

These small grey particles were difficult to identify usingSEM back-scatter electron imaging (BEI), and a chemical mapwas constructed in order to identify the locations of theparticles. The results can be seen in Figure 12.



Table V

Results of chemical analysis, hardness tests, and phase analysis

Sample no. Design Si content Design Cu content Hardness (HV) OES analysed Si OES analysed Cu Phases presentin clad (wt %) in clad (wt %) content (wt %) content (wt %)

1 2.3 44.3 243 2.2 33.3 Al+θ2 1.1 37.1 234 1.1 28.5 Al+θ3 0 30 214 0.2 22.3 Al+θ4 3.4 51.5 370 3.1 39.7 Al+θ+Si5 0 42 277 0.2 31.6 Al+θ6 2.3 44.3 225 2.1 31.8 Al+θ7 4.5 46.5 259 4.2 35.4 Al+θ+Si8 6.8 48.8 250 6 33.5 Al+θ+Si9 0 54 260 0.2 41.5 Al+θ10 1.1 49.2 265 1.2 39.2 Al+θ11 3.4 39.5 203 3.2 32 Al+θ+Si12 2.3 44.3 199 2.1 33 Al+θ13 Control Control 116

Figure 9—Sample 9, displaying primary theta solidification

Figure 10—Sample 12 displaying primary alpha solidification

Figure 11—Sample 4, containing tertiary grey phase

Figure 12—Chemical map of sample 8


It is clear that there are areas of high silicon concentrationin this sample and these areas cannot be distinguished fromthe rest of the microstructure by SEM. The particles could notbe accurately analysed using EDS due to their small size, andthermodynamic modelling using FactSage™ was used topredict the equilibrium phases that form on solidification inthe samples. The third phase that forms in the high-siliconsamples was identified as high-purity (>99%) siliconparticles.

When the fusion line was inspected, it was evident thatvery good metallurgical bonding between the clad layer andthe metal substrate had been achieved. This is evident inFigure 13.

The formation of grain boundary precipitates can also beobserved; however, these precipitates could be eliminatedthrough solution annealing.

Mechanical properties

Hardness is believed to have the largest influence on thewear resistance of the material, and thus all the samples weresubjected to fifteen hardness measurements using the micro-Vickers hardness test. The results were collated and used togenerate a hardness contour map with relation to individualsample chemistries, using ‘Design Expert’ (Figure 14).

It can be seen that high harness is due to a synergisticeffect between copper and silicon.

Wear testing

Based on the results obtained from the above tests, samples12 and 4 as well as the unclad control sample were subjectedto wear testing. Sample 12, which solidified as primary α-Al,displayed the lowest hardness, and sample 4, whichsolidified as primary θ-intermetallic, displayed the highesthardness. A slurry erosion wear test proved to be toouncontrollable to provide consistent results, and therefore anon-standard, wet two-body abrasion wear test wasperformed. Tables VI and VII describe the test conditions andresults.

It is important to note that the wear test was extremelyaggressive and high mass loss was observed. Sample 12displayed the highest wear resistance, while the uncladsample displayed the lowest wear resistance. Although

sample 4 had a higher hardness than sample 12, it had ahigher mass loss during the wear test. This can probably beattributed to spalling at the sample edges (as shown in Figure 15).

▲


Figure 13—Fusion line between cladding layer (top) and substrate(bottom)

Figure 14—Hardness (HV) contour map based on chemistries obtainedfrom OES results

Table VI

Test conditions in non-standard wet two-bodyabrasion wear test

Parameter Value

Abrasive medium Diamond-impregnated resin (220 grit)Force applied to sample 100 NRun time 10 minRotation speed 310 r/min

Table VII

Mass loss in two-body abrasion wear test

Sample no. Mass loss per unit area exposed (g/cm2)

Control sample (unclad) 0.162Sample 4 (hardest) 0.156Sample 12 (softest) 0.134

Figure 15—Spalling of sample 4

Conclusions

The surface hardness of AA2014 was successfully increasedby laser cladding. The effects of the cladding operation wereexamined using a variety of tests, and an understanding ofthe mechanisms by which the cladding layer will protect thebase material from an abrasive environment has beengained.

This cladding process was successful in providingadditional wear resistance as well as hardness to thesubstrate material without reducing the strength-to-weightratio of the material. The wear resistance of the materialshowed a strong correlation to the primary phase thatsolidifies, and the solidification mechanism is linked to boththe thermodynamics and kinetics. It was determined thatprimary alpha aluminium solidifying samples are more wear-resistant, but have lower hardness than the primary thetaphase intermetallic solidifying samples. The reason for thelow wear resistance of the hard primary theta solidifyingsamples is most likely due to spalling. Furthermore, it wasdetermined that both the primary intermetallic and primaryalpha aluminium samples were harder and more wear-resistant than the substrate. This implies that the processcould be applied to a wide range of aluminium alloys and willbroaden the range of application of aluminium alloys byincreasing the lifespan of these materials under severe wearconditions.

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DUBOURG, L., PELLETIER, H., VAISSIERE, D., HLAWKA, F., and CORNET, A. 2002.Mechanical characterisation of laser surface alloyed aluminium–coppersystems. Wear, vol. 253, no. 9. pp. 1077–1085.

HYATT, C.V. 1998. The Effect of Heat Input on the Microstructure and Propertiesof Nickel Aluminum Bronze Laser Clad with a Consumable of CompositionCu-9.0Al-4.6Ni-3.9Fe-1.2Mn. A Metallurgical and MaterialsTransactions.

JEONG, D.H. 2003. The relationship between hardness and abrasive wearresistance of electrodeposited nanocrystalline Ni–P coatings. ScriptaMaterialia. pp. 1067–1072.

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KAUFMAN, J.G. 2002. Aluminum alloys. Handbook of Materials Selection. Kutz,M. (ed.). John Wiley & Sons. Ch. 4. p. 104.

KIM, K.T. 2006.

KYUNG TAE KIM, S. I. 2006. Hardness and wear resistance of carbon nanotubereinforced Cu matrix nanocomposites. Materials Science and Engineering.pp. 46–50.

LIU, Y. 1995. Microstructural Study of the Interface in Laser-Clad Ni-AI Bronzeon AI Alloy AA333 and Its Relation to Cracking. A Metallurgical andMaterials Transactions.

Murray, J. 1992. Binary alloy phase diagrams. Introduction to Alloy PhaseDiagrams. Vol. 3. Baker, H. (ed.). ASM Handbook, ASM International,Materials Park, OH. p. 2.44.

RAGHAVAN, V. 2007. Al-Cu-Si. Journal of Phase Equalibria and Diffusion. pp. 180–181.

TYLCZAK, J.H. 1992. Abrasive wear. Friction, Lubrication, and WearTechnology. ASM Handbook, vol. 18. Materials Park, OH. pp. 184–190.

VILAR, R. 1999. Laser cladding. Journal of Laser Applications, vol. 11, no. 64.pp. 64-81. ◆



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Introductions

It is important to determine whether iron oreused as feed to a sinter plant will produce asinter suitable for use in a blast furnace. Thefeed to the blast furnace should be permeable(slightly porous, but not too porous since thiscould have an adverse effect on sinterproperties) and have a high strength. A fewstandard tests are available to evaluate rawmaterials for use in the sinter process. Thecurrent method used can be difficult tointerpret, time-consuming, and providesmostly limited information. A new testmethod, which allows calculation of thereactivity of iron ore with lime with increasingtemperatures, was investigated. The newmethod should be easy to implement,reproducible, quick, and give valid results.

Background

Kumba Iron Ore identified a need to develop anew test method to evaluate iron ore in termsof lime assimilation. Lime is added to iron oreto improve the porosity and to obtain the

correct sinter strength and permeability for usein the blast furnace. If the sintered ore doesnot have adequate strength, a large amount offines will be produced during stockpiling andhandling of ore before use in the blast furnace.These fines will then be blown out by the off-gas from the top of the blast furnace.Insufficient porosity of the sintered ore, on theother hand, will not allow sufficient flow ofgas through the ore, causing insufficientreduction of iron in the blast furnace andresulting in a product that is not suitable forsteelmaking. The porosity of the sintered oreshould, however, not be too high since thiscan have an adverse effect on the sinterproperties.

Higuchi et al. (2003, p. 1388) found thatdifferent iron ore types react differently duringlime assimilation, depending on surfacemorphology as well as chemical composition.The standard method for determining whetheran ore type is suitable for use in the sinterplant is the penetration test. A 5 mm limetablet is placed on top of a 10 mm iron oretablet, both pressed from powder. The ore andlime tablets are heated in a furnace and thelime penetrates into the ore tablet. The depthof penetration is used as an indication of thereactivity of the ore type.

Steelmaking

Steel is produced by three processes, of whichthe blast furnace (BF) and basic oxygenfurnace (BOF) combination has been mostpopular since the nineteenth century. Liquidpig iron (hot metal) produced by the BF isused as feed to the BOF. The solids feed to theBF include iron oxide, metallurgical coke,sintered ore, manganese ore, and dolomite(Buschow, 2001, pp. 4293–4296).

Development of a method for evaluating rawmaterials for use in iron ore sinter in termsof lime assimilationby W. Ferreira*, R. Cromarty*, and J. de Villiers*Paper written on project work carried out in partial fulfilment of B. Eng (Metallurgy)

SynopsisSteel is produced in a basic oxygen furnace from hot metal obtainedfrom a blast furnace. A sintered iron ore with good high-temperature properties (strength and permeability) should be usedas feed to the blast furnace. The quality of this sintered ore dependson the reactivity of the iron ore used as feed to the sinter plantduring the lime assimilation step in the sintering process. Thepenetration test is the standard method for evaluating the reactivityof iron ore with lime. It is, however, difficult to determine the exactdepth of penetration from the standard test. A new test method isproposed that allows automatic evaluation of iron ores in terms oflime assimilation with increasing temperature. A comparison of thecoefficients of variation for the new and standard methods for eachore type demonstrates that the results of the new test are morereproducible and more precise than those of the standard method.The test is also less time-consuming and easier to implement.

Keywordsiron ore, sintering, properties, lime assimilation

* Department of Materials Scnience andMetallurgical Engineering, University of Pretoria,Pretoria, South Africa.



Development of a method for evaluating raw materials for use in iron ore sinter

Blast furnace

Solid material is fed into the BF from the top while hot gas isblown from the bottom upwards, as can be seen in Figure 1.This gas, which consists of a mixture of air and pure oxygen,is used to improve combustion efficiency to reduce the ironore with metallurgical coke. Molten metal with a slag layer ontop collects at the bottom of the BF and the liquid metal istapped to be used as feed in the BOF.

In order to obtain sufficient contact between the gas andsolid particles in the BF, a permeable burden is required toallow a high and uniform gas flow rate (Barker et al (2006,p.1393). The iron feed material should not contain excessfines, since the fines will be lost to the top gas.

Barker et al (2006, p.1393) also mention that sinterstrength is an important characteristic since the sinter usedas feed to the BF will be subjected to stockpiling, handling,and transportation. During all of these steps the sinter shouldnot degrade and produce fines, which will be blown out of theBF with the top gas. To ensure that the iron ore feed hassufficient permeability, strength, and correct size, the fine oreis sintered.

Iron ore sintering

Iron ore sintering can be described as the controlled burningof a fuel mixed with iron ore (Barker et al, 2006, p. 1393).The process converts natural fine iron ore material, screenediron ore fines, coke, and lime into a fused clinker-likeaggregate that can be effectively used in the BF. Iron orefines are mixed with 5% anthracite, which acts as fuel and isconveyed in the sinter through the process. The mixture ofsinter material is fed onto a moving grate, and at the feed endgas burners are used to ignite the top of the bed. As themixture moves forward, the combustion zone progressesdownward due to air flow through the permeable bed. Thisresults in a temperature profile through the thickness of thebed. Temperatures as high as 1300°C to 1480°C are reachedat the hottest spot in the bed, causing the particles to fusetogether into porous clinker. At the discharge end, sintering

would have occurred through the thickness of the bed. Thematerial is then crushed and screened. The oversize materialis sent to the BF stockpiling yard and the undersize isreturned to the sinter process.

Sintering of the ore has further benefits to the BFoperation since the flux is incorporated into the sinter mixinstead of being added separately to the BF feed, and it alsoproduces a sized sinter as feedstock with better high-temperature properties (Barker et al, 2006, p. 1394).

Lime assimilation during iron ore sintering

With the world-wide increasing demand for iron ore, morelower-quality grades of iron ore are being produced than inthe past. As sinter plants have to make use of the ore athand, it is important to have a quick pre-production test toevaluate whether an ore will be suitable for use in the BFafter sintering.

The iron ore sintering process consists of three stages.The first stage is the heating of the burden before meltformation takes place. The second stage is the primary meltformation of pseudo-particles with an adhering fines layer,and the last stage is the assimilation of lime with the nucleusore (Hida and Nosaka, 2007, p. 103). Various tests are usedto determine the properties of the ore during each step. In thisproject the focus will be on testing the properties obtainedduring the last stage of the sintering process.

Three tests currently are used in industry to evaluate ironore in terms of lime assimilation. These include the smallpacked bed sintering test and the penetration test asdescribed by Higuchi et al (2004, pp. 1385-1386), and alsothe variation of the small packed bed sintering test (Hida andNosaka, 2007, p. 104).

Since the penetration test makes use of a tablet of ironore and a second tablet of a combination of lime and iron ore,the conditions in this method are similar to those found inthe sintering process. This method was therefore chosen asthe standard to be used to evaluate the new developed testmethod, and is hence the only method described here in moredetail.

The standard penetration test

The standard penetration test is used to evaluate the meltingof fines into the adhering layer of pseudo-particles duringsintering. The test uses two pressed tablets: an ore tabletconsisting of an equivalent mass ratio of size fractions -0.25mm and +0.25–0.5 mm and a primary melt tablet consistingof 26 mass% CaO-FeO). The ore of the two size fractions ismixed and pressed into a tablet 10 mm in diameter and 5 mmin height, while the primary melt reagents are hand-mixedfor 20 minutes and pressed into a tablet with diameter andheight of 5 mm. Both tablets are produced using an ironmould at a pressure of 0.314 kN.

The primary melt tablet is placed on top of the ore tabletin the centre of a nickel vessel. The sample is heated fromambient temperature to 800°C within 3 minutes, then from800°C to 1300°C in 2.5 minutes. The temperature ismaintained at 1300°C for 2 minutes, after which the furnaceis then set to cool over 10 minutes.

A vertical section of the sample is mounted and polished.Macro images at 5× magnification are used to measure thedepth of penetration, which is defined as the distance from

▲


Figure 1—Blast furnace feed materials (Buschow et al. 2001, p. 4296)

the top rim of the tablet to the tip of the reaction zone(Higuchi et al, 2004, p. 1386). Figure 2 shows the heatedsample with the penetration depth indicated by the arrows.This measurement is used to compare the reactivity ofdifferent ore types (Higuchi and Okazaki, 2004, p. 432).

New length reducibility test method

Industry requires an easy test method to determine theassimilation of iron ore with lime during sintering, as there isa need to determine the reactivity of a given type of iron orewith lime and the subsequent porosity and sinter strength.Since the grade of iron ore being mined is decreasing due tothe increase in demand, it is becoming even more importantto establish whether sintering of a particular ore will result ina suitable product for use in the blast furnace.

A disadvantage of the penetration test is that there is nocertainty on the measured depth of penetration, sincepenetration invariably does not proceed in the perfectlyhemispherical fashion described in the literature. The depthoften varies across the section and there is uncertaintywhether to measure the deepest, the shallowest, or averagedepth of penetration. Furthermore, the samples sometimesreact to such an extent that a completely molten mass isobtained, and no conclusion can be drawn from such results.This test is also not considered to be fully reproducible, sincemeasurement of the penetration depth is done manually.

