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VOLUME 114 NO. 5 MAY 2014

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VOLUME 114 NO. 5 MAY 2014

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Tega Industries (South Africa) Pty LtdP.O Box 17260, Benoni West, 1503, South Africa,

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Email: [email protected],www.tegaindustries.com

Tega offers value added consultancy services and solutions TOTAL : Solution

With focus on core engineering applications in the Mining and Mineral Processing Industry, Steel plants, Power, Port and Cement Industries.

in Mineral Beneficiation, Bulk Solids handling, Wear andAbrasion customised to suit specific applications.

TM

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Collision Avoidance Systems

Safety Access Systems

Safety Isolation Boxes

Asset Protection for vehicles, infrastructure and personnel

Risk Mitigation aid that delivers improved safety and productivity performance

Real-time risk monitoring and reporting of “High Potential Interactions”

Improved “situational awareness” reduces operator stress

Complies with MDG15 standards

Safe easy access

Robust, able to withstand vibration

Interlock Protection

Emergency egress

Optional roll over isolation switch

Ridgid folded corner construction

Easy access- top, front or back

Customisable

Complies with EMESRT recommendations

Corner Hans Strydom & Van Rensburg Street Witbank, Gauteng Tel: +27 (0) 13 692 4161 Fax: +27 (0) 13 692 3337

www.probeIMT.co.za

SAFETY AND PRODUCTIVITY

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ii MAY 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]

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ContentsJournal Commentby P. Smith . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . v

President’s Corner by M. Dworzanowski . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . vii

Special ArticlesSouth African Mineral Resource Committee (SAMREC): Re-write of the SAMREC Code (2014)by K. Lomberg. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . viii

Sulphuric acid plant water saving optionsby R.J. Forzatti, I. Natha, L. Roux, and D.A. van den Berg . . . . . . . . . . . . . . . . . . . . . . . . . . . . 355

Challenges and successes at the Nkomati Nickel JV: pit-to product process improvementsby G. Cockburn . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 365

Evaluation of a Versatic 10 acid/Nicksyn™ synergistic system for the recovery of nickel and cobalt from a typical lateritic leach liquorby A.C. du Preez and M.H. Kotze . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 375

Evaluation of different adsorbents for copper removal from cobalt electrolyteby V. Yahorava, M. Kotze, and D. Auerswald . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 383

Thermodynamic analysis and experimental study of manganese ore alloy and dephosphorization in converter steelmakingby G. Chen and S. He. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 391

Atmospheric oxidative and non-oxidative leaching of Ni-Cu matte by acidified ferric chloride solutionby L.M. Sekhukhune, F. Ntuli, and E. Muzenda . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 401

Universities and decision-making: programme and qualification mix – four learning pathwaysby W.P. Nel . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 411

Erratum: ‘A study on the effect of coke particle size on the thermal profile of the sinters produced in Esfahan Steel Company (ESCO)’, by A. Dabbagh, A. Heidary Moghadam, S. Naderi, and M. Hamdi . . . . . . . . . . . . . . . . . . . . . . 418

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 MAY 2014

VOLUME 114 NO. 5 MAY 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. StaceyR.J. Stewart

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 th is publ icat ion 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 appropr iate, the copyr ight) isacknowledged. Apart from any fair dealingfor the purposes of review or criticismunder The Copyright Act no. 98, 1978,Section 12, of the Republ ic 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 art icle appear ing in th is journalind icates that the copyr ight 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. Th is consent does notextend to other kinds of copying, such ascopy ing for general d istr ibut ion, foradvertising or promotional purposes, forcreat ing new collect ive works, or forresale.

VOLUME 114 NO. 5 MAY 2014

Base Metals papers

General papers

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

The 7th Southern African Base Metals Conference washeld in Mpumalanga from 2 to 3 September 2013, witha visit to Nkomati Nickel Mine on 3 September. TheConference attracted 22 papers from sub-Saharan Africa(DRC, Zambia, Namibia, and South Africa) as well asfrom Finland and Australia. The scope of the papers waswide-ranging, including geology, engineering design,and process metallurgy. This edition of the SAIMMJournal includes a selection of four papers from theconference. It is regrettable that four papers werewithdrawn from publication for a variety of reasons atthe request of the authors.

Sulphuric acid plant water savings options by R.J.Forzatti, et al. considers strategies for saving water insulphur-burning acid plants, although the principlesalso apply to acid plants burning sulphide minerals suchas pyrite, sphalerite, etc. Acid plants basically involveexothermic reactions on a huge scale, and requireextensive cooling. The paper presents an economicevaluation of various options to achieve this goalagainst various backgrounds of localized power andwater costs. Given the current sensitivity toenvironmental factors throughout the world, this paperprovides an important contribution to the debate. Ingeneral, mines tend to be found in remote locationswhere both power and water are at a premium andindeed, sometimes not available at all. The northernregions of Chile provide a good example of this with thechallenges of the Atacama Desert.

Challenges and successes at the Nkomati Nickel JV:pit to product process improvements by G. Cockburn isof particular interest in that the deposit was firstsubjected to a feasibility study in the 1970s(INCO/Anglo American). Several subsequent attemptsall failed to build a viable case. However, today thereexists a successful operating mine with a growth profilethat can only be described as spectacular – 10 000t/month to 700 000 t/month. A mine-to-milloptimization programme is described along with theconsequent benefits – significantly the effect of blast-hole patterns and explosive powder factors on primarycrusher feed size distribution and the critical role that astable plant throughput plays in milling and flotationperformance.

Evaluation of a versatic 10 acid/Nicksyn™synergistic system for the recovery of nickel and cobaltfrom a typical lateritic leach liquor by A.C. Du Preez andM.H. Coetzee is a valuable contribution to high-pressureacid leach (HPAL) nickel technology, given the

probability that nickel lateritic ores will be increasinglyexploited in the future in comparison to sulphide ores.Although HPAL was first implemented many decadesago at Moa Bay in Cuba, the technology has faced manytechno-economic challenges. One has been thedifficulties associated with calcium as encountered inthe Australian ‘dry lateritic’ projects during the 1990s.This paper presents the results of laboratory test workwhich demonstrates that calcium co-extraction can beavoided in the extraction and stripping stages of thesolvent extraction plant by judicious selection ofreagents and operating conditions.

Evaluation of different absorbents for copperremoval from cobalt electrolyte by V. Yahorava, et al.Several fibrous ion exchangers were investigated andcompared to the more conventional granular ionexchangers for the removal of copper from cobaltelectrolytes. A comparison of the design parameters andindicative costs for the impurity removal process ispresented for the two alternatives.

P. Smith

Journal Comment

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vi MAY 2014 The Journal of The Southern African Institute of Mining and Metallurgy

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

The South African mining industry has been mainly associatedwith gold, which is understandable given that it was the goldmines of the late nineteenth century that were really the

beginning of the industry as we know it. Currently platinum mining ismaking headlines regarding long strikes, and coal mining and Eskomare also much in the news. Diamonds have always featured in the

media, to an extent that varies with time. However, base metals havenever received any prominence, although they have formed part of the industry’s

contribution for many decades. Copper, nickel, lead, and zinc are produced in South Africa.Palabora Mining Company produces copper, Nkomati Nickel produces nickel, and Black Mountainproduces copper, zinc, and lead. In addition, copper, nickel, and cobalt are by-products from the basemetals refineries associated with the four major platinum producers.

Copper is produced in the form of cathodes from electrorefining and electrowinning. Nickel isproduced in the form of metal cathodes by electrowinning, or as metal briquettes or nickel sulphate.Cobalt is produced in the form of metal briquettes or cobalt sulphate. There is no primary production oflead metal, and there is no longer any primary production of zinc metal since the closure of Zincor at theend of 2011. With the exception of lead and zinc, the beneficiation of base metal ores in South Africa isthus well developed.

The beneficiation of base metal ores will always provide a significant challenge to extractivemetallurgists – I speak from personal experience with many base metal projects in southern Africa. Theflow sheet options are considerable, and in many instances mineral processing, pyrometallurgy, andhydrometallurgy need to be applied.

If we broaden our view of base metals to the southern African region, then we see a copper andcobalt industry of global significance. The Central African Copperbelt that spans Zambia and theDemocratic Republic of Congo (DRC) is a world-class mineral province. Although mined tonnage doesnot compare with the copper porphyries of North and South America, the higher copper grades mean thatactual copper production is not that far behind. Zambia and the DRC are 5th in global copper productionwhen their output is combined. Beneficiation of copper ores, both sulphide and oxide, in Zambia and theDRC is well developed, with most of the copper being produced as cathode metal via electrorefining orelectrowinning. There is a significant diversity of copper concentrators, smelters, and refineries withinZambia and the DRC. When the Nchanga tailings leach complex in Zambia was originally built close to40 years ago it boasted the world’s largest copper solvent extraction and electrowinning plant. At Ndolain Zambia is one of the world’s few refineries processing copper refinery anode slimes, producingselenium, tellurium, silver, and gold as by-products.

The Copperbelt constitutes the world’s largest deposit of cobalt, which is associated with the copperin oxide and sulphide minerals. There are a number of cobalt plants which beneficiate oxide andsulphide cobalt concentrates into cathode metal via electrowinning. Zambia and the DRC produce abouthalf the world’s cobalt.

There are nickel mines in Botswana and Zimbabwe, and zinc, lead, and copper mines in Namibia.Zinc and lead are also produced on a small scale in Zambia and the DRC. This all highlights the extent ofbase metals production in southern Africa and illustrates why the SAIMM organizes a base metalsconference every two years. The next conference will be held in Zambia in 2015, with a copper / cobalttheme. The event will be convened in conjunction with our Zambia and DRC branches, and will serve asa vehicle for promoting them, as well as provide a means to motivate the hosting of the InternationalCopper Conference in Southern Africa.

M. DworzanowskiPresident, SAIMM

Presidentʼs

Corner

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

South African Mineral Resource Committee(SAMREC)

Re-write of the SAMREC Code (2014)

The mining industry is a vital contributor to national and global economies; never more so than at present with soaringdemand for the commodities that it produces. It is a truly international business that depends on the trust and confidenceof investors and other stakeholders for its financial and operational well-being. Unlike many other industries, it is based

on depleting mineral assets, the knowledge of which is imperfect prior to the commencement of extraction. It is thereforeessential that the industry communicates the risks associated with investment effectively and transparently in order to earn thelevel of trust necessary to underpin its activities. (CRIRSCO website)

The SAMREC Code, which sets out minimum standards, recommendations, and guidelines for Public Reporting ofExploration Results, Mineral Resources and Mineral Reserves in South Africa, is being reviewed and improved to ensure that itremains relevant to the minerals industry and keeps abreast with recent developments. This revision is considered necessarybecause as the guidelines of the Code are used, various issues and practical realities have become apparent that require furtherguidance from the Code. This rewrite is designed to improve the Code and eliminate possible contradictory reporting practices,and align SAMREC with recent changes to international codes in keeping with international best practice.

The SAMREC Code is one of seven codes that are affiliated under the CRIRSCO family of reporting codes. As a result of theCRIRSCO/CMMI initiative, considerable progress has been made towards widespread adoption of globally consistent reportingstandards. These are embodied in similar Codes, guidelines, and standards published and adopted by the relevant professionalbodies around the world. The definitions in this edition of the SAMREC Code are either identical to, or not materially differentfrom, existing international definitions. In recent years the Russian Code (NAEN) (2011) was added to the original Codes.Various Codes have been revised and reissued – CIM of Canada (2010), PERC representing Europe (2013), JORC representingAustralia and New Zealand (2012), and SME representing the USA (under review for issuing in 2014).

Various aspects of the Code remain unchanged. Because SAMREC is part of the CRIRSCO family, there are 15 coredefinitions e.g. Mineral Resource, Mineral Reserve etc. that are common between the international codes. These are not beingchanged. Rather, the guidance and interpretation is being improved so that the Code is relevant. The Code remains a guidelinefor minimum public reporting of Exploration Results, Mineral Resources, and Mineral Reserves. The desire of the SAMRECWorking Group is that the Code is used for all forms of reporting of Exploration Results, Mineral Resources, and MineralReserves, both public and private. The principles that underpin the code remain Transparency, Materiality, and Competence.The Code requires that anyone who uses the Code and asserts themselves as a Competent Person (CP) in accordance with theCode needs to have five years’ relevant experience and be registered with SACNASP or ECSA or be a member of GSSA, SAIMM,or PLATO or a recognized professional organization (RPO).

A body whose members put themselves forward as CPs is required to have a code of ethics and a disciplinary code.Scientists working in South Africa are required to comply with the Natural Scientific Professions Act of 2003. However, wherethe SAMREC Code is used as the basis for a mineral resource or reserve declaration that falls outside of the jurisdiction of SouthAfrica laws and the CP declares his/her membership of GSSA or SAIMM in support of the declaration, then these organizationsrequire the CP to follow the newly instituted procedure.

Because the GSSA and SAIMM are not statutory bodies and represent broader interests than just minerals reporting, theGSSA and SAIMM have introduced by-laws that require individuals who utilize their membership as a credential for reportingpurposes to notify the societies and subject themselves to a peer review prior to the publication of the work. This peer reviewentails confirming that they are members of the societies in the category they claim and have the necessary qualification andexperience to undertake this assignment as a CP. However, this does not militate against the individual producing work that issubstandard. Should the individual complete substandard work and a complaint is laid, they will be subject to the disciplinaryprocess.

Issues regarding the rewrite are discussed at a monthly meeting of the SAMREC Working Group (WG) chaired by KenLomberg ([email protected]) and held on the last Thursday of each month at the Military Museum in Saxonwold. Allinterested parties are invited to participate. These meetings also provide an opportunity for industry to highlight aspects thatmay need to be reviewed or improved. We would like to encourage all interested parties to submit any issues relevant to therewrite of the SAMREC Code via the SAIMM ([email protected]) by 30 June 2014. The intention is to complete a draft for publiccomment by the end of Q3 2014.

Once a draft has been finalized it will be issued for comment prior to being ratified by the SAMREC/SAMVAL Committee(SSC). It is also the intention of the SAMREC WG to prepare a companion volume that would include the practical application ofthe Code and assist in providing a benchmark for all industry practices. This volume is likely to be produced after the launch ofthe Code as the proceedings of a SAMREC conference.

K. Lomberg

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PAPERS IN THIS EDITIONThese papers have been refereed and edited according to internationally accepted standards and areaccredited for rating purposes by the South African Department of Higher Education and Training

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

Base Metals papersSulphuric acid plant water saving optionsby R.J. Forzatti, I. Natha, L. Roux, and D.A. van den Berg . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 355This paper considers several alternative cooling water systems for a conceptual 2000 t/d sulphuric acid plant. A proprietary design tool is used to compare design options on both an economic and a weighted sustainability scale.

Challenges and successes at the Nkomati Nickel JV: pit-to product process improvementsby G. Cockburn . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 365The metallurgical, operational, and management challenges involved in a number of milling and flotation optimizationinitiatives at the Nkomati Nickel JV operation are discussed, together with the outcomes obtained.

Evaluation of a Versatic 10 acid/Nicksyn™ synergistic system for the recovery of nickel and cobalt from a typical lateritic leach liquorby A.C. du Preez and M.H. Kotze. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 375A synergistic solvent extraction system is evaluated on a laboratory scale for the recovery of nickel and cobalt from a synthetic lateritic sulphate leach liquor, without the co-extraction of calcium.

Evaluation of different adsorbents for copper removal from cobalt electrolyteby V. Yahorava, M. Kotze, and D. Auerswald . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 383Granular and fibrous ion exchangers are compared for the removal of copper from cobalt advance electrolyte. The results are used to size a full-scale operation and carry out a techno-economic evaluation of the two ion exchange systems.

General papersThermodynamic analysis and experimental study of manganese ore alloy and dephosphorization in converter steelmakingby G. Chen and S. He . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 391In this study, the effects of slag compositions, slag amount, temperature, and the carbon content of steel on the manganese and phosphorus distribution ratios during converter steelmaking are analysed. The results of the research could be useful in deciding on the application of manganese ore in alloying and identifying the slagging regime in converter steelmaking.

Atmospheric oxidative and non-oxidative leaching of Ni-Cu matte by acidified ferric chloride solutionby L.M. Sekhukhune, F. Ntuli, and E. Muzenda . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 401The atmospheric leaching of nickel from a copper-bearing matte by acidic ferric chloride solution was studied at the laboratory scale. Leaching was found to be diffusion-controlled, and took place via three separate mechanisms that occurred simultaneously throughout the process. Oxidative leaching yielded higher nickel recoveries than non-oxidativeleaching.

Universities and decision-making: programme and qualification mix – four learning pathwaysby W.P. Nel . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 411Many factors have to be considered when deciding on a Programme Qualification Mix (PQM). This paper focuses on the Higher Education Qualifications Sub-Framework (HEQSF) requirements, Engineering Council of South Africa (ECSA) standards and registration, and how these, together with the various qualifications and educational Learning Programmes (LPs) provided for by the HEQSF, may impact on the PQM decision taken by engineering departments and schools at South African universities.

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Introduction

Conventional sulphuric acid plants requirewater for their cooling systems. Cooling isrequired to reject surplus heat not recovered assteam. Reducing the water consumption lowersthe cost of sourcing reliable supplies of cleanwater as well as the cost associated withtreating effluent streams. It also helps improvethe sustainability of the acid plant operationby reducing the impact on surroundingcommunities.

This paper focuses on two broad categoriesof water-saving options:

➤ Pretreatment of the make-up waterrequired for evaporative cooling systems

➤ Replacement of evaporative cooling withdry cooling.

The sustainability aspect of these optionsis analysed using Hatch’s 4 Quadrant designtool.

Technologies integrated with the acid plantdesign that recover heat from acid will reducethe overall cooling water demand (and hencemake-up water consumption). For thepurposes of this evaluation, the following acidplant technology options are not considered:

➤ HRS (by MECS) and HEROS (byOutotec), which produce useful low-pressure steam

➤ Heat recovery acid coolers, whichpreheat boiler feed water.

Overview

A conventional 2000 t/d sulphur-burningsulphuric acid plant is considered in this paper.

Process description

Solid sulphur is delivered in bulk bags orcontainers and is stored and transferred to amelting and filtration circuit. The sulphur-melting system uses low- and medium-pressure (LP and MP) steam, usually providedfrom the acid plant steam system. Dirty moltensulphur is filtered to remove ash and othersolid impurities. Molten sulphur is alsosometimes received instead of solid sulphur ifthe sulphur source is nearby.

Clean sulphur is transferred to the acidplant where sulphuric acid is produced. Thesulphur is burned in a furnace at approxi-mately 1200°C in contact with dry air toproduce SO2 gas (approx. 12 vol.%) (King,Davenport, and Moats, 2013). The SO2 gas isoxidized to SO3 gas in contact with avanadium pentoxide-type catalyst. The SO3 isthen absorbed and reacts with the aqueouscomponent of strong sulphuric acid to produceH2SO4. Circulating and product acid cooling isachieved in heat exchangers supplied withcooling water, typically provided from anevaporative cooling tower.

The acid plant steam system is designed torecover the heat generated by the exothermic

Sulphuric acid plant water saving optionsby R.J. Forzatti*, I. Natha†, L. Roux†, and D.A. van den Berg*

SynopsisThe production of sulphuric acid from sulphur generates heat. The majorityof this heat is recovered as steam and is often used to generate electricity.Heat not recovered as steam is rejected to cooling water systems. Thedesign of the turbine, condenser, and cooling water systems impacts theoverall water, energy, and environmental footprint of the plant. This reviewconsiders a conceptual 2000 t/d sulphuric acid plant with severalalternative cooling water systems. The review utilizes Hatch’s 4 Quadrantsustainable design tool to compare the alternatives on both an economicand a weighted sustainability scale.

Keywordssulphuric-burning acid plant, steam, power generation, cooling systems,sustainable design.

* Hatch Associates, Perth, Western Australia.† Hatch Goba, Woodmead, South Africa.© The Southern African Institute of Mining and

Metallurgy, 2014. ISSN 2225-6253. This paperwas first presented at the, Base Metals Conference2013, 2–4 September 2013, IngwenyamaConference & Sports Resort, Mpumalanga.

355The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 MAY 2014 ▲

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Sulphuric acid plant water saving options

reactions within the acid plant. Heat is recovered from thesulphur burner off-gas via the production of saturated high-pressure (40 barg or 60 barg) steam in a waste heat boiler.Saturated steam from the boiler flows through a superheaterto produce superheated steam, which is fed to a steamturbine generator to produce electricity.

Exhaust steam from the acid plant turbine is condensed,re-pressurized, and returned to the acid plant as boiler feedwater. The turbine steam condenser (‘surface’ condenser)uses cooling water from an evaporative cooling tower. Theevaporative cooling tower loses water through evaporation,drift (entrainment), and blowdown. A continuous supply offresh water is required to make up for these losses.

Demineralized water is used as make-up for losses withinthe steam circuit (e.g. boiler blowdown and deaerator vent)and for dilution water within the acid plant.

Typical operating parameters for a 2000 t/d sulphur-burning acid plant are shown in Table I.

Electricity generation

Steam produced by the acid plant can be:

➤ Used to generate electricity in a steam turbinegenerator

➤ Supplied to other plant consumers (e.g. heating forhydrometallurgical equipment)

➤ Exported to other customers➤ Condensed.

Sulphur-burning acid plants produce more electricity thanthey consume when all of the steam is sent to a steamturbine generator. This excess electricity can be:

➤ Used to operate other facilities within the plant➤ Sold to the market.

Water balance

The water balance for a conventional 2 000 t/d sulphur-burning acid plant based on evaporative cooling is given inTable II.

The following is noted:

➤ The single largest water loss is due to waterevaporation in the cooling tower

➤ Other losses include drift and blowdown. Theblowdown indicated is calculated based on three cyclesof concentration, assuming fresh water has a totaldissolved solids (TDS) of 300–400 ppm.

Water-saving options

Make-up water pretreatment

Treatment of make-up water to the cooling tower can be usedto change the water chemistry to achieve higher cycles ofconcentration, thereby reducing blowdown.

Softening

Softening of the cooling water make-up can be used toremove several dissolved salts that cause scale formationsuch as calcium, magnesium, barium, strontium, and iron.Other scale-forming components, such as silica, are notremoved.

A water softener consists of a vessel filled with cationicresin that exchanges (removes) the dissolved species fromthe water and replaces these with sodium. Cooling systemsfed with high hardness water sources will benefit most fromhaving the make-up water softened.

As an example, a water source with a feed hardness ofapproximately 300 mg/L (expressed as CaCO3), pH of approx-imately 8, and alkalinity of approximately 300 mg/L as CaCO3might normally be concentrated three times; a cycles ofconcentration (COC) of 3. The scaling tendency of this water,at a COC of 3, is within the typical range that can be managedwith a scale inhibitor. This same make-up water source could

356 MAY 2014 VOLUME 114 The Journal of The Southern African Institute of Mining and Metallurgy

Table I

Typical 2 000 t/d acid plant operating parameters

Parameter Units Value

Acid production (100% H2SO4 basis) t/d 2 000

Sulphur consumption t/d 660

Steam production (superheated) t/h 110

Electricity generation (steam turbine) MWe 23

Electricity consumption MWe 5

Cooling requirements

Main acid coolers 106 kJ/h 144

Product acid cooler 106 kJ/h 7

Turbine surface condenser 106 kJ/h 234

Other coolers 106 kJ/h 11

Nominal cooling duty 106 kJ/h 396

Design cooling duty* 106 kJ/h 468

*Design cooling duty includes an additional 72 x 106 kJ/h installedcapacity for when the steam turbine is bypassed

Table II

Sulphuric acid plant water balance

Inputs Inflow H2O (t/h) Outputs Outflow H2O (t/h)

Air moisture (to sulphur burner) 4 Steam deaerator vent 2Cooling tower make-up water 220 Water in product acid 1Water to demin plant (for acid dilution) 24 Water converted to H2SO4 by reaction 15

Cooling tower evaporation and drift loss 147Cooling tower blowdown 73

Other effluent (steam system and demin plant) 10Total Inputs 248 Total outputs 248

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be concentrated more than seven times if first softened,representing a reduction in cooling water make-up of approximately 20%.

Filtration

Filtration to remove suspended solids from water can beapplied to either the entire fresh water make-up stream forthe acid plant or to the cooling tower water recirculationstream. Filtration of the cooling circuit make-up water isgenerally considered where there is a high level of suspendedsolids in the feed. These solids, if not removed, can causefouling within the cooling water circuit, which lowers coolingefficiency and increases the pressure drop through the pipingsystem. The solids can also accelerate corrosion within thewater circuit if they are abrasive. Unfiltered particles canserve as nucleation sites for biological growth. Filtration isrequired ahead of a cooling water softening system.

It should be noted that the majority of suspended solidsin the cooling circuit are generated within the cooling circuitrather than introduced in the make-up water, as the coolingwater is in contact with surrounding air in open-circuitevaporative cooling towers. Internal sources of solids includepipe corrosion products, biological growth material, and dustintroduced from the air as it contacts the water in the coolingwater tower. For this reason there is often more merit infiltering the cooling water itself rather than the make-upwater alone.

Demineralized water treatment

Demineralized water is used as make-up for losses within thesteam circuit (e.g. boiler blowdown and de-aerator vent) andfor dilution water within the acid plant. Typical demineralizedwater system configurations include:

➤ Reverse osmosis (RO) only➤ Ion exchange (IX) with a decarbonator tower➤ RO followed by polishing IX.

Waste generation as a percentage of feed is typically 30%.The selected configuration is dependent on the site raw waterquality, and can be optimized to provide water savings. Thesesavings will, however, be small compared to potential savingsin the cooling system.

Cooling technologies

The following cooling technologies are discussed:

➤ Evaporative cooling towers➤ Dry cooling technologies➤ Hybrid cooling towers.

Evaporative cooling towers

Evaporative cooling tower designers have identified manyways to reduce the overall water losses from these systems.Some of these include (EnduroSolv, 2012):

➤ Optimizing water chemistry to reduce scaling,corrosion, and biological growth, subsequentlyincreasing the cycles of concentration and decreasingblowdown. This includes the use of automatedchemical dosing systems

➤ Better operating procedures and equipment to monitorand control blowdown

➤ Use of high-efficiency drift eliminators and equipmentto recapture drift

➤ Optimizing the selection and amount of fill inside thetower, which affects the heat transfer efficiency of thetower

➤ Automatic blowdown based on conductivity to avoidunnecessary blowdown in cases where the feed waterquality is better than initially anticipated

➤ Minimizing unintentional water losses from leaks oroverflow (i.e. faulty level control resulting in additionof excess make-up water)

➤ Special tower design considerations to reduce partic-ulates, debris, and cooling water exposure to sunlight.

These advances in cooling tower design and control haveresulted in minor water savings. Fundamentally, the coolingis provided through the evaporation of water, and hencethere is an inherent loss of water when adopting thistechnology.

Dry cooling technologies

Dry cooling technologies work by heat exchange to air and donot rely on the evaporation of water to provide cooling.Applicable technologies for a sulphur-burning acid plantinclude:

➤ Air-cooled condenser (ACC) on the steam turbineexhaust,

➤ Fin-fan coolers to supply cooling water to theabsorbing acid coolers.

Dry cooling technologies are dependent on the differencebetween ambient temperature and the cooling watertemperature. In locations with high ambient temperatures,the temperature difference will be lower, leading to signifi-cantly increased dry cooling unit size.

Air-cooled condenser (ACC)

An ACC (Figure 1) is comprised of finned tube bundlesgrouped together into modules and mounted in an A-frameconfiguration on a concrete or steel support structure. Steamfrom the turbine exhaust enters the top of the condenser viaa steam duct and manifold. Steam flows downward throughtwo or three rows of finned tubes. Condensate is recoveredinside bonnet header boxers connected to a hot water tank.The axial-flow forced-draught fan is fixed in the module andforces the atmospheric cooling air across the condensate areaof the fin tubes.

Sulphuric acid plant water saving options

357The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 MAY 2014 ▲

Figure 1—Schematic of an ACC (SPX, 2012)

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Sulphuric acid plant water saving options

Dry cooling in an ACC requires a significant temperaturedifference to provide adequate heat exchange to thesurrounding air. Typically, the cooling water supplytemperature will be 25°C to 30°C higher than the ambient airtemperature. This results in a higher condenser outlettemperature which in turn raises the condenser pressure,causing the steam turbine to operate less efficiently. An ACCcan also be impacted by wind direction and speed as well asproximity to large buildings. More recent advancements inACC technology include (Mortensen, 2011; Maulbetsch,DiFilippo, and Zammit, 2008):

➤ Wind guide vane technologies to mitigate wind impactsincluding walls, screens, lips, and louvers. CFD windflow modelling is also used to optimize the location andarrangement of the ACCs

➤ Improved finned tube bundle designs for higher heattransfer efficiency and lower pressure drop

➤ Pre-cooled ACC, which uses the evaporative coolingeffect of a fog spray into the upstream side of the ACCfans. The expected water consumption is approximately75% less than equivalent evaporative-only cooling. Theadvantage of this system is that it reduces air temper-atures to the fans on very hot days.

The application of an ACC for cooling of turbine exhaustis widely adopted on many steam turbine systems. Eskom,the South African power utility, has adopted the largest ACCscurrently in operation in the world for the Matimba, Kendal,and Majuba power stations (Eskom, 2010).

Fin-fan coolers

Fin-fan coolers (as depicted in Figure 2) include one or morebundles of finned tubes connected by headers with an air-moving device such as an axial fan located above (induceddraught) or below (forced draught) the tube bundle. Coolingwater flows through the tubes and heat is exchanged toambient air. The fin-fan circuit uses demineralized qualitywater and is closed-loop (not open to atmosphere),eliminating the need for a continuous water supply.

Fin-fan coolers require a significant temperaturedifference to provide adequate heat exchange to thesurrounding air. Typically, the cooling water supplytemperature will be 15°C higher than the ambient airtemperature. A fin-fan cooler can be used to provide cooling

for the drying and absorption sections of an acid plantbecause the absorbing acid heat exchangers target approxi-mately 70°C. The product acid heat exchangers target35–40°C, which cannot be consistently achieved in mostlocations using fin-fan coolers.

Hybrid cooling towers

Hybrid cooling towers have an air-to-air (dry cooling) sectionand evaporative cooling section operating in series. As shownin Figure 3, heated cooling water first passes through the drysection, where part of the heat load is removed by an aircurrent, typically induced via fans. After passing the drysection, water is further cooled in the wet section of thetower, which can be cooled in a conventional openevaporative circuit or closed circuit (tubes are cooled withwater on the outside).

The resulting heat transfer performance is similar to a wetcooling tower, with the dry cooler providing the advantage ofprotecting the working fluid from environmental exposureand contamination. Depending on the hybrid tower configu-ration, the water consumption lies between the wet and drycircuit options reported in this evaluation.

Cooling technology and steam turbine electricitygeneration

The turbine exhaust cooling system performance directlyaffects the amount of electricity produced by the steamturbine generator. The lower the condenser outlet

358 MAY 2014 VOLUME 114 The Journal of The Southern African Institute of Mining and Metallurgy

Figure 2—Fin-fan cooler (Wilson, 2011) Figure 3—Hybrid cooler examples (after EPRI, 2002)

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temperature, the lower the condenser outlet pressure andturbine exhaust back-pressure. A lower turbine exhaustback-pressure increases turbine output.

The impact of increasing ambient temperature (dry-bulbtemperature) on the turbine output is shown in Figure 4,similar to the loss of electricity generation reported by others(US Department of Energy. 2009). It compares the base case(turbine surface condenser on evaporative cooling) to theturbine ACC option. For both options, the equipment is sizedto remove the full heat load over the full ambient temperaturerange. As can be seen, with an ACC, the power generation islower because an ACC runs at a higher temperature than asurface condenser.

Summary

The following cooling circuit water-saving options arecompared in this paper:

➤ Base case – evaporative cooling (no pretreatment)➤ Evaporative cooling (with pretreatment)➤ Dry cooling (no pretreatment) with the following

variants:– ACC on turbine exhaust– ACC on turbine exhaust, and fin-fans on

absorbing acid circuit.

These options are shown in the schematic in Figure 5. All options assume steam is generated in the sulphur

burner waste heat boiler and electricity generated in a steamturbine generator. Where dry cooling options are considered,the balance of cooling is provided by evaporative cooling.

Table III summarizes the make-up water consumptionand heat removal duties of the cooling circuit options.

Evaluation of options

The Hatch 4QA approach compares the economic andsustainability impacts of alternative project options.

Economic impact

Order–of-magnitude capital and operating cost estimates weredeveloped for each cooling system option and compared tothe base case (evaporative cooling with no pretreatment). Asummary of the comparison is presented in Table IV.

Sustainability impacts

Sulphur-burning acid plant emissions include:

➤ Gaseous emissions—sulphur dioxide, nitrogen oxides,and acid mist in tail gas

➤ Liquid effluents—waste heat boiler and cooling circuitblowdown, demineralized water treatment plant waste,plant washings, spillages and leakages

➤ Solid effluent—sulphur filter cake residues and spentconverter catalyst

➤ Noise pollution—main blower and turbine.

Sulphuric acid plant water saving options

The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 MAY 2014 359 ▲

Figure 4—Impact of ambient temperature on turbine electricitygeneration

Figure 5—Water saving options schematic

Table III

Summary of cooling circuit options

Parameters Units Evaporative cooling Dry cooling (no pretreatment)

No pretreatment Pretreatment included ACC ACC +Fin-fan

Make-up water consumptionMake-up water t/h 220 172 86 7

Heat removal duty (nominal)Evaporative cooling tower 106 kJ/h 396 396 162 18ACC 106 kJ/h 0 0 234 234Fin-fan 106 kJ/h 0 0 0 144Total (nominal) 106 kJ/h 396 396 396 396Total (design)* 106 kJ/h 468 468 468 468

*Design cooling duty includes an additional 72 x 106 kJ/h capacity for when the steam turbine is bypassed

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Sulphuric acid plant water saving options

Commitments by major corporations as well asgovernment regulatory requirements have resulted in thedevelopment of several new and cost-effective technologies toefficiently reduce gaseous emissions from an acid plant.Noise pollution has been addressed with suitable soundreduction measures such as acoustic insulation, enclosures,and silencers.

The treatment and disposal of liquid and solid effluents isfacing more stringent controls through public awareness andgovernment regulations. Reduction in water consumption ispotentially the greatest beneficial impact on localcommunities and the environment. Furthermore, a reductionin the cooling water blowdown will reduce the plant effluentand ultimately reduce the impact on the overall plant effluentcatchment area.

Water, footprint, power, and waste

Four sustainability criteria were identified to quantitativelycompare the different water saving options, namely: waterintensity, power intensity, waste intensity, and footprint. Acomparison of these criteria is given in Table V. Water andwaste intensity are calculated from the mass balance; powerand footprint values are estimated from recent projectexperience. The values shown are for the cooling system onlyand exclude the criteria associated with the remainder of theacid plant.

The following is noted with respect to each of theintensity factors:

➤ Water intensity– Pretreatment provides water savings when applied

to the base case– Dry cooling options provide the lowest overall

water consumption.➤ Power intensity

– Evaporative cooling includes power to operate thecooling tower fans and the cooling water supplypumps

– Base case with pretreatment has a slightly higherpower usage due to more pumping requiredbetween upfront unit operations

– Dry cooling includes power to operate the fansonly. Although electricity consumption is lowerfor dry cooling options, the power intensity shownin Table V is higher because it has been calculatedtaking into account a reduction in turbineelectricity generation of 1.5 MWe (see previously).

➤ Footprint intensity – The footprint of the base case with pretreatment is

comparable with the base case– Dry cooling options require more footprint for the

same cooling duty.➤ Waste intensity

– Evaporative cooling options generate moreblowdown and therefore increased waste foreffluent treatment

– There is a large reduction in waste generatedwhen pretreatment is included ahead of theevaporative cooling tower.

Four Quadrant Analysis

Hatch developed the Four Quadrant Analysis (4QA) approachto compare project options using the economic and sustain-ability impacts. The 4QA tool plots each option compared to abase case:

➤ The x-axis is a cost ratio, with a lower cost ratiorepresenting a lower cost option relative to the basecase

➤ The y-axis is a sustainability ratio, which compares theintensity of the option to the base case. A lowersustainability ratio is preferred, which indicates a lowerimpact on the environment.

The cost ratio (CR) is calculated for each option, relativeto the base case (BC), using the following equation:

360 MAY 2014 VOLUME 114 The Journal of The Southern African Institute of Mining and Metallurgy

Table V

Sustainability criteria values for cooling system options

Sustainability criteria values for Base case ACC ACC + Fin-fan

cooling system options No pretreatment With pretreatment No pretreatment No pretreatment

Water intensity m3/t acid 2.6 2.3 1.0 0.1Power intensity kWh/t acid 36 38 44 48Footprint intensity m2/t acid 10 12 18 23Waste intensity m3/t acid 0.9 0.5 0.3 0.03

Table IV

Cooling system capital and operating cost comparison

Cost parameters Base case ACC ACC + Fin-fan

No pretreatment With pretreatment No pretreatment No pretreatment

CAPEX (relative to base case) 1.00 1.21 1.33 1.43OPEX (relative to base case) 1.00 1.18 0.99 0.87

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[1]

The CR is the sum of the annual operating cost and theannual capital cost repayments, based on a nominal 5-yearrepayment (compounded monthly). Any credits received forselling electricity to the market have not been included intothe evaluation.

The sustainability ratio (SR) is calculated for each option,relative to the base case (BC), using the following equation:

[2]

The weightings can be adapted based on the generalimportance of each criterion. The weightings used in thisevaluation (Table VI) have placed a high importance on waterintensity as many plants strive to reduce water consumption.The relative weightings will be site-specific, for example aridlocations may consider water impacts more important andfootprint less important.

The relative cost ratio and sustainability ratio of thewater- saving options are given in Table VII, and the 4QAplot is in Figure 6.

The 4QA shows that for typical site locations:

➤ Base case with pretreatment has a better sustainabilityratio than the base case; however, the cost ratio willincrease by 19%. This is mainly due to the highoperating cost associated with reagent consumption forcationic resin regeneration. Alternative pretreatmentoptions can be investigated with lower reagent usageand allowing increased cycles of concentration in thecooling tower

➤ Dry cooling offers considerable improvements to thesustainability ratio, but the cost ratio will increase by10% for ACC and 4% for ACC and fin-fan.

Sensitivity analysis

The cost and sustainability ratios can be affected by severalfactors, some of which are briefly considered in the followingsensitivity analyses.

Water costs

Fresh water supply costs are location-dependent. The relativewater supply costs used for this sensitivity are based on:

➤ Low water cost (US$0.2 per m3)—typical for locallyavailable water source of good quality (e.g. dam locatedclose to plant) with no additional extraction charges

➤ Average water cost (US$1.0 per m3)—typical for watersources located a reasonable distance from the plant,requiring minor infrastructure to be built and someminor water treatment on site (e.g. sand filtration)

➤ High water cost (US$3.0 per m3)—typical for watersources located at a considerable distance or water ofpoor quality requiring significant treatment (e.g.reverse osmosis). High water cost would also apply forwater that is local and of good quality, but with a highextraction charge.

Table VIII summarizes the cost ratios for the water costsensitivity analyses. The sustainability ratios remainunchanged.

The sensitivity analysis shows that:

➤ Reduced water costs (US$0.2 per m3) increase the costratio of dry cooling options, making them unfavourablecompared to the base case

Sulphuric acid plant water saving options

The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 MAY 2014 361 ▲

Table VII

Sustainability and cost ratio summary

Ratio Base case ACC ACC + fin-fan

No pretreatment With pretreatment No pretreatment No pretreatment

Cost ratio 1.00 1.19 1.10 1.04Sustainability ratio 1.00 0.87 0.77 0.67

Table VI

Weightings of sustainability criteria

Sustainability criterion Weighting

Water intensity 40%Power intensity 15%Footprint intensity 20%Waste intensity 25%

Figure 6—Hatch 4Q Analyses (average fresh water and power cost)

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Sulphuric acid plant water saving options

➤ Increased water costs (US$3 per m3) reduce the costratio of dry cooling options to less than the base case(by as much as 23%)

➤ This supports the general observation that as waterextraction costs increase, dry cooling options arepreferred

➤ The base case with pretreatment has the highest overallcost ratio, which corroborates the data in Table VII. At reduced water costs (US$0.2 per m3) thebase case with pretreatment compares well with the drycooling options, but becomes the least favourableoverall at increased water costs (US$3 per m3).

The 4QA was updated with the low and high water costsin Figure 7.

Electricity costs

Electricity supply costs are also location-dependent. Therelative electricity supply costs used for this sensitivity arebased on:

➤ Low electricity cost (US$0.05 per kWh): for locationswith abundant low cost electricity, e.g. hydroelectricity

➤ Average electricity cost (US$0.1 per kWh): for locationswith a typical mixed electricity supply, e.g. a mix ofcoal, renewable, and gas-fired power stations

➤ High electricity cost (US$0.3 per kWh): for locationswhere electricity is generated on site, e.g. local diesel orgas-fired generators.

Table IX summarizes the cost ratios for the electricity costsensitivity analyses.

The sensitivity analysis shows that:

➤ Lower electricity costs have a minor impact on the 4QAplot

➤ At increased electricity cost, the advantage ofevaporative cooling is clear, due to the increasedturbine electricity output

➤ The base case with pretreatment compares well with

the dry cooling options at increased electricity costs;however, it becomes the least favourable at the lowerelectricity costs.

The 4QA was updated with the low and high electricitycosts in Figure 8.

Sustainability criteria weightings

The sustainability criteria weightings can be adjusted to suitthe plant location and requirement for generating electricity,thereby impacting the sustainability ratio. As an example, theweightings can be adjusted as water or power becomes moreor less important to the local community. Furthermore,additional sustainability criteria can be included, such as:

➤ Specific reagent consumption (e.g. high RO membranecosts)

➤ Downstream impact and stewardship (qualitative)➤ Operability and maintainability (qualitative)

362 MAY 2014 VOLUME 114 The Journal of The Southern African Institute of Mining and Metallurgy

Figure 7—Hatch 4Q Analyses for water cost sensitivity

Table IX

Cost ratios for electricity cost sensitivity

Cost sensitivity Base case ACC ACC + fin-fan

No pretreatment With pretreatment No pretreatment No pretreatment

Low electricity cost ratio US$0.05 per kWh 1.00 1.21 1.08 1.00Average electricity cost ratio US$0.1 per kWh 1.00 1.19 1.10 1.04High electricity cost ratio US$0.3 per kWh 1.00 1.13 1.13 1.14

Table VIII

Cost ratios for water cost sensitivity

Cost sensitivity Base case ACC ACC + fin-fan

No pretreatment With pretreatment No pretreatment No pretreatment

Low fresh water cost ratio US$0.2 per m3 1.00 1.24 1.22 1.22Average fresh water cost ratio US$1 per m3 1.00 1.19 1.10 1.04High fresh water cost ratio US$3 per m3 1.00 1.10 0.91 0.77

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➤ Social, for example using local labour (qualitative)➤ Government and externals (qualitative)➤ Emissions (qualitative).

The 4QA approach is flexible and can be customized andadapted to meet specific project criteria.

Conclusions

The Hatch 4QA approach compares the economic andsustainability impacts of alternative project options. It can beused for technology and site selection from concept throughto feasibility studies and beyond. The 4QA additionallyserves as a risk management tool to quantify the impacts ofvarying sustainability criteria and input costs.

For the water-savings options considered in this paper, itis the acid plant location that largely determines the sustain-ability and cost ratios. Key findings include:

➤ The cost ratio of evaporative cooling is generally lower,provided there is good-quality and low- to medium-cost water available. The sustainability ratio isgenerally higher due to the high water consumption,which can make evaporative cooling unfavourable evenat sites with low water costs (Maulbetsch, DiFilippo,and Zammit, 2008)

➤ The cost ratio of the base case with pretreatment is thehighest overall due to the high reagent usage.Optimizing the make-up water chemistry by adjustingthe pH and adding scale inhibitors might be a moreefficient way of increasing the cycles of concentration,but needs to be investigated on a case-by-case basis

➤ Reverse osmosis (RO) can also offer water savings, butthese could be offset by the RO waste generation

➤ The sustainability ratio of the base case withpretreatment is lower than without pretreatment due tothe decreased water usage

➤ Dry cooling options have a higher capital outlay, butcan have lower operating costs in locations wherewater extraction costs are high. The sustainability ratio

is generally lower for dry cooling, due to lower waterconsumption and similar power consumption toevaporative cooling

➤ Acid plant electricity generation capacity is lower withdry cooling options and is worsened during higherambient temperature conditions. For large acid plantsgenerating electricity that is sold to market, this willadversely impact plant revenue.

Acknowledgements

The authors wish to acknowledge the contributions of DrMatthew King and Rusi Kapadia, and Hatch for permission topublish this paper.

References

ENDUROSOLV. 2012. Cooling Tower Water Saving Strategies,

http://endurosolv.com/pdf/cooling_tower_savings_strategies.pdf

[Accessed 25 Apr. 2014].

EPRI. 2002. Comparison of Alternate Cooling Technologies for California Power

Plants: Economic, Environmental and Other Trade-offs. Electric Power

Research Institute, Aplo Alto, CA, and Californian Energy Commission,

Sacramento, CA.

ESKOM. 2010. Factsheet, General Communication CO 0005, Revision 7, October

2010.

http://www.eskom.co.za/AboutElectricity/FactsFigures/Documents/CO000

5CoolingTechniquesRev10.pdf [Accessed 12 June 2013].

KING, M.J., DAVENPORT, W.G., and MOATS, M.S. 2013. Sulphuric Acid

Manufacture: Analysis, Control and Optimization. 2nd edn. Elsevier,

Burlington, MA. ISBN: 978-0-08-098220-5.

MAULBETSCH, J.S., DIFILIPPO, M.N., and ZAMMIT, K.D. 2008. spray cooling for

performance enhancement of air-cooled condensers. EPRI Advanced

Cooling Workshop, Carolina.

MORTENSEN, K. 2011. Improved Performance of an Air Cooled Condenser (ACC)

Using SPX Wind Guide Technology at Coal-Based Thermoelectric Power

Plants. US Department of Energy.

http://www.netl.doe.gov/File%20Library/Research/Coal/ewr/water/Proj51

9.pdf [Accessed 24 April 2014].

SPX. 2012. Air Cooled Condensers. www.spx.com/en/assets/pdf/A4_ACC-

12.pdf [Accessed 18 June 2013].

US DEPARTMENT OF ENERGY. 2009. Concentrating Solar Power Commercial

Application Study: Reducing Water Consumption of Concentrating Solar

Power Electricity Generation. Report to Congress in response to Energy

Independence and Security Act of 2007. (Pub. L. No. 110-140),

http://www.nrel.gov/csp/pdfs/csp_water_study.pdf [Accessed 25 April

2014].

WILSON, B. 2011. Detail Engineering and Layout of Piping Systems. 1st edn. On

Demand Books, New York. ISBN : 9781926633183. ◆

Sulphuric acid plant water saving options

The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 MAY 2014 363 ▲

Figure 8—Hatch 4Q Analyses for electricity cost sensitivity

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Introduction

Nkomati Nickel JV has experienced aphenomenal growth rate over the last fewyears, from a 10 kt/month operation in 2006to a 700 kt/month complex in 2013. Thisgrowth required the re-engineering of virtuallyevery aspect of the operation, from miningnew ore types with new methods, orepreparation and processing, to tailingsdeposition.

Many alternative processing methods wereconsidered before the current flow sheets wereadopted. Production of separate nickel andcopper concentrates through successiveselective flotation, Activox® leaching ofconcentrates, and local smelting of concen-trates were assessed among other options.Ultimately, economic factors resulted in thecurrent circuit choices. This expansionoccurred against the backdrop of the ongoing

global economic crisis, with persistent lowmetal prices.

This paper discusses many of the effortsthat contributed the turnaround of NkomatiNickel. Emphasis is placed on the challengesand successes at the MMZ plant, though thePCMZ plant showed similar improvements.

The Uitkomst deposit

Nkomati Nickel JV exploits the Uitkomstdeposit in South Africa’s Mpumalangaprovince, in the mountains between WatervalBoven, Machadodorp, and Badplaas (Figure 1).

The orebody is an early-age (2044 Ma)Bushveld layered lenticular mafic-ultramaficintrusion into the basal sediments of theTransvaal Supergroup, approximately 9 kmlong and 1500 m wide (Figure 2). The depositdips north-east at about 4 degrees. The depositwas exploited by AngloVaal with variouspartners since the early 1990s. Nkomati NickelJV is a 50/50 partnership between AfricanRainbow Minerals and Norilsk Nickel Africa.

The orebody has multiple zones ofsulphide mineralization:

➤ MSB: Massive Sulphide Body with Nigrades in excess of 2%. Mined since1997, now mined out

➤ MMZ: Main Mineralized Zone. Headgrades 0.3–0.7% Ni, approximately 0.37% Ni average

➤ PCMZ: Peridotitic, ChromititicMineralized Zone. Chrome-rich ore withgrades of 0.2–1% Ni, 0.23% Ni average,chrome grades of 10-15% Cr2O3.

➤ Massive chromitite (often called PCR):stockpiled for a separate chromewashing plant, currently mothballed

➤ Basal Mineralized Zone: unexploited atpresent.

Challenges and successes at the NkomatiNickel JV: pit-to product processimprovementsby G. Cockburn*

SynopsisNkomati Nickel JV exploits the ores of the Uitkomst Complex nearMachadodorp in the Waterval Boven district in South Africa’s MpumalangaProvince.

Due to factors such as the remote location, stellar growth inproduction, opencast mining methods, and ore characteristics, a number ofinnovative processing options were selected.

Nkomati has undertaken numerous initiatives over the last few yearsto improve plant running times, metallurgical performance, and operationalprofitability. Great emphasis has been placed on effectiveness ofmanagement control systems. A number of initiatives such as shortinterval control and time–in-state metrics have been implemented. A focuson improvement on availability and asset utilization of key items ofequipment has been particularly effective.

While the ores are remarkably similar to the Merensky and UG2 reefs,the relatively high base metal sulphide content and mineralogical charac-teristics make metallurgical treatment somewhat different to the ores of theBushveld Complex. The low head grades and flotation kinetics distinguishNkomati from other base metal operations. Numerous milling and flotationoptimization initiatives have resulted in dramatic improvements inthroughput, recoveries, and concentrate grades.

This paper discusses the metallurgical, operational, and managementchallenges and the outcomes obtained.

KeywordsUitkomst, Nkomati Nickel, sulphide mineralization, grade control, orefragmentation, problem solving methodologies.

* Nkomati Nickel JV.© The Southern African Institute of Mining and

Metallurgy, 2014. ISSN 2225-6253. This paperwas first presented at the, Base Metals Conference2013, 2–4 September 2013, IngwenyamaConference & Sports Resort, Mpumalanga.

365The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 MAY 2014 ▲

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Challenges and successes at the Nkomati Nickel JV: pit to product process improvements

Ore production

Currently only the MMZ and PCMZ ores are mined. Oreproduction from the open pit is approximately 650 kt/month,of which approximately 300 kt is PCMZ and 350 kt MMZ.The MMZ is also mined in the underground mining section,

producing approximately 50 kt per month by bord and pillarand longhole open stoping methods.

It must be noted that the production profile (Figure 3) isroutinely optimized and updated as models are tuned andimproved based on the outcomes of the RC drillingprogramme discussed below.

366 MAY 2014 VOLUME 114 The Journal of The Southern African Institute of Mining and Metallurgy

Figure 1—Location of Nkomati Nickel JV

Figure 2—Idealized cross-section of Uitkomst deposit

Figure 3—Life-of-mine ore production profile

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Mineralogy

The pyroxenites and peridotites of the Lower Pyroxenite thathosts the MMZ consist mainly of clinopyroxene, olivine, andplagioclase. Hydrothermal action has resulted in extensivealteration of these minerals to amphibole, serpentine, biotite,and talc. Contamination of the ultramafic suite by countryrocks accounts for the most of the calcite, dolomite, quartz,and plagioclase. Talc content is highly variable within thedeposit and irregular (Brits, 2008)

Highly altered talc-rich zones are often associated withpyrite-rich zones, and the high flotation kinetics of boththese minerals complicates the flotation process, resulting in lower pentlandite recoveries, dilution of the concentratewith pyrite, and reduced concentrate quality due to higherMgO levels.

Nickel is mainly contained within pentlandite, although asignificant proportion (as much as 15%) occurs in solidsolution within pyrrhotite and 1–2% within chlorite. Copperoccurs almost exclusively as chalcopyrite, with someoccurring as bornite (1–2%).

The MMZ in many ways resembles Merensky Reef,although it contains substantially lower platinum groupmetals (PGMs) and higher base metal sulphides (typically5–8%) with traces of PGMs (1 g/t head grade, predominantlyMerenskyite). The PCMZ resembles the UG2 Reef, withchrome grades of 7–15% Cr2O3. From the geologist’sperspective, the ores are effectively the same with theexception of the chrome grades. The boundaries of the twoore types are not clearly delineated, making segregation ofore and prevention of cross-contamination challenging.

From a processing perspective, however, the ores aresignificantly different. The target liberation grind for MMZore is 67% -75 μm, although recoveries are relativelyinsensitive to grind. PCMZ ore is extremely sensitive to grind,with a target grind of 80% -75 μm, and drastic losses inrecovery occur at lower grind values. Misplaced ore thusdirectly affects plant performance.

Grade control

An extensive reverse circulation (RC) drilling programme atan initial hole spacing of 25 m × 25 m, and subsequently at12 m × 12 m, has greatly enhanced the ability to model theorebody and so allow far better head grade control.

This is critical, considering the variability of gradeswithin the orebody, and the fact that a substantial amount ofPCMZ ore in particular is below economically viable grade.Management of the resource is thus a vital aspect ofmaximizing the value of the mine. RC drilling data and theresource models derived from it are extensively used in mine-to-mill reconciliations as well.

Processing challenges

Primary gyratory crusher

With the open pit supplying the vast bulk of the ore, aprimary gyratory crusher at the pit was selected withoverland conveyors to transport to the two plants. Loadingand crushing are alternated between the two ore types, withcrushed ore transported by conveyors approximately 3 km tothe respective conical stockpiles located at the plants.

The Metso 54 × 75 Mk2 gyratory crusher was initiallyviewed as something of an Achilles’ heel of the operation.The crusher suffered numerous breakdowns and trips andbecame the major process bottleneck. Although designedwith an F80 of 450 mm and an F 100 of 1000 mm and a feedrate of 1600–1800 t/h, rocks substantially larger than designwere routinely crushed, resulting in trips and mechanicalfailures. Tramp steel was also a major contributor todowntime.

Great focus was placed on preventing large rocks fromentering the crusher. A ‘SPLIT’ camera and image analysissystems were introduced to monitor and record the size ofrocks on trucks prior to tipping, with an additional systemmonitoring rock size during tipping. These systems provide avital service in monitoring crusher feed PSD (Figure 4). Largerocks are prevented from entering the crusher largely throughvisual observation by control room operators, who rejecttruckloads with large rocks.

Interestingly, analysis of data indicated that high-amperage trips (and damage) on the crusher were not causedby large rocks alone. Correlation of SPLIT rock size images,vibration, and ampere readings indicated that smallerfootball-sized rocks, mostly from re-broken boulders fed tothe crusher in the absence of fines, were as much of achallenge as very large rocks. This was exacerbated by wearon the liners, where the crusher cavity would wear to a‘hockey stick’ shape, trapping critical-sized rocks, akin tobearings in a race.

Key to improving throughput and availability was theimplementation of strict planned maintenance systems. Keyperformance indicators such as overall equipment efficiency(OEE) were introduced and proved very effective inmonitoring the actual crusher performance. OEE is calculatedfrom actual tons processed divided by the theoreticalmaximum the equipment can process over the period ofconsideration. This measure cuts out all the clutter andconfusion of allocation of downtime, and provides a ‘bottomline’ performance value.

The overland conveyor system capacity was increased toaccommodate the increase in crushed tonnages from 1300 to2000 t/h. This necessitated the installation of larger headpulleys, shallower troughing angles on belts, and faster beltspeeds to reduce persistent belt splice failures. Attention tobest practices in splicing the steel-cored belts was vital toincrease availability.

Challenges and successes at the Nkomati Nickel JV: pit to product process improvements

367The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 MAY 2014 ▲

Figure 4—PSD analysis of crusher feed using the SPLIT camera system

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Challenges and successes at the Nkomati Nickel JV: pit to product process improvements

Figure 5 clearly indicates the step change in total monthlymilled tons for the complex, subsequent to the resolution ofthroughput constrains around the primary crusher inJuly/August 2012.

Ore fragmentation

Fragmentation pattern improvements within the pit wererealized through redesigned drilling patterns, and changes tothe blast timing and blast direction. Interestingly, improvedfragmentation was achieved at an increased hole spacing andmuch reduced powder factor. Production hole spacing wasincreased from 3.0 m × 3.5 m to 3.5 m × 3.5 m, whilemaintaining the 10.7 m hole depth and 3 m stemmingmaterial depth. Powder factor was reduced from 1.9 kg/m2

to 1.4 kg/m2. Analysis of blasted material indicated that the majority of

large rocks originated from the collar, close to the surface.The introduction of 3 m stab holes (with a relatively lightcharge) between production holes resolved this problem.

Fragmentation was further improved through changingthe direction of the blast from north to south (up-dip) to westto east (cross-dip) and introducing substantially slower blasttiming. The introduction of hole depth counters to ensure

consistent hole depths as well as rigorous quality assuranceinspections contributed to consistency in fragmentation aswell as increased production.

MMZ comminution circuit

The MMZ circuit design (Figure 6) employs a FAG primarymill in closed circuit with a vibrating screen and pebblecrushers. A secondary ball mill in closed circuit with cyclonessupplies feed to the float circuit at SG 1.3–1.34, 70% -75 μmat 620 t/h. Design considerations and alternativecomminution options considered were discussed byWolmarans and Morgan (2009).

The choice of this circuit caused extensive debate, as theFAG mill was viewed by some as a ‘stone washer’, that wouldresult in excessive metal losses through the sliming of softernickel minerals. The proponents advocated that the reducedoperating cost far outweighed potential recovery losses.