The alternative test method, which is aimed at achievingconditions similar to those found in the sintering process,makes use of a single cylindrical sample consisting of amixture of 26 mass% lime and 74 mass% iron ore which isplaced in the furnace.

Hewakandamby et al. (2013, p. 456) described a testmethod to determine the behaviour of ash from biomass andcoal at elevated temperatures. A cylindrical sample is heatedin an ash fusion furnace, and images of the sample obtainedat fixed temperature intervals are used to determine thetemperatures at certain states, for example the softeningtemperature.

In the new length reproducibility test method used here,the furnace is equipped with a digital camera to obtainimages of the sample with increasing temperature. MatlabTM

is then used to convert the digital images to a greyscaleimage to find a matrix equivalent to the size of each sample.As the samples are heated and started to react, the height ofthe samples decreases and can be measured as a function oftemperature by the software program and a plot of heightfraction versus temperature obtained. The samples are largeenough to allow most samples to be submitted subsequentlyfor additional testing such as sinter strength tests or opticalexamination to determine the porosity.

Hewakandamby et al (2013, p. 454) concluded that thenormal ash fusion test used in the past does not produceconsistent results as it is based on visual inspection and theresults are therefore dependent on the subjective judgementof the person carrying out the experiment. The original ashfusion test as applied to iron ore sintering cannot, therefore,be regarded as inherently reproducible. The method describedhere is an automated one, with a higher precision than thatof the original visually evaluated ash fusion test.


The validity of the new length reducibility test method wasevaluated by comparing the results with those from astandard penetration test, using various ore types.

Three ore types were tested: two haematite ores, a typicalNorthern Cape ore in South Africa, a West-African ore, and agoethitic iron ore from Marra Mamba in Australia. All threeore types were tested under the same conditions using bothmethods, including the quantities of ore and lime used toproduce the samples. Eight samples of each ore type weretested in each of the two methods.

Standard test procedure

The standard penetration test procedure used was based onthe conditions given by Higuchi et al. (2004, pp. 1385-1386). The test was carried out in an infrared furnace, whichhas a fast enough heating rate to simulate the sintertemperature profile.

Samples and sample preparation

All sampling methods and preparations were done accordingto ISO 10836 (ASTM part 12), which is equivalent to ASTME877 (ASTM part 12). The two tablets were placed in thefurnace simultaneously, one on top of the other. The toptablet was 5 mm in diameter and height and comprised 26mass% lime and 74 mass% Fe ore. The bottom tablet, 10 mmin diameter and 5 mm in height, consisted of equal amountsof -0.25 mm and 0.25–0.5 mm pure iron ore.

Given the density of iron ore fines of 5 g/cm3 (Fe oreMSDS), the total mass of iron ore needed for a 10 mm oretablet was 1.96 g. In order to limit the number of variables inthe test work, all the samples of each ore type were preparedfrom a single batch mixture. The total mass of iron orerequired for the samples for each ore type was milled to therequired particle size distribution (equal amounts of -0.5 +0.25 mm and -0.25 mm) and compressed at 0.314 kNin a hand-operated hydraulic press.

The melt tablet consisted of 0.113 g lime and 0.322 g ironore (5x5 mm site consisting out of 26% lime and 74% ironore, The ore and lime for all of the samples for each ore typewas hand-mixed in a single batch and pressed into a greenpellet. When the correct particle size distribution of equalamounts of -0.5 mm +0.25 mm and -0.25 mm, was obtained,the mixed product was pressed into pellets in a hand-operated hydraulic press at a compressive load of 0.314 kN.

The samples were heated in an infrared furnace for thestandard penetration test. The ore sample was placed in thefurnace with the 5 mm melt tablet on top of the 10 mm oretablet. The furnace was set to simulate the sinter process. Thetemperature was increased to 800°C over 3 minutes, then to



Figure 2—Vertical section through lime and iron ore pellet aftersintering (Hida and Nosaka, 2007, p. 104)


1300°C in over 2.5 minutes, and maintained at 1300°C for 2minutes. The temperature was then decreased to ambientover 10 minutes and the sample removed from the furnace.

The reacted samples were sectioned vertically through theaxial centre and mounted in resin. The bottom half of thesample was removed to allow inspection of the centre of thesample by conventional and stereo-optical microscopy. Sincethe 5× magnification of the sample surface did not allow fullinspection of the surface, the stereo microscope was used totake three successive images of the entire surface area, whichwere combined into a single collage. A distinct defect or grainwas used as reference point to allow the entire penetrationdepth to be measured from the bottom of the sample (oretablet) up to the point of maximum penetration from the top.This total was then subtracted from the original height of theore tablet to find the depth of penetration, which could beidentified as a darker uniform phase at the top of the tablet,as seen in Figure 3. The bottom granular part consisted ofiron ore particles with size fractions of equal amounts -0.25 mm and -0.5 + 0.25 mm.

Length reducibility test procedure

These tests were carried out in an ash fusion furnaceequipped with a digital camera to capture the change inheight of the various samples as a function of temperature.

Samples and sample preparation

The 10 mm diameter by 5 mm height tablets for these testsconsisted of single samples with 26 mass% CaO and 74mass% iron ore, i.e. a lime to ore ratio equivalent to thepenetration tests. The ore comprised equal amounts of -0.25mm and +0.25 -0.5 mm material, as was the case for thepenetration test. The procedure for preparing the powders,mixing them, and pressing them into green pellets wasidentical to that used in the penetration test.

Heating procedure

Up to four samples were placed simultaneously in the ashfusion furnace by balancing them on the sample carrier in thefurnace, with the camera focused on all four samples. Thefurnace was set to increase the temperature at 7°C/minuteand the computer set to take an image of the four sampleswith each 2°C increase in temperature, starting at 1150°C.

Once the furnace reached 1204°C it was switched off andallowed to cool to 400°C before the samples were removed.The temperature of 1204°C was chosen to ensure that thesamples did not melt completely and molten material leakonto the furnace tube, thereby cracking it.

Analyses of the samples

The digital images of the samples in the furnace wereanalysed using a program written in ImageJ specifically forthis project. The program requires a manual but accuraterectangular selection of each sample in the image to be drawnbased on the initial width and height of the sample. Theselection allows the program to set boundaries on wherethese measurements need to be taken, which allows it tofocus exactly on where the sample to be measured is locatedin the bigger image. The pixels in this rectangle will becounted. After this initial set-up, the image acquired at themaximum temperature is opened together with the imageimmediately preceding. The program then automaticallydetermines the difference between the two images, thresholdsthe image, converts it to a binary image, and removes anyoutliers. These steps are illustrated in Figures 4 and 5.

It can be seen that in the temperature increment from1202°C to 1204°C, only the Northern Cape ore tablet showeda measurable change in area (Figure 5). After this step theprogram makes use of a matrix to measure the number ofpixels that make up the difference in surface area measured(red pixels in Figure 5). This gives an indication of the totalreactivity of the ore type during that temperature increment.The amount of pixels counted can now be divided by theinitial width measured by the program from the rectangular

▲


Figure 3—Marra Mamba pellet with darker melt phase penetrating intogranular ore phase

Figure 5—Difference between images A and B in Figure 4 (indicated byred) as measured by the macro written in Image J

Figure 4—Typical Northern Cape ore (top left on stand) and West-African ore (top right on stand ) iron ore at 1202°C (A) and 1204°C (B) inthe ash fusion furnace

selection to determine the height change during thetemperature step. The data is displayed in a text file and canbe imported directly into Excel® to construct a graph and toquantitatively determine the results. The greater the changein height for the sample, the more reactive the ore type.

Since the reactivity analysis is done by making use of theimages as well as the written program, the samples removedfrom the furnace can be subjected either to a strength test ora porosity measurement, as no further dimensionalmeasurements need to be taken.


Eight samples of each of the three different ore types weretested by the two test methods. The average penetrationdepth measured for the penetration tests and average changein height for the length reducibility test were used to quanti-tatively compare and analyse the two test methods. Theresults are given in Table I

Standard test method and new test results

Table I shows that the trends of the results obtained for bothtest methods were similar. Both methods indicated that theWest-African ore had the lowest reactivity, followed by theNorthern Cape ore with medium reactivity, and then MarraMamba with the highest reactivity. These results show thatthe new length reducibility test can therefore be used toobtain the same type of information as the standardpenetration test used in industry.

Reproducibility of test methods

To determine the reproducibility of both methods, thecoefficient of variation was plotted for each ore type. Thecoefficient of variation, Cv, is calculated by:

[1]

where σ is the standard deviation and μ is the mean. A lowercoefficient of variation indicates a higher reproducibility andmore precise results.

Figure 6 shows that the new length reducibility test has aconsistently lower coefficient of variation than the standardpenetration test, indicating that the new method has a higherreproducibility than the standard test, potentially leading tomore precise results.

Additional results

Figure 7 indicates that the Marra Mamba and Northern Capeores start reacting at about 1185°C, while the West-Africanore starts reacting only above 1200°C. This informationcannot be obtained from the standard test.

Advantages and disadvantages

In order to determine which test method would produceresults faster and easier, the disadvantages and advantagesof both methods are listed.

Standard test

Advantages:➤ Since this test is carried out in the infrared furnace, a

high heating rate can be used, thus simulating thetemperature profile of the sintering process moreclosely

➤ The infrared furnace cools down, and the sample canbe removed, within an hour



Table I

Comparison of results

Total change in height (New Method) (mm) Total Penetration (Old Method) (mm)

Typical Northern Cape ore West-African ore Marra Mamba Typical Northern Cape ore West-African ore Marra Mamba

1 1.44 0.04 1.55 0.36 0.22 0.102 0.23 0.22 0.74 0.64 0.66 0.593 1.41 0.16 1.21 0.62 0.22 1.824 0.56 0.05 1.27 1.38 0.29 0.495 1.82 0.02 0.83 0.48 0.48 1.246 1.52 0.15 1.02 0.48 0.69 3.647 0.79 0.19 0.86 1.21 1.74 1.878 0.20 0.02 1.07 2.69 0.27 4.49Average 1.00 0.11 1.07 0.98 0.57 1.78Standard deviation 0.63 0.08 0.27 0.78 0.51 1.56σ/μ 0.63 0.77 0.25 0.80 0.89 0.88

Figure 6—Coefficient of variation for each ore type tested using bothmethods

Figure 7—Average change in height as a function of temperature forthe three ore types tested as determined by the length

Typical NorthernCape oreWest-African ore

Marra Mamba

Northern Cape ore West-African ore Marra Mamba


➤ Since the sample has to be mounted and sectioned inorder to determine the reactivity, a porositymeasurement can easily be carried out.

Disadvantages:➤ Only one sample can be tested at a time, making this a

time-consuming method.➤ All samples are mounted and vertically sectioned, and

thus are not available for strength tests➤ Working with the mould to produce the melt tablet

requires great care since the 5 mm mould is extremelydelicate and can resist hardly any pressure

➤ When placing the tablets in the tube of the infraredfurnace, balancing the small melt tablet on top of thelarger ore tablet is very difficult

➤ The depth of penetration is hard to measure. For someore types a definite penetration can be seen as auniform and darker phase penetrating into the granularphase of the ore, but this is not necessarily the case forall ore types

➤ The depth of penetration is measured by hand and thefinal value is often subjective.

New length reducibility test

Advantages:

➤ The ash fusion furnace allows testing of up to foursamples per test run. The length reducibility test istherefore a more productive test

➤ The analysis of the reactivity is done automatically by acomputer program and this removes operator subjec-tivity

➤ No vertical cross-section of the sample after the test isnecessary, since the reactivity is a function of thechange in height of the sample. The samples are thusavailable for either a strength test or a porositymeasurement

➤ The sample consists of a single tablet that is easy toproduce to the required weight and dimension specifi-cations

➤ This method does not require a 5 mm diameter tabletmade in a fragile mould.

Disadvantages:➤ The maximum temperature for these three ores could

not be set higher than 1204°C, since at 1206°C theMarra Mamba ore melted completely. The melt couldleak into the furnace tube, causing it to crack

➤ The ash fusion furnace has a very slow heating rateand cannot replicate the actual heating rate in thesinter process.

Comparison of methods

Apart from its proven greater reproducibility and precision,the length reducibility test also provides the followingadvantages:

1. The time required for sample preparation is reduced,since only one tablet is required instead of two foreach ore type and test run. Four samples can beanalysed simultaneously during a single 4-hour run,including cooling time, compared with two hours for asingle sample sample by the standard method. Thenew test method also requires only one cylindrical

tablet which is simple to reproduce. There is no needto produce two tablets, each with a different mixtureof raw materials

2. Results are obtained automatically by making use ofthe images taken with the digital camera and theimage analysis program. The results are thereforeobtained easily and the error associated with humanjudgement is avoided.

Future work and recommendations

In order to obtain international accreditation for this newmethod, the following aspects are of particular importanceaccording to ISO 17025:

➤ Reproducibility/repeatability➤ Discrimination of samples➤ Sensitivity➤ Detection threshold➤ Comparison against existing methods➤ Inter-laboratory tests.

Conclusion

In a comparison of the standard test with the newly provenlength reducibility test it was found that:

1. The new length reducibility test yields morereproducible results

2. The results obtained by the new test method are moreprecise

3. The new length reducibility test is less time-consuming and easy to implement

4. The new test can therefore be used with confidence toevaluate and quantify the reactivity of iron ore interms of lime assimilation in the sintering process.

References

ASTM Standards, Chemical Analysis of Metals; Sampling and analysis of metal

bearing ores. 1982. Part 12 E877. ASTM, Philadelphia.

BARKER, J., KOGEL, J., KRUKOWSKI, S., and TRIVEDI, N. 2006. Industrial Minerals

and Rocks - Commodities, Markets, and Uses. 7th edn. Society for Mining,

Metallurgy, and Exploration (SME), Warrendale, PA..

BUSCHOW, K., CAHN, R., FLEMINGS, M., ILSCHNER, B., KRAMER, E., and MAHAJAN, S.

2001. Encyclopedia of Materials - Science and Technology, vols. 1-11.

Elsevier, Amsterdam.

FE ORE MSD UNIVERSAL MINERALS.

HEWAKANDAMBY, B., LESTER, E., PANG, C.H., and WU, T. 2012. An automated ash

fusion test for characterisation of the behavior of ashes from biomass and

coal at elevated temperatures. Fuel, vol. 103. pp. 454–466.

HIDA, Y. and NOSAKA, N. 2007. Evaluation of iron ore fines from the viewpoint

of their metallurgical properties in the sintering process. Transactions of

the Institute of Mining and Metallurgy C, Mineral Processing and

Extractive Metallurgy, vol.116. pp. 101–107.

HIGUCHI, K., HOSOTANI, Y., OKAZAKI, J., and SHINAGAWA, K. 2003. Influence of iron

ore characteristics on penetrating behavior of melt into ore layer. ISIJ

International, vol. 43. pp. 1384–1392.

HIGUCHI, K. and OKAZAKI, J. 2004. Marra Mamba ore, its mineralogical properties

and evaluation for utilization. ISIJ International, vol. 45. pp. 432.

POVEROMO, J.J. 2005. IMAR 7: Industrial rocks and minerals, 7th ed. Society for

Mining, Metallurgy, and Exploration. 1568 pp. ◆

▲


Introduction

A high manganese recovery is vital inrendering the ferromanganese productionprocess economically viable. It is particularlycrucial in recovering manganese products from

secondary sources such as ferromanganeseslag. This industrial waste material containsan appreciable amount of residual manganesethat can be economically exploited.Ferromanganese slag is a waste product fromthe production of ferromanganese metal inblast furnaces and submerged arc furnaces.There is an estimated 53 Mt of slag dumped inSouth Africa (Parker and Loveday, 2006).Discarding the slag not only entails disposalcosts, but also poses an environmental andhealth threat, mainly due to the leaching ofmobile manganese in the slag.