This circuit is very sensitive to feed particle size distri-bution as well as ore hardness. Excessive fine material (andcoarse material) results in overloading of the mill.Maintaining a full stockpile of 20–30 kt live capacitycontributes greatly to mill stability through consistent feed.

368 MAY 2014 VOLUME 114 The Journal of The Southern African Institute of Mining and Metallurgy

Figure 5—Metallurgical complex milled throughput

Figure 6—MMZ milling circuit flow sheet with sampling and monitoring points

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Various regimes of feeder operation have been attemptedto supply consistent PSD material to the mill, with operationof one inner and one outer feeder generally used. On-lineimage analysis of mill feed to determine PSD, with automatedfeeder selection, has been less successful. Segregationpatterns vary as the stockpile is loaded or depleted, andprediction of individual feeder PSD has not yet beensuccessful.

MMZ FAG circuit pebble crushers

Optimization of the MMZ milling circuit highlighted thecritical importance of the gap setting of the pebble crushers.While the process design criteria specified a closed sidesetting of 10–13 mm, realistically 19–21 mm was the bestthat could be achieved without causing frequent mechanicalfailures to the crushers.

Bypassing the pebble crushers or running with excessivegap setting would increase milling power from 24 to 30 kwh/t, with the mill feed rate cut back from 600 t/h toapproximately 450 t/h due to overloading of the primary millwith pebbles. In addition, recoveries are affected by as muchas 5–10%. While crusher comminution energy contributesjust 400 kW to an average total of 16 MW for the FAG/ballmill/pebble crusher circuit (less than 2.5%), the pebblecrusher is vital in removing critical-sized material from thecircuit.

Extensive re-engineering of the crusher bushes andrigorous attention to planned maintenance, which wascontracted to the OEM, allowed the crusher gap to be reducedgradually to 13 mm. The correct running in of the liners overa 4-week period was vital to prevent metal-on-metal contactand bush damage. Redesigned liners that do not need theextended running-in period are under development and areexpected shortly.

A direct result of the higher-than-design crusher gapsetting was an increased pebble crusher throughput, whichexceeded the pebble production rate from the FAG mill. Thisresulted in stop/start operation of the crushers that threwripples through the primary mill, mill discharge sump, andthe cyclone circuits. Cyclical changes in froth stability wereregularly observed in flotation, and were attributed toshifting grind as the milling circuit flows changed.

Data analysis indicated that stop/start operation of thecrusher increases milling circuit power requirements byapproximately 1 kWh/t, or R2.5 million per year at currentpower costs. Lost recovery costs are substantially higher(Van der Merwe, 2013).

Solving the crusher circuit instability was thus essential.Various initiatives were tested to balance crusher rate feedwith capacity. Although manipulating the screen panel sizeon the primary mill discharge was effective, knock-on effectson secondary mill performance resulted in less than optimumsecondary mill performance.

A novel approach was to increase the crusher speed byapproximately 15%. While it may sound counter-intuitive toincrease a crusher speed to reduce capacity, this encourageschoke conditions and reduced throughput, although the exactchange in capacity has yet to be quantified. Furtherrefinements have been to re-set pebble bin low and highlevels to encourage more frequent (but shorter) stops and

starts. Downstream surge bins and other options arecurrently under consideration.

Milling circuit expert system tuning

Tuning of the PxP mill expert control system and othercontrol loops has assisted in improving stability. Assistancefrom the OEM, FL Smidth, was essential in understandingand optimizing the mill circuit control system in particular.An important change in philosophy was the fixing of thenumber of operating cyclones and varying sump wateraddition, rather than allowing the control system to open andclose cyclones to maintain constant pressure. The effects ofvaried density of the cyclone feed appears to be lessdisruptive than opening and closing cyclones.

MMZ comminution circuit operating costsWhile detailed modelling of the milling circuit is yet to bedone, indications are that milling efficiency on the FAG/ballmill circuit is close to what would be expected in a crushingand ball milling circuit. Comminution circuit operating costsare lower on the MMZ plant than the PCMZ plant (whichutilizes conventional crushing and two stages of ball milling)by approximately R10 per ton. MMZ plant recoveries areclose to or in excess of design figures. The selection of thiscircuit design over others considered by DRA appears to havebeen justified.

MMZ flotation circuitThe MMZ flotation circuit (Figure 7) is a relatively standardrougher-cleaner-recleaner configuration. A combined nickel-copper-cobalt and PGM concentrate is collected from therecleaner. Currently the pyrrhotite scavengers function asextended rougher capacity.

One of the immense challenges on the flotation circuit hasbeen to achieve operational stability, and the resolution of anumber of issues has contributed to the current relativelysmooth operation.

Flotation cell level controlAnalysis of the cumulative level control valve outputs of thecells in the two parallel rougher banks indicated that the onebank received more flow than the other. In addition, the celldischarge valves would saturate at 100% output moreregularly on the one bank, resulting in excessive rougherconcentrate volumetric flow to the cleaner circuit. Theunequal flow split was attributed to an inherent flaw in thedesign of the splitter between the two banks.

Excellent results were obtained with the implementationof the Gipronickel mass pull control algorithm. This relativelysimple loop adjusts air to the flotation bank based onvolumetric flow of concentrate. Although not as advanced assome concentrate mass pull control models, this system hasbeen very effective in breaking the cyclical swings in recircu-lating load that characterized the circuit.

Flotation cell level control issues were exacerbated by thepiping arrangement of the tailings from the rougher andscavenger banks. The gooseneck discharge pipes collect gritunable to be carried over with the tailings. Insufficient headon the gravity flow arrangement to the tailings thickenerrestricts further easy solutions, and alternative solutions arebeing considered.

Challenges and successes at the Nkomati Nickel JV: pit to product process improvements

The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 MAY 2014 369 ▲

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Challenges and successes at the Nkomati Nickel JV: pit to product process improvements

Reagent optimization

Talc and serpentine are the main contributors to high MgOlevels in the concentrate, which incur severe smeltertreatment penalties. Both these minerals have fast flotationkinetics and a strong froth stabilizing action. Under-depression results in vastly increased mass pulls, increasedcirculating loads, and low concentrate grades while alsoresulting in recovery losses.

Extensive tuning of the flotation circuit reagents hasimproved concentrate grades dramatically. Depressantallocation within the circuit has largely been shifted to thecleaners and re-cleaners. Careful control of collector haslimited the over-collection of gangue.

The main value minerals recovered from the MMZ ore arepentlandite and chalcopyrite, while substantial (and variable)amounts of pyrite and pyrrhotite occur. Pyrite andchalcopyrite have substantially higher flotation kinetics thanpentlandite, and mass pull control during periods of abnormalore conditions is critical in maintaining recoveries.

Operator training

It is an easy mistake for inexperienced operators to produce,for example, a high-grade pyrite concentrate at the expenseof nickel recovery. While many operators had extensiveexposure to PGM flotation, fewer had base metal flotationexperience. Training of operators, control room operators,and metallurgists has paid off very well in terms of reducingand largely eliminating poor plant performance due tomisdiagnosis or incorrect response to operational upsetconditions. In combination with the ‘process recipes’discussed below, operation of the MMZ plant has vastlyimproved.

Plant operational management

Process recipes

Of all the improvements implemented at the MMZ plant, theintroduction of ‘process recipes’ has to rank among the mostsuccessful in ensuring consistent operation.

Process recipes stipulate the ranges within which keyprocess variable must be run. These include froth depths, air

addition rates, mill power draw, cyclone feed pressures, andSGs and other metallurgical variables. These have addedimmensely to the stability of the operation of the plant, andhave prevented operator-induced instability due to incorrector excessive corrections of perceived process upsets. Processrecipes are adjusted infrequently, and only in consultationwith senior management, metallurgists, and production staffand are signed off at senior level.

‘Short-interval control’ (SIC) was introduced as a way ofguiding operators into checking the key process variables toassess the plant performance on a regular routine basis. Thisconsists of a set of focused log sheets that required routinechecks and monitoring, as well as monitoring of 2-hourlyquality control assay results. These consist of a set ofcalculations attached to the 2-hourly quality control assayreporting sheets. Flotation circuit recoveries, upgrade ratiosacross flotation banks, ratios between Fe/MgO, Ni, and Fe,and other indicators of concentrate quality or ore qualitywarn operators of current of looming plant performanceissues.

SIC is being extended with the introduction of time-in-state (TIS) monitoring. TIS presents the operator with a ‘dial’dedicated to each section of the plant (Figure 8). Dataanalysis has been used to assess the key parameters thatinfluence the performance of a specific section, as well at thewhole plant. An algorithm monitors multiple contributingfactors, such as froth depths, aeration rate, concentrate sumplevels, and flotation cell feed and product grades

The operator is presented with and easy–to-read dial,with bar chart indicating what factor is out of range, andcomments as to what ‘lever’ to pull to correct the situation.Essential to the deployment of the TIS system has been thedevelopment and implementation of valid models thatunderlie the ‘idea state’ index.

Performance mapping

Extensive analysis of data using ‘performance mapping’ isused. Historical data pertaining to key variables consideredby plant operational and metallurgical staff to be the maindrivers and indicators of ideal operating performance isanalysed, and the correlations are plotted in 2-dimensionalspace.

370 MAY 2014 VOLUME 114 The Journal of The Southern African Institute of Mining and Metallurgy

Figure 7—MMZ flotation circuit flow diagram

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These performance maps have been extremely useful inindicating the correlations between various operationalvariables and process performance. They also indicate howoften the process is in various ‘states’, which are often notthe ‘ideal state’. The technique has been very useful inindicating appropriate process recipe set-points.

It is important to note that this method is used inconjunction with the plotting of trends and conventionalmetallurgical evaluation methods.

Graphics such as those shown in Figure 9 were generatedfor all major variables deemed to potentially impact plantperformance, or to be indicators of conditions that wouldaffect plant performance. The example shown correlates therougher bank flotation performance with key parameterssuch as mill power, as well as other indicators such as therougher motor power draw. While rougher cell power draw isnot considered to be a key variable, it is an indicator of cellaeration, flotation feed densities, or wear on the flotation cellmechanism that requires scrutiny. These indicators are usedtogether with the TIS dials to advise operators on whatactions should be taken to rectify sub-optimum operatingconditions.

Fluid Reports

Fluid Reports is an automated reporting system developed byBlue Nickel in conjunction with U-Drive. This system drawsdata from the SCADA’s historian, and generates automatedreports and trends on the performance of key processvariables, including feed tons, key flow rates, powerconsumption, or similar process variables that directly affectthroughput, recoveries, or product quality (Figure 10).

While not containing any information not alreadyavailable on the SCADA, the system has proved extremelyuseful in producing easy-to-read trends tailored to anaudience who do not have ready access to SCADA viewers.An added benefit is the interpretation of trends relating toprocess control issues, inserted into the trends as ‘stickynotes’, and detailed in a separate report. Many of themetallurgical and control issues resolved were identified inFluid Reports and resolved with the assistance of the BlueNickel team.

A similar user-configurable web-based application allowsreal-time trends to be viewed from any computer orsmartphone. These features are aimed at providing better

Challenges and successes at the Nkomati Nickel JV: pit to product process improvements

The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 MAY 2014 371 ▲Figure 8—Time-in-state performance indication dials

Figure 9—Performance maps

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Challenges and successes at the Nkomati Nickel JV: pit to product process improvements

support to operational staff outside office hours without theneed for engineering, instrument, or metallurgical staff to becalled out in the case of upset conditions.

Management systems

Lastly, the benefits of strong focused leadership have to berecognized. Recent management changes, with fresh ideasand methods, have had a very positive influence on produc-tivity.

Some of the most effective tools implemented were the‘Gap’ list and ‘5 Why’ problem-solving methodologies. Thesecontributed to a halving of the monthly mill trip rate in onemonth, from approximately 60 to less than 30 trips. Currentefforts are aimed at dropping this to below 10 trips permonth.

The importance of housekeeping on morale and disciplinecannot be understated. The mills are arguably the largest,most expensive, and most visible items on a plant.

The condition and maintenance standards on the mills(Figure 11) provide immediate visual reference on thehousekeeping and maintenance standards set for the rest ofthe operation.

Plant performance improvements

Numerous efforts were conducted simultaneously to resolvethe metallurgical, operational, and management issuesdiscussed above. Isolation of the contribution of individualprocess changes is thus difficult.

The impact of the introduction of new managementtechniques can probably best be seen in the reduction in milltrips. A step change in mill trips coincided with newmanagement appointments in the middle of 2012 (Figure 12).

It can be seen that mill stoppages due to instrumentationissues persisted throughout the period ending June 2013. It

must highlighted that this is not due to a failing of themanagement systems implemented, but rather the time takento resolve systemic instrumentation issues.

Reduced mill breakdowns as well as the resolution ofmaintenance and process throughput issues at the primarycrusher can largely be credited with the improved MMZ plantthroughput, as can be seen in Figure 13. Step changeimprovements in milled tons can be seen from mid-July 2012.

Poor mill throughput in July 2012 was due to downtimeattributed to pre-existing damage to FAG mill bearings andlube system. Throughput on the MMZ plant has exceedednameplate tonnages regularly since, and is expected to do somore consistently on the conclusion of current improvementprojects.

372 MAY 2014 VOLUME 114 The Journal of The Southern African Institute of Mining and Metallurgy

Figure 11—Mills and lube rooms

Figure 10—Fluid Reports trend reporting

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The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 MAY 2014 373 ▲

Figure 13—MMZ plant thoughput

Figure 12—MMZ mill breakdown analysis

The metallurgical impact of the efforts to improve millingcircuit stability, as well resolve the issues concerning thepebble crushers, are difficult to isolate from the impact ofefforts to optimize the flotation circuit control.

While step change improvements in throughput areapparent the case of crusher and milled tons, gradualimprovements in recovery are apparent. Figure 14 showsmonthly nickel recovery figures.

Conclusions

Despite a number of challenges, the Nkomati Nickel JV hasseen very strong growth in production over the last year.

A number of major equipment reliability and throughputissues at the mills and primary crusher were resolved,allowing for increased concentrate production. This has beenachieved by optimization of the existing equipment, withoutmajor capital investment.

The successful introduction of problem-solving method-ologies such Gap lists and 5 Whys played a significant role inidentifying and addressing equipment and operational issues.

Close cooperation with OEMs was crucial in resolving theissues around both the Metso primary crusher and FLSpebble crushers.

Improved fragmentation was achieved at lower powderfactors, allowing improved primary crusher capacity and lessdowntime on the primary crusher.

Improved mill throughput and more stable milling circuitoperation were achieved, with increased milling energyefficiency, through resolution of maintenance and operationalissues around the pebble crushers. The importance ofachieving design operational set-points on key equipmentsuch as pebble crushers in a FAG circuit is clear. Theoperating cost of the FAG mill is approximately R10 per toncheaper than the conventional crushing and ball millingcircuit of the PCMZ plant.

Numerous flotation circuit optimization projects,including the implementation of process recipes, Gipronickelmass pull control, IME’s time-in-state monitoring, andperformance mapping, as well as the use of Blue Nickel’sFluid Reports, have collectively contributed to recoveryimprovements of 13.1% year–on-year (62.3% in 2011/2012to 75.4% in 2012/2013).

Production trends indicate that the improvements aresustainable and that the figures for the year ahead shouldsurpass the previous year’s performance.

This effort cannot be attributed to a single department,but is rather the culmination of the combined efforts of

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Challenges and successes at the Nkomati Nickel JV: pit to product process improvements

mining, engineering, production, and metallurgy. The inputof various consultants, including the development ofanalytical and information systems tools, has beeninvaluable, as has been that of the engineering consultantsand head office advisors.

ReferencesBRITZ, J. 2008. High level review on the mineralogical variability of the Main

Mineralized Zone at Nkomati Mine. Internal memorandum, Norilsk NickelAfrica, 18 July 2008.

VAN DER MERWE, K. 2012. Primary Crusher Report dd12Jul2012. Report, IME, 12July 2012.

VAN DER MERWE, K. 2013. Pebble Crusher Performance March 2013. Report,IME, 13 March 2013.

WOLMARANS, E. AND MORGAN, P. 2009. Milling circuit selection for Nkomati 375ktpm concentrator. 5th Southern African Base Metals Conference, Kasane,Botswana, 27–31 July 2009. Southern African Institute of Mining andMetallurgy, Johannesburg. pp. 269–290. ◆

374 MAY 2014 VOLUME 114 The Journal of The Southern African Institute of Mining and Metallurgy

Figure 14—MMZ plant nickel recovery

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Introduction

The economic recovery of nickel from lateriteores has been pronounced for some years andwill become more critical in the future, aslateritic ores constitute most of the world’sknown nickel and cobalt resources, with nickelproduction from sulphide deposits progres-sively decreasing. High pressure acid leaching(HPAL) is being used for the recovery of nickelfrom nickel laterite ores, and increasinglyatmospheric leaching is also being considered.The quantity of laterite resources amenable tohydrometallurgical processing (limonite,nontronite/smectite) is almost twice thatamenable to pyrometallurgical processing(saprolite, garnierite) (Bacon and Mihaylov,2002).

Various hydrometallurgical flow sheets arebeing used for the recovery of nickel fromlaterite ores. Most plants use the Caron orHPAL processes, for which simple block flowdiagrams are given in Figure 1. The Caronprocess is considered primarily for limoniticores to avoid the high acid consumptionassociated with the iron content in the ore. Theore is calcined reductively to reduce the ferric(associated with goethite), prior to ammoniacal

leaching of the nickel and cobalt. YabuluNickel Refinery, Queensland, Australiaimplemented the Caron process.

The HPAL process can be used forlimonitic as well as saprolitic ores (<4% Mg)and has been installed on numerous plants,including Moa Bay, Bulong, Murrin Murrin,and Goro (Bacon and Mihaylov, 2002).However, the actual recovery of nickel andcobalt and their separation primarily fromcalcium, magnesium, and manganese is doneemploying different flow sheets. In the MurrinMurrin flow sheet the pH of the HPALpregnant solution is adjusted to pH 3.5–4 toneutralize excess acid and precipitate mostferric, aluminium, and chrome. This isfollowed by sulphide precipitation of nickeland cobalt, which is the primary technologyemployed to separate the nickel and cobaltfrom manganese, magnesium, and calcium.

The Bulong flow sheet (Figure 2) alsoneutralized the free acid and precipitated ferric,aluminium and chromium, but it employeddirect solvent extraction (SX) for the recoveryof nickel and cobalt (Flett, 2005). Cyanex 272(2,4,4-trimethylpentyl phosphinic acid) wasused to recover cobalt from the neutralizedstream, followed by nickel SX using Versatic10 acid (V10, a tertiary-branched carboxylicacid). One of the major drawbacks of direct SXas operated at Bulong was that the selectivityof the V10 extractant was inadequate toprevent calcium loading during extraction.This resulted in gypsum precipitation in the SXcircuit, and hence major operationaldifficulties.

The most recent major nickel lateriteproject, namely Vale Inco’s Goro nickel project,was commissioned during 2012. Nickel is alsorecovered via direct SX (no prior precipitation

Evaluation of a Versatic 10 acid/Nicksyn™synergistic system for the recovery of nickeland cobalt from a typical lateritic leach liquorby A.C. du Preez* and M.H. Kotze*

SynopsisMintek has been involved in extensive test work since the early 1990s onthe recovery of nickel and cobalt from leach liquors saturated in calcium,using synergistic solvent extraction systems. During this period theNicksyn™ reagent was developed, optimized, commercially manufactured,and tested by Tati Nickel on a demonstration plant for more than 2800operating hours. Efficient recovery of nickel without the co-extraction ofcalcium, thus avoiding gypsum formation in the extraction and strippingcircuits, was illustrated. This synergistic system was recently evaluated ona laboratory scale for the recovery of nickel and cobalt from syntheticlateritic sulphate leach liquor containing about 3 g/L nickel, 0.5 g/L cobalt,0.7 g/L manganese, 20 g/L magnesium, and with calcium at saturation.Extraction and stripping parameters were determined for this feed liquorand are discussed in this paper.

Keywordssolvent extraction, nickel laterites, reagents, synergistic systems.

* Mintek, Randburg, South Africa.© The Southern African Institute of Mining and

Metallurgy, 2014. ISSN 2225-6253. This paperwas first presented at the, Base Metals Conference2013, 2–4 September 2013, IngwenyamaConference & Sports Resort, Mpumalanga.

375The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 MAY 2014 ▲

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Evaluation of a Versatic 10 acid/Nicksyn™ synergistic system for the recovery of nickel

of nickel and cobalt) using Cyanex 301 [bis(2,4,4-trimethylpentyl) dithiophosphinic acid]. Metal distributionequilibria generated for pH vs. extraction (Figure 3) showthat nickel and cobalt can be extracted without neutralizationduring extraction (Mihaylov et al., 2000). However, thisreagent has a number of drawbacks, namely:

(i) The very strong extraction of copper requiresefficient removal of copper from the full flow of thepregnant leach solution via ion exchange (IX),which would be expensive. Furthermore, if anybreakthrough from the IX circuit occurs, the copperwould report to the Cyanex 301 circuit, requiringthe copper to be stripped with thiourea in sulphuricacid medium

(ii) Strong nickel and cobalt extraction makes strippingdifficult. The Goro process was designed to usehydrochloric acid stripping in four stripping stages,each with 5 minutes’ residence time in the mixer, atan operating temperature of 60°C and a residualhydrochloric acid concentration of 3 M

(iii) Due to the high residual acid concentration requiredfor stripping, pyrohydrolysis is used for nickelrecovery as NiO

(iv) The introduction of chloride into the systemrequires more sophisticated materials ofconstruction, and can have environmental concernsamong other complications

(v) Cyanex 301 is chemically unstable and in thepresence of oxygen and metals such as ferric, thereagent is oxidized to form a disulphide. Hence, airhas to be excluded from the operating system,which Goro achieves by employing Bateman PulsedColumns. The reagent can be regenerated usingsulphuric acid and zinc powder.

Mintek developed the Nicksyn™ reagent during the1990s, and together with V10 it provides an alternative andmore cost-effective approach to direct nickel and cobalt SXfrom laterite processing liquors. This synergistic system hasbeen demonstrated over 2800 hours on the Tati NickelActivox® demonstration plant in Botswana, and has sincebeen commercialized (Du Preez et al., 2007; Masiiwa et al.,2008). This paper describes the technical performance of theV10 and Nicksyn™ synergistic system for the recovery ofnickel and cobalt from neutralized HPAL lateritic leach liquor.

376 MAY 2014 VOLUME 114 The Journal of The Southern African Institute of Mining and Metallurgy

Figure 1—Primary hydrometallurgical processing options for lateriticnickel ores (Dalvi, Bacon, and Osborne, 2004)

Figure 2—Simplified block flow diagram of the Bulong flow sheet

Figure 3—Metal extraction by 15 vol.% Cyanex 301 at an O:A phaseratio of 0.5 using NaOH for pH adjustment

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Experimental procedures

Laboratory tests

Analytical methods

All metal analyses were done by inductively coupled plasma-optical emission spectroscopy (ICP-OES) with a detectionlimit of 2 mg/L for all metals. Organic samples were strippedwith sulphuric acid (approx. 1 M) at an organic-to-aqueous(O:A) phase ratio of 0.5, after which the strip liquors weresubmitted for analysis.

Reagents and solutions

Versatic 10 acid (V10, a tertiary-branched carboxylic acid)was obtained from Chemquest (produced by ResolutionPerformance Products Ltd.), while Nicksyn™ was prepared forMintek by an international, reputable manufacturer. Thechemical composition and technical information on Nicksyn™

remain the proprietary information of Mintek and cantherefore not be disclosed. Appropriate dilutions of V10 aloneand V10 mixtures with Nicksyn™ were done using analiphatic hydrocarbon diluent, Shellsol D70, which wasobtained from Shell Chemicals.

Metal distribution studies

Metal distribution equilibria (pH vs. extraction profiles,extraction and stripping isotherms) were generated bycontacting the required organic phase with the appropriatefeed solution at various O:A phase ratios, using rapidmagnetic stirring and controlling the temperature in a water-jacketed glass vessel at 25ºC. An equilibrium time of 10 to 15minutes was allowed to ensure steady state was reached.

The pH value of the aqueous phase (in the case ofextraction isotherms) was adjusted or controlled by theaddition of sodium hydroxide solution (approx. 1 to 10 M),using a calibrated combined glass reference electrode.Samples of the organic phase were taken immediately afterthe aqueous samples to prevent possible re-equilibration aftereach pH adjustment. Aqueous samples were submitted foranalyses. Organic samples were stripped with 1 M H2SO4(O:A phase ratio of 0.5), after which the strip liquors wereanalysed for the relevant elements via ICP-OES.

Organic phase (0.5 M V10 plus 0.25 M Nicksyn™) wasbatch-loaded for stripping purposes by contacting portions offresh organic phase with synthetic laterite solution (at an O:Aphase ratio of 0.45) at pH 6.0 and at 25ºC. Samples ofaqueous and organic phases were analysed by ICP-OES. Thisprocedure was repeated two more times to simulate the threestages required according to the McCabe-Thiele construction(Figure 7). The batch-loaded organic phase obtained wasthen contacted with a synthetic nickel spent electrolyte(approx. 71 g/L nickel in 40 g/L H2SO4) at different O:Aphase ratios at 25ºC, measuring the final pH values of theloaded strip liquors. Samples of the loaded strip liquors andorganic phases were analysed as previously described.

For the batch countercurrent extraction experiment,organic and aqueous phases were contacted (at 25°C), usingmagnetic stirring at an O:A phase ratio of 0.45. A sequenceof batch contacts that simulates the conditions of a four-stagecontinuous flow process was used as illustrated in Figure 4.Six full cycles (D to I, see Figure 4) were completed in this

way to ensure steady-state conditions. Samples of theraffinates of the fourth stages (4D to 4I), and the aqueousphases of the first (1I), second (2I), and third stage (3I) ofthe last cycle (I) were submitted for analyses. Portions of theloaded organic phases of the first stages (1D to 1I) as well asthe loaded organic phases of the second (2I), third (3I) andfourth (4I) stages of the last cycle were taken and stripped asdescribed above, after which the strip liquors were analysedby ICP-OES.

Results and discussion

Direct recovery of nickel and cobalt from a synthetic solutionrepresenting a nickel HPAL laterite leach liquor after iron,aluminium, and chromium removal was tested using theV10/Nicksyn™ system. Nicksyn™ is now commerciallyavailable, and hence offers a very attractive option to beconsidered for HPAL leach liquor.

Feed solution

A synthetic feed solution was made up from metal sulphatesalts to contain nickel, cobalt, manganese, calcium (atsaturation), and magnesium. The average of variousanalyses of the synthetic laterite feed solution is given inTable I.

Evaluation of a Versatic 10 acid/Nicksyn™ synergistic system for the recovery of nickel

377The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 MAY 2014 ▲

Figure 4—Scheme of contacts for batch countercurrent extractionexperiment

Table I

Average composition of synthetic laterite feedsolution

Feed Concentration, g/LNi Co Mn Mg Ca

Laterite leach liquor (synthetic) 3 0.5 0.66 20.2 0.46

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Evaluation of a Versatic 10 acid/Nicksyn™ synergistic system for the recovery of nickel

Organic phase compositions

The different concentrations of V10 and the molar ratios ofV10:Nicksyn™ diluted in Shellsol D70 are given in Table II.

Extraction metal-distribution equilibria (pH vs.extraction)

The origin of the synergistic effect for nickel by a carboxylicacid (such as V10, which exists in the form of dimers H2A2),with the addition of a synergistic compound (L) such asNicksyn™, has been discussed previously in terms ofcompeting equilibria, and is given in Equations 1 to 3 (Du Preez et al., 2007; Masiiwa et al., 2008):

[1]

[2]

[3]

where H2A2 denotes the carboxylic acid dimer and L denotesthe synergist.

Results for selected pH vs. extraction isotherms areshown in Figure 5 and Figure 6. The pH50 values (the pH atwhich 50% of the metal originally present in the aqueousphase is extracted under a given set of conditions) aresummarized in Table III.

Synergistic shifts in the pH50 values for the extraction ofnickel (i.e. the difference in pH50 value for V10 alone and thepH50 value for the appropriate synergistic mixture) increasedfrom 1.23 to 1.68 units when Nicksyn™ addition wasincreased from 0.125 to 0.5 M, whilst the shifts for cobaltincreased from 0.65 to 1.15 units with the same Nicksyn™

additions. The extraction of manganese was largely

unaffected, hence the separation (ΔpH50Mn-Co) between cobaltand manganese increased from 0.75 to 1.13. Extractions ofcalcium and magnesium were negligible (<1%) under theseconditions. This synergistic system therefore not onlyprovides an option for the recovery and separation of nickeland cobalt efficiently from calcium and magnesium, but alsogives the option of selecting a degree of manganese removal,with ease of pH control in practical flow sheets.

Extraction isotherms

The distribution isotherms and McCabe-Thiele constructionsfor the extraction of nickel and cobalt from synthetic lateriteleach solution generated using 0.5 M V10 plus 0.25 M

378 MAY 2014 VOLUME 114 The Journal of The Southern African Institute of Mining and Metallurgy

Table II

V10 concentrations and molar ratios ofV10:Nicksyn™

V10 Nicksyn™ V10:Nicksyn™

Vol.% M M Molar ratio

9.6 0.50 - -9.6 0.50 0.125 4:19.6 0.50 0.25 2:19.6 0.50 0.50 1:1

Table III

pH50 values for the extraction of metals from synthetic laterite leach solution using V10 alone and with Nicksyn™in Shellsol D70 at 25ºC

V10, M Nicksyn™, M pH50

Mg Ca Mn Co Ni Mn-Co Ca-Co

0.5 - >7.17 >7.17 >7.17 6.52 6.37 >0.65 >0.650.5 0.125 >7.19 >7.19 6.62 5.87 5.14 0.75 >1.320.5 0.25 >7.43 >7.43 6.47 5.53 4.97 0.94 >1.900.5 0.5 >7.30 >7.30 6.50 5.37 4.69 1.13 >1.93

* A > sign preceding a pH value indicates that 50% metal extraction was not reached at this pH value

Figure 5—Extraction of metals from synthetic laterite leach solution by0.5 M V10 alone and 0.5 M V10 plus 0.125 M Nicksyn™ in Shellsol D70 at25ºC

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Nicksyn™ (at pH 6.0) and 0.5 M V10 plus 0.5 M Nicksyn™

(at pH 5.8) in Shellsol D70 at 25ºC are shown in Figure 7 andFigure 8, respectively. The McCabe-Thiele construction wasdrawn for optimum recovery of nickel with the idea to gaugewhat the possible recovery for cobalt could be under thechosen conditions. In order to determine the effect ofinsufficient aluminium removal prior to the SX circuit, about200 mg/L aluminium (as sulphate) was added to the leachsolution used for the generation of the extraction isotherms.

The McCabe-Thiele construction on the isothermindicated that a loading of about 7.1 g/L nickel could beachieved in three countercurrent extraction stages at an O:Aphase ratio of 0.45. The maximum loading of cobalt underthese conditions was about 300 mg/L. Calcium andmanganese co-extraction were about 6 mg/L and 34 mg/L,respectively.

The results for the McCabe-Thiele construction shown inFigure 8 indicated that a slightly higher loading of about 9.6 g/L nickel could be achieved with the increased ratio ofV10:Nicksyn™ of 1:1 in three countercurrent extractionstages at an O:A phase ratio of 0.33. Cobalt loading underthese conditions was slightly lower than 400 mg/L. The co-extractions of calcium and manganese under these conditionswere about 4 mg/L each.

Although slightly better separation (1.13 vs. 0.94 pHunits) between cobalt and manganese, and higher loading ofcobalt together with nickel (7.2 vs. 9.7 g/L) were achieved

using a V10:Nicksyn™ ratio of 1:1 compared with theV10:Nicksyn™ ratio of 2:1, respectively, the additional costsof the increased Nicksyn™ concentration should beconsidered on an economic basis for each application.

In order to recover all cobalt together with nickel, the O:Aphase ratio (and possibly the number of stages) would haveto be adjusted to compensate for cobalt being ‘crowded off’ asillustrated in the unfavourable isotherms obtained for cobalt(Figure 9 and Figure 10).

The McCabe-Thiele construction redrawn for optimumcobalt recovery under the conditions tested indicated that ahigher O:A phase ratio (1.86 vs. 0.45 as previously drawn fornickel recovery) should be employed to ensure a loading ofapproximately 257 mg/L cobalt without being ‘crowded off’by nickel, using two countercurrent extraction stages. Underthese conditions, nickel would still be recovered (leaving <2 mg/L in the raffinate) with minimum co-loading ofcalcium (5 mg/L), magnesium (4 mg/L), and manganese (7 mg/L).

For optimum cobalt recovery under these conditions theMcCabe-Thiele construction indicated that an O:A phase ratioof 1.27 (instead of 0.33 previously drawn for nickel recovery)

Evaluation of a Versatic 10 acid/Nicksyn™ synergistic system for the recovery of nickel

The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 MAY 2014 379 ▲

Figure 6—Extraction of metals from synthetic laterite leach solution by0.5 M V10 plus 0.25 and 0.5 M Nicksyn™ in Shellsol D70 at 25ºC

Figure 8—Distribution isotherm for the extraction of nickel and cobaltfrom synthetic laterite leach solution by 0.5 M V10 plus 0.5 M Nicksyn™in Shellsol D70 at 25ºC and pH 5.8

Figure 7—Distribution isotherm for the extraction of nickel and cobaltfrom synthetic laterite leach solution by 0.5 M V10 plus 0.5 M Nicksyn™in Shellsol D70 at 25ºC and pH 6.0

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Evaluation of a Versatic 10 acid/Nicksyn™ synergistic system for the recovery of nickel

would be required to achieve a loading of about 400 mg/Lcobalt without ‘crowding off’ by nickel in two to threecountercurrent extraction stages. Under these conditions, asimilar nickel recovery would still be expected (<2 mg/L inthe raffinate) with minimum co-loading of calcium (4 mg/L),magnesium (4 mg/L), and manganese (14 mg/L). In boththese cases optimum design should include economic cobaltrecovery without calcium co-extraction.

Batch countercurrent test work

A batch countercurrent test was performed for the extractionof cobalt and nickel with 0.5 M V10 plus 0.25 M Nicksyn™ inShellsol D70 at 25ºC. Four extraction stages were used withan O:A phase ratio of 0.45 and an equilibrium pH value of6.0 in each stage (i.e. a flat pH profile) over a total of six fullcycles (see Figure 4). The results are illustrated in Figure 11.

A loaded organic phase (Stage 1I) containing about 7.7 g/L nickel was obtained at steady state (Cycle I, Figure 4),leaving about 5 mg/L nickel in the raffinate, which related to>99% extraction. Cobalt was loaded up to approximately 1.4 g/L in stage 3I, after which it was ‘crowded off’ by nickelto only about 176 mg/L on the loaded organic phase (Stage1I) under these conditions. In order to recover both cobaltand nickel, a higher O:A phase ratio and probably more

stages would be required, as shown in Figure 9 and Figure10. A pH profile (and not a flat profile as was employed here)over all the stages could also assist to enhance cobaltrecovery, providing no calcium is co-extracted.

The co-loading of some impurities at steady state of thebatch countercurrent experiment is shown in Figure 12.

The efficient removal of aluminium prior to nickel andcobalt recovery is strongly indicated, as aluminium wasstrongly extracted by the synergistic mixture (from the feedsolution containing approximately 213 mg/L only <2 mg/Lwas left in the raffinate). Magnesium extraction was low (<4 mg/L on the loaded organic for stages 1I, 2I, and 4I) and the anomaly observed in the higher loading in stage 2I(36 mg/L on the loaded organic phase) was most likely dueto analytical error. Manganese was co-loaded (between 40and 62 mg/L), with calcium co-loading minimal (approx. 7 mg/L) based on loaded organic phase analyses. Phaseseparations were clear in all stages and no crud formationwas observed.

Further optimization of the O:A phase ratio, number ofstages, and pH profile across the extraction circuit wouldhave to be done to optimize cobalt recovery (including nickel)as well as limiting the co-loading of unwanted impuritiessuch as manganese. This test work has to be performed foreach individual lateritic feed solution.

380 MAY 2014 VOLUME 114 The Journal of The Southern African Institute of Mining and Metallurgy

Figure 9—Distribution isotherm for the extraction of cobalt fromsynthetic laterite leach solution by 0.5 M V10 plus 0.25 M Nicksyn™ inShellsol D70 at 25ºC and pH 6.0

Figure 10—Distribution isotherm for the extraction of cobalt fromsynthetic laterite leach solution by 0.5 M V10 plus 0.5 M Nicksyn™ inShellsol D70 at 25ºC and pH 5.8

Figure 11—Batch countercurrent extraction of nickel and cobalt fromsimulated laterite solution by 0.5 M V10 plus 0.25 M Nicksyn™ inShellsol D70 in 4 stages and an O:A phase ratio of 0.45 and at 25ºC

Figure 12—Batch countercurrent extraction of impurities fromsimulated laterite solution by 0.5 M V10 plus 0.25 M Nicksyn™ inShellsol D70 in 4 stages and an O:A phase ratio of 0.45 and at 25ºC

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Stripping test work

In order to conduct stripping test work, a batch of freshorganic phase (0.5 M V10 plus 0.25 M Nicksyn™ in ShellsolD70) was preloaded to reasonably represent a loaded organicphase that would be expected from an extraction circuit.Synthetic spent nickel electrolyte was prepared to containapproximately 71 g/L nickel and 40 g/L H2SO4 (which shouldrelate to a delta of approximately 24 g/L nickel). A strippingisotherm was generated and the results are shown in Figure13.

Loaded strip liquor containing about 95 g/L nickel (at apH value between 1.1 and 4.2) could be generated. TheMcCabe-Thiele construction indicated that two to three stagesand an O:A phase ratio of 3.3 would be required to achievethis. Strip liquor pH values measured for stripping done atO:A phase ratios of 0.1 to 2.0 varied between 0.5 and 1.1,which indicates that adequate sulphuric acid was availablefor complete stripping of nickel. Stripping at O:A phase ratiosof 5.0, 8.0 and 10.0 resulted in strip liquors exhibiting pHvalues of 4.2, 4.6, and 4.8 respectively, which indicatedunfavourable stripping conditions for nickel. If plantoperation required these operating conditions, it would beadvisable to employ pH control (at about 3) in order tofacilitate efficient stripping of nickel in a minimum number ofstages and to provide an advanced electrolyte suitable forelectrowinning.

Any co-loaded manganese could be removed byscrubbing with pH-adjusted water or a portion of the loadedstrip liquor, depending on downstream requirements. Theaddition of a washing stage is recommended for removal ofentrained aqueous phase from the loaded organic phase.

Conclusions

➤ The V10 plus Nicksyn™ synergistic mixture wasevaluated for the use of direct solvent extraction ofnickel and cobalt from lateritic leach liquor

➤ Nicksyn™ is now commercially available via a secured,reputable international supplier. It is a very attractiveoption and should be included in the evaluation for allHPAL hydrometallurgical projects

➤ For a solution containing about 3 g/L nickel, 0.5 g/Lcobalt, 0.7 g/L manganese, 20 g/L magnesium, and

with calcium at saturation, the following parameterswere found:

– Three to four extraction stages would be requiredfor optimum nickel and cobalt recovery. A pHprofile (instead of the same pH value in everystage) could be employed to assist in the recoveryof cobalt. The O:A phase ratio could vary andwould be determined by optimum cobalt recoveryexcluding co-extraction of calcium

– If required, a scrub stage could be included forremoval of co-extracted impurities such asmanganese and calcium. Typical scrub solutionswould be pH-adjusted water, or a portion of theloaded strip liquor

– One wash stage is recommended for removal ofany entrained aqueous phase from the loadedorganic phase

– Two to three stripping stages would be requiredusing spent electrolyte at an O:A phase ratio ofabout 3. The pH of stripping should be maintainedat about 3 to facilitate a suitable advancedelectrolyte for subsequent electrowining.

➤ Nicksyn™ was found to be stable over a testing periodof 100 days (representing stripping conditions) in thelaboratory. The solubility was found to be between 1and 3 mg/L under specific conditions tested (Du Preezand Preston, 2004).

➤ Mintek is currently involved in an international projectwith a commercial client where Nicksyn™ will beincluded in the definitive feasibility study for nickeland cobalt recovery from laterite solutions.

Acknowledgements

This paper is published by permission of Mintek. The authorswould like to thank Nosipho Cola for her contribution to thegeneration of experimental data.

ReferencesBACON, G. and MIHAYLOV, I. 2002. Solvent extraction as an enabling technology

in the nickel industry. Journal of the Southern African Institute of Miningand Metallurgy, November/December 2002. pp. 435–443.

DALVI, A.D., BACON, W., and OSBORNE, R. 2004. The past and the future of nickellaterites. PDAC 2004 International Convention, Trade Show & InvestorsExchange, March 2004.

FLETT, D. 2005. Solvent extraction in hydrometallurgy: the role of organophos-phorus extractants. Journal of Organometallic Chemistry, vol. 690. pp. 2426-2438.

MIHAYLOV, I., KRAUSE, E., COLTON, D.F., and OKITA, Y. 2000. The development ofa novel hydrometallurgical process for nickel and cobalt recovery fromGoro laterite ore. CIM Bulletin, vol. 93, no. 1041. pp. 124–130.

DU PREEZ, R., KOTZE, M., NEL, G., DONEGAN, S., and MASIIWA, H. 2007. Solventextraction test work to evaluate a Versatic 10/Nicksyn™ synergisticsystem for nickel-calcium separation. Proceedings of the Fourth SouthernAfrican Conference on Base Metals, Swakopmund, Namibia, 23-27 July2007. Southern African Institute of Mining and Metallurgy, Johannesburg.pp. 193–210.

MASIIWA, H., MATHE, O., DONEGAN, S., NEL, G., DU PREEZ, R., KOTZE, M., andIRELAND, N. 2008. Evaluation of Versatic 10 and Versatic 10/Nicksyn™

synergistic systems for nickel-calcium separation on the Tati BMRhydrometallurgical demonstration plant, Proceedings of ISEC 2008,Tucson, Arizona, September 2008. pp. 169-175.

DU PREEZ, A.C. and PRESTON, J.S. 2004. Separation of nickel and cobalt fromcalcium, magnesium and manganese by solvent extraction withsynergistic mixtures of carboxylic acids. Journal of the South AfricanInstitute of Mining & Metallurgy, vol. 104, no. 6. pp. 333–338. ◆

Evaluation of a Versatic 10 acid/Nicksyn™ synergistic system for the recovery of nickel

The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 MAY 2014 381 ▲

Figure 13—Distribution isotherm for the stripping of nickel from batch-loaded 0.5 M V10 plus 0.25 M Nicksyn™ in Shellsol D70 with syntheticspent electrolyte at 25ºC

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Introduction

Mintek conducted a significant amount ofwork on the development of ion exchangefibres (IXFs) during the 1980s and 1990s.Scale-up of production of these materials was,however, not feasible at that stage. Similardevelopment work was done at the Institute ofPhysical and Organic Chemistry (IPOC) inBelarus, who have commercialized theproduction of these materials more recently(Shunkevich et al., 2004; Soldatov et al.,2004; Vatutsina et al., 2007). This providedMintek with the opportunity to resume itsdevelopment on the use of IXFs formetallurgical processes.

An important feature of fibrous ionexchangers is their extremely high osmoticstability, which allows multiple cycles ofdrying and moistening as well as conversion

from one ionic form to the other withoutdestruction of the filaments. The filteringlayers of ion exchange fabrics do not signifi-cantly change their volume and permeabilitydue to swelling of the filaments, which occursas ion exchange proceeds and ionic statechanges.

This paper demonstrates the possibility ofusing fibrous ion exchangers for hydrometal-lurgical applications such as copper removalfrom cobalt electrolyte. A comparison ispresented of the design parameters andindicative costs for impurity removal usingfibrous ion exchangers and granular resinshaving similar functionality.

Experimental procedures

A number of IXFs were tested for theirselectivity, maximum copper capacity, andequilibrium behaviour in order to choose themost suitable fibre for the subsequent investi-gations, which consisted of split elution, mini-column, and mini-pilot plant tests. A similartest programme was carried out on a granularresin having the same functionality as theselected fibre, namely a resin containingimino-diacetic acid groups.

Maximum loading capacities for divalentmetals

A maximum loading capacity test wasperformed in order to determine the maximumcapacity of the fibres at different pH values fordivalent metals using copper. A fibre-to-solution ratio was used that would provide anexcess of metal in solution with respect to the

Evaluation of different adsorbents for copperremoval from cobalt electrolyteby V. Yahorava*, M. Kotze*, and D. Auerswald†

SynopsisIon exchange is considered to be an effective technology for the removal ofvarious impurities from cobalt advance electrolytes. With the correct choiceof resin, ion exchange can consistently remove the required impurities tothe levels for the production of high-grade cobalt metal. Although ionexchange was in the past used primarily for nickel removal, more recentlyit has been also considered for the removal of copper, zinc, and cadmium.Generally, granular ion exchange products are used, but Mintek is currentlyevaluating ion exchange fibres for a number of applications, including theremoval of copper from cobalt advance electrolytes. Fibrous ion exchangershave major advantages compared to granular resins in that they havesignificantly higher reaction rates, and wash water volumes could belimited.

Granular and fibrous ion exchangers were evaluated and compared forthe removal of copper from cobalt advance electrolyte. A syntheticelectrolyte containing 50 g/L cobalt and 50 mg/L copper was used for thetest work. Equilibrium isotherms, mini-column tests, and split elution testswere done. The results were used to size a full-scale operation to treat 100 m3/h of electrolyte. The potential cobalt losses or recycle requirementswere estimated, and data to calculate indicative operating costs for eachadsorbent was generated. This information was used for a techno-economiccomparison of granular and fibrous ion exchange systems for the removalof copper from cobalt advance electrolyte.

Keywordsion exchange, impurity removal, copper, cobalt electrolyte.

* Mintek, South Africa.† TENOVA Bateman Africa.© The Southern African Institute of Mining and

Metallurgy, 2014. ISSN 2225-6253. This paperwas first presented at the, Base Metals Conference2013, 2–4 September 2013, IngwenyamaConference & Sports Resort, Mpumalanga.

383The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 MAY 2014 ▲

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Evaluation of different adsorbents for copper removal from cobalt electrolyte

amount of fibre used. This was done to ensure that there wasadequate copper available, so that there were no equilibriumconstraints during the determination of the maximum loadingcapacity. The tests were done by contacting air-dried fibre (inthe H+ form) with the calculated volume of a 1.3 g/L copperfeed solution at pH values of 2, 2.5, 3, 3.5, 4, and 4.5. Fibresamples were removed from the solution by filtration afterthe pH stabilized, washed with water, and stripped with anexcess of 2 M HCl solution.

‘S-curves’

The effect of pH on the extraction of metals from anequimolar mixture of Cu2+, Mg2+, Mn2+, Ca2+, Ni2+, Zn2+, andCo2+ was determined to evaluate the relative affinity of eachmaterial tested for the metals of interest.

The mass of fibre used was sufficient (in capacity) toadsorb all the metals from solution. A feed solution wasprepared from metal sulphate salts. One batch test wascarried out for each fibre, and samples were taken at pHvalues of 2, 2.5, 3, 3.5, 4, 4.5, 5, 5.5, 6, 6.5, and 7. Thecontact period for each batch test was dictated by the timerequired for the pH to stabilize at the level at which it wascontrolled. The test was stopped when the pH had been stablefor at least 1 hour. A solution sample was taken and theloaded fibre was filtered out of the barren solution. Theloaded fibre was washed and stripped by contacting it with300 mL of 2M HCl solution in a beaker while stirring using amagnetic stirrer for approximately 1 hour.

Similar experiments were done with the resin, but thefeed solution did not contain Mn since it started to precipitate(as MnO2) in the presence of the resin at a pH of 3. Thecontact period for each pH point was 24 hours. The loadedresin was separated from the solution, washed, and elutedusing 10 bed volumes (BVs) of 2 M H2SO4.

Equilibrium isotherms

Adsorption equilibrium isotherms were generated for theloading of Cu and Co onto the adsorbents by batch contact ofportions of the fibre/resin in the H+ form and syntheticsolution (100 mg/L of Cu and 50 g/L of cobalt in a sulphatemedia) at different fibre to solution volume ratios. The pHvalues for the individual experiments were controlled at 3, 4,4.5, and 5 for Fiban X-1 fibre, Lewatit TP 207 resin, Fiban K-3, K-4, and AK-22 (3) fibres respectively.

Efficiency of fibres in numerous operational cycles

The objective of this experiment was to establish theefficiency of selected materials in numerousadsorption/elution cycles. A small column was packed with afibre sample in the di-sodium form and synthetic solutionsimulating cobalt electrolyte containing 50 mg/L Cu asimpurity was passed through the column to fully load thefibre. After loading, the fibre in the column was washed withexcess water prior to elution with sulphuric acid. Eluatesamples were analysed for their Co and Cu concentrations.After elution, the fibre was converted back into the di-sodiumform for the next adsorption cycle. These adsorption andelution cycles were repeated from 10 to 20 times.

After the last cycle, a portion of the stripped fibre wasdried and analysed for any residual copper and cobalt.

Mini-column tests

Breakthrough tests for TP 207 resin and X-1 fibre

For the purpose of a techno-economic comparison betweenthe FIBAN X-1 fibre and granular TP 207 resin, mini-columnbreakthrough tests were conducted employing the conditionspresented in Table I.

Optimization of fibre column operating parameters

Adsorption within a packed-bed column is a process in whichcontinuous mass transfer occurs between two phases (themobile phase containing the solution and the solid phase ofthe packed bed). The solution concentration in both phases is a function of the contact time and the depth of theadsorption zone. The mass balance of the packed-bed reactorcan be described by the Thomas model (Thirunavukkarasu et al., 2002):

[1]

where:Ce is the effluent adsorbate concentration (mg/L) C0 the influent adsorbate concentration (mg/L) k the Thomas rate constant (L/min·mg) q0 the maximum solid phase concentration of solute

(mg/g) m the mass of the adsorbent (g) V the throughput volume (mL) Q the volumetric flow rate (mL/min).

According to the mass balance of a specific packed bed,the determining factors of the mass balance for a given beddepth of the column are the volumetric flow rate and theinitial solution concentration. For fibres it was also necessaryto check the impact of fibre packing density because whilethis parameter changes, the filtering layer resistance alsovaries and this can influence the column efficiency.Therefore, in order to optimize the adsorption process in apacked-bed fibre column, the following parameters wereexamined and their influence on the column efficiency wasestimated:

➤ Flow rate, which translated to an increase in the linearvelocity for the specific experimental set-up

➤ Packing density.

384 MAY 2014 VOLUME 114 The Journal of The Southern African Institute of Mining and Metallurgy

Table I

Mini-column test parameters

Input Lewatit TP 207 Fiban X-1

Material form di-sodium

Density, g/cm3 (dry) 0.42 0.23

Absolute dry mass, g 119 5

Flow rate, mL/h BV/h 100 902 4

Aspect ratio of resin bed (height:diameter) 4.1 1.8

Volume of adsorbent (H+ form), mL 50 22

pH 4.5

Cu feed, mg/L 100

Co feed, g/L 50

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While studying the influence of a certain parameter thevalues of the other parameters were kept constant. For all thetests 4 cm diameter columns were used. The tests wereconducted on synthetic solution containing Cu at approxi-mately 50 mg/L and Co at approximately 50 g/L. The feedsolution pH was adjusted to a value of approximately 3 bythe addition of 1 M NaOH solution. Input parameters forthese tests are presented in Table II.

Mini-pilot plant

The operational sequence for a lead-lag-lag configuration ofcolumns is illustrated in Figure 1. Initially, all three columnsare in adsorption. Transfer (loaded or lead resin/fibre columnleaves adsorption circuit for elution) occurs when the bulk ofthe mass transfer zone (MTZ) has moved through the leadcolumn (C1) into the lag columns (C2 and C3). The feed istransferred from the lead column (C1) to column C2, whichthen becomes the lead column. Column C1 is stripped and/orregenerated and returned to the adsorption circuit in the lag(or last) position. The solution is passed downflow throughthe resin bed, but upflow through the fibre bed. The mainparameters of the process employing the fibre are presentedin Table III.

The entire operation could be divided into three parts:(1) Conversion of stripped fibre into the di-sodium form.

The dry X-1 staple was packed into the columns.Regenerant (2 BVs of 1 M NaOH) was passedthrough the column at a flow rate of 10 BV/h (toconvert fibre to the di-sodium form), followed by 1 Lof wash water (upflow) to remove entrained base.

(2) Adsorption. Feed solution was passed through theadsorption column in an upflow direction. Once thefibre bed was filled with solution, the piston waspushed down to adjust the packing density to thatrequired for the specific test (a sketch of theadsorption column with fibrous ion exchanger ispresented in Figure 2). Solution was passed throughat a flow rate of 4 L/h. Solution samples were taken

from the first and last columns every 15 minutes.The first column was removed from the adsorptioncircuit to be eluted once the barren had reached 80%breakthrough, and the second column wastransferred to the lead position.

(3) Elution. Air was pumped through the column toremove entrained solution from the fibre, then 200mL of 0.2 M H2SO4 was passed downflow throughthe column at 14.2 mL/min (10 BV/h) to elute any Coloaded onto the fibre. After Co elution, 100 mL of 1M H2SO4 was passed downflow through the fibrecolumn at 14.2 mL/min (10 BV/h) to elute anyremaining metals loaded on to the fibre. The strippedfibre was washed with 500 mL of water (upflow,

Evaluation of different adsorbents for copper removal from cobalt electrolyte

385The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 MAY 2014 ▲Table II

Input parameters for optimization of the fibre columnoperation

Material Fiban X-1

Fibre form Di-sodium

Parameter to be optimized Solution flow rate

Flow rate BV/h 11 16 22 34 44Fibre volume mL 95 95 95 95 95Mass of absolute dry fibre, g 26.6 25.7 25.7 25.7 25.7Packing density g/cm3 0.28 0.27 0.27 0.27 0.27H/D 2 2 2 2 2

Parameter to be optimized Packing density*

Packing density g/cm3 0.37 0.43 0.53Flow rate BV/h 24 24 26Fibre volume mL 88 63 60Mass of absolute dry fibre, g 32.3 25.7 32.1H/D 1.8 1.3 1.2

*Fibre with different capacity was used

Table III

Parameters for fiber mini-pilot campaign

Feed flow rate L/h 4Cu loading g/kg 25Packing density g/cm3 0.4Cu feed mg/L 50Number of columns # 3Fibre volume per column mL 83Mass fibre per column g 33

Figure 1—Illustration of the operating sequence of a lead-lag-lag ionexchange circuit

Figure 2—Ion exchange fibre column for adsorption

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Evaluation of different adsorbents for copper removal from cobalt electrolyte

50 mL/min) after elution. The piston compressing the fibrewas lifted up and 200 L of 1 M NaOH was passed upflowthrough the column at 20 mL/min (the colour of the fibrebecame bright orange) to convert the fibre to the di-sodiumform. The converted fibre was washed with 1 L of water toremove entrained NaOH (upflow, 67 mL/min). The entireelution and regeneration procedures of the fibre took 1 hour.The column was returned to the system as the lag column inthe third position, as shown in Figure 1. The fibre wascompressed until the required volume was attained only afterit was filled with solution. This was necessary because thefibre swelled in the base form and the resistance of thefibrous layer was high enough to block the solution flowthrough the column.

Results and discussion

Characteristics of materials tested

Four fibrous ion exchange materials (FIBAN®) were tested incomparison with a conventional granular ion exchanger. Theimino-diacetic acid resin tested, Lewatit TP207, is a productfrom Lanxess. Functional groups and backbone structures ofthese materials are presented in Figure 3.

Although FIBAN® K-3 and K-4 have the same carboxylicfunctionality, their matrices are different. FIBAN® K-3 wassynthesized from a polyacrylonitrile (PAN) backbone, whileK-4 is a product of radiation grafting of acrylic acid to apolypropylene (PP) backbone (Shunkevich et al., 2004).Generally, the difference in matrices of ion exchangers has animpact on their physical properties, and not their chemicalproperties. The main distinctive feature of the fibressynthesized on the PAN backbone is possible hydrolysis ofthe nitrile groups of the matrix over a period of time,resulting in the formation of additional carboxylic acid groups(Soldatov et al., 2004; Vatutsina et al., 2007).

Maximum copper loading capacities

Maximum copper loading capacities as a function of pH forthe various fibres indicated that the fibres function optimallyat relatively high pH values. However, Cu could start precipi-tating at pH 5, so the maximum Cu loadings were done at pH4.5. The results for the maximum copper loading capacitiesfor the various fibres and the TP 207 resin tested arepresented in Table IV.

Loading capacities found experimentally were somewhatlower than the metal loading capacities that would have been

predicted from the theoretical exchange capacities. Thissuggests that not all the functional groups within thestructure of the fibres and the resin are available for theextraction of copper at lower pH values. Fiban AK-22(3)showed a higher capacity than was expected, whichpresumably was caused by complexation of copper by thepolyamine groups on the fibre.

pH vs extraction isotherms

The ‘S-curves’ (pH vs extraction) were constructed underconditions whereby the fibres/resins had sufficient capacitywithin one batch test in order to adsorb all the metals fromsolution (i.e. excess fibre/resin and limited metals insolution). The results of ‘S-curve’ tests can be used toestablish the metal selectivity order, and the aim of this testwork was to select the fibre that would provide the bestcopper selectivity over cobalt. The metal loading capacitiesincreased with pH, and the S-curves achieved for Fiban X-1are presented in Figure 4 as an example.