The slag can be fed back into the furnace.This option, however, is not favoured due tothe build-up of alkali metals and zinc in thesmelting process. The slag can also be crushedand ground to produce cement aggregate orslag cement. The disadvantage to this,however, is that there is a danger thatmanganese can leach out, and only freshlytapped slag can be utilized. A hydrometal-lurgical route that can be adopted to recycleboth the discarded and freshly tapped slagwould be advantageous. The slag would besubjected to a water-starved digestion andleached to produce a pregnant leach solutionfrom which manganese could be recoveredusing various methods, including precipitationand electrowinning, and a residue that can besold as an additive for Portland cement.

The research focus was therefore aimed atdeveloping economically and technically viableprocesses to produce a pregnant leach solutionwith a high manganese content and lowimpurity concentration. The pregnant leachsolution could be utilized to produce saleableproducts that meet the quality requirements, orproducts containing manganese in a lowoxidation state that could be fed back into the

The recovery of manganese products fromferromanganese slag using ahydrometallurgical route

by S.J. Baumgartner* and D.R. Groot†

Paper written on project work carried out in partial fulfilment of B. Eng. (Metallurgical Engineering)

SynopsisThe ferromanganese industry is under pressure to deal with the slagarising from the production of ferromanganese, which is discarded inlandfills or slag heaps. This material poses an environmental and healthrisk to surrounding ecosystems and communities, and disposal costs areincreasing. Ferromanganese slag contains an appreciable amount ofresidual manganese metal, which can be exploited. Previous work hasshown that the slag can be leached fully, while rejecting the silica to aresidue. The methods that were investigated to recover manganese fromthe leach solution included hydroxide precipitation to upgrade the leachsolution followed by manganese carbonate precipitation to produce apure manganese carbonate product or a manganese carbonate furnacefeed material, which would be recycled to increase manganese recoveriesin the production of ferromanganese. In addition, electrowinning ofelectrolytic manganese dioxide from the leach solution was studied. Themethods were compared in terms of selectivity, costs, and productquality. Co-recovery of the leach residue, which is a potential cementadditive, is discussed.

Among the methods investigated to upgrade the pregnant leachsolution, hydroxide precipitation utilizing ammonia to adjust the pHappears to be the most effective in removing major impurities such asiron, aluminium, and silica to less than 1 ppm. The manganesecarbonate and impure manganese carbonate furnace feed products metquality specifications. However, although the production of thesematerials was technically viable, the large amounts of base reagent thatwere required to raise the pH, and the associated high operating costs,rendered the process uneconomic.

An optimization study was therefore carried out with the primaryobjective to determine the ideal acid amount to be utilized in the water-starved digestion stage, thereby decreasing acid and base consumptionwhile optimizing the quality of the pregnant leach solution, andproducing a leach residue that contained <1% Mn. The outcome was aneconomically viable process. Additional benefits included an increase inthe manganese content of the impure manganese carbonate furnace feedmaterial, and a substantial reduction in the dilution of the pregnantleach solution, thereby maintaining high manganese concentrations thatrendered the solution viable for electrowinning of electrolyticmanganese dioxide, the production of which yielded a current efficiencyof 74%.

Keywordsenvironment, ferromanganese slag, manganese products, water-starveddigestion, precipitation.

* Palabora Copper Ltd, South Africa.† Department of Materials Science and Metallurgical

Engineering, University of Pretoria, Pretoria, SouthAfrica.



The recovery of manganese products from ferromanganese slag

ferromanganese production process as a sweetener, thuscircumventing the environmental and health impacts ofprocessing ferromanganese slag. In addition, a saleablebyproduct such as a cement additive would be produced.

Leaching of ferromanganese slag

Various authors have undertaken prior investigations into theleaching of ferromanganese slag. The leaching processesdiscussed include ferric chloride in the presence of sucrose,carbamate leaching, and the quick leach (also known as thewater starved digestion method) with concentrated sulphuricacid (98–99%).

Naganoor et al. (2000) used the ferric chloride route toleach slag. It was deduced that roasting prior to leaching wasessential to ensure that the manganese was converted into asoluble form. A manganese recovery of 82% was achieved in2 hours at a temperature of 80˚C. The oxidation of sucrose,however, increased the rate of leaching, yielding an 86%recovery in 1 hour. The pregnant leach solution can befurther treated to produce electrolytic manganese dioxide(EMD) and electrolytic manganese metal (EMM). Mcintoshand Baglin (1992) investigated the feasibility of usingammonium carbamate as a reagent to recover manganesefrom slag. A recovery of 99% was achieved in 3 hours at atemperature of 65˚C. The manganese carbamate can be addedas a sweetener in the steelmaking industry or can be furtherpurified.

The slag that was used to produce the pregnant leachsolution in previous investigations was first crushed andmilled, then premixed with deionized water followed by waterstarving (digesting) the slurry in concentrated sulphuric acid(98%) (Groot et al., 2013). A solidified cake was produced,which was left to age for 24 hours; this resulted in anincrease in manganese recovery. Thereafter, the cake waswater-leached in a 1–3 stage water leach, recovering between75–95% of the manganese into solution. These results,however, were dependent on the number of leaching stages.

Recovery of manganese as manganese carbonate

Zhang et al. (2010) investigated carbonate precipitation ofmanganese from a pregnant leach solution at 60°C usingsodium carbonate (Na2CO3 ) at different pH values. Theexperiments were conducted under ambient conditions, butwith slow agitation to reduce oxidation by air. It wasobserved that at pH>7.5 a 90% manganese recovery wasobtained with co-precipitation of calcium and magnesium atrecoveries of 43% and 13%, respectively. At pH>8.5, 99.5%of the manganese was recovered, with co-precipitation of97% of the calcium. According to Zhang et al. (2010) theprecipitation of manganese carbonate is thermodynamicallyfavoured over magnesium carbonate and (but to a lesserdegree) calcium carbonate. This is attributed to the smallerlog K(Ca/Mn) value measured in comparison to the theoreticalK value. This is an important consideration when attemptingto recover manganese carbonate that meets specificationsfrom a solution that contains high magnesium and lowcalcium concentrations. The results obtained by Zhang et al.suggested that the recovery of manganese via a carbonateprecipitation yielded a good quality product. This, however,was dependent on the manganese to magnesium ratio.

An alternative process to produce manganese carbonate(MnCO3) from a manganese solution was developed by INCOfor the recovery of metal values from ocean-floor nodules inthe presence of 3–8% NH3 and 1–6% CO2. The reagent actsboth as a lixiviant and as a precipitant for manganese asMnCO3, which is recovered in the residue; nickel, cobalt, andcopper are solubilized as stable ammines (Illis and Brandt,1975; Zhang and Cheng., 2007). Kono et al. (1986)conducted similar studies into the recovery of manganesefrom nodules using a sulphurous acid leach, with precipi-tation of manganese as MnCO3 in the presence of (NH4)2CO3while copper, nickel, and cobalt remained in the solution asammine complexes.

Production of electrolytic manganese dioxide (EMD)

Electrolytic manganese dioxide (EMD) is a vital ingredient inLeclanche-type dry cells. Owing to modern advances in theelectronics industry, which have necessitated a greatercapacity for manganese dioxide production, naturalmanganese dioxide (NMD) has been replaced by electrolyticmanganese dioxide (EMD) (Rethinaraj and Visvanathan,1993).

The half-cell reactions are given below:[1]

[2]

[3]

The current efficiency at the anode is affected by thefollowing parasitic reactions (Te Riele, 1983):

The evolution of oxygen:

[4]

The oxidation of Mn2+ to Mn3+, which diffuses into theelectrolyte solution:

[5]

Mn3+ is an unstable ion that disproportionates to produceMn2+ and MnO2, which forms sludge:

[6]

An additional parasitic reaction that can occur if there isan appreciable amount of iron in solution, and which isresponsible for decreasing current efficiencies and contami-nating the final EMD product, is given by:

[7]

Table I shows the effect that different concentrations ofiron in solution have on the final product (Te Riele, 1983).

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Table I

Effect of iron content of solution on iron contami-nation in the EMD (Te Riele, 1983)

Fe in electrolyte (ppm) Fe in final EMD (%)

0–10 0.02110–50 0.09845–80 0.14960 –90 0.19

There are several factors that affect the quality of EMDproduced, one of which is the acid concentration of theelectrolyte. A high acid concentration (>50 g H2SO4 per litre)has a detrimental effect on the amount of battery-active EMDplated. In addition, the morphology is affected, resulting in aless compact deposit, and current efficiency is decreased bythe formation of Mn3+ which is promoted by the higher acidconcentration (Te Riele, 1983). Temperature is anothercritical factor. A temperature above 90°C results in a desirablecurrent efficiency, which ranges between 90–97%, providedthe system is operated at 100 A/m2 (Te Riele, 1983;Rethinaraj and Visvanathan, 1992). At lower temperatures,however, the current efficiency will be variable depending onthe type of materials utilized for the electrodes, and the cellvoltage will increase, thereby increasing energy consumption.In addition, lower temperatures result in shorter dischargetimes on the EMD battery cell (Te Riele, 1983).


The flow diagram in Figure 1 summarizes the experimentalroutes that were investigated to determine the mosttechnically and economically viable methods of producingvarious manganese products.


Materials

A South African ferromanganese producer in KwaZulu-Natalsupplied the ferromanganese slag utilized in previous investi-gations. The chemical and phase compositions of the slagwere determined using X-ray fluorescence spectrometry(XRF) and X-ray diffraction spectrometry (XRD), which wereperformed by Scrooby's Laboratory Service and theDepartment of Geology at the University of Pretoria.

Chemical composition of the ferromanganese slag

The typical composition of the slag, determined by XRFanalysis, is given in Table II.

The major elements were manganese, silicon, andcalcium. The iron content was low, which was ideal as ironresults in difficulties in the downstream purification stages.The major phases present in the material includedglaucochroite, manganosite, and gehlenite as shown in Table III.

The manganese content of the slag (approx. 30%) wassufficient for the investigation to proceed further.

All reagents were of analytical grade, and supplied byMerck Millipore, and distilled water was used throughout. Allglass and metal ware was decontaminated with detergent andrinsed thoroughly with distilled water.

Sample preparation

The work was performed on 100 g ferromanganese slagsamples that had been pre-crushed to 45 mm. The materialwas further reduced in size using a gyratory crusher, thenmilled to a cut size of 600 μm utilizing ball milling, and dry-screened.

Purification of the manganese pregnant leachsolution

The objective of the purification stage was to obtain apregnant leach solution (PLS) that contained high manganeseand low impurities concentrations, from which manganeseproducts that meet the quality specifications could beproduced.



Figure 1—Flow diagram illustrating the basic process routesinvestigated for the recovery of manganese products

Table II

Typical composition of the ferromanganese slag

Element Mass %

C 1Mn 28P 0.01Si 21Cr <0.005Ni <0.005Cu 0.006Al <0.005V <0.005Ti 0.14Co <0.005Ca 19Mg 3Fe 1

Table III

Phase analysis of the ferromanganese slag (Groot et al., 2013)

Phase Chemical formula Mass %

Glaucochroite CaMnSiO4 55.00 – 65.10Manganosite MnO 2.28 – 4.50Gehlenite Ca2Al[AlSiO7] 4.00 - 8.49Monticellite CaMgSiO4 0.00 – 2.00Quartz SiO2 0.00 – 5.10Amorphous - 21.00 – 30.0


The reagents used to increase the pH for impurityremoval included sodium hydroxide (NaOH) or ammonia(NH3), which were employed in separate purification investi-gations. The oxidant was hydrogen peroxide (34% H2O2), tooxidize Fe2+ to Fe3+ so that iron was precipitated within a pHwindow of 3–5.5.

A PLS sample size of 100 ml was used throughoutexperimental runs. Before base reagent was added, 34%H2O2 was added dropwise until the solution turned from pinkto a mustard colour. For purification using NaOH, 24 g ofNaOH pellets were weighed, and dissolved in 200 ml distilledwater to obtain a solution of 120 g/l of NaOH, which wasadded dropwise to the solution to increase the pH from lessthan 0 to approximately 5.6. The solution was kept at roomtemperature and continuously stirred.

A similar procedure was followed for the experiments thatinvolved the upgrading of the PLS using 25% ammoniasolution, which was added dropwise to increase the pH toapproximately 5.6.

The upgraded PLS was analysed for major elements byinductively coupled plasma – optical emission spectrometry(ICP-OES).

Production of manganese products

Precipitation of manganese carbonate

A 100 ml sample of PLS was used for the investigationsutilizing Na2CO3 or (NH4)2CO3 as base reagents. The Na2CO3,at 150 g/l, was added dropwise to the PLS; the (NH4)2CO3was added to the PLS in its solid form. Approximately 15 g of(NH4)2CO3 was added to adjust the pH from approximately5.5 to 8.5, the pH range in which >99% of the manganesewas recovered. The system was open to the atmosphere, andthe procedure was conducted at room temperature (approx.25°C) while the solution was agitated. Once precipitation wascomplete, the product was filtered, washed with distilledwater, and dried at 30°C.

The manganese and impurities contents of the barrensolution obtained after precipitation were determined by ICP-OES. The precipitate was analysed using XRF and XRD.

Production of a furnace feed material (impure MnCO3)

The apparatus and reagents used to produce the furnace feedmaterial were the same as for the purification of the PLS andthe pure MnCO3 product, except no filtration step wasincluded. The product was dried in an oven at 30°C.

Electrowinning of electrolytic manganese dioxide(EMD)

The electrodes were graphite rods with a diameter of 7.5 mm.A current density of 100 A/m2 was adopted, with a platingtime of 72 hours to allow a deposit thickness of approxi-mately 2.4 mm, which would produce a mass of EMD >3 g,and a temperature of 90°C. The electrolyte volume was 1.1 litres.

The set-up of the electrowinning experiment for theproduction of EMD is shown in Figure 2.

The EMD was analysed for gamma and alpha manganesedioxide (MnO2) by XRD, and an elemental analysis wascarried out by XRF. The manganese content of the solutionafter electrowinning was analysed using ICP-OES.


Pregnant leach solution and residue

To produce the pregnant leach solution and residue the milledslag material was subjected to a water starved digestion andthereafter water-leached for two hours. Table IV shows amass balance giving the general recoveries of manganese andimpurity elements to the PLS and residue respectively. Themass balance was based on a 200 g slag sample.

The discrepancies in the percentage recovery of elementswere probably due to the inhomogeneity of the ferroman-ganese slag sample. It was assumed that the chemicalcomposition of the slag was as given in Table IV.

Purification of the pregnant leach solution (PLS)

Hydroxide purification

It was essential to obtain a purified PLS in order to produceMnCO3 and EMD that adhered to the chemical specifications.The PLS had to contain >25 g/l Mn for the production ofMnCO3, and 27–66 g/l Mn for the production of EMD (TeRiele, 1983). In addition, the PLS should adhere to thespecifications in Table V.