Results indicated that TP 207 and Fiban X-1would appearto function optimally at about pH 3 for Cu removal from Coelectrolyte. The other fibres generally would requiresomewhat higher pH values. The order of selectivity for thevarious fibres and TP 207 were determined as follows:

386 MAY 2014 VOLUME 114 The Journal of The Southern African Institute of Mining and Metallurgy

Table IV

Copper maximum loading capacity (pH 4.5)

Adsorbent Maximum theoretical capacity, meq/g* Theoretical Cu maximum loading, mmol/g Cu maximum loading, mmol/g

Fiban AK- 22(3) 1.0 0.5 1.3Fiban K-3 5.4 2.7 1.9Fiban K-4 5.0 2.5 2.0Fiban X-1** 3.7 1.9 1.8Lewatit TP207 5.2 2.62 2.57

*Based on content of –COOH groups**In some tests X-1 with 3.2 meq/g maximum theoretical capacity was used, namely, for the mini-pilot plant, re-usability testing and design

Figure 3—Chemical structures of materials tested

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Fiban X-1: Cu > Ni > Zn Co > Mn > Ca > MgFiban K-3: Cu > Zn > Mn Ni Co > Ca > MgFiban K-4: Cu > Zn > Ni Co Ca > Mn > MgFiban AK-22(3): Cu > Ni > Co Zn > Mg Mn CaLewatit TP 207: Cu > Ni > Zn > Co > Ca > MgThe selectivity orders observed for Lewatit TP 207 and

fibre Fiban X-1 were similar, which could be expected due tothe similarity of their functional groups (Figure 3).

Equilibrium isotherms

Equilibrium isotherms were constructed using syntheticcobalt electrolyte solution containing 100 mg/L Cu. The pHvalues for the individual experiments were controlled at 3, 4,4.5, and 5 for Fiban X-1 and TP 207, K-3, K-4, and AK-22(3) respectively. Results are presented in Figure 5.

The Langmuir equilibrium model (Soldatov et al., 2004)was fitted to the equilibrium data obtained and selectivitycoefficients were calculated. The main results of theequilibrium tests are presented in Table V.

Fiban AK-22(3) provided the highest loading of copperand the lowest co-loading of cobalt, while Fiban X-1, K-3,and K-4 had lower copper loadings and significantly higherCo co-loadings. Unnecessarily high Co co-loadings wouldgenerally increase the reagent amounts required to load andstrip the fibre, and might cause a higher Co loss.

Langmuir parameters indicate that the resin had a highermaximum copper loading (the value of parameter a charac-terizes the saturation adsorption capacity). The highestaffinity between the ion exchanger and copper was shown byAK-22(3) (b = 0.38 L/mg).

Based on these results the various adsorbents testedcould be arranged in the following order with regard to theirselectivity for copper over cobalt:

Fiban AK-22(3) > Fiban X-1 > Lewatit TP207 > Fiban K-4 > Fiban K-3Fibres with a selectivity for copper over cobalt higher

than that of the Lewatit TP 207 resin were chosen for furthertest work.

Re-usability of Fiban X-1 and AK-22(3)

Fiban AK-22(3) was noticed to change colour from white topinkish, which indicated the possibility of cobalt poisoning ofthe fibre. Following this observation, Fiban X-1 and AK-

22(3) were selected and subjected to numerous cycles ofadsorption/elution to establish if any poisoning was evident.

The results of loading and elution cycles of Fiban X-1 and AK-22(3) are presented in Figures 6 and 7 respectively.Analysis of residual Cu/Co at termination of the tests is listed in Table VI.

Results indicate that Fiban AK-22(3) was poisoned withcobalt and completely lost its capacity within only 10 cyclesof adsorption and elution. Hence, this material was excludedfrom further test work.

Fiban X-1 retained its selectivity and capacity over 20cycles of adsorption and desorption (on average 0.82 mmol/gof divalent metals was loaded with only 4% variation in the

Evaluation of different adsorbents for copper removal from cobalt electrolyte

The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 MAY 2014 387 ▲

Table V

Main results of adsorption equilibrium tests

Materials X-1 K-3 K-4 AK-22(3) TP 207

Langmuir a 26 47 39 32 164b 0.04 0.01 0.01 0.38 0.01

RSQ 0.99 0.98 0.98 0.92 0.95

Maximum copper loadings mg/g 20 24 21 31 97Cobalt co-loading 10 72 55 13 82

Selectivity Cu/Co* 1000 175 197 1304 591

Figure 4—Extraction versus pH for Fiban X-1

Figure 5—Copper adsorption equilibrium isotherms

Figure 6—Cobalt and copper loading of FIBAN X-1 over 20adsorption/elution cycles based on elution data

whereCuresin, Coresin - metal loaded onto the material; mg/g

Cubarren, Cobarren – metal concentration in solution, mg/L.

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Evaluation of different adsorbents for copper removal from cobalt electrolyte

loading capacity, average selectivity coefficient obtained was996), and no residual copper/cobalt was found in the fibreafter the final stripping cycle, as shown in Table VI.

Mini-column tests

Comparison of breakthrough curves for TP 207 resin andX-1 fibre

Mini-column breakthrough tests were done to allow a techno-economic comparison between FIBAN X-1 fibre and thegranular TP 207 resin. Synthetic solution containing 100mg/L Cu and 50 g/L Co at pH 4.5 was passed downflowthrough the column containing fibre/resin at a flow rate of 2 BV/h. Results obtained are presented in Table VII andFigure 8.

Results indicated that the fibrous ion exchanger had aconsiderably shorter mass transfer zone compared to theresin, which would reduce the size of the plant significantly.It also had a higher selectivity coefficient for Cu over Co,which should result in lower operating costs.

Influence of column parameters on performance

Column parameters, including flow rate, packing density, andaspect ratio, were varied in order to optimize the designparameters for the fibre ion exchange column.

Fiban X-1 loading efficiency was tested at five differentflow rates, and three different packing densities or aspectratios. Results are depicted in Figure 9.

An increase in the linear flow rate increased the height ofthe MTZ, but decreased its residence time and thereforedecreases the service time of the bed. However, the total

388 MAY 2014 VOLUME 114 The Journal of The Southern African Institute of Mining and Metallurgy

Figure 7—Cobalt and copper loading of FIBAN AK-22(3) over 10adsorption/elution cycles based on elution data

Table VII

Mini-column breakthrough results

TP 207 resin Fiban X-1 fibre

Cu loading, g/kg 97 35Co loading, g/kg 82 28Selectivity for Cu/Co 588 638Flowrate, BV/h 2a 4.7b

Linear velocity, mm/sec 0.057 0.051Height of MTZ, cm 8.2 2.8

a Resin volume was measured via tapped wet-volume methodb Fibre volume was controlled by pressing with a piston and could be

varied depending on desired packing density

Table VI

Co/Cu content in Fiban X-1 and AK-22(3)

FIBAN® Co, % Cu, %

X-1 <0.05 <0.05AK-22(3) 2.26 <0.05

Figure 8—Copper breakthrough curves for TP 207 and Fiban X-1

Figure 9—Influence of the column parameters on bed volumes that canbe treated for 1% Cu breakthrough: (a) flow rate at 0.27 g/cm3 fibredensity; (b) packing density at 24 BV/h flow rate

Packing density, g/cm3

BV

till

1%

bre

akth

rou

gh

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loading capacity of the fibre bed was constant (20 mg/g ofCu, 8 mg/g of Co). The choice of the optimum flow rate wouldthen be dictated by design requirements.

Increasing of fibre packing density leads to an increase inthe hydrodynamic resistance of the filtering layer to solutionflow. At a packing density >0.4 g/cm3 a problem with thefiltering layer resistance was observed. After conversion ofthe fibre into the basic form the fibre swells and blocks thefree flow of solution at a high packing density. This is typicalfor carboxylic ion exchange fibres. Thus, the optimumpacking density (taking into account the need to convert thefibre into basic form during current tests) was 0.4 g/cm3 forthe laboratory equipment.

The following conclusions were drawn based on theresults obtained during these tests where the parameterswere varied:

➤ The influence of flow rate on fibre efficiency wasdescribed by exponential decay with variation ofresidence time for the mass transfer zone from 1.2 to 3 minutes

➤ The optimum packing density, where the pressure dropwas reasonable, was 0.4 g/L. However, it wasnecessary to maintain a lower packing density duringconversion of the stripped fibre into the basic form,rinsing of excess NaOH, and the initial adsorptionphase. This was required as the resistance of the fibrebed hampered solution flow under these conditions.The fibre should be pressed to a higher density onlyafter adsorption has started.

Mini pilot-plant campaign results

The results of the countercurrent process are presented inFigure 10 and Table VIII. A lead-lag-lag fixed bedarrangement was used. Synthetic solution containing 50 mg/L Cu and 50 g/L Co at pH 3 was passed through thesystem at 4 L/h flow rate (total mass of absolute dry fibrewas 94.5 g, total volume per 3 columns was around 240 mL).

As can be seen from the graph at the bottom of Figure 10,no copper breakthrough (> 0.5 mg/L in effluent stream) afterthe last column was observed during the mini-pilot planttests. Ten transfers (of the lead columns) were done.

The average cobalt loss per column with respect to thedelta between cobalt advance and spent electrolytes was0.14±0.05 %. This could probably be lowered duringoptimization of the technique to selectively strip co-loaded Coprior to Cu stripping,, to achieve the levels that were obtainedduring the mini-column test work.

Comparison between fibre and resin plant designs

Based on the results of the mini pilot-plant campaign, andequilibrium and mini-column tests, the plant sizing, reagentconsumption, and Co recycle/loss for TP207 resin and fibreFiban X-1 were compared in order to establish theadvantages or disadvantages of fibrous ion exchanger forhydrometallurgical applications. The results of calculationsfor plants using resin or fibre are presented in Table IX.

In spite of the fact that the loading capacity of TP207 forcopper is double that of the fibre, the use of Fiban X-1 hasadvantages compared with the resin:

➤ Somewhat better selectivity for copper over cobalt➤ Shorter mass transfer height allows an increase in the

productivity of the adsorbent➤ It would allow a decrease in the total volume of fibre

used per column; faster fibre regeneration also allows adecrease in fibre volume per column

➤ The total volume of the plant can be reduced by morethan 90% using fibre, reducing the costs of theadsorbent as well as CAPEX

➤ Backwash might not be necessary➤ Cobalt losses per annum are around 50% less for the

fibre plant➤ Various shapes and sizes of columns can be used.

CAPEX requirements

The data presented in Table IX was used to size an ionexchange plant using either TP207 resin or the Fiban X-1fibre as the adsorbent. An ’order-of-magnitude’ mechanicalequipment cost was then calculated for each scenario.

The lower volume of fibre required results in considerablysmaller columns, while the shorter cycle time reduces thevolumes of all related tankage. The result is that themechanical equipment cost required for the fibre plant isabout 25% of that required for a resin-based plant.Adsorbent cost can generally form a significant portion of thetotal capital cost of an ion exchange plant, especially wherechelating resins such as the iminodiacetic acid resins areused. Because the adsorbent volume/mass is far lower for thefibre plant, further savings will be achieved in the cost of the’first fill’.

Evaluation of different adsorbents for copper removal from cobalt electrolyte

The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 MAY 2014 389 ▲

Table VIII

Mini pilot-campaign results

Average height of MTZ cm 5Average cobalt loss per column % 0.14±0.05Average copper loading mg/g 24

Figure 10—Cu adsorption profiles (lead-lag-lag system)

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Evaluation of different adsorbents for copper removal from cobalt electrolyte

Conclusions

Several fibrous ion exchangers were investigated. Fiban X-1with iminodiacetic acid groups proved to be the bestadsorbent for copper removal from cobalt electrolyte. Testresults obtained during evaluations of FIBAN® X-1 fibrousion exchanger and granular ion exchanger Lewatit TP 207were compared to assess the potential cost implications of afibrous ion exchanger. This comparison established that thefibrous ion exchanger FIBAN® X-1 could be suitable forhydrometallurgical processes. Preliminary indications are thatit could offer major cost savings compared to conventionalgranular iminodiacetic acid resin. Further work on Curemoval is focused on evaluation of the fibre and resin withlimited or no regeneration included. Ultimately, the form inwhich the fibre or resin is to be employed will be an economicdecision. Removal of entrained and co-loaded cobalt from thefibre is being optimized to further limit cobalt loss/recycle.

Other fibres investigated have promising characteristicsfor selective Zn and Ni removal from cobalt electrolyte.FIBAN® X-1 could also be used as a substitute for TP 207,

and for nickel-cobalt separation when these metals arepresent in solution in similar orders of concentration.

References

SHUNKEVICH, A.A., MARTSINKEVICH, R.V., MEDYAK, G.V., FILANCHUK, L.P., and

SOLDATOV, V.S. 2004. Comparison of fibrous carboxylic ion exchangers in

water treatment to remove heavy metal ions. Russian Journal of Applied

Chemistry, vol. 77, no. 2. pp. 249–253.

SOLDATOV, V., PAWłOWSKI, L., SHUNKEVICH, A., AND WASąG, H. 2004. New material

and technologies for environmental engineering. Part 1. Syntheses and

structures of ion exchange fibers. Drukarnia Liber Duo Kolor, Lublin.

127 pp.

THIRUNAVUKKARASU, O.S., VIRARAGHAVAN, T., SUBRAMANIAN, K.S., and TANJORE, S.

2002. Organic arsenic removal from drinking water. Urban Water, vol. 4.

pp. 415–421.

VATUTSINA, V.M.., SOLDATOV, V.S., SOKOLOVA, V.I., JOHANN, J., BISSEN, M., and

WEISSENBACHER, A. 2007. A new hybrid (polymer/inorganic) fibrous

sorbent for arsenic removal from drinking water. Reactive and Functional

Polymers, vol. 67, no. 3. pp. 184–201. ◆

390 MAY 2014 VOLUME 114 The Journal of The Southern African Institute of Mining and Metallurgy

Table IX

Comparison of TP 207 resin with Fiban X-1 fibre

Parameter Fibre Resin

Input

Feed flow rate m3/h 100

Cu loading g/L 10 20

Co loading g/L 10 45

Res time MTZ min 1.6 30

Cu in feed mg/L 50

Co in feed g/L 50000

# columns 2 3 4 2 3 4

Elution time h 1.5 10

Output

Upgrade 200 400

Transfer time h 6.8 3.4 2.3 210 105 70

Fibre flow rate m3/h 0.5 0.25

Fibre for MTZ m3 2.7 50.0

Fibre for elution m3 0.75 2.5

Fibre/column m3 3.4 1.7 1.1 52.5 26.3 17.5

Total fibre volume m3 6.8 5.1 4.6 105.0 78.8 70.0

No. of transfers/year 1218 2436 3654 40 79 119

Time/year h 8322

Co loss/recycle

Co treated (based on 5 g/L delta) kg/a 4161

Co loaded per cycle kg/cycle 34 17 11 2363 1181 788

% 20

Co loss/recycle per cycle (of the Co loaded) kg/cycle 6.8 3.4 2.3 472.5 236.3 157.5

Co loss/recycle per annum kg/a 8322 18725

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Introduction

The increase in the international iron ore pricein recent years has forced Chinese steelcompanies to seek domestic sources of iron orein order to reduce costs. The availability oflarge amounts of phosphorus- andmanganese-rich ores in the Three GorgesReservoir region has made the development ofa characteristic metallurgical technology inlocal steel plants possible (Wang and Dong,2009). In the conventional steelmakingprocess, one converter must fulfill severalfunctions, such as dephosphorization,decarburization, and raising temperature. It isunreasonable and uneconomical to decreasethe phosphorus content using only oneconverter when high-phosphorus molten ironis smelted. Separation of the decarburizationprocess from dephosphorization (using De-Pand De-C converters) in the duplex meltingprocess is advantageous, as it allows for the

use of phosphorus-rich iron ore and relativelylower amounts of slag, as well as directalloying with manganese ore in the converter.Consequently, iron procurement costs andmetal losses are reduced to a considerableextent. For these advantages to be translatedinto commercial benefits, the dephospho-rization converter should be fully exploited toproduce low-phosphorus, high-carbon, high-temperature semi-steel, and the semi-steelsmelting process should be further optimizedby manganese ore alloying and by improvingthe manganese yield. A number of studieshave been carried out on the improvement ofthe manganese yield and its impact on theconverter, both in China and abroad (Suito andInoue, 1995; Gao, Zhao, and Xing, 2011;Kaneko et al., 1993; Min and Fruehan, 1992;Lv et al., 2010; Soifer, 1958).

In this study, the main factors affecting themanganese yield are systematicallyinvestigated by thermodynamic analysis andindustrial tests, and the relationship betweenmanganese alloying and dephosphorization inthe converter is discussed in detail. Finally, thefinal slag composition and control ranges forconverter steelmaking are proposed.

Thermodynamic calculations

The reaction of manganese ore during thealloying process in the converter is describedas follows for a given slag system (Gao, Zhao,and Xing, 2011; Kaneko et al., 1993; Huang,X.H. 2008).

[1]

Thermodynamic analysis and experimentalstudy of manganese ore alloy anddephosphorization in converter steelmakingby G. Chen* and S. He*

SynopsisIn this study, the effects of slag compositions, slag amount, temperature,and carbon content of steel on the manganese and phosphorus distributionratios during converter steelmaking were analysed using the classicalregular solution theory, and industrial tests were performed using two 80 ttop-and-bottom combined blown converters (duplex melting process). Theresults indicate that the slag amount, temperature, and carbon content insteel are the main factors affecting the manganese yield when converterslag compositions remain constant. The FeO content of the slag has astrong impact on the manganese distribution ratio, while the slag basicityand MgO content have no obvious effect. The calculations and experi-mental results show that the phosphorus distribution ratio increasessharply with increasing slag basicity R, but then decreases with theincrease of MgO and MnO contents in the slag. The final slag in convertersteelmaking should have the following characteristics: 3.5 < R < 4.5, 15% <(FeO) < 20%, and 6% < (MgO) < 8%. The slag amount should be controlledappropriately at the same time. The results of this investigation would beuseful in deciding on the application of manganese ore in alloying andidentifying the slagging regime in converter steelmaking.

Keywordsslag compositions, distribution ratios, classical regular solution theory,slagging regime.

* College of Materials Science and Engineering,Chongqing University, Chongqing, China.

© The Southern African Institute of Mining andMetallurgy, 2014. ISSN 2225-6253. Paper receivedMar. 2013; revised paper received Feb. 2014.

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Thermodynamic analysis and experimental study of manganese ore alloying

[2]

The equilibrium constant of the above reaction is

[3]

(Yang and Cao). The effect of [C] content of steel on themanganese distribution ratio (LMn = (%Mn)/[%Mn]) is givenby Equation [1] at PCO = 1, f[C] = 1, and f[Mn] = 1. The activityof (MnO), a[Mn], can be obtained by using the regular ionicsolution model (Huang, 2008).

[4]

[5]

[6]

[7]

where x(i) is the mole fraction of positive ion i.The (FeO) content plays a critical role in the control of

manganese ore reduction during the later stage of blowing inthe converter (Suito and Inoue, 1995; Gao, Zhao, and Xing,2011; Morales and Fruehan, 1997; Takaoka et al., 1993).This reaction is shown in Equation [8].

[8]

[9]

Equation [10] is derived from Huang (2008).

[10]

The effects of slag basicity R = (%CaO)/(%SiO2) and(FeO) and (MgO) contents of the slag on LMn are studiedthrough the reaction in Equation [8] at f[Mn] = 1. Similarly,the activity coefficients of (Fe2+), γ(Fe2+) and (Mn2+), γ(Mn2+)can be obtained from Equations [4] and [5] respectively.

Moreover, the effects of temperature on LMn are calculatedseparately by Equations [1] and [8], and the carbon contentin Equation [1] is set as 0.08%.

The major dephosphorization reaction between moltensteel and slag in the converter is described by Equation [11](Basu, 2007; Ikeda and Matsuo, 1982).

[11]

In order to calculate the phosphorus distribution ratio Lp = (%P)/[%P], the activity of the complex ion ofphosphorous and oxygen can be expressed by the simplifiedreaction Huang (2008) :

[12]

[13]

[14]

[15]

where the KP value is 0.0234 (Huang (2008). The values ofx(P5+), x(Fe2+), γ(Fe2+), and γ(P5+) can also be obtained from theregular ionic solution model (Huang (2008).

Figures 1 and 2 show that R has no obvious effect onLMn. In addition, compared with (FeO), (MgO) has a lessobvious effect on the change in LMn. LMn tends to increasewith an increase in (FeO) but decrease with an increase in(MgO). Therefore, LMn is more strongly affected by the (FeO)content rather than R or (MgO) content. Figure 3 shows thevariation of the calculated activity coefficient of (MnO) as afunction of (FeO) content. When (FeO) content in slagincreases from 15% to 35%, the activity coefficient of (MnO)decreases from 1.71 to 1.29, which can also be observed inthe results of Jung et al. (1993), Jung (2003), and Suito andInoue (1984). Obviously, the results obtained in this workare in agreement with the data from the literature. Inaddition, as can be seen in Figure 4, the activity of FeO inslag increases strongly with an increase of (FeO) in slag overthe calculated concentration range, which is in good

392 MAY 2014 VOLUME 114 The Journal of The Southern African Institute of Mining and Metallurgy

Figure 1—Effect of FeO content of slag on LMn

Figure 2—Effect of MgO content of slag on LMn

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accordance with the results of other investigators (Moralesand Fruehan, 1997; Huh and Jung, 1996; Sobandi,Katayama, and Momon, 2002). The distribution ratio LMnthus increases with the increase of (FeO), resulting in a lowreduction efficiency of (MnO) when (the FeO) content isincreased, as shown by Equation [9]. Hence, in order toimprove the manganese yield in the converter, the (FeO)content of the final slag should be maintained at a low levelin the alloying process performed using manganese ore. Thishas been affirmed by previous experiments (Suito and Inoue,1995; Morales and Fruehan, 1997; Suito and Inoue, 1984;Jung, Rhee, and Min, 2002). However, the dephosphorizationin the converter requires a high (FeO) content. As illustratedin Figure 5, the increase in (FeO) content initially enhancesLP, but the trend is reversed beyond a certain level, and theoptimal (FeO) content decreases with increasing R. Generally,LP has already reached the maximum level when the (FeO)content approaches 20% at a relatively high R value. Theseresults are similar to those assessed thermodynamically andexperimentally by previous researchers (Basu, 2007; Ikedaand Matsuo, 1982; Sobandi, Katayama, and Momon, 2002;Suito and Inoue, 1995). Therefore, the (FeO) content couldbe controlled between 15% and 20% to achieve a highmanganese yield.

It is well known that slag with higher R and higher (FeO)content is required in the later stage of the smelting processfor dephosphorization (Basu, 2007; Jeong et al., 2009;Nozaki et al., 1983). Consequently, the conditions used for

converter steelmaking are suited for dephosphorization. Asshown in Figures 5–7, it is obvious that the dephospho-rization effect increases sharply as R is increased, which isbelieved to dramatically reduce the activity coefficient of(P2O5) in slag, as recognized (Basu, 2007; Sobandi,Katayama, and Momon, 2002; Suito and Inoue, 1995;Turkdogan, 2000; Suito and Inoue, 1982, 1984; Nakamura,Tsukihashi, and San, 1993). A higher R is thusindispensable for dephosphorization in the converter.Conversely, an excessively high R will worsen the kineticconditions for manganese ore reduction and dephospho-rization. Simultaneously, taking previous studies (Lv et al.,

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Figure 3—Effect of FeO content of slag on r(MnO)

Figure 4—Effect of FeO content of slag on a(FeO)

Figure 5—Effect of FeO content of slag on LP

Figure 6—Effect of MnO content of slag on LP

Figure 7—Effect of MgO content of slag on LP

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Thermodynamic analysis and experimental study of manganese ore alloying

2010; Basu, 2007; Tabata et al., 1990) into consideration, Rshould be high when the (FeO) content is relative low and, ingeneral, the R of the final slag should be controlled between3.5 and 4.5 to achieve a higher degree of dephosphorization.

Furthermore, LP clearly decreases with an increase in the(MgO) and (MnO) contents of the slag at a given R (Figures 6and 7). Also, the influences of (MgO) and (MnO) contents onLP are gradually enhanced with the increase of R from 2 to 4;the effects of R on LP become progressively weaker withincreasing (MgO) and (MnO) contents. Specifically,calculations show that increasing the (MgO) and (MnO)contents of the slag can increase the activity of (P2O5), whichis detrimental for LP, as confirmed by Figure 8. Separately,the dephosphorization effects weaken significantly when theslag has a high (MnO) content, as has been reported by manyinvestigators (Suito and Inoue, 1995; Mukherjee andChatterjee, 1996; Simeonov and Sano, 1985) and observedby previous researchers, all of whom have advised againstthe addition of manganese oxide to converter slag. As aresult, manganese ore alloying during converter smeltingwould become advantageous when low-phosphorus hot metalis used as raw material. Also, the experiments of Halder andco-workers demonstrated conclusively that 2CaO·SiO2 existsunder conditions of higher slag basicity, lower steel tappingtemperatures, and higher phosphorus contents of the hotmetal, which comprised the majority of the solid part of theslag and also had greater solubility for phosphorus than theliquid part of the slag (Deo et al., 2004) . In addition, Suito,Inoue, and Takada (1981) also proved that slag containing2CaO·SiO2 had a higher LP in the MgO-saturated slag of thesystem CaO-MgO-FeOx-SiO2, and the same result was alsoobtained in the CaO-CaF2-SiO2 system by Muraki,Fukushima, and Sano, (1985). However, the increase of(MgO) content in slag could result in a reduction in both thesize of the 2CaO·SiO2 grains and the dissolution ofphosphorus in 2CaO·SiO2. Thus, dephosphorization wasgreatly hindered (Deo et al., 2004). Therefore, taking onlydephosphorization into consideration, the lower the MgOcontent the better. However, since (MgO) plays an importantrole in protecting the furnace lining, it is imperative that theappropriate amount of (MgO) should be present in the slag.We can manipulate this relationship by using slag-splashingprotection technology for the converter. In general, the (MgO)content should be controlled between 6% and 8%.

From Equations [1] and [8], the values of LMn arecalculated separately as a function of steel temperature inFigure 9. The results of the authors and of Jung et al. (1993)and Jung, Rhee, and Min (2002) are plotted for comparison.It is seen that most values of log(LMn) in this work areslightly higher compared with previous results. This may be aresult of the slightly different components of the slag. Ingeneral, log(LMn) decreases linearly with increasingtemperature. Equations [3] and [10] clearly indicate that thereaction in Equation [1] is endothermic and Equation [8] isexothermic. The equilibrium [Mn] content in steel is thereforeexpected to increase with increasing temperature. On theother hand, many studies indicate that the dissolutionreaction (MnO(s) = MnO(slag) of (MnO) in slags isendothermic, so that (MnO) dissolution in slag increases withtemperature (Jung et al., 1993; Suito and Inoue, 1984;Simeonov and Sano, 1985; Suito and Inoue, 1984.Accordingly, the activity of (MnO) will increase withincreasing mole fraction of (MnO) in the slag, and theseresults are widely accepted in previous studies (Ding andEric, 2005). Thus, the equilibrium [Mn] content increasesnaturally.

Figure 10 shows that the final [C] carbon content of thesteel has a great influence on LMn. With a decrease in thefinal carbon content [C] in steel, LMn increases distinctly, andthis has been experimentally confirmed by previous investi-gators (Kaneko et al., 1993; Yang and Cao, 2009; Tabata et

394 MAY 2014 VOLUME 114 The Journal of The Southern African Institute of Mining and Metallurgy

Figure 8—Effect of MgO and MnO contents of slag on log(a(P2O5)

Figure 9—Effect of temperature on log(LMn)

Figure 10—Effect of [C] of steel on LMn

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al., 1990; Matsuo, Fukagawa, and Ikeda, 1990). Clearly, the[C] in the molten steel can accelerate the reduction of (MnO)and improve the manganese yield. The [C] content of steelacts as a heat source and as well as a reducing agent, andcarbon is thus consumed in large amounts. At the same time,carbon reacts with oxygen to form CO gas, which canenhance the fluidity of the entire slag system, which drivesthe reactions towards equilibrium (Keum et al., 2007).

The effect of the amount of slag on the manganese yield ((Ws * ([Mn]f - [Mn]s))/(0.3 * WMn), where [Mn]f and [Mn]sare the [Mn] content in the final steel and semi-steel respec-tively; Ws and WMn are the yield of final steel (tons) andmanganese ore charged (tons), respectively) is shown inFigure 11. While the slag amount affects the manganeseyield, it does not influence LMn. The manganese ore yielddecreases sharply with an increase in the slag amount, due tothe fact that the (MnO) content in the slag will decrease andthe balanced [Mn] content in steel also will decline. As can beseen, since the manganese yield is less than 35% when theslag amount exceeds 60 kg/t, the slag amount must bemaintained at a value less than 40 kg/t to ensure a high Mnyield (>45%). Similar results have been reported by otherresearchers (Kaneko et al., 1993; Tabata et al., 1990;Mukherjee and Chatterjee, 1996).

In conclusion, steelmaking by a process that involvesmanganese ore alloying and the use of low quantities of slagis an effective measure for lowering the consumption of rawmaterials and raising the manganese yield, which are thetypical advantages of the De-P/De-C steelmaking process.

Industrial tests

Industrial process description

To verify the results of the thermodynamic calculations, wecarried out industrial tests using two 80 t converters at asteel plant in China. The blast furnace, De-P converter, De-Cconverter, refining, and continuous casting route has beenbeen adopted. The experimental conditions used for theindustrial-scale tests are shown in Table I. Tables II and IIIshow the compositions of the molten iron and manganese oreused in the tests, respectively; the additions of manganeseore and compositions of semi-steel, final steel, and final slagare shown in Table IV. In this exploratory study, only a small

amount of manganese ore (0~810 kg) was added to ascertainthe factors affecting alloying, which can provide referencedata for further industrial-scale production.

Results and analysis

The values of LMn obtained from the results are shownagainst (FeO) contents in Figure 12, where the R valuesrange from 3.0 to 5.5. The values of LMn change irregularly

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Table I

Operating conditions of De-P/ De-C converters

De-P converter De-C converter

Molten iron/semi-steel 77~83 t 84~93 tscrap 7~10 t 0

LimeSlagging agent Lime

Slagging elements Returned slag Slagging agentFluorite Manganese oreBauxite

Flow rate of top gas (O2) 10000–13000 Nm3/h 16000–16500 Nm3/h

Flow rate of bottom gas (N2/Ar) 300–600 Nm3/h 300–600 Nm3/h

Blowing time 7–9 min 10~12 min

Table II

Compositions of molten iron (mass %)

C Si Mn P S

4.0–4.2 0.4–0.6 0.2–0.4 0.11–0.45 0.036–0.06

Table III

Compositions of manganese ore (mass, %)

CaO MgO SiO2 Al2O3 Mn TFe S P

8–12 4–8 10–15 2–5 28–32 1–3 0.08–0.15 0.01–0.02

Figure 11—Effect of slag amount on manganese yield

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Thermodynamic analysis and experimental study of manganese ore alloying

with the increase of R at a given (FeO) content, and LMnincreases with an increase in (FeO) content at a given R,which shows a relatively consistent tendency with thermody-namics (Figure 1). This is also confirmed in Figures 13 and14. LMn increases as the FeO content increases when [C]content and temperature remain constant. In addition, LMnshows a rapid increase as the [C] in the steel decreases at thesame FeO content (Figure 13), and the temperature has asimilar impact on LMn, as illustrated in Figure 14. In Figure13, the experimental values of LMn locate between 25 and 45,indicating that these values agree well with the calculatedvalues (which are between 20 and 47) according to Equation[1] (Figure 10), under the condition of 0.04–0.09% [C]carbon content. However, these experimental values turn outto be somewhat less than the calculated values according toEquation [8] (Figure 10), probably because the reduction ofMnO in slag by carbon [C] in steel leads to a significantincrease of the [Mn] content of the steel, and in turn, theoxidization of [Mn] by FeO in the slag may not reach athermodynamic equilibrium. As shown in Figures 12–14, asexpected, low LMn values result when the (FeO) content isless than 25%, the [C] content exceeds 0.06%, and thetemperature is higher than 1920K. As is well known, theincrease in temperature depends on the oxidization of carbonduring blowing in the BOF, and there is a strong negativecorrelation between [C] carbon content and temperature. Onthe other hand, the oxidation of [C] carbon will be depressedwhile the oxidation of [Fe] will be facilitated when the [C]content is less than about 0.05% (Huang, 2008), which willcause excessive (FeO) content in slag. As demonstrated inFigure 15, when the final carbon content is less than 0.06%,the majority of the corresponding (FeO) content in slag

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Table IV

Manganese ore additions and compositions of semi-steel, final steel, and final slag (mass %)

Semi-steel, % Mn Ore, Final steel, % Temp., Final slag, %

C Mn P S kg C Mn P S °C SiO2 Al2O3 CaO MgO TFe P2O5 MnO FeO

1 2.8 0.11 0.13 0.057 0 0.023 0.01 0.011 0.04 1612 8.34 2.05 40.66 8.61 24.02 2.87 2.22 21.72 2.8 0.1 0.163 0.046 0 0.047 0.02 0.016 0.043 1638 16.9 2.57 46.09 8.01 18.43 3.61 2.22 16.883 2.7 0.1 0.15 0.041 0 0.044 0.02 0.008 0.036 1608 8.12 2.57 43.27 7.4 22.9 3.34 1.93 21.24 2.84 0.08 0.14 0.042 0 0.042 0.02 0.01 0.047 1624 8.08 2.05 44.82 6.18 21.5 3.9 1.93 20.195 2.82 0.04 0.128 0.046 0 0.037 0.01 0.01 0.042 1630 8.51 1.54 43.41 6.08 23.32 3.52 2.22 21.846 2.75 0.07 0.1 0.048 0 0.063 0.02 0.013 0.048 1627 9.52 2.05 42.15 7.9 20.94 3.57 1.93 20.197 2.73 0.1 0.12 0.048 0 0.067 0.07 0.01 0.045 1638 8.98 2.05 38.34 8.72 23.04 3.55 3.86 20.988 2.8 0.07 0.06 0.05 0 0.15 0.1 0.018 0.06 1666 12.2 1.94 38.76 12.67 20.25 2.52 2.53 19.269 2.94 0.08 0.1 0.055 0 0.046 0.06 0.021 0.057 1642 10.58 2.62 44.26 8.82 16.34 1.76 2.97 17.1710 2.87 0.07 0.077 0.053 0 0.053 0.05 0.018 0.052 1645 10.02 3.05 40.74 11.15 18.71 1.76 2.82 18.7511 2.7 0.08 0.088 0.048 70 0.078 0.09 0.014 0.048 1632 10.24 3.19 41.72 10.34 18.29 1.51 4.46 17.5312 2.88 0.09 0.071 0.045 400 0.095 0.12 0.015 0.048 1629 10.78 2.9 41.16 9.12 17.03 1.26 5.94 17.6813 2.78 0.09 0.08 0.043 480 0.117 0.16 0.015 0.048 1640 11.24 3.2 40.45 8.82 17.87 1.76 6.54 17.8914 2.85 0.08 0.077 0.049 490 0.048 0.11 0.021 0.058 1646 12.62 3.07 43.7 11.25 14.92 2.01 4.31 14.3715 2.86 0.1 0.196 0.046 510 0.07 0.06 0.019 0.039 1624 8.84 1.05 41.3 10.95 22.06 2.89 3.4 20.5516 2.73 0.13 0.152 0.048 510 0.052 0.1 0.015 0.039 1655 8.12 2.05 38.48 7.9 22.06 3.58 4.75 21.7717 3.01 0.14 0.18 0.051 530 0.083 0.14 0.02 0.044 1653 9.38 2.57 40.88 8.41 17.59 4.07 4.75 17.2418 2.88 0.11 0.138 0.047 600 0.062 0.11 0.019 0.051 1664 8.14 2.57 34.39 8.92 24.02 3.57 5.2 22.4219 2.87 0.1 0.101 0.049 600 0.062 0.14 0.021 0.058 1634 11.7 2.88 42.01 8.82 17.45 2.01 5.5 16.9620 2.86 0.12 0.14 0.045 630 0.08 0.18 0.023 0.044 1645 9.86 1 42.43 7.8 17.45 2.18 6.06 15.4521 2.74 0.1 0.071 0.046 670 0.082 0.13 0.01 0.038 1622 10.46 2.64 40.45 9.12 18.71 1.51 6.13 19.422 2.9 0.16 0.146 0.038 680 0.06 0.16 0.019 0.04 1629 9.12 1 38.9 10.54 18.29 4.04 6.65 16.3123 2.6 0.18 0.201 0.044 690 0.04 0.11 0.018 0.038 1653 8.42 1.8 39.33 10.84 19.97 3.43 5.02 16.8124 2.44 0.15 0.18 0.045 700 0.05 0.1 0.015 0.038 1630 8 1.05 40.6 8.82 23.04 3.46 5.62 19.7625 2.78 0.06 0.05 0.052 710 0.056 0.1 0.012 0.048 1621 9.86 2.01 37.49 9.93 21.78 1.26 6.09 21.1226 2.88 0.13 0.145 0.048 760 0.07 0.16 0.018 0.04 1656 9.46 1.04 38.9 7.09 18.99 4.13 6.21 15.5927 2.8 0.18 0.22 0.04 810 0.07 0.17 0.02 0.037 1660 8.5 1.04 42.15 8.61 17.87 4.41 6.35 15.59

Figure 12—Effect of FeO content of slag on LMn

Figure 13—Effect of [C] on LMn

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exceeds 28%. However, a high [C] carbon content, hightemperature, and low (FeO) content are required in order toattain a low LMn. More intensive research is required todetermine the appropriate amount of carbon powder thatshould be added to the slag when the [C] carbon content ofthe semi-steel is insufficient to provide adequate heat andreduce the manganese ore to a large extent.

The influence of the slag amount on the manganese yieldis shown in Figure 16. The yield of the manganese oredecreases sharply with an increasing slag amount at similarvalues of LMn, and it falls below 25% when the slag amountis 60 kg/t. Generally, the manganese yields in the industrial

tests are fairly low (usually below 20%); this is the inevitableconsequence of the presence of a large amount of slag. Themain reason is that the iron ore used by the aforementionedsteel plant is rich in phosphorus, and the [P] content of thesemi-steel charged in the De-C furnace is still high and alarge amount of time and energy will lost in dephospho-rization. Consequently, a large amount of slag is generated.At the same time, a large amount of slag is required for deepdephosphorization, which in turn has an enormous negativeinfluence on manganese yield.

The influences of R and the (FeO), (MnO), and (MgO)contents of the slag on LP are shown in Figures 17–19. Fromthe experimental results, it is obvious that LP increases with

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The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 MAY 2014 397 ▲

Figure 14—Effect of temperature on LMn

Figure 15—Relation between of FeO content and slag amount

Figure 16—Effect of slag amount on manganese yield

Figure 17—Effect of FeO content on LP

Figure 18—Effect of MnO content on LP

Figure 19—Effect of MgO content on LP

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Thermodynamic analysis and experimental study of manganese ore alloying

increasing R from 3.0 to 5.5, regardless of whether the (FeO),(MnO), or (MgO) contents are held constant. It is also foundthat the higher the (FeO) content of slag, the more the LPincrease at a given R. However, the optimum (FeO) content isnot observed. This is corroborated by the theoreticalcalculations and has been reported in previous investigations(Basu, 2007; Ikeda and Matsuo, 1982; Sobandi, Katayama,and Momon, 2002; Suito and Inoue, 1995). It is consideredthat this phenomenon is due to the increase in the amount ofslag, even if the LP is decreased slightly with an (FeO)content higher than the optimum. As shown in Figure 15, theresults confirm a strong positive correlation between (FeO)content and slag amount. In addition, as expected, thedephosphorization effect weakens markedly with increasing(MgO) and (MnO) contents in slag when R is greater than 4.However, when R is less than 4, no regular relationship hasbeen found. Furthermore, the impact of R on LP is weakenedwhen the (MgO) and (MnO) contents are high, respectively.

From the data pertaining to the smelting test conducted in27 heats, we can conclude that the average final [P] contentof the steel and the average degree of dephosphorization are0.016% (0.008–0.023%), and 87.4%, respectively, when theaverage [P] content of the semi-steel charged in the De-Cfurnace is 0.126% (0.05–0.22%). Since a larger amount ofslag (on average, about 68.3 kg/t) with high (FeO) content(mean, 25.6%) is used in the converter to decrease the [P]content to the required steel grade, a relatively good degree ofdephosphorization is achieved, but the manganese yield isonly 17.2% on average. Hence, for manganese ore alloying tobe beneficial, the process in the De-P furnace should beoptimized to decrease the [P] content of the semi-steelfurther.

Obviously, the data obtained in the test work is basicallyin good agreement with the results of the thermodynamiccalculations. This, in turn, shows that the choice of themethod of calculation is reasonable.

On the basis of the thermodynamic calculations and theindustrial test results, it is concluded that coordinated controlbetween the dephosphorization ability and manganese orealloying technology in the De-C converter should beconsidered carefully. The characteristics of the final slag forconverter steelmaking should be controlled in the followingranges: 3.5 < R < 4.5, 15% < (FeO) <20%, and 6% < (MgO) <8%.

Conclusions

A thermodynamic analysis and industrial tests of manganeseore alloying and dephosphorization in converter steelmakingwere carried out. The conclusions can be drawn as follows.

(1) The main factors affecting the alloying processperformed using manganese ore in the converter arethe slag amount, temperature, and the [C] content ofthe steel with a given slag system

(2) The (FeO) content of the slag has an enormousimpact on LMn but shows no clear relationship withthe slag basicity or the (MgO) content of the slag

(3) The LP increases sharply with increasing slag basicity,but weakens with increasing (MgO) and (MnO)contents in the slag

(4) The characteristics of the final slag for convertersteelmaking should be controlled in the following

ranges: 3.5 < R < 4.5, 15% < (FeO) < 20%, and 6% <(MgO) < 8%. The slag amount should be controlledappropriately at the same time.

Acknowledgements

The authors greatly appreciate the funding support fromChongqing Science and Technology Key Project(CSTC2008AB4018)

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SUITO, H., INOUE, R., and TAKADA, M. 1981. Phosphorus distribution betweenliquid iron and MgO saturated slags of the system CaO-MgO-FeOx-SiO2.Transactions of the Iron and Steel Institute of Japan, vol. 21. pp. 250–259.

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Thermodynamic analysis and experimental study of manganese ore alloying

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Introduction

Hydrometallurgy has become one of the mosteconomical processes for recovering metalsfrom low-grade matte. Matte is a mixture ofmetal sulphides and precious metals producedfrom the smelting of sulphide ores. Acidleaching unselectively separates the basemetals from the precious metals, producing aliquor rich in base metals and leaving a solidresidue rich in precious metals. Solventextraction (SX) and electrowinning (EW) canbe used to recover marketable base metalproducts from the liquor. Gaseous reductioncan also be employed to produce base metalpowders (Agrawal et al., 2006).

There are different types of leachingtechniques and each type has requirements interms of particle size, leaching time, agitationrate, etc. to ensure an efficient process andhigh percentage metal extraction. Agitationleaching in batch vessels operated underatmospheric or high pressure has been widelyused commercially for processing Ni, Cu, andCo sulphide concentrates. Leaching of these

mattes has been conducted using ammonia oracid solutions following the developmentalwork of Sherritt Gordon in the period1950–1969 (Forward and Mackiw, 1955;Pearce et al., 1960). The choice between acidand ammonia leaching has been based on localconditions or the composition of the matte.Acid leaching has been used for mattes withsubstantial cobalt contents (above 3%), whileammonia leaching has been widely used forNi-Cu mattes with low cobalt contents (Pearceet al., 1960). However, with furtherdevelopments in acid pressure leachingtechnology by Sherritt Gordon in the 1960smost Ni-Cu mattes have been treated usingacid leaching. Impala Platinum commissionedthe first commercial acid pressure leachingprocess for treating Ni-Cu mattes containingplatinum group metals (PGMs) in horizontalautoclaves in 1969, and since then acidleaching has been widely used in the SouthAfrican platinum industry (Plasket andRomanchuk, 1978).

Acid leaching can be conducted in asulphate or chloride medium. The sulphatemedium has been widely used, especially inSouth Africa, because the process equipmentused is more adaptable to sulphate rather thanchloride systems and there is no possibility ofplatinum dissolution as in chloride systems(Brugman and Kerfoot, 1986). However,chloride systems have the followingadvantages; most metal chlorides are moresoluble than sulphate salts, leaching can beperformed at moderate temperatures, and theoxidation process yields elemental sulphur,which is environmentally more acceptable thansulphate from sulphuric acid leaching (Park etal., 2006). Given these advantages and thedevelopment of corrosion-resistant material,

Atmospheric oxidative and non-oxidativeleaching of Ni-Cu matte by acidified ferricchloride solutionby L.M. Sekhukhune*, F. Ntuli*, and E. Muzenda*

SynopsisThe atmospheric leaching of copper-bearing matte by acidic ferric chloridesolution was studied at the laboratory scale. The aim was to achievemaximum copper and nickel recovery by investigating the mechanisms ofleaching, as well as identifying the effect of temperature, and concentrationof ferric chloride and oxygen. Djurleite (Cu1.96S), hazelwoodite (Ni3S2),and Ni alloy were the primary phases detected in the matte. The quanti-tative composition of the matte was Cu 31%, Ni 50%, S 13%. Fe and Coconstituted 2%, with platinum group metals (PGMs) accounting for 0.5%. Amaximum nickel extraction of 98% was achieved using two-stage oxidativeleaching at 90°C and 11 g/L Fe3+ as compared to 65% under non-oxidativeconditions. A copper extraction of 99% was achieved in the first 45 minutesusing two-stage non-oxidative leaching, and copper was recovered fromsolution by cementation. Three processes took place simultaneouslythroughout the leaching process, namely: dissolution, cementation/metathesis, and oxidation. The leaching process was found to be diffusion-controlled.

KeywordsNi-Cu matte, acid leaching, cementation, ferric chloride, leachingmechanism.

* Department of Chemical Engineering, University ofJohannesburg, Johannesburg, South Africa.

© The Southern African Institute of Mining andMetallurgy, 2014. ISSN 2225-6253. Paper receivedJun. 2013; revised paper received Nov. 2013.

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Atmospheric oxidative and non-oxidative leaching of Ni-Cu matte

the leaching behaviour of typical South African PGM-containing Ni-Cu mattes in chloride media is worth investi-gating.

This study investigated the atmospheric leaching of PGM-containing Ni-Cu matte by acidic ferric chloride solution. Theaim of this project was to achieve maximum Ni and Curecovery from the Ni-Cu enriched matte by investigating themechanisms of leaching, as well as identifying the effect ofprocessing variables; namely, temperature and concentrationof lixiviant under oxidative and non-oxidative conditions.

The sulphide chemistry is very complex due to the factthat sulphide concentrates usually consist of highlyintergrown sulphide minerals. This is one of the reasons whythere is a lack of understanding and knowledge of themechanism of leaching. These processes are thus often notoperated at optimum conditions (Rademan, 1999).Furthermore, the morphology and mineralogy of complexsulphide ores differs in nature, thus there is a need for aseparate study for each type of matte. Leaching of Ni-Cumattes is normally conducted in two stages, with the firststage aimed at removing most of the nickel and cobalt andprecipitating the Cu by cementation. The second stage isnormally aimed at removal of Cu and the remaining basemetals to produce a platinum-rich residue (Lamya andLorenzen, 2009). Although the entire leaching process can beachieved in a single stage, multiple stage leaching is highlyselective, enabling the recovery of the base metals insubsequent processes (Hofirek and Kerfoot, 1992). Inaddition, higher metal extraction levels are achieved. Two-stage leaching studies were therefore conducted in this work.

Methodology

The parameters investigated were temperature and concen-tration of lixiviant. Single- and two-stage leaching processesunder oxidative and non-oxidative conditions wereconducted. The effect of temperature was investigated usingsingle-stage non-oxidative leaching at temperatures of 50,70, and 90°C ± 5°C for each experimental run, with aconstant Fe3+ concentration of 5 g/L. Non-oxidative two-stage leaching was used to investigate the effect of ferricchloride concentration (lixiviant), with varying concen-trations of 5, 8, and 11 g/L Fe3+ and a constant temperatureof 90°C for each experimental run. The ferric chloride concen-tration and temperature were the same for both the first- andsecond -stage leaching. Based on the results of thetemperature investigations, a temperature of 90 ± 5°C waschosen for further investigations.

Apparatus and reagents

The non-oxidative atmospheric leaching experiments wereperformed in a 5 L glass vessel fitted with four baffles, avariable-speed overhead stirrer with a flat blade turbine-typeimpeller, and a heating element with a temperature controller.The cover of the vessel had two ports for holding apH/temperature probe and for taking samples. These portsallowed for minimal air ingression into the reactor during theexperiment.

The matte was obtained from Impala Platinum Refinery,and ferric chloride was used as the leaching agent.Hydrochloric acid was used to control the pH below 3.

Characterization

The matte before and after leaching was characterized by X-ray diffractometry (XRD) with a Cu Kα radiation source(Xpert Phillips) to determine the mineralogical phases presentin the matte, and X-ray fluorescence (XRF) (PW 2540 VRCSample changer) to determine the elemental composition ofthe matte. Scanning electron microscopy (SEM) (JOEL JSM5600) coupled with energy dispersive spectroscopy (EDS)was used to confirm the mineralogical phases in the matteand capture the particle morphology. A laser diffractiontechnique (Malvern Mastersizer 2000) coupled to a liquiddispersing unit (Hydro 2000G) was used to determine theparticle size distribution (PSD) of the matte before and afterleaching. Ni and Cu concentrations in the leach liquor weremeasured by inductively coupled plasma-optical emissionspectrometry (ICP-OES (Spectro Arcos Fsh12) and atomicabsorption spectroscopy (AAS) (Thermo Scientific ICE 3000instrument).

Experimental procedure

Leaching solution (4 L) was heated in the reactor withagitation until the required temperature was reached; 40 g ofmatte of particle size -105+75 μm was then added into thevessel. The stirrer was set to the required constant speed of600 r/min and timing of the experiment started. HCl solution(1.5 M) was added stepwise to the mixture to keep the pHbelow 3. Temperature and pH at t=0 were recorded andthereafter monitored every 15 minutes. Liquor samples of 25 mL were taken after every 15 minutes for the 200-minuteduration of leaching; which was considered a sufficient timeto yield a high percentage extraction. The liquor samples werefiltered and the filtrate was kept in sample bottles forchemical analysis. At the end of each experiment, the pulpwas allowed to cool to room temperature and then filtered.The filter cake was washed with demineralized water anddried overnight. The residue samples were also kept forchemical analysis.

For non-oxidative two-stage leaching, the generalprocedure as outlined above was followed. However, after200 minutes of first-stage leaching, the leaching solution wasdecanted from the reactor and then filtered. Thereafter 4 L offresh leaching solution of the desired concentration was thenadded to the reactor and the filtered cake returned to thereactor to commence second-stage leaching. The concen-trations of Fe3+ used in the non-oxidative two-stage leachingtests were 5, 8, and 11 g/L Fe3+. The concentration was keptconstant for both first- and second-stage leaching.

For the two-stage oxidative leaching the experimentalprocedure was similar to that of the non-oxidative two-stageleaching, with the exception of the introduction of oxygen,and a fixed Fe3+ concentration of 11 g/L Fe3+ and a constanttemperature at 90 ± 5°C were used based on the findings ofthe previous experiments. In the first stage, fresh matte wascontacted with fresh lixiviant for 200 minutes, and in thesecond stage, solid residue from the first stage was contactedwith fresh solution and further leached for 200 minutes.Oxygen below 5 kPa was spurged into the solutionthroughout the 200-minute experimental run. The oxygenpressure was kept constant.

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Results and discussion

Matte characterization before leaching reaction

The XRD pattern of the matte before leaching is shown inFigure 1. The major phases detected in the matte weredjurleite (Cu1.96S) and hazelwoodite (Ni3S2). Hazelwooditehad the highest peak intensity of 100% at 31.0532 angleposition, and djurleite had the highest relative intensity of25% at 32.5175 angle position.

The phases identified by SEM-EDX were Ni alloy,hazelwoodite, djurleite, and some PGMs (Figure 2).

The elemental composition of the matte before leachingas determined by XRF is shown in Table I. Copper constituted31 mass % while nickel (i.e. Ni alloy and Ni3S2) was themajor element in the matte, representing nearly half of thematte. Since the matte is a sulphide, sulphur also constituteda significant amount (13 mass %). The remaining 2%consisted of Fe and Co, while PGMs like Pt and Pd accountedfor only 0.5%.

SEM of the matte showed that the particles wereirregularly shaped with a smooth outer surface layer i.e. nocracks and veins on the particles (Figure 3).

The PSD of the matte generated by a laser diffractiontechnique is shown in (Figure 4). The modal size of the matte was 138 μm before leaching. The volume distribution

(Figure 4a) was transformed into the number distribution(Figure 4b) and shows that the matte consisted of a largernumber of smaller sized particles, which however;contributed a small volume percentage.

The d(0.1) showed that 10% of the particles were smallerthan 86.27 μm before leaching, whereas the d(0.9)demonstrated that 90% of the particles were smaller than209.58 μm.

Atmospheric oxidative and non-oxidative leaching of Ni-Cu matte

403The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 MAY 2014 ▲

Table I

Elemental composition of the matte before leaching,mass %

Co Cu Fe Ni Pb Pd Pt S Se Si Other

0.4 31 1.4 50 0.081 0.21 0.27 13 0.09 1.2 2.35

Figure 2—SEM-EDX micrographs of the matte before leaching

Figure 1—XRD spectra of the matte before leaching

Figure 4a—Particle size distribution (by volume) of the matte beforeleaching

Figure 3—SEM images of the matte particle before leaching (magnifi-cation 100×)

PGMʼs

Hazelwoodite

Chalcocite

Ni Alloy

a = Ni3S2

b = Cu2S

0 10 20 30 40 50 60 70 80

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Atmospheric oxidative and non-oxidative leaching of Ni-Cu matte

Effect of leaching variables

Temperature

Experiments were conducted at temperatures of 50, 70, and90°C ± 5°C, at a constant concentration of 5 g/L Fe3+. Therecoveries of Ni and Cu as a function of leaching time areshown in Figure 5. Maximum percentage extractions of Ni fortemperatures of 50, 70, and 90°C were 18, 27, and 55%respectively. The highest Ni recovery was obtained at 90°C.

A temperature of 90°C was therefore chosen as theoptimum temperature in this study, since it is advantageousto leach at a temperature near the boiling point of ferricchloride of approximately 105°C (Dutrizac, 1992).Researchers such as Park et al. (2006), Rademan et al.(1999) and Dutrizac. (1992) have demonstrated thattemperature has a direct influence on leaching. The higherthe temperature, the higher the leaching rate, therefore thehigher the metal recovery. The results obtained in this studyare in agreement with this conclusion. The extraction ofcopper, however, decreased with increasing temperature(Figure 5b) due to cementation. At 50, 70, and 90°C themaximum percentage extractions of Cu were 14, 4, and 2%respectively. The cementation process was favoured by highertemperatures. Cu extraction was highest at 50°C because theNi extraction rate was the lowest at this temperature, thusallowing a small amount of copper to be leached. Cu wasrejected at this temperature from 14% to approximately 4% inthe leach liquor by the end of 200 minutes.

The elemental composition of the matte before and afterleaching at 90ºC is shown in Table II.

The copper content was 31% before leaching and 42%after leaching. This increase was due to the increased surfaceexposure of copper sulphide in the residue as a result ofleaching. The relative proportion of Ni in the matte decreasedby 14% after leaching (Table II). This seemed very low at thisstage, but was attributed to the fact that Ni alloy was thephase that leached the most. This finding was substantiatedby the SEM-EDS analysis after leaching at 90°C. Figure 6shows that hazelwoodite and copper sulphide are theremaining minerals in the matte after leaching, thus proving

that almost all the Ni alloy was extracted. This finding agreeswith that of Rademan et al. (1999), who stated that Ni alloywas the most reactive phase, followed by Ni3S2, and lastlyCu2S.

Micro-cracks on the particles can be seen in Figure 6.Micro-cracking enhances the leaching rate by increasing thesurface area available for reaction; thus increasing the rate ofdiffusion of the lixiviant and products. The metathesisprocess was also believed to be taking place simultaneouslywith cementation according to Equation [1].

Ni3S2 + 2Cu2+ → Cu2S + NiS + 2Ni2+ [1]

404 MAY 2014 VOLUME 114 The Journal of The Southern African Institute of Mining and Metallurgy

Table II

Quantitative elemental composition of the mattebefore and after single-stage non-oxidative leachingwith 5 g/L [Fe3+] at 90º C, mass%

Co Cu Fe Ni Pb Pd Pt S Other

Before 0.4 31 1.4 50 0.081 0.21 0.27 13 2.35 After 0.06 42 0.67 36.0 0.213 0.306 0.405 7.6 12.16

Figure 4b—Particle size distribution (by number) of the matte beforeleaching

Figure 5b—Percentage extraction of Cu after single-stage non-oxidative leaching with 5 g/L [Fe3+]

Figure 5a—Percentage extraction of Ni after single-stage non-oxidativeleaching with 5 g/L [Fe3+]

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However, the XRD analysis (Figure 7) shows that nomineral transformation took place during the leachingprocess.

The inability of XRD to detect NiS can be attributed tothree reasons: (1) the leaching reaction had not run tocompletion, (2) the NiS concentration was too low to bedetected, or (3) NiS was oxidized by ferric ions according toEquation [2] thus liberating elemental sulphur.

NiS + 2Fe3+ → Ni2+ + 2Fe2+ + S0 [2]

Table II also shows that other base metals, such as Coand Fe, are oxidized by Fe3+, depleting the concentration ofFe3+. These elements show a decrease in concentration afterleaching (Table II). Little attention was paid to these metalsin this study because they are present in small amountscompared to Ni and Cu, and their mineralogical phases werenot detectable by either SEM or XRD. Based on the initialfindings, it was proposed that to obtain maximum Ni and Cuextraction, it is advantageous to leach in two stages at 90°C.The first stage results in a higher extraction of Ni. Second-stage leaching was carried out using fresh leaching solutionon the partially leached solid residue to extract the remainingNi and Cu. It was also concluded that the ferric ion solutionshould be prepared at higher concentrations, so that the Niand Cu dissolution rates and extractions are not compromisedby other Fe3+-depleting agents.

Concentration

The concentrations of Fe3+ investigated were 5, 8, and 11 g/L, at a constant temperature of 90°C. It was thereforeadvantageous to leach in two continuous stages because thepercentage Ni extraction increased from 18% for single-stageleaching (Figure 5) to 26% (Figure 8) for two-stage leachingat 90°C using 5 g/L Fe3+.