During the addition of the base reagent, oxygen wasbubbled through the solution at 60°C to oxidize iron from theFe2+ to the Fe3+ state, thus allowing iron to precipitate as a

▲


Figure 2—Experimental set-up for the production of EMD

Table IV

Mass balance and recovery of elements to the leachresidue and PLS

Element Slag (feed), g Residue (g) Leach solution (g) % Recovery

Mn 55.4 9.344 31.7 74.1P 0.02 - 0.115 >100Si 40.8 13.120 0.034 32.2Cr - - 0.025 >100Ni - - - -Cu 0.012 - 2.927 >100Al - 0.346 - >100V - - - -Ti 0.274 0.057 - 20.8Co - - - -Ca 38.2 21.71 0.367 57.8Mg 6.34 1.872 6.307 >100Fe 1.86 0.330 1.205 82.5

hydroxide below a pH of approximately 5.6. However, thiswas ineffective due to the slow kinetics of dissolution ofoxygen in the leach solution, and was further hindered by thesolution pH < 3 (Zhang and Cheng, 2007). The oxygen wasreplaced with hydrogen peroxide, which was effective inoxidizing iron to the ferric state; therefore, all test work wascarried out using hydrogen peroxide as the oxidant at roomtemperature where applicable.

A window of optimal precipitation of the majorimpurities, including Fe, Al, and Si, lay in a pH range of 2.5and 5.6–6. Complete precipitation of impurities occurredbetween pH 5.5–6. From the results in Table VI, both NaOHand NH3 were effective reagents in removing impurities fromsolution. However, NH3 appeared to be the more effective,resulting in a purified PLS with Fe, Al, and Si contents of lessthan 1 ppm, in contrast to NaOH where the solution

contained 604 ppm Fe, 8 ppm Si, and 140 ppm Al afterprecipitation.

Manganese products

Precipitation of manganese carbonate

Manganese precipitated from the purified PLS at pH >6, andmanganese recovery reached 99.8% and 98.9% using(NH4)2CO3 and Na2CO3, respectively, at a pH >8.5. Theresults of the ICP-OES analyses before and after carbonateprecipitation are given in Table VII. The results agreed withthose obtained by Zhang et al. (2010), who achieved amanganese recovery of >99.5% pH >8.5.

The results in Table VIII indicate that using either(NH4)2CO3 or Na2CO3 as the precipitating reagent resulted ina MnCO3 product that meets the quality specifications, withthe exception of the calcium content, which exceeded themaximum allowable amount. This finding is in agreementwith Zhang et al. (2010), who argued that carbonate precipi-tation is more selective for magnesium than calcium, due tothe thermodynamics favouring the co-precipitation ofcalcium, which is attributed to the low log K(Ca/Mn) valuedetermined in comparison to the theoretical K value.

The lower MnCO3 yields from Tests 1, 3, and 5 were dueto the co-precipitation of additional phases such as Na2CO3and (NH4)2Mg(SO4)2.6H2O. These phases, however, arewater-soluble; therefore the product was washed with waterto yield products that contained >92% MnCO3.



Table V

Standard specifications for a viable manganese PLS

Typical PLS specifications, ppm

Ni 1Co 0.3Fe 15Si 10Cu 5Zn 10

Table VI

Element concentrations in the PLS before and after hydroxide precipitation with NH3 or NaOH

Species Units Reagent

NaOH NH3

PLS pH ~-0.3 PLS pH ~5.6 PLS pH ~-0.3 PLS pH ~5.6

Mn g/L 36 19 24 15Cr ppm 40 <1 20 <1Fe ppm 1568 17.8 842 <1K ppm 117 4 240 262Mg g/L 7 3 4 2Al ppm - 56 2038 <1Al g/L 3 - - -Na ppm 366 - 115 86Na g/L - 56 - -Ni ppm 21 7 11 5Si ppm 157 <1 <1 <1Zn ppm 11.3 <1 9 <1

Table VII

Chemical composition of the PLS before and after carbonate precipitation

Species Unit Reagent

Na2CO3 (NH4)2CO3

PLS pH ~5.6 PLS pH ~8.5 PLS pH ~5.6 PLS pH ~8.5

Mn g/L 18.65 - 14.76 -Mn ppm - 650.3 - 4.8Ca ppm 209.1 10.8 181.6 14.1Fe ppm 17.8 9.4 <1 <1Mg ppm 3437.50 643 2321.69 1608.70


A disadvantage of using Na2CO3 as a base reagent in thissystem was that between pH 7.3–7.5 buffering appeared tohave occurred, due to the buffering action of the bicarbonatethat formed from the dissociation of carbonic acid. Toovercome this, a stoichiometric amount of NaOH was addedto increase the pH to approximately 8.5.

After the addition of the NaOH the pH was raised to >8.5,resulting in 98.9% Mn recovery, in comparison to 96% Mnrecovery at pH 7.5.

Furnace feed material

The production of the MnCO3 furnace feed material involvedthe precipitation of impurities followed by the precipitation ofmanganese as MnCO3, without a filtration step.

As in the production of pure MnCO3, the manganeserecovery reached >99% using (NH4)2CO3 or Na2CO3 as thebase reagent at pH >8.5. The XRD and XRF results are givenin Tables IX and X.

These results suggest that using NaOH and Na2CO3 orNH3 and (NH4)2CO3 resulted in a product that can potentiallybe fed into a furnace (>35% Mn), as shown in Tests 1 and 3in Table X. The MnCO3 furnace feed produced by the NaOHand Na2CO3 precipitation method contained appreciableamounts of sodium (approx. 13 %), which would bedetrimental to furnace operations because of the build-up ofalkali metals in the furnace, which may lead to explosionsand refractory attack. With the use of NH3 and (NH4)2CO3,approximately 45% of the product contained ammoniumphases, which would pose a problem to furnace operationsowing to the dissociation of the ammonium phases to NH3and the further dissociation of ammonia to H2 and N2 attemperatures above 400°C.

The ammonium and sodium phases are water-soluble; torid the material of these phases a water wash step wasemployed in order to render the product viable as a furnacefeed.

Economic analysis

To determine the net present values (NPVs), it was decided toutilize past and present production and plant history of thesupplier of the slag. In addition, the capex and opex valueswere calculated based on 2013 expenditures. The NPVs forthe products investigated were negative, due to the high acidcontent of the PLS, which required a large base reagent inputto neutralize the solution which in turn increased theoperating costs.

Optimization study

An optimization study was undertaken with the primaryobjective of decreasing the base reagent addition required toneutralize the PLS by investigating the ideal acid amount tobe used. The benefits of this included a reduced acid andbase reagent consumption, which resulted in a reduction in

▲


Table VIII

Chemical composition of the MnCO3 produced and the standard specifications (www.alibaba, 2013)

Parameter Standard specification (%) Test 1 (%) Test 2 (%) Test 3 (%) Test 4 (%) Test 5 (%)

MnCO3 >92 56.48 98.7 94.15 98.1 79.54Mn >43 44 43.5 42.6 43 44Iron as Fe 0.08% max. 0.31 <0.01 1.02 0.93 0.02Nitric acid-insoluble 0.1% max. - - - - -Chloride as Cl 0.1% max. - - - - -Ca 0.1% max. 1.02 1.27 2.34 1.86 1.91Sulphate as SO4 1% max. - - - - -

Table IX

XRD results for the MnCO3 furnace feed materialproduced via NaOH and Na2CO3, or NH3 and(NH4)2CO3 precipitation

Test Reagent Phase Mass %

1 NaOH and Na2CO3 MnCO3 n.d.

Na2SO4 n.d.

(Ca,Mn)(Ca,Mn,Fe,Mg)(CO3)2 n.d.

2 NaOH and Na2CO3 MnCO3 60.26

Na2SO4 39.74

3 NH3 and (NH4)2CO3 MnCO3 56.48

(NH4)2Mg(H2O)6(SO4)2 35.88

NH4SO4 7.64

Table X

Standard specifications (Gous, 2013) and results oftests carried out to produce MnCO3 furnace feed

Precipitation method

NaOH and NH3 and

Na2CO3 (NH4)2CO3

Parameter Specification (%) Test 1 (%) Test 2 (%) Test 3 (%)

MnCO3 - - 60.26 56.48

Mn >35 44 27.7 36.2

Fe - 0.31 1.22 0.9

Na <0.8 # 12.9 *

K <0.9 * * *

Zn <0.07 Trace Trace Trace amounts amounts amounts

# Not analysed for * Element not contained in product

n.d. A quantitative analysis could not be carried out

operating costs, thereby resulting in a positive net profitvalue and rendering the production processes economicallyviable. In addition, the technical viability was improved asthe amounts of impurities in the leach solution were reduced,as most were precipitated and reported to the residue. Thiswas beneficial for the production of the impure MnCO3furnace feed material, which had a higher manganese contentand lower impurities, thereby yielding a product with a highMn:Fe mass ratio, which was ideal. In addition, due to thereduction in base reagent consumption, dilution of the PLSwas reduced and a high manganese concentration wasmaintained. This was beneficial for the electrowinning ofEMD, since high manganese concentrations (>27 g/l) wererequired. The results of the study are summarized in Figures 3 and 4.

From the results the ideal acid addition to produce a PLSwith high manganese and low impurities content appears tobe 40 ml; the pH at this amount is approximately 4.3 and anappreciable amount of impurities had precipitated out, as isevident in Figure 4. The concentrations of Al, Fe, and Siobtained were 5 ppm, 684 ppm, and 39 ppm respectively, incontrast to an acid addition of 100 ml, which resulted in Al,Fe, and Si concentrations of 4 g/l, 1.6 g/l, and 157 ppm. Thisreduced the Al, Fe, and Si contents of the PLS by 99.88%,42.8% and 75%, respectively, compared with the originalwater-starved digestion acid amount of 100 ml. Themanganese concentrations at 40 ml acid addition were >29 g/l, which rendered the leach solution amenable to theproduction of quality manganese products and was suited forthe electrowinning of EMD.

The manganese concentration appeared to decrease withacid additions of less than 50 ml. This was attributed toincomplete reaction of the slag, as less than 50 ml is belowthe stoichiometric amount, and thus the silicate matricescould not be completely destroyed. The slight decrease inmanganese concentration between 40–60 ml acid wasprobably due to material variance; this is further substan-tiated by the results given in Table XI, which clearly illustratethe difference between the tests conducted with 50 ml acidunder identical conditions and with the same slag sample.The manganese concentrations obtained in test 1 and test 2differed by 6 g/l, and variances existed between the majorimpurity elements also.

A countercurrent three-stage water wash was employedon the material produced after the slag material was digestedusing a 40 ml acid addition. It was vital that the manganesecontent of the residue was brought below 1% for it to be aviable cement additive. Figure 7 illustrates the extent towhich manganese was leached from the material using 40 mlof acid.

From the results in Figure 5 it is evident that the three-stage water wash was able to reduce the manganese contentof the residue to approximately 2.7 %, which did not meet thespecification of < 1% manganese, thus a 40 ml acid additionwas not ideal. A further study was therefore carried out todetermine the optimum acid addition required to render theresidue suitable for a quality cement additive containing <1% Mn.

From the results in Figure 6 it appears that an acidaddition of 50—55 ml would be ideal, and would yield amanganese content of less than 1% in the residue while stillproducing a pregnant leach solution of pH 3.5–4 (Figure 3),from which products with a high manganese content can beproduced.



Figure 3—pH of the pregnant leach solution as a function of acidaddition

Figure 4—Effect of acid addition on the concentrations of Mn, Mg, Al,Fe, and Si in the PLS

Table XI

Example of elemental variance in the PLS at 50 mlacid volume in the water-starved digestion

Acid volume 50mlElement Test 1 Test 2

Mn 30 g/L 36 g/LAl 313 ppm 359 ppmFe 917 ppm 883 ppmMg 6 g/L 7 g/L

Figure 5—Manganese concentration as a function of the number ofwater wash stages carried out on the cake produced with 40 ml acid inthe water-starved digestion


Production of a furnace feed material

The pH of the PLS utilized to produce the impure manganesecarbonate furnace feed product was approximately 4.3. Themanganese and impurities contents are given in Table XII.

Since reducing the acid amount from 100 ml to 40 mlresulted in an increase in pH to 4.3, it was decided to addNa2CO3 or (NH4)2CO3 to increase the pH to >8 in order toprecipitate manganese as MnCO3. However, phases such as(NH4)2Mg(SO4)2.6H2O and Na2SO4 co-precipitated andcontaminated the product, rendering it unsuitable as afurnace feed. The product was therefore subjected to awashing step with distilled water, as these phases are water-soluble, in an attempt to produce a viable furnace feed. Apure manganese carbonate (>92%) resulted; the XRFanalysis is given in Table XIII. The resulting manganesecarbonate precipitates yielded >95% MnCO3, with >45% Mncontent, and contained less than 9% impurities, of which ironconstituted 1–1.2%. Manganese and iron co-precipitated at apH of approximately 7 when iron was not oxidized andcarbonate was added to the PLS (Pakarinen, 2011). Asshown by Figure 7, the formation of FeCO3 is impossible dueto the high redox potential; however, iron will most likelyprecipitate as ferric oxohydroxide (FeO(OH)).

It is evident from the results that the composition of thefurnace feed material was improved by increasing the pH.Furthermore, the water-soluble phases were effectivelyremoved by washing with water. It can be concluded that theprocess conditions resulted in a technically viable product.

The XRF analyses in Table XIII indicate that a productwith high manganese content and high Mn:Fe ratio ofapproximately 48:1 was obtained, which acceptable as anadditive in ferromanganese production.

Economic outcome

To render the production of the various manganese productsviable, changes were made to the water starved digestionprocess after an optimization study had been undertaken todetermine the ideal acid amount that would be required todecrease the base consumption. It was determined that anacid amount of 50 ml would be ideal to achieve this. Theeconomic outcome of this alteration is that it rendered theNPVs of the various manganese products positive.

In addition, the change to acid amount yielded an impureMnCO3 furnace feed material with a manganese content of > 45%, which was higher than the ideal manganese content

▲


Figure 6—Manganese content in the leach cake as a function of acidamount employed in the water-starved digestion

Table XII

Chemical composition of the PLS using 40 ml acidaddition in the water-starved digestion

Major element analysis on PLS – 40 ml acid volume

Element Unit Concentration

Mn g/l 29.85Al ppm 6B ppm 240Ca ppm 637Cr ppm <1Fe ppm 663K ppm 104Mg ppm 5411Na ppm 143Ni ppm 7Si ppm 52Sr ppm 9V ppm <1Zn ppm 2

Figure 7—Eh-pH diagram for manganese and iron compounds(Pakarinen, 2011)

Table XIII

XRF analyses of impure MnCO3 furnace feedproduced using (NH4)2CO3 or Na2CO3 as basereagents

Base reagent

Element (%) (NH4)2CO3 Na2CO3

Mn 48.8 47.3P <0.005 <0.005Si <0.005 <0.005Cr 0.09 0.09Ni <0.005 <0.005Cu <0.005 <0.01Al <0.01 <0.01V <0.005 <0.005Mo <0.005 <0.005Co <0.005 <0.005Ca 4.43 4.83Mg 1.71 3.72Fe 1.13 1.07

of between 35–40%, which is sought after by the ferroalloyindustry. The benefits of the feed material produced via thehydrometallurgical route include a high quality feed materialwhich will result in an increase in the production offerromanganese metal per day as the recoveries ofmanganese will be increased, and thereby industry will havethe option of purchasing lower grade ore (high Fe ore) at areduced cost (Steenkamp, 2013).

Electrowinning of EMD

We decided to electrowin electrolytic manganese dioxide(EMD) from the PLS by water-starved digestion of the slagwith 40 ml of sulphuric acid. The cake was then water-leached with 350 ml of distilled water to ensure that themanganese concentration in the leach solution was >27 g/l,allowing it to be amenable to electrowinning of the EMDproduct. This method was employed due to the chemicalvariance of the slag sample, and to ensure that the recoveryof manganese to the leach solution was >27 g/l. The concen-trations of the manganese and impurity elements in the leachsolution before and after purification with ammonia solutionare given in Table XIV.

The mass of EMD produced over a 72 hour period was8.3 g, thus yielding a current efficiency of approximately74%. A current density of 100 A/m2 should result in currentefficiencies that typically lie between 90–97%, as wasachieved by by Te Riele (1983). The current efficiency in thisinvestigation, however, may have been affected by theincrease in acid concentration, which promoted the formationof Mn3+. This can be overcome by increasing the pH by

adding a base reagent such as ammonia, in order to increasethe pH to 5–6, thereby maintaining an ideal currentefficiency.

The results in Table XV indicate that a product that meetsspecifications can be produced. The iron content was high;however, this result is not in agreement with the ICP-OESanalysis of < 1 ppm Fe (Table XIV). According to Te Riele(1983) a pregnant leach solution containing < 1 ppm Feshould yield a product with only 0.021 % Fe. Furthermore,SEM-EDS analysis of the material did not detect iron.

The micrograph in Figure 8 clearly show slight pitting ofthe surface of the EMD.

Figure 9 is considered to show the typical morphology ofEMD. The morphology appears to be similar to that obtainedby Liu et al. (2007). Thus, good quality EMD can beproduced from the leaching of ferromanganese slag.

Conclusions and recommendations

The manganese content of the slag (>30%) rendered thematerial attractive for further processing. A pregnant leachsolution with typical manganese concentration of 25–35 g/land low impurities content was obtained. Impurities wereeffectively removed using ammonia or sodium hydroxide.Ammonia was the most effective reagent, removing Fe, Si,and Al to levels of <1 ppm.



Table XIV

Chemical composition of the PLS used forelectrowinning of EMD before and after purificationwith NH3

Species Unit Before purification pH ~ 4.5 After purification pH ~5.6

Mn g/L 30.8 30.8Cr ppm 1 < 1Fe ppm 674 < 1K ppm 162 234Mg g/L 5.6 5.1Al ppm 81 < 1Na ppm 181 237Ni ppm 6 6Si ppm 79 22Zn ppm 2 < 1

Table XV

Analysis of the EMD product

Parameter Battery-active EMD specifications Results

MnO2 91% > 95%

Mn >59% 68%Fe <10 ppm 700 ppmAl <100 ppm 5 ppmCu <5 ppm < 1 ppmPb < 5 ppm Trace

Figure 9—SEM micrograph of the surface of the EMD deposit at 2.5 μmresolution

Figure 8—SEM micrographs of the surface of the EMD deposit at 500 μm and 250 μm resolution


Water-starved digestion of the slag with 100 ml sulphuricacid and purification of the resulting solution led to amanganese recovery of >99.7% into a manganese carbonateproduct at a pH >8.5. However, co-precipitation of calciumcarbonate resulted in the product exceeding the maximumspecified calcium content. Further investigations will beaimed at investigating the kinetics of co-precipitation in orderto decrease the amount of calcium carbonate precipitated.

A viable MnCO3 furnace feed material with a highmanganese (> 45% Mn) and low iron content (< 2% Fe) wasproduced from the pregnant leach solution utilizing NH3 and(NH4)2CO3 or NaOH and Na2CO3 to increase the pH to >8.5.Water-soluble phases containing ammonia and sodium thatformed could be detrimental to furnace operations. To rid thefeed material of these phases and to yield a quality productthe material was washed with water, which yielded a productcontaining >95% MnCO3.

Although the processes were shown to be technicallyfeasible, the large amounts of the base reagent required toadjust the pH of the pregnant leach solution from less than 0 to 5.6 when 100 ml of sulphuric acid was utilized in thedigestion stage resulted in an increase in operating costs,rendering the production of the manganese productsunviable. Therefore, to improve economic viability a studywas undertaken to optimize the quality of the PLS, and toproduce a leach residue (cement additive) that contained < 1% Mn. It was found that decreasing the sulphuric acidaddition to approximately 50 ml resulted in an increase of thepH of the PLS to approximately 3.4–4, and within this pHrange most of the major impurity elements precipitated outand reported to the residue. This adjustment resulted in theproduction of a MnCO3 furnace feed material with amanganese content >45% and a high Mn: Fe ratio.

In addition, to produce pure manganese carbonate, asmaller amount of neutralizing reagent would be required toadjust the pH to 5.6 when utilizing an acid addition of 50 ml.This adjustment rendered the process economically viable.Additional benefits of using 50 ml of acid in the digestionstage include a significant reduction in the dilution of thepregnant leach solution with large amounts of base reagent,thereby maintaining a high manganese concentration in theleach solution, and eliminating potential gelling problems.

The use of a hydrometallurgical process to recovermanganese products and produce a leach residue that issaleable as a cement additive is beneficial to the environmentand surrounding communities in ferromanganese-producingareas. Both slag dumps and current arisings can be utilized,thereby reducing the environmental and health risks posedby the leaching of mobile manganese contained in the slag. Inaddition, all potential waste streams produced during thehydrometallurgical process, such as the barren solution, canpotentially be recycled to produce additional marketable by-products, including ammonium sulphate or sodium sulphate.However, further investigations would have to be carried outto confirm this.

References

ALIBABA GROUP. 2013. Manganese product specifications and price.

http://sjzbcchem.en.alibaba.com/product/517509869-

213404979/sell_manganese_carbonate_MnCO3.html [Accessed 15 August

2013].

GROOT, D., KAZADI, D., POLLMANN, H., DE VILLIERS, J., REDTMANN, T., and

STEENKAMP, J. 2013. Utilization of ferromanganese slags for manganese

extraction and as a cement additive. Advances in Cement and Concrete

Technology in Africa, Emperor's Palace, Johannesburg, 28– 30 January

2013. pp. 984–985.

GOUS, J. 2013. Transalloys. Personal Communication.

ILLIS, A. and BRANDT, B.J. 1975. Selective process for the recovery of metal

values from sea nodules. CA Patent no. 974371.

KONO, Y., MIZOTA, T., and FUJII, Y. 1986. A precipitation separation method for

copper, nickel, and cobalt recovery from sulfurous acid leach liquor of sea

manganese nodules. Nippon Kogyo Kaishi, vol. 102, no. 1183.

pp. 585–590.

LIU, B., THOMAS, P.S., RAY, A.S., DONNE, S.W., and WILLIAMS, R.P. 2007. DSC

characterisation of chemically induced electrolytic manganese dioxide.

Journal of Thermal Analysis and Calorimetry, vol. 88. pp. 177–180.

MCINTOSH, S.N. and BAGLIN, E.C. 1992. Recovery of manganese from steel plant

slag by carbamate leaching. Report of investigations, US Department of

the Interior, Bureau of Mines.

NAGANOOR, P.C., PRASANN, A.S.R., SHIVAPRASAD, K.H., and BHAT, K.L. 2000.

Extraction of manganese from ferro-manganese slag. International

Symposium on Processing of Fines, Jamshedpur, India, 2–3 November

2000. National Metallurgical Laboratory, Jamshedpur. pp. 300–306.

PAKARINEN, J. 2011. Recovery and refining of manganese as by-product from

hydrometallurgical processes. Colloqium, Lappeenranta University of

Technology, Finland. p. 29.

PARKER, J. and LOVEDAY, G. 2006. Recovery of metal from slag in the ferro-alloy

industry. Hidden Wealth. Southern African Institute of Mining and

Metallugy, Johannesburg. pp. 7-15.

RETHINARAJ, J.P. and VISVANATHAN, S. 1993. Preparation and properties of

electrolytic manganese dioxide. Journal of Power Sources, vol. 42.

pp 335–343.

STEENKAMP, J. 2013. Department of Materials Science and Metallurgical

Engineering, University of Pretoria. Personal Communication.

STEENKAMP, R. 2013. Exxaro. Personal Communication

TE RIELE, W.A.M. 1983. Electrowinning of manganese dioxide. SAIMM

Vacation School Electrometallurgy, Randburg. South African Institute of

Mining and Metallugy, Johannesburg. pp. 238-261.

WELLBELOVED, D.B, CRAVEN, P.M., and WAUDBY, J.W. 2003. Ullmann's

Encyclopedia of Industrial Chemistry. Wiley and Sons, New York.

Chapter 6.

ZHANG, W. and CHENG, C. 2007. Manganese metallurgy review Part II:

Manganese separation and recovery from solution. Hydrometallurgy,

vol. 89, no. 3-4. pp 160–170.

ZHANG, W., CHENG, C., and PRANOLO, Y. 2010. Investigation of methods for

removal and recovery of manganese in hydrometallurgical processes.

Hydrometallurgy, vol. 101, no. 1–2. pp 58–63. ◆

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Introduction

The Bushveld Complex in South Africa is thelargest known layered igneous complex andcontains two-thirds of the world’s PGMreserves (Jacobs, 2006). Platinum-groupmetals (PGMs) are concentrated in threenarrow layers in the Bushveld Complex,namely the Merensky Reef, the Platreef, andthe UG2 chromitite layer (Jones, 2005). PGMsoccur together with base metal sulphides in theMerensky and Platreef, while the UG2 layerhas a high chromite content and low concen-trations of base metal sulphides. UG2 orescontain approximately 30% Cr2O3 comparedwith 0.1 % Cr2O3 in the Merensky deposit(Jones, 2005). Mines are increasinglyexploiting the UG2 as the Merensky Reefbecomes depleted.

The chromite content of the UG2 orepresents a major challenge for the PGMsmelting process as high levels of chromiumincrease the slag liquidus temperature, whichnecessitates higher operating temperatures. If,

under non-reducing conditions, the Cr2O3content of the concentrate exceeds approxi-mately 1.8% the chromite spinels do notdissolve in the slag at conventional slagtemperatures (1500–1550°C), but build up inthe furnace hearth and form a ‘mushy’ three-phase layer at the slag-matte interface(Eksteen 2011; Hundermark, 2011). Thisleads to inefficient matte segregation anddifficulties in matte tapping. To achieveadequate suspension of the spinels, thefurnace temperature is increased by increasingthe electrode immersion depth and powerdensity (Eksteen, 2011). Chromium control istherefore an important factor when consideringfurnace operating temperature. This is done byeffectively controlling the concentrate blendthat is fed into the furnace and adjusting theoperating power accordingly (Jacobs, 2006).

Of great concern when smelting concen-trates of high chromite content is high mattetemperatures that can destroy the protectiveslag freeze lining when the matte temperatureis higher than the slag liquidus temperature(Eksteen 2011; Hundermark 2011; Warner etal., 2007). Slag temperatures of between1600°C and 1700°C and matte temperatures ofbetween 1450°C and 1500°C were estimatedby Snyders et al. (2006) at the Polokwanesmelter. These high operating temperaturesresult in highly fluid matte, which causes hightap rates and severe matte penetration into thetapping channel and endwall. Failure of thetaphole bricks can cause major damage to thefurnace and poses a great safety risk tooperators.

FactSAGE® calculations performed byEksteen (2011) predicted that at temperaturesabove 1500°C the matte will react chemicallywith the refractory lining. In this processchromium will be picked up in the matte and

Wear of magnesia-chrome refractory bricksas a function of matte temperatureby M. Lange*, A.M. Garbers-Craig*, and R. Cromarty*Paper written on project work carried out in partial fulfilment of B. Eng. (Metallurgy)

SynopsisThe postulation that primary platinum group metal (PGM) mattewill chemically react with magnesia-chrome bricks when temper-atures exceed 1500°C was tested. Magnesia-chrome brick sampleswere heated in contact with matte at 1300°C to 1750°C for 30minutes, after which the refractory samples were analysed usingreflected light microscopy and scanning electron microscopy. Thesamples were all completely penetrated by matte. As thetemperature increased the matte also penetrated the fusedaggregate grains and disintegrated them. The chromium concen-tration of the matte inside the refractory samples was found to beslightly higher than that of the bulk matte. At temperatures of1500˚C and higher, MgO, FeO, and magnesium-rich silicate crystalscould be identified in the matte directly adjacent to the refractory-matte interface. Phase relations clearly indicated that chemicalreactions take place between primary PGM matte and the magnesia-chrome refractory material at temperatures above 1500°C, but thatthese reactions are more complex than expected from FactSAGE®

calculations.

Keywordsrefractory, magnesia-chrome bricks, penetration, fused grains.

* Centre for Pyrometallurgy, Department ofMaterials Science and Metallurgical Engineering,University of Pretoria, Pretoria, South Africa.



Wear of magnesia-chrome refractory bricks as a function of matte temperature

oxygen will be transferred from the refractory to the matteaccording to reactions [1] and [2]:

[1]

With ΔG ° = -8.478 kJ/mol at T=1500°C (ΔG ° < 0, when T > 1450°C)

[2]

With ΔG ° = +11.03 kJ/mol at T=1500°C (ΔG ° < 0, when T > 1650°C)

Eksteen (2011) concluded that the combination of hightemperatures and sulphidation reactions would cause disinte-gration of the brick due to the destruction of its spinelbonding phase.

This project subsequently investigated whether chemicalinteraction between magnesia-chrome bricks and primaryPGM matte takes place at temperatures above 1500°C, aspredicted by Eksteen (2011).

Materials

Vereeniging Refractories supplied samples of reconstitutedfused-grain magnesia-chrome refractory bricks, while furnacematte and slag samples were obtained from the AngloPlatinum Waterval Smelter. According to the supplier’sspecifications, the brick had a chemical composition of 57.4%MgO, 21.3% Cr2O3, 11.2% Fe2O3, 7.8% Al2O3, 1.5% SiO2,and 0.8% CaO, and an apparent porosity of 16%. The mattewas analysed using powder X-ray diffraction (XRD),scanning electron microscopy–energy dispersive X-rayspectroscopy (SEM-EDS), inductively coupled plasma opticalemission spectrometry (ICP-OES), and sulphur analysis(Leco-S). The matte sample consisted mainly of nickel,copper, and iron sulphides, and contained 39% Fe, 15% Ni,11% Cu, 34%S, 0.61% Cr, and 1% SiO2.

Experimental

Laboratory-scale experiments were performed in an inductionfurnace consisting of a graphite susceptor in an insulated,gas-tight, chamber. The furnace was purged with argon toprevent oxidation of the graphite susceptor. Power wassupplied by an Ambrell Ekoheat 15/100 radio frequencypower supply. Temperature was controlled by a Eurotherm2416 temperature controller with a Type B thermocouple.

Two sample types were used to study matte–refractoryinteractions:

➤ The first design consisted of crucibles and lids madefrom the supplied magnesia-chrome brick. Matte andslag were placed into a cavity drilled into a core cutfrom the refractory brick (Figure 1). It was intended toinvestigate the interaction between the matte and therefractory brick crucible. In this design the refractorycrucible was in direct contact with the graphitesusceptor

➤ In the second design, matte was contained in a closedalumina crucible. A cylindrical sample of refractorybrick was inserted into the matte (Figure 1). There wasa limited volume of refractory in contact with the matte,and the refractory sample and matte were isolated fromthe graphite susceptor.

The induction furnace was purged with technical-gradeargon (99.9% Ar) for 30 minutes, after which it was heatedto temperatures ranging from 1300°C to 1750ºC at 25°C/min.The samples were kept at temperature for 30 minutes. Thefurnace was cooled down at 25°C/min under argon.Experimental runs were randomized to improve statisticalaccuracy.

The cooled samples were visually inspected, measured,cut, and polished sections prepared. Samples were cold-mounted in epoxy resin, ground, and finally polished to a 1 μm finish. The polished sections were examined usingreflected light microscopy and scanning electron microscopy.SEM-EDS was used to evaluate the amount of mattepenetration and the possible chemical interaction between thematte and refractory.