The Ni percentage extractions for 5, 8, and 11 g/L of Fe3+

were approximately 26, 29, and 62% respectively. Thusfurther tests to establish the leaching mechanisms andparticulate processes were conducted for samples obtainedwhen leaching with 11 g/L Fe3+.

XRF analysis (Table III) showed a decrease in initial Nicontent of the matte from 50% to approximately 17% afterleaching.

Sulphur mass percentage also decreased, implying thatsulphur was liberated into the solution. XRD identified the

mineralogical phases containing Ni after leaching with Fe3+

11 g/L as NiS, Ni3S2, and Ni3S4 (Figure 9). Ni7S6 was, however, detected only after leaching with

5 g/L Fe3+. This can be explained by the fact that this Fe3+

concentration resulted in the slowest leaching rate, allowingthe quasi-intermediate product Ni7S6 (a mineral species thatis stable for only a short period during the reaction) to bedetected. XRD also detected elemental sulphur; this is thesulphur produced from the possible oxidation of NiS(Equation [2]).

According to Rademan et al. (1999) base metals aregradually leached out of the sulphide lattice to form specieswith lower metal-to-sulphur ratios. It is proposed that thetransformation proceeded as follows: Ni3S2 is rapidly alteredto Ni7S6, then forms NiS, and finally Ni3S4. Sulphide minerals

Atmospheric oxidative and non-oxidative leaching of Ni-Cu matte

The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 MAY 2014 405 ▲

Table III

Quantitative elemental analysis after two-stage non-oxidative leaching with 11 g/L [Fe3+] at 90ºC, mass%

Co Cu Fe Ni Pb Pd Pt S Other

Before 0.4 31 1.4 50 0.081 0.21 0.27 13 2.35After 0.074 46.63 7.54 17.38 0.104 0.184 0.231 5.84 22.01

Figure 6—SEM-EDS image of the matte after single-stage non-oxidativeleaching with 5 g/L [Fe3+] at 90°C

Figure 7—XRD pattern after single-stage non-oxidative leaching with 5g/L [Fe3+] at 90°C

Figure 8—Percentage extraction of Ni after two-stage non-oxidativeleaching with varying [Fe3+] at 90°C

PGMʼs

Hazelwoodite

Chalcocite

Micro cracks

a = Ni3S2

b = Cu2S

0 20 40 60 80

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Atmospheric oxidative and non-oxidative leaching of Ni-Cu matte

can be solubilized in acid or basic medium but form elementalsulphur only under oxidizing conditions. The leaching ofsulphides in acid medium in the presence of ferric chloridetakes place readily with the liberation of elemental sulphur(Havlík, and Kammel, 1995). The tests performed in thisstudy were non-oxidative; but elemental sulphur wasliberated, as shown in Figure 9. This demonstrates that ferricions performed two functions – one of oxidizing the sulphidelattice and the other of attacking (acid leaching) the ore. Thisfinding agrees with that of Park et al. (2006). The onlydisadvantage with non-oxidative ferric chloride leaching(Equation [3]) is that the ferrous ions formed cannot beregenerated back to ferric ions according to Equation [4]:

MS + 2Fe3+ → M2+ + 2Fe2+ + S M = Cu, Ni [3]

2Fe2+ + 2H+ + 0.5O2 → 2Fe3+ + H2O [4]

The leaching of copper sulphides is slightly more complexthan that of nickel sulphide, as seen in the single-stagetemperature leaching studies. A similar pattern is observed inthe two-stage leaching process, where higher extractions areobtained in the first few minutes of the experiment, followedby Cu cementation as early as 40 minutes into the reactiontime (Figure 10). At the end of the two-stage leaching

process using 5, 8, and 11 g/L of Fe3+, the Cu extractionswere 3, 11, and 11% respectively. However, the highestextractions, of 35, 67, and 96% for Fe3+ concentrations of 5,8, and 11 g/L respectively, were obtained in the first 15minutes, before the cementation process began.

Table III shows an increase in the Cu content; this isbecause a Cu–rich cake was created as a result ofcementation, while more Ni was dissolved into the solution.Rademan et al. (1999) stated that Cu2S is leached to formdigenite (Cu1.8S), with (Cu31S16) forming as a quasi-intermediate product. The digenite leaches further to formcovellite (CuS). Figure 9 shows that Cu2S was transformed toCu1.8S, with Cu31S16 detected only with leaching at 5 g/LFe3+. The leaching of copper proceeded as shown inEquations [5–7]:

5Cu2S + 2Fe3+ → Cu1.8S + Cu2+ + 2Fe2+ [5]

Cu1.8S + 8Fe3+ → 5CuS + 4Cu2+ + 8Fe2+ [6]

CuS + 2Fe3+ → Cu2+ + 2Fe2+ + S [7]

However, CuS formed by Equation [6] was not detectedby XRD (Figure 9), possibly as a result of copper rejectionand the slow leaching rate. An oxidative two-stage leach wastherefore conducted at the selected temperature of 90°C and11 g/L Fe3+ concentration to increase the overall leaching rateand base metal extraction. During first-stage leaching i.e. theleaching of nickel alloy, micro-cracking of the particlesoccurred, exposing the Ni3S2 and Cu2S on the edges of thecracks (barely visible) and on the outside of the particles(Figure 6), thus increasing the surface area available forleaching. In the second stage the size of the micro-cracksincreased as the leaching process continued until the particlesbecomes lightly porous, depositing PGMs on their surface,and finally particle breakage occurred as shown in Figure 11.

Effect of oxygen

According to Qui et al. (2007) oxygen diffuses initially intothe solution from the gas/liquid interface and then diffusesfurther into the solid/liquid interface. Oxygen participates inthe reaction after it contacts the ore surface. The solubility ofoxygen in water is affected by the temperature and partialpressure. Increasing the partial pressure of oxygen increasesits solubility. However; Deng et al. (2001) stated that

406 MAY 2014 VOLUME 114 The Journal of The Southern African Institute of Mining and Metallurgy

Figure 9—XRD pattern after two-stage non-oxidative leaching with 11 g/L [Fe3+] at 90°C

Figure 10—Percentage extraction of Cu after two-stage non-oxidativeleaching with varying [Fe3+] at 90°C

Figure 11—Electron micrograph of the matte after two-stage non-oxidative leaching with 11 g/L [Fe3+] at 90°C

a = Ni3S4

b = NiSc = Cu1.8Sd = Se = Ni3S2

f = Cug = Cu2S

0 20 40 60 80

PGMʼs

Hazelwoodite

CopperSulphide

Porous particle

Micro cracks

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leaching at a temperature above 85°C is not desirable,because oxygen solubility in the slurry decreases at highertemperatures, especially near the boiling point of thesolution. Since the oxygen partial pressure for this study waslow and the leaching temperature was high, oxygen solubilitywas enhanced by the high impeller speed of 600 r/min,leading to good dispersion and dissolution of oxygen in theleach slurry. Figure 12 illustrates how effective and importantoxygen is to the leaching process.

Approximately 98% Ni was extracted under oxidativeleaching. Ni alloy was extracted simultaneously with almostall of the Ni3S2. Under non-oxidative conditions, only a totalof 62% Ni was extracted. Nickel was dissolved via metathesis(Equation [9]) and oxidation (Equations [10–12]).

2Fe3+ + 3H2O → Fe2O3 + 6H+ [8]

Ni3S2 + 2Cu2+ → Cu2S + NiS + 2Ni2+ [9]

Ni0 + 2H+ + 0.5O2 → Ni2+ + H2O [10]

3Ni3S2 + 4H+ + O2 → Ni7S6 + 2Ni2+ + H2O [11]

Ni7S6 + 2H+ + 0.5O2 → 6NiS + Ni2+ + H2O [12]

Equation [10] shows that oxygen is responsible for Nialloy dissolution, and copper rejection is a result of Nimetathesis (Equation [9]). The intermediate product Ni7S6formed (Equation [11]) quickly transformed into NiS(Equation [12]). This confirms the findings of Rademan et al.(1999). It can thus be concluded that oxygen increasesliberation from the intricate sulphide bonds, thus enhancingthe dissolution process. Cu extraction from a copper–nickelcomplex sulphide ore was more complex than Ni extraction.Figure 13 depicts the extraction of Cu.

The leaching chemistry changed and the solubilized Cuwas re-deposited on the residue. It was evident that undernon-oxidizing conditions Cu is extracted rapidly andsuddenly undergoes cementation. The same process occurredunder oxidizing conditions. The nickel remaining in thepartially leached residue of the first stage was furtherdissolved by metathesis and oxidation in the second-stageleaching. It was concluded that no copper was leached frommatte in the two stages. The majority of the copper originallypresent in the matte remains intact through the oxidativeleaching, because copper is not as easily oxidized as nickeland as a result of copper cementation by nickel. To enhancethe extraction of Cu, a third oxidative leaching stage isnecessary, as Figure 5 shows that in the absence ofcementation most of the Cu is extracted within the first 50minutes of leaching. Chalcocite will be leached and

transformed to covellite (CuS). It is predicted that coppersulphide would be attacked by ferric ions (Equations[13–16]), and oxygen in this instance would serve as anoxidant responsible for the regeneration of ferrous ions toferric ions, according to Equation [13].

2Fe2+ + 2H+ +0.5O2 → 2Fe3+ + H2O [13]

5Cu2S + 2Fe3+ → Cu1.8S + Cu2+ + 2Fe2+ [14]

Cu1.8S + 8Fe3+ → 5CuS + 4Cu2+ + 8Fe2+ [15]

CuS + 2Fe3+ → Cu2+ + 2Fe2+ + S [16]

After the first-stage non-oxidative leaching, micro-cracking of the matte was observed (Figure 6), and after the two-stage non-oxidative leaching, extensive micro-cracking occurred with minor pores formed on the particles(Figure 11). Figure 14 shows extensive pores formed on theparticle under two-stage oxidative leaching, exposing morePGMs. It is believed that in the third stage, the particleswould proceed to complete breakage as a result of increasedoxidation and dissolution.

The XRF results agreed with those from SEM, as theyshowed that after oxidative leaching the relative proportionsof Pd and Pt in the matte was increased enabling them to bedetected by XRF (Table IV).

Atmospheric oxidative and non-oxidative leaching of Ni-Cu matte

The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 MAY 2014 407 ▲

Figure 12—Ni extraction after two-stage oxidative and non-oxidativeleaching with 11 g/L [Fe3+] at 90°C

Figure 13—Cu extraction at 90°C with 11 g/L [Fe3+]

Figure 14—Electron micrograph of the matte after two-stage oxidativeleaching with 11 g/L [Fe3+] at 90°C

PGMʼs

Hazelwoodite

Porous particle

Chalcocite

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Atmospheric oxidative and non-oxidative leaching of Ni-Cu matte

Since XRF expresses the elemental compositions asfractions of the total mass of the sample, during leachingsome constituents of the matte are removed, changing therelative proportion of the elements in the matte Coppermass% increased as a result of cementation, while Ni contentdecreased because of dissolution. Sulphur was also liberatedinto the solution; however, no evidence of sulphurpassivation in the leaching process, as observed by Park etal. (2006), was found in this study. XRD did not detect anyhazelwoodite (Ni3S2) or chalcocite (Cu2S) in the leachedresidue (Figure 15). The Ni transformed as far as NiS and apart of the Ni was oxidized to NiO. XRD confirmed thatcopper underwent dissolution in the first 45 minutes, as thecopper phase detected was Cu1.8S.

Particulate processes

The PSD of the matte obtained by a laser diffractiontechnique is shown in Figure 16.

The peaks represent the modal size (the particle size withthe highest volume percentage). The modal size of Ni-Cumatte before leaching was 138 μm. The modal size afterleaching at 90°C using 5 g/L Fe3+ (single-stage LX) reducedto 107 μm. This suggests a decrease in particle size either asa result of dissolution or breakage of the particles in thelarger size fraction. There was no significant shift in thevolume distribution and modal size after two-stage non-oxidative leaching using 11 g/L Fe3+; however, there was adecrease in the volume percentage of the mode. This indicatesa decrease in the proportion of larger particles as the numberof leaching stages was increased. The modal size afteroxidative leaching was the same as that reported after non-oxidative leaching, but shifted slightly to the left anddecreased in height, implying a greater number of particlesdecreased in size.

Kinetics

Various researchers such as Fan et al. (2010) and Jin et al. (2009) have found that the dissolution kinetics of Ni-Cu matte during the leaching process follow the shrinkingcore model (SCM). Only nickel dissolution kinetics wereinvestigated in this study, because nickel was the only metaldissolved during the leaching process since copper wasprecipitated. The shrinking core model gave a satisfactory fitto the experimental data only at 90°C, and the most suitablemodel was the product layer/ash diffusion model (Equation[17]) which had a correlation coefficient of 0.8 (Figure 17a).The experimental data at 90°C was fitted to the three-dimensional diffusion model by Jander (Equation [18]) andyielded a correlation coefficient of 0.93 (Figure 17b):

[1–(1–X)1/3]2 = k1t [17]

[1–3(1–X)2/3 + (1–X) + α (1–(1–X)1/3)]2 = k2t [18]

where X is the fraction reacted, k1 and k2 are reaction rateconstants, and α is a constant.

These results show that the leaching mechanism isdiffusion controlled as result of the formation of a poroussulphur layer on the surface of the particles during leaching.

408 MAY 2014 VOLUME 114 The Journal of The Southern African Institute of Mining and Metallurgy

Table IV

Quantitative elemental analysis after two-stage,oxidative leaching with 11 g/L [Fe3+] at 90ºC, mass%

Co Cu Fe Ni Pb Pd Pt S Other

Before 0.4 31 1.4 50 0.081 0.21 0.27 13 2.35 After oxidation 0.068 48.0 1.90 26.0 0.090 0.350 0.560 6.90 16.13

Figure 17a—Plot of the mixed (surface and product layer/ash diffusion)controlled process at various temperatures

Figure 15—XRD pattern after two-stage oxidative leaching with 11 g/L[Fe3+] 90°C

Figure 16—PSD (by volume) curves for the matte before and afterleaching

a = NiSb = Cu1.8Sc = NiO

0 10 20 30 40 50 60 70 80

4000

3500

3000

2500

2000

1500

1000

500

0

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For a better fit it is recommended that the PSD of the mattebe included in the SCM. Due to non-fitting models at lowertemperatures, the activation energy could not be determined.

Conclusion

Ni was selective leached from a Ni-Cu matte using a two-stage oxidative leaching process at 90°C and 11 g/L Fe3+. TheNi extraction was 98%. Non-oxidative leaching was found tobe less effective than oxidative leaching, as only 62% of theNi was extracted under the same conditions. Cu was entirelyrejected from solution by cementation, which would enable itto be selectively extracted in a third leaching stage. Oxygenwas found to enhance the liberation of metals from theintricate sulphide bonds, thus enhancing the dissolutionprocess. The major disadvantage of non-oxidative leaching isthat ferrous ions can not be regenerated back to ferric ions,leading to lower metal extractions. The shrinking core modeldid not give an acceptable fit to the experimental data attemperatures below 90°C, with the three-dimensionaldiffusion model by Jander giving the best fit at 90°C.Leaching resulted in extensive micro-cracking and poreformation on the particle surfaces, resulting in particlebreakage. Although a porous sulphur layer was formed onthe surface of the particles during leaching, sulphurpassivation was not found to occur.

Acknowledgements

The authors would like to thank Impala Platinum forsupplying the Ni-Cu matte, and the University ofJohannesburg URC for financial support.

References

AGRAWAL, A., KUMAR, V., PANDEY, B.D., and SAHU, K.K. 2006. A comprehensivereview on the hydro metallurgical process for the production of nickel andcopper powders by hydrogen reduction. Materials Research Bulletin, vol. 41. pp. 879–892.

BRUGMAN, C.F. and KERFOOT D.G. Treatment of nickel-copper matte at WesternPlatinum by the Sheritt acid leach process. Proceedings of the 25thAnnual Conference of Metallurgists, Toronto, Canada, 17-20 August 1986.Canadian Institute of Mining, Metallurgy and Petroleum, Toronto. pp. 512–531.

DENG, T., LU, Y., WEN, Z., and LIU, D. 2001. Oxygenated chloride-assistedleaching of copper residue. Hydrometallurgy, vol. 62, no. 1. pp. 23–30.

DUTRIZAC, J.E. 1992. The leaching of sulphide minerals in chloride media.Hydrometallurgy, vol. 29, no. 1–3. pp.1–45.

FAN, C., LI, B., FU, Y., and ZHAI, X. 2010. Kinetics of acid-oxygen leaching oflow-sulfur Ni-Cu matte at atmospheric pressure. Transactions ofNonferrous Metal Society of China, vol. 20, no. 6. pp. 1166–1170.

FORWARD, F.A., and MACKIW, V.N. 1955. Chemistry of the ammonia pressureprocess for leaching Ni, Cu, and Co from Sheritt Gordon sulphide concen-trates. Journal of Metals, vol. 7. pp. 457–463.

HABASHI, F. 1993. A Textbook of Hydrometallurgy. Metallurgie ExtractiveQuebec. Quebec, Canada. p. 99.

HAVLÍK, T. and KAMMEL, R. 1995. Leaching of chalcopyrite with acidified ferricchloride and carbon tetrachloride addition. Minerals Engineering, vol. 8,no. 10, pp. 1125–1134.

HOFIREK, Z., and KERFOOT, D.G.E. 1992. The chemistry of the nickel-coppermatte leach and its application to process control and optimization.Hydrometallurgy, Theory and Practice, Proceedings of the Ernest PetersInternational Symposium. Cooper, W.C. and Dreisinger, D.B. (eds.).Hydrometallugy, vol. 29. pp. 357–381.

JIN, B., YANG, X., and SHEN, Q. 2009. Kinetics of copper dissolution duringpressure oxidative leaching of lead-containing copper matte.Hydrometallurgy, vol. 99, no. 1–2. pp.119–123.

LAMYA, R.M., and LORENZEN, L. 2009. A semi-empirical kinetic model for theatmospheric leaching of a Ni-Cu converter matte in copper sulphate-sulphuric acid solution. Journal of the Southern African Institute ofMining and Metallurgy, vol. 109. pp. 755–760.

PARK, K.H., MOHAPATRA, D., and REDDY B.R. 2006. A study on the acidifiedferric chloride leaching of a complex (Cu–Ni–Co–Fe) matte. Separation andPurification Technology, vol. 51, no. 3. pp. 332–337.

PEARCE, R.F., WARNER, J.P., and MACKIW, V.N. 1960. A new method of matterefining by pressure leaching and hydrogen reduction. Journal of Metals,vol. 12. pp. 28–31.

PLASKET, R.P. and ROMANCHUK, S. 1978. Recovery of nickel and copper fromhigh grade matte at Impala Platinum by the Sherritt Process.Hydrometallugy, vol. 3. pp. 135–151.

QUI, T., NIE, G., WANG, J., and CUI, L. 2007. Kinetic process of oxidative leachingof chalcopyrite under low oxygen pressure and low temperature.Transactions of Nonferrous Metal Society of China, vol. 17, no. 2. pp. 418–422.

RADEMAN, J.A.M., LORENZEN, L., and VAN DEVENTER, J.S.J. 1999. The leachingcharacteristics of Ni–Cu matte in the acid–oxygen pressure leach process atImpala Platinum. Hydrometallurgy, vol. 52, no. 3. pp. 231–252. ◆

Atmospheric oxidative and non-oxidative leaching of Ni-Cu matte

The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 MAY 2014 409 ▲

Figure 17b—Plot of the data at 90°C on the three-dimensional diffusion model by Jander

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Introduction

Many university departments are rethinkingtheir programme and qualification mixes(PQMs) as a result of the implementation ofthe new Higher Education Qualifications Sub-Framework (HEQSF). Deciding on a viablePQM for a university department is a complexmatter that is influenced by many differentfactors, requirements and constraints. Inaddition to the requirements stated in theHEQSF, many other factors, such as marketrequirements, economic viability, pathdependency, articulation, mode of teaching(online learning vs classroom-basededucation), and university type (e.g. compre-

Universities and decision-making: programmeand qualification mix — four learningpathwaysby W.P. Nel*

Synopsis

The introduction of the Higher Education Qualifications Framework(HEQF) and the updated Higher Education Qualifications Sub-Framework(HEQSF) has caused many South African university departments torethink their programme qualification mixes (PQMs). In addition to therequirements stated in the HEQSF, a number of other factors have to betaken into consideration by a university department. These factorsinclude, for example, the standards generated by the EngineeringStandards Generating Body (ESGB) and subsequently approved by theEngineering Council of SA (ECSA) and the need to prepare students forvarious categories of professional registration with ECSA. This means thata university department has to choose the correct mix of LearningProgrammes (LPs) from the HEQSF menu (which consists of 13 types ofLPs). Preparing students for ECSA registration is aligned with the missionof universities, which is to teach and undertake research. However,research and the LPs associated with research go beyond the requirementsfor current ECSA registration. Assuming that universities offeringengineering LPs would elect to prepare students for both ECSA registrationand teach them to produce research outputs, which is mostly done atMaster and Doctorate levels (NQF Levels 9 and 10), then it follows thatacademics are more interested in NQF Level 5 to 10 pathways (abbreviatedas ‘L5-10’) rather than the shorter pathways required towards profes-sional registration. (For example, ECSA requires an NQF L5-L7 pathwayfor registration as a candidate professional technologist. This specificpathway may consist, for example, of two LPs, namely the 360-creditDiploma and the Advanced Diploma.) A L5-L10 pathway is a combinationof LPs that will prepare the learner with a NSC (or equivalent qualificationat level 4) to Doctoral level (level 10). Universities may choose at leastfour major pathways from the HEQSF menu in order to educate anddevelop students from NQF Level 5 to 10. However, various pathwaystowards registration in the category of candidate with ECSA are alsoembedded into these four NQF L5-L10 pathways, where each consist of aunique combination of LPs. Each of these pathways has an opportunitycost, and economic reality means that smaller departments may have tochoose between the four pathways. Of all the many factors involved inPQM decision-making, the focus of this paper is on the HEQSFrequirements, ECSA standards, and ECSA registration and how these,together with the various qualifications and educational LPs provided forby the HEQSF may impact on the PQM decision taken by engineeringdepartments and schools at South African universities. The proposed fourNQF L5-L10 ‘pathway tool’ for PQM decision-making may be useful forpointing out the advantages, disadvantages, and applications of thevarious pathways and combinations of pathways. Rather than decidingfrom a menu of thirteen qualifications and associated LPs, this articleproposes that decision-making be undertaken on the basis of a menu offour main articulated ‘NQF L5-L10’ pathways (which also include one ormore of the ECSA’s pathways for professional registration). The proposed‘NQF L5-L10 pathway’ tool is an attempt to move one step closer to theaim of achieving a structured decision-making approach for designing aPQM at departmental level.

* University of South Africa (UNISA).© The Southern African Institute of Mining and

Metallurgy, 2014. ISSN 2225-6253. Paper receivedFeb. 2014; revised paper received Apr. 2014.

411The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 MAY 2014 ▲

The ECSA designed standards for someof the 13 LPs that form part of the HEQSF -ECSA have to date not developedcompetency standards for Levels 9(Masters) and L10 (Doctorate). If univer-sities design LPs according to thesestandards, then learners would be eligible tocomply with ECSA’s educationalrequirements to register in the categories ofcandidate Pr Techn., Pr Tech., Pr Cert. Eng.,and Pr Eng.

KeywordsHigher Education Qualifications Sub-Framework (HEQSF), educational learningprogrammes (LPs), educational pathways,programme and qualification mix (PQM),registration in the appropriate candidatecategory with the ECSA approvedstandards, PQM decision-making,articulation.

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Universities and decision-making: programme and qualification

hensive, research-intensive, or technology-focused) also haveto be considered. One of the implications of the mergers ofRand Afrikaans University with Technikon Witwatersrand toform the University of Johannesburg and the old Universityof South Africa (Unisa) with Technikon Southern Africa toform the new Unisa is that these comprehensive universitiesmay now offer the B Eng degree. Should they do that? Thisdecision forms part of the decision-making that willdetermine their future PQMs, the combination of HEQSF-aligned programmes to offer in the future.

Competency standards approved by ECSA are the primerequirement to be considered by departments offeringengineering-related programmes and an important factorwhen deciding on a PQM. The inclusion or exclusion ofWork-Integrated Learning (WIL) is also an important HEQSF-related provision that has to be taken into consideration byuniversity departments.

The word ‘pathway’ is to be found in a number of ECSAdocuments. ‘Pathway’ means the combination of educationalLPs that must be followed to attain a specific goal (e.g.compliance with the educational requirements for profes-sional registration as engineer, certificated engineer,technologist, and technician). Academic staff at universitieshave two major responsibilities: tuition and research.Research and the LPs associated with research go beyond therequirements for current ECSA registration. Assuming thatuniversities offering engineering LPs would elect to bothprepare students for ECSA registration and teach them toproduce research outputs, which is mostly done at Mastersand Doctorate Levels (NQF Levels 9 and 10), then it followsthat academics are more interested in NQF Level 5 to 10pathways (abbreviated as ‘L5-10’) rather than the shorterpathways required towards professional registration. (Forexample, ECSA requires an NQF L5-L7 pathway forregistration as a candidate professional technologist. Thisspecific pathway may consist, for example, of two LPs,namely the 360-credit Diploma and the Advanced Diploma.)A L5-L10 pathway is a combination of LPs that will preparethe learner with a National Senior Certificate (NSC) (orequivalent qualification at exit level 4) to Doctoral level (exitlevel 10).

This paper does not focus on all the factors, requirements,and constraints that may impact, in various ways, on thePQM decision, but attempts to investigate the proposed ‘NQFL5-L10 pathways method’ or tool to simplify PQM decision-making while considering only some of the factors,requirements, and constraints such as the requirementsspecified by the ECSA and the HESQF. For this purpose, Ihave identified four main pathways (for NQF L5-L10) fromthe HEQSF document for developing and educating someonefrom National Senior Certificate (NSC) to doctoral level. Inthis article, I shall show that these NQF Level 5 to 10pathways, combined with the various pathways leading toregistration with the appropriate candidate category with theECSA, are important tools that can be used in the complexPQM decision-making process. These pathways are proposedas a method for reducing the complexity of decision-making(from the broad menu of thirteen HEQSF qualifications) to

one of deciding between articulated pathways to achievevarious goals while considering various factors,requirements, and constraints that may impact on the PQMdecision of a department. In essence, I have generalized andexpanded ECSA’s pathways for professional registration to aproposed ‘pathway tool’ in order to reduce the complexity ofPQM decision-making by one level. Instead of deciding froma menu of qualifications and associated programmes allowedby the HEQSF, I shall propose, in this article, that decision-making be done from a menu of articulated NQF L5-L10pathways that incorporate HEQSF-compliant ECSA pathwaysfor engineering practitioners.

Background information

A PQM of a university is a list, menu, or mix of approved LPsand qualifications that will be subsidized by the Departmentof Higher Education and Training (DHET). A universitydepartment or school has to follow a certain procedure to getits programmes on to such a list. In the future, the PQM ofSouth African universities may consist only of programmesprovided for by the HEQSF.

The HEQF was gazetted by the South African Minister ofEducation in 2007 and is an integral part of the NationalQualifications Framework (NQF). It was updated in January2013 and is now called the HEQSF. Universities are allowedto offer qualifications at NQF Levels 5 through 10. Thethirteen types of qualifications that form part of the HEQSFmenu are listed in Table I. It is important to note thatUniversities of Technology (UoTs) – the former technikons –and comprehensive universities are particularly affected bythe HEQSF because national diplomas have been replaced bydiplomas and the B Tech-degree has been excluded from theHEQSF. See Tables I and II.

ECSA designed standards for some of the 13 LPs thatform part of the HEQSF. ECSA has, however, not developedcompetency standards for the Level 8 Postgraduate diploma,Level 9 Professional Masters, and L10 Professional Doctoraldegrees. The HEQSF has significant implications foreducational providers. Changes to Work-integrated Learning(WIL) are also described in the HEQSF (2013, p. 11).

A number of South African universities have designedtheir new HEQSF-aligned PQMs and are in the process ofobtaining approval. It may still take a while until the HEQSF-aligned PQMs will be implemented. Currently, therequirements for registration with the ECSA for various typesof candidacy programmes are as described in Table II. In thecase of the technician, certificated engineer, and engineer asingle LP is required to meet the academic requirements forprofessional registration. In the case of the technologist thispathway is longer in terms of the number of LPs that must becompleted - it consist of at least two LPs, namely a NationalDiploma and BTech. See Table II.

In the past the UoTs offered cooperative education,meaning that the educational institution and industrycooperated to provide a joint educational programme, whichmight have included work-integrated learning (WIL). Thispractice has been largely continued by UoTs and compre-hensive universities (CUs) that offer vocational programmes.

412 MAY 2014 VOLUME 114 The Journal of The Southern African Institute of Mining and Metallurgy

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CUs like UJ and Unisa may in future for example, decide toinclude both traditional academic and vocation-basedprogrammes in their PQMs.