Microstructure of unreacted magnesia-chromerefractory brick

The as-supplied magnesia-chrome brick consisted of largemagnesia grains (dark grey phase), which contained smallamounts of iron and chrome in solid solution (Figure 2).Finely exsolved spinel crystals (light grey phase) wereobserved in the magnesia grains. The spinels on the grainboundaries had roughly the same composition as the spinelswithin the magnesia grains.

Design 1 – magnesia-chrome crucible containingmatte and slag

Design 1 (magnesia-chrome crucible, filled with matte andslag) was used for tests at 1300°C and 1500°C. This design,however, resulted in the refractory soaking up all the matte,leaving no matte in the crucible for analysis. Themicrostructure of the magnesia-chrome brick with soaked-upmatte is shown in Figures 3 and 4.

▲


Figure 1—Crucible design 1 (left) and 2 (right)

The white phase is a penetrated Fe-rich sulphide phase,while the very light grey phase represents a Cu-Fe-Ni-basedsulphide (Figure 4). The light grey phase in Figure 4 is thetypical chromite spinel that is found in the refractory, withthe dark grey phase surrounding it consisting of mainly

magnesium oxide. The penetration that took place wasminimal but uniform throughout the refractory. This experi-mental set-up presented numerous difficulties as there wasno matte left for chemical analysis due to the completepenetration. This design also allowed the refractory to be indirect contact with the graphite susceptor of the inductionfurnace, thereby exposing it to a very reducing atmosphere.These reducing conditions resulted in the formation of aniron-nickel alloy within the refractory closest to the graphitesusceptor.

It was subsequently decided to change the design to aset-up in which a small piece of refractory, surrounded with alarge amount of matte, was placed in a large alumina crucible(Design 2). The crucible was closed with an alumina lid tolimit the sulphur losses from the matte. No slag was used inthis design as the slag tended to react with the aluminacrucible. Alloy formation in the matte was not observed whenDesign 2 was used.

Design 2 – alumina crucible containing matte andrefractory sample

Observed matte-refractory interaction is discussed in terms ofmatte penetration into the open porosity, matte penetrationinto the fused aggregate grains, and matte-refractoryinteraction at the interface.

Matte penetration into open porosity

All the examined refractory samples from Design 2 werecompletely penetrated by matte, similar to samples fromDesign 1. A penetration profile could not be distinguished asthe refractory samples had a ‘wicking’ effect on thesurrounding matte. All the samples had roughly the sameoutward appearance (Figure 5).

The extent of penetration as a function of temperaturecan be seen in Figure 6. The porosity of the unreacted brick(A) became completely filled with matte as the refractorypiece came in contact with the matte at the different temper-atures. The large fused grains began disintegrating as thetemperature was raised and finally the matte even penetratedinto the fused grains themselves.

Matte penetration into the fused aggregate grains

Examination of the aggregate grains at higher magnificationsgave important insight into the reactions taking place at



Figure 2—Backscattered electron image of the microstructure of theas-received magnesia-chrome brick

Figure 3—Backscattered electron image of matte penetration (whitephase) into the magnesia-chrome crucible at 1300°C

Figure 4—Backscattered electron image of matte penetration into themagnesia-chrome crucible at 1300°C (high magnification)

Figure 5—Refractory sample tested at 1400˚C


increasing temperatures. Except for penetration, the tests at1300°C and 1400°C showed no clear signs of chemicalinteraction. It was only at temperatures of 1500°C and higherthat signs of interaction were noticed. This agrees withFactSAGE® predictions.

At 1500°C and higher, spinels migrated out of themagnesia grains and collected on the grain boundaries. Thisleft large areas of the magnesia grains exposed as theybecame depleted in spinels. Furthermore, matte started topenetrate into the fused grains and disintegrated them(Figure 7). EDS analysis of this penetrating matte indicatedthat it locally contained up to 0.7% Cr, compared to 0.3% Crin the bulk matte.

Disintegration of the fused grains was more pronouncedin the refractory samples that were reacted at 1600°C and1700°C (Figures 8 and 9). The spinel crystals migrated out ofthe magnesia grains, while the matte completely disintegratedthe large fused grains. At 1600°C the penetrating matte had aCr content of 0.5% compared to 0.3% in the bulk matte,which is not a significant increase.

At 1750˚C (Figure 10) the penetrating matte contained0.8% Cr compared with a bulk matte concentration of 0.3%.A sudden increase in porosity was also observed (Figure 10),together with a large amount of sulphur gas evolution duringthe experiment.

As the experimental temperature increased the spinelcrystals increasingly migrated out of the magnesia grains andgrouped together as the magnesia grains disintegrated. Thisdestruction of the fused grains would make the bricks morevulnerable to the effects of thermal shock, as without thefinely dispersed spinel crystals the bricks would be moresusceptible to crack propagation.

Matte-refractory interaction at the refractory-matteInterface

It was important to carry out a detailed analysis of theinterface between the refractory and the surrounding matte,as this would indicate how the actual furnace lining wouldreact when brought in contact with the superheated matte.Tests at 1300˚C and 1400˚C again resulted in mattepenetration only.

▲


Figure 6—Reflected light micrographs of matte penetration into themagnesia-chrome refractory as a function of temperature. A: Unreactedbrick, B: 1300˚C, C: 1400˚C, D: 1500˚C, E: 1600˚C, F: 1700˚C)

Figure 7—Backscattered electron image of the microstructure of themagnesia-chrome brick reacted at 1500˚C



At temperatures of 1500ºC and higher a magnesium-richsilicate phase could be observed to form a boundary layer atthe refractory-matte interface (Figures 11 to 13). Thisboundary phase could have been established through thepredicted formation of MgO that reacted with the silica ineither the matte or refractory, as both contained smallamounts of silica. On either side of the boundary, FeOcrystals started to form, together with spinel crystals. Thestoichiometry of these crystals started to change from M3O4to MO. Some large spinel crystals could also be observed atthe refractory-matte interface, as well as magnesia grainswith finely dispersed exsolved spinel crystals.

At temperatures of 1700°C and 1750°C more FeO crystalsstarted to form at the interface, and the stoichiometry of themagnesium-rich silicate phase was calculated to be(Mg,Fe)2SiO4 (Figures 13 and 14). The extent of spinelsegregation out of the magnesia grains close to the interfacewas also higher at these temperatures. At 1750°C the spinelswere completely removed from the fused magnesia grainsclose to the refractory-matte interface (Figure 14).

Conclusions

The observed amount of matte penetration into themagnesia-chrome brick samples was substantial. All therefractory samples were completely penetrated by matte. As

the temperature increased the matte also started to penetratethe fused aggregate grains, and disintegrated them.

CrS was not observed as a separate phase. It was noticed,however, that the chromium content of the matte inside therefractory sample was slightly higher than that of the bulk



Figure 14—Backscattered electron image of the matte-refractoryinterface at 1750˚C



Figure 12—Backscattered electron image of the matte-refractoryinterface at 1600˚C (dark grey phase in the light grey matte is Mg-richsilicate crystals)



matte. At temperatures of 1500˚C and higher, definite signsof MgO and FeO formation were found, as well as the gradualdisintegration of the grains inside the refractory at therefractory-matte interface.

The formation of a magnesium-rich silicate boundaryphase at the matte-refractory interface at temperatures above1500ºC was unexpected and needs further investigation.Results obtained in this study clearly indicate that thereactions that took place between primary PGM matte and themagnesia-chrome refractory material are more complex thanthose predicted through FactSAGE® modelling.

Acknowledgements

Technical and financial support from Anglo AmericanPlatinum is gratefully acknowledged. This work is based onresearch supported in part by the National ResearchFoundation of South Africa (Grant number TP1208219517).

References

EKSTEEN, J. 2011. A mechanistic model to predict matte temperatures during the

smelting of UG2-rich blends of platinum group metal concentrates.

Minerals Engineering, vol. 24. pp. 676–687.

HUNDERMARK, R., MNCWANGO, S., DE VILLIERS, L., AND NELSON, L. 2011. The

smelting operations of Anglo American's platinum business: an update.

Southern African Pyrometallurgy, Cradle of Humankind, South Africa, 6-9

March 2011. Jones, R. and den Hoed, P. (eds). Southern African Institute

of Mining and Metallurgy, Johannesburg. pp. 295–308.

JACOBS, M. 2006. Process discription and abbreviated history of Anglo

Platinum’s Waterval Smelter. Southern African Pyrometallurgy, Cradle of

Humankind, South Africa, 5-8 March 2006. Jones, R. (ed.). Southern

African Institute of Mining and Metallurgy, Johannesburg. pp. 17–28.

JONES, R. 1999. Platinum smelting in South Africa. South African Journal of

Science, vol. 95. pp. 529–539.

JONES, R. 2005. An overview of Southern African PGM smelting. Mintek,

Randburg, South Africa.

NELL, J. 2004. Melting of platinum group metal concentrates in South Africa.

Journal of the South African Institution of Mining and Metallurgy,

vol. 104, no. 7. pp. 423–429.

SNYDERS, C., EKSTEEN, J., and MOSHOKWA, A. 2006.. The Polokwane smelter

mattte tapping channel model. Fifth International Conference on CFD in

the Process Industries, Melbourne. pp. 3-6.

WARNER, A.E.M., DIAZ, C.M., DALVI, A.D., MACKEY, P.J., TARASOV, A.V., and JONES

R.T. 2007. World nonferrous smelter survey part IV: nickel: sulfide.

Journal of Metals, April 2007. pp. 58–72. ◆

▲


The SAIMM LibraryThe library has been indexed and sorted.Although it is not as big as we would like it tobe, we have a fair number of reference booksand a certain amount of resource material.

Access to the library and the control of theborrowing process is in the hands of Kea Shumba.

The titles are available on the website atthe following link: http://www.saimm.co.za/saimm-library?task=showCategory&catid=32

Sam Moolla, Manager: SAIMM Secretariat

Introduction

Copper and cobalt are commodities of greateconomic value. Copper derives its value fromits outstanding properties which include highelectrical and thermal conductivity,malleability, and toughness. Copper is widelyused in electrical cabling, piping, and theconstruction industry. Cobalt owes its value tothe fact that it can maintain its ferromagnetismup to temperatures higher than any othermetal, and to its superior catalytic qualities.Cobalt is widely used in the manufacture ofmagnets, catalysts, and batteries. It is alsoused as a pigment in paints and as an alloyingelement in steel production.

The Central African Copperbelt hosts 40%of the world’s cobalt reserves and 10% of theworld’s copper reserves. The two countries thatlie on the Copperbelt, the Democratic Republicof Congo (DRC) and Zambia, produce abouthalf of the world’s copper-cobalt ore(Crundwell et al., 2011). The DRC is said to bethe biggest miner of these ores, although thelargest cobalt refinery is in China (Miller,2009). Zambia and the DRC have a veryimportant role to play in supplying globalcopper and cobalt requirements.

Escalating global demand for copper andcobalt has forced companies to increasinglyexploit lower-grade ores. However, most ofthese low-grade ores are processed at veryhigh costs, since large volumes have to beprocessed. The methods applied for theextraction of copper and cobalt from thecopper-cobalt oxide ores of the Central AfricanCopperbelt include (i) heap leaching of thelow-grade ores, (ii) upgrading of the ores byflotation prior to leaching, and (iii) direct oreleaching using sulphuric acid. All thesemethods have inherent problems. For instance,heap leaching often has low rates of recovery,the inclusion of a flotation stage prior toleaching is costly, and direct ore leachingconsumes large volumes of acid.

The main cobalt-bearing mineral in thecopper-cobalt oxide ore deposits of the DRC isheterogenite (CoO.2Co2O3.6H2O). Coppertypically occurs as chrysocolla (CuOSiO2.2H2O)and malachite [CuCO3.Cu(OH)2] (Crundwell etal., 2011). Minor amount of copper silicatessuch as dioptase (CuSiO3.H2O), katangite(CuS13O9.nH2O), and carbonates such asazurite (Cu3(OH)2(CO3)2 also occur (Prasad,1989). The main oxides are listed in Table I.

Comparing the extent of the dissolution ofcopper-cobalt ores from the DRC Regionby S. Stuurman*, S. Ndlovu*, and V. Sibanda*Paper written on project work carried out in partial fulfilment of BSc Engineering (Metallurgical and Materials Engineering)

SynopsisInorganic acids such as sulphuric acid have found use together withcertain reducing agents in leaching of copper-cobalt oxide ores. Thesereagents are not ideal due to the adverse effect the inorganic acidsgenerally have on the environment and the high costs of the reducingagents. In this study a copper-cobalt oxide ore from the CentralAfrican Copperbelt was leached in two different environments;sulphuric acid in conjunction with hydrogen peroxide as a reducingagent and tartaric acid. The effects of acid concentration, reducingagent concentration, and temperature were independently determinedfor both leaching environments. The sulphuric acid concentration wasvaried between 0.4 M and 1.2 M and the concentration of hydrogenperoxide between 4.0 M and 6.5 M, while the tartaric acid concen-tration was varied between 0.15 M and 0.35 M. The temperature wasvaried between 20°C and 50°C. The results showed that the extractionof both copper and cobalt increased with sulphuric acid concen-tration, reaching a peak at approximately 0.8 M and then decreasingat higher acid concentrations. A similar increase and decrease inmetal extraction was observed when the reducing agent wasincreased. In leaching with tartaric acid, the extraction of cobalt wasmuch higher than that of copper, although extraction of both metalsincreased with acid concentration. Additions of small amounts ofhydrogen peroxide were found to increase cobalt extraction in tartaricacid but had a minimal effect on copper. An increase in the solutiontemperature had a significant effect in the organic acid environment,with the effect on cobalt extraction being much more pronouncedthan on copper.

Keywordsleaching, copper, cobalt, sulphuric acid, hydrogen peroxide, tartaricacid, reducing agent.

* School of Chemical and Metallurgical Engineering,University of the Witwatersrand, Johannesburg,South Africa.

© The Southern African Institute of Mining andMetallurgy, 2014. ISSN 2225-6253. Paper receivedMar. 2014.


Comparing the extent of the dissolution of copper-cobalt ores from the DRC Region

In the Katanga province of the DRC, cobalt is commonlyproduced from heterogenite (CoO.2Co2O3.6H2O) as a by-product of copper production. Cobalt in heterogenite occurs inboth the 2+ and 3+ oxidation states. In contrast to copperoxide minerals, which readly dissolve in sulphuric acidsolution, cobalt is difficult to leach from heterogenite. This isbecause when heterogenite is leached in an inorganic acidsuch as sulphuric acid, Co2+ easily goes into solution, butCo3+ does not dissolve:

[1]

The insoluble Co3+ becomes soluble only after reductionto Co2+. Therefore the hydrometallurgical dissolution of Co3+

can take place only in the presence of a reducing agent. Thereducing agents commonly used in the process includeferrous ions (present in the leach solution as a result ofleaching of iron minerals in the ore and leaching of ironscrap), sodium metabisulphite (Na2S2O5), and hydrogenperoxide (Apua and Mulaba-Bafubiandi, 2011; Lydall andAuchterlonie, 2011; Seoa et al., 2013). The consumption of these reducing agents is generally high and these make up 47% of the total operation which makes the production of copper and cobalt very expensive (Mulaba-Bafubiandi et al., 2007).

The use of sulphuric acid as the main leaching agent alsoposes a threat to the environment. Sulphuric acid waste leachsolution, because it is non-biodegradable and toxic, is storedin on-site holding ponds, which are securely lined. Eventhough the holding ponds are considered environmentallysafe there have been reports of spills impacting groundwatereven miles away (Freeman, 2005). Therefore, an economicalmethod of extracting the minerals that uses a more environ-mentally friendly leaching reagent is required in order for theprocessing of these low-grade ores to be profitable.

Since inorganic acids have not performed well ineconomic terms or in meeting the standards for ‘green’chemistry and environmental impact, this has prompted theneed to consider organic acids as lixiviants. To date, organicacids have not been used widely as leachants due to theirreported low leaching efficiencies. They are also veryexpensive and are, as a result, most unlikely to be used onlow-grade ores in conventional processing routes. However,they are attractive due to the fact that they have been foundto cause less harm to the environment as they arebiodegradable. In addition, organic acids are also recyclable

(Gharabaghi et al., 2010). Organic acids extract metals byforming complexes. The more stable the complex the higherthe extraction rate, the smaller the ionic radius of the targetmetal the more stable the complex formed, and the higher theoxidation state of the metal the more stable the metalcomplex (Weisstein, 2011).

Currently there are no documented studies on theleaching of copper-cobalt oxide ores from the Central AfricanCopperbelt by organic acids. In the present work, preliminarystudies have been conducted to establish the feasibility ofleaching using an organic acid. The leaching behaviour of acopper-cobalt oxide ore in an organic acid environment(tartaric acid) was compared with that in an inorganicenvironment (sulphuric acid) fortified with a reducing agent(hydrogen peroxide). Tartaric acid has been observed tochelate metal ions and is relatively cheap compared to otherorganic acids. This choice was further reinforced by studiescomparing the leaching rates of heavy metals from contam-inated soils and spent batteries using different organic acids,which showed that tartaric acid and citric acid could removeheavy metals from contaminated soils and waste materialmore efficiently and rapidly than all other potential organicextractants (Wasay et al., 2001; Li et al., 2010).


Ore sample

The ore used in the test work is a copper-cobalt oxide orefrom the Katanga Province in the DRC. The ore was crushedand milled to 80% –150 μm. The composition of the ore isgiven in Table II.

Reagents

The main reagents included analytical grade sulphuric acid(98%) as a leaching reagent, analytical grade hydrogen

▲


Table I

Primary copper and cobalt oxides found in copper-cobalt ores of the Democratic Republilc of Congo (Prasad,1989)

Oxides of copper Oxides of Cobalt

Malachite Cu2(OH)2CO3 Heterogenite (CO2O3.CuO)H2O

Cuprite Cu2O Kolwezite (Cu, Co)2(CO3)(OH)2

Libethenite Cu2(OH)PO4 Stainierite Co2O3.H2O

Pseudomalachite Cu2(OH)4PO4H2O Amorphous CoO.2Co2O3.6H2Oheterogenite

Tenorite CuO

Table II

Chemical composition of the Cu-Co ore sample usedin the test work

Element Cu Co Zn Ni Fe Mn Mg SiO2

Wt% 4.53 0.3 0.029 0.003 1.4 0.22 2.48 84.5

peroxide (30%) as the reducing agent, 100% tartaric acid,and distilled water for dilution. All reagents were sourcedfrom Merck Millipore.

Experimental method

Inorganic acid leaching

Dilute solutions of sulphuric acid and hydrogen peroxidewere prepared separately. The dilute solutions were mixed inbeakers accordingly to obtain the desired pH. The beakerswere then placed in a shaking water bath that was set at thedesired temperature and the solutions were given time toreach a steady temperature. The pH of the solutions wasmonitored using a calibrated pH meter and the temperaturewas measured using a thermometer.

Once the solution pH and temperature of the solutionwere stable, the ore sample was introduced into the solutionin the beakers. The pulp in the beaker was continuouslystirred for 3 hours to facilitate leaching. Table III summarizesthe test conditions for the inorganic acid leaching tests.

The experiments were all undertaken over a 3 hourperiod.

Organic acid leaching

Dilute acid solutions of different concentrations wereprepared by dissolving tartaric acid powder in distilled water.The pH of the solutions was measured. The solutions werepoured into beakers and the beakers were placed in ashaking water bath to achieve the desired temperature. Oncethe leaching solution in the beaker reached a steadytemperature, the ore sample was introduced. The pulp wasleft to leach for 24 hours while being continuously stirred. Asample for analysis was taken after the first 3 hours.

Table IV summarizes the test conditions for the organicacid leaching tests.

The experiments were all undertaken over a 3 hourperiod.

Leachate analysis

Leachate samples were analysed to determine the concen-

tration of copper and cobalt after each experiment. In bothsets of experiments, the leached pulp was left to settle inorder to allow for solid/liquid separation to occur. The pulpwas then filtered using filter paper, a Bu–

..chner funnel, and a

filter flask to obtain the leach solution. Samples of the filteredsolution were poured into sample bottles and sent for copperand cobalt analysis by inductively coupled plasma massspectrometry (ICP-MS) using a Perkin Elmer Nexlon 300Dinstrument.


Inorganic acid leaching

Effect of sulphuric acid concentration

Figure 1 shows the extraction of the metals as a function ofsulphuric acid concentration. The results show similar trendsfor the extraction of both copper and cobalt. At 0.4 M acidconcentration, cobalt and copper extractions were 34.0% and35.0% respectively. Extraction increased to a maximum of97.4% (Cu) and 78.2% (Co) at 0.8 M sulphuric acid concen-tration, and then declined with further increases in acidconcentration to 78.3% for copper and 63.3% for cobalt at 1.2 M.

One would expect that with an increasing acid concen-tration the dissolution of the copper and cobalt would



Table III

Test conditions for inorganic acid leacing in a reducing environment

Test Temperature (°C) Sulphuric acid concentration (M) Hydrogen peroxide concentration (M)

Effect of sulphuric acid concentration 25 0.4, 0.8, 1.2 3.0Effect of hydrogen peroxide (reducing agent) concentration 25 0.4 4.0, 5.5, 6.5Effect of temperature 20, 30, 40 0.8 4.0

Table IV

Test conditions for inorganic acid leaching

Test Temperature (°C) Tartaric acid concentration (M) Hydrogen peroxide concentration (M)

Effect of tartaric acid concentration 25 0.15, 0.25, 0.35 0.0Effect of hydrogen peroxide (reducing agent) concentration 25 0.35 3.0, 4.5, 5.5, 6.5Effect of temperature 20, 30, 40, 50 0.35 0.0

Figure 1—Effect of sulphuric acid concentration on copper and cobaltextraction (temperature 25°C, hydrogen peroxide concentration 3M,leaching time 3 hours


increase, as there are more H+ ions available. However, thisis not the case. The Eh-pH diagrams for copper and cobalt inFigure 2 and Figure 3 show that copper and cobalt ionizationis favoured at relatively low pH levels. The diagrams indicatethat in order to solubilize copper as Cu2+, the copper oxideminerals have to be leached in highly acidic conditions (pH <1). At pH conditions >1, a solid CuOFeO2 is formed(Figure 2).

Similarly, cobalt is best leached at pH <1. At pH levels >1,a solid CoO.Fe2O3 is formed, as seen in Figure 3, which has apassivating effect on the leaching of cobalt.

Thus as the pH increases beyond about 2 there is apossibility of formation of solid compounds of copper(CuFeO2) and cobalt (CoO.Fe2O3). These solids have apassivating effect on the leaching reactions, resulting in adecrease in the dissolution of both copper and cobalt.Solutions >1.0 M in concentration had a starting pH greaterthan 1. Addition of ore may have increased the pH to levelsthat favour passivation.

The shift from increasing extraction to decreasingextraction (Figure 1) may also be due to a change in the rate-controlling step. At acid concentrations between 0.4 M and0.8 M, the reaction kinetics may be controlled by the masstransfer of acid from the bulk solution to the particles. Atconcentrations from 0.8 M to 1.2 M, the controlling

mechanism may have changed, so that further increases inacid concentration would have no effect in the dissolution ofboth cobalt and copper ores.

Effect of the concentration of the reducing agent

The reaction of Co3+ oxides in the presence of hydrogenperoxide as the reducing agent is expected to proceed by thefollowing reactions:

[2]

[3]

The effect of a reducing agent on metal recoveries wasinvestigated in preliminary test work that focused on thesolution Eh. In the absence of reducing agent the Eh of thesolution varied between 600 and 900 mV vs SCE (0.242 V)over the leaching period. However, on addition of about 3.0M hydrogen peroxide the Eh of the solution decreaseddrastically to values around 300 to 500 mV vs SCE.According to Figure 1, the lower Eh of the solutionsubsequently enhances the dissolution of Co (III) phase.

The effect of hydrogen peroxide concentration inconjunction with sulphuric acid was then tested. Figure 4shows the metal extraction as a function of reducing agentconcentration.

The results show similar trends for both copper andcobalt extraction. Maximum extractions of 95.1% for copperand 79.4% for cobalt were obtained at around 4 M hydrogenperoxide. Extraction of copper and cobalt then decreased to59.3% and 63.8% respectively at about 6.5 M hydrogenperoxide. It is important to note that copper extraction washigher than that of cobalt at all hydrogen peroxide concen-trations except 6.5 M. This possible indicates differentreaction mechanisms controlling the Cu and Co extractionprocesses. Another point is the sharp increase in metalextraction from 3.0 M to 4.0 hydrogen peroxide concen-tration, which is followed by a decrease in extraction at 5.5 Mand above. This clearly indicates that the concentration ofreducing agent has a positive effect on metal extraction onlyup to a certain extent; above this range the reaction ratebecomes less dependent on hydrogen peroxide concentration.At reducing agent concentrations above 4 M, the masstransfer of reducing agent from solution to particles may nolonger be the rate controlling mechanism.

▲


Figure 2—Eh-pH diagram of the Cu-O-Fe system at 25°C (Schlesinger et al., 2011)

Figure 3—Eh-pH diagram of the Co-O-Fe system at 25°C (Schlesingeret al., 2011)

Figure 4—Effect of hydrogen peroxide concentration on copper andcobalt extraction (temperature 25°C, sulphuric acid concentration 0.4 M, leaching time 3 hours)

Effect of temperature

Since the optiuum cobalt and copper extractions wereobtained with 4.0 M hydrogen peroxide and 0.4 M sulphuricacid, these conditions were used in the temperature tests.Figure 5 shows the metal extractions as a function oftemperature. There is a marginal increase in the extraction ofboth copper and cobalt as the leaching temperature isincreased. It is known from the Arrhenius equation that anincrease in temperature generally enhances reaction kinetics.Increasing temperature may thus increase the mass transferkinetics of acid and reducing agent from solution to particles.The reaction at the surface of the particles is also enhancedby increasing temperature, thus leading to increased metaldissolution and higher extraction efficiency.

However, the temperature increase does not seem to haveas great an effect on the extraction efficiency of copper andcobalt as compared to acid and reducing agent concentration.The temperature effect was therefore capped at 40°C.

Organic acid leaching

Effect of tartaric acid concentration

Figure 6 shows the extraction of copper and cobalt as afunction of tartaric acid concentration.

The copper and cobalt extractions are both much lowerthan those achieved with the inorganic acid (Figure 1), withthe difference being more significant for copper than forcobalt. The maximum copper extracted within the range oftartaric acid concentration used was about 5% at 0.35 M,compared to about 90% in sulphuric acid media. Themaximum cobalt extracted was about 38% in tartaric acidcompared to about 80% in sulphuric acid media. Thedissolution of copper and cobalt is governed by the extent towhich the two acids dissociate, which is quantified by theirrespective dissociation constants (pKa). The lower the pKavalue, the higher the dissociation rate. Sulphuric acid has apKa of 1.99, compared with the pKa1 of 2.98 and pKa2 of4.34 for tartaric acid at 25°C (Murthy, 2008). It is alsonoticeable from Figure 6 that the maximum leaching abilityof tartaric acid was not reached within the selected concen-tration range. The effect of increased tartaric acid concen-trations on cobalt extraction should be investigated in futurestudies.

Cobalt extraction was much higher than that of copper atall acid concentrations tested. This is notably the reverse ofwhat was observed with the sulphuric acid leaching process,

where the extraction of copper was significantly higher thanthat of cobalt Furthermore, the level of copper extraction didnot change significantly with increasing acid concentration.These observations indicate that the cobalt mineral has afaster reaction rate with the organic acid than the coppermineral, and that the leaching reactions of the two metalsystems may be controlled by two different mechanisms.

Extraction of metals from ores by organic acids generallytakes place through protonation, chelation, and ligandexchange reactions. Thus the organic acid supplies bothprotons and metal-complexing organic acid anions, with theprotons contributing to proton-promoted mineral dissolution.The major factor with organic acids, however, may be thatmetal-organic complexes can form at the solid-solutioninterface, weakening cation-oxygen bonds and catalysing thedissolution reaction. The greater the stability of the metalcomplex formed; the higher the metal dissolution. Thestability of the complex depends on the ionic radius and theoxidation state of the metal ion. A smaller ionic radius and ahigher oxidation state both increase the stability of the metalcomplex formed..

Unlike Co (II) complexes, Co (III) complexes undergoligand substitution reactions relatively slowly and so tend tobe stable to ligand exchange (Kim et al., 1993). This couldexplain the lower cobalt extraction in the organic acid ascompared to sulphuric acid enhanced with a reducing agent.This might indicate that the Co (II) mineral species undergodissolution through the protonation mechanism. Theresulting Co (III) species, however, did not undergo muchdissolution due to the slow ligand exchange reactions. Inview of the low extractions observed with tartaric acid, theeffect of additions of a reducing agent was investigated.



Figure 5—Effect of temperature on copper and cobalt dissolution.Sulphuric acid concentration 0.4 M, hydrogen peroxide concentration 4 M, leaching time 3 hours

Figure 6—Effect of tartaric acid concentration on copper and cobaltextraction (temperature 25°C, leaching time 3 hours)

Figure 7—Effect of reducing agent concentration on copper and cobaltextraction (tartaric acid concentration 0.35 M, temperature 25°C,leaching time 3 hours)


Effect of addition of a reducing agent

Figure 7 shows the effect of hydrogen peroxide additions tothe organic acid leach solution. The addition of a reducingagent had a significant effect on cobalt extraction, but littleeffect on copper. As outlined in the preceding section it ispossible that the addition of a reducing agent will reduce theCo (III) species to Co (II).Furthermore, the uncomplexed Co(III) ion itself is not stable in water due to thehydrolysis/reduction reaction:

[4]

The generated Co (II) ions undergo ligand substitutionmuch faster, and as a result cobalt dissolution will be greatlyenhanced.

Effect of temperature

Figure 8 shows the effect of temperature on the extraction ofcobalt and copper in tartaric acid in the absence of hydrogenperoxide. Tests were carried out from 20°C to 50°C over a 3 hour leaching period.

As with all chemical reactions, the extraction of bothcobalt and copper increases with increasing temperature.Figure 8 shows that in general, an increase in temperaturehas a much more positive effect on the extraction of cobaltthan that of copper. The highest cobalt extraction (62%) wasrecorded at the highest temperature used (50°C), while only10% copper extraction was recorded at the same temperature.According to Shabani et al. (2012) the dissolution of cobalt ishighly dependent on temperature, with the effect being morenoticeable in the presence of tartaric acid than in sulphuricacid (Figure 5). Organic acids are weak acids and their metaldissolution abilities are affected by the extent of dissociation.The more the acid dissociates, the greater its ability tosolubilize metals. An increase in temperature results inincreased dissociation of tartaric acid, increasing the numberof hydrogen and ligand ions in the acid and thus enhancingthe cobalt extraction.

One other noteworthy aspect is that the temperature testswere carried out in the absence of a reducing agent. It ispossible therefore that an increase in temperature beyond50°C could result in the extraction efficiency for cobaltreaching similar levels to those observed at lower temper-atures in the presence of hydrogen peroxide. This aspectshould be investigated further in future work, as it couldhave an impact on the potential applicability and economicsof the process.