The HEQSF is one of three sub-frameworks. The othersare the framework for General and Further Education Sub-framework (GFETSF) and the framework for OccupationalQualifications (QCTO [sa], p. 3). The HEQSF provides forthree progression routes, namely vocational, professional,and general academic. These are listed in Table III and arerelated to some extent to the ‘types’ of universities found inSouth Africa. According to the HEQSF (2013, p. 9),undergraduate certificates and diplomas are usually foundwithin the vocational route, while the professional Bachelor,Master’s, and Doctorate degrees are characteristic of theprofessional route. The general academic learning routefocuses primarily on theoretical knowledge and research athigher levels.

It is important to note that the HEQSF specifies certainminimum requirements for qualifications. The minimumnumber of credits may, however, be exceeded by universities.For example, the HEQSF specifies that the Higher Certificateshould consist of at least 120 credits. ECSA’s minimumrequirement is, however, 140 and multiples of 140 credits.See Table IV (HEQSF, 2013, pp. 21, 22, 25, 28, 30; VanNiekerk, 2013: slide 11). It is not currently clear whether(public) universities will receive subsidy for those credits thatexceed the minimum number as specified by the HEQSF.

Four articulated (NQF L5-10) pathways under theHEQSF

Figure 1 illustrates four main L5-L10 higher educationalpathways, with the National Senior Certificate (NSC) or itsequivalent as the admission requirement and the Doctorate asthe exit level (Actually, there are more than four suchpathways if the differences at Master’s and Doctoral level arealso considered.) Please note that Figure 1 relates toengineering, since it uses some of ECSA’s names for thevarious HEQSF qualifications.

If the sole purpose is to develop a student from NSC toDoctoral level (and not for a specific ECSA registrationcatergory), then pathway 1 is a substitute for pathways 2, 3,and 4. Similarly, pathway 2 is a substitute for pathways 1, 3,and 4, and so on (for developing a student with a NSC rightup to Doctoral level). Note that each qualification in a specificpathway complements the others in the same pathway. Thismeans that all the qualifications, except for the Doctorate, areprerequisites for (subsequent) others in the same pathway.The removal of any one of the qualifications in a specificpathway will result in articulation deficiency. The view thatsome programmes or pathways may be substitutes for otherspoints to the fact that they may compete for the samestudents, assuming that such students meet the admissionrequirements of the different programmes.

Universities and decision-making: programme and qualification

413The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 MAY 2014 ▲

Table I

The HEQSF qualifications menu

Undergraduate Postgraduate

Qualification Level Qualification Level

Higher certificate 5 Postgraduate diploma 8

Bachelor honours degree 8

Advanced certificate 6

Diploma (240 credits) 6

Diploma (360 credits) 6

Advanced diploma 7 Master’s degree 9

Bachelor’s degree 7 (Professional) Master’s degree 9

(Professional) Bachelor’s degree 8 Doctoral degree 10

(Professional) Doctoral degree 10

Table II

Pre-HEQSF pathways towards professional registration

Category of professional registration Pre-HEQSF academic requirements for professional registration in the various categories

Technician National Diploma

Certificated Engineer Recognised certificate of competency (COC) for example the COC Mine Manager

(Mine Health and Safety) (Metalliferous)

Technologist National Diploma + B Tech (Eng), or M Tech (Eng) and pre-requisite LPs

Engineer B Eng/BSc (Eng)

Source: ECSA (http://www.ecsa.ac.za)

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Universities and decision-making: programme and qualification

The concept of articulation is an important factor thatshould guide PQM decision-making. The White Paper forPost-school Education and Training stresses that articulationmust be provided between various qualifications, thatstudents should not experience any ‘dead ends’, and thatpeople should be able to improve their qualifications withoutunnecessary repetition/duplication (DHET, 2013: viii).

LPs approved by the ECSA

The introduction of the HEQF set into motion a chain ofevents. The ECSA approved a number of genericHEQF/HEQSF-aligned competency standards for LPs thatform part of the pathways for professional registration as anengineer, technologist, technician, certificated engineer, andthe category of candidate with the Engineering Council ofSouth Africa (ECSA). The Engineering Standards GeneratingBody (ESGB) is a committee of ECSA that develops andrecommends relevant competency standards (qualification)for engineering practitioners for approval by to ECSA (VanNiekerk, 2013: slide 8).

ECSA pathways obviously have an influence on a facultyor school of engineering’s PQM and the individualdepartments and/or sections that make up this faculty orschool. In a number of cases, the ECSA introduced additionalrequirements (in addition to those required by the HEQSF)for those qualifications that form pathways to the variouscategories of professional registration. ECSA’s professionaldevelopment model towards registration is a two-stageprocess. Obtaining a relevant, engineering-accredited qualifi-cation is the first stage. Stage 2, (professional development ofengineering practitioners) consists of a candidacyprogramme. The various forms of professional registration(and their pathways) with the ECSA are shown in Figure 2prescribing learning outcomes.

414 MAY 2014 VOLUME 114 The Journal of The Southern African Institute of Mining and Metallurgy

Figure 1—From National Senior Certificate (NSC) to PhD - four articulated pathways under the HEQSF (ECSA qualification standards exist for those initalics and underlined)

Table III

Vocational (V), professional (P), and general (G)academic routes provided for in the HEQSF (ECSAqualification standards exist for those in italics)

NQF Type of qualification and route V P Glevel according to the HEQSF

5 Higher Certificate – ‘primarily vocational’ X(HEQSF, 2013: 21)

6 Advanced Certificate – ‘primarily Xvocational’ (HEQSF, 2013: 21); ‘particular career or professional context’ (HEQSF, 2013: 23)

6 Diploma (minimum, 240 credits) and XDiploma (minimum, 360 credits) – ‘primarily has a vocational orientation, which includes professional, vocational, or industry specific knowledge’ (HEQSF, 2013: 24)

7 Advanced Diploma – ‘vocational or X Xprofessional preparation or specialisation’ (HEQSF, 2013: 26)

7 (General) Bachelor’s degree – minimum X360 credits (HEQSF, 2013: 26)

8 Professional Bachelor’s degree – minimum X480 credits (HEQSF, 2013: 26)

8 Bachelor Honours degree – ‘broad and X Xgeneric areas of study, disciplines or professions’ (HEQSF, 2013: 30)

8 Postgraduate Diploma (HEQSF, 2013: 31) X X

9 Master’s degree (HEQSF, 2013: 32) X(Professional) Master’s degree X(HEQSF, 2013: 33)

10 Doctoral degree (HEQSF, 2013: 36) X(Professional) Doctoral degree (HEQSF, 2013: 38) X

Note that the HEQSF provides for two variants of the diploma, oneconsisting of 240 credits and another consisting of 360 credits (HEQSF,2013, p. 24).

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It is important to note that ECSA’s pathways are notcomprehensive NQF L5-10 pathways. (Currently, the B Engdegree at NQF Level 8 is the highest qualification that ECSAacknowledge for professional registration. This does not,however, prevent ECSA from adding further pathways forprofessional registration in the future.) It may be useful forengineering departments to incorporate one or more ofECSA’s pathways into one or more NQF L5-10 pathways inorder to prepare students not only for professionalregistration, but also to provide them with a route via whichthey can become involved in research (if research outputs areimportant for a specific department).

‘NQF L5-10’ pathways as a decision-making tool

Departments have one of the following three options as far asthe selection of NQF L5-10 pathways are concerned:

➤ To offer part of a NQF L5-10 pathway only. In thiscase, the four main ‘NQF L5-10’ LP are not useful as atool. In a specific area of study a department maydecide to offer one or two programmes on the ECSA

framework (Figure 2) only. This may be done for anumber of reasons (e.g. addressing a specific industryneed). Another reason may be that such an area ofstudy may not be earmarked for research activities. Adecision has been made, for example, at Unisa’sDepartment of Electrical and Mining Engineering, tooffer a Higher Certificate and Advanced Certificate inMine Surveying only. (Note that PLATO is the profes-sional body for mine surveyors.) ECSA’s competencystandards do not cater for programmes in the area ofmine surveying, since surveying is not primarilyconsidered as an engineering field of study

➤ To offer one ‘comprehensive’ (NQF L5-10) pathwayonly. The depth of the field of study may, for example,justify offering one comprehensive pathway. Studentnumbers, economic viability, staff capacity, and otherfactors may create an environment in which it isadvisable to offer one comprehensive pathway only.The White Paper for Post-school Education andTraining expresses the need for a greater focus onresearch and innovation, the building of research

Universities and decision-making: programme and qualification

The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 MAY 2014 415 ▲

Table IV

Differences between the HEQSF’s and ECSA’s requirements for selected qualifications

Qualification HEQSF’s minimum specified NQF credits ECSA’s minimum required NQF credits

Higher Certificate 120 140 – HCert(___Eng)

Advanced Certificate 120 140 – AdvCert(___Eng)

Diploma 240 280 - Dip (Eng Tech)360 360 – Dip (Eng)

Advanced Diploma 120 140 – AdvDip(Eng)

B degree 360 420 – BEng TechB degree (Professional) 480 560 – BEng

Bachelor Honours 120 140 – BEng Tech(Hons)

Figure 2—Pathways towards professional registration with the ECSA (Van Niekerk, 2013: slide 12)

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capacity, and the creation of more Master’s and PhDlearners (DHET, 2013: xiv, 33, 35). It is for thesereasons that a full NQF level 5 to 10 pathway isproposed as the PQM for mining engineering at Unisa

➤ To offer multiple ‘comprehensive’ (NQF L5-10)pathways. This may be done if the appropriateconditions exist. This option is discussed in more detailbelow.

The following provides some information and guidelineson the usefulness, advantages, disadvantages, andapplications of various NQF L5-10 pathways andcombinations of pathways.

Offering pathway 1 (starting with the B Eng)

Pathway 1 consists of three LPs, namely the B Eng, Master’s,and Doctoral degrees (see Figures 1 and 2). It thereforeincludes the (ECSA) pathway for developing a person with aNSC to the level where the academic requirements forregistration as a candidate professional engineer are met.

Pathway 1 is offered by many engineering faculties attraditional academic universities. This pathway has basicallynot been affected by the HEQSF. Some of these universitiesmay be research-intensive institutions that want to focus onthe postgraduate level. One of the advantages of this pathwayis that the B Eng degree is a well-established qualificationthat has been offered by a number of universities for aconsiderable period of time. Two (potential) disadvantagesof this pathway are as follows:

➤ The admission requirements for the B Eng degree atmost universities preclude many students fromregistering for this LP, e.g. symbols for Mathematicsand Physical Science.

➤ Offering a B Eng degree by means of distance learningis problematic due to the fact that it is a 560-creditqualification and a student in full-time employmentmay take very long to complete this degree. Employedstudents will also have to spent time away from theworkplace to attend laboratory sessions and non-work-based WIL if incorporated into such a programme.

Offering pathway 2 (starting with the Dip [Eng Tech])

Pathway 2 consists of five LPs, namely the 240-credit Dip(Eng Tech), Adv Dip (Eng), B Eng Tech Hons, M Eng, PhD,as well as a WIL component (HEQSF, 2013: 24). See Figures1 and 2. This pathway includes a number of the vocationalprogrammes listed in Table III. It is therefore an appropriateoption for a department at a comprehensive university thatwants to offer both vocational and academic programmes andthat has a history of offering vocational programmes.Pathway 2 will allow such a department to keep its currentstudent body (since a switch to pathway 1, for example, willresult in higher admission requirements (for the B Engdegree) which only a much smaller percentage of its currentstudent body is likely to meet). (This example points to the‘path-dependency’ factor, which is not discussed in thispaper.) One would expect more diversity in the staff profile ofa department offering this pathway: some members will needto have gained industrial experience and done vocational

programmes themselves, while some staff will, in addition,have to be recognized researchers with the ability tosupervise research at Master’s and Doctorate levels. One of itsmain strengths is that this pathway could prepare studentsfor three of ECSA’s four categories of professionalregistration, namely Certificated Engineer, Technologist, andTechnician (Figure 2). It is therefore the single most flexiblepathway in terms of ECSA’s requirements for professionalregistration, and should be of great use to the general studentpopulation and, indeed, most industries.

One of the advantages of pathway 2, which is particularlyrelevant to distance learning, is its multiple exit levelscombined with the fact that it includes some of the smallestprogrammes (in terms of credit value). This is veryimportant, since a learner who is in full-time employmentmay take twice as long to complete a programme comparedwith a student at a class-based institution. More exit levelsand shorter programmes in a pathway therefore reduce therisk of (1) non-completion on the part of the student, and (2)reduced government subsidy on the part of the university.

An important advantage (from a university’s perspective)of pathway 2 is that WIL does not form a compulsoryrequirement for the 240-credit diploma, but follows after thediploma has been completed – this means that the 240 creditdiploma is a prerequisite for the WIL component for learnersaspiring to achieve access to the Advanced Diploma. In termsof the HEQSF, universities will be held responsible for placingstudents in industry for work-based, WIL (HEQSF, 2013: 11,24). In 2011, 860 students were registered for the currentNational Diploma in Mining Engineering at Unisa. Thenumber of students that complete the National Diplomasuccessfully are fewer and more manageable as far as WILplacement is concerned. There are a number of reasons forthis relatively low throughput rate, but two are worthmentioning:

1) Distance learning requires the student to be bothhighly disciplined and independent

2) The admission requirements for mining at Unisa aremuch lower compared with those of the University ofPretoria and the University of the Witwatersrand.

In pathway 2, an additional 120 credits to the 240-creditDiploma must be completed in order to obtain admission tothe level 7 Advanced Diploma in Engineering. This isdescribed in the HEQSF (2013: 25) as follows: ’Candidateswho complete the 240-credit Diploma may enter an AdvancedDiploma upon successful completion of a work-integratedlearning component or a combination of work-integratedlearning and coursework equivalent to 120 credits that isapproved and accredited by an education provider and/or aprofessional body and a QC.’ (See Figure 1.)

Offering pathway 3 (starting with the Dip [Eng])

Pathway 3 consists of five LPs, namely the 360-creditDiploma, Adv Dip (Eng), B Eng Tech Hons, M Eng, and PhD.See Figures 1 and 2. Like pathway 2, it includes a number ofthe vocational programmes, listed in Table III.

Pathway 3 may achieve the same basic objective aspathway 2, but includes WIL in the (360-credit) diploma. The

416 MAY 2014 VOLUME 114 The Journal of The Southern African Institute of Mining and Metallurgy

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disadvantage of pathway 3 is that the lack of sufficientplacement opportunities for the 360-credit diploma mayprevent some students from graduating in the shortest timepossible and thus influence a university’s throughput rateand government subsidy. Pathway 3 – where the diplomaincludes a WIL component (as indicated in Figure 2) – isrecommended only in cases where the university canguarantee WIL placement for students. (See ‘Brief commentregarding WIL placement’.)

Offering pathway 4 (starting with the B Eng Tech)

Pathway 4 consists of four LPs, namely the B Eng Tech, BEng Tech Hons, M Eng, and PhD. See Figures 1 and 2.Pathway 4 is an option for a UoT or CU that elects toimplement the new B Eng Tech as an ‘alternative’ to thecurrent B Tech degree/national diploma combination.(Traditional academic universities are not excluded from this pathway.) One advantage is that this could preparestudents for two categories of ECSA professional registration(Table V). Another is that students obtain a ‘degree’ and nota ‘diploma’ at undergraduate level – this is important forsome students. (In pathways 2 and 3 students will obtain a‘degree’ for the first time only at postgraduate level – thehonours degree.)

Offering pathways 1 and 2

The combination of these two pathways is an appropriateoption for a fairly well-resourced department that has enoughstudents to ensure the viability of both pathways. It combinesthe strengths of pathway 2 with the more ambitious andelitist pathway 1. This may be a good option for a largedepartment at a comprehensive university. All four levels ofprofessional registration with ECSA can be covered by meansof this combination (Table V).

Offering other combinations of pathways

Other combinations of pathways are also possible (e.g.offering pathways 1, 2, 3, and 4). Offering all four pathways

is recommended only for very large, extremely well-resourceddepartments with very high numbers of students.

General remark regarding the need for a diversity ofeducational pathways

Engineering departments in the country will have to collec-tively offer a diversity of LPs to ensure that suitable numbersof technicians, technologists, certificated engineers, andprofessional engineers are educated and trained to cater forthe country’s need for scarce skills i.e. range of engineeringpractitioners.

Brief comment regarding WIL placement

The HEQSF (2013: 11) provides for at least five types of WIL:simulated learning, work-directed theoretical learning,problem-based learning, project-based learning, andworkplace-based learning. One type of WIL which is currentlybeing used is workplace-based learning. As far as this type oflearning is concerned, the HEQSF (2013: 11) states that:‘Where the entire WIL component or any part of it takes theform of workplace-based learning, it is the responsibility ofinstitutions that offer programmes requiring credits for suchlearning to place students into appropriate workplaces.’ Theimplication of this requirement is different from currentpractice (at Unisa) where it is also the student’s responsi-bility to try and find WIL placement, an internship,employment, or a bursary with a mining company. The WILoffice at Unisa enables students to upload CVs to a databasethat is available to employers. Unisa currently providesstudents with an experiential learning guide and mentor’sguide for WIL and relies on mines, mentors at mines, andheads of department to ensure that students are exposed tothe various areas prescribed. The new HEQSF expect univer-sities to guarantee workplace-based learning. For everylearner enrolling on the respective programme, given the factthat universities do not own mines or other workplaces,industry’s participation and support for WIL is crucial. It is,however, also important to note that the HEQSF provides forgreater diversity of WIL activities. That said, it is extremely

Universities and decision-making: programme and qualification

The Journal of The Southern African Institute of Mining and Metallurgy VOLUME 114 MAY 2014 417 ▲

Table V

Individual NQF L5-10 pathways and their relationship with pathways that lead to professional registration with theECSA

NQF L5-10 pathway Inclusion of the ECSA pathways for registration as professional ...

Engineer Certificated Engineer Technologist Technician

1 Yes No No No

2 No Yes Yes Yes

3 No Yes Yes Yes

4 No Yes Yes No

1 and 2 Yes Yes Yes Yes

... - - - -

(Other combinations, e.g. 1, 2, and 3) Yes Yes Yes Yes

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Universities and decision-making: programme and qualification

expensive for universities to provide facilities such as mock-up tunnels and stopes and pilot plants, and to develop avirtual reality capability and various forms of simulatedlearning.

If one assumes that industry may still prefer that themajor part of WIL must be workplace-based learning, thenthis raises the question of how universities can guaranteeworkplace-based placement for students. This is becauseproviding such placements involves scarce industry resourcesand not all employers in need of qualified persons may havethe capacity to provide students with suitable workplace-based support. One option may be to limit student numbersto what industry is capable of committing to. But shoulduniversities follow the example of the various faculties ofmedicine and veterinary sciences, both of which havestringent pre-screening mechanisms? A university thatapplies such screening mechanisms needs to know how manystudents can be absorbed by industry and various otheremployer organizations. Following the example of thefaculties of medicine means that engineering departments inuniversities will need to have suitable screening and selectionmechanisms in place to determine which of the applicants willhave the best chances of success in their future careers.Perhaps another solution is for mining companies to re-introduce the well-known and respected Learner Officialprogrammes.

Pathway 2 has the potential to overcome the aboveproblem (at least, to some extent). The 240-credit Dip (EngTech), which will include theoretical and laboratory modulesbut no compulsory work-placed based modules, can act as a‘screening mechanism’ before successful (but unemployed)students are employed by industry. The Stage 2 structuraland mentored professional development of engineeringgraduates could be designed in such a way as to incorporate WIL.

Conclusion and the way forward

Decision-makers in university departments have to considermany factors, requirements, and constraints when decidingon a PQM. In this paper, I have generalized and expandedECSA’s pathways for professional registration to suggest a

proposed ’pathway tool’ that will reduce the complexity (byone level) of this decision-making process. Instead of havingto decide on the basis of a menu of thirteen qualifications andassociated programmes, I propose that decision-making bedone from a menu of four main articulated ‘NQF L5-10’pathways that also include one or more of the ECSA’spathways leading to professional registration. The proposed‘NQF L5-10 pathways’ tool is an attempt to move one stepcloser to the aim of achieving a structured decision-makingapproach for designing a PQM at departmental level. Aholistic evaluation of the strengths and weaknesses ofvarious ‘NQF L5-10’ pathways and combinations of suchpathways are required when deciding on a PQM. The miningsection at the Department of Electrical and MiningEngineering at Unisa proposes pathway 2 as its PQM for thediscipline of Mining Engineering.

References

Department: Higher Education and Training (DHET). 2013. White Paper for

Post-school Education and Training: Building an Expanded, Effective and

Integrated Post-School System. Pretoria.

ECSA. Candidate Engineer, Candidate Certificated Engineer, Candidate

Engineering Technologist, Candidate Engineering Technician.

http://www.ecsa.ac.za [Accessed 1 Apr. 2014].

QUALITY COUNCIL FOR TRADES AND OCCUPATIONS. [Sa]. Introduction to the Quality

Council for Trades and Occupations (QCTO).

HIGHER EDUCATION QUALIFICATIONS FRAMEWORK (HEQF). Higher Education Act, No.

101 of 1997. Government Gazette 303533, Notice 928. 6 October 2007.

HIGHER EDUCATION QUALIFICATIONS SUB FRAMEWORK (HEQSF). 2013 (as revised).

http://sun025.sun.ac.za/portal/page/portal/Administrative_Divisions/INB/

Home/New%20Modules/Revised%20HEQSF%20Jan2013%20FINAL.pdf

[Accessed 3 Oct. 2013].

VAN NIEKERK, D. 2013. Engineering qualifications and the Higher Education

Qualifications Sub-Framework (HEQSF). Presentation at the Science

Campus of Unisa. ◆

418 MAY 2014 VOLUME 114 The Journal of The Southern African Institute of Mining and Metallurgy

ErratumThe affiliation for the author A. Heidary Moghadam published in the SAIMM Journal vol. 113, no. 12, pp. 941–945entitled: ʻA study on the effect of coke particle size on the thermal profile of the sinters produced in EsfahanSteel Company (ESCO)ʼ, by A. Dabbagh*, A. Heidary Moghadam†, S. Naderi*, and M. Hamdi* was incorrectly listedby the author.The correct affiliation of the author should be Metallurgy Department, Engineering Faculty, Islamic Azad University,Dezful Branch, Dezful, Iran

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Technical Conference andIndustry day

20–21 August 2014—Conference22 August 2014—Industry day

OBJECTIVESThe conference will focus on improvinghealth, safety and the environmentalimpact in the mining and metallurgyindustry and highlight actions to betaken. It will act as a platform for learningand allow people to share ideas onsafety, health and the environment.

This conference aims to bringtogether management, DMR, Chamberof Mines, Unions and health and safetypractitioners at all levels from theindustry to share best practice andsuccessful strategies for zero harm and avalue-based approach to health andsafety. It will address the mainchallenges in the mining industry such aslogistics, energy and safety ofemployees, contractors and thecommunities.

For further information contact:Head of Conferencing

Raymond van der Berg, SAIMM,P O Box 61127, Marshalltown 2107

Tel: +27 11 834-1273/7Fax: +27 11 833-8156 or +27 11 838-5923

E-mail: [email protected]: http://www.saimm.co.za

WHO SHOULD ATTEND

The conference should be of value to:˙ Safety practitioners˙ Mine management˙ Mine health and safety officials˙ Engineering managers˙ Underground production supervisors˙ Surface production supervisors˙ Environmental scientists˙ Minimizing of waste˙ Operations manager˙ Processing manager˙ Contractors (mining)˙ Including mining consultants, suppliers

and manufacturers˙ Education and training˙ Energy solving projects˙ Water solving projects˙ Unions˙ Academics and students˙ DMR˙ Acid mine drainage.

Emperors Palace, Hotel Casino ConventionResort, Johannesburg

SPONSORSHIPSponsorship opportunities are available.Companies wishing to sponsor or exhibitshould contact the Conference Co-ordinator.

SUPPORTED BY:

Sustaining Zero Harm

‘It always seems impossible until it’s done’Nelson Rolihlahla Mandela

SPONSORS:

MMMA

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

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

19–23 May 2014 — Fundamentals of Process SafetyManagement (PSM)Johannesburg, South Africa Contact: RDC PriorTel: +27 (0) 825540010, E-mail: [email protected]

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/

27–29 May 2014 — Furnace Tapping Conference 2014Misty 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

11–12, June, 2014 — AIMS 2014: 6th InternationalSymposium ‘High Performance Mining’Aachen, GermanyContact: Sandra ZimmermannTel: +49-(0)241-80 95673, Fax: +49-(0)241-80 92272 E-Mail: [email protected]: http://www.aims.rwth-aachen.de

26–30 June 2014 — Society of Mining Professors A Southern African Silver AnniversaryThe Maslow Hotel, Sandton, Gauteng, South AfricaContact: 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

15–16 July 2014 — Mine Planning SchoolMine Design Lab, Chamber of Mines Building,The University of the WitwatersrandContact: Camielah JardineTel: +27 11 834-1273/7, Fax: +27 11 838-5923/833-8156 E-mail: [email protected], Website: http://www.saimm.co.za

4–5 August 2014 — Pyrometallurgical ModellingPrinciples and PracticesEmperors Palace Hotel Casino Convention Resort,JohannesburgContact: Camielah JardineTel: +27 11 834-1273/7, Fax: +27 11 838-5923/833-8156 E-mail: [email protected], Website: http://www.saimm.co.za

6–8 August 2014 — MinPROC 2014Lord Charles Hotel, Somerset West, Cape Town

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.za

1–2 September 2014 — Drilling and BlastingSwakopmund Hotel & Entertainment Centre, Swakopmund, Namibia 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

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.za

16–17 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

23–24 September 2014 — Grade control andreconciliationMoba Hotel, Kitwe, Zambia 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

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The Journal of The Southern African Institute of Mining and Metallurgy MAY 2014 ▲xi

2014 (contineud)

29–30 September 2014—SHAPE: 1st InternationalConference on Solids Handling and Process EngineeringUniversity of Pretoria, South AfricaContact: 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

20–24 October 2014 — 6th International PlatinumConferenceSun City, South AfricaContact: 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

12 November 2014 — 12th Annual Southern AfricanStudent ColloquiumMintek, RandburgContact: 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

18–19 November 2014 — Third InternationalEngineering Materials and Metallurgy Conference andExhibition (iMat 2014)Shahid Beheshti International Conference Center, Tehran, Iran Contact: Kourosh Hamidi E-mail: [email protected]

19–20 November 2014 — Accessing Africa’s MineralWealth: Mining Transport LogisticsEmperors 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.za

201514–17 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 forMetallurgical Technology 2015Dusseldorf, GermanyWebsite: http://www.metec-tradefair.com6–8 July 2015—Copper Cobalt Africa IncorporatingThe 8th Southern African Base Metals Conference

The Falls Resort, Victoria Falls, Livingstone, Zambia Contact: Raymond van der BergTel: +27 11 834-1273/7, Fax: +27 11 838-5923/833-8156E-mail: [email protected], Website: http://www.saimm.co.za13–17 July 2015—School on Production of Clean SteelMisty Hills Conference Centre, MuldersdriftContact: 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.za12–14 August 2015—The Seventh InternationalHeavy Minerals Conference ‘Expanding the horizon’12–13 August 2015—Conference14 August 2015—Technical VisitSun City, South AfricaContact: 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.za30 September-2 October 2015—WorldGold Conference201530 September–1 October–Conference2 October 2015–Technical VisitsJohannesburg, South AfricaContact: 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.za5–9 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.za12–14 October 2015—Slope Stability 2015:International Symposium on slope stability in open pitmining and civil engineeringCape Town Convention Centre, Cape TownContact: 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.za8–13 November 2015—MPES 2015: Twenty ThirdInternational Symposium on Mine Planning &Equipment Selection 8–11 November 2015—Conference12–13 November 2015—Tours and Technical VisitsSandton Convention Centre, Johannesburg, South AfricaContact: Raj Singhal, E-mail: [email protected] or E-mail: [email protected], http://www.saimm.co.za

INTERNATIONAL ACTIVITIES

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xii MAY 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

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2014◆ SYMPOSIUM

6th South African Rock Engineering Symposium SARES 2014 —Creating value through innovative rockengineering12–14 May 2014, Misty Hills Country Hotel and Conference Centre, Cradle of Humankind

◆ CONFERENCE

27–28 May 2014,29 May 2014,Misty Hills Country Hotel and Conference Centre, Cradle of Humankind

◆ SEMINAR

26–30 June 2014,

◆ SCHOOLMine Planning School15–16 July 2014, Mine Design Lab, Chamber of MinesBuilding, The University of the Witwatersrand

◆ CONFERENCE

4–5 August 2014, Emperors Palace, Hotel Casino ConventionResort, Johannesbur

◆ CONFERENCEMineSafe Conference 2014Technical Conference and Industry day20–21 August 2014, Conference22 August 2014, Industry dayEmperors Palace, Hotel Casino Convention Resort,Johannesburg

1–2 September 2014,

◆ SCHOOL3rd Mineral Project Valuation School9–11 September 2014, Mine Design Lab, Chamber of MinesBuilding, The University of the Witwatersrand

◆ CONFERENCE

16–17 September 2014,

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]

F

Website: http://www.saimm.co.za

EXHIBITS/SPONSORSHIP

Companies wishing to sponsor

and/or exhibit at any of these

events should contact the

conference co-ordinator

as soon as possible

Page 84: Saimm 201405 may

®

Crest Chemicals

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