Conclusions

Sulphuric acid in the presence of hydrogen peroxide as areducing agent was able to extract a large amount of bothcopper and cobalt. About 95% copper and 80% cobaltextractions were achieved in 0.4 M sulphuric acid and 4 Mhydrogen peroxide in 3 hours’ leaching time at 25°C.Extractions of both copper and cobalt increased with anincrease in sulphuric acid concentration up to 0.8 M. Additionof hydrogen peroxide to the sulphuric acid leaching solutionhad a positive effect on both copper and cobalt dissolution upto 4 M hydrogen peroxide. An increase in temperature,however, did not have as an significant effect on theextraction efficiency of copper and cobalt as the reducingagent concentration.

With tartaric acid as the lixiviant, about 40% cobalt and5% copper extraction were realized at 25°C. Thus morecopper and cobalt were extracted in the inorganic acidenvironment than in the organic acid environment. However,the addition of hydrogen peroxide to the tartaric acid leachingsolution resulted in an 80% cobalt extraction and about 10%copper extraction under the same temperature conditions. Inthe tartaric acid leaching environment, the change intemperature had a much more pronounced effect on cobaltextraction than that of copper, with about 60% cobalt and10% cobalt extracted in the absence of hydrogen peroxide. Inaddition, changes in temperature had a more significanteffect on the extraction of cobalt in organic solutions than inthe inorganic environment.

The results obtained in these two leaching environmentsindicate the potential of tartaric acid to extract cobalt, ratherthan copper, from the copper-cobalt ores. It is recommendedthat further investigations be carried out with higher concen-trations of tartaric acid, since in the current test work theconcentration of tartaric acid was too low to achievemaximum leaching capability. Tests involving othercommonly used organic acids such as citric and oxalic acidwould also add value to this research area.

Acknowledgements

The authors would like to thank Gecamines in DRC for theprovision of the ore samples used in this study and theSchool of Chemical and Metallurgical Engineering at theUniversity of the Witwatersrand for technical assistance.

References

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GHARABAGHI, M., IRANNAJAD, M., and NOAPARAST, M. 2010. A review of thebeneficiation of calcareous phosphate ores using organic acid leaching.Hydrometallurgy, vol. 103. pp. 96–107

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▲


Figure 8—Effect of temperature on copper and cobalt extraction(tartaric acid concentration 0.35 M, leaching time 3 hours)

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[Accessed 23 October 2013]. ◆



SAIMM 120th Anniversary

T he Southern African Institute of Mining and Metallurgy (SAIMM) has redesigned our logo to coincide with our 120thAnniversary. This logo is more aligned with the changes over the last two decades, while maintaining the professionalism

that the SAIMM is renowned for. We have also emphasized the fact that we are 120 years old, and have continued tomaintain our technical excellence with regard to our Journal and the events that we organize. To add to these achievementswe continue to increase our membership.

The Parts of an Achievement of Arms and their SignificanceThe arms under consideration comprise separate parts, viz., Shield, Helm, Mantling,Crest, Supporters, Compartment, and Motto.

The Shield: This is in blue divided by a golden chevron, to represent the major sectionsof the industry. The flaming crucibles in the upper section represent Metallurgy and thecrossed pick and shovel in the lower section represent Mining.

The Helm: This is an Esquireʼs Helmet, which is the customary type of use for the arms of corporate bodies.

The Wreath and the Mantling: These are always in the two main ʻcoloursʼ of the shield, in this case gold as a metal and blueas the colour. The mantling was originally a short cloak draped from the helmet as a protection against the sun, and thewreath helped to hold the crest in place.

The Crest: This served as an additional mark of distinction. In this case the demi-lion represents strength and holds thenational flower of South Africa in his left Claw.

The Supporters: In this case heraldic beasts have been chosen, symbols of these ancient professions, the black lionrepresenting mining and the golden dragon representing metallurgy. The ʻdifferentʼ marks on their shoulders are carried overfrom the shield of the Chemical, Mining and Metallurgical Society, and their colours and the diamonds in their collars areintended to represent the main fields of mining in South Africa, namely gold, coal, and diamonds.

The Compartments: This is, appropriately, an outcrop of rock.

The Motto: ʻCapaci Occasioʼ has been taken over from the Instituteʼs predecessor, the Chemical, Mining and MetallurgicalSociety, with the exhortation, ʻto the capable the opportunityʼ.

2014◆ SEMINAR

Society of Mining Professors—A Southern African Silver Anniversary26–30 June 2014, The Maslow Hotel, Sandton, Gauteng

◆ SCHOOLMine Planning School15–16 July 2014, Mine Design Lab, Chamber of MinesBuilding, The University of the Witwatersrand

◆ CONFERENCEPyrometallurgical Modelling Principles and Practices4–5 August 2014, Misty Hills Country Hotel and Conference Centre, Cradle of Humankind

◆ CONFERENCEMinPROC 20146–8 August 2014, Cape Town

◆ CONFERENCEMineSafe Conference 201420–21 August 2014, Conference22 August 2014, Industry dayEmperors Palace, Hotel Casino Convention Resort,Johannesburg

◆ SCHOOL3rd Mineral Project Valuation School9–11 September 2014, Mine Design Lab, Chamber of MinesBuilding, The University of the Witwatersrand

◆ CONFERENCESurface Mining 201416–17 September 2014, The Black Eagle Room, Nasrec Expo Centre

◆ SCHOOLGrade control and reconciliation23–24 September 2014, Moba Hotel, Kitwe, Zambia

◆ CONFERENCESHAPE: 1st International Conference on Solids Handling andProcess Engineering29–30 September 2014, University of Pretoria, South Africa

◆ CONFERENCE6th International Platinum Conference20–234 October 2014, Sun City, South Africa

SAIMM DIARY

Forthcoming SAIMM events...

For further information contact:Conferencing, SAIMM

P O Box 61127, Marshalltown 2107Tel: (011) 834-1273/7

Fax: (011) 833-8156 or (011) 838-5923E-mail: [email protected]

For the past 120 years, the SouthernAfrican Institute of Mining andMetallurgy, has promoted technical

excellence in the minerals industry. Westrive to continuously stay at thecutting edge of new developments inthe mining and metallurgy industry.The SAIMM acts as the corporatevoice for the mining and metallurgyindustry in the South African economy.We actively encourage contact andnetworking between members and thestrengthening of ties. The SAIMMoffers a variety of conferences that aredesigned to bring you technicalknowledge and information of interestfor the good of the industry. Here is aglimpse of the events we have lined upfor 2013. Visit our website for moreinformation.

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2014

12–14 May 2014 — 6th South African Rock EngineeringSymposium SARES 2014Creating value through innovative rock engineeringMisty Hills Country Hotel and Conference Centre,Cradle of HumankindContact: Raymond van der BergTel: +27 11 834-1273/7, Fax: +27 11 838-5923/833-8156 E-mail: [email protected], Website: http://www.saimm.co.za

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24–31 May 2014 — ALTA 2014 Nickel-Cobalt-Copper,Uranium-REE and Gold-Precious Metals Conference & ExhibitionPerth, Western AustraliaContact: Allison Taylor E-Mail: [email protected], Tel: +61 (0)411 692-442 Website: http://www.altamet.com.au/conferences/alta-2013/

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20–22 August 2014 — MineSafe Conference 2014Technical Conference and Industry day20–21 August 2014: Conference22 August 2014: Industry dayEmperors Palace, Hotel Casino Convention Resort, JohannesburgContact: Raymond van der BergTel: +27 11 834-1273/7, Fax: +27 11 838-5923/833-8156 E-mail: [email protected], Website: http://www.saimm.co.za18–19 November 2014 — Third International EngineeringMaterials and Metallurgy Conference and Exhibition (iMat2014)Shahid Beheshti International Conference Center, Tehran, Iran Contact: Kourosh Hamidi E-mail: [email protected]

9–11 September 2014 — 3rd Mineral Project Valuation SchoolMine Design Lab, Chamber of Mines Building,The University of the WitwatersrandContact: Camielah Jardine, Tel: +27 11 834-1273/7, Fax: +27 11 838-5923/833-8156 E-mail:[email protected], Website: http://www.saimm.co.za16–17 September 2014 — Surface Mining 2014The Black Eagle Room, Nasrec Expo CentreContact: Camielah Jardine, Tel: +27 11 834-1273/7, Fax: +27 11 838-5923/833-8156, E-mail: [email protected], Website: http://www.saimm.co.za

2015March 2015 — PACRIM 2015Hong Kong, ChinaTel: +27 11 834-1273/7, Fax: +27 11 838-5923/833-8156, E-mail: [email protected]: http://www.pacrim2015.ausimm.com.au

14–17 June 2015 — European Metallurgical ConferenceDusseldorf, GermanyWebsite: http://www.emc.gdmb.de

14–17 June 2015 — Lead Zinc Symposium 2015Dusseldorf, GermanyWebsite: http://www.pb-zn.gdmb.de

16–20 June 2015 — International Trade Fair for MetallurgicalTechnology 2015Dusseldorf, GermanyWebsite: http://www.metec-tradefair.com

5–9 October 2015 — MPES 2015: 23rd InternationalSymposium on Mine Planning & Equipment SelectionSandton Convention Centre, Johannesburg, South AfricaContact: Raj SinghaiE-mail: [email protected] or Contact: Raymond van der BergTel: +27 11 834-1273/7, Fax: +27 11 838-5923/833-8156 E-mail: [email protected], Website: http://www.saimm.co.za

INTERNATIONAL ACTIVITIES

▲

xii APRIL 2014 The Journal of The Southern African Institute of Mining and Metallurgy

Company AffiliatesThe following organizations have been admitted to the Institute as Company Affiliates

AECOM SA (Pty) Ltd

AEL Mining Services Limited

Air Liquide (Pty) Ltd

AMEC GRD SA

AMIRA International Africa (Pty) Ltd

ANDRITZ Delkor(pty) Ltd

Anglo Operations (Pty) Ltd

Anglogold Ashanti Ltd

Arcus Gibb (Pty) Ltd

Atlas Copco Holdings South Africa (Pty) Limited

Aurecon South Africa (Pty) Ltd

Aveng Mining Shafts and Underground

Aveng Moolmans (Pty) Ltd

Bafokeng Rasimone Platinum Mine

Barloworld Equipment -Mining

BASF Holdings SA (Pty) Ltd

Bateman Minerals and Metals (Pty) Ltd

BCL Limited (BCL001)

Becker Mining (Pty) Ltd

BedRock Mining Support Pty Ltd

Bell Equipment Limited

BHP Billiton Energy Coal SA Ltd

Blue Cube Systems (Pty) Ltd

Bluhm Burton Engineering Pty Ltd

Blyvooruitzicht Gold Mining Company Ltd

BSC Resources Ltd

CAE Mining (Pty) Limited

Caledonia Mining Corporation

CDM Group

CGG Services SA

Chamber of Mines

Concor Mining

Concor Technicrete

Council for Geoscience

CSIR Natural Resources and theEnvironment

Department of Water Affairs and Forestry

Deutsche Securities (Pty) Ltd

Digby Wells and Associates

Downer EDI Mining

DRA Mineral Projects (Pty) Ltd

Duraset

E+PC Engineering and Projects Company Ltd

Elbroc Mining Products (Pty) Ltd

eThekwini Municipality

Evraz Highveld Steel and Vanadium Limited

Exxaro Coal (Pty) Ltd

Exxaro Resources Limited

Fasken Martineau

FLSmidth Minerals (Pty) Ltd (FFE001)

Fluor Daniel SA ( Pty) Ltd

Franki Africa (Pty) Ltd-JHB

Fraser Alexander Group

Goba (Pty) Ltd

Hall Core Drilling (Pty) Ltd

Hatch (Pty) Ltd

Herrenknecht AG

HPE Hydro Power Equipment (Pty) Ltd

Impala Platinum Holdings Limited

IMS Engineering (Pty) Ltd

JENNMAR South Africa

Joy Global Inc.(Africa)

Leco Africa (Pty) Limited

Longyear South Africa (Pty) Ltd

Lonmin Plc

Ludowici Africa (Pty) Ltd

Wekaba Engineering (Pty) Ltd

Magnetech (Pty) Ltd

MAGOTTEAUX (PTY) LTD

MBE Minerals SA Pty Ltd

MCC Contracts (Pty) Ltd

MDM Technical Africa (Pty) Ltd

Metalock Industrial Services Africa (Pty)Ltd

Metorex Limited

Metso Minerals (South Africa) (Pty) Ltd

Minerals Operations Executive (Pty) Ltd

MineRP

Mintek

Modular Mining Systems Africa (Pty) Ltd

MSA Group (Pty) Ltd

Multotec (Pty) Ltd

Murray and Roberts Cementation

Nalco Africa (Pty) Ltd

Namakwa Sands (Pty) Ltd

New Concept Mining (Pty) Limited

Northam Platinum Ltd - Zondereinde

Osborn Engineered Products SA (Pty) Ltd

Outotec (RSA) (Proprietary) Limited

PANalytical (Pty) Ltd

Paterson and Cooke Consulting Engineers

Paul Wurth International SA

Polysius A Division Of ThyssenkruppEngineering

Precious Metals Refiners

Rand Refinery Limited

Redpath Mining South Africa (Pty) Ltd

Rosond (Pty) Ltd

Royal Bafokeng Platinum

Roymec Technologies (Pty) Ltd

RSV Misym Engineering Service (Pty) Ltd

RungePincockMinarco Limited

Rustenburg Platinum Mines Limited

SAIEG

Salene Mining (Pty) Ltd

Sandvik Mining and Construction Delmas(Pty) Ltd

Sandvik Mining and Construction RSA(Pty)Ltd

SANIRE

Sasol Mining (Pty) Ltd

Scanmin Africa (Pty) Ltd

Sebilo Resources (Pty) Ltd

SENET (Pty) Ltd

Senmin International (Pty) Ltd

Shaft Sinkers (Pty) Limited

Sibanye Gold Limited

Smec SA

SMS Siemag

SNC Lavalin (Pty) Ltd

Sound Mining Solution (Pty) Ltd

SRK Consulting SA (Pty) Ltd

Time Mining and Processing (Pty) Ltd

Tomra Sorting Solutions Mining (Pty) Ltd

TWP Projects (Pty) Ltd

Ukwazi Mining Solutions (Pty) Ltd

Umgeni Water

VBKOM Consulting Engineers

Webber Wentzel

Weir Minerals Africa (Pty) Ltd

Xstrata Coal South Africa (Pty) Ltd

Universiteit van Pretoria • University of Pretoria • Yunibesithi ya PretoriaPrivaatsak • Private Bag • Mokotla wa Poso X20 Hatfield 0028Suid-Afrika • South Africa • Afrika BorwaTel: +27 (0) 12 420 4111 • Fax • Fekse: +27 (0) 12 420 4555 www.up.ac.za

University of PretoriaDepartment of Mining Engineering

Educating and leading mining engineersinto the futureOur broad-based curriculum emphasises the interna-tional engineering education model of conceptualising, designing, implementing and operating mines. The undergraduate programme is accredited by the Engineering Council of South Africa (ECSA).

We offer a range of programmes:• Undergraduate• Postgraduate• Research• Continued Professional Development (CPD) courses

A degree at the University of Pretoria covers all aspects of mining engineering from pre-feasibility assessments

to closure of mines, whether it be hard rock or coal deposits, surface or underground operations. The development of management/people skills is incorpo-rated in some of the final year modules in the under-graduate programme.

Our portfolio of undergraduate, postgraduate and research degree programmes provide an excellent basis for a rewarding career in mining.

For more information contact the head of department, Prof Ronny Webber-Youngman on 012 420 3763 or e-mail [email protected] or visit our website www.up.ac.za.

